Sanbrado Gold Project, Burkina Faso

NI 43-101 Technical Report: Sanbrado Gold Project, Burkina Faso

Effective Date: 13 December 2017

Principal Office: 14 Southbourne Street, Scarborough WA 6019, Western Australia T: + 61 8 9481 7344 F: + 61 8 9481 7355 E: [email protected] www.westafricanresources.com ACN: 121 539 375 West African Resources Limited

Document Information Page

Principal Mining Engineer Qualified Persons Stuart Cruickshanks FAusIMM SCME Executive Engineer Tom Kendall FAusIMM, MMICA, Mintrex Pty Ltd FIEAust, CPEng Manager: Projects and Studies Leon Lorenzen FAusIMM, FSAIMM, Mintrex Pty Ltd FIEAust, CPEng Managing Director David Morgan MAusIMM, MIEAust, Knight Piesold Consulting Pty Ltd CPEng Principal Consultant Brian Wolfe MAIG International Resource Solutions Pty Ltd

Effective Date 13 December 2017

Versions / Status FINAL REPORT

Print Date 13 December 2017

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Table of Contents

Executive Summary ...... 1

Introduction ...... 32

Reliance on Other Experts ...... 34 Property Description and Location ...... 35

Accessibility, Climate, Local Resources, Infrastructure and Physiography ...... 38

History ...... 41

Geological Setting and Mineralisation ...... 46

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Deposit Types ...... 52 Exploration ...... 54

Drilling ...... 56

10.2.1 Hole Planning and Set-up ...... 58 10.2.2 Surveying and Orientations ...... 59 10.2.3 Hole Logging Procedures ...... 59 10.2.4 AC and RC Logging ...... 59 10.2.5 DC Logging ...... 59

10.3.1 Historical Dry Bulk Density Testwork by Channel ...... 60 10.3.2 WAF Dry Bulk Density Testwork ...... 60

Sample Preparation, Analyses and Security ...... 65

11.1.1 Historical Work ...... 65 11.1.2 West African Resources Ltd ...... 65

11.3.1 Channel Resources Ltd ...... 66 11.3.2 West African Resources Ltd ...... 66

11.4.1 Channel Resources Ltd ...... 66 11.4.2 West African Resources Ltd ...... 68

Data Verification ...... 69

12.3.1 Thompson and Howarth Plot (T & H Plot) ...... 70 12.3.2 Rank % HARD Plot ...... 70 12.3.3 Mean vs. % HARD Plot ...... 70 12.3.4 Mean vs. %HRD Plot ...... 70 12.3.5 Correlation Plot ...... 70 12.3.6 Quantile-Quantile Plot ...... 70 12.3.7 Standard Control Plot ...... 71 12.3.8 Cusum Plots ...... 71

12.4.1 Standards ...... 71 12.4.2 Blanks ...... 87 12.4.3 Pulp (Laboratory) Repeats ...... 87 12.4.4 Field Duplicates ...... 87

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Metallurgy and Process Development ...... 94

13.5.1 Samples Testwork Program A15684 ...... 98 13.5.2 Results Testwork Program A15684 ...... 98 13.5.3 Abrasion Index and Ball Mill Work Index Program A15684 ...... 100 13.5.4 A15684 Summary ...... 100

13.6.1 Samples Testwork Program A16326 ...... 101

13.7.1 Summary of Testwork Program A16326 Part A ...... 104

13.8.1 Samples Utilised for Testwork Program 2016 - A16326 Part B ...... 104 13.8.2 Results Testwork Program A16326 Part B ...... 105 13.8.3 Bulk Tests for Testwork Program A16326 Part B ...... 106 13.8.4 Summary of Testwork Program A16326 Part B ...... 111

13.9.1 Samples Utilised for Testwork Program A16557 ...... 112 13.9.2 Results of Testwork Program A16557 ...... 112 13.9.3 Summary of Testwork Program A16557 ...... 125

13.10.1 Samples Utilised for Testwork Program A17283 ...... 126 13.10.2 Results Testwork Program A17283 ...... 126 13.10.3 Summary of Testwork Program A16283 ...... 127

13.11.1 Samples Utilised for Testwork Program A17341 ...... 128 13.11.2 Results Testwork Program A17341 (Comminution) ...... 128 13.11.3 Results Testwork Program A17431 (Gravity and Leaching) ...... 131 13.11.4 Summary of Testwork Program A17431 ...... 134

13.12.1 Samples Utilised for Testwork Program A17750 ...... 135 13.12.2 Results Testwork Program A17750 ...... 135 13.12.3 Summary of Testwork Program A17750 ...... 137

13.14.1 Channel Resources/SGS Work ...... 139 13.14.2 WAF Mineralogical Work ...... 140

13.15.1 Evaluation of Key Gold Leaching Processing Variables ...... 142 13.15.2 Gravity Gold Recovery ...... 150 13.15.3 Reagent Consumptions ...... 157 13.15.4 Silver Metallurgy ...... 159 13.15.5 Sulphide Influences and Leach Feed Head Grade ...... 160

13.16.1 Methodology ...... 164 13.16.2 Final Algorithms ...... 169

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Mineral Resource Estimates ...... 172

14.2.1 Database Validation ...... 172 14.2.2 Summary of Data Used in Estimate ...... 173 14.2.3 Interpretation and Modelling ...... 174 14.2.4 Data Flagging and Compositing ...... 176 14.2.5 Statistical Analysis ...... 179 14.2.6 Variography ...... 186 14.2.7 Block Modelling ...... 196 14.2.8 Dry Bulk Density Data ...... 196 14.2.9 Grade Estimation ...... 197 14.2.10 Resource Reporting ...... 205

14.3.1 Database Validation ...... 207 14.3.2 Summary of Data Used in Estimate ...... 208 14.3.3 Interpretation and Modelling ...... 208 14.3.4 Data Flagging and Compositing ...... 212 14.3.5 Statistical Analysis ...... 215 14.3.6 Variography ...... 222 14.3.7 Block Modelling ...... 235 14.3.8 Dry Bulk Density Data ...... 236 14.3.9 Grade Estimation ...... 236 14.3.10 Resource Reporting ...... 243 Mineral Reserve Estimates ...... 245

15.3.1 Pit Optimisation Assumptions ...... 250

Mining Methods ...... 257

16.1.1 Drilling & Blasting ...... 257 16.1.2 Loading & Hauling ...... 257 16.1.3 Grade Control ...... 258 16.1.4 Auxiliary Equipment ...... 258 16.1.5 Pit Dewatering ...... 258 16.1.6 Waste Rock Storage ...... 258

Recovery Methods ...... 260

17.2.1 Blending of Process Plant Feed ...... 262 17.2.2 Comminution Circuit Design Basis ...... 263 17.2.3 Comminution Circuit Selection ...... 264

17.3.1 Primary Crushing ...... 264 17.3.2 Crushed Ore Stockpile ...... 265 17.3.3 Grinding and Classification Circuit ...... 266 17.3.4 Gravity Recovery Circuit ...... 267 17.3.5 Carbon in Leach Circuit ...... 267

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

17.3.6 Elution, Electrowinning and Smelting ...... 268 17.3.7 Tailings Disposal ...... 269

17.4.1 Lime ...... 270 17.4.2 Cyanide ...... 270 17.4.3 Caustic Soda/Sodium Hydroxide ...... 271 17.4.4 Hydrochloric Acid ...... 271 17.4.5 Activated Carbon ...... 272 17.4.6 Oxygen ...... 272 17.4.7 Flocculant ...... 272

17.7.1 Compressed Air ...... 275 17.7.2 Process Water ...... 276 17.7.3 Raw Water ...... 276 17.7.4 Potable Water ...... 276 17.7.5 Sewage ...... 276 Project Infrastructure ...... 277

18.2.1 Upgrade of Existing Gravel Road from Zempasgo ...... 277 18.2.2 Site Access Road ...... 277 18.2.3 Mine Access and Haulage Roads ...... 278

18.3.1 Process Plant ...... 278 18.3.2 Mining Contractor’s Area ...... 279 18.3.3 Tailings Storage Facility ...... 279 18.3.4 Water Storage Facility ...... 281 18.3.5 Gibgo River Water Abstraction ...... 281

18.4.1 Main Entrance Gatehouse and First Aid ...... 282 18.4.2 Administration Building ...... 283 18.4.3 Laboratory and Core Shed ...... 283 18.4.4 Crib Rooms ...... 283 18.4.5 Security, Laundry and Change Room ...... 283 18.4.6 Plant Workshop ...... 284 18.4.7 Warehouse and Reagent Storage ...... 284 18.4.8 MCC Building (Switchroom Buildings) ...... 284 18.4.9 Process Control Rooms ...... 284 18.4.10 Plant Office ...... 284 18.4.11 Gold Room Building ...... 285

18.5.1 Raw Water ...... 285 18.5.2 Potable Water ...... 285 18.5.3 Water Management ...... 285 18.5.4 Sewage ...... 286

18.9.1 Process Plant ...... 287

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

18.9.2 Accommodation Camp ...... 288 18.9.3 Infrastructure and Mining Contractor’s Area ...... 288

18.11.1 Project Management Plan ...... 290 18.11.2 Work Breakdown Structure ...... 293 18.11.3 Project Schedule ...... 294 18.11.4 Operational Readiness ...... 294

18.12.1 Introduction ...... 295 18.12.2 Transition from Construction to Operational Start-Up ...... 295 18.12.3 Site Operations Roster ...... 295 18.12.4 Operations Organisation Structure and Manning ...... 296 18.12.5 Workforce Make-Up ...... 303 18.12.6 Contractors ...... 304 18.12.7 Occupational Health and Safety ...... 305 18.12.8 Pre-Employment Procedures and Recruitment ...... 306 18.12.9 Training and Development ...... 308 Market Studies and Contracts ...... 310 Environmental Studies, Permitting and Social or Community Impact ...... 311

20.4.1 Closure and Rehabilitation ...... 312 20.4.2 Site Monitoring ...... 314

20.5.1 Introduction ...... 315 20.5.2 Legislation ...... 316 20.5.3 Mining Code ...... 316 20.5.4 Environmental Code ...... 317 20.5.5 Regulations ...... 317 20.5.6 Environmental Protection Policies and Strategies ...... 318 20.5.7 Rural and Land Development Policies and Strategies ...... 319 20.5.8 International Agreements, Protocols and Conventions ...... 319

20.7.1 General Approach ...... 321 20.7.2 Closure Costs ...... 323 20.7.3 Cost Provisions for Mine Closure ...... 324 Capital and Operating Costs ...... 325

21.1.1 Qualifications and Exclusions ...... 325 21.1.2 Mining Costs ...... 326 21.1.3 Processing and Maintenance Costs ...... 327 21.1.4 Labour ...... 327 21.1.5 Mobile Equipment ...... 329 21.1.6 Power ...... 329 21.1.7 Consumables ...... 330 21.1.8 Maintenance Cost ...... 331 21.1.9 Laboratory ...... 331 21.1.10 Water Supply ...... 331

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

21.1.11 General and Administration Costs ...... 331 21.1.12 Personnel ...... 333 21.1.13 Sustaining Capital ...... 334 21.1.14 Owner’s Mining Costs ...... 336 21.1.15 Summary ...... 336

21.2.1 Estimate Basis ...... 337 21.2.2 Estimate Assumptions and Clarifications ...... 337 21.2.3 Exclusions ...... 338 21.2.4 Estimate Detail ...... 338 21.2.5 Contingency ...... 345 21.2.6 Exchange Rates ...... 345 21.2.7 Statistical Analysis ...... 346 Economic Analysis ...... 348

22.4.1 Capital and Operating Costs ...... 350 22.4.2 Closure and Salvage Value ...... 351 22.4.3 Working Capital ...... 351

22.5.1 Surface Taxes ...... 351 22.5.2 Government Royalty ...... 351 22.5.3 Local Development Fund ...... 352 22.5.4 Duties and Levies ...... 352 22.5.5 Value Added Tax ...... 352 22.5.6 Corporate Income Tax ...... 352 22.5.7 Withholding Tax ...... 352

Adjacent Properties ...... 358

Other Relevant Data and Information ...... 359 Interpretation and Conclusions ...... 360 Recommendations ...... 361 References ...... 362

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

List of Figures

Figure 1.1 – Sanbrado Gold Project Location ...... 2 Figure 1.2 – Gold Deposits of West Africa on Regional Geology ...... 4 Figure 1.3 – Sanbrado Gold Project – Drillhole Summary Plan ...... 7 Figure 1.4 – Plan View of Drilling ...... 11 Figure 1.5 – Isometric NE View of Mankarga 1 ...... 12 Figure 1.6 – Isometric SE View of Mineralisation Domains M5 Deposit ...... 14 Figure 1.7 – M5 Pit Design showing Initial Pit Stages and Waste Rock Dump ...... 19 Figure 1.8 – Sanbrado Gold Project Process Plant Flow Diagram ...... 22 Figure 1.9 – Site Development ...... 24 Figure 4.1 – Sanbrado Gold Project Location ...... 35 Figure 4.2 – Corporate Structure - West African Resources Ltd ...... 37 Figure 5.1 – Sanbrado Gold Project Location Map Showing Major Secondary Access Roads ...... 38 Figure 5.2 – Sanbrado Gold Project: Photograph showing Typical Physiography of the Sanbrado Permit ...... 40 Figure 6.1 – Tanlouka Permit : Total Gradient of Total Magnetic Intensity (TMI) ...... 43 Figure 6.2 – Tanlouka Permit : Historic Soil Geochemistry overlying TMI Magnetic image ...... 44 Figure 6.3 – Tanlouka Permit: Historic Channel RC and DC Drilling ...... 45 Figure 7.1 – Generalised Geological and Structural Map of the West African Craton ...... 46 Figure 7.2 – Gold Deposits of West Africa on Regional Geology ...... 48 Figure 7.3 – Regional Geological Map-Eastern Burkina Faso ...... 49 Figure 8.1 – Location of the Major Gold Mines in the West African Region ...... 52 Figure 9.1 – Sanbrado Gold Project – Completed Soil Auger Drilling Program ...... 55 Figure 10.1 – Location of the Major Gold Mines in the West African Region ...... 58 Figure 10.2 – Sanbrado Gold Project – Completed Soil Auger Drilling Program ...... 62 Figure 10.3 – Sanbrado Gold Project – Mankarga 5 Typical Cross-section (SW650) ...... 62 Figure 10.4 – Sanbrado Gold Project – Mankarga 1 North and South Drillhole Summary Plan ...... 63 Figure 10.5 – Sanbrado Gold Project – Mankarga 1 South Section SE0450 ...... 64 Figure 12.1 – Standard G310-1 ...... 72 Figure 12.2 – Standard G310-6 ...... 73 Figure 12.3 – Standard G311-1 ...... 74 Figure 12.4 – Standard G313-6 ...... 75 Figure 12.5 – Standard G901-1 ...... 76 Figure 12.6 – Standard G901-3 ...... 77 Figure 12.7 – Standard G901-7 ...... 78 Figure 12.8 – Standard G908-7 ...... 79 Figure 12.9 – Standard G909-5 ...... 80 Figure 12.10 – Standard G913-9 ...... 81 Figure 12.11 – Standard G913-9 ...... 82 Figure 12.12 – Standard G995-1 ...... 83 Figure 12.13 – Standard G996-4 ...... 84 Figure 12.14 – Standard GS-1P5F ...... 85 Figure 12.15 – Standard GS-P7H ...... 86 Figure 12.16 – BlankWAR4 ...... 88 Figure 12.17 – BlankWAR7 ...... 89 Figure 12.18 – Laboratory Repeats ...... 90 Figure 12.19 – Field Duplicates ...... 91 Figure 12.20 – Laboratory Duplicates ...... 92 Figure 13.1 – SOX Master Composite Gravity Performance ...... 108 Figure 13.2 – MOX Master Composite Gravity Performance ...... 108 Figure 13.3 – Residue Grade by Size Fraction – Grade 2016/17 A16557 ...... 121

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Figure 13.4 – Residue Grade by Size Fraction – Percentage 2016/17 A16557 ...... 121 Figure 13.5 – WOX Master Composite Gravity Performance 2016/17 A16557 ...... 123 Figure 13.6 – FRS Master Composite Gravity Performance 2016/17 A16557 ...... 123 Figure 13.7 – Residue Grade and Sulphide Influence A17341...... 133 Figure 13.8 – MOX Flocculant Screening Results – Extract from Outotec Report ...... 138 Figure 13.9 – Gold Inclusion – First Image ...... 140 Figure 13.10 – Gold Inclusion – Second Image ...... 141 Figure 13.11 – Gold Inclusion – Third Image ...... 141 Figure 13.12 – Grind Sensitivity, M4 MOX 6escription ...... 143 Figure 13.13 – Grind Sensitivity, WOX Master Composite ...... 143 Figure 13.14 – Grind Sensitivity, FRS Master Composite ...... 144 Figure 13.15 – FRS Master Composite Oxygen and NaCN Sensitivity – 75µm ...... 146 Figure 13.16 – FRS Master Composite Oxygen and NaCN sensitivity – 90µm ...... 147 Figure 13.17 – FRS Master Composite Variable Sensitivity ...... 147 Figure 13.18 – WOX Master Composite Oxygen and NaCN Sensitivity ...... 148 Figure 13.19 – WOX Master Composite Oxygen Sensitivity ...... 148 Figure 13.20 – FRS Composites – Averaged Data ...... 149 Figure 13.21 – Select Gravity Recovery Data ...... 151 Figure 13.22 – SOX Free Gold Recovery ...... 152 Figure 13.23 – SOX Free + Intensive Leach Gold Recovery ...... 153 Figure 13.24 – MOX Free Gold Recovery ...... 153 Figure 13.25 – MOX Free + Intensive Leach Gold Recovery ...... 154 Figure 13.26 – WOX Free Gold Recovery ...... 154 Figure 13.27 – WOX Free + Intensive Leach Gold Recovery ...... 155 Figure 13.28 – FRS Free Gold Recovery ...... 155 Figure 13.29 – FRS Free + Intensive Leach Gold Recovery ...... 156 Figure 13.30 – SOX Reagent Consumption ...... 157 Figure 13.31 – MOX Reagent Consumption ...... 158 Figure 13.32 – WOX Reagent Consumption ...... 158 Figure 13.33 – FRS Reagent Consumption ...... 159 Figure 13.34 – Grind Sensitivity and Sulphide Relationship ...... 161 Figure 13.35 – Grind Sensitivity, FRS Master Composite (106µm) ...... 162 Figure 13.36 – Grind Sensitivity, FRS Master Composite (106µm) ...... 162 Figure 13.37 – Grind Sensitivity, FRS Master Composite (75µm) ...... 163 Figure 13.38 – Grind Sensitivity, FRS Master Composite (75µm) ...... 163 Figure 13.39 – Residue Grade as a Function of Calculated Head Grade - FRS ...... 164 Figure 13.40 – Residue Grade as a Function of Leach Feed Grade - FRS ...... 165 Figure 13.41 – FRS Family of Curves – Residue Ratio ...... 166 Figure 13.42 – FRS Algorithm Actual versus Modelled Leach Feed Grade ...... 167 Figure 13.43 – FRS Algorithm Actual versus Modelled Calculated Head Grade ...... 167 Figure 13.44 – WOX Algorithm Actual versus Modelled Calculated Head Grade ...... 168 Figure 14.1 – Plan View of Drilling ...... 173 Figure 14.2 – Isometric NE View of Mankarga 1 ...... 175 Figure 14.3 – M1 South Section SE 0450 ...... 175 Figure 14.4 – Typical Cross Section (Section NE 350 local grid) of Mineralisation Domains ...... 177 Figure 14.5 – Isometric SE View of Mineralisation Domains ...... 178 Figure 14.6 – Log Histograms of Uncut Gold Grade by Domain - Mankarga 1 ...... 179 Figure 14.7 – Log Histograms of Uncut Gold Grade by Domain - Mankarga 5 ...... 181 Figure 14.8 – Mankarga 1 South Variogram ...... 187 Figure 14.9 – Mankarga 1 South High Grade Variogram ...... 187 Figure 14.10 – M3 West Variogram ...... 188 Figure 14.11 – Zone 200 – Back-transformed Gaussian Variogram ...... 191

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Figure 14.12 – Zone 300 – Back-transformed Gaussian Variogram ...... 194 Figure 14.13 – Zone 400 – Back-transformed Gaussian Variogram ...... 194 Figure 14.14 – Plan View of Drilling ...... 207 Figure 14.15 – Isometric NE View of Mankarga 1 ...... 209 Figure 14.16 – M1 North Section SW 0200 (Zone 700) ...... 210 Figure 14.17 – M1 South Section SE 0450 ...... 211 Figure 14.18 – Isometric SSW View of Mankarga 3 ...... 211 Figure 14.19 – Typical Cross Section (Section NE 350 local grid) of Mineralisation Domains ...... 213 Figure 14.20 – Isometric SE View of Mineralisation Domains ...... 214 Figure 14.21 – Log Histograms of Uncut Gold Grade by Domain - Mankarga 1 ...... 216 Figure 14.22 – Log Histograms of Uncut Gold Grade by Domain - Mankarga 3 ...... 216 Figure 14.23 – Log Histograms of Uncut Gold Grade by Domain - Mankarga 5 ...... 217 Figure 14.24 – Mankarga 1 North Variogram ...... 223 Figure 14.25 – Mankarga 1 North Variogram ...... 224 Figure 14.26 – Mankarga 1 South High Grade Variogram ...... 224 Figure 14.27 – M3 East Variogram ...... 226 Figure 14.28 – M3 West Variogram ...... 226 Figure 14.29 – M3 West Variogram ...... 227 Figure 14.30 – Zone 200 – Back-transformed Gaussian Variogram ...... 228 Figure 14.31 – Zone 300 – Back-transformed Gaussian Variogram ...... 232 Figure 14.32 – Zone 400 – Back-transformed Gaussian Variogram ...... 232 Figure 15.1 – Mankarga 1 North Deposit, Summary Optimisation Results ...... 248 Figure 15.2 – Mankarga 1 South Deposit, Summary Optimisation Results ...... 248 Figure 15.3 – Mankarga 3 Deposit, Summary Optimisation Results ...... 249 Figure 15.4 – Mankarga 5 Deposit, Summary Optimisation Results ...... 249 Figure 15.5 – M5 Pit Design showing Initial Pit Stages and Waste Rock Dump ...... 255 Figure 15.6 – M1 and M3 Pit and Waste Rock Dump Designs ...... 255 Figure 17.1 – Sanbrado Process Plant Flow Diagram ...... 261 Figure 17.2 – Mining Schedule Feed Blend ...... 262 Figure 18.1 – Existing Gravel Road from Zempasgo ...... 277 Figure 18.2 – Site Development ...... 278 Figure 18.3 – Proposed Process Plant Site ...... 279 Figure 18.4 – Proposed Gibgo River Water Abstraction Area ...... 281 Figure 18.5 – Project Organisation Chart ...... 290 Figure 18.6 – General Management Organisation Chart ...... 297 Figure 18.7 – Mining Organisation Chart ...... 298 Figure 18.8 – Process Organisation Chart ...... 298 Figure 18.9 – Maintenance Organisation Chart ...... 299 Figure 18.10 – Commercial and Administration Organisation Chart ...... 301 Figure 18.11 – Sustainability Organisation Chart ...... 302 Figure 18.12 – Security Organisation Chart ...... 302 Figure 22.1 – Historical Gold Price ...... 350 Figure 22.2 – Project, Post-Tax, All Equity Basis PVNCF5% ($M) - Sensitivity to Key Input Parameters ...... 357 Figure 22.3 – Project, Post-Tax, All Equity Basis IRR (% pa) - Sensitivity to Key Input Parameters ...... 357

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

List of Tables

Table 1.1 – Sanbrado Gold Project Open Pit Feasibility Study Highlights ...... 1 Table 1.2 – Design Values Based on Testwork ...... 10 Table 1.3 – Mineral Resource Statement, Sanbrado Gold Project, August 1, 2016 ...... 16 Table 1.4 – Selected Optimum Pit Shells ...... 18 Table 1.5 – Average Calculated Cutoff Grade (g/t Au) ...... 19 Table 1.6 – Pit Inventories ...... 19 Table 1.7 – Mine Production Schedule ...... 21 Table 1.8 – Sustaining Capital Estimate ...... 29 Table 1.9 – LOM Operating Costs ...... 29 Table 1.10 – Capital Cost Summary ...... 29 Table 1.11 – Operating Costs During the Production Period ...... 31 Table 10.1 – Summary of AC, RC and DC Drilling at the Mankarga Prospects ...... 57 Table 10.2 – WAF AC, RC and DC Drilling included in Resource Estimate to 11 December 2016 ...... 61 Table 12.1 – WAF Internal Standards ...... 71 Table 12.2 – WAF Blanks ...... 87 Table 13.1 – Metallurgical Testwork Hole Locations and Intervals ...... 96 Table 13.2 – Composite Details (2014-A15684) ...... 99 Table 13.3 – Summary Leach Results (2014 -A15684) ...... 99 Table 13.4 – Abrasion and Ball Mill Work Index for MOX and FRS (2014-A15684) ...... 100 Table 13.5 – Gold Head Assay Data 2015 - A16326 ...... 102 Table 13.6 – UCS Results 2015 -A16326 ...... 103 Table 13.7 – Crushing Work Indices 2015-A16326 ...... 103 Table 13.8 – Abrasion Indices 2015-A16326 ...... 104 Table 13.9 – Assay of Master Composites A16326 ...... 105 Table 13.10 – Rod and Ball Mill Work Indices 2016 - A16326 ...... 105 Table 13.11 – Levin Test Results on SOX Material 2016 - A16326 ...... 105 Table 13.12 – Master Composite Grind Sensitivity 2016 - A16326 ...... 106 Table 13.13 – Master Composite Bulk Gravity-Leach 2016 - Test A16326 ...... 107 Table 13.14 – Master Composite Bulk Gravity-Leach Test 2016 - A16326 ...... 109 Table 13.15 – Fleming Kinetic Parameter Determination 2016 - A16326 ...... 109 Table 13.16 – Carbon Equilibrium Parameter Determination 2016 - A16326 ...... 109 Table 13.17 – Oxygen Uptake Rate Determination 2016 - A16326 ...... 110 Table 13.18 – Master Composite Viscosity Data 2016 - A16326 ...... 110 Table 13.19 – Master Composite Variability Testing 2016 - A16326 ...... 111 Table 13.20 – UCS Testwork Summary 2016/17 A16557 ...... 113 Table 13.21 – Crushing Work Index Summary 2016/17 A16557...... 114 Table 13.22 – Bond Abrasion Test Summary 2016/17 A16557...... 114 Table 13.23 – Bond Rod and Ball Mill Work Index Testwork Summary 2016/17 A16557 ...... 115 Table 13.24 – Composite Head Assay Data 2016/17 A16557 ...... 115 Table 13.25 – Grind Size Optimisation Testwork 2016/17 A16557 ...... 116 Table 13.26 – NaCN Sensitivity Testwork 2016/17 A16557 ...... 117 Table 13.27 – Influence of High Pulp Density 2016/17 A16557 ...... 117 Table 13.28 – Influence of CIL Leaching 2016/17 A16557 ...... 118 Table 13.29 – Influence of Oxygenation versus Air 2016/17 A16557 ...... 118 Table 13.30 – Influence of Oxygenation versus Air 2016/17 A16557 ...... 118 Table 13.31 – Intensive Leaching of Gravity Concentrate 2016/17 A16557 ...... 119 Table 13.32 – Variability Testing 2016/17 A16557 ...... 120 Table 13.33 – High Grade Residue Re-Leach Testwork 2016/17 A16557 ...... 120 Table 13.34 – Variability Sample Optimisation Leach Series 2016/17 A16557 ...... 122

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Table 13.35 – WOX and FRS Sequential Carbon Triple Contact Testwork 2016/17 A16557 ...... 124 Table 13.36 – WOX and FRS Carbon Equilibrium Loading Testwork 2016/17 A16557 ...... 124 Table 13.37 – WOX and FRS Oxygen Uptake Testwork 2016/17 A16557 ...... 125 Table 13.38 – WOX and FRS Slurry Rheology Testwork 2016/17 A16557 ...... 125 Table 13.39 – Head Assays A17283 ...... 126 Table 13.40 – M1 Composite Gravity / Leach /Results A17283 ...... 127 Table 13.41 – UCS Summary 2016/7 - A17341 ...... 129 Table 13.42 – Crushing Work Index Determinations A17341 ...... 129 Table 13.43 – SMC Determinations 2016/7 - A17341 ...... 130 Table 13.44 – Abrasion Index Determinations 2016/7 - A17341 ...... 130 Table 13.45 – Bond Rod and Ball Work Index Determinations 2016/7 - A17341 ...... 131 Table 13.46 – Head Assay Results 2016/7 - A17341 ...... 132 Table 13.47 – Summary Gravity-Leach Data 2016/7 - A17341 ...... 132 Table 13.48 – Diagnostic Leach Results 2016/7 - A17341 ...... 134 Table 13.49 – Elevated NaCN Gravity Leach Tests (106µm) 2016/7 - A17341 ...... 134 Table 13.50 – High Grade Composite Head Assays 2016/7 - A17750 ...... 135 Table 13.51 – High Grade Gravity/Leach Summary Results 2016/7 - A17750 ...... 136 Table 13.52 – High Grade Diagnostic Leach Results RC Comps 2016/7 - A17750 ...... 137 Table 13.53 – High Grade Diagnostic Leach Results DD Comps 2016/7 - A17750 ...... 137 Table 13.54 – Grind Sensitivity ...... 142 Table 13.55 – Varying Test Conditions ...... 145 Table 13.56 – Cyanide Sensitivity – FRS, MOX and WOX Samples ...... 145 Table 13.57 – FRS Composite Air/Oxygen Data ...... 149 Table 13.58 – Influence of Residence Time ...... 150 Table 13.59 – Reagent Consumption Considerations ...... 159 Table 13.60 – Influence of Sulphides ...... 160 Table 13.61 – Design Values Based on Testwork ...... 171 Table 14.1 – Summary Statistics Grouped by Domain for 2m and 3m Composites of Uncut Gold Grade (g/t) ...... 180 Table 14.2 – Summary Statistics Grouped by Domain for Composites of Uncut, Declustered Gold Grade (g/t) ...... 182 Table 14.3 – Composite Top Cut Statistics - M1 South High Grade Domains ...... 183 Table 14.4 – Summary Statistics Grouped by Domain for 2m and 3m Composites of Top cut Gold Grade (g/t) ...... 183 Table 14.5 – Summary Statistics Grouped by Drillhole Type ...... 184 Table 14.6 – Indicator Class Statistics M1 ...... 185 Table 14.7 – Indicator Class Statistics M5 ...... 185 Table 14.8 – M1 South Variogram Models ...... 189 Table 14.9 – M5 Zone 100 Variogram Models ...... 190 Table 14.10 – M5 Zone 200 Variogram Models ...... 192 Table 14.11 – M5 Zone 300 Variogram Models ...... 193 Table 14.12 – M1 Variogram Models for OK Estimates ...... 195 Table 14.13 – M3 Variogram Models for OK Estimates ...... 195 Table 14.14 – Block Model Parameters ...... 196 Table 14.15 – Measured Average Bulk Density Value ...... 197 Table 14.16 – Measured Average Bulk Density Value ...... 198 Table 14.17 – M1 South MIK Sample Search Criteria ...... 201 Table 14.18 – M1 South MIK Sample Search Criteria ...... 201 Table 14.19 – Mankarga 1 South OK Sample Search Criteria ...... 202 Table 14.20 – Mankarga 5 OK Sample Search Criteria ...... 202 Table 14.21 – Variance Adjustment Ratios ...... 204 Table 14.22 – Confidence Levels by Key Criteria ...... 205 Table 14.23 – Mineral Resource Statement, Sanbrado Gold Project, 13th December 2017 ...... 206 Table 14.24 – Summary Statistics Grouped by Domain for 2m and 3m Composites of Uncut Gold Grade (g/t) ...... 215 Table 14.25 – Summary Statistics Grouped by Domain for 3m Composites of Uncut, Declustered Gold Grade (g/t) ...... 218

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Table 14.26 – Summary Statistics Grouped by Domain for 2m and 3m Composites of Uncut Gold Grade (g/t) ...... 219 Table 14.27 – Summary Statistics Grouped by Drillhole Type ...... 220 Table 14.28 – Indicator Class Statistics M1 ...... 221 Table 14.29 – Indicator Class Statistics M5 ...... 221 Table 14.30 – M1 North Variogram Models ...... 225 Table 14.31 – M1 South Variogram Models ...... 229 Table 14.32 – M5 Zone 100 Variogram Models ...... 230 Table 14.33 – M5 Zone 200 Variogram Models ...... 231 Table 14.34 – M5 Zone 300 Variogram Models ...... 233 Table 14.35 – M1 Variogram Models for OK Estimates ...... 234 Table 14.36 – M3 Variogram Models for OK Estimates ...... 234 Table 14.37 – M3 Variogram Models for OK Estimates ...... 234 Table 14.38 – Block Model Parameters ...... 235 Table 14.39 – Measured Average Bulk Density Value ...... 236 Table 14.40 – M1 North MIK Sample Search Criteria ...... 237 Table 14.41 – M1 South MIK Sample Search Criteria ...... 238 Table 14.42 – M1 South MIK Sample Search Criteria ...... 239 Table 14.43 – Mankarga 1 South OK Sample Search Criteria ...... 239 Table 14.44 – Mankarga 3 OK Sample Search Criteria ...... 240 Table 14.45 – Mankarga 5 OK Sample Search Criteria ...... 240 Table 14.46 – Variance Adjustment Ratios (5mE x 12.5mN x 5mRL SMU) ...... 241 Table 14.47 – Measured Average Bulk Density Value ...... 242 Table 14.48 – Mineral Resource statement, Sanbrado Gold Project, August 1, 2016 ...... 243 Table 15.1 – Mineral Reserve by Category ...... 245 Table 15.2 – Selected Optimum Pit Shells ...... 247 Table 15.3 – Mankarga 5 Pit Slope Angles ...... 250 Table 15.4 – Mankarga 1 and 3 Pit Slope Angles ...... 250 Table 15.5 – Mankarga 5 Average Contract Mining Costs ...... 251 Table 15.6 – Mankarga 1 & 3 Average Contract Mining Costs ...... 251 Table 15.7 – Process and Fixed Costs ...... 252 Table 15.8 – Process Recovery ...... 253 Table 15.9 – Process Recovery for M1 and M5 Sulphide Ore ...... 253 Table 15.10 – Process Recovery for M1 and M5 Sulphide Ore ...... 254 Table 15.11 – Average Calculated Cutoff Grade (g/t Au) ...... 254 Table 15.12 – Comparison of Optimised Pit Shells and Pit Designs ...... 254 Table 15.13 – Pit Inventories ...... 256 Table 16.1 – Mine Production Schedule ...... 259 Table 17.1 – Comminution Circuit Operating Design Criteria ...... 263 Table 17.2 – Ore Specific Comminution Parameters ...... 263 Table 17.3 – SAG Mill Specifications ...... 264 Table 18.1 – Estimated Power Demand ...... 286 Table 18.2 – Operational Manning Summary ...... 296 Table 18.3 – Labour Classification System ...... 296 Table 18.4 – Total Project Manning Level Year 1 ...... 303 Table 18.5 – Personnel Breakdown Year 1 Onwards ...... 303 Table 18.6 – Personnel Breakdown Year 3 Onwards ...... 303 Table 18.7 – Personnel Breakdown Year 6 Onwards ...... 304 Table 20.1 – Communities Affected by the Project Operating License Boundaries ...... 320 Table 21.1 – Average LOM Operating Cost Summary ...... 326 Table 21.2 – Processing and Maintenance Costs ...... 327 Table 21.3 – Annual Labour Cost ...... 328 Table 21.4 – Labour Classification System ...... 329

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

Table 21.5 – Power Costs ...... 330 Table 21.6 – Consumable Costs ...... 330 Table 21.7 – Maintenance Costs ...... 331 Table 21.8 – Estimated General & Administration Cost ...... 332 Table 21.9 – Sustaining Capital Estimate ...... 334 Table 21.10 – LOM Operating Costs...... 336 Table 21.11 – Capital Cost Summary ...... 337 Table 21.12 – Capital Cost Estimate Contingencies ...... 345 Table 21.13 – Exchange Rates (Q12017) ...... 346 Table 21.14 – Capital Costs by Discipline ...... 347 Table 22.1 – Mine Production and Mill Feed Schedule ...... 349 Table 22.2 – Operating Costs During the Production Period ...... 351 Table 22.3 – Operating Costs During the Production Period ...... 354 Table 22.4 – Base Case Cashflow Details ...... 355 Table 22.5 – Project Economics Sensitivity to Gold Price ...... 356 Table 26.1 – Recommendations and Associated Costs ...... 361

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso West African Resources Limited

EXECUTIVE SUMMARY Introduction West African Resources (WAF) is a gold exploration company based in Perth, Australia, dedicated to creating shareholder value through the acquisition, exploration and development of resource projects in Burkina Faso, West Africa. Its exploration assets consist of several Burkina Faso properties including the Sanbrado Gold Project, previously the Tanlouka Gold Project.

This report relates only to updated resource estimates for the M1 South and M5 deposits. Mineral Reserves and associated economic data including operational and capital expenses have been restated for completeness and remain unchanged from the February 2017 open pit feasibility study. Work is on-going on the Sanbrado Gold Project, and the Company expects to further update mineral resources in early 2018, followed by completion of a Definitive Feasibility Study incorporating both open-pit and underground mining in mid-2018.

WAF acquired a 90% interest in the Sanbrado property through the acquisition of Channel Resources Ltd on 21 January 2014, and acquired the remaining 10% of the project from GMC SARL, a Burkina Faso registered entity, in September 2015.

The Sanbrado Gold Project incorporates one granted exploration licence and one granted mining licence covering some 116km2.

The Burkina Faso Ministry of Mines and Energy grants an Exploration Permit for a three year term, with option to renew for two further three year terms. An Exploitation Permit is valid for 10 years. The Burkina Government maintains a 10% free carry interest in the project and a free on board royalty of 4% for base metals and 3% for precious metals.

Since February 2017, WAF has continued exploration activities and has significantly enlarged the resource base, which now includes fresh rock resources and higher grade zones at M1 South. This Report details the updated resource model for the M1 South and M5 deposits and restates the optimised mining and processing plans reported in February 2017. Mineral Reserves and associated project economics have not been updated in this report. This work is in progress and is expected to be completed in mid-2018. The updated resource as well as all other necessary components of project definition have been compiled into this report (the Sanbrado Gold Project).

Property Description and Location The Sanbrado Gold Project is located in the Burkina Faso, West Africa, approximately 90km east southeast of the capital Ouagadougou, and approximately 20km south of the regional centre of Mogtedo (Figure 1.1).

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Figure 1.1

Sanbrado Gold Project Location

Accessibility, Climate, Local Resources, Infrastructure and Physiography The Sanbrado Gold Project is located adjacent to the major towns of Boulsa and Zorgho and is well serviced by grid power and mobile phone coverage. The project is accessed via the bitumen RN4 highway, which links Burkina Faso and Niger and crosscuts the southwestern corner of the project area.

The Sanbrado site is accessible by lateritic secondary roads from Mogtedo and Zempsago, a distance of about 25km. The secondary roads are relatively flat and accessible for most of the year.

The project is located south of the sub-Saharan Sahel of West Africa, and vegetation is predominantly open-forested savannah and grassland. The regional climate is sub-tropical and semi-arid, with warm, dry winters and hot, wet summers. The wet season extends from mid- April until mid-October. Annual rainfall is generally between 700mm and 1,000mm. Peak rainfall months are August, September and October. Daytime temperatures range from 35˚C to 45˚C in the dry season and from 30˚C to 35˚C in the wet season.

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The property is flat to gently rolling. Crops under cultivation include millet, rice, cotton, peanuts and maize. Wildlife is limited to livestock, birds and reptiles.

History Channel Resources commenced exploration work over the Tanlouka area in 1994. In the period from 1994 to 1996, Channel undertook several reconnaissance style mapping and prospecting traverses.

Regional soil sampling was initiated in 1996 which covered the Tanlouka permit at 500m centres in a triangular grid. Anomalous gold-in-soil zones were defined over areas of 2km by 5km at Mankarga and Koloba Samba (Guérard, 1997). More detailed soil sampling in 2000 resulted in the definition of a continuous gold soil anomaly over 3km long and 200m to 300m wide in the southern part of the Mankarga East grid.

Since 2000, various drilling programmes have been completed on the Sanbrado permit. Reverse Air Blast (RAB) drilling was undertaken over the soil anomalies at Sanbrado in April 2000. In 2010, Channel completed wide-spaced reverse circulation (RC) drilling at Mankarga 1, Mankarga 2, Mankarga 3, Mankarga 4 and Mankarga 5. Follow-up RC drilling commenced in November 2010 through to April 2011, with most holes drilled at Mankarga 5. Diamond Core (DC) drilling commenced in July 2011 on the Mankarga 5 deposit and was completed in February 2012.

A mineral resource estimate was prepared for the Mankarga 5 Zone by AMEC in 2012.

Since the merger between Channel and WAF in 2013, WAF has carried out a number of drilling and project development programs, concentrating on the Mankarga series of deposits and culminating in this Feasibility Study.

Geological Setting and Mineralisation The gold deposits of West Africa largely lie within the Proterozoic domain of the Man Shield, the southernmost subdivision of the West African Craton. The principal gold producing areas are associated with the Lower Proterozoic systems of the Birimian (2.17-2.18 billion years) comprising metavolcanic (arc) and metasedimentary (basin) rocks, and the marginally younger, unconformably overlying rocks of the Tarkwaian epiclastic system.

The spatial distribution of gold mineralisation in West Africa (Figure 1.2) is closely governed by north to northeast trending belts of Upper Birimian metavolcanic rocks, ranging from 15km to 40km in width. Almost without exception, the major gold deposits lie at, or close to the margins of the belts in proximity to the strongly deformed contacts between the Upper and Lower Birimian sequences.

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Figure 1.2 Gold Deposits of West Africa on Regional Geology

The Sanbrado Gold Project is located within a strongly arcuate volcano-sedimentary northeast trending belt that is bounded to the east by the Markoye Fault. The Markoye Fault is interpreted to be the main conduit for the gold mineralisation found at Sanbrado, accommodating large scale fluid migration.

The generalised lithologic sequence of the Sanbrado area is characterised by minor metabasalts, metasedimentary and volcanosedimentary rocks of the Birimian System, intruded by Eburnean granodiorites, as well as granitic, dioritic and mafic intrusions.

The metasedimentary package has been deformed and variably metamorphosed under greenschist and amphibolite facies conditions, resulting in schists, psammites and pelites. The metasedimentary package is largely constrained by the felsic intrusives.

Sanbrado hosts shear zone type quartz-vein gold mineralisation, similar to that found elsewhere in late Proterozoic Birimian terrains of West Africa. These orogenic type deposits exhibit strong relationships with regional arrays of major shear zones.

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The structural environment can be described as brittle-ductile, coincident with a greenschist-amphibolite metamorphic facies. High grade mineralisation occurs along the fold hinges and is marked by an intense silica-albite alteration assemblage. Gold is free and is mainly associated with minor pyrite and pyrrhotite disseminations and stringers that are foliation controlled.

Known mineralisation at Mankarga 1 (M1) extends along strike for approximately 1km, is up to 50m wide and 150m in depth. The Mankarga 5 (M5) mineralisation extends along strike for approximately 3km, is up to 100m wide and 300m in depth. The Mankarga 3 (M3) mineralisation extends along strike for 750m, is up to 50m wide and 75m in depth. Mineralisation at all prospects remains open at depth.

Deposit Types Most West African gold deposits occur in the Lower Proterozoic part of the Craton, in sedimentary and volcano-sedimentary formations, which have undergone multiphase deformation, a setting analogous to many other gold-producing Precambrian greenstone belts in other parts of the world, such as the Abitibi Belt in Canada and the Norseman- Wiluna Belt in Australia.

The West African Lower Proterozoic greenstone belts host a number of world class gold deposits such as Obuasi in Ghana (>20Moz), Sadiola in Mali (8Moz) and Morila in Mali (5.9Moz).

Deposits comprise numerous mineralisation styles, including quartz reefs hosted within frequently carbonaceous phyllites and greywackes associated with major semi-conformable shear structures and subsidiary oblique faults. Lower grade mineralisation may also be present as disseminations or associated with sheeted quartz veining within tuffs, greywackes and mafic dykes situated in proximity to major structures.

Gold mineralisation is also associated with sheeted vein swarms and stockwork zones within granitoids. These deposits are typically lower grade than reef style mineralisation and appear to be confined to the smaller belt-type or Dixcove Suite granitoids and their regional equivalents.

Exploration Exploration activities on the Sanbrado permit have included geological mapping, soil sampling, rock and chip sampling, geophysical surveys and drilling. Drilling programs have included auger, air core (AC), reverse circulation (RC) and diamond core (DC) drilling techniques.

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From 1994 to 2012, all exploration activities were completed by Channel personnel and their consultants. From early 2014, exploration has been undertaken by WAF following its acquisition of Channel and the Sanbrado permit, which was part of the Channel tenement portfolio.

No exploration activities of substance were undertaken from mid-2012 until early 2014 when WAF began an infill drilling program at the M5 prospect. WAF also completed an auger drilling program encompassing the entire original Tanlouka permit on a 400m by 100m grid to validate existing anomalies generated via historic soil sampling.

Drilling Drilling completed by Channel included two phases of RC drilling (Jun 2010 to April 2011) and a DC drilling programme (Phase 3, Jul 2011 to Feb 2012). Since 2014, all drilling work has been managed directly by WAF. Only Channel’s Phase 1, 2 and 3 drill programs, as well as AC, RC and DC drilling by WAF have been used for estimating mineral resources.

The area of the Mankarga 5 resource was drilled using Reverse Circulation (RC), Aircore (AC) and Diamond core (DC) and RC PreCollar/Diamond Tail (DT) drillholes on a nominal 50m x 25m grid spacing with infill to 25m x 25m in the far south portion. A total of 723 AC holes (23,314m), 71 DC holes (14,537m), 41 DT holes (8,739m) and 69 RC holes (7,078m) were drilled by WAF between 2013 and 2017 (as at 11 October 2017). A total of 60 RC holes (7,296.2m) and 71 DC holes (15,439.6m) were drilled by Channel Resources Ltd (CHU) between 2010 and 2012. Holes were angled towards 120° or 300° magnetic at declinations of between -50° and -60°, to optimally intersect the mineralised zones.

The area of the Mankarga 1 South resource was drilled using RC, AC and DC drillholes on a nominal 25m x 20m grid spacing. A total of 217 AC holes (3,290m), 42 DC holes (12,041m), 52 DT holes (14,113m) and 115 RC holes (12,602m) were drilled by WAF between 2015 and 2017. A total of 23 RC holes (3,060.0m) and 7 DC holes (1,199.0m) were drilled by CHU between 2010 and 2012. Holes were angled towards 020°, 045°, 180°or 225° magnetic at declinations of between -50° and -60°, to optimally intersect the mineralised zones.

A drillhole summary plan is given in Figure 1.3.

A hand-held GPS was used to locate and prepare the pads for the planned holes. Drillhole collars were surveyed using a DGPS in UTM grid WGS84 Zone 30 North. Downhole surveys comprised single shot readings and were taken during drilling at 25m depth and every 50m thereafter.

All drilling is logged to a standard that is appropriate for the category of resource which is being reported.

Dry bulk densities were completed by WAF (carried out internally) and Channel (carried out by SGS laboratories). Both utilised industry standard core immersion techniques.

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Figure 1.3 Sanbrado Gold Project – Drillhole Summary Plan

Sample Preparation, Analyses and Security All sample preparation and analyses were carried out at independent laboratories in Ouagadougou, Burkina Faso.

Channel submitted a total of 555 samples from the Phase 1 RC program and a total of 3,446 samples from the Phase 2 RC program to Abilab Burkina SARL (Abilab). A total of 11,128 samples from the Channel Phase 3 DC program were assayed, primarily at the SGS Burkina Faso SA (SGS) in Ouagadougou.

A total of 18,825 RC samples and 5,602 DC samples (including QAQC samples) from the WAF drilling programs were collected and submitted to BIGS Global Burkina SARL (BIGS).

No aspect of laboratory sample preparation or analysis was conducted by any employee, officer, director or associate of WAF. Chain of custody and sample security protocols used by Channel and WAF comply with international standards.

WAF and predecessors have used a combination of duplicates, checks, blanks and standards to ensure quality control of sampling methods and assay testing. Procedures and QA/QC management are consistent with good industry practice and are deemed fit for purpose.

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Data Verification Mr Brian Wolfe is the Qualified Person responsible for the Mineral Resource estimates for the Sanbrado Gold Project. Mr Wolfe visited the Project in May 2014, May 2016 and April 2017. Steps undertaken to verify the integrity of data used in this report include:

. Field visits to the areas described in this Report, including the Sanbrado permit and M1, M3 and M5 deposits; . Inspection of drill core and RC drill chips; . Inspection of AC, RC and DC drilling activities, sampling and logging procedures; . Review of WAF’s data collection, database management and data validation procedures; and . Review of the previous NI 43-101 report for the project titled “NI 43-101 Technical Report on Mineral Resources for the Mankarga 5 Gold Deposit Sanbrado Property” dated September 2014. Additional data verification steps undertaken during the resource estimate process include:

. Validation of drilling, geology and assay database, including checks overlapping intervals, samples beyond hole depth and other data irregularities; . Review of WAF QAQC charts for standards, blanks and duplicates; . Visual and statistical analysis of resource estimate model outputs versus primary data; and . Random cross checks of assay hardcopy reports against the database. Based on this review work, the Qualified Person is of the opinion that the dataset provided by WAF is of an appropriate standard to use for resource estimation work.

Mineral Processing and Metallurgical Testing WAF has completed a metallurgical testwork program designed to determine the amenability of the various feed materials to conventional gravity separation followed by agitated leaching and the associated carbon in pulp/carbon in leach (CIP/CIL) processing methods.

Metallurgical samples were selected from the M1 and M5 deposits and subjected to testwork programmes. No metallurgical testwork was conducted on the M3 deposit, which comprises less than 2% of the resource, as the mineralogy and assaying from the geological drilling indicated similar mineralogical behaviour as oxide material from deposits M1 and M5.

More than 100 separate gravity/leach tests have been conducted on a range of samples. Variations in grind sizes, cyanide concentrations, oxygen sources and leaching times were trialled. Comminution work was conducted to evaluate grinding options, such as Semi- Autogenous Grinding (SAG) methods and most of the testwork was undertaken by ALS in Perth. Thickening testwork was undertaken by Outotec.

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The Sanbrado materials tested present as free-milling, providing high gold leach extractions and medium gold leaching rates with relatively low reagent consumption. The liberation size required is typical for such ores and a conventional grinding flowsheet applies. Moderate levels of free gold can be recovered using conventional centrifugal type concentrators.

The free gold component is amenable to gravity recovery. The bulk of the non-free gold is contained in non-sulphide minerals as composite or complex particles. Some gold is associated with sulphides and can be expected to result in variable leaching rates, due to un-liberated gold and possible lock-up in final residues.

Composite samples were developed to reflect a spread of head grades, mineralogy, host lithology, and spatial and physical location. There does not appear to be any significant difference in metallurgical performance as a function of location or lithology. However, there are differences in performance for the four metallurgical (oxidation state) domains identified, namely SOX, WOX, MOX and FRS.

The materials do not show high oxygen demands, nor do they present problematic rheological characteristics. Thickening testwork was successful but required high rates of flocculant. However, the flocculant used in the test work is considered sub-optimal and there is opportunity to improve in this area.

The progression in oxidation from SOX, MOX, WOX to FRS comes with a general increase in ore hardness and toughness, along with a reduction of total gold recovery when applied to the proposed flowsheet. This is typical behaviour for such deposits. However, FRS recoveries remain in the mid-80% range across the general head grade range.

Testwork undertaken to date supports application of the proposed processing technology to oxidised, partially oxidised and primary feed types, with low reagent consumptions and moderate leach times. Design-specific testing will be undertaken to optimise detailed design parameters and specifically address the behaviour of deeper high grade samples from the M1 deposit.

Testwork indicates that the range of gravity recovery values in the full-scale operation can be expected to be as follows:

. SOX: 5% to 10%. . MOX, WOX and FRS: 10% to 20%, with a design consideration for electrowinning (FRS) at the top of the range. SAG milling characterisation testwork suggests that the FRS ore is very hard and energy intensive. Milling circuit design will consider the practicality of blending ore types and the risk profile associated with single stage SAG milling on such materials.

Additional test work is in progress and will be updated in the Definitive Feasibility Study technical report due for completion in mid-2018.

Table 1.2 presents a summary of proposed key design criteria based on testwork results.

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Table 1.2 Design Values Based on Testwork

Design SOX MOX WOX Fresh (50% SOX/MOX, 50% WOX/FRS) Gravity Gold Recovery, nominal 7.5% 15% 15% 20% 20% Leach Gold Recovery (2.8g/t Au) 87% 79% 77% 69% 87% Total Gold Recovery 95% 94% 92% 89% 95% Leach Feed pH 10.5 Oxygen Consumption kg/t 0.05 0.05 0.058 0.068 0.2 Loaded Carbon Grade 3,500 Tailings Thickener Settling Rate t/m².h 0.7 Flocculent consumption g/t 70 50 40 40 70 Lime Consumption kg/t 2.5 1.1 0.65 0.5 3.5 Cyanide Consumption kg/t 0.2 0.2 0.15 0.15 0.35

Mineral Resource Estimates The Mineral Resource Estimate section is presented in two parts, the first relates to the mineral resource updates for M1 South and M5 as at the effective date of the report and the second relates to the previous Mineral Resource Estimates as at February 20th 2017. The previous Mineral Resource Estimate is described as the Mineral Reserve Estimates and subsequent sections of the report are based on the February 20th 2017 Mineral Resource and have not been updated to reflect the updates to the M1 South and M5 Mineral Resources.

The Mineral Resource Estimate for the Sanbrado Gold Project has been updated as at the effective date of the 13th December 2017. Grade estimates were updated for, M1 South and M5 (Figure 1.4). Gold grade estimation was completed using a combination of Multiple Indicator Kriging (MIK) and Ordinary Kriging (OK). This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralisation, and the style, geometry and tenor of mineralisation. The estimation was constrained with geological and mineralisation interpretations.

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Figure 1.4 Plan View of Drilling

The resource estimation was based on the available exploration drillhole database which was compiled in-house by WAF. The database was reviewed and validated prior to commencing the resource estimation study.

Oxidation and lithology boundaries were interpreted based on grade information and geological observations, resulting in modelled wireframes used to constrain resource estimations.

At M1 North, mineralisation largely lies within a moderately dipping shear zone corridor (Zone 700) which can be traced over a strike of at least 300m (Figure 1.5). Overall strike direction of the shear zone is approximately 135° with a dip of approximately 60° to the northeast. A smaller, less well mineralised parallel structure exists to the northwest (Zone 800). Mineralised domains were constructed on approximately 25m spaced cross sections orientated perpendicular to drilling using a nominal 0.3g/t Au edge cutoff to define overall shear zone mineralisation. Good continuity of the mineralised envelopes was observed at the scale of the sectional spacing with moderate grade variability typical for deposits of this type.

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Figure 1.5 Isometric NE View of Mankarga 1

Mineralisation at M5 largely lies within a broad, steeply dipping low grade shear zone corridor that can be traced over a strike length of 3,000m (Figure 1.6). The overall strike direction of the shear zone is approximately 035° and is sub vertical to steeply west dipping. Distinctive higher grade hanging wall and footwall lodes can be identified within this broader lower grade halo.

Mineralised domains were constructed on approximately 50m spaced cross sections orientated perpendicular to drilling using a nominal 0.3g/t Au edge cutoff for overall shear zone mineralisation. Good continuity of the mineralised envelopes was observed at the

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scale of the sectional spacing with moderate grade variability, typical for deposits of this type.

Three main mineralisation estimation domains (Zones 100, 200, 300) were defined, with a fourth domain (Zone 400), to the south. Zone 102 is a minor sub-domain located adjacent to and in the hangingwall of Zone 100. Zone 301 is a minor domain located in the footwall of Zone 300.

The composites were used for subsequent statistical, geostatistical and grade estimation investigations. The grade distributions are typical for gold deposits of this style and show a positive skew or near lognormal behaviour. The coefficient of variation is moderate to high, consistent with the presence of high grade outliers that potentially require cutting (capping) for grade estimation.

Visual inspection of the available datasets for each of the estimation domains indicated some clustering of the data within higher grade regions of the deposit. Data clustering often occurs when drilling campaigns selectively target higher grade regions of the deposit, resulting in an artificially high mean grade in many cases. Declustering was completed to remove any effects of preferential sampling of high grade areas that may have occurred.

The variography was calculated and modelled in the geostatistical software, Isatis. The rotations are tabulated as dip and dip direction of major, semi-major and minor axes of continuity. Modelled variograms were generally shown to have good structure and were used throughout the MIK estimation and for the change of support process.

3-D block models were created for the four separate deposit areas in the Universal Transverse Mercator (UTM) grid projection, Zone 30 North, using the World Geodetic System 1984 datum (WGS84). The parent block size was selected on the basis of the average drill spacing. Sub- blocking was applied to an appropriate dimension to ensure adequate volume representation. The models covered all the interpreted mineralisation zones and included suitable additional waste material to allow later pit optimisation studies.

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Figure 1.6 Isometric SE View of Mineralisation Domains M5 Deposit

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An extensive dry bulk density database was supplied containing a total of 6,648 dry bulk density determinations widely dispersed throughout the different mineralisation domains. The database was subdivided on the basis of oxidation, resulting in 4,952 determinations of fresh rock, 593 of transition material and 1,103 of oxide material.

MIK was used as the grade estimation method within the defined mineralisation wireframes at M1 and 5. OK was used as the grade estimation method for the minor domains at M5, the high grade domains at M1 and for all domains at M3. Estimation was completed in the mining package Vulcan using the GSLib geostatistical software. Change of support was applied to the MIK grade estimates to produce ‘recoverable’ gold estimates targeting a selective mining unit (SMU) of 5mE by 12.5mN by 5mRL.

All relevant statistical information was recorded to enable validation and review of the MIK and OK estimates. The recorded information included:

. Number of samples used per block estimate; . Average distance to samples per block estimate; . Estimation flag to determine in which estimation pass a block was estimated; and . Number of drillholes from which composite data were used to complete the block estimate. The estimates were reviewed visually and statistically prior to being accepted. The review included:

. Comparison of the E-type or OK estimate versus the mean of the composite dataset, including weighting where appropriate to account for data clustering; . Comparison of the reconstituted cumulative conditional distribution functions of the MIK estimated blocks versus the input composite data; and . Visual checks of cross sections, long sections, and plans. Comparative estimates using different grade estimation methods were also completed to test the sensitivity of the reported model to selected MIK and OK interpolation parameters. Variation in overall grade noted in the comparative estimations was insignificant.

The resource categorisation was based on the robustness of the various data sources available, including:

. Geological knowledge and interpretation; . Variogram models and the ranges of the first structure in multi-structure models; . Drilling density and orientation; and . Estimation quality statistics. The resource estimates for the Sanbrado Gold Project have been classified as Indicated and Inferred Mineral Resources. Applying these confidence levels, resource classification codes were assigned to the block model using the following criteria:

. Indicated Mineral Resources

 Blocks are predominately estimation pass 1;

 Distance to nearest data of 35m or less;

 Drillhole section spacing predominantly 50m; and

 Appropriate quality of estimate statistics.

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. Inferred Mineral Resources

 Blocks are predominantly estimation pass 2; and

 Blocks not assigned as Indicated. The preferred reporting cutoff grade for the targeted open pit resources is 0.5g/t Au, which was selected based on previous cutoff grade calculations and reporting cutoffs applied to similar gold deposits in the region. The Mineral Resource estimate has also been reported at a 3g/t Au cutoff grade to demonstrate grade tonnage relationships at higher cutoff grades for potential underground mining methods.

The summary total resource for the Sanbrado Gold Project is provided in Table 1.3. The Mineral Resource estimates are predicated on an open pit mining and conventional CIL processing scenario and have been reported within optimised pit shells using a gold price of US$1,650/oz and a cutoff grade of 0.3g/t Au for oxide mineralisation and 0.4g/t Au for fresh rock. Key pit optimisation input parameters are:

. Metallurgical recovery for CIL processing of 95% for oxide and transitional material, and 90% for fresh rock; . Mining costs: $1.50/t oxide; $1.90/t transitional; $2.50/t fresh; . Process costs: $9.00/t for oxide; $12.00/t for transitional and fresh; . Pit slope angles of 45° for oxide and 50° for transitional and fresh; and . Conceptual annual production rate of 2.0Mtpa.

Table 1.3 Sanbrado Gold Project, December 2017 Resource

Indicated Resource Inferred Resource Cutoff

(Au g/t) Grade Grade Tonnes Au Oz Tonnes Au Oz (g/t) (g/t) O/P <120m 0.5 730,000 6.8 161,000 70,000 5.1 11,000 M1 South U/G >120m 3 470,000 26.4 395,000 350,000 16.1 180,000 Total Combined 1,200,000 14.4 556,000 410,000 14.4 191,000 M5 O/P 0.5 35,890,000 1.3 1,461,000 11,950,000 1.1 412,000 M1 North* O/P 0.5 780,000 1.9 49,000 660,000 1.9 41,000 M3* 0.5 170,000 2.0 11,000 260,000 1.4 12,000

Note: Mineral resources are not mineral reserves and have not demonstrated economic viability. All figures have been rounded to reflect the relative accuracy of the estimates. No changes have been made to the M1 North and M3 Mineral Resources since the February 2017 estimates.

Mineral Reserve Estimates Mineral Reserves and associated project economics have not been updated in this report. Information from the February 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018.

Mineral Reserves for the Sanbrado Gold Project have been estimated in accordance with NI 43-101 guidelines, and have excluded the use of Inferred Mineral Resources. Inferred Mineral Resources are too low confidence to be converted to Mineral Reserves.

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Mineral Reserves were estimated based on the resources contained within pit designs, including an allowance for mining dilution. Only Indicated Mineral Resources were used in the pit optimisation and design process. Consequently, the Mineral Reserve comprises only Probable Reserves.

Probable Mineral Reserves contained within the pit designs comprise 16.8Mt at 1.7g/t Au for 894koz contained gold. These Mineral Reserves were used in the preparation of production schedules and cashflow analyses.

Whittle 4D software was used to determine the optimum shell upon which the pit design was based. It is important to note that the capital costs associated with construction of the project are ignored in the optimisation process. Optimisations were run for the M1 North, M1 South, M3 and M5 deposits. Table 1.4 shows the optimisation results for the selected pit shells.

Pit slope angles varied between deposits, depending on the oxidation state of the pit walls. Mining costs were sourced by quotation from mining contractors experienced in West Africa.

MIK grade estimates provide a recoverable resource estimate that takes mining selectivity into account, and no further mining dilution or losses need be applied. OK was used to estimate the gold grade in the M3 deposit and a high grade zone in the M1 deposit. For the high grade zone in M1, the mining approach would attempt to minimise ore losses at the expense of some extra dilution, and a 20% dilution with no ore losses were applied. For the M3 deposit, the resource model was regularised to a SMU size of 5mE by 5mN by 5mRL. The regularisation of the block model results in the calculation of weighted average gold grades for the total block volume, which are essentially diluted grades. Ore losses will occur where a block contains a small proportion of mineralised material and the resultant weighted average block grade falls below the cutoff grade.

The process throughput rate varies for differing ore types, resulting in differing unit rates for the fixed costs for each ore type.

Metallurgical testwork has indicated that the recovery of gold is grade dependant. Process recoveries used for the pit optimisations are based on recovery algorithms as determined by testwork results as shown in Section 1.11. A Government Royalty of 4% and an additional 1% community development levy were applied to the optimisation process.

A $5.00/oz refining charge was allowed for in the optimisations. A gold price of $1,200/oz was assumed for pit optimisation. As recovery is based on a grade dependant formula, there is no in-situ cutoff grade applied. The cutoff grade is determined by the cost of processing divided by the realisable sale value of 1 gram of recovered gold. The average cutoff grades calculated from the recovery data are shown in Table 1.5.

Final pit designs were prepared for each deposit to enable practical and efficient access to each bench. The designs were based on the selected optimised shells and geotechnical design criteria. The initial and final pit outlines for the M5 deposit are shown in Figure 1.7. The final pit inventories are provided in Table 1.6.

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Table 1.4 Selected Optimum Pit Shells

Processe Costs Cashflow d Ore Total Strip Pit Shell Waste Revenue Deposit Material Ratio Disc. (US$M) Undiscou Disc. Disc. # (Mt) Ore Au Cont. Au Mining Process Royalty Avg. (Mt) (w:o) nted Best Worst (Mt) (g/t) (koz) (US$M) (US$M) (US$M) (US$M) (US$M) (US$M) (US$M)

M1 North 20 5.7 5.1 8.8 0.6 2.05 38.4 12.2 11.4 2.2 40.5 14.7 13.7 13.7 13.7 M1 South 17 32.6 31.8 39.8 0.8 8.48 217.5 78.2 16.6 12.5 231.4 120.0 114.3 111.9 113.1 M3 Pit 20 0.6 0.5 3.4 0.1 1.77 8.2 1.3 2.1 0.5 9.2 5.4 5.3 5.3 5.3 M5 Pit 21 65.1 50.3 3.4 14.8 1.33 633.0 146.8 253.1 36.6 674.8 238.3 193.6 176.6 185.1 Total 104.1 87.7 5.4 16.4 1.71 897.2 238.5 283.2 51.8 955.9 382.4 327.0 307.5 317.3

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Table 1.5 Average Calculated Cutoff Grade (g/t Au)

M1 M3 M5 Strongly Oxidised 0.37 0.37 0.37 Moderately Oxidised 0.43 0.43 0.43 Transitional 0.62 0.63 0.63 Sulphide 0.63 0.72 0.70

Table 1.6 Pit Inventories

Processed Ore Total Waste Strip Au Grade Cont. Au (Mt) (Mt) Ratio (Mt) (g/t) (koz) M5 64.1 48.7 3.2 15.3 1.30 639 M1Sth 30.3 29.5 35.6 0.8 7.96 212 M1Nth 5.6 5.1 9.9 0.5 2.07 34 M3 1.0 0.9 6.0 0.1 1.75 8 Total 101.0 84.2 5.1 16.8 1.65 894

Figure 1.7 M5 Pit Design showing Initial Pit Stages and Waste Rock Dump

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Mining Methods The material to be excavated will predominantly be free dig from surface, with blasting required deeper in the oxidation profile. Given these conditions, conventional open pit mining techniques using drill and blast with material movement by hydraulic excavator and trucks will be employed. The project scale suits 120t to 250t class excavators in a backhoe configuration matched to 95t class mine haul trucks.

Mining activities will be undertaken by an experienced contractor, with WAF retaining responsibility for technical services comprising mine planning, production scheduling, grade control, surveying, and supervision and management of contract mining operations.

Production schedules were completed with the primary aim of providing the highest value ore to the mill as early as possible, in order to maximise the value to the project. To achieve this, the higher grade and higher strip ratio material from the M1 South pit and the southern part of M5 are mined first. This results in a total material movement of 25Mtpa to 30Mtpa for the first two years of production, reducing to 13Mtpa to 14Mtpa for the remainder of the mine life.

In order to define the boundaries between high grade, low grade and waste material, it is necessary to drill RC holes on a close spacing and assay the mineralised intercepts. It is proposed that rows of RC holes be drilled with a 6m spacing across strike and an 8m spacing along strike for grade control purposes.

An allowance has been included in the operating and capital costs for a pit dewatering system to pump water from pit sumps, although further groundwater studies are being undertaken to provide an estimate of groundwater inflows.

Waste rock dumps have been designed with 15m batter heights with angles of 20°. The batters are separated by 15m wide berms. The overall slope angle of the waste dump faces is 15°. Over 50% of waste material mined will be moderately or strongly weathered, and should be benign from an acid mine drainage perspective. Of the balance, 15% is WOX material whilst the remaining 35% is FRS material. Further work is underway to determine the acid generating potential of rock types. Construction of the dump will be scheduled to encapsulate any acid generating rock within non-acid generating or acid neutralising rock.

The resultant mine production schedule is shown in Table 1.7. Rather than reducing the mining rate to match the milling rate from Year 4 onwards, the pits will be mined at a constant rate and excess ore will be stockpiled. This has been done to minimise the fixed mining costs over the life of the project. It has been estimated that compressing the mining schedule saves in excess of US$10M in mining costs.

Further work is in progress and is expected to be completed in mid-2018.

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Table 1.7 Mine Production Schedule

Pre- Total Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Prod

Mining Total Material (Mt) 101.0 5.3 30.8 23.6 14.3 14.3 12.7 Waste (Mt) 84.2 4.9 28.1 20.9 12.0 10.0 8.1 (Mt) 16.8 0.4 2.7 2.7 2.3 4.3 4.5 Total Ore (au g/t) 1.7 1.66 2.03 2.36 2.11 1.16 1.23 Processing Total Milled (Mt) 16.8 2.1 2.0 1.6 1.9 2.0 2.0 2.0 1.9 1.3 Head Grade (au g/t) 1.7 2.34 2.87 2.84 1.53 1.10 1.07 1.07 1.07 0.90 Recovered Au (koz) 810.2 145.4 166.7 137.9 81.3 64.3 62.4 62.4 55.6 34.2

Recovery Methods The Sanbrado Gold Project process plant will have a design throughput of 2.0Mtpa.

The proposed plant design comprises the following:

. Primary Crushing; . Crushed Ore Stockpile; . Grinding and Classification; . Gravity Recovery; . Leaching and Adsorption; . Elution; . Electrowinning; and . Smelting. Figure 1.8 provides a flowsheet diagram.

Haul trucks operating directly from the open pit will deliver Run of Mine (ROM) ore to the ROM pad. Ore will be stored on the ROM pad in separate stockpiles of varying ore types and grades to facilitate blending of the feed into the crushing plant. The comminution circuit selected is a single stage SAG mill, fed with primary crushed ore.

The crushing circuit will be designed with a throughput of 450tph and availability of 68.5%. The milling circuit is designed for a nominal throughput of 250tph. It will operate at 91.3%

availability, and achieve a design grind product size of 80% passing (P80) 90 micron.

The gravity circuit will take a cut from the cyclone underflow and will consist of two centrifugal concentrators and an intensive leach reactor for treatment of the gravity concentrate. The gravity circuit is expected to treat up to 50% of the cyclone underflow.

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Figure 1.8

Sanbrado Gold Project Process Plant Flow Diagram

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The CIL circuit will consist of seven adsorption tanks, treating the cyclone overflow. The CIL circuit is designed for a design leach feed grade of 2.8g/t Au.

Metal recovery and refining will consist of an elution circuit, electrowinning cells and smelting.

A tailings storage facility (TSF) will be used for tailings deposition to a paddock type storage dam located 1.5km from the Plant Site.

The Plant Control Systems (PCS) will be a network of Process Logic Controllers (PLC) sitting beneath a Supervisory Control and Data Acquisition (SCADA) network layer. The PLC’s will perform the necessary controls and interlocking, whilst the SCADA terminals will monitor the PLC’s and provide an interface for operator interaction.

Power will be accepted at the terminals of the plant high voltage feeder housed in the Plant Main Substation. This will house the Plant Main 11 kilovolt distribution board. Power distribution within the plant area and vicinity will be at 11 kilovolts and 415 volts.

Plant air and instrument air will be supplied from two air compressors operating in duty/standby mode.

Process water will be stored in and delivered to the plant from the process water tank. The sources of the process water inputs will be from the TSF decant return water and from the tailings thickener overflow. Potable water will be produced from local water bores.

Raw water will be sourced primarily from the Gibgo River, located 6.5km southwest of the plant site. The river water will be pumped to a water storage facility (WSF) located approximately 800m from the plant, adjacent to the TSF. Further work is in progress and is expected to be completed in mid-2018.

Project Infrastructure There is limited existing infrastructure and no existing services that are suitable to support the Sanbrado Gold Project. The majority of existing infrastructure supports local subsistence and small-scale agricultural practices.

Figure 1.9 presents the overall project arrangement.

The TSF will be located 1.5km northeast of the process plant, outside the 500m blast zone around the M5 pit. The proposed TSF is designed with a storage capacity of approximately 25Mt and consists of two cells of equal 12.5Mt capacity. The TSF will comprise a four-sided paddock storage facility formed by a multi-zoned earthfill embankment and a HDPE geomembrane basin liner. Tailings within the two TSF cells will be raised at similar rates. Raises will be completed using full downstream embankment construction (i.e., no embankments constructed on the tailings surface). The TSF design incorporates an underdrainage system to reduce the pressure head acting on the HDPE liner, reduce seepage, increase water recovery, increase tailings densities, and improve the geotechnical stability of the embankments. Tailings deposition into the TSF will be sub-aerial from a HDPE pipeline located around the perimeter of each cell.

The WSF will be located immediately to the southwest of the TSF in the area of a small creek line that intermittently flows to the southeast and towards the River Gibgo.

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Figure 1.9 Site Development

The WSF is designed to store up to 15Mm³ of water at the maximum operating level. The WSF will be primarily supplied with water from the Gibgo River over a four month period during the wet season. Current estimates indicate that a water harvesting dam (WHD) constructed on the river with a storage capacity of approximately 200,000m³ will provide sufficient attenuation of flows for the required 1.5Mm³/yr of water to be harvested from the system, pumping continuously at 140L/s.

Administrative, maintenance and support functions will be housed in blockwork constructed buildings and steel-frame construction sheds.

Potable water will be produced by two dewatering bores located west of the M5 pit. Water will be delivered from the bores to a tank located at the accommodation camp, from which water will be fed to a 170m³ per day water treatment plant. Potable water for use in the process plant will be pumped from the camp to a 160m³ tank for use in the process plant and for distribution to the mining contractor’s area.

One sewage treatment system, located at the camp site, will be installed to service the plant buildings and the accommodation camp. All sewage water will be treated before the treated effluent is pumped to the TSF.

A heavy fuel oil (HFO) power station will be constructed at the process plant by an independent power provider (IPP) under a build-own-operate (BOO) agreement. The power station will be fitted with three duty HFO generator engines and one standby generator to maintain supply in the event of a shutdown of one of the engines.

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There is no significant or reliable telecommunication infrastructure in the immediate mine site area. Telecommunications will be established by satellite link which will include voice, email and internet traffic for the process plant, camp and main office. A conventional UHF radio system with hand held radios and chargers will be provided for site coverage.

Project security will be provided by a specialist local security provider and will comprise:

. Mining lease access control; . Read in/read out access control; . Fencing (double and single layer) and electronic security gates; . Electronic surveillance including CCTV; . Physical and visual barriers; . Lighting; and . Patrols. The accommodation camp will be constructed prior to the commencement of process plant construction in order to accommodate construction personnel. The accommodation camp and supporting facilities will be designed for an accommodation capacity of 172 people.

A project implementation strategy has been developed for the effective design, engineering, construction and commissioning of the process plant and associated infrastructure and services, together with the detailed schedules for each phase of the project development up to plant operation at rated capacity.

The design, construction and operation of the Project will conform to the requirements of the various regulations in Burkina Faso, or requirements within Australian Standards, ISO Standards, European Standards and WAF internal standards.

Project implementation will utilise an Engineering, Procurement and Construction Management (EPCM) approach where possible.

Construction is expected to take 19 months to complete from award of the process plant EPCM contract. Mobilisation of operations personnel will occur progressively over a 15 month period prior to production. Operations readiness will commence early in the project to initially accommodate the needs of open pit mining (commencing five months prior to production). This will incorporate infrastructure and the processing plant as the construction program progresses.

The total operating manpower for the project will be 636, including all staff and contractors. It is expected about 168 of that number will be WAF employees and the remaining 468 will be Contractors including the mining contractor staff, security personnel, medical staff and accommodation/catering contract staff.

Where possible, the workforce will be drawn from the local area proximal to the project, however, the population of Nedogo village and surrounding region is not large enough to support an adequate workforce for the operation. It is anticipated that up to 50% of personnel will be drawn from the local area and more than 75% of all personnel will be Burkina Faso Nationals. Some personnel will be sourced internationally and operate on a fly-in, fly-out (FIFO) basis.

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The intended approach to nationalisation of the workforce will be to attempt reduction of start-up expatriate (international) numbers by 10% each year. A substantial recruiting and training effort will be required to ensure that qualified operators are available for mining, processing and support operations. Supervisory roles within the operations departments will be filled by skilled expatriate workers who, along with vendor trainers, will supervise the development of operating personnel and the structuring of training to ensure the quality of operating personnel.

Environmental Studies, Permitting and Social or Community Impact An Environmental and Social Impact Assessment (ESIA) was prepared for WAF during 2015 and 2016 for the Tanlouka Heap Leach Gold Project. This ESIA and associated baseline surveys predicted that there will be generally minor impacts on local flora and fauna, air quality and visual amenity. This ESIA was approved by the Burkina Faso government in November 2017.

A planned review of the Sanbrado Gold Project ESIA will also determine whether or not the updated project will reduce some overall impacts, including a reduction in the number of households subject to relocation due to changes in buffer zone requirements. The re- submitted ESIA will also identify community benefits associated with infrastructure remaining at cessation of mining such as the WSF, Gibgo River Dam and water pipeline, the camp and electrical power line.

The local community has a relatively low standard of living with significant challenges educating and retaining younger members who could provide the energy and knowledge to assist in the improvement of socio-economic conditions. Local infrastructure is not well developed and the proposed mine will aid the upgrading of transport, power and water infrastructure in the area. It will also offer skills development and indirect business opportunities which can be sustained beyond the life of the mine. The mine will also provide a stimulus for educated and trained young people to remain in the community.

Control measures will be identified for all potentially negative environmental and social impacts to mitigate negative effects and enhance positive effects of the Project. Monitoring and management plans will be developed to address key environmental issues such as air quality, surface and groundwater, noise and vibration and erosion and sediment control. A similar management plan is under development for the identified socio-economic impacts of the project.

Hazardous and non-hazardous solid and liquid wastes generated at the mine site will be managed in accordance with local regulations. A modular sewage treatment plant will be installed onsite and operated and maintained during the life of the mine.

Surface water diversion channels will be required around all open pits. A sediment control structure (SCS) is located at the site boundary of each discharge point for sediment control and compliance monitoring.

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At the end of the TSF and WSF operation, the external embankments will have an average downstream slope of 1:3.5. The profiles will be stable under both normal and seismic loading conditions, provide a stable drainage system, and will allow for revegetation.

The rehabilitation of the WSF is to be agreed. It is possible that the WSF will be retained as a community facility and the embankment slopes of the WSF have been designed to provide long term stability.

The monitoring of the mine area is vital to ensure that the environment impact is managed effectively and that negative impacts are minimised. Regular sampling and assessment of groundwater, surface water, air quality and soil resources will be undertaken throughout the life of the mine and conducted post-closure at defined intervals sufficient to satisfy legal requirements and the relevant conditions of Project approval. Specific monitoring and management plans will be developed and implemented to address each of these areas.

The Burkina Faso Environmental Code stipulates that submission of an ESIA is mandatory for any mine whose production capacity is greater than 100tpd. The Burkina Faso ESIA development, review and approval procedure includes the following steps:

. Preparation of a Project Description and draft ESIA Terms of Reference (proponent); . Review, approval and finalisation of the ESIA Terms of Reference (Burkina Faso Ministry of Environment, Green Economy and Climate Change (MEEVC)); . ESIA development (proponent); . Public consultation and information sessions (proponent); . Submission of the ESIA to MEEVC for review (MEEVC); . Review and preliminary analysis of the ESIA (National Office of Environmental Assessments (BUNEE)); . Consideration of the ESIA (Technical Committee on Environmental Assessments (COTEVE)); . Public inquiry (MEEVC); . Report of the Public Inquiry (MEEVC); . Decision on the approval of the project (MEEVC); and . Granting of a Mining Licence, if the project is approved (Ministry of Mines). The 1995 Mining Policy Statement stresses the importance of the private sector as a driver of economic development and the mining sector as an important factor in ensuring the country’s growth.

In line with Burkina Faso ESIA requirements, potentially affected populations and their representatives have been consulted during the ESIA process for the Tanlouka Gold Project and will continue to be consulted as part of the Sanbrado Gold Project.

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Potential socio-economic impacts identified during the consultation process include:

. Reduced access and quarantine of approximately 749ha of agricultural and pasture lands of which 253ha are in the mining permit area; . The loss of homes on approximately 20 concessions including 122 households requiring the relocation of approximately 674 people; and . The social impacts of resettlement including community disturbance, land use conflicts and integration issues. The development of a Preliminary Closure Plan allows integration with operational design concepts and ensures that the design and operation of the site are compatible with the closure plan. A Detailed Closure Plan will be developed after the Project has been commissioned. The Detailed Closure Plan will be reviewed periodically.

As the Closure Plan is developed during the mine life, a set of completion criteria for rehabilitation will be derived from site trials and community consultation. The criteria will be consistent with overall site closure objectives and will be determined and agreed with the regulator and relevant key stakeholders. Through long-term monitoring of the site, it will be demonstrated that the development of rehabilitated areas is consistent with agreed completion criteria.

Capital and Operating Costs Mineral Reserves and associated project economics have not been updated in this report. Information from the February 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018.Operating cost estimates for the Sanbrado Gold Project have been prepared to allow for modelling of the proposed processing schedule, taking into account the variability of the ore feed from the different pit domains and the proposed blended mill feed.

The operating cost estimate covers the administration, owners mining, processing and maintenance costs. The operating costs have been estimated from a variety of sources, including:

. First principle estimates; . Consumable consumption rates as provided by metallurgical testwork results; . Grinding power and media consumptions as provided by OMC; . Quotations for the supply of consumables and services; . Advice provided by WAF; and . Published and unpublished costs for operations of a similar scale and location. All costs are in United States Dollars (USD) and reflect an estimate accuracy of ±15% at Quarter 1 2017 (Q12017) with a 90% level of confidence.

Table 1.8 summarises the sustaining capital cost estimate.

The average All-in Sustaining Cost for the project during the production period is estimated to be $759/oz, as summarised in Table 1.9.

The capital cost estimate is based upon an EPCM approach whereby the owner assumes the builder’s risk. As a result, the capital cost estimate does not include a builder’s margin.

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Table 1.8 Sustaining Capital Estimate

Sustaining Capital Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Mining $1.2M - $0.6M ------Tailings Storage Facility $6.6M $2.8M $2.4M $2.7M $2.9M $2.9M $2.9M - $4.4M Tailings Discharge $1.2M ------Access Road - $0.1M ------Surface Water Management $0.2M ------Water Storage Dam ------$0.0M Mobile Equipment & Vehicles - - $0.1M - $0.3M - - - - Process Plant $0.3M $0.3M $0.5M $0.5M $0.5M $0.5M $0.3M - - Camp & Administration $0.0M $0.0M $0.1M $0.1M $0.1M $0.1M $0.1M - - Total $9.4M $3.1M $3.6M $3.2M $3.7M $3.4M $3.2M - $4.4M

Table 1.9 LOM Operating Costs

Operating Costs US$/t Ore (processed) US$/Oz (produced) Mining 15.82 329 Processing 10.77 224 G & A 4.92 102 Cash Operating Cost 31.51 655 Royalties 2.89 60 Refining 0.10 2 Total Cash Cost 34.50 717 Sustaining Capital 2.03 42 All-in Sustaining Cash Cost 36.53 759

The capital cost estimate has been prepared as a bankable level feasibility study and is presented in United States Dollars (USD) to an accuracy level of ±15%, with a 90% level of confidence, as at Quarter 1 2017 (Q1 2017).

Table 1.10 summarises the capital cost estimate for the project, including contingency.

Table 1.10 Capital Cost Summary

Sub-Total Contingency Total Cost Area US$ US$ US$ Process Plant $53,247,461 $4,968,966 $58,216,427 Infrastructure $33,130,909 $3,569,207 $36,700,116 Owner’s Costs $25,773,608 $3,476,159 $29,249,767 Total Project Costs $112,151,978 $12,014,332 $124,166,310

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Economic Analysis Mineral Reserves and associated project economics have not been updated in this report. Information from the February 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018.

The Sanbrado Gold Project has been evaluated on a discounted cashflow basis. The results of the analysis show the Sanbrado Gold Project to be economically very robust. The present value of the net cashflow with a 5% discount rate (PVNCF5%) is $143 million on a pre-tax, project basis, using a base gold price of $1,200/oz. Project post-tax PVNCF5% at a $1,200 gold price is $100 million on an all equity basis. Project internal rates of return (IRR) are 27% pre-tax and 21% post-tax.

The Burkina Faso government is entitled to a 10% free-carried interest in the mine operating entity. The economic analysis assumes that WAF will provide all development funding via loans to the mine operating entity, which will be paid back with interest from future gold sales. On this basis, WAF’s 90% interest in the project is expected to provide a post-tax PVNCF5% of $91 million and a post-tax IRR of 20% at a gold price of $1,200/oz.

Project payback period at a gold price of $1,200/oz is expected to be 2.1 years on a pre-tax basis and 2.3 years on a post-tax, all equity basis. Payback period is defined as the time after process plant start-up that is required to recover the initial expenditures incurred developing the project. This is the point in time when the cumulative undiscounted net cashflow is zero.

Like most gold mining projects, the key economic indicators of PVNCF5% and IRR are most sensitive to changes in revenue parameters such as the gold price. A $100/oz reduction in the gold price would reduce the post-tax PVNCF5% by $44 million, and reduce the post-tax IRR by 7%. A $100/oz increase in the gold price would increase the post-tax PVNCF5% by $43 million, and increase the post-tax IRR by 6%.

The cashflow analysis has been prepared on a constant first quarter calendar 2017 US dollar basis. No inflation or escalation of revenue or costs has been incorporated.

Table 1.11. summarises base case economic results.

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Table 1.11 Sanbrado Gold Project Base Case Economic Results Summary

Units LOM Total

Gold Revenue Gold Price $/oz 1,200 Gold Sales ozs 810,156 Gold Sales Revenue $M 972 Pre-Production Costs Development Capital $M -124 Pre-production Mining $M -7 Pre-production General and Administrative $M -2 Production Period Costs Mining $M -266 Deferred Payment of Pre-production Mining Costs $M -5 Processing $M -181 General and Administrative $M -83 Refining $M -2 Royalties $M -49 Sustaining Capital $M -34 Project Net Cashflow, pre-tax $M 219 PVNCF5% $M 143 IRR % pa 27 Payback Period Years 2.1 Project Net Cashflow, Post-Tax, All Equity Basis Project Net Cashflow, pre-tax, from above $M 219 Income Tax, all equity basis $M -56 Project Net Cashflow, post-tax, all equity basis $M 163 PVNCF5% $M 100 IRR % pa 21 Payback Period Years 2.3 WAF Net Cashflow, Post-Tax Project Net Cashflow, pre-tax, from above $M 219 Income tax, with interest and management fee deduction $M -42 Withholding taxes on interest and dividends $M -11 Dividends to Government (Government 10% carried interest) $M -10 WAF Net Cashflow, post-tax $M 156 PVNCF5% $M 91 IRR % pa 20

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INTRODUCTION Terms of Reference The Sanbrado Gold Project (formerly Tanlouka Gold Project) is owned and operated by West African Resources Limited (WAF). WAF is a public company listed on the Australian Securities Exchange (ASX) and Toronto Stock Exchange (TSXV), and registered in Australia (ASX and TSX code: WAF).

WAF has prepared a Technical Report for the Sanbrado Gold Project to the standard of the Canadian National Instrument 43-101 “Standards of Disclosure for Mineral Projects” following the completion of a Resource Estimation Studies covering the Mankarga 1 South and Mankarga 5 (M5) deposits in October 2017. Mineral Reserves and associated project economics have not been updated in this report. Information from the February 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018.

Purpose of Report This report details the results of a Feasibility Study for the Sanbrado Gold Project as outlined in the 30 October 2017 news release by West African Resources Ltd. This technical report conforms to the standards prescribed by NI 43-101, Standards and Disclosure for Mineral Projects, and pursuant to Form 43-101F.

Cautionary Notes This report has been compiled based on information available up to and including the date of this report. The status of agreements, royalties or tenement standing pertaining to the assets, have not been investigated by contributing consultants and were not required to be. All matters relating to ownership are to be directed to WAF for clarification if required.

Sources of Information This report relies on historic and recent data generated by WAF. WAF has engaged a number of specialist consultants and information from these reports prepared by previous independent consultants has been utilised in the compilation of this report. Section 27 provides a list of references relied upon in preparation of this report.

Site Visits The Independent Qualified Person (Mining Engineer) Stuart Cruickshanks, Principal Consultant at SCME, visited the Sanbrado Gold Project January 2017, and carried out meetings and discussions with key exploration personnel for project overview, field verification of drillhole locations and review of drill core. Field inspections of proposed mining and plant areas and potential water storage sites as well as other locations for potential site infrastructure were completed.

The Independent Qualified Person (Resource Geologist) Brian Wolfe, Principal Consultant at International Resource Solutions Pty Ltd, visited the Sanbrado Gold Project Site in May 2014, May 2016 and April 2017. These visits included:

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. Visits to the exploration sites, outcrop exposures, and observation of surface drilling, review of drill core from several diamond holes drilled at M5, M3 and M1 that form part of the Project resource estimate; . Review of the exploration procedures used by WAF at the Sanbrado and Boulsa Project; . Review of WAF exploration database; and . Review of geological setting of the deposit and surrounding area. The Independent Qualified Person (Civil Engineering) Dave Morgan, Managing Director of Knight Piésold Pty Ltd, visited the Sanbrado Gold Project in March 2015 and inspected proposed sites for water storage and sediment control dams, reviewed civil and road construction locations and reviewed environmental and social data.

The Independent Qualified Person (Minerals Processing) Tom Kendall, Principal Consultant at Mintrex Pty Ltd, visited the Sanbrado Gold Project in March 2015 and met with key exploration personnel for project overview, field verification of drillhole locations and review of drill core. Field inspections of proposed mining and plant areas and potential water storage sites were completed, as well as visits to other locations for potential site infrastructure.

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RELIANCE ON OTHER EXPERTS While information provided by WAF relating to mineral rights, surface rights and permitting has been reviewed, no opinion is offered in these areas. The Qualified Person is not expert in land, legal, permitting, and related matters and therefore has relied upon, and is satisfied, there is a reasonable basis for this reliance on the information provided by the WAF management regarding mineral rights, surface rights and permitting in Section 4 of this Technical Report. The qualified persons have relied upon September 2014 report entitled: “NI 43-101 Preliminary Economic Assessment, Mankarga 5 Gold Deposit, Tanlouka Gold Project, Burkina Faso” and April 2015 report entitled: “NI 43-101 Pre-Feasibility Study, Mankarga 5 Gold Deposit, Tanlouka Gold Project, Burkina Faso.” The authors of this Technical Report, state that they are Qualified Persons for the areas as identified in the Certificate of Qualified Person attached to this report.

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PROPERTY DESCRIPTION AND LOCATION Property Location The Sanbrado property is located in the West African country of Burkina Faso, approximately 90km east southeast of the capital Ouagadougou, and approximately 20km south of the regional centre of Mogtedo. Centroid co-ordinates for the Project are longitude 0° 51‟ W and latitude 12° 10‟ N. The project location is shown in Figure 4.1.

Figure 4.1 Sanbrado Gold Project Location

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Mineral Tenure The Burkina Faso Ministry of Mines and Energy grants an Exploration Permit for a three year term, renewable for two further three year terms providing all permit conditions are met. The Exploration Permit gives the title holder the exclusive right to explore for the nominated commodity and to apply for an Exploitation Permit, should a deposit of economic significance be identified. The Burkina Government maintains a 10% free carry interest in projects and a free on board royalty of 5% for precious metals.

The Sanbrado Gold Project comprises one granted Exploration Licence (Manesse Exploration Permit - N2017/014/MEMC/SG/DGCMIM) granted on 13/01/17 for 3 years and one granted Mining Licence (Sanbrado Mining Permit -No 2017 –104/PRES/PM/MEMC/MINEFID/MEEVCC) granted 13/03/2017 for 8 years, covering an aggregate area of 116km2. The Company is in the process of updating the Mining Licence from heap leach processing to CIL processing, which is expected to be completed in early 2018.

Ownership WAF is a Perth, Australia, based gold exploration company dedicated to creating shareholder value through the acquisition, exploration and development of resource projects in Burkina Faso, West Africa. Its exploration assets comprise several Burkina Faso properties, including the Sanbrado Gold Project. WAF acquired a 90% interest in the Sanbrado property through the acquisition of Channel Resources Ltd on 21 January 2014, and acquired the remaining 10% of the project from GMC SARL, a Burkina Faso registered entity, in September 2015.

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Figure 4.2 provides WAF’s current corporate structure.

Figure 4.2 Corporate Structure - West African Resources Ltd

Environmental An Environmental and Social Impact Assessment (ESIA) was completed over the Sanbrado Gold Project area. The results of this study are discussed in Section 20 of this report.

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ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY Accessibility The Sanbrado Gold Project is located in the Commune du Boudry, close to the major towns of Boulsa and Zorgho. The area is well serviced by mobile phone coverage and low voltage grid power is being rolled out. The Project is accessed via the bitumen RN4 highway, which links Burkina Faso and Niger and crosscuts the southwestern corner of the project area.

The Sanbrado Gold Project site is accessible by lateritic secondary roads from Mogtedo and Zempsago, a distance of about 25km. The secondary roads are relatively flat and accessible for most of the year. (Figure 5.1).

Figure 5.1 Sanbrado Gold Project Location Map Showing Major Secondary Access Roads

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Access within the project is excellent via all-weather formed gravel and sand roads and village tracks. Access during the rainy season, from July to September, can be restricted depending on rainfall.

Climate Sanbrado is located south of the sub-Saharan Sahel of West Africa, and vegetation is predominantly open-forested savannah and grassland. The climate of the region is sub- tropical and semi-arid, with warm, dry winters and hot, wet summers. The wet season extends from mid-April until mid-October. Annual rainfall is generally between 700mm and 1,000mm. Peak rainfall months are August, September and October. Daytime temperatures range from 35˚C to 45˚C in the dry season and from 30˚C to 35˚C in the wet season.

Exploration activities can be curtailed during periods of heavy rain, however it is expected that mining operations would be conducted year-round.

Vegetation Vegetation in uncultivated areas comprises savannah woodlands with dense bush common along the seasonal watercourses.

Crops under cultivation include millet, rice, cotton, peanuts and maize. Wildlife is restricted to livestock, birds and reptiles.

Local Resources and Infrastructure The project is located in a relatively sparsely inhabited area of Burkina Faso. As such, infrastructure and local resources are modest and limited to communities close to the major roads. Project execution would require building a greenfields project with attendant infrastructure. Power for mining operations will have to be generated on site.

Rudimentary supplies to support exploration activities are available in the village of Mogtedo (population ~15,000), approximately 20km northwest of the Sanbrado property. Mogtedo is the location of WAF’s exploration compound and the location of archived sample and pulp reject storage facilities, as well as a yard where drilling materials are stored. In 2011, Channel built an exploration camp about 5km from the M5 deposit that includes sleeping quarters, bathrooms, showers and a kitchen. All major items and equipment must be imported or obtained from Ouagadougou.

Three water wells were drilled on the property. Two wells, located along M5, were drilled to supply water for core drilling and are accompanied by holding ponds. The third well was drilled to provide potable water for the exploration camp. The Company proposes to build a water pipeline from the Nakambé River some 15km south of the project, which will provide adequate water supplies for operations.

Physiography The property is flat to gently rolling with sparse outcrop (Figure 5.2). Locally, hills and ridges a few tens of meters high are capped by a hard, ferruginous laterite crust (cuirasse). The remainder of the terrain is composed of clay-rich saprolitic profiles that are up to 50m deep. Drainage patterns are rectangular to dendritic, reflecting late fracture systems trending mainly north, east and northwest.

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Watercourses are dry for much of the year, but can quickly become wide and flooded during the rainy season.

Figure 5.2 Sanbrado Gold Project: Photograph showing Typical Physiography of the Sanbrado Permit

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HISTORY GMC undertook the first known organised exploration over the Bombore permit from 1989 to 1993, which comprised prospecting, hand panning of crushed rock and soils, some systematic rock sampling, hand excavation of trenches with limited mapping and site descriptions (Guérard, 1997).

Starting in 1994, the exploration work done over the Tanlouka area was by Channel, as part of its Bombore permit. In the period from 1994 to 1996, Channel undertook several reconnaissance style mapping and prospecting traverses. This work covered the entire Bombore permit, which included the Tanlouka permit in the east and High River Gold’s Mango permit in the west. All areas of artisanal workings and significant outcrop were described and classified for further work.

During this period, the main emphasis of work was the central area of the permit which became known as the Bombore First Target, located northwest of the Tanlouka permit. At the time of this work, there were no known artisanal workings on the Tanlouka permit, but workings were subsequently found in the southwest corner of the Tanlouka property. A historical mineral resource estimate was carried out on the M5 zone in 2012 by AMEC.

Geological Mapping and Rock Chip Sampling Following positive soil geochemical results in the Mankarga area in September 1996, Channel’s geologists recommended that this area be geologically mapped at 1:50,000 scale. The first mapping and prospecting undertaken over the area of Sanbrado, with the exception of the early work by Anderson (1995), was completed in 1999. Most of the outcrops observed in the area of the gold-in-soils geochemical anomaly are composed of weakly to moderately deformed metagabbro with an associated major structural trend varying from 010˚ to 040˚. These gabbros are intruded by metre scale, fine-grained felsic dykes, which are in turn fractured and injected by quartz stringers.

A total of 14 rock samples were taken during this mapping program, mostly of quartz stringers and a few intrusive rocks, and yielded uniformly low assay results for gold with a highest value of 0.063g/t Au in a grab sample. Between 2006 and 2012, a number of consultants (Anderson, Ouattara and SRK) were appointed to carry out mapping and desktop reviews to ascertain a geological interpretation of the area and establish the mineralisation potential of the Tanlouka permit, and more specifically the M5 deposit.

Geophysical Surveys In March 1996 a fixed wing airborne geophysical survey was flown by Aerodat Inc. for magnetics, EM-VLF and Radiometrics, and SPOT images provided by MIR Teledetection Inc. Survey specifications comprised 8,548 line km based on a 300m line spacing with tie lines every 5km.

Due to proximity to the equator, the EM-VLF data was not very useful and only the magnetics and radiometrics data were used. Both datasets show the presence of northeast trending lithostructural domains. The far southeast corner of the Tanlouka permit is underlain by a separate domain, characterised by a low magnetic and a very high radiometric signature, in an area thought to be dominantly post-tectonic K-granites and granitic gneisses.

In March and April, 2008, SAGAX carried out a ground geophysical survey for induced polarisation (IP) and magnetics for the area around M5. The survey comprised ten 300m

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long profiles with a station reading conducted every 25m in an east-west direction. Each profile was 200m apart in a north-south direction. Three main domains corresponding to the main geological units were interpreted as a highly conductive zone in the southwestern part of the grid correlated with volcano-sedimentary schists, a highly resistive zone in the south-eastern portion of the grid thought to be related to granitic rocks, and intermediate resistivities in the central part of the grid thought to be related to basaltic and amphibolitic formations.

In February 2011, UTS Aeroquest was contracted to carry out a low level magnetic airborne survey. The survey specifications comprised 1,994 line km on 50m spaced lines (north-south) with 500m spaced tie lines (east-west).

A map showing the total gradient of the total magnetic intensity over the original Tanlouka permit is shown in Figure 6.1.

Soil Geochemistry Regional soil sampling was initiated in 1996 which covered the Tanlouka permit at 500m centres in a triangular grid. Assay results for Au, Cu, Pb and Zn highlighted two gold anomalies on the portion of Bomboré now covered by the Tanlouka permit.

A total of 765 samples at 100m spacing along lines 500m apart in the southeast corner of the original Bomboré permit were collected by December 1996 (Guérard, 1997). Anomalous gold values were associated with the western contact of the Mankarga sub-domain (Guérard, 1997). Anomalous gold-in-soil zones were defined over areas of 2km by 5km at Mankarga and Koloba Samba (Guérard, 1997).

In 2000, three separate grids, called Mankarga East, West and South were sampled in more detail (Learn, 2000). A continuous gold soil anomaly over 3km long and 200m to 300m wide was defined over the southern part of the Mankarga East grid. In addition, an oblique northeast trending zone of anomalous gold values extended into the granite terrain to the east, suggesting the presence of an oblique structure, possibly a splay or fold in association with the principal deformation zone.

In early 2010, Channel conducted a detailed soil sampling program on and around the M5 area. A total of 1,321 soil samples and several termite mound samples were collected. Two large soil grids were established in the central (Manesse grid, 4,800 soil samples) and northern (Tanwaka grid 3,754 soil samples) portions of the Tanlouka permit. Targets near the centre of each of these grids had been identified in a reconnaissance soil sampling program completed in 1998.

WAF has subsequently auger drilled the entire Tanlouka permit to verify some of the historic anomalies and results are discussed in Section 9. A summary of the historic soil geochemistry is shown in Figure 6.2.

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Figure 6.1 Tanlouka Permit : Total Gradient of Total Magnetic Intensity (TMI)

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Figure 6.2 Tanlouka Permit : Historic Soil Geochemistry overlying TMI Magnetic image

Petrographic Studies Three petrographic studies were completed between November 2010 and August 2011 by Geoconsult. The first was based on nine rock samples collected by Channel in and around the artisanal workings. The suite of rocks included four samples of variably altered intrusive rocks, four samples of variably altered and deformed sediments and one rhyodacite. Gold was identified in one of the altered intrusive rocks that showed evidence of silicification.

The second study was completed in March 2011 on thin sections made from Reverse Circulation (RC) chip samples. Of the 12 RC samples submitted, seven were from M5 and five were from M1. Geoconsult completed the requested analysis, but remarked that due to the small size of the chips provided, the analysis may not be entirely representative of the parent rock. This suite of rocks was divided into four different lithological units comprising intermediate intrusives, feldspar porphyry, schists and extensively altered porphyritic volcanic rock. Silicification was the primary alteration observed in the schists and shearing was the prominent deformation observed. Sulphide mineralisation was primarily pyrrhotite with minor pyrite and arsenopyrite.

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The third study was completed in August 2011 on twelve drill core samples from a single drillhole, DDH Tan11 DD 02. The mineralogy and texture of the rocks suggest that they are fine-grained, variably silicified and metamorphosed argillaceous and arenaceous sediments, except for one sample of quartz diorite. Metamorphic grade is middle greenschist facies.

A petrographic study commissioned by WAF was completed in October 2016. The study was based on 22 core samples collected by WAF from M1 South that typified the different lithologies and areas of geological uncertainty.

A second set of 21 rock chip samples of mineralised rocks from M5 has recently been collected with a view to determining detailed mineralisation processes.

Historic Drilling Since 2000 various drilling programmes have been completed on the Sanbrado Gold Project permit. Rotary Air Blast (RAB) drilling was undertaken over the soil anomalies at Sanbrado in April 2000. In June 2003 an RC drilling campaign of 5 holes was completed by Orezone. In 2010 Channel completed wide-spaced drilling at Mankarga 1, Mankarga 2, Mankarga 3, Mankarga 4 and Mankarga 5. Follow-up drilling commenced in November 2010 through to April 2011, comprising 81 RC holes at Mankarga 1, Mankarga 1 Nth, Mankarga 2 and Mankarga 5. Most of the holes were drilled at Mankarga 5. Diamond Core (DC) drilling commenced in July 2011 on the Mankarga 5 deposit and was completed in February 2012.

Historic Channel Resources drillhole locations are shown in Figure 6.3.

Figure 6.3 Tanlouka Permit: Historic Channel RC and DC Drilling

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GEOLOGICAL SETTING AND MINERALISATION Regional Setting The gold deposits of West Africa largely lie within the Proterozoic domain of the Man Shield, the southernmost subdivision of the West African Craton (Figure 7.1). Successions comprising the Man Shield overlie the Archaean Liberian Craton. The principal gold producing areas are associated with the Lower Proterozoic systems of the Birimian (2.17-2.18 billion years) comprising metavolcanic (arc) and metasedimentary (basin) rocks, and the marginally younger, unconformably overlying rocks of the Tarkwaian epiclastic system.

Figure 7.1 Generalised Geological and Structural Map of the West African Craton

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The Birimian Greenstone Belts of the West Africa Craton are argued to be generated by either collisional tectonics or mantle plume activity (Goldfarb et al., 2001). Archean and Paleoproterozoic gold-forming events correlate with subduction-style plate tectonics and the episodic growth of juvenile continental crust.

The Birimian System can be broadly subdivided into the Lower Birimian phyllites, tuffs and greywackes; and the Upper Birimian basaltic to andesitic lavas and volcanoclastics. These subdivisions are largely believed to be coeval and have been deformed and regionally metamorphosed from lower greenschist to lower amphibolite facies grade (Bossière et al., 1996).

The Birimian System has been intruded by two distinctive granitoid types. The larger basin- type granitoids are associated with the initial Eburnean event, a major phase of crustal thickening as a result of shortening, folding and granitoid emplacement, followed by regional north to northeast trending transcurrent faulting (e.g., Markoye Fault). Large scale fluid migration along these deep-seated structures is inherent to most orogenies. Hydrothermal gold-bearing fluids follow secondary and tertiary fault systems, adjacent to the main fault at shallower crustal levels.

The smaller belt-type (arc related) granitoids lack the characteristic foliation of the basin- type granitoids, and are generally interpreted to be syn or post-tectonic.

The younger Proterozoic Tarkwaian sediments consist of a thick series of arenaceous and minor argillaceous sediments, believed to be derived from erosion of the Birimian rocks.

The geology of Burkina Faso can be subdivided into three major litho-tectonic domains comprising a Paleo-Proterozoic basement underlying most of the country, a Neo- Proterozoic sedimentary cover developed along the western, northern and south-eastern portions of the country, and a Cenozoic mobile belt forming small inliers in the north-western and extreme eastern regions of the country.

The Paleo-Proterozoic basement comprises Birimian volcano-sedimentary and plutonic rock intruded by large batholiths of Eburnean granitoid. The overall structure of this basement is defined by two major north-northeast trending sinistral shear zones subdividing the country into three domains comprising an eastern domain cut by a series of northeast trending structures, a central domain characterised by arcuate structural patterns, and a western domain hosting north to northeast trending structural features.

The spatial distribution of gold mineralisation in West Africa (Figure 7.2) is closely governed by north to northeast trending belts of Upper Birimian metavolcanic rocks, ranging from 15km to 40km in width. Almost without exception, the major gold deposits lie at, or close to the margins of the belts in proximity to the strongly deformed contacts between the Upper and Lower Birimian sequences.

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Figure 7.2 Gold Deposits of West Africa on Regional Geology

Banket deposits represent the third significant style of mineralisation, hosted by quartz pebble conglomerates towards the base of the Tarkwaian Series. The gold is thought to be of detrital origin, derived from erosion of the Birimian Series upon which the Banket Group lie.

Sanbrado Gold Project Geology The Sanbrado Gold Project is located within a strongly arcuate volcano-sedimentary northeast trending belt in the central domain that is bounded to the east by the Tiébélé- Dori-Markoye Fault (Markoye Fault), one of the two major structures subdividing Burkina Faso into three litho-tectonic domains (Figure 7.3).

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Figure 7.3 Regional Geological Map-Eastern Burkina Faso

Airborne geophysics data show the presence of northeast trending litho-structural domains.

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Alteration assemblages are determined by metamorphic conditions, host lithologies of variable bulk compositions, as well as spatial association relative to regional structures and their second and third order splays.

Alteration assemblages observed at Sanbrado are typical of greenschist facies domains and dominated by chlorite, biotite, epidote, silica, sericite, calcite, albite, pyrite and +/-tourmaline.

Structure The Sanbrado Gold Project area displays a strong east-northeast structural trend, governed by the regional tectonic strain in east Burkina Faso (Anderson, 2006 and Barnett, 2010). A number of interpreted shear zones and faults occur within the Sanbrado Gold Project area and the fault network can be subdivided both in terms of age and fault order (scale) (Morel and Hughes, 2014). A deformation history for the Mankarga deposits has largely been established from overprinting and geometric relationships in outcrop and drill core, and comprises three events (D1 to D3) followed by brittle faulting (Morel and Hughes, 2014; Hughes, 2014).

D2 has been established as the most intense and complex event, comprising substantial fluid flow, ductile deformation, intense alteration, emplacement of igneous intrusions and the introduction of gold (Davis, 2014; Hughes, 2014).

The Mankarga deposits are hosted within the D2 Mankarga Shear, which has accommodated both sinistral and dextral shearing (S2) during regional transpression (Davis, 2014). On an orebody scale, D2 structures at M5 strike 040° and dip 70°to 80° northwest, which is concordant with most regional structural orientations in the Sanbrado Gold Project area. At M1, D2 structures typically strike 315° and dip 65°to 80° northeast. This is discordant with the regional structural orientations, and is interpreted as being the result of folding and/or conjugate faulting around the granodiorite.

At M5, a number of oblique dextral shear zones were developed across the Mankarga Shear coeval with sinistral D2 shearing in response to volume changes and the necessity to compensate for the intense strain during transpression (Davis, 2014). These dextral shears have been reactivated as brittle faults during D3. D3 manifests as chevron-style folds, crenulations and a commonly developed crenulation cleavage.

At M1, moderately to steeply plunging F2 folds are evident and are parallel to S2 shear orientation. D3 is dominated by dextral faulting with a strike direction ranging from northeast to east.

Structural geological controls show that gold mineralisation in the Sanbrado area was deposited during D2. The geometry of relatively higher grade mineralisation appears to define linear shoots parallel to F2 folds of lithological layering and S1 within overall S2- parallel envelopes. Flexures in the trend of D2 structures, such as S2 trend lines and D2 shears at major lithological boundaries (e.g., granodiorite-metasediment contacts), have also been important for focusing gold emplacement.

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Mankarga Mineralisation The Markoye Fault is interpreted to be the main conduit for the gold mineralisation found at Sanbrado, accommodating large scale fluid migration. Sanbrado hosts shear zone type quartz-vein gold mineralisation, similar to that found elsewhere in late Proterozoic Birimian terrains of West Africa. These orogenic type deposits exhibit strong relationships with regional arrays of major shear zones.

The structural environment can be described as brittle-ductile, coincident with a greenschist- amphibolite metamorphic facies. Arrays of transtensional features are apparent within a preferred S2 shear corridor and a strong shear controlled folding component is evident. High grade mineralisation occurs along the fold hinges and is marked by an intense silica-albite alteration assemblage. Gold is free and is mainly associated with minor pyrite and pyrrhotite disseminations and stringers that are foliation controlled.

The mineralised structures occasionally display syn to post mineralisation displacement due to plunging fold noses and minor late faulting.

Known mineralisation at M1 extends along strike for approximately 1km, is up to 50m wide and 150m in depth. The M5 mineralisation extends along strike for approximately 3km, is up to 100m wide and 300m in depth. The M3 mineralisation extends along strike for 750m, is up to 50m wide and 75m in depth. Mineralisation at all prospects remains open at depth.

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DEPOSIT TYPES Most West African gold deposits occur in the Lower Proterozoic part of the Craton, in sedimentary and volcano-sedimentary formations, which have undergone multiphase deformation, a setting analogous to many other gold-producing Precambrian greenstone belts in other parts of the world, such as the Abitibi Belt in Canada and the Norseman- Wiluna Belt in Australia.

Previous reviews of the geology of gold deposits in the West African region include those of Milési et al. (1989, 1992), Foster and Piper (1993), and Béziat et al. (2008). The most gold- endowed regions are spatially associated with the Baoulé-Mossi domain, the Archaean Kénéma-Man domain and the Kédougou-Kéniéba inlier, in the southwestern part of the West African Craton (Figure 8.1).

Figure 8.1 Location of the Major Gold Mines in the West African Region

(Source: modified after Holliday, 2014)

The West African Lower Proterozoic greenstone belts host a number of world class gold deposits such as Obuasi in Ghana (>20Moz), Sadiola in Mali (8Moz) and Morila in Mali (5.9Moz). Various gold deposit types have been recognised, namely orogenic, intrusion- related, skarn-hosted, porphyry-hosted, palaeoplacer and gold contained within iron oxide- copper-gold (IOCG) deposits (Markwitz et al., 2015).

Supergene or oxidised gold deposits are also known in the region, e.g., Yatela and Siguiri. The distribution of gold deposits is shown in Figure 8.1, with examples of different styles given below:

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. Shear Zone-Hosted Vein and hydrothermal alteration: Obuasi, Prestea, Poura, Belahouro. . Disseminated, Shear-Zone Hosted: Bogosu, Konongo. . Breccia Disseminated: Syama, Sabodala. . Intrusive Disseminated: Ayenfuri. . Porphyry copper-gold: Diénémera. . Tourmalinite: Loulo. . Intrusive Contact: Ity, Morila, Sadiola. . Stockwork: Bouda, Guibare, Essakane, Djambaye II in the Tabakoto camp . Extension Vein Stockwork: Kalana . Skarn: Ity . Palaeoplacer: Tarkwa . Disseminated sulphides and associated hydrothermal alteration in volcanic or plutonic rocks: Yaoure-Angovia Notwithstanding the variation in deposit types listed above, two main types of mesothermal gold mineralisation occur within the Birimian, namely quartz vein-hosted and disseminated sulphide types, with most having developed during the Eburnean Orogeny between 2.2 and 1.8Ga (Béziat et al., 2008; Leube et al., 1990).

The majority of gold deposits on the West African Craton are orogenic-style deposits (Robertson and Peters, 2016). Orogenic-style deposits are sited proximal to major accretionary structures within, or at the boundaries of, composite metamorphosed volcanic- plutonic or sedimentary terrains. Deposits are structurally hosted, associated with second or higher order splays of translithospheric faults (McCuaig and Kerrich, 1998). Mineralisation can be generally described as gold-bearing quartz veins, stringers and wall rock replacement, accompanied by minor sulphides and localised by brittle to ductile structures within various host rock types (McCuaig and Kerrich, 1998; Goldfarb et al, 2001). The majority of orogenic gold deposits occur within rocks that have been metamorphosed to greenschist facies, within a metamorphic pressure-temperature regime broadly corresponding to the brittle-ductile transition (McCuaig and Kerrich, 1998; Moritz, 2000).

Although a large number of such deposits share broad similarities, the nature of the gold distribution is highly variable at a local scale and between deposits. Mineralisation is typically associated with swarms of discontinuous veins of varying thickness and extent and as disseminations in sheared and altered wall rock. Gold occurs as native gold and/or associated with sulphides, mainly pyrite, pyrrhotite and arsenopyrite. Common criteria to all styles include a structural framework, receptive host rocks with strong competency/rheological contrasts and evidence of polyphase intrusives, deformation and alteration events (Robertson and Peters, 2016).

Gold mineralisation at the Mankarga deposits is associated with sheared and boudinaged quartz vein and veinlet arrays exhibiting silicification with sulphide, carbonate-albite and tourmaline-biotite alteration assemblages consistent with an orogenic gold deposit. Typically for gold deposits in tropical climates, the weathering profile is a well developed saprolitic cover consisting of extensive surface oxidation of bedrock. Gold deposits frequently display a surface oxide zone, a transition zone and a deeper sulphide zone. Gold is often free milling in the oxide zone.

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EXPLORATION Introduction Exploration activities on the Sanbrado Gold Project have included geological mapping, soil sampling, rock chip sampling, geophysical surveys and drilling. Drilling programs have included auger, air core (AC), reverse circulation (RC) and diamond core (DC) drilling techniques. From 1994 to 2012, all exploration activities were completed by Channel personnel and their consultants. From early 2014 exploration has been undertaken by WAF following its acquisition of Channel and the Sanbrado permit, which was part of the Channel tenement portfolio.

No exploration activities of substance were undertaken from mid-2012 until early 2014 when WAF began an infill drilling program at M5. WAF also completed an auger drilling program encompassing the entire original Tanlouka permit on a 400m by 100m grid to validate existing anomalies generated via historic soil sampling.

Survey Coordinate System The coordinate system used for all data collection and surveying on the Sanbrado property is the Universal Transverse Mercator (UTM) projection, Zone 30 North, using the World Geodetic System 1984 datum (WGS84).

Auger Drilling Auger drilling was completed across the entire Tanlouka permit in early 2014, in order to verify the existing soil geochemical anomalies and to cover unsampled areas in the western and southern parts of the tenement. The auger drilling was based on a 400m by 100m spaced grid. Areas of interest were infilled to 200m by 50m spacing. A figure showing the completed auger program is shown in Figure 9.1.

A total of 4,352 auger holes were completed for 22,915 meters drilled, resulting in 4,352 three meter interface sample composites.

Surface Mapping and Structural Review During 2015, WAF carried out detailed mapping of the southern area of the Tanlouka permit covering the Mankarga deposit areas at scales of 1:5,000 and 1:1,000. The results of the mapping were integrated with the geology observed in the drillholes.

A review of the structural history was carried out by Dr Davis of Orefind in June 2014 across the Mankarga, Manesse and Tanwaka prospects in an attempt to establish the relationship between gold mineralisation and structural history.

The geological history for Tanlouka was determined to comprise a three-stage deformation history. Gold mineralisation was emplaced during the second deformation (D2), which was a protracted and complex event comprising more than one stage of igneous intrusion, contact metamorphism, ductile structure formation, and extensive hydrothermal fluid flow.

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Figure 9.1 Sanbrado Gold Project – Completed Soil Auger Drilling Program

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DRILLING Introduction Since 2014, all drilling work has been managed directly by WAF’s personnel. WAF owns and operates a fleet of seven drill rigs including three auger rigs, one rotary air blast (RAB) rig, two multi-purpose RC-DC rigs and one dedicated DC rig capable of drilling to 500m. In Burkina Faso, WAF has a local exploration, drilling and support team of more than 50 people.

Three areas were taken into consideration for the resource evaluation: M5, M1 and M3.

The area of the M5 resource was sampled using RC, AC and DC drillholes on a nominal 100m by 25m grid spacing. A total of 723 AC holes (23,314m), 71 DC holes (14,537m), 41 DT holes (8,739m) and 69 RC holes (7,078m) were drilled by WAF between 2013 and 2017.. A total of 60 RC holes (7,296.2m) and 71 DC holes (15,439.6m) were drilled by Channel between 2010 and 2012. The strike of the mineralisation is approximately 030° and the dip is approximately 70°to 80° to the northwest. The majority of the DC holes were drilled at 120° azimuth and all DC holes have a declination of -50°. The majority of the RC holes were drilled at 300° azimuth and declination of -50°, in order to optimally intersect the mineralised zones.

The area of the M1 resource was sampled using a combination of RC, AC and DC drillholes on a nominal 100m by 25m grid spacing. . A total of 217 AC holes (3,290m), 42 DC holes (12,041m), 52 DT holes (14,113m) and 115 RC holes (12,602m) were drilled by WAF between 2015 and 2017. A total of 23 RC holes (3,060m) and 7 DC holes (1,199m) were drilled by Channel between 2010 and 2012. The strike of the mineralisation is approximately 315° and the dip is approximately 70°to 80° to the northeast. The majority of the DC and RC holes were drilled at 225° azimuth and declination of -50°, in order to optimally intersect the mineralised zones.

The area of the M3 resource was sampled using a combination of AC and DC drillholes on a nominal 50m by 25m grid spacing. A total of 269 AC holes (9,008m) and 4 DC holes (262.7m) were drilled by WAF in 2015. The strike of the mineralisation is approximately 360° and the dip is approximately 70° to 80° to the east. The majority of the DC and AC holes were drilled at 225° azimuth with a declination of -50°.

From 2000 to 2012 there were five drilling campaigns on the Sanbrado permit, including a RAB program in 2000 by Channel and a RC program in June 2003 by Orezone Resources (Orezone). Drilling completed by Channel included two phases of RC drilling (Jun 2010 to April 2011) and a DC drilling programme (Phase 3, Jul 2011 to Feb 2012). Only Channel’s Phase 1, 2 and 3 drill programs, as well as AC, RC and DC drilling by WAF have been used for estimating mineral resources.

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Drilling completed by Channel from 2000 to 2012 is well documented in the report authored by Smith (2012) titled “NI 43-101 Technical Report on Mineral Resources for the Mankarga 5 Gold Deposit Sanbrado Property, Burkina Faso for Channel Resources”. The drilling programs by Channel summarised in this section are based on that report. For additional information on drilling methods, procedures and results from Channel’s drilling please refer to the aforementioned report.

Table 10.1 shows all the drilling carried on the Mankarga prospects to date. Figure 10.1 shows the drillhole locations.

Table 10.1 Sanbrado Gold Project Summary of AC, RC and DC Drilling at the Mankarga Prospects

Year Company Prospect Drill Type Holes Meters M1 RC 5 768 2010 Channel M2 RC 3 506 M5 RC 15 2170.7 M1 RC 18 2292 2011 Channel RC 51 5667.5 M5 DD 61 13378.96 M1 DD 7 1199 2012 Channel M5 DD 15 2962.5 AC 109 4313 2013 WAF M5 DD 1 350.3 AC 393 13392.5 2014 WAF M5 DD 25 4630.4 M5 AC 125 2790.4 AC 356 6898.2 M1 DD 3 167.3 2015 WAF M2 AC 40 1541.85 AC 236 7788.65 M3 DD 3 167.16 AC 4 52 M1 RC 147 14450.6 DD 50 6957.6 AC 33 1219.15 2016 WAF M3 RC 9 962 DD 1 95.4 AC 48 1592.5 M5 RC 31 3514.7 DD 13 1358.55 AC 38 530 M1 RC 18 1699.4 DD 45 14976.2 2017 WAF AC 49 1226 M5 RC 37 3399.1 DD 50 13331.9

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Figure 10.1 Location of the Major Gold Mines in the West African Region

Drilling Procedures The drilling procedures used by WAF at the project from early 2014 are detailed in this section. WAF has established drilling and logging protocols and procedures to which all the company geologists adhere. These include systematic drillhole planning, site preparation, drilling and logging.

10.2.1 Hole Planning and Set-up New drilling programs and budgets are meticulously planned using all available exploration data on GIS and exploration modelling software. The programs are implemented subject to final revisions and approvals.

Holes were generally drilled perpendicular to the strike of mineralisation. At M5, holes were drilled on both southeast and northwest azimuths to intersect the steeply northwest dipping, northeast striking shear zone.

A hand-held GPS was used to locate and prepare the pads for the planned holes. The hole collars were spotted in the field and pegged using a GPS with an accuracy of measurement generally to within a few metres. The field geologist uses a compass to mark the fore and back sights to ensure correct drill rig alignment and orientation. The geologist also sets the hole declination angle.

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10.2.2 Surveying and Orientations Drillhole collars are surveyed using a DGPS in UTM grid WGS84 Zone 30 North. The DGPS measurements are validated using a known control station and generally have an accuracy of less than 10cm.

Downhole surveys used to map actual hole traces comprised single shot readings taken during drilling at 25m depth and every 50m thereafter.

DGPS RL data was acquired by WAF and is used for topographic control.

10.2.3 Hole Logging Procedures All drilling is logged to a standard that is appropriate for the category of resource which is being reported.

All measurements taken in the field, as well as all data obtained from sampling and logging of drillholes, are captured directly in fixed forms with menus on an Excel platform loaded in Toughbook computers. All parameters are logged using standard codes. The logs are checked daily by the senior geologist for completeness and accuracy.

Relevant non geological data such as hole ID, declination, azimuth, hole depth, bit size, date, water ingress, etc., are recorded on datasheets.

10.2.4 AC and RC Logging Logging of AC and RC chips is done at the drill site so that drillholes can be extended as necessary to ensure potential mineralised zones are intersected. The detailed logging of the cuttings records geological, mineralogical and structural features.

Small samples of screened and washed chips from every 1m depth are saved in clearly labelled plastic boxes (chip trays) bearing the depth markers and hole identification. Subsequent to representative sample collection, bags containing the remainder of the AC and RC chips are transported from the drill site to a bag farm close to the Poussighin campsite.

Bags are weighed directly after collection from the cyclone, prior to assay sample splitting, so that recoveries can be estimated accurately.

10.2.5 DC Logging Drill core is recovered by the drillers and stored in boxes with markers inserted after each run to indicate the depth and any core loss or gain. At the end of the shifts, the boxes are closed and transported to an enclosed storage area at the Poussighin core shed.

Geotechnical logging is carried out on all DC holes to obtain rock quality and structural information. The data are stored in the structural/geotechnical table of the database.

A geologist determines the Rock Quality Designation (RQD) at the core shed. Prior to logging, the core pieces are fitted together on rails in order to check the depth markers placed by the drillers and possible core mix-ups, and to calculate the core recovery. The reference line of the oriented core is drawn from the spear imprint and the geotechnical parameters are logged. The core is placed back in the boxes to clearly display the line of reference or the angle between the dominant fabric (bedding, foliation) and the core axis. Subsequently the lithological contacts, structural features, mineralised zones and the samples are marked up.

The following information is recorded in the core logs in an Excel spreadsheet:

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. Geology – Rock type, colour (using a standard colour chart), texture, grain size, weathering (oxide [strong/moderate/weak], transition, fresh), alteration, veins, sulphides, mineralogy. . Structure – Azimuth/dip and dip direction, shear, fracture, joint, infill, colour, thickness, bedding, crenulation, veins, quality of the measurement. . Sample sheet - Number, weight, mineralogy and abundance (volume%) of veins and mineralisation. . Geotechnical - Rock strength, weathering, joint sets with type, count, angle, alteration, infill, roughness. . Bulk Density. All core is photographed wet and dry before being marked and cut for assaying. The photographic records are downloaded as individual computer files, and relabelled according to their depth.

All drill core is laid out in clearly marked one meter long moulded plastic boxes stored at the core shed near the Poussighin campsite.

Dry Bulk Density The prospect area is moderately to deeply weathered (oxidised). The depth of the top of fresh rock over mineralised zones is generally around 50m to 60m below surface. Dry bulk densities are based upon density measurements completed by WAF (carried out internally) and Channel (carried out by SGS laboratories). Both utilised industry standard core immersion techniques.

Regolith domains were constructed, coded into the resource model and used to assign appropriate dry bulk densities. Dry bulk densities used for resource estimation are described in Section 14.

10.3.1 Historical Dry Bulk Density Testwork by Channel A total of 3,494 drill core density determinations were performed on drill core using water immersion methods at the SGS laboratory. Data were collected from a representative suite of all rock types, mineralisation types and grade ranges. A total of 3,372 drill core density determinations were made. There were 700 density determinations coded as saprolite and 2,672 density determinations coded as sulphide.

10.3.2 WAF Dry Bulk Density Testwork Dry bulk densities were estimated from density measurements carried out on more than 4,300 10cm core billets using the Archimedes method of weight in air versus weight in water. Over 750 determinations were from oxide material and over 3,500 determinations were from fresh rock samples. Core samples were dried and wax was used to seal the core prior to immersion. Steel billet density standards were used for QAQC.

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Drilling Results WAF commenced drilling on the Sanbrado Gold Project permit, focusing on upgrading the M5 deposit in early 2014. Drilling at M1 and M3 commenced in 2015. Drilling programmes remain ongoing in these areas.

Table 10.2 provides a summary of drilling completed by WAF to update the M5 Resource and also for the maiden resources at M1 and M3 as at 11 December 2016. Drillhole summary plans and typical cross-sections of M5 and M1 are presented as Figure 10.2 to Figure 10.5.

Table 10.2 Sanbrado Gold Project WAF AC, RC and DC Drilling included in Resource Estimate to 11 December 2016

AC RC Diamond Prospect Holes Metres Holes Metres Holes Metres Mankarga 1 North 177 4158 50 4417 9 768.3 Mankarga 1 South 183 2793 97 10766 44 6357.2 Mankarga 2* 40 1542 Mankarga 3 269 9007 9 962 4 384 Mankarga 5 675 22088 31 3515 40 7480 Total 1344 39588 187 19660 97 14898.5

* Note: Mankarga 2 was not part of any resource evaluation

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Figure 10.2 Sanbrado Gold Project – Mankarga 5 Drillhole Summary Plan

Figure 10.3 Sanbrado Gold Project – Mankarga 5 Typical Cross-section (SW650)

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Figure 10.4 Sanbrado Gold Project – Mankarga 1 North and South Drillhole Summary Plan

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Figure 10.5 Sanbrado Gold Project – Mankarga 1 South Section SE0450

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SAMPLE PREPARATION, ANALYSES AND SECURITY Four commercial laboratories are established in Ouagadougou, namely ALS-Chemex (formerly Abilab), SGS, Intertek and BIGS GLOBAL BURKINA SARL (BIGS), all of which offer sample preparation services, with three of them offering a range of analytical services including bottle-roll cyanidation and fire assay.

Sampling and assaying completed by Channel from 2000 to 2012 is well documented in the report authored by Smith (2012) entitled “NI 43-101 Technical Report on Mineral Resources for the Mankarga 5 Gold Deposit Sanbrado Property, Burkina Faso for Channel Resources”. The sampling and assaying programs by Channel summarised in this section are based on that report.

RC Sampling Methods

11.1.1 Historical Work During the Phase 1 and Phase 2 RC drilling programs undertaken by Channel, samples were collected at 2m intervals from the cyclone attached to the RC rig. Each sample was then passed through a multistage riffle splitter which effectively divided the sample into 1/8th and 7/8th portions. The 1/8th portion was passed through a single-stage riffle splitter to yield two equal samples of between 2kg and 3kg each.

Each split was placed in plastic sample bags. One bag was designated the archive portion and the other as the laboratory sample split. Stubs of the same sample ticket number were stapled to the top of both split portions. Sample numbers were also written on the plastic bags with red marker in the case of the archive split and black marker for the laboratory split. At the end of each day all samples were locked in a secure room.

All of the Phase 2 RC drillholes were downhole surveyed using a GYRO SMART instrument.

11.1.2 West African Resources Ltd AC and RC samples were collected on the rig at 1m intervals using a three tier splitter. All samples were dry, with samples weighed prior to splitting. The 1m splits were then numbered, bagged and dispatched to the laboratory for analysis. The remaining sampled material was archived securely on site.

Field QAQC procedures included the insertion of blank samples and duplicate samples as well as certified reference material as assay standards. The insertion rate of these QAQC samples averaged 3 in 20. Field duplicates were taken on 1m composites using a riffle splitter.

DC Sampling Methods Both the historic and the more recent WAF DC drilling sampling methods were very similar. Core was cut in half lengthways using a diamond saw.

Core was placed in the saw cradle facing downhole. For consistency, the orientation line was placed on the left side of the saw blade, with the right side of the core taken for assay sample. The half core with the orientation line was retained in the core box. If there was no orientation line for reference, the core was still placed in the saw cradle facing downhole with the right side taken for assay and the left side retained in the core box. Where possible core was sampled representatively without bias for mineralisation or structure.

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Assay samples of half core were generally taken at 1m intervals. Blanks and standard reference samples were inserted at a frequency of not less than one every 20 samples. Half core samples were placed in a calico sample bag and identified with the sample number. Sample duplicates were split from the original sample at the laboratory following primary crushing and inserted into a specifically numbered sample bag provided to the laboratory.

The drill core sampling methods and associated quality control protocols used by Channel and WAF are industry standard and appropriate.

Chain of Custody and Transport

11.3.1 Channel Resources Ltd The chain-of-custody for RC and DC samples collected and shipped from the exploration compound was as follows:

. RC samples or core boxes were transported to the exploration compound by the drill contractors or Channel geologists and placed in the logging area. . The logging and sample preparation area is a fenced compound and all prepared samples were locked in the sample holding room. Only authorised personnel had access to the sample holding room. . RC samples or split core samples were placed in sealed rice sacks in groups of about 15 to 20 samples per sack for shipping. . A sample submission form accompanied each batch of samples. . Sample batches were transported to Ouagadougou by Channel employees in company- owned vehicles. All pulp rejects and coarse rejects from the laboratories were retained by Channel and stored in two secure warehouses in Mogtedo.

11.3.2 West African Resources Ltd The AC and RC samples and the drill core retrieved by the drillers were collected and handled at the drill site by WAF personnel. The samples were transported to the core shed area at the basecamp at Poussighin for logging and sampling prior to delivery by WAF personnel to BIGS Ouagadougou for sample preparation. Whilst in storage, samples were kept under guard in a locked yard.

Tracking sheets were used to track the progress of batches of samples. The samples were continually under the direct control of WAF, including during preparation and shipment of the samples. This ensured effective chain of custody by WAF from the drill sites to the analytical laboratory.

Assay Sample Preparation

11.4.1 Channel Resources Ltd A total of 555 samples from the Phase 1 RC program and a total of 3,446 samples from the Phase 2 RC program were collected and submitted to Abilab Burkina SARL (Abilab). Samples, representing 2m intervals, were prepared using the following laboratory codes at Abilab:

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. WEI-21: all samples weighed as received from client. . DRY-21: high temperature drying (max temp=105°C) for rock chip and core samples. . LOG-22: sample received without barcode. . LOG-24: sample received without barcode. At least one out of every 50 samples is selected at random for routine pulp QC tests (LOG-QC). For routine pulps, the specification is 85% passing a 75µm screen. . CRU-31: used if rock particles are too large for PREP-31. . PREP-31: 250g to 500g are split off and pulverised in an LM2 mill to better than 85% passing 75µm. . SPL-21: crushed sample is split using a riffle splitter. . PUL-31: Pulverise a split or total sample of up to 250g to 85% passing 75µm or better. A total of 11,128 samples from the Phase 3 DC program were assayed, primarily at the SGS Burkina Faso SA (SGS) in Ouagadougou. Towards the end of the Phase 3 program, Channel sent some of the remaining samples to Abilab, due to the slow turnaround time experienced at SGS. Samples, representing 1.5m intervals, were prepared using the following laboratory codes at SGS:

. WGH79: weighing of samples and reporting of weight. . DRY10: Dry samples <3.0kg at 105°C. . SPL26: Sample volume reduction using riffle splitter; . CRU21: Crush <3.0kg to 75% passing 2mm. . PUL45: Pulverise 250g in Cr steel to 85% passing 75µm. All Phase 1 and 2 RC samples were then analysed by Abilab laboratory code Au-AA26 in which a 50g charge is subjected to standard fire assay (FA) with an atomic absorption spectrometry (AAS) finish detection limit for gold is 0.01g/t Au (0.01ppm).

In addition, Channel sent all of the Phase 1 RC coarse rejects for multi-element analyses at an ALS Chemex laboratory in Johannesburg, South Africa. The samples underwent a 35 element analysis by inductively coupled plasma mass spectrometry (Code ME-ICP41). No significant results were obtained.

The majority of the Phase 3 DC samples were prepared and analysed at SGS using laboratory code FAA505, in which a 50g charge is subjected to standard FA followed by an AAS finish.

Sample preparation checks for fineness were carried out by the laboratory as part of their internal procedures to ensure the grind size of 85% passing 75µm was being attained. Laboratory QAQC involves the use of internal lab standards using certified reference material, blanks, splits and duplicates as part of in house QAQC protocols.

Certified reference materials, having a good range of values, were inserted blindly and randomly. Results highlighted that sample assay values were accurate and that contamination had been minimised.

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Repeat or duplicate sample analyses revealed that precision of samples was within acceptable limits. For diamond core, one blank and one standard was inserted every 18 core samples. For RC samples, one blank, one standard and one duplicate were inserted every 17 samples.

11.4.2 West African Resources Ltd A total of 18,825 RC samples and 5,602 DC samples (including QAQC samples) from the WAF drilling programs were collected and submitted to BIGS. Samples were prepared using the following laboratory codes at BIGS:

. PRG100: Sample reception/registration, including weight of sample as received. On registration, a RFID label is created (no barcode used). . PRD100: Drying, in stainless trays at 90°C max (unless other temperature requested). . PRJ200: Samples that are >10mm particle size are jaw crushed to <2mm before pulverising. . PRR200: Sample mass reduction using a rotary divider (single split). . CPO-B22: Combination Code. Includes weighing, drying, jaw crushing (<3kg), rotary split, 1 portion pulverised (LM2) to >95% passing 75µm. . PREPLOG: A repeat is taken before pulverising for 1 in 20 samples. Pulverising (75µm wet sieving) and jaw crushing (2mm dry sieve) logs are created and available. Conclusions All sample preparation and analyses were carried out at independent laboratories in Ouagadougou, Burkina Faso. No aspect of laboratory sample preparation or analysis was conducted by an employee, officer, director or associate of WAF.

WAF and predecessors have used a combination of duplicates, checks, blanks and standards to ensure suitable quality control of sampling methods and assay testing. The procedures and QA/QC management are consistent with good industry practice and are deemed fit for purpose. Results of recent sampling have not identified any issues which materially affect the accuracy, reliability or representativeness of the results.

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DATA VERIFICATION Independent Qualified Person Review and Verification Mr Brian Wolfe visited the Sanbrado Gold Project in May 2014, May 2016 and April 2017. Steps undertaken to verify the integrity of data used in this report include:

. Field visits to the areas outlined in this report including the Sanbrado Gold Project permit and M1, M3 and M5 deposits. . Inspection of drill core and RC drill chips. . Inspection of AC, RC and DC drilling activities, sampling and logging procedures. . Review of WAF’s data collection, database management and data validation procedures. . Review of the previous NI 43-101 report for the project titled “NI 43-101 Technical Report on Mineral Resources for the Mankarga 5 Gold Deposit Tanlouka Property” September 2014. The Qualified Person has reviewed and cross-checked sections of this Report prepared by WAF.

The Qualified Person also completed the updated resource estimate for the M5 and M1 South Deposits in addition to the resource estimates for M1 North and M3. Additional data verification steps undertaken during this estimate process included the following:

. Validation of drilling, geology and assay database (including checks overlapping intervals, samples beyond hole depth and other data irregularities. . Review of WAF QAQC charts for standards, blanks and duplicates. . Visual and statistical analysis of resource estimate model outputs versus primary data. . Random cross checks of assay hardcopy reports against the database. Based on this review work, the Qualified Person is of the opinion that the dataset provided by WAF is of an appropriate standard to use for resource estimation work.

QAQC Data Analysis QAQC data from WAF’s 2014 to 2016 drilling programs is presented in the following sections. QAQC data from Channel’s drilling from 2000 to 2014 is documented in the report authored by Smith (2012) entitled “NI 43-101 Technical Report on Mineral Resources for the Mankarga 5 Gold Deposit Tanlouka Property, Burkina Faso for Channel Resources”. For information on the QAQC programs from Channel’s programs please refer to the aforementioned report.

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The following Sanbrado Gold Project data has been assessed:

. Standards: The standards assays provide a measure of relative analytical precision. Sample batches which fall outside specified tolerance ranges are flagged and investigated. . Laboratory Repeats (i.e. represents a second 50g pulp sample taken from the first 200g pulp). . Laboratory Duplicates (i.e. represents a second 200g pulp sample taken from the initial pulverised material). . Field Duplicates (i.e. represents a second 3kg sample split taken from the original RC field sample interval. This sample is taken at the same time as the original sample). Statistical Evaluation The quality control data has been statistically evaluated, and summary plots have been produced for interpretation using the functions available in QC Assure software. The types of plots produced are summarised below:

12.3.1 Thompson and Howarth Plot (T & H Plot) Shows the mean relative error in percent of grouped assay pairs across the entire grade range. Used to visualise precision levels by comparing against given control lines.

12.3.2 Rank % HARD Plot Ranks all assay pairs in terms of precision levels measured as half of the absolute difference from the mean of the assay pairs (% HARD). Used to visualise relative precision levels, and to determine the percentage of the assay pairs population occurring at a certain precision level.

12.3.3 Mean vs. % HARD Plot Used as another way of illustrating relative precision levels by showing range of % HARD over the grade range.

12.3.4 Mean vs. %HRD Plot Similar to Mean vs. %HARD Plot, but the sign is retained, thus allowing negative or positive differences to be computed. This plot gives an overall impression of precision, and also shows whether or not there is significant bias between the assay pairs by illustrating the mean percent half relative difference between the assay pairs (mean % HRD).

12.3.5 Correlation Plot This is a simple plot of the value of assay 1 against assay 2. This plot allows an overall visualisation of precision and bias over selected grade ranges. Correlation coefficients are also used.

12.3.6 Quantile-Quantile Plot This is a similar plot to the correlation plot, but smooths the data by reducing the effects of outliers. This is achieved by grouping the assay pairs into quantiles of grade. The Q-Q plot is a good indicator of bias (accuracy).

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12.3.7 Standard Control Plot This plot shows the assay results of a particular reference standard over time. The results can be compared to the expected value, and the ±10% precision lines were also plotted. This gives a good indication of precision and accuracy.

12.3.8 Cusum Plots These plots illustrate the cumulative sum of the deviation from the expected value of a particular reference standard or from the mean of the assays over time. Used to determine direction and severity of bias, and to illustrate changes in grade over time.

Results of the analysis of the BIGS laboratory data are presented the following sections.

Fire Assay Analyses

12.4.1 Standards Summary plots for the standards are presented in Figure 12.1 to Figure 12.10. The standard details and comments are presented in Table 12.1.

Table 12.1 WAF Internal Standards

Expected No of Standard % Samples in % Standard Value EV Range Min Max Mean Assays Deviation EV Range bias Au g/t

G310-1 112 4.94 4.446-5.434 4.795 5.040 4.915 0.052 100 -0.5 G310-6 102 0.65 0.585-0.712 0.587 0.666 0.615 0.009 100 -5.4 G311-1 839 0.52 0.468-0.572 0.375 0.619 0.510 0.028 92.49 -1.9 G313-6 94 4.94 4.446-5.434 4.67 5.586 4.943 0.139 98.9 0.1 G901-1 543 2.58 2.322-2.838 2.257 2.925 2.644 0.064 99.3 2.5 G901-3 108 2.87 2.583-3.157 2.75 3.19 2.975 0.08 99.1 3.7 G901-7 1,281 1.52 1.368-1.672 1.368 1.829 1.491 0.051 97.7 -1.9 G908-7 293 4.82 4.338-5.302 3.866 5.422 4.773 0.164 97.3 -1.0 G909-5 340 2.63 2.367-2.893 2.440 3.078 2.6423 0.097 96.8 -0.3 G913-9 209 4.91 4.419-5.401 4.759 5.110 4.945 0.063 100 0.7 G914-10 388 10.26 9.234-11.286 8.62 10.900 10.155 0.287 97.7 -1.0 G995-1 47 2.75 2.475-3.025 2.514 3.346 2.815 0.19 85.1 2.4 G996-4 305 0.51 0.459-0.561 0.387 0.657 0.475 0.04 65.6 -6.8 GS-1P5F 31 1.40 1.26-1.54 1.115 1.757 1.375 0.14 77.4 -1.5 GS-P7H 36 0.799 0.719-0.879 0.696 0.98 0.771 0.05 88.9 -3.6

In general, the standard assay results indicate that acceptable accuracy was achieved, with all but one of the standards (G996-4) showing an acceptable range of results, given that two other standards (G310-6 and G311-1) with a similar Expected Value (EV) demonstrate good performance in comparison to G996-4. Use of G996-4 was discontinued in September 2016. Standard G311-1 demonstrates a high range of values from batches since the end of June 2017. It is recommended to review whether substitution of standards has erroneously taken place.

Overall the variability is within acceptable limits and indicates a high level of accuracy for the analytical laboratory and the assay method.

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Figure 12.1 Standard G310-1

Summary (Standard: G310-1)

Standard: G310-1 No of Analyses: 112 Element: Au-FPF500_ppm Minimum: 4.795 Units: ppm Maximum: 5.040 Detection Limit: 0.005 Mean: 4.915 Expected Value (EV): 4.940 Std Deviation: 0.052 E.V. Range: 4.446 to 5.434 % in Tolerance 100.000 % % Bias -0.510 % % RSD 1.050 %

Standard Control Plot (Standard: G310-1)

5.4

5.2

5.0

4.8

4.6 Au-FPF500_ppm (ppm)

4.4 228760 234520 234720 234920 235200 235600 236120 237380 241320 241620 241840

CheckID

Au-FPF500_ppm Expected Value = 4.940 EV Range (4.446 to 5.434) Mea n of Au-FPF500_ppm = 4.915

Cumulative Deviation from Assay Mean (Standard: G310-1)

1.0

0.5

0.0

-0.5

-1.0

-1.5 228760 234520 234720 234920 235200 235600 236120 237380 241320 241620 241840 CumulativeAu-FPF500_ppmof Sum Mean (ppm) - CheckID

Au-FPF500_ppm Mea n of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = -0.331

Cumulative Deviation from Expected Value (Standard: G310-1)

-1

-2

-3

-4 228760 234520 234720 234920 235200 235600 236120 237380 241320 241620 241840

CheckID CumulativeofAu-FPF500_ppm Sum Expected - Value(pp

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Expected Value (ppm) = -1.756

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Figure 12.2 Standard G310-6

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Figure 12.3 Standard G311-1

Summary (Standard: G311-1)

Standard: G311-1 No of Analyses: 839 Element: Au-FPF500_ppm Minimum: 0.375 Units: ppm Maximum: 0.619 Detection Limit: 0.005 Mean: 0.510 Expected Value (EV): 0.520 Std Deviation: 0.028 E.V. Range: 0.468 to 0.572 % in Tolerance 92.491 % % Bias -1.883 % % RSD 5.413 %

Standard Control Plot (Standard: G311-1)

0.70

0.60

0.50

0.40 Au-FPF500_ppm (ppm)

0.30 172240 178000 183400 207460 214100 226940 237700 252120

CheckID

Au-FPF500_ppm Expected Value = 0.520 EV Range (0.468 to 0.572) Mean of Au-FPF500_ppm = 0.510

Cumulative Deviation from Assay Mean (Standard: G311-1)

-1

-2

-3

-4

-5 172240 178000 183400 207460 214100 226940 237700 252120 Cumulative Sum of Au-FPF500_pp m - Mean (ppm) CheckID

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = -2.885

Cumulative Deviation from Expected Value (Standard: G311-1)

-2

-4

-6

-8

-10

-12 172240 178000 183400 207460 214100 226940 237700 252120

CheckID Cumulative Sum of Au-FPF500_ppm Expected- Value (pp

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Expected Value (ppm) = -6.998

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Figure 12.4 Standard G313-6

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Figure 12.5 Standard G901-1

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Figure 12.6 Standard G901-3

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Figure 12.7 Standard G901-7 Summary (Standard: G901-7)

Standard: G901-7 No of Analyses: 1281 Element: Au-FPF500_ppm Minimum: 0.936 Units: ppm Maximum: 1.829 Detection Limit: 0.005 Mean: 1.491 Expected Value (EV): 1.520 Std Deviation: 0.051 E.V. Range: 1.368 to 1.672 % in Tolerance 97.736 % % Bias -1.917 % % RSD 3.403 %

Standard Control Plot (Standard: G901-7)

1.8

1.6

1.4

1.2

Au-FPF500_ppm (ppm) Au-FPF500_ppm 1.0 163000 173680 181920 188500 190520 200340 207760 217240 237540 245220 250920 255100

CheckID

Au-FPF500_ppm Expected Value = 1.520 EV Range (1.368 to 1.672) Mean of Au-FPF500_ppm = 1.491

Cumulative Deviation from Assay Mean (Standard: G901-7)

-2

-4 163000 173680 181920 188500 190520 200340 207760 217240 237540 245220 250920 255100 Cumulative SumofAu-FPF500_ppm Mean - (ppm) CheckID

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = 1.283

Cumulative Deviation from Expected Value (Standard: G901-7)

-10

-20

-30

-40 163000 173680 181920 188500 190520 200340 207760 217240 237540 245220 250920 255100

CheckID Cumulative SumofAu-FPF500_ppm Expected - Value(pp

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Expected Value (ppm) = -17.390

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Figure 12.8 Standard G908-7 Summary (Standard: G908-7)

Standard: G908-7 No of Analyses: 293 Element: Au-FPF500_ppm Minimum: 3.866 Units: ppm Maximum: 5.422 Detection Limit: 0.005 Mean: 4.773 Expected Value (EV): 4.820 Std Deviation: 0.164 E.V. Range: 4.338 to 5.302 % in Tolerance 97.270 % % Bias -0.970 % % RSD 3.444 %

Standard Control Plot (Standard: G908-7)

5.5

5.0

4.5

4.0 Au-FPF500_ppm (ppm)

3.5 171260 205100

CheckID

Au-FPF500_ppm Expected Value = 4.820 EV Range (4.338 to 5.302) Mean of Au-FPF500_ppm = 4.773

Cumulative Deviation from Assay Mean (Standard: G908-7)

-2

-4

-6

-8

-10 171260 205100 Cumulative Sum of Au-FPF500_ppm Mean- (ppm) CheckID

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = -5.241

Cumulative Deviation from Expected Value (Standard: G908-7)

-5

-10

-15

-20 171260 205100

CheckID Cumulative Sumof Au-FPF500_ppm - Expected Value (pp

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Expected Value (ppm) = -12.113

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Figure 12.9 Standard G909-5 Summary (Standard: G909-5)

Standard: G909-5 No of Analyses: 340 Element: Au-FPF500_ppm Minimum: 2.440 Units: ppm Maximum: 3.078 Detection Limit: 0.005 Mean: 2.623 Expected Value (EV): 2.630 Std Deviation: 0.097 E.V. Range: 2.367 to 2.893 % in Tolerance 96.765 % % Bias -0.264 % % RSD 3.707 %

Standard Control Plot (Standard: G909-5)

3.0

2.8

2.6

2.4 Au-FPF500_ppm (ppm)

2.2 228780 248600 256460

CheckID

Au-FPF500_ppm Expected Value = 2.630 EV Range (2.367 to 2.893) Mean of Au-FPF500_ppm = 2.623

Cumulative Deviation from Assay Mean (Standard: G909-5)

-2 228780 248600 256460 Cumulative Sum of Au-FPF500_ppm - Mean (ppm) CheckID

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = 4.039

Cumulative Deviation from Expected Value (Standard: G909-5)

-2

-4 228780 248600 256460

CheckID Cumulative Sum of Au-FPF500_ppm - Expected Value (pp

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Expected Value (ppm) = 2.857

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Figure 12.10 Standard G913-9

Summary (Standard: G913-9)

Standard: G913-9 No of Analyses: 209 Element: Au-FPF500_ppm Minimum: 4.759 Units: ppm Maximum: 5.110 Detection Limit: 0.005 Mean: 4.945 Expected Value (EV): 4.910 Std Deviation: 0.063 E.V. Range: 4.419 to 5.401 % in Tolerance 100.000 % % Bias 0.706 % % RSD 1.278 %

Standard Control Plot (Standard: G913-9)

5.4

5.2

5.0

4.8

4.6 Au-FPF500_ppm (ppm)

4.4 217300 218160 219300 219660 220640 224160 227140 227340 228120 231120 245100 245780 246180 246720 247420 248680 249500 249920 257980 265680

CheckID

Au-FPF500_ppm Expected Value = 4.910 EV Range (4.419 to 5.401) Mea n of Au-FPF500_ppm = 4.945

Cumulative Deviation from Assay Mean (Standard: G913-9)

-1

-2

-3

-4

-5 217300 218160 219300 219660 220640 224160 227140 227340 228120 231120 245100 245780 246180 246720 247420 248680 249500 249920 257980 265680 Cumulative Sum of Au-FPF500_ppm - Mean (ppm) CheckID

Au-FPF500_ppm Mea n of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = -2.265

Cumulative Deviation from Expected Value (Standard: G913-9)

-2 217300 218160 219300 219660 220640 224160 227140 227340 228120 231120 245100 245780 246180 246720 247420 248680 249500 249920 257980 265680

CheckID Cumulative Sum of Au-FPF500_ppm Expected- Value (pp

Au-FPF500_ppm Mea n of Cumulative Sum of Au-FPF500_ppm - Expected Va lue (ppm) = 1.374

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Figure 12.11 Standard G914-10

Summary (Standard: G914-10)

Standard: G914-10 No of Analyses: 388 Element: Au-FPF500_ppm Minimum: 8.620 Units: ppm Maximum: 10.900 Detection Limit: 0.005 Mean: 10.155 Expected Value (EV): 10.260 Std Deviation: 0.287 E.V. Range: 9.234 to 11.286 % in Tolerance 97.680 % % Bias -1.025 % % RSD 2.824 %

Standard Control Plot (Standard: G914-10)

12.0

11.0

10.0

9.0 Au-FPF500_ppm (ppm)

8.0 238220 244840 259020

CheckID

Au-FPF500_ppm Expected Value = 10.260 EV Range (9.234 to 11.286) Mean of Au-FPF500_ppm = 10.155

Cumulative Deviation from Assay Mean (Standard: G914-10)

-10

-20

-30 238220 244840 259020 Cumulative Sum of Au-FPF500_ppm - Mean (ppm) Au-FPF500_ppm of- Sum Cumulative CheckID

Au-FPF500_ppm Mea n of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = -8.070

Cumulative Deviation from Expected Value (Standard: G914-10)

-10

-20

-30

-40

-50 238220 244840 259020

CheckID Cumulative Sumof Au-FPF500_pp Expected m - Value(pp

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Expected Value (ppm) = -28.519

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Figure 12.12 Standard G995-1

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Figure 12.13 Standard G996-4

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Figure 12.14 Standard GS-1P5F

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Figure 12.15 Standard GS-P7H

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12.4.2 Blanks Blanks are used to test for contamination during the sample preparation process, and are usually composed of clean silica sands, such as pool filter sand or beach sands. Nonmineralised material can also form a good medium for blanks.

The blank material used by WAF at M5 has been sourced by the company from southwestern Burkina Faso, and the material was tested by two different laboratories. The quality of the blanks is deemed adequate, and gives an appropriate measure of the quality of the sample preparation at BIGS Global.

Figure 12.16 and Figure 12.17 show that the blank materials BLANKWAR4 and BLANKWAR7 generally give consistent results. A number of outliers are observed, however the overall results are within acceptable limits. Table 12.2 presents WAF Blank Sample results.

Table 12.2 WAF Blanks

Expected No of Standard Value EV Range Min Max Mean % Samples in EV Range Assays Au g/t

BLANKWAR4 1,193 0.005 0.001-0.1 0.002 0.133 0.003 97.8 BLANKWAR7 4,589 0.005 0.001-0.1 0.002 0.12 0.002 99.5

12.4.3 Pulp (Laboratory) Repeats Statistics and summary plots for the FA repeats are presented in Figure 12.18. A total of 4,800 data pairs are available however this reduces to 1,529 when a lower cutoff grade of 0.1g/t Au is applied.

Precision of the repeats for the FA method is considered good, with 99.1% of the assay pairs reporting within acceptable limits (10% HARD), and no bias present in the dataset.

12.4.4 Field Duplicates Field duplicates are used to determine sampling error and also give an indication of the precision of the data pairs (original vs. duplicate). The quality of the data will depend greatly on the quality of the actual duplicate prepared in the field. RC samples are relatively easy to duplicate by producing a second split on site, however representative DC duplicates are difficult to prepare in the field. The archived portion of the half core is often used as a field duplicate, however, two halves of a length of core are not really comparable and produce poorly correlated results.

DC duplicates are best prepared by producing a second sample split after primary crushing at the laboratory, which is the method adopted by WAF.

Statistics and summary plots for the RC field duplicate data are presented in Figure 12.19. A lower cutoff grade of 0.1g/t Au was applied to the dataset and 1,307 pairs were available for analysis (4,785 uncut).

Precision of the RC field duplicates for the FA method is within acceptable limits. The duplicates correlate well with the original and 93.9% of all the data fall within an acceptable 30% precision limit.

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Figure 12.16 BLANKWAR4

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Figure 12.17 BLANKWAR7 Summary (Standard: BLANKWAR7)

Standard: BLANKWAR7 No of Analyses: 4589 Element: Au-FPF500_ppm Minimum: 0.002 Units: ppm Maximum: 0.120 Detection Limit: 0.005 Mean: 0.002 Expected Value (EV): 0.002 Std Deviation: 0.003 E.V. Range: 0.001 to 0.010 % in Tolerance 99.542 % % Bias 9.218 % % RSD 119.071 %

Standard Control Plot (Standard: BLANKWAR7)

0.10

0.08

0.06

0.04

0.02 Au-FPF500_ppm (ppm)

0.00 197428 215061 232061 251521

CheckID

Au-FPF500_ppm Expected Value = 0.002 EV Range (0.001 to 0.010) Mean of Au-FPF500_ppm = 0.002

Cumulative Deviation from Assay Mean (Standard: BLANKWAR7)

0.2

0.1

0.0

-0.1

-0.2

-0.3

-0.4 197428 215061 232061 251521 Cumulative Sum of Au-FPF500_ppm - Mean (ppm) CheckID

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Mean (ppm) = -0.123

Cumulative Deviation from Expected Value (Standard: BLANKWAR7)

-1

-2

-3 180861 182537 184221 187801 189441 191121 192801 194481 196101 197428 200341 201821

CheckID Cumulative Cumulative ofSum Au-FPF500_ppm Expected- Value (pp

Au-FPF500_ppm Mean of Cumulative Sum of Au-FPF500_ppm - Expected Value (ppm) = 0.300

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Figure 12.18 Laboratory Repeats

Summary LabRpt (> 0.1 ppm Au)

Au_Sample Au_Repeat Units Result No. Pairs: 1,529 1,529 Pearson CC: 0.996 Minimum: 0.101 0.101 ppm Spearman CC: 0.999 Maximum: 320.000 330.555 ppm Mean HARD: 1.684 Mean: 3.470 3.492 ppm Median HARD: 0.989 Median 0.376 0.380 ppm Std. Deviation: 17.804 17.875 ppm Mean HRD: -0.043 Coefficient of Variation: 5.130 5.119 Median HRD -0.137

Mean vs. HARD Plot Rank HARD Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

20 100 15 80 10 60 99.084% of data are with in Precision limits 40 5 HARD (%)

HARD (%) 20 0 0.1 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)

Mean HARD: 1.684 Median HARD: 0.989 Precision: 10% Precision: 10%

HRD Histogram Mean vs. HRD Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

100 20 80 10 60 0 40

20 HRD (%) -10

Frequency (%) Frequency 0 -20 -1.0 0.0 1.0 0.1 1 10 100 1000 HRD (/100) Mean of Data Pair (ppm)

Mean HRD: -0.043 Median HRD: -0.137 Mean HRD: -0.043 Median HRD: -0.137 Precision: +/-10.000% Precision: +/-10%

T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (> 0.1 ppm Au) (> 0.1 ppm Au)

1000 100 10 10 1 0.1 0.1 0.01

0.001 Median AD (ppm) 0.001 0.1 1 10 100 1000 0.1 1 10 100

Absolute Difference (ppm) Difference Absolute Mean of Data Pair (ppm) Grouped Mean of Pair (ppm)

10% 20% 30% 50% 10% 20% 30% 50%

Correlation Plot QQ Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

400 400

300 300

200 200 100 100 0

Au_Repeat (ppm) Au_Repeat 0 0 100 200 300 400 (ppm) Au_Repeat 0 100 200 300 400 Au_Sample (ppm) Au_Sample (ppm)

P.CC= 0.996 S.CC= 0.999 Ref. Line y = 1.000x + 0.020 Ref. Line y = 1.003x + 0.010

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Figure 12.19 Field Duplicates

Summary FieldDuplicate (> 0.1 ppm Au)

Au_Sample Au_FD Units Result No. Pairs: 1,307 1,307 Pearson CC: 0.973 Minimum: 0.101 0.101 ppm Spearman CC: 0.943 Maximum: 112.261 93.201 ppm Mean HARD: 8.411 Mean: 1.344 1.329 ppm Median HARD: 4.043 Median 0.332 0.345 ppm Std. Deviation: 5.968 5.484 ppm Mean HRD: -0.179 Coefficient of Variation: 4.441 4.125 Median HRD 0.000

Mean vs. HARD Plot Rank HARD Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

100 100

80 80 93.879% of data are within Precision limits 60 60 40 40

HARD (%) 20

HARD (%) 20 0 0.1 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)

Mean HARD: 8.411 Median HARD: 4.043 Precision: 30% Precision: 30%

HRD Histogram Mean vs. HRD Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

50 100 40 50 30 0 20

10 HRD (%) -50

Frequency (%) 0 -100 -1.0 0.0 1.0 0.1 1 10 100 1000 HRD (/100) Mean of Data Pair (ppm)

Mean HRD: -0.179 Median HRD: 0.000 Mean HRD: -0.179 Median HRD: 0.000 Precision: +/-30% Precision: +/-30%

T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (> 0.1 ppm Au) (> 0.1 ppm Au)

100 10

10 1 1 0.1 0.1 0.01 0.01

0.001 Median AD (ppm) 0.001 0.1 1 10 100 1000 0.1 1 10 100

Absolute Difference (ppm) Difference Absolute Mean of Data Pair (ppm) Grouped Mean of Pair (ppm)

10% 20% 30% 50% 10% 20% 30% 50%

Correlation Plot QQ Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

150 150

100 100

50 50 Au_FD (ppm) 0 Au_FD (ppm) 0 20 40 60 80 100 120 140 0 0 20406080100120140 Au_Sample (ppm) Au_Sample (ppm)

P.CC= 0.973 S.CC= 0.943 Ref. Line y = 0.894x + 0.128 Ref. Line y = 0.913x + 0.102

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DC duplicates, where two splits are taken after primary crushing of the sample, were also analysed. A total of 3,951 splits were taken however applying a lower cutoff grade of 0.1g/t Au reduces the number of available pairs to 1,153. Statistical results are shown in Figure 12.20.

Figure 12.20 Laboratory Duplicates Summary LabDup (> 0.1 ppm Au)

Au_Duplicat Au_Sample e Units Result No. Pairs: 1,153 1,153 Pearson CC: 1.000 Minimum: 0.101 0.101 ppm Spearman CC: 0.999 Maximum: 349.153 343.219 ppm Mean HARD: 1.019 Mean: 2.469 2.465 ppm Median HARD: 0.662 Median 0.325 0.326 ppm Std. Deviation: 18.104 18.062 ppm Mean HRD: -0.059 Coefficient of Variation: 7.332 7.326 Median HRD -0.119

Mean vs. HARD Plot Rank HARD Plot (> 0.1ppm Au ppm) (> 0.1 ppm Au)

10 100 80

5 60 40 HARD (%)

HARD (%) 20 0 0.1 1 10 100 1000 0 0 102030405060708090100 Mean of Data Pair (ppm) Rank (%)

Mean HARD: 1.019 Median HARD: 0.662 Precision: 10% Precision: 10%

HRD Histogram Mean vs. HRD Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

100 10 5 50 0

HRD (%) -5

Frequency (%) Frequency 0 -10 -1.0 0.0 1.0 0.1 1 10 100 1000 HRD (/100) Mean of Data Pair (ppm)

Mean HRD: -0.059 Median HRD: -0.119 Mean HRD: -0.059 Median HRD: -0.119 Precision: +/-10% Precision: +/-10%

T & H Precision Plot (Assay Pairs) T & H Precision Plot (Grouped Pairs) (> 0.1 ppm Au) (> 0.1 ppm Au)

1000 100 10 10 1 0.1 0.1 0.01 0.001 0.001 Median AD (ppm) 0.1 1 10 100 1000 0.1 1 10 100

Absolute Difference (ppm) Difference Absolute Mean of Data Pair (ppm) Grouped Mean of Pair (ppm)

10% 20% 30% 50% 10% 20% 30% 50%

Correlation Plot QQ Plot (> 0.1 ppm Au) (> 0.1 ppm Au)

400 400 300 300

200 200 100 100 0 0

Au_Duplicate (ppm) Au_Duplicate 0 100 200 300 400 Au_Duplicate (ppm) 0 100 200 300 400 Au_Sample (ppm) Au_Sample (ppm)

P.CC= 1.000 S.CC= 0.999 Ref. Line y = 0.998x + 0.002 Ref. Line y = 0.998x + 0.002

Printed: 16-Oct-2017 12:02:45 Data Imported: 13-Oct-2017 16:37:01Page 1

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 92 West African Resources Limited

The graphs show that 100% of the pairs are within a 10% HARD.

Similarly, the T&H plots show a mean relative error for most of the pairs of well below 10%, which is good. The correlation plots show that there is no marked bias in the data.

Conclusions These methods of data verification are considered at or above industry standard. The results of the QAQC data analyses discussed in the preceding sections demonstrate that the quality of the data is acceptable for use in mineral resource estimation.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 93 West African Resources Limited

METALLURGY AND PROCESS DEVELOPMENT Introduction WAF is currently completing a metallurgical testwork program designed to determine the amenability of the various feed materials to conventional gravity separation followed by agitated leaching and the associated carbon in pulp/carbon in leach (CIP/CIL) processing methods. The metallurgical samples subjected to testwork programs were selected from the M1 and M5 deposits. No metallurgical testwork was conducted on the very small M3 deposit (less than 2% of the resource), as the mineralogy and assaying from the geological drilling indicated similar mineralogical behaviour to oxide material from deposits M1 and M5.

Testwork undertaken to date supports the application of the proposed processing technology to the oxidised, partially oxidised and primary feed types, with low reagent consumptions and moderate leach times. Some design-specific testing will be undertaken to confirm a number of detailed design parameters and specifically address the behaviour of deeper high grade samples from the M1 deposit. Additional test work is in progress and will be updated in the Definitive Feasibility Study technical report due for completion in mid-2018.

History of Testwork Channel completed preliminary mineralogical and metallurgical testwork at SGS Laboratories in Vancouver, BC, in 2012. This work is documented in the report authored by Smith (2012) entitled “NI 43-101 Technical Report on Mineral Resources for the Mankarga 5 Gold Deposit Tanlouka Property” which is summarised in Section 13.4.

In 2014 a testwork program was undertaken by WAF at the ALS Global facility located in Perth Western Australia. This program investigated the amenability of the Sanbrado ores to heap leach processing methods. Initially samples from 4 drillholes were subjected to testing. A further 6 holes were drilled and the program was expanded and progressed into 2015.

Whilst this heap leach program is not directly influencing the current project design and economic considerations due to the different technology proposed, selective data was generated which is applicable to the comminution and agitated leaching sections of the plant. Relevant data from this initial heap leach program is detailed in Section 13.5.

As exploration drilling progressed, the additional resource was considered to be capable of supporting an agitated leaching flowsheet. Consequently, testwork was undertaken at the ALS Global facility in Perth to investigate conventional leaching behaviour under selected conditions. Comminution work was also conducted to evaluate various grinding options such as Semi-Autogenous Grinding (SAG) methods. Thickening testwork was undertaken by Outotec.

Dedicated metallurgical testwork drillholes were drilled and samples collected for a number of testwork programs, which were undertaken though 2016 and into early 2017. The results of these testwork programs are contained in ALS report numbers A15684, A16557, A17283, A17341 and A17750.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 94 West African Resources Limited

Background to the Metallurgical Sample Collection by WAF Three metallurgical sample drilling programs were undertaken from late 2013 to mid-2016. In addition, high grade assay rejects from a number of RC and DC drillholes were provided to the ALS laboratory in January 2017 to further evaluate the behaviour of high grade material.

The first program targeted predominantly oxide and partially oxidised material at the M5 deposit for heap leaching testwork. Drilling was undertaken during December 2013 through to January 2014. A series of deeper holes were drilled during September and October 2014 to collect fresh and weakly oxidised samples. These holes are referred to as Tanmet 1 to Tanmet 10.

The second program, targeting geotechnical and metallurgical samples, was conducted during February and March 2015.These holes are referred to as Tan15–GT-Met01 to Tan15- GT-Met10.

The third program targeted the M1 deposit (4 holes), and to a lesser extent fresh material from the M5 deposit (2 holes) in June 2016. This program provided full core for comminution and metallurgical performance testing. These holes are referred to as Tan16- Met11 to Tan16-Met16.

To evaluate high grade material from the M1 deposit, a series of high grade crushed intervals were sourced from DC holes 44, 52 and 55 and RC holes 122 and 131.

As noted, the M3 deposit has not been tested metallurgically. It is a very small proportion of the resource and geologically presents the same characteristics as the strongly oxidised material from M5 and M1, suggesting that material from M3 will be amenable to the same flowsheet proposed for M5 and M1. It is also envisaged in the mining schedule that the M3 deposit will be blended with M1 and/or M5 material during operations.

Four metallurgical domains were identified during the pre-feasibility stage of the program, namely Fresh (FRS), Weak Oxidation (WOX), Mild Oxidation (MOX) and Strong Oxidation (SOX). Each of these domains have been represented along strike and down dip by the various drilling programs targeting specific domains.

The various testwork programs have shown consistent and similar metallurgical response for all of the domains in all of the deposits tested, suggesting low sample variability and a reduced number of metallurgical samples required to define the metallurgical behaviour of the mineralised material.

Table 13.1 summarises the drillholes used for metallurgical testwork. Note that not all intervals in the ranges were employed. Typically, contiguous intervals were used to make short range composites. There is reasonable coverage along strike and down dip and spatial distribution of the samples is considered adequate for the purposes of this study.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 95 West African Resources Limited

The different shadings in Table 13.1 represent the different phases of drilling and sample supply for testing during the development of the metallurgical testwork programs.

Table 13.1 Metallurgical Testwork Hole Locations and Intervals

Drillhole Easting Northing From To Resource Domain

TanMet001 743180 1337344 6 63 M5 MOX TanMet002 743353 1337474 20 29 M5 SOX, MOX, FRS TanMet003 742999 1337097 3 66 M5 SOX, MOX, FRS TanMet004 742631 1336681 0 72 M5 SOX, MOX TanMet005 742296 1336235 12 55 M5 SOX, MOX, WOX TanMet006 742429 1336341 9 51 M5 SOX, MOX TanMet007 742513 1336520 7 35 M5 SOX TanMet008 742598 1336641 0 31 M5 SOX, MOX TanMet009 742701 1336817 7 39 M5 SOX, MOX TanMet010 742813 1336984 5 54 M5 SOX, MOX, WOX TAN15-GT-Met01 742670 1336660 58 71 M5 WOX, FRS TAN15-GT-Met02 742892 1336931 90 99 M5 FRS TAN15-GT-Met03 742870 1336961 Not used M5 NA TAN15-GT-Met04 743042 1337078 55 71 M5 MOX, FRS TAN15-GT-Met05 743205 1337329 38 53 M5 WOX, FRS TAN15-GT-Met06 742768 1337006 95 106 M5 FRS TAN15-GT-Met07 742407 1336239 135 144 M5 FRS TAN15-GT-Met08 742360 1336220 Not used M5 NA TAN15-GT-Met09 742315 1336179 116 124 M5 FRS TAN15-GT-Met10 742222 1336119 62 84 M5 WOX TAN16-Met11 741389 1337626 54 75 M1 FRS TAN16-Met12 741612 1336928 28 90 M1 FRS TAN16-Met13 741483 1337218 70 130 M1 FRS TAN16-Met14 741519 1337091 53 84 M1 MOX, FRS TAN16-Met15 742156 1336222 97 110 M5 FRS TAN16-Met16 742317 1336356 87 102 M5 FRS TAN16-DD034 741458 1337632 109 133 M1 FRS TAN16-DD044 741557 1336848 114 128 M1 FRS TAN16-DD052 741531 1337232 124 173 M1 FRS TAN16-DD055 741667 1337017 199 220 M1 FRS TAN16-RC122 741643 1336958 86 128 M1 FRS TAN16-RC131 741483 1337217 76 128 M1 FRS

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 96 West African Resources Limited

Historical Testwork Program (2012) On January 30, 2012, SGS of Vancouver, BC, Canada received a shipment of 139 core samples from Channel to perform preliminary mineralogical and metallurgical testwork. The scope of the program included the following:

. Mineralogy - to investigate the liberation, association and nature of the contained gold. . Gravity Testing - to explore the option of recovering gravity recoverable gold as a preliminary step before leaching. . Leach Testing - Including fine and coarse leaching kinetics to compare heap leaching amenability versus direct cyanidation. . Environmental Testing - which includes ABA (acid base accounting), NAG (net acid generation) and MWMP (meteoric water mobility procedure). The mineralogical examination of the two samples was carried out with X-ray diffraction (XRD), quantitative evaluation of materials by scanning electron microscopy (QEMSCANTM), electron microprobe analysis (EMPA) and chemical analysis. The XRD analysis showed that the sulphide sample consisted of quartz, plagioclase, mica, pyrite, amphibole, calcite, chlorite and trace pyrrhotite. The oxide sample consisted of quartz, plagioclase, mica and trace kaolinite and amphibole. The QEMSCANTM analysis was in agreement with the XRD analysis.

In terms of gold mineralogy, the average gold grade for the sulphide sample was 1.58g/t Au with 142 gold grains identified occurring mainly as native gold and electrum. Gold content averaged 86.8% in the 19 gold grains analysed. The gold particles ranged in size from 1µm to 60µm, with the highest percentage (30.5 weight %) of grains between 10µm and 15µm.

The oxide sample average gold grade was 1.79g/t Au with 73 gold grains identified occurring mainly as native gold. Gold content averaged 87.8% in the 14 gold grains analysed. The gold particles ranged in size from 1µm to 60µm, with the highest percentage (22.0 weight %) of grains between 2µm and 5µm.

Six cyanide bottle roll leach tests were conducted on both oxide and sulphide samples. The tests assessed the effects of primary grind, cyanide dosage and density at a leach time of 72 hours and maintained a pH range of 10.5 to 11. The oxide composite produced a final pregnant leach solution with recoveries ranging from 92.9% to 95.3%. The sulphide composite produced a final pregnant solution with recoveries ranging from 84.7% to 92.3%.

Historical Testwork Program A15684 (2014) A testwork program was developed to investigate the amenability of the M5 deposit to heap leaching processing (pre-feasibility study). The testwork was carried out at ALS Global in Perth on four metallurgical holes drilled at various locations in the M5 deposit.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 97 West African Resources Limited

In total, 1.6 tonnes of whole NQ core were shipped to ALS Global for metallurgical testwork including the following:

. Coarse feed intermittent bottle rolls (IBR) at 100% passing 6.25mm and 12.5mm crush sizes to determine suitability of oxide ore to heap leach processing.

. Direct cyanidation bottle rolls at P80 100µm of all ore types. . Size by size gold analysis on IBR residue. . Percolation rate testwork. . Agglomeration testwork. . Column cyanidation leach testwork. . Comminution testwork.

13.5.1 Samples Testwork Program A15684 Four NQ drillholes were located by the WAF geological team to represent those areas of the M5 deposit where most of the reserve and associated gold ounces were considered to reside. These holes were designated TanMet001 to TanMet004.

To retain the maximum mass of core the samples were sent to ALS complete (full core) without assay. The geological logging of the core included a grade estimate as a function of the mineralisation based on historical experience and observation. Grade ranges were limited to high, medium, low and very low (waste).

On receipt the samples were dried, weighed and crushed to nominally 19.5mm. The samples were grouped as a function of anticipated head grade ranges as well as the oxidation state of the core. Oxidation states logged were as follows:

. SOX Saprolite, being totally oxidised showing little or no primary rock texture. Complete oxidation of all primary minerals. . MOX Moderately oxidised. This material shows some primary rock texture, total oxidation of feldspar to clay, and total oxidation of sulphides. . WOX Weakly oxidised. Material showing strong primary rock textures, partial oxidation of feldspars to clay, partial oxidation of sulphides (often showing iron oxide staining). . FRS Fresh ore - no oxidation present.

13.5.2 Results Testwork Program A15684 As this program’s main focus was on heap leaching during the pre-feasibility study phase of the project, the amount of data collated relevant to the currently proposed CIL flowsheet is limited. Consequently, only select results relevant to a CIL flowsheet are presented in this section.

Contiguous sub-composites were combined to generate 10 composites representing the various metallurgical domains. Table 13.2 summarises the composites, oxidation state represented and the corresponding head assays.

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The 10 composites were subjected to various intermittent bottle roll leaching tests at two different crush sizes as well as a conventional bottle roll test at a grind size of nominally

P80 100µm.

Table 13.2 Composite Details 2014-A15684

Composite Head Assay Au g/t Oxidation MET 1A 1.17 MOX MET 1B 0.90 WOX MET 1C 1.36 FRS MET 2A 0.80 MOX MET 3A 0.92 SOX MET 3B 1.15 MOX MET 3C 0.87 WOX MET 3D 1.84 FRS MET 4A 2.78 SOX MET 4C 0.97 FRS

The conventional bottle roll tests investigated the sensitivity of the composites to grinding and agitated tank leach (akin to CIL or CIP processing), as well as any refractory behaviour.

Table 13.3 presents a summary of the leach results. Tests all show low sodium cyanide (NaCN) consumption and moderate to high quicklime (lime) consumption.

Table 13.3 Summary Leach Results 2014 -A15684

% Au Extraction Calc’d Leach Consumption Sub- Test @ hours Au Head Residue (kg/t) Comp. ID No. Grade Au 2 4 8 24 48 (g/t) (g/t) NaCN Lime

MET 1A JS3391 66.41 77.39 91.47 96.33 97.13 1.74 0.05 0.26 1.96 MET 1B JS3392 59.71 70.47 83.65 89.46 94.51 1.09 0.06 0.40 0.56 MET 1C JS3393 72.04 80.71 93.45 94.04 95.79 1.19 0.05 0.25 0.36 MET 2A JS3394 70.53 86.39 94.16 96.33 98.46 0.65 0.01 0.25 1.45 MET 3A JS3395 77.93 86.51 94.14 97.88 97.88 0.94 0.02 0.30 3.23 MET 3B JS3396 68.21 83.59 94.72 96.64 97.27 1.10 0.03 0.30 1.49 MET 3C JS3397 77.43 85.02 92.46 95.58 95.58 1.36 0.06 0.23 1.15 MET 3D JS3398 72.93 78.83 93.98 93.98 93.98 1.00 0.06 0.25 0.35 MET 4C JS3399 65.75 69.28 72.73 73.58 73.58 0.83 0.22 0.29 0.45

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 99 West African Resources Limited

Leaching was conducted for 48 hours, at 1000mg/L NaCN concentration and with oxygen sparging to maximise extraction. Whilst the conditions may be considered aggressive, the material represented by these samples is considered non-refractory and presents low NaCN consumption, suggesting that the material is amenable to conventional processing techniques utilising crush, grind and CIL/CIP methods.

The cyanide concentration used was high by industry standards and the consumption values low, suggesting that the material did not contain any notable cyanicides.

13.5.3 Abrasion Index and Ball Mill Work Index Program A15684 Abrasion indices (Ai) and Ball Mill Work Indices (BWI) were determined for the MOX and FRS samples. The results are summarised in Table 13.4.

Table 13.4 Abrasion and Ball Mill Work Index for MOX and FRS 2014-A15684

Abrasion Index Ball Mill Work Index (BWi) Sample (Ai) kWh/t (at 106 µm closing size) Moderately oxidised (MOX) 0.0071 7.9 Fresh (FRS) 0.0997 11.9

Both samples provided very low Ai and BWI values. This suggested that the ore types will be amenable to conventional comminution circuits. However, the BWi for the FRS sample was considered very low for this type of material and the value of 11.9kWh/t was considered potentially non-representative. As this program was not aligned to determining grinding circuit design, the BWi values and the issue with the FRS result were of little consequence and not investigated further.

13.5.4 A15684 Summary The A15684 program had shown that the various materials were amenable to both heap leaching and agitated leaching practices. Whilst the non or partially oxidised material did not provide high heap leach extractions, they did provide high agitated leach extractions, suggesting that high extractions at low reagent consumptions were obtainable provided that the materials were subjected to adequate liberation.

Pre-Feasibility Testwork Program A16326 Part A (2015) and Feasibility Testwork Program A16326 Part B (2016) A testwork program was developed to further investigate the amenability of the M5 material to heap leaching processing. The testwork was again carried out at the ALS Global facility in Perth and was reported in ALS report A16326 Part A and A16326 Part B.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 100 West African Resources Limited

Testwork for A16326 Part A included the following:

. Moisture determinations. . ICP determinations and head assays including screen fire assay (SFA) methods. . Unconfined compressive strength (UCS), crushing work index (CWi) determinations and abrasion index (Ai) testing. . Coarse feed intermittent bottle rolls (IBR) at 100% passing 6.25mm and 12.5mm crush sizes. One test was conducted at 2mm to 3mm crush to simulate potential high pressure grinding rolls application. . Size by size gold analysis on IBR residues. Subsequent development of the CIL/CIP options resulted in a selection of the samples used for the heap leach work subjected to additional testwork commencing mid-2016. The CIL/CIP focused work is reported in ALS report 2016 - A16326 Part B.

The sections below describing the A16326 Part A results only consider those aspects of the heap leach work relevant to the current CIL/CIP flowsheet option. Given that the A16326 samples were originally selected for heap leach work, there is a bias toward the more oxidised materials.

The A16326 Part B stage of work included the following:

. Bond Rod Mill Work Index determinations. . Bond Ball Mill Work Index determinations. . Levin test determinations (on the SOX material). . CIL optimisation testwork. . Gravity gold recovery testwork. . Bulk sample leaches to provide sample for carbon characterisation and thickening and tailings characterisations. Thickening and tailings characterisation were undertaken by others. . Oxygen uptake (demand) and slurry rheology testwork.

13.6.1 Samples Testwork Program A16326 A further 6 NQ drillholes (TanMet005 to TanMet010) were drilled during September and October 2014 to extend metallurgical sample coverage of the M5 deposit. The holes targeted fresh and weakly oxidised material along with additional oxidised material. Approximately 880kg of sample were provided to ALS.

The samples were processed utilising the same methods for TANMET001-004 and the same oxidation designations were retained. Thirty-five short range composites were generated representing all oxidation states. The composites comprised contiguous drill core intervals ranging from a few meters to several metres.

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Results of Testwork Program A16326 Part A The composites were subjected to head analyses including ICP analysis. Gold analysis was undertaken by duplicate fire assay (FA) as well as SFA. Generally, the assays were aligned, however there were some composites, particularly the higher grade composites, where the duplicate FA results (on the same pulp sample) differed from the SFA results.

High levels of visible gold had not been observed during the drilling program, yet the SFA results suggested some gold was typically reporting to the +75µm fraction.

Table 13.5 presents the various assay determinations. The percentage of gold in the +75µm fraction generally increases as the head grade increases. This would suggest some free gold is present and could influence assay repeatability.

Table 13.5 Gold Head Assay Data 2015 - A16326

Composite Au g/t Au g/t, Repeat Au g/t SFA 75µm % +75µm SOX (5-1) 1.39 1.58 1.29 10% SOX (5-2) 1.93 2.26 0.99 21% MOX (5-3) 0.95 0.91 1.24 23% MOX (5-4) 0.51 0.46 0.39 8% WOX (5-5) 8.12 8.01 30.89 40% SOX (6-1) 0.89 0.94 0.92 5% SOX (6-2) 0.95 1.08 0.94 20% MOX (6-3) 5.83 4.94 5.07 46% MOX (6-4) 2.34 2.50 3.42 13% MOX (6-5) 0.97 0.93 0.92 6% MOX (6-6) 1.43 1.48 3.46 9% SOX (7-1) 1.44 1.41 0.88 7% SOX (7-2) 0.65 0.65 0.83 5% SOX (7-3) 1.98 2.01 3.14 14% SOX (7-4) 0.47 0.50 0.42 3% SOX (8-1) 2.86 2.84 2.51 5% SOX (8-2) 3.64 3.93 4.11 6% SOX (8-3) 1.23 1.27 1.30 7% MOX (8-4) 2.35 2.58 3.54 8% MOX (8-5) 2.04 1.95 0.88 14% MOX (8-6) 0.61 0.61 0.51 4% SOX (9-1) 1.00 0.99 0.55 8% SOX (9-2) 0.47 0.45 0.56 7% SOX (9-3) 0.65 0.70 1.06 SOX (9-4) 1.12 1.04 1.15 6% MOX (9-6) 0.05 1.12 1.91 5% SOX (10-1) 1.46 1.66 2.42 6% MOX (10-2) 1.79 1.80 1.10 27% MOX (10-3) 0.66 0.76 0.65 2% MOX (10-4) 0.66 0.64 0.61 4% MOX (10-5) 1.93 1.79 1.71 12% WOX (10-6) 0.20 0.17 0.18 1% WOX (10-7) 16.20 16.10 9.72 WOX (10-8) 1.76 1.50 1.77 9% WOX (10-9) 0.54 0.52 0.27 5%

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 102 West African Resources Limited

There were low levels of total sulphur determined, and no notable contaminants or cyanicides detected. Given the good leach performance observed in the earlier testwork program, these observations align.

UCS determinations were undertaken, although most of the samples had some level of oxidation present. The one fresh UCS sample showed only moderate strength. A number of samples failed at the start of the test and were therefore very weak. Table 13.6 presents the data and detail of the sample locations.

Table 13.6 UCS Results 2015 -A16326

Oxidation Sample Description UCS (mPa) Strength Failure Mode MOX TANMET005 TRAY#11 36.49-36.58m 7.7 Weak Shear FRS TANMET005 TRAY#19 64.31-64.46m 31.4 Medium Strong Shear MOX TANMET005 TRAY#14 46.95-47.02m NA NA NA WOX TANMET005 TRAY#15 50.66-50.81m 4.2 Very Weak Shear MOX TANMET006 TRAY#8 31.94-32.19m 3.1 Very Weak Shear WOX TANMET006 TRAY#14 58.64-58.71m 4.8 Very Weak Shear MOX TANMET006 TRAY#10 42.0-42.14m 6.8 Weak Shear WOX/MOX TANMET010 TRAY#13 0 50.94-51.09m (2) 14.7 Weak Shear WOX TANMET009 TRAY#1 42.93-43.06m NA NA NA

CWi determinations were undertaken. Mean values were found to be low to moderate, signifying crushing duties will not be arduous on these materials. As would be expected, the less oxidised materials show greater resistance to crushing. There is also some variation in the data, with peak values in the order of 10kWh/t being reported. This suggests that some materials will sporadically require higher energy levels, but still easily within the realms of jaw and cone crushing applications.

Results are summarised in Table 13.7.

Table 13.7 Crushing Work Indices 2015-A16326

Impact Crushing Work Index Sample ID (kWh/t) Average Maximum Minimum Std. Dev. MOX TANMET005 (36.0-38.9m) 3.1 3.6 1.8 0.9 MOX TANMET005 (46.2-50.3m) 3.9 5.5 1.8 1.6 WOX TANMET005 (50.3-54.1m) 4.2 5.5 3.6 0.9 FRS TANMET005 (64.15-66.0m) 6.2 10.5 1.7 2.5 MOX TANMET006 (30.6-35.9m) 4.0 5.4 3.5 0.8 MOX TANMET006 (39.4-43.33m) 3.9 5.4 1.7 1.1 WOX TANMET006 (55.8-59.2m) 5.6 10.5 1.7 2.7 WOX TANMET009 (41.93-43.29m) 2.7 3.6 0.9 1.1 WOX/MOX TANMET010 (50.4-52.2m) 2.9 5.0 1.7 1.1

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Ai were determined on a number of MOX and WOX samples. Non-abrasive or moderately abrasive readings were recorded, suggesting these ore types will generally result in low wear rates for equipment such as crushers, mill media, mill liners and pumps. It should be noted the WOX sample from TANMET006 54.2 to 55.0 meters is borderline on being classified as abrasive. Table 13.8 provides a summary of the data.

Table 13.8 Abrasion Indices 2015-A16326

Oxidation Sample Description Ai Classification MOX TANMET006 32.4-33.4m 0.0137 Moderately Abrasive WOX TANMET010 49.9-51.48m 0.0303 Moderately Abrasive MOX TANMET009 41.0-43.8m 0.0021 Non Abrasive MOX TANMET006 38.0-41.0m 0.0075 Non Abrasive WOX TANMET006 54.2-55.0m 0.1983 Moderately Abrasive

13.7.1 Summary of Testwork Program A16326 Part A Testwork Program A16326 Part A provided data regarding the physical characteristics of the Sanbrado materials. UCS, CWi and Ai values were all within the ranges that could be expected for these types of materials relevant to the oxidation states.

Of note, the SFA results show that some localised regions have a high proportion of gold in the +75µm fraction. The deportment of this coarse/free gold is inconsistent and can therefore expect to provide some variability in metallurgical performance.

Feasibility Testwork Program A16326 Part B Program A16326 Part B involved continued investigations aimed at refining the understanding of how the Sanbrado materials behave as well as the generation of key criteria required for the CIL process flowsheet development. The work included the following:

. Bond Rod and Ball mill index determinations. . Levin test determinations (SOX). . Gravity/cyanide leach testing. . Bulk gravity/cyanidation testing on larger sample sizes. . Carbon characterisations testwork (carbon kinetics and equilibria loadings). . Slurry rheology testwork. . Oxygen uptake testwork.

13.8.1 Samples Utilised for Testwork Program 2016 - A16326 Part B Larger samples were needed for thickening and tailings characterisation testwork. Three master composites were generated from the A16326 Part A composites, one for each of the SOX, MOX and WOX domains (referred to as SOX TANMET MC, MOX TANMET MC, and WOX TANMET MC).

Additional work was undertaken on select composites already prepared as part of the A16326 Part A program.

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13.8.2 Results Testwork Program A16326 Part B A series of head assays including SFA were undertaken on the three master composites and are summarised in Table 13.9. The MOX and WOX master composites showed a high grade coarse gold component. The SOX master composite showed little coarse gold by comparison.

Table 13.9 Assay of Master Composites A16326

+75 µm 75 µm All Sample Calc'd Description Weight Au Weight Au1 Au2 Au3 Au4 Ave Au +75 head (g) (g/t) (g) (g/t) (g/t) (g/t) (g/t) (g/t) um (g/t) SOX TANMET MC 24.9 1.1 904 1.2 1.0 1.1 1.1 0% MOX TANMET MC 17.3 34.1 979 1.5 1.9 1.4 1.7 1.6 2.2 26% WOX TANMET MC 9.1 51.8 987 1.7 1.2 1.9 1.1 1.5 2.0 23%

RWi and BWi were determined for two MOX composites as well as the MOX master composite. The results are consistent and show a low grinding power demand for these MOX materials. The results are presented in Table 13.10.

Table 13.10 Rod and Ball Mill Work Indices 2016 - A16326

RWi BWi Sample ID (kWh/t) (kWh/t) MOX 5-3 8.5 4.8 MOX 6-5 8.9 6.3 MOX TANMET MC 8.5 7.2

Open circuit Levin ball mill grindability testwork was carried out on selected sub- composites including SOX 5-1, SOX 6-2, SOX 8-1, SOX 9-4 and the SOX TANMET Master Composite, using the method proposed by J. Levin (“A proposal for the determination of Grindability of fine materials”, Mintek Report No. M177, Council for Mineral Technology). This open circuit method provides a more robust method of determining energy demands for soft or high clay ores such as the SOX materials.

The results of this work are summarised in Table 13.11.

Table 13.11 Levin Test Results on SOX Material 2016 - A16326

Power Applied SOX 5-1 SOX 6-2 SOX 8-1 SOX 9-4 SOX TANMET MC kWh/t P80 µm P80 µm P80 µm P80 µm P80 µm 0 1056 1605 302 630 786 2 173 67 137 156 3 142 121 86 4 85 81 5 79 129 51 69 66 6 114 7 90 8 73

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The Levin tests show there is some variation in the composite performance. Composite

SOX 6-2 reported a value of 7kWh/t to achieve a P80 of 90µm whereas the other samples, including the master composite, require values of around 3kWh/t to 4kWh/t to achieve similar grinds. It is noted that SOX 6-2 is bordering the MOX zone in hole TANMET 006 and so it is quite probable that the competence of this sample is greater as a consequence. It is also significantly coarser.

A number of agitated cyanide leach tests were undertaken at various grind sizes to assess grind sensitivity of the master composites. A gravity/amalgamation step was undertaken at the reported grind size and the gravity and amalgam tails were subsequently leached for 48 hours. Cyanide concentration was 500mg/L (initial dose controlled to a minimum of 250mg/L as NaCN was utilised) and oxygen for leaching was provided by air sparging.

Results are summarised in Table 13.12.

Table 13.12 Master Composite Grind Sensitivity 2016 - A16326

% Au Extraction Consumption Calc'd Leach Composite P80 Grind @ hours (kg/t) Au Head Residue ID Size (µm) (g/t) (g/t) Grav. 2 12 24 48 NaCN Lime 125 9.13 67.05 85.59 91.53 92.98 1.42 0.10 0.22 1.30 SOX TANMET 106 16.80 75.08 91.46 93.43 94.40 1.43 0.08 0.22 1.27 Master Comp 75 12.92 76.62 91.53 91.53 93.23 1.62 0.11 0.18 1.48 125 31.47 71.79 85.71 88.59 92.34 2.94 0.23 0.18 0.86 MOX TANMET 106 38.67 78.89 94.02 94.02 94.53 2.74 0.15 0.22 0.82 Master Comp 75 38.97 78.05 91.99 93.91 95.47 2.21 0.10 0.22 0.82 WOX TANMET 106 46.89 67.58 90.99 90.99 93.66 1.81 0.12 0.18 0.77 Master Comp 75 61.36 77.00 93.31 95.76 96.00 2.88 0.12 0.18 0.80

In line with the SFA results reported for the SOX master composites, the quantity of gravity gold (coarse gold) was found to be low.

As the grind size was decreased, generally more gold was liberated and collected by the amalgamation process. This indicates that the free gold present is fine grained.

Leach kinetics improve as the grind size is reduced. However, ultimate extractions after 48 hour are similar. The sensitivity to grind is not strong, but there is improvement in

extraction when grinding down to the P80 of 75µm.

The cyanide consumptions are relatively low whilst quicklime consumptions are moderate for oxide or partially oxidised material.

The results show the SOX, MOX and WOX materials are all amenable to agitated cyanidation with acceptable extractions being achieved even at the coarser grind sizes.

13.8.3 Bulk Tests for Testwork Program A16326 Part B To evaluate gravity recovery performance as a function of yield as well as to prepare bulk samples for tailings and thickener testwork, 20kg samples of the SOX and MOX master

composites were prepared at a P80 of 75µm. This grind was selected to present a “worst case” scenario for thickening and tailings characterisation and to provide contingency should the grind size be reduced.

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Multiple passes of the prepared samples were made through a Knelson concentrator and the concentrates subjected to amalgamation. Amalgam tails were then subjected to a small scale intensive leach to estimate potential gold extraction should an intensive leach reactor be employed as part of the flowsheet.

The gravity tails were leached using 250mg/L NaCN for 48 hours with oxygen sparging.

Low residue grades of 0.11g/t Au were achieved for both samples at low cyanide consumption, with total extractions in the region of 94%. Gravity recovery was reported as 14% and 22% for the SOX and MOX composites respectively. Total gold extraction from intensive leach of the gravity concentrate resulted in a total extraction via gravity amalgam/gravity leach of 25% and 39% for SOX and MOX respectively. Results are summarised in Table 13.13.

Table 13.13 Master Composite Bulk Gravity-Leach 2016 - Test A16326

% Au Extraction Consumption Calc'd Leach Composite @ hours (kg/t) Au Head Residue ID (g/t) (g/t) Total Gravity 2 12 24 48 NaCN Lime SOX TANMET 25.22 73.43 89.60 90.45 94.12 1.81 0.11 0.24 1.52 Master Comp MOX TANMET 39.16 82.87 90.98 93.09 94.82 2.13 0.11 0.25 0.83 Master Comp

The influence of yield (mass pull to concentrate) on amalgam (free) gold recovery for the SOX and MOX master composites are presented in Figure 13.1 and Figure 13.2 respectively.

Typically for a process facility, the mass pull to concentrate would be in the 0.03% range. Extrapolating the graphs in Figures 13.1 and 13.2 suggest free gold recovery of less than 5% for SOX material and possibly less than 10% for MOX material in a full-scale operation. Such an extrapolation may be considered crude, however the process facility has the ability to cater for higher or lower gravity recoveries due to the inherent behaviour of carbon based flowsheet applications. As such, the key function of a gravity circuit is to recover the slow leaching coarser gold particles to minimise losses post leaching.

These plots consider the free gold, or gold that reports to amalgam. If the full-scale flowsheet includes intensive leaching of the gravity concentrate, then the free gold plus the soluble gold in the concentrate has to be considered for the circuit design. Additional discussion is presented regarding the influence of this soluble gold in Section 13.16.2 which further discusses gravity recovery aspects.

Bulk leaching tests were also undertaken on the SOX and MOX master composites to prepare

sample for carbon characterisation testwork. 10kg samples were ground to a P80 of 75µm and leached with 250mg/L NaCN using oxygen sparging for 24 hours without gravity recovery to provide higher grade leach solutions. The leach results compared very well with similar bulk leach test results that included gravity recovery. The similar leach extractions with and without gravity recovery suggest that the free gold present is fine and that the bulk of this gold leaches within a 24 hour period. Results are summarised in Table 13.14.

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Figure 13.1 SOX Master Composite Gravity Performance

SOX TANMET MC Gravity Performance - Free Gold 30.00%

25.00%

20.00%

15.00%

y = 0.0283ln(x) + 0.2238 R² = 0.9897 10.00% %Recovery of gold toamalgam

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

Figure 13.2 MOX Master Composite Gravity Performance

MOX TANMET MC Gravity Performance - Free Gold 30.00%

25.00%

20.00% y = 0.0195ln(x) + 0.2359 R² = 0.9827

15.00%

10.00% % Recovery Recovery % of gold toamalgam

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

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Table 13.14 Master Composite Bulk Gravity-Leach Test 2016 - A16326

% Au Extraction Consumption Calc'd Leach Composite @ hours (kg/t) Au Head Residue ID (g/t) (g/t) 2 6 12 24 48 NaCN Lime SOX TANMET 65.54 85.67 91.30 92.12 92.91 1.86 0.13 0.21 1.77 Master Comp MOX TANMET 47.56 85.51 92.66 94.11 94.81 2.09 0.11 0.18 0.97 Master Comp

Carbon kinetic and equilibrium testing were carried out on the bulk leach products. The Fleming gold adsorption constants are summarised in Table 13.15 and the equilibrium data in Table 13.16.

Table 13.15 Fleming Kinetic Parameter Determination 2016 - A16326

Fleming Gold Composite ID Adsorption Constants k (hr-1) n SOX TANMET Master Comp 182.25 0.738 MOX TANMET Master Comp 246.05 0.678

Table 13.16 Carbon Equilibrium Parameter Determination 2016 - A16326

Equilibrium Carbon Loading Au (g/t) Composite ID @ Solution Concentration 1.0mg/L 0.50mg/L 0.20mg/L SOX TANMET Master Comp 7,623 5,885 4,180 MOX TANMET Master Comp 27,901 18,023 10,114

The kinetic data present typical values, indicating that typical CIL/CIP design considerations are applicable with regard to kinetics. The equilibrium values, however are high, particularly for the MOX master composite. This work will need to be repeated as part of detailed design when the final blend strategy has been adopted for the plant. In the interim, lower values will be applied in line with typical industry values due to the adopted blending strategy of 50% SOX/MOX and 50% FRS/WOX. Loadings are assumed to be lower than those presented by the testwork, resulting in higher elution rates and therefore a larger elution plant than may be necessary.

Samples of the SOX and MOX Tanmet Master Composites were subjected to oxygen uptake determinations and viscosity testing at various pulp densities and shear rates. The data is presented in Table 13.17 and Table 13.18.

The results show that these samples have very low oxygen demands and that oxygen inputs to maintain elevated levels will be low. This data was used to estimate the oxygen demands and associated operating costs.

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Table 13.17 Oxygen Uptake Rate Determination 2016 - A16326

Sample ID

Time SOX TANMET Master Comp MOX TANMET Master Comp (hours) Oxygen Uptake Rate* Oxygen Uptake Rate* (mg/L/min) (mg/L/min) 0** -0.0024 -0.0021 1 -0.0019 -0.0016 2 -0.0010 -0.0004 3 -0.0014 -0.0015 4 -0.0008 -0.0015 5 0.0000 -0.0008 6 0.0000 -0.0004 24 -0.0009 -0.0008

Table 13.18 Master Composite Viscosity Data 2016 - A16326

Bohlin Visco 88 Viscosity @ Shear Rate (sec-1 ) (cps) Composite ID Pulp Density P80:90µm) % Solids (w/w) 4.2 7.4 13.1 21.9 38.9 67.4 119.2 209.5 60 5049 3124 1986 1333 860 574 381 322 SOX TANMET 50 599 363 220 144 93 80 89 106 40 0 0 0 0 0 33 45 64 60 882 668 450 338 234 169 184 193 MOX TANMET 50 0 0 0 0 0 38 54 79 40 0 0 0 0 0 0 34 58

The viscosity testwork presents low to moderate viscosities for those parameters relevant to pumping and agitation. The more oxidised SOX material presents a shear thinning slurry as does the MOX, however, the SOX material viscosity is higher. This is expected, given that the SOX material contains more clay and also requires higher doses of quicklime to achieve elevated pH values. None of the values recorded are considered problematical for process design or plant operation.

A number of composites were subjected to testwork designed to represent the flowsheet proposed for the economic treatment of the Sanbrado materials. This work included a gravity recovery step followed by kinetic leaching using the conditions summarised as follows:

. Primary grind size: P80:90µm . Initial NaCN Dosages: 0.030% (w/v) . Maintained NaCN Levels: 0.025% (w/v) . Sparging Gas: Oxygen . Leach Duration: 48h

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Table 13.19 presents the results of the testwork for SOX samples originating from 5 different drillholes, and for MOX samples originating from 3 different drillholes. Two MOX composites originate from hole TANMET 6 due to the high number of MOX intercepts in this hole.

Table 13.19 Master Composite Variability Testing 2016 - A16326

% Au Extraction Consumption Calc'd Leach Composite @ hours (kg/t) Au Head Residue (P80:90µm) (g/t) (g/t) Grav. 2 12 24 48 NaCN Lime

SOX 5-1 18.94 84.97 95.45 92.57 95.39 0.98 0.05 0.07 1.73 SOX 6-2 11.07 83.69 93.82 91.87 93.78 0.72 0.05 0.07 2.30 SOX 7-3 7.45 60.90 70.99 73.04 74.55 2.75 0.70 0.07 0.94 SOX 8-1 8.17 74.95 90.50 90.50 91.63 2.45 0.21 0.10 2.78 SOX 10-1 28.91 90.13 95.62 95.62 96.09 2.94 0.12 0.10 3.23 MOX 5-3 33.31 78.07 88.23 88.72 90.63 1.44 0.14 0.07 0.77 MOX 6-4 26.12 82.74 93.19 94.34 94.90 2.45 0.13 0.10 0.66 MOX 6-6 18.33 78.68 89.82 89.46 89.82 3.93 0.40 0.14 0.86 MOX 10-2 35.45 86.31 94.22 95.85 96.92 1.30 0.04 0.07 2.95

The results show that the leaching is almost complete in 18 hours. Gravity recoveries are moderate, with the SOX material gravity component typically lower than the MOX material, as had been noted previously.

Of note, SOX 7-3 and MOX 6-6 samples presented high grade leach residues, particularly SOX 7-3. Leaching kinetics for these two samples were very slow suggesting that these high grade residues are probably due to gold lock-up and/or poor gold liberation. This is supported by the fact that neither sample showed a high proportion of +75µm gold in the SFA results (Table 13.5). In addition, the lack of slow leaching kinetics specifically post 18 hours also suggests that the high residue grades are not due to coarse gold. NaCN levels were considered adequate as were oxygen levels, suggesting the high grade residues are not due to inadequate leach conditions.

Additional discussion on these aspects as relevant to the testwork outcomes and leach extraction expectations are presented in Section 13.14.

13.8.4 Summary of Testwork Program A16326 Part B Program A16326 Part B confirms that the SOX and MOX materials are amenable to conventional gravity and cyanide leaching processes. Oxygen uptake rates are low (low oxygen demand), slurry rheology is not problematical and reagent demands are low. Comminution characteristics for the MOX material indicate that energy demands will be low.

One SOX variability test presented a high leach residue grade, suggesting that lock up of gold due to associated minerals may be possible at times even in oxidised samples.

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Feasibility Testwork Program A16557 (2016/17) ALS program A16557 was developed to provide the bulk of the data necessary to support the proposed flowsheet. As stated previously, earlier programs had provided most data on the SOX and MOX materials. Program A16557 focussed on the more competent and less oxidised WOX and FRS materials. The testwork program covered the following:

. Compositing and head assaying. . Bond Rod and Ball Mill Work Index determinations. . UCS and Bond Crushing Work Index determinations. . Gravity/cyanidation leaching on various composites including variability samples. . Pulp density and CIL variability testwork. . Oxygenation sensitivity testwork. . Mercury investigation (assessment of presence of soluble mercury). . Rheology and oxygen uptake determinations. . Bulk sample gravity and leaching to provide samples for testing by others as well as understand gravity/yield relationship for WOX and FRS. . Submission of bulk samples to others for associated testwork (thickening and tailings characterisation).

13.9.1 Samples Utilised for Testwork Program A16557 The program utilised samples from the TAN15-GT-Met01 to TAN15-GT-Met10 series of holes (with the exclusion of TAN15-GT-Met03) from the M5 deposit.

Intervals were selected with the assistance of ALS personnel for UCS and CWi testing which was completed first. These intervals post testing were utilised in the make-up of other composites including those used for Bond RWi and Bond BWi tests.

Six composites were made up for work index testing and a further 19 were made up to spatially represent WOX and FRS material types throughout the deposit. One composite of MOX material was also collected.

A WOX and a separate FRS master composite were made up for general scoping testing as well as to generate sample for larger scale tests. A second FRS master composite was made up later in the program as the original FRS master composite was depleted.

13.9.2 Results of Testwork Program A16557 Table 13.20 presents a summary of the UCS data compiled for various oxidation types in the TAN15-GT series of drillholes. The FRS material shows strong competence as had previously been observed for this material. Some FRS material also presented as low values and this is a consequence of failure along planes observable in the core, at times associated with changes in mineralisation.

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Table 13.20 UCS Testwork Summary 2016/17 A16557

Interval UCS Failure Sample ID Strength Description (m) (MPa) Mode TAN15-GT-MET01 (WOX) 59.0-60.0 6.46 Shear Weak TAN15-GT-MET01 (WOX) 63.95-65 7.65 Shear Weak TAN15-GT-MET02 (FRS) 90.35-91.12 49.1 Shear Med. Strong TAN15-GT-MET02 (FRS) 97.4-98.4 12.7 Shear Weak TAN15-GT-MET04 (MOX) 50.2-51.2 6.46 Shear Weak TAN15-GT-MET04 (FRS) 64.7-65.75 72.2 Shear Strong TAN15-GT-MET05 (FRS) 41-42 32.1 Shear Med. Strong TAN15-GT-MET05 (FRS) 53-54 141.2 Shear Strong TAN15-GT-MET06 (FRS) 103-104 51.5 Shear Med. Strong TAN15-GT-MET06 (FRS) 114.3-115.1 19.5 Shear Weak TAN15-GT-MET07 (FRS) 131.8-132.8 45.5 Shear Med. Strong TAN15-GT-MET07 (FRS) 140.0-141.0 24.3 Shear Med. Strong TAN15-GT-MET09 (FRS) 118.0-118.8 66.0 Shear Strong TAN15-GT-MET09 (FRS) 125.0-126.0 94.4 Shear Strong TAN15-GT-MET10 (WOX) 63.0-64.0 10.6 Shear Weak TAN15-GT-MET10 (WOX) 69.7-71.0 63.2 Shear Strong TAN15-GT-MET10 (WOX) 86.0-87.0 21.1 Shear Med. Strong

The WOX material presented lower UCS values as anticipated. The WOX material results seems to be quite variable with values generally between those obtained for FRS and MOX material.

The results suggest that the WOX and FRS material is amenable to conventional crushing. The low WOX UCS values suggest some consideration should be given to the final design of chutes, hoppers and similar equipment to mitigate materials handling issues.

Table 13.21 presents the results of the CWi determinations on a selection of intervals from the TAN15-GT drillhole series. The values are generally aligned with the UCS data. TAN15- GT-MET09 (118m to 118.8m) and TAN15-GT-MET10 (86m to 87m) being two notable exceptions. It should be noted that the UCS test can present differing values depending on the structure of the rock and is sensitive to planes of weakness and changes in mineralogy within the ore. The CWi test can be sensitive to similar features within the ore body. However, if the more competent rock clusters are of the same range of size as the CWi samples, it is generally common to obtain results for CWI much higher than might be suggested by the more macro scale UCS tests.

The CWi values recorded again suggest conventional crushing and impact breakage will be applicable to process these materials.

Table 13.22 presents the Ai test results. The values obtained show a mix of abrasion characteristics, with WOX materials ranging from non-abrasive to moderately high abrasion. The FRS material typically shows moderate to high abrasion characteristics. Abrasion indices are more of a function of mineralogy, such as quartz content, and so a material high in quartz may show higher abrasion than another sample, but have lower UCS or CWi values.

Table 13.21

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Crushing Work Index Summary 2016/17 A16557

Bond Impact Crushing Work Index (kWh/t) Sample ID Standard Average Maximum Minimum Deviation TAN15-GT-MET01 (63.95-65m) WOX 2.99 3.5 1.74 0.85 TAN15-GT-MET02 (97.4-98.4m) FRS 4.21 10.4 1.72 2.61 TAN15-GT-MET04 (64.7-65.75m) FRS 10.01 15.5 5.17 3.02 TAN15-GT-MET05 (53-54m) FRS 12.58 18.6 5.08 4.46 TAN15-GT-MET09 (118-118.8m) FRS 7.55 13.7 3.43 3.63 TAN15-GT-MET10 (63-64m) WOX 5.99 10.6 1.78 2.84 TAN15-GT-MET10 (69.7-71m) WOX 8.82 15.6 1.73 4.27 TAN15-GT-MET10 (86-87m) WOX 10.67 14.2 7.11 2.31

Table 13.22 Bond Abrasion Test Summary 2016/17 A16557

Drillhole Interval Bond Abrasion Index Composite ID (m) (Ai) TAN15-GT-MET01 (m) WOX 63.9-65.0 0.0330 TAN15-GT-MET02 (m) FRS 97.4-98.4 0.2987 TAN15-GT-MET04 (m) FRS 64.7-65.7 0.2173 TAN15-GT-MET05 (m) FRS 53.0-54.0 0.4415 TAN15-GT-MET09 (m) FRS 118.0-118.8 0.3518 TAN15-GT-MET10 (m) WOX 63.0-64.0 0.1197 TAN15-GT-MET10 (m) WOX 69.7-71.0 0.2161 TAN15-GT-MET10 (m) WOX 86.0-87.0 0.3021

The UCS, CWi and Ai values suggest that the ores cannot be categorised as having a tight range of characteristics based on oxidation state alone. A general range is obtained, however outliers are common.

Table 13.23 presents the Bond Rod and BWi test results for the various comminution composites. The WOX BWi values are generally lower than the FRS values. Comparing these values to the MOX results in Table 13.10 and the SOX Levin results in Table 13.11, a progression of increasing work index with lower oxidation levels is observed. Similarly, the earlier unexpectedly low BWi value for the first FRS sample tested (result presented in Table 13.4) would in fact seem to be an error.

BWi values in excess of 20kWh/t usually suggest some FRS material will be energy intensive, and that attempting higher gold leach extractions with a finer grind will be costly.

The RWi values present as moderate to high, again suggesting high energy intensity at the coarser end of the milling spectrum.

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Table 13.23 Bond Rod and Ball Mill Work Index Testwork Summary 2016/17 A16557

Bond Work Indices Composite ID (kWh/t) RWi *BWi M1 WOX WI 1 - 15.9 M2 FRS WI 2 - 25.0 M4 FRS WI 3 - 17.0 M5 FRS WI 4 16.4 11.3 M9 FRS WI 5 - 22.1 M10 WOX WI 6 - 16.6 M6 FRS 13 19.8 16.5 M7 FRS 14 21.6 20.2 M10 WOX 18 19.1 16.2

Table 13.24 presents the head assay data for the various composites. Silver assays are generally low except for sample M1WOX1. Copper (cyanicide) levels are low as are the organic carbon and sulphide sulphur assays.

Table 13.24 Composite Head Assay Data 2016/17 A16557

Au (SFA) Ag Cu C S Composite ID organic Sulphur (g/t) (g/t) (g/t) (%) (%) M4 MOX 6 0.72 <0.3 30 0.03 <0.02 WOX MC 1.68 0.6 35 <0.03 0.26 FRS MC 2.66 <0.3 75 0.03 0.62 M2 FRS 4 1.72 <0.3 45 <0.03 0.52 M7 FRS 14 3.45 1.2 45 <0.03 0.38 M10 WOX 18 3.57 <0.3 20 <0.03 0.44 M1 WOX 1 1.03 1.8 60 0.03 0.22 M2 FRS 5 4.50 0.6 35 <0.03 0.44 M4 FRS 7 1.79 <0.3 40 <0.03 0.48 M5 WOX 8 3.19 0.6 60 <0.03 0.84 M5 FRS 9 1.81 0.3 105 <0.03 0.98 M5 FRS 10 2.36 0.6 95 <0.03 0.68 M5 FRS 11 1.57 <0.3 85 <0.03 0.60 M6 FRS 12 0.87 <0.3 70 0.06 0.78 M6 FRS 13 1.35 <0.3 195 <0.03 0.92 M9 FRS 15 5.98 <0.3 25 <0.03 0.28 M10 WOX 16 3.27 <0.3 20 0.03 0.10 M10 WOX 17 1.28 <0.3 20 <0.03 0.38 M10 WOX 19 1.48 <0.3 15 <0.03 0.28

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Grind sensitivity testwork undertaken on a selection of the composites is presented in Table 13.25. Tests were conducted at 40% solids w/w, leached for 24 hours in total with an initial NaCN concentration of 300mg/L with oxygen to give a DO of <20ppm.

Table 13.25 Grind Size Optimisation Testwork 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) P Grind Size Composite ID 80 (µm) Gravity Overall Calc. Head Tail NaCN Lime 150 19.28 83.93 0.93 0.15 0.24 1.32 125 17.22 84.05 0.78 0.13 0.25 1.30 M4 MOX 6 106 23.96 86.82 0.83 0.11 0.25 1.36 75 21.46 89.58 0.82 0.09 0.22 1.25 150 19.46 83.68 1.59 0.26 0.29 0.39 125 31.04 89.09 1.47 0.16 0.25 0.39 WOX MC 106 21.22 89.64 1.98 0.21 0.25 0.39 75 24.32 91.50 1.71 0.15 0.25 0.39 150 15.77 76.22 2.00 0.48 0.29 0.31 125 24.00 82.93 2.17 0.37 0.29 0.31 FRS MC 106 35.63 85.89 2.83 0.40 0.29 0.31 75 31.84 88.03 2.34 0.28 0.25 0.31 150 3.29 71.89 1.37 0.39 0.24 0.33 125 2.92 71.13 1.54 0.45 0.29 0.34 M2 FRS 4 106 4.69 77.90 1.49 0.33 0.31 0.35 75 0.19 80.97 1.34 0.26 0.22 0.35 150 12.66 82.09 2.76 0.50 0.22 0.34 125 17.55 81.48 2.59 0.48 0.17 0.27 M7 FRS 14 106 19.15 86.32 2.92 0.40 0.21 0.25 75 20.60 88.23 2.72 0.32 0.22 0.27 150 40.90 92.34 2.87 0.22 0.29 0.32 125 41.86 94.66 2.81 0.15 0.22 0.32 M10 WOX 18 106 41.93 97.05 3.22 0.10 0.25 0.32 75 34.74 95.57 2.26 0.10 0.22 0.32

The results reveal the anticipated trend of decreasing residue grade as the grind becomes finer. However, some samples show minimal grind sensitivity such as M10 WOX 18 in comparison with other samples such as M2 FRS 4. Evaluation of the grind sensitivity is made in Section 13.16.

The FRS samples present a trend of higher residue grades compared to the WOX samples in line with previous data.

Gravity gold recovery varies from as low as a few percent to values in excess of 40% revealing a large variance in the gravity/free gold component of the ore body. There is

some suggestion that some free gold is still being liberated even at the P80 75µm range.

Reagent consumption is relatively low and consistent across all samples.

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Table 13.26 presents the results from a series of tests where the NaCN concentrations were varied. Initial cyanide concentrations of 250mg/L, 500mg/L and 750mg/L NaCN were trialled with constant 250mg/L NaCN concentrations maintained during leaching. A 48 hour leach period was applied and oxygenation was also undertaken controlled to nominally 20ppm DO.

Table 13.26 NaCN Sensitivity Testwork 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Initial NaCN Composite ID % Gravity Overall Calc. Head Tail NaCN Lime 0.025 42.47 93.44 2.44 0.16 0.14 0.58 WOX MC 0.050 22.70 91.99 1.50 0.12 0.16 0.71 0.075 35.83 93.59 2.11 0.14 0.28 0.45 0.025 21.33 88.00 1.88 0.23 0.14 0.42 FRS MC 0.050 11.88 87.08 1.94 0.25 0.26 0.33 0.075 16.78 88.56 1.97 0.23 0.35 0.36 0.025 30.57 91.27 0.92 0.08 0.12 1.65 M4 MOX 6 0.050 35.49 91.13 1.01 0.09 0.22 1.63

The elevated NaCN levels presented more rapid leach kinetics. However, the final leach extractions after 14 hours and 48 hours were similar. It could be argued there might be a slight benefit in elevated NaCN levels, but it is far from conclusive that these samples are sensitive to NaCN levels. It should also be noted that NaCN consumption was elevated with elevated initial NaCN dosages, which suggests any minor potential benefit could be lost with increased operating costs in a full-scale operation.

Two tests, one on the WOX MC and one on the FRS MC were undertaken at 50% solids w/w pulp density. The outcome of this work is summarised in Table 13.27. No difference was observable compared to similar tests at 40% solids.

Table 13.27 Influence of High Pulp Density 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Slurry % Composite ID Solids (w/w) Gravity Overall Calc. Head Tail NaCN Lime WOX MC 50 32.91 94.43 1.98 0.11 0.14 0.49 FRS MC 50 19.92 85.89 1.81 0.26 0.13 0.23

The WOX and FRS master composites were subjected to leaching under CIL conditions. The results are summarised in Table 13.28. The only observable difference was an increase in NaCN consumption. This is often observed in such testwork as the carbon is fresh and exacerbates the NaCN consumption and in fact overstates it. There was no apparent advantage or disadvantage in using CIL apart from the NaCN observation.

Use of oxygen versus air was evaluated for different NaCN conditions. The initial dose was 250mg/L NaCN in each test. For two of the air sparged tests, the concentration was retained at 250mg/L, but maintained at 150mg/L for the other tests. The results are summarised in Table 13.29.

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Table 13.28 Influence of CIL Leaching 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Carbon Composite ID Addition (g/L) Gravity Overall Calc. Head Tail NaCN Lime WOX MC 15 22.74 93.43 1.67 0.13 0.61 0.42 FRS MC 15 23.54 89.81 2.23 0.27 0.62 0.29

Table 13.29 Influence of Oxygenation versus Air 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Composite Sparging Maintenance ID Gas NaCN (%) Gravity Overall Calc. Head Tail NaCN Lime Air 0.025 26.41 89.67 1.74 0.18 0.15 0.50 WOX MC Air 0.015 29.39 91.65 1.50 0.13 0.15 0.54 Oxygen 0.015 29.39 91.65 1.50 0.13 0.15 0.54 Air 0.025 40.51 89.45 3.22 0.34 0.19 0.31 FRS MC Air 0.015 18.83 86.22 2.18 0.30 0.15 0.43 Oxygen 0.015 10.41 85.54 1.73 0.25 0.11 0.41

The tests suggest that over a 48 hour period, there may be some advantage using oxygen for the FRS material, including potential NaCN savings. The MOX sample outcomes are not consistent compared to the FRS results. Due to the variability in the calculated head grades, it is difficult to draw any clear conclusions regarding the benefit of oxygen for either FRS or MOX. These points are further discussed in Section 13.16.

The WOX and FRS master composites were leached (whole ore leach, no gravity) and the results compared to the conventional gravity, amalgam, gravity tail and amalgam tail leach. The final solutions were sampled for mercury to assess if mercury may be a potential issue.

The result are summarised in Table 13.30. The results suggest that for the non-gravity tests mercury levels reporting to solution were very low and inconsequential. Even when amalgamation was used, the WOX MC test had a mercury level just on detection whilst the FRS MC remained below detection. These results suggest mercury should not be an issue when processing such materials.

Table 13.30 Influence of Oxygenation versus Air 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Final Solution Composite ID Gravity Overall Calc. Head Tail NaCN Lime Hg (mg/L) N/A 77.69 2.15 0.48 0.08 0.79 <0.002 WOX MC 47.53 90.24 1.95 0.19 0.08 0.60 0.002 N/A 85.34 2.05 0.30 0.08 0.54 <0.002 FRS MC 30.96 81.82 2.06 0.38 0.11 0.33 <0.002

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Work was undertaken to assess leaching of gravity concentrates at a fine grind of nominally 10µm and a high NaCN concentration of 1000mg/L. This was to simulate a flowsheet where a high mass pull of concentrate was recorded and then subjected to fine grinding to maximise leach extraction. The test also assesses the potential refractory nature of the sulphide mineralisation. Table 13.31 presents the results of this work.

Table 13.31 Intensive Leaching of Gravity Concentrate 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Composite ID Gravity Overall Calc. Head Tail NaCN Lime WOX MC – Gravity Conc. 74.79 99.67 30.63 0.10 2.11 3.69 FRS MC - Gravity Conc. 58.45 99.56 27.36 0.12 4.54 3.87 WOX MC 37.20 89.12 1.70 0.19 0.17 0.53 FRS MC 31.45 84.17 2.18 0.35 0.18 0.44

Gold extractions of over 99% on nominal 30g/t head grades were recorded for both concentrates. The NaCN consumptions were substantially lower than anticipated, at 2.1kg/t and 4.5kg/t for the WOX material and FRS material respectively. Gold residue grades of the concentrate leaches were 0.10g/t and 0.12g/t for the WOX material and FRS material respectively, suggesting that the gold in the residues is not recovered as a function of gold liberation size, thus the ores are not refractory. The tests show that there is an alternative flowsheet option which can be expected to produce improved gold leach extractions, albeit at elevated capital cost and operating cost.

The variability samples were subjected to the selected nominal proposed flowsheet of

gravity followed by a 24 hour leach, grind P80 of 90µm, 250mg/L NaCN (controlled at a minimum of 150mg/L) with oxygen addition. The results are presented in Table 13.32.

The results in general show high gold extractions, variable gravity recoveries and low to very low reagent consumption, however samples M2 FRS 5, M5 WOX 8, M5 FRS 9 and M5 FRS 10 all show elevated gold residue grades. Lithologies, mineral assemblages and accessories were not noted as being peculiar when compared to the other samples generated. M2 FRS 5 had a sulphide assay of 0.44%. M5 WOX 8, M5 FRS 9 and M5 FRS 10 had elevated levels of sulphide of 0.84%, 0.98% and 0.68%, respectively. These levels, however, are not as high as presented by other samples which gave high gold extractions (M6 FRS 12 and M6 FRS 15 had sulphide levels of 0.78% and 0.92% respectively, but produced high gold extractions).

The more problematical samples were subjected to a 1,000mg/L NaCN re-leach of the residues for 48 hours with oxygen sparging. The results are presented in Table 13.33.

The M2 FRS 5 sample did not release much more gold, however, the M5 based composites gave around 23% to 29% additional gold extraction with final residue grades in the range of 0.2g/t to 0.27g/t. Total leach extractions were therefore in the range of 88% to 93% for these M5 based composites.

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Table 13.32 Variability Testing 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Composite ID Calculated Assay Gravity Overall Tail NaCN Lime Head Head M1 WOX 1 18.67 85.90 0.21 1.26 1.03 0.10 0.96 M2 FRS 5 5.11 77.39 1.22 4.50 4.50 0.11 0.54 M2 FRS 5 9.68 80.85 1.01 4.68 4.50 0.39 0.33 M4 FRS 7 26.32 89.32 0.27 2.13 1.79 0.14 0.45 M5 WOX 8 24.11 91.20 0.30 2.86 3.19 0.18 0.70 M5 FRS 9 30.73 87.03 0.33 2.16 1.81 0.16 0.62 M5 FRS 9 17.12 88.06 0.23 1.78 1.81 0.40 0.53 M5 FRS 10 23.30 84.72 0.44 2.36 2.36 0.18 0.50 M5 FRS 10 24.97 89.02 0.38 3.12 2.36 0.41 0.32 M5 FRS 11 11.22 84.81 0.25 1.38 1.57 0.17 0.54 M6 FRS 12 9.90 85.20 0.15 0.86 0.87 0.11 0.31 M6 FRS 13 6.42 84.58 0.22 1.17 1.35 0.18 0.27 M9 FRS 15 57.74 96.59 0.26 6.38 5.98 0.11 0.31 M10 WOX 16 52.26 94.02 0.22 3.01 3.27 0.15 0.52 M10 WOX 17 26.69 93.16 0.10 1.18 1.28 0.11 0.32 M10 WOX 19 23.87 93.67 0.11 1.47 1.48 0.11 0.42 WOX MC 32.10 93.83 0.14 1.92 1.68 0.15 0.49 FRS MC 25.74 86.95 0.36 2.33 2.66 0.15 0.39 M1 WOX 1 18.67 85.90 0.21 1.26 1.03 0.10 0.96

Table 13.33 High Grade Residue Re-Leach Testwork 2016/17 A16557

Au Grade (g/t) Consumption (kg/t) Overall Au Composite ID Extraction % Calc. Head Tail NaCN Lime M2 FRS 5 7.83 0.98 0.90 1.60 0.32 M5 WOX 8 22.99 0.27 0.21 1.66 0.52 M5 FRS 9 23.43 0.26 0.20 1.58 0.84 M5 FRS 10 28.78 0.37 0.27 1.25 1.28

The leach residues from the variability testing summarised in Table 13.32 were subjected to size by size analysis. The data relevant to those samples having high head grade and that were re-leached (Table 13.33) have been summarised by two graphs presented in Figure 13.3 and Figure 13.4.

The first graph shows the assay by size fraction. The various samples are compared to the size by size data of the WOX MC and FRS MX. The distributions are generally aligned with the master composite data except for sample M2 FRS 5 which has very high grades in the coarse fractions and M5 FRS 10 which has elevated fine fraction grades.

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Figure 13.3 Residue Grade by Size Fraction – Grade 2016/17 A16557

Residual Gold - Grade 2.5

1.5

1 Assay Assay by Fraction,g/t Au 0.5

0 0 20 40 60 80 100 120 140 Size fraction, um

M2 FRS 5 M5 WOX 8 M5 FRS 9 M5 FRS 10 WOX MC FRS MC

Figure 13.4 Residue Grade by Size Fraction – Percentage 2016/17 A16557

Residual Gold - % 35.0

30.0

25.0

20.0

15.0 Residual Gold % Gold Residual

10.0

5.0

0.0 0 20 40 60 80 100 120 140 Size fraction, um

M2 FRS 5 M5 WOX 8 M5 FRS 9 M5 FRS 10 WOX MC FRS MC

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When the data is compared as a percentage of the total residual gold per size fraction (Figure 13.4), the different size distributions come into play. As the mass of -38µm is high, the percentage of gold in this fraction is elevated even at low grade. M5 FRS 10 has an atypical mass of fine gold losses due to the elevated assay grade. It has the lowest residual gold in the coarse fractions. Interestingly M2 FRS 5 has comparable gold losses in most fractions on a percentage basis, but does have an elevated loss in the 60µm to 90µm range suggesting that this sample would be grind sensitive at the nominal grind of 90µm.

It would appear that high residual gold grades are not due to one common cause. It can be theorised that in the case of M2 FRS 5 there is a lot of gold lock-up in some sulphides and other minerals and the grind for this sample might not be optimum at 90µm. The other samples responded to extended leach times and higher NaCN levels, suggesting potential complex composites that are sensitive to rate of dissolution and therefore cyanide concentration.

There is opportunity to further explore the metallurgical characteristics of such composites and perhaps more importantly, the variability of the intervals used to make these composites. As this program (A16557) was paused for inclusion in the feasibility study, this work will be completed during the optimisation phase post the BFS.

Table 13.34 summarises a series of gravity/leach tests undertaken at a NaCN concentration of 300mg/L for 48 hours duration using various combinations of oxygen and air sparging. This data was compiled to allow additional evaluation of the NaCN level and the use of oxygenation. These tests can be directly compared with other tests in this program or alternatively compared with previous tests where the conditions were varied for the same samples.

Table 13.34 Variability Sample Optimisation Leach Series 2016/17 A16557

Au Extraction % Au Grade (g/t) Consumption (kg/t) Sparging Composite ID Gas Gravity Overall Calc. Head Tail NaCN Lime Air 21.26 88.71 2.26 0.26 0.11 0.54 FRS MC 2 Oxygen 27.55 90.29 2.21 0.22 0.15 0.42 FRS MC Air 27.41 88.74 2.04 0.23 0.08 0.43 M1 WOX 1 Air 10.60 82.49 1.08 0.19 0.13 0.75 M10 WOX 16 Air 46.88 92.13 3.18 0.25 0.20 0.59 M10 WOX 17 Air 20.80 91.82 1.47 0.12 0.19 0.48 Air 7.87 80.97 1.52 0.29 0.08 0.74 M2 FRS 4 Oxygen 4.56 78.32 1.31 0.29 0.11 0.64 Air 16.40 85.82 1.80 0.26 0.08 0.30 M4 FRS 7 Oxygen 28.10 89.24 2.28 0.25 0.11 0.37 M6 FRS 12 Air 15.54 84.46 0.87 0.14 0.16 0.32 Air 16.68 85.43 2.61 0.38 0.15 0.34 M7 FRS 14 Oxygen 17.02 88.13 3.20 0.38 0.23 0.33 Air 48.73 92.67 4.64 0.34 0.16 0.32 M9 FRS 15 Oxygen 54.65 95.99 5.23 0.21 0.18 0.29

For those tests that can be compared there appears to be varied responses regarding air versus oxygen sparging due to the influence of head grade variations. Additional analysis is presented in Section 13.16 including comparison of earlier results with those contained in Table 13.34.

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The influence of yield (mass pull to concentrate) on amalgam (free) gold recovery for the WOX and FRS master composites are presented in Figure 13.5 and Figure 13.6 respectively.

Figure 13.5 WOX Master Composite Gravity Performance 2016/17 A16557

WOX MC Gravity Performance - Free Gold 30.00%

25.00%

20.00% y = 0.0155ln(x) + 0.2194 R² = 0.9979

15.00%

10.00% % Recovery of gold to amalgam

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

Figure 13.6 FRS Master Composite Gravity Performance 2016/17 A16557

FRS MC Gravity Performance - Free Gold 30.00%

25.00%

20.00%

y = 0.0149ln(x) + 0.1894 R² = 0.9966 15.00%

10.00% % Recovery of gold to amalgam to gold of % Recovery

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

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As mentioned previously, the mass pull to concentrate would be in the 0.03% range for the type of process facility under consideration. Extrapolating the graphs in Figure 13.5 and Figure 13.6 suggests free gold recovery of nominally 10% for WOX material and less than 10% for FRS material in a full-scale operation.

These plots consider the free gold, or gold that reports to amalgam. If the full-scale flowsheet includes intensive leaching of the gravity concentrate, the free gold plus the soluble gold in the concentrate has to be considered for the circuit design. Additional discussion is presented regarding the influence of this soluble gold in Section 13.16.2 which further discusses the gravity recovery aspects.

Sequential carbon triple contact and equilibrium work was undertaken to provide carbon performance characteristics. The results as reported in Table 13.35 and Table 13.36 can be considered typical for all instances with the exception of the FRS MC equilibrium data. These values are considered high as had been reported for the earlier SOX and MOX values as summarised by Table 13.15 and Table 13.16 (A16326 program). It would seem that the lack of soluble contaminants in the Sanbrado material provides for elevated carbon loadings, which suggests elution capacity can be sized cost effectively.

Table 13.35 WOX and FRS Sequential Carbon Triple Contact Testwork 2016/17 A16557

Fleming Gold Composite ID Adsorption Constants K (hr-1) n WOX MC 162.59 0.770 FRS MC 137.02 0.797

Table 13.36 WOX and FRS Carbon Equilibrium Loading Testwork 2016/17 A16557

Equilibrium Carbon Loading Au (g/t) Composite ID @ Solution Concentration 1.0mg/L 0.50mg/L 0.20mg/L WOX MC 3,172 2,489 1,806 FRS MC 5,767 4,061 2,554

Table 13.37 presents the oxygen uptake data for the WOX and FRS master composites. Both show very low oxygen uptake rates, indicating that a small oxygen supply system would be required in the full-scale plant. As the ores are low in sulphides, in particular reactive sulphides, the low oxygen uptake rates are as expected.

Viscosity testwork was undertaken at a range of pulp densities. The results are summarised in Table 13.38. Even at high pulp densities the viscosity values are low for both master composites, suggesting no potential pumping issues. The low viscosities for the low shear/low pulp density cases will need to be considered for agitation design. The low values are not problematical, but the ability to suspend carbon in the pulp is more difficult at low viscosities.

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Table 13.37 WOX and FRS Oxygen Uptake Testwork 2016/17 A16557

Sample ID

Time WOX MC FRS MC (hours) Oxygen Uptake Rate* Oxygen Uptake Rate* (mg/L/min) (mg/L/min) 0 -0.0170 -0.0188 1 -0.0084 -0.0136 2 -0.0093 -0.0175 3 -0.0066 -0.0149 4 -0.0065 -0.0118 5 0.0076 -0.0095 6 0.0079 -0.0108 24 -0.0035 -0.0048

Table 13.38 WOX and FRS Slurry Rheology Testwork 2016/17 A16557

Bohlin Visco 88 Viscosity @ Shear Rate (sec-1 ) (cps) Composite ID Pulp Density P80:75µm) % Solids (w/w) 4.2 7.4 13.1 21.9 38.9 67.4 119.2 209.5 60 886 540 367 260 179 139 144 168 WOX MC 50 0 0 0 0 0 37 55 78 40 0 0 0 0 0 0 33 55 60 0 0 0 0 90 107 125 143 FRS MC 50 0 0 0 0 0 26 42 64 40 0 0 0 0 0 0 34 54

13.9.3 Summary of Testwork Program A16557 Testwork program A16557 has provided sufficient reliable data regarding comminution and leach characteristics of the various ore domains. It has highlighted variability in some leach performance and provided a means to estimate the likely gravity recoveries for WOX and FRS ore materials.

The program highlights a need to further investigate this variability as there is an opportunity to increase gold extractions over those predicted by the recovery algorithms which have used some of this sub-optimal extraction data.

Testwork Program A17283 ALS program A17283 was undertaken to obtain some preliminary metallurgical results from the M1 deposit which, at the time, had recently been defined by drilling but not metallurgically assessed.

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Testwork included the following:

. Composite preparation and head assays. . Bond Ball Mill Work Index determination. . Screen fire assays. . Grind establishment. . Gravity/cyanide leaching testwork. Testwork included the necessary sample preparation, Bond BWi determination, head assays, SFA and gravity/cyanidation using the same general conditions as the earlier testwork, being a conventional gravity step followed by leaching of the gravity tail.

13.10.1 Samples Utilised for Testwork Program A17283 Half NQ drill core was selected from fresh material from one drillhole (TAN16 DD034) by the WAF geological team and sent to ALS where it was received June 2016. Sample was limited and as a result only two metallurgical testing composites were generated, comprising two lots of contiguous core from 109m to 114m downhole (Composite 1) and 128m to 133m downhole (Composite 2). A dedicated Bond BWi composite was also generated from select half core from 106m to 136m downhole.

13.10.2 Results Testwork Program A17283 Head assays were determined by SFA at a grind size of 75µm. The results are presented in Table 13.39. The gold head grades of the +75µm and –75µm fractions are similar, suggesting that little coarse gold was present. Silver content is minor, similar to the M5 results. Most sulphur is present as sulphides, with Composite 1 containing a much higher level of sulphides.

Table 13.39 Head Assays A17283

Analyte Unit Composite 1 Composite 2 Au (via SFA 75um) g/t 1.98 1.44 Ag g/t 0.6 <0.3 STOTAL % 2.30 0.62 SSULPHIDE % 2.18 0.56 Cu g/t 90 95 Zn g/t 95 65

A single Bond BWi was determined and reported at a moderate value of 14.4kWh/t, which is at the lower end of other fresh ore determinations made for the M5 material.

Gravity/leaching testwork was conducted at grinds of P80 106µm and 75µm and at NaCN concentrations of 500mg/L and 250mg/L to encompass the ranges investigated in previous M5 testwork. Gravity recovery was achieved by passing the ground sample through a 3 inch Knelson concentrator, amalgamating for free gold recovery, combining the amalgam and gravity tails and then agitation leach in an oxygen sparged bottle roll.

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The results of the testwork are cutoff summarised in Table 13.40.

Table 13.40 M1 Composite Gravity / Leach /Results A17283

Au Extraction % Au Grade Consumption Composite P80 Grind Sodium Cyanide @ hours (g/t) (kg/t) ID Size (µm) Target (mg/L) Gravity 48 Calc. Head Tail NaCN Lime 500 4.91 84.05 2.04 0.33 0.37 0.35 106 250 7.46 82.14 2.21 0.40 0.17 0.42 Composite 1 500 7.89 89.05 1.96 0.22 0.37 0.30 75 250 8.82 87.13 2.10 0.27 0.13 0.58 500 21.35 90.68 1.29 0.12 0.29 0.22 106 250 14.10 88.59 1.49 0.17 0.07 0.29 Composite 2 500 28.32 92.53 1.61 0.12 0.25 0.22 75 250 20.86 90.96 1.44 0.13 0.15 0.24

Both composites gave results in line with previous M5 testwork results, suggesting that the M1 material was amenable to processing using the same flowsheet as M5 material. Some grind sensitivity was similarly presented. There appears to be some sensitivity to NaCN concentration at around 0.05g/t benefit for the higher NaCN value of 500mg/L, suggesting that higher NaCN concentrations may be economically viable for Mi material.

The residue grades for Composite 1 are around double those of Composite 2. Similarly, the NaCN consumption is slightly higher than Composite 2, albeit low compared to many fresh ore types. The notable difference between these two composites is the level of sulphide present and gold head grade, both of which are higher for Composite 1. This sulphide, if having similar mineralogy to M5 material, can be expected to hold more gold post leaching. The higher head grade would also suggest a higher lock up of gold in other minerals simply due to the higher head grade itself.

Elevated sulphide concentrations often result in elevated NaCN consumption due to the elevated concentration of thiocyanates. This, along with a slightly higher concentration of zinc (a potential cyanicide) would explain the slightly higher NaCN consumption seen for Composite 1.

Silver results are contained in the A17283 report. Extractions were again similar to those observed in the M5 leaching, with extractions in the mid 40% to high 60% range.

13.10.3 Summary of Testwork Program A16283 This preliminary testwork program on M1 FRS samples showed that the material had very similar characteristics as the M5 FRS material. Consequently, this allowed the feasibility study to progress based on the premise M1 material would be amenable to the flowsheet proposed. Further M1 work could be conducted during detailed design to confirm this assumption.

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Testwork Program A17341 (2016/17) ALS program A17341 was developed to provide comminution data as well as additional data on both the M1 and M5 material characteristics.

Testwork included the following:

. Composite preparation. . UCS and CWi testwork. . JK and SMC comminution testwork. . Bond Rod Mill and Ball Mill Work Index determinations. . Abrasion Index determinations. . Grind establishment. . Gravity/cyanide leach testwork. . Other sundry diagnostic testwork.

13.11.1 Samples Utilised for Testwork Program A17341 Sample comprised full core from 5 DC drillholes drilled specifically for metallurgical sample selection. The holes were designated TAN16 MET11 to TAN 16 MET16 and were a mix of PQ and HQ holes which allowed for sample to be submitted for JKTech comminution testing as well as UCS and other comminution tests. Drillholes TAN16 MET11 to TAN16 MET14 were from the M1 deposit and drillholes TAN16 MET15 and TAN16 MET16 were from the M5 deposit.

Intervals were selected and subjected to UCS testing and CWi determinations. Select intervals were also subjected to JK and SMC testing, with an overlap of the JK and SMC samples to allow comparison between JK and SMC data outputs.

Five discrete comminution composites were compiled for Bond RWi and Bond BWi as well as Ai testing. These composites included overlap with intervals used for the UCS, JK and SMC work to give a set of spatially comparable comminution data.

The samples were predominantly FRS material, with minor amounts of MOX material. Following assay determinations, a total of 10 FRS composites and one MOX composite were also prepared for gravity and leach testing.

13.11.2 Results Testwork Program A17341 (Comminution) Samples were taken from defined FRS intervals and subject to UCS testing. Table 13.41 summarises the results, which show that the FRS material is classified as medium hard to hard. These values do not show any issue with regards to conventional crushing applications.

CWi determinations were conducted using the same general interval ranges as the UCS tests to allow direct comparison. The summary results are presented in Table 13.42. Similar to the UCS tests, the CWi values suggest no issue with regard to applying conventional crushing techniques.

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Table 13.41 UCS Summary 2016/7 - A17341

Interval UCS Strength Sample ID Failure Mode (m) (MPa) Description 65.60 – 65.92 26.387 Shear Med. Hard TAN16 MET11 68.32 – 68.50 135.410 Cataclasis Hard 40.30 – 40.70 57.285 Shear Med. Hard TAN16 MET12 44.90 – 45.54 37.598 Shear Med. Hard 72.38 – 72.63 49.233 Shear Med. Hard TAN16 MET13 79.52 – 79.85 56.219 Shear Med. Hard 66.25 – 66.56 41.074 Shear Med. Hard TAN16 MET14 72.50 – 72.80 83.797 Shear Hard 101.90 – 102.23 49.393 Columnar Med. Hard TAN16 MET15 108.14 – 108.46 65.369 Shear Hard 101.60 – 101.90 61.457 Shear Hard TAN16 MET16 89.80 – 90.03 27.368 Shear Med. Hard

Table 13.42 Crushing Work Index Determinations A17341

Bond Impact Crushing Work Index (kWh/t) Sample ID Standard Average Maximum Minimum Deviation TAN16 MET11 (64.00 – 70.68m) 6.6 13.3 2.4 3.6 TAN16 MET12 (35.55 – 47.00m) 10.8 14.6 8.5 1.8 TAN16 MET13 (70.75 – 80.52m) 6.3 8.3 4.9 1.5 TAN16 MET14 (63.10 – 73.20m) 5.2 8.4 3.4 1.5 TAN16 MET15 (97.31 – 110.20m) 9.9 16.9 6.4 3.7 TAN16 MET16 (85.56 – 104.50m) 7.7 11.5 2.5 2.6

A series of intervals generally aligned with the UCS and CWi intervals were subjected to SMC testing. This test regime provides data used in sizing and selecting comminution equipment. The data was provided to Orway Mineral Consultants Pty Ltd (OMC) who have sized and selected equipment per the flowsheet and as further described in Section 17. The detailed testwork and application of this data is described in Appendix II of ALS report A17341. Summary results are presented in Table 13.43.

The results show that the FRS material is quite hard/tough as represented by the A*b values in the 30 to 36 range and will require a relatively high amount of energy for comminution. The data also suggests that single stage SAG milling as proposed for the comminution flowsheet is a valid selection. However, the application will require a blending strategy of the softer material types. Detailed discussion regarding the application of this data can be found in the process plant section (Section 17).

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Table 13.43 SMC Determinations 2016/7 - A17341

Derived Values TAN 16 DWi SG Mia Mih Mic Composite ID (kWh/m³) A b A*b ta (kWh/t) (kWh/t) (kWh/t) MET 11 8.99 2.8 71.3 0.43 30.7 24.0 18.8 9.7 0.29 (65-68m) MET 12 8.87 2.8 61.4 0.52 31.9 23.4 18.3 9.5 0.29 (33-36m) MET 12 8.06 2.8 59.7 0.58 34.6 21.8 16.8 8.7 0.32 (40-43m) MET 12 8.96 2.8 69.5 0.44 30.6 24.0 18.8 9.7 0.29 (63-66m) MET 13 8.66 2.8 77.1 0.41 31.6 23.4 18.2 9.4 0.30 (70-74m) MET 14 8.95 2.7 75.5 0.40 30.2 24.3 19.1 9.9 0.29 (74-78m) MET 15 7.56 2.7 61.3 0.59 36.2 21.2 16.1 8.3 0.34 (89-92m) MET 15 8.98 2.7 77.3 0.39 30.1 24.6 19.3 10.0 0.29 (104-107m) MET 16 8.04 2.7 69.7 0.48 33.5 22.7 17.4 9.0 0.32 (90-93m)

Ai tests were conducted on the 5 comminution composites prepared. These comminution composites cover the same general range of drillhole intervals as subjected to the UCS, CWi and SMC testing.

The composites subjected to Ai testing were also subjected to Bond RWi and BWi determinations. The Ai values are directly associated with the RWi and BWi values.

Table 13.44 summarises the Ai results. The values show moderate to high abrasion which will be reflected in moderate to high wear costs associated with crushing liners, mill liners, mill media and screening and feeder equipment when processing FRS ore types.

Table 13.44 Abrasion Index Determinations 2016/7 - A17341

Composite ID Bond Abrasion Index (Ai) TAN16 MET11 0.1330 TAN16 MET12 0.1128 TAN16 MET14 0.3498 TAN16 MET15 0.2657 TAN16 MET16 0.3590

The more abrasive composites have generally a higher proportion of sedimentary schist whereas the less abrasive materials are dominated by psammite or other minor rock types.

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The five comminution samples were subjected to Bond RWi and BWi testing. Results are summarised in Table 13.45. Results show the material to again be hard/tough with moderate to high work indices. This aligns with the SMC results and also with the Ai results to a lesser degree.

Table 13.45 Bond Rod and Ball Work Index Determinations 2016/7 - A17341

Bond RWi Bond BWi Composite ID (kWh/t) (kWh/t) TAN 16 MET11 21.5 16.4 TAN 16 MET12 20.7 15.8 TAN 16 MET14 20.3 17.0 TAN 16 MET15 22.4 18.4 TAN 16 MET16 22.0 21.3

The comminution data on the FRS material shows it is competent and hard/tough with moderate to high abrasion characteristics, which is common for such material. The values obtained suggest that these materials are inside the bounds of conventional comminution devices such as jaw and cone crushers, SAG and ball mills.

The SOX, MOX and to a lesser degree WOX materials are less competent, and if blended with the FRS material allow for some reduction in the power demands and equipment sizing compared to a full FRS feed to the comminution circuit. Due to the proposed mining schedule, softer ores will be available for most of the mine life allowing for such a philosophy to be applied. This opportunity forms the basis of the OMC modelling and is described in more detail in Section 17.

The early Ai and BWi values presented in Table 13.4 from testwork program A15684 gave low values for the FRS sample tested. These results are not reflected in the more definitive program conducted as part of A17341. The earlier values are considered dubious and have not been used as design values.

13.11.3 Results Testwork Program A17431 (Gravity and Leaching) Nine composite samples were submitted for gravity and leaching testwork. The composites selected had head grades in the ranges expected to be experienced by the proposed processing facility.

Table 13.46 summaries the head assay data. Several points are noted as follows:

. There is some scatter in the gold head assays suggesting the presence of free gold. . Copper grades for some composites are greater than 100ppm suggesting the possibility of elevated NaCN consumption if the copper is soluble. . The sulphide sulphur grades in some composites are elevated. Given previous performance of other samples with elevated sulphide grades, some gold lock-up in these composites may be expected.

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Table 13.46 Head Assay Results 2016/7 - A17341

Au Au (SFA) Ag Cu C S Composite ID organic Sulphur (g/t) (g/t) (g/t) (g/t) (%) (%) M11 FRS30 3.57/3.82 3.40 1.2 75 0.09 2.36 M11 FRS31 1.59/1.54 1.74 0.6 45 0.06 1.58 M12 FRS32 2.96/3.05 3.05 0.6 70 0.03 1.90 M12 FRS33 6.49/6.25 5.02 1.2 50 0.09 1.80 M12 FRS34 1.84/2.00 2.60 0.3 65 0.03 0.98 M13 FRS35 2.41/2.24 2.14 <0.3 85 0.06 0.86 M13 FRS36 8.55/8.44 10.4 0.9 135 0.03 0.92 M15 FRS38 3.42/3.51 2.58 1.2 120 0.03 0.44 M16 FRS39 3.98/4.62 5.19 0.9 205 0.03 0.40

The composites were subjected to the same generic flowsheet as previously, consisting of gravity via Knelson concentrator, amalgamation followed by agitated leaching of the

amalgam and gravity tails. Grind sizes of P80 of 75µm and 106µm were evaluated for each sample. A NaCN concentration of 300mg/L was maintained with oxygen sparging for all the tests. Kinetic data was compiled by way of intermediate solution sampling for a leach period of 24 hours.

A summary of the results of the program are presented in Table 13.47.

Table 13.47 Summary Gravity-Leach Data 2016/7 - A17341

Au Extraction % Au Grade (g/t) Consumption (kg/t) P Grind Composite ID 80 Size (µm) Gravity 2hrs 6hrs 24hrs Calc. Head Tail NaCN Lime 106 35.03 81.00 88.57 90.47 3.88 0.37 0.27 0.63 M11 FRS30 75 30.64 66.77 81.41 92.56 3.97 0.30 0.22 0.63 106 45.47 62.06 79.41 88.94 2.03 0.23 0.19 0.30 M11 FRS31 75 52.80 75.99 89.98 94.05 2.10 0.13 0.11 0.40 106 28.28 48.39 69.82 80.70 3.39 0.66 0.21 0.35 M12 FRS32 75 29.03 60.26 77.97 79.55 2.74 0.56 0.19 0.38 106 26.00 66.68 88.49 92.10 6.27 0.50 0.14 0.39 M12 FRS33 75 62.25 74.11 91.33 97.07 10.2 0.30 0.14 0.47 106 31.37 76.76 90.86 92.43 2.71 0.21 0.14 0.28 M12 FRS34 75 46.93 74.43 93.62 96.19 3.41 0.13 0.11 0.41 106 16.67 74.65 82.13 85.19 2.16 0.32 0.11 0.30 M13 FRS35 75 18.21 71.57 86.51 87.72 2.36 0.29 0.19 0.28 106 12.14 39.50 72.29 88.60 8.11 0.93 0.11 0.30 M13 FRS36 75 19.03 46.66 79.99 91.36 11.5 1.00 0.16 0.26 106 47.33 84.03 89.76 92.73 2.82 0.21 0.08 0.35 M15 FRS38 75 50.38 87.46 92.26 93.24 2.14 0.15 0.11 0.31 106 15.32 77.12 84.58 85.42 4.73 0.69 0.11 0.49 M16 FRS39 75 18.59 80.88 83.31 88.02 4.84 0.58 0.11 0.52

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A number of observations can be made from the FRS Composite results in Table 13.28, as follows:

. Gravity gold recovery is variable and does not appear to be directly related to the head grade. For example, the high grade M13 FRS36 sample has presented low gravity recovery. . Gold leaching extractions are generally high (early 90’s). The exception is sample M13 FRS36. . Reagent consumption is relatively low. It is noted that the higher sulphide content samples tend to have higher NaCN consumptions.

. There is some grind sensitivity presented between a P80 of 106µm and a P80 of 75µm. Incremental gains averaging 0.076g/t Au were achieved. Figure 13.7 presents the residue grade for each test against the sulphide sulphur assay. The grind sensitivity is also represented by the off-set in the 106µm to 75µm data. This graph suggests that there is no direct association between sulphide sulphur level and residue grade. However, the plot does not consider the head grade of the leach test which might also influence the outcome. Additional analysis is presented in Section 13.16.

Figure 13.7 Residue Grade and Sulphide Influence A17341

Residue grade and sulphide influence

1.2 Res 106 um Res 75 um 1

0.8

0.6

0.4

Residue Residue GradeAu g/t 0.2

0 00.511.522.5 Sulphide sulphur %

A simplified diagnostic leach evaluation was undertaken on the 106µm grind residues from the leach tests of samples M12 FRS32, M12 FRS33, M13 FRS36 and M16 FRS39. An aqua regia leach was completed, followed by a FA of the aqua regia leach residue. Results are presented in Table 13.48 and show that the bulk of the residual gold is contained in acid digestible minerals, possibly sulphides. Lock up in non-acid digestible (mostly silicates) is minor, suggesting that even at a 106µm grind, liberation is good and that additional grinding may potentially benefit gold extractions by releasing gold locked in sulphides.

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Table 13.48 Diagnostic Leach Results 2016/17 - A17341

M12 FRS 32 M12 FRS 33 M13 FRS 36 M16 FRS 39 Gold Content Au Distribution Description g/t % g/t % g/t % g/t % Acid Digestible Mineral 0.60 95.2 0.44 91.7 0.82 92.1 0.68 81.9 Locked Silicate (Gangue) 0.03 4.8 0.04 8.3 0.07 7.9 0.15 18.1 Encapsulated Residue Assay 0.66/0.65 - 0.48/0.51 - 0.93/0.92 - 0.65/0.73 - Total - 100.00 - 100.00 - 100.00 - 100.00

Earlier testwork had suggested that the samples were not sensitive to cyanide concentrations. However, it was decided to assess whether or not an increase in NaCN concentrations would be beneficial to gold extractions, specifically for the samples observed to have elevated residue grades. Consequently, the same samples subjected to the short diagnostic leaching tests were subjected to a repeat gravity and leach test but at a constant NaCN concentration of 500mg/L throughout the test.

The results, presented in Table 13.49, show little if any variance in the outcomes which suggests that the high residue grades are not a function of low cyanide concentrations. The NaCN consumptions were marginally higher than previous tests, which is to be expected. The consumptions are still considered low even at these elevated concentrations. Consequently, the diagnostic work and this repeat leach work suggested that these samples have gold lock-up in acid digestible minerals which is responsible for the elevated residue grades.

Table 13.49 Elevated NaCN Gravity Leach Tests (106µm) 2016/17 - A17341

Au Extraction % @ hrs Au Grade (g/t) Consumption (kg/t) Composite ID Gravity 2 6 24 Calc. Head Tail NaCN Lime M12 FRS32 38.63 69.57 77.25 80.34 3.64 0.72 0.23 0.32 M12 FRS33 33.77 63.16 83.79 90.57 6.84 0.65 0.27 0.37 M13 FRS36 20.05 68.83 84.08 88.98 9.35 1.03 0.17 0.28 M16 FRS39 19.60 78.99 83.86 86.15 5.13 0.71 0.14 0.58

13.11.4 Summary of Testwork Program A17431 Program A17431 effectively confirmed that the M1 FRS material behaves very similar to the M5 FRS material.

JKTech and SMC testing showed that the FRS material is tough but can be processed via conventional grinding methods. The UCS and CWi data also showed that conventional crushing methods are applicable.

Gravity and leach behaviour is in line with previous FRS observations. Some higher grade residues present are due to residual gold mainly associated with acid digestible minerals, possibly sulphides. However, variations in gravity gold recoveries and sulphide assays suggest that gold deportment is not the only reason.

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Reagent consumption remains low for both NaCN (<0.3kg/t) and quicklime, and leaching is mostly complete in 12 to 18 hours with 24 hour extractions, ranging from uncharacteristically lower values of 80% to highs of +95%.

Testwork Program A17750 (2016/17) This testwork program focused on the M1 deposit, which has a number of very high gold grade areas. These gold grades are substantially higher than the grades represented by the various composites in the previous testwork programs. WAF dispatched a number of high grade RC and DC drill samples to ALS Perth for basic evaluation.

The work involved the following:

. Head assay determinations. . Compositing. . Gravity recovery evaluation. . Agitated bottle roll leaching, both direct leach and with carbon present to deal with the high grades.

13.12.1 Samples Utilised for Testwork Program A17750 Samples used in the program originated from DC drillholes 44, 52 and 55 and RC holes 122 and 131. Some additional intervals were selected and held for possible future work.

Five composites were compiled. Two from RC chips (HG RC 1 and 2) and three from DC drillhole crushed remnants post assay split (HG DD 3, 4 and 5).

13.12.2 Results Testwork Program A17750 Head assays were determined for each of the composites. Results are presented in Table 13.50. Repeat gold assays showed repeatable values. Given the high grades, this was not expected as it was suggested that the high grade is due to coarse gold blebs or “spotty” gold and thus more variable results would have been expected.

Table 13.50 High Grade Composite Head Assays 2016/17 - A17750

Element Units HG RC 1 HG RC 2 HG DD 3 HG DD 4 HG DD 5

Au1 g/t 152 167 145 75.5 114

Au2 g/t 154 158 141 83.9 124 Ag g/t 25 20 20 10 15 Cu g/t 50 85 50 100 45 Fe % 4.26 4.32 2.7 5.92 2.94 S2- % 1.38 1.12 1.42 0.82 1.26

As sample availability was limited, no SFA were undertaken to establish coarse/free gold influences.

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Of interest, sulphide levels are low as were copper and iron assays, generally in line with the earlier lower grade M1 and M5 samples tested. This suggests that the high gold values are not associated with elevated sulphide content.

Following grind establishment, the samples were ground to a P80 of 90µm. A gravity step was conducted using a 3 inch Knelson concentrator and amalgam step as had been maintained thought the previous testwork programs. Amalgam and gravity tails were combined for leaching.

One leach test was conducted without carbon present, and a duplicate test was conducted with carbon present. It is often observed with high grade leaching that the elevated solution tenors can retard further gold leaching. The use of carbon negates this influence.

Elevated NaCN concentrations of 1000mg/L were used, due to the high grades and short leaching times. The leach tests were oxygen sparged and ran for 24 hours. The results of the program are summarised in Table 13.51.

Table 13.51 High Grade Gravity/Leach Summary Results 2016/17 - A17750

Au Head Grade Au Extraction Au Tail Reagents Composite ID (g/t) (%) Grade (kg/t) Assay Calc. Gravity Gold Total Gold (g/t) NaCN Lime

Direct Cyanidation P80: 90µm HG RC 1 152 / 154 142 13.66 96.91 4.39 0.48 0.40 HG RC 2 167 / 158 178 15.32 96.93 5.46 0.47 0.35 HG DD 3 145 / 141 147 42.26 94.77 7.67 0.59 0.36 HG DD 4 75.5 /83.9 113 36.79 94.29 6.46 0.43 0.34 HG DD 5 114 / 124 115 8.78 96.62 3.89 0.44 0.32

CIL Cyanidation P80: 90µm HG RC 1 152 / 154 127 15.27 96.11 4.94 0.60 0.41 HG RC 2 167 / 158 121 22.49 94.66 6.48 0.60 0.38 HG DD 3 145 / 141 145 42.70 94.01 8.70 0.66 0.37 HG DD 4 75.5 /83.9 94.7 43.97 95.39 4.37 0.49 0.36 HG DD 5 114 / 124 103 9.83 96.02 4.09 0.65 0.30

The results of the testwork report elevated residue grades, which is to be expected given the exceptionally high head grades. Leach extractions are all around 94% and similar with or without carbon present. Other points of note are as follows:

. Assay head and calculated head grades align well for such high grade material, which again show limited influence of possible ”spotty” gold. Some deviation has occurred in the HG RC 1 and 2 CIL tests. . Gravity recovery is not particularly high, suggesting that the gold remains associated with composite particles. . The gravity recovery of +40% for sample HG DD 3 presents a high proportion of free gold. However, the repeatability of the head assays and calculated head grades suggest that the gold is fine and well dispersed.

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. Reagent consumptions are still relatively low, especially when considering the elevated NaCN concentration used. The higher consumption in the CIL tests will be due to the carbon oxidising the cyanide, which is a common observation with CIL tests. . Kinetic results (refer to the ALS report in Appendix 13.7) show that most of the leaching is concluded in the first 6 to 12 hours, again supporting the presence of fine gold. A diagnostic leach test was performed on three of the composites, HG RC 1, HG RC 2 and HG DD 3. The process involved intensive cyanidation for 24 hours at 2% NaCN with LeachWell followed by an aqua regia digestion and finally a FA of the final residue.

The RC composite results are summarised in Table 13.52 and DC composite results are summarised in Table 13.53.

Table 13.52 High Grade Diagnostic Leach Results RC Comps 2016/17 - A17750

HG RC 1 HG RC 2 Diagnostic Gold Content Description Au Distribution Stage g/t % g/t % 1 Free/Cyanidable 1.05 27.93 1.87 36.06 2 Acid Digestible Mineral Locked 2.04 54.25 2.46 47.52 3 Silicate (Gangue) Encapsulated 0.67 17.82 0.85 16.42 Calculated Head 3.76 - 5.18 - Total Gold Content - 100.00 - 100.00

Table 13.53 High Grade Diagnostic Leach Results DC Comps 2016/17 - A17750

HG DD 3 HG DD 4 HG DD5 Diagnostic Gold Content Description Au Distribution Stage g/t % g/t % g/t % 1 Free/Cyanidable 1.17 16.46 0.82 21.12 0.58 17.04 2 Acid Digestible Mineral Locked 5.49 77.47 2.12 54.83 2.44 71.27 3 Silicate (Gangue) Encapsulated 0.43 6.07 0.93 24.05 0.40 11.68 Calculated Head 7.09 - - 3.42 - Total Gold Content 100.00 - 100.00 - 100.00

13.12.3 Summary of Testwork Program A17750 Program A17750 has shown that high grade gold samples of +100g/t gold from the M1 FRS region can provide leach extractions in the mid 90% range. Of note, these high grade samples do not necessarily have high gravity gold components, suggesting that the gold deportment is similar to the lower grade M1 and M5 FRS material. Even at elevated NaCN dose rates, NaCN consumptions are relatively low, and CIL processing is not necessary to maintain high leach extractions.

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Outotec Thickening Testwork – Outotec Report S2921TA (2016/17) Outotec undertook limited flocculant screening followed by dynamic thickener testing on bulk samples of SOX, MOX, WOX and FRS samples from the M5 deposit that had been prepared from the bulk leach tests undertaken by ALS.

The flocculant screening tests were conducted on the WOX bulk sample. Figure 13.8 presents the results of this screening work and shows a mix of behaviour and in some instances very high flocculant dosage rates. Outotec recommended the use of flocculant M10 at a pH of 11 to achieve improved overflow clarity.

Figure 13.8 MOX Flocculant Screening Results – Extract from Outotec Report

The performance of flocculant M155 was very good with regard to settling but clarity was poor. This flocculant was not tested at elevated pH levels, which would appear to be a shortcoming given the improved settling at much lower dose rates than the M10 option.

Testing of the various material types progressed using M10 flocculant. All samples were successfully thickened under a range of different flux rates but at high flocculant dosages and with varying overflow clarity as outlined below:

. SOX material required very high flocculant dosages in the range of 70ppm to 90ppm to achieve nominally 50% to 55% solids underflow density at flux rates of 0.7t/m²h. . MOX material required flocculant dosages in the range of 50ppm to 60ppm to achieve nominally 55% underflow density at flux rates of 0.5t/m²h to 1.0t/m²h. . WOX material required flocculant dosages in the range of 40ppm to 50ppm to achieve nominally 60% underflow density at flux rates of 0.5t/m²h. . FRS material required flocculant dosages in the range of 50ppm to 70ppm to achieve nominally 57% to 60% underflow density at flux rates of 0.7t/m²h.

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This data has been used for the thickener sizing and operating costs estimates. However, it is possible that the flocculant selection is not fully suited for the Sanbrado material and should be investigated further during the detailed design stage of the project. Additional flocculant screening and repeat dynamic thickener testing is warranted to optimise the thickener needed. In addition, a series of blends of material types to improve the oxide material performance is warranted. The thickener sizing and operating cost estimate associated with this area of the flowsheet might thus be somewhat conservative.

Mineralogical Analysis

13.14.1 Channel Resources/SGS Work Comment regarding the mineralogical characteristics of the Sanbrado materials was made in Section 13.4 when discussing earlier testwork undertaken by Channel. In the previous work, a sulphide (FRS) sample and an oxide (SOX/MOX) sample were evaluated using XRD and QEMSCAN techniques. The work was undertaken by SGS on behalf of Channel.

The samples were prepared by grinding to nominally P80 of 150µm and were screened at 106µm and 38µm fractions, which is coarser than the nominal grind of P80 90µm proposed for the process facility.

The average gold grade for the sulphide sample was 1.58g/t Au with 142 gold grains identified occurring mainly as native gold and electrum. Gold content averaged 86.8% for the 19 gold grains analysed. The gold particles ranged in size from 1µm to 60µm, with the highest percentage (30.5 weight %) of grains between 10µm and 15µm. The sulphide sample contained around 1% pyrite and around 0.8% pyrrhotite.

The oxide sample average gold grade was 1.79g/t Au with 73 gold grains identified occurring mainly as native gold. Gold content averaged 87.8% for the 14 gold grains analysed. The gold particles range in size from 1µm to 60µm, with the highest percentage (22.0 weight %) of grains between 2µm and 5µm. The oxide sample contained 0.3% pyrite.

Regarding gold liberation, the sulphide sample contained pure, free and liberated gold minerals accounting for 65% of the gold. The main association of gold minerals is as middlings with quartz/feldspar (14.1%), followed by complex particles (12.1%), as middlings with other silicates (6.9% and Fe Sulphides (1.1%). Trace amounts (<1%) occur as middlings with other minerals.

In the oxide sample, pure, free and liberated gold minerals account for 42.7% of the gold. The main association of gold minerals is as middlings with quartz/feldspar (44.0%), followed by middlings with mica (8.3%), complex particles (3.6%) and other silicates (1.5%). Trace amounts (<1%) occur as middlings with other minerals.

These observations, whilst only coming from two samples from the earliest defined auriferous areas, correlate well with the metallurgical observations, as follows:

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. Most of the gold in these samples is liberated or associated with non-sulphide minerals. High gold extractions are therefore expected, particularly at the finer grinds as tested in the subsequent metallurgical program which will have addressed some of the complex particle issues. . The amount of coarse gold is low, which is reflected in the faster kinetics observed during the subsequent metallurgical testing. . No notable cyanicides were identified, again reflected in subsequent testwork. . The effect of gold locked in sulphides appears to be minor. . No carbonaceous materials are present, suggesting that no preg-robbing issues would be expected, as observed by subsequent metallurgical testwork.

13.14.2 WAF Mineralogical Work WAF has undertaken a number of petrological studies. In November 2016 Minerex Petrographic Services evaluated a number of drill core sample fragments to identify rock types. This work also resulted in observation of some free gold associated with sulphides from drillhole TAN15 – DD029 at 53.4m to 53.5m downhole.

Fine gold inclusions of less than 20µm were also found contained within arsenopyrite grains. The arsenopyrite itself was associated with pyrrhotite or pyrite (Figure 13.9, Figure 13.10 and Figure 13.11).

Figure 13.9 Gold Inclusion – First Image

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Figure 13.10 Gold Inclusion – Second Image

Figure 13.11 Gold Inclusion – Third Image

Such inclusions could be expected to provide some of the residual gold in leach tails. Noting the scale in these figures and considering that preferential grinding of sulphides will occur, there is a reasonable probability that a majority of such gold grains would be liberated in the full-scale plant.

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These images suggest that there will be some areas of the deposits where lock up of gold may be more prevalent than other areas if the ratio of gold contained in such sulphides is higher. This may also explain some of the higher grade residues observed from time to time in the testwork. Optimisation work will need to be conducted during operations when these areas are targeted for processing.

Metallurgical Testwork Data Analysis In excess of 100 separate gravity/leach tests have been conducted on a range of samples from the Sanbrado Gold Project. Different grind sizes, cyanide concentrations, oxygen sources and leaching times have been trialled. The dataset has allowed for various analyses to be conducted, a number of which are summarised in the sections below.

13.15.1 Evaluation of Key Gold Leaching Processing Variables Table 13.54 presents the grind sensitivity data for a mix of MOX, WOX and FRS samples and

master composites. Grinds of P80 150µm, 125µm, 106µm and75 µm were evaluated. An initial NaCN concentration of 500mg/L controlled to 250mg/L and oxygen sparging at 40% solids w/w were utilised in a 48h leach.

Table 13.54 Grind Sensitivity

Grind Size Gravity Gold Total Gold Residue Grade Calc Head Assay Head Sample ID (um) (%) Extrn (%) (Au g/t) (Au g/t) (Au g/t)

P80 : 150 19.28 83.93 0.15 0.93 0.72

P80 : 125 17.22 84.05 0.13 0.78 0.72 M4 MOX 6 P80 : 106 23.96 86.82 0.11 0.83 0.72

P80 : 75 21.46 89.58 0.09 0.82 0.72

P80 : 150 19.46 83.68 0.26 1.59 1.68

P80 : 125 31.04 89.09 0.16 1.47 1.68 WOX MC P80 : 106 21.22 89.64 0.21 1.98 1.68

P80 : 75 24.32 91.50 0.15 1.71 1.68

P80 : 150 15.77 76.22 0.48 2.00 2.66

P80 : 125 24.00 82.93 0.37 2.17 2.66 FRS MC P80 : 106 35.63 85.89 0.40 2.83 2.66

P80 : 75 31.84 88.03 0.28 2.34 2.66

P80 : 150 40.90 92.34 0.22 2.87 3.57

P80 : 125 41.86 94.66 0.15 2.81 3.57 M10 WOX 18 P80 : 106 41.93 97.05 0.10 3.22 3.57

P80 : 75 34.74 95.57 0.10 2.26 3.57

P80 : 150 3.29 71.89 0.39 1.37 1.72

M2 FRS 4 P80 : 125 2.92 71.13 0.45 1.54 1.72

P80 : 106 4.69 77.90 0.33 1.49 1.72

M2 FRS 4 P80 : 75 0.19 80.97 0.26 1.34 1.72

P80 : 150 12.66 82.09 0.50 2.76 3.45

P80 : 125 17.55 81.48 0.48 2.59 3.45 M7 FRS 14 P80 : 106 19.15 86.32 0.40 2.92 3.45

P80 : 75 20.60 88.23 0.32 2.72 3.45

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Figure 13.12, Figure 13.13 and Figure 13.14 present the data schematically for the M4 MOX 6 composite, WOX MC and FRS MC samples. All show an increasing benefit with

finer grind sizes down to a P80 of 75µm. At the time of undertaking this work, the availability of MOX material was limited and no MOX MC was evaluated.

Figure 13.12 Grind Sensitivity, M4 MOX 6

M4 MOX 6 90% 0.16

89% 0.14

0.12 88%

0.10 87% 0.08 86%

Total Extraction Total 0.06 Residue Residue Grade,g/t 85% 0.04

84% 0.02

83% 0.00 60 80 100 120 140 160 P80 Grind, um

Extraction, % Residue Au g/t

Figure 13.13 Grind Sensitivity, WOX Master Composite

WOX MC 92% 0.30

91% 0.25 90%

89% 0.20

88% 0.15 87% Total Extraction

86% 0.10 Residue Grade,g/t

85% 0.05 84%

83% 0.00 60 80 100 120 140 160 P80 Grind, um

Extraction, % Residue Au g/t

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Figure 13.14 Grind Sensitivity, FRS Master Composite

FRS MC 90% 0.50

0.45 88% 0.40 86% 0.35 84% 0.30

82% 0.25

0.20 Total Extraction Total 80% Residue Residue Grade,g/t 0.15 78% 0.10 76% 0.05

74% 0.00 60 80 100 120 140 160 P80 Grind, um

Extraction, % Residue Au g/t

In addition, the SOX material was already known to have very low grind sensitivity based on the earlier heap leach testwork where +90% extractions were achieved at a fine crush size.

Based on the data only, there seemed to be adequate argument to justify a grind of 75µm, which was the grind size selected for subsequent testwork. However, as additional data came to hand, the optimum economic grind size was determined to be between 75µm and 106µm, and a design grind size of 90µm was selected as the basis going forward in the latter stages of the testwork. Variability of metallurgical performance between 75µm, 90µm and even 106µm grinds is more pronounced when treating the harder ores like FRS and less prominent when treating the softer oxides ores (SOX and WOX). However, the harder the ore, the larger the effect of grinding on capital and operating cost and thus on the economic viability of the project.

Table 13.55 presents a summary of various testwork outcomes for differing pulp density, use of CIL, air and oxygen applications (note all other leach tests were conducted at

40% solids w/w, P80 of 75µm, target NaCN concentration of 250mg/L or 150mg/L and leaching times of 48 hours).

The results do not present any obvious advantages or disadvantages under the various conditions tested. Whilst some residue grades show improvement, when other aspects such as the head grade or gravity component are considered, there is no clear benefit. For example, the FRS MC test utilising air presents the worst residue grade. However, a +40% gravity recovery was achieved and the calculated head grade is at times almost double the other FRS MC calculated head grades. This makes these results difficult to compare directly.

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Table 13.55 Varying Test Conditions

Gravity Gold Total Gold Residue Grade Calc Head Assay Head Test Variation Sample ID (%) Extrn (%) (Au g/t) (Au g/t) (Au g/t) WOX MC 32.91 94.43 0.11 1.98 1.68 50% Solids FRS MC 19.92 85.89 0.26 1.81 2.66 WOX MC 22.74 93.43 0.13 1.67 1.68 CIL FRS MC 23.54 89.81 0.27 2.23 2.66 WOX MC 26.41 89.67 0.18 1.74 1.68 AIR <9ppm FRS MC 40.51 89.45 0.34 3.22 2.66 AIR <9ppm, WOX MC 29.39 91.65 0.13 1.50 1.68 Min 150ppm CN FRS MC 18.83 86.22 0.30 2.18 2.66 O2 <20ppm, WOX MC 29.45 92.70 0.15 2.05 1.68 Min 150ppm CN FRS MC 10.41 85.54 0.25 1.73 2.66

What does appear to be advantageous is the use of oxygen. The kinetic data for these oxygen tests show improved initial gold leaching kinetics which is discussed further below.

As series of cyanide sensitivity tests were conducted using moderate to high initial NaCN concentrations of 250mg/L to 700mg/L. The results are summarised in Table 13.56. All

tests were conducted at 40% solids, P80 of 75µm, leach time of 48 hours with oxygen sparging. The various tests were conducted to allow the NaCN concentration to decay without further additions except those 500mg/L tests highlighted in yellow (Table 13.56), where additional NaCN was added to maintain a minimum cyanide concentration of 250mg/L.

Table 13.56 Cyanide Sensitivity – FRS, MOX and WOX Samples

Target Cyanide Gravity Gold Total Gold Residue Grade Sample ID (ppm) (%) Extraction (%) (Au g/t) FRS MC 250 21.33 88.00 0.23 FRS MC 500 11.88 87.08 0.25 FRS MC 750 16.78 88.56 0.23 FRS MC 500 14.83 85.73 0.26 M4 MOX 6 250 30.57 91.27 0.08 M4 MOX 6 500 35.49 91.13 0.09 M4 MOX 6 500 22.37 88.51 0.10 WOX MC 250 42.47 93.44 0.16 WOX MC 500 22.70 91.99 0.12 WOX MC 750 35.83 93.59 0.14 WOX MC 500 35.97 92.70 0.14

None of the 3 samples tested showed any benefit of elevated cyanide concentrations in the ranges tested. Consequently, lower cyanide concentrations were applied to the bulk of the remaining testwork.

From the data produced during the various testwork programs, it is possible to generate various graphs comparing the variables evaluated.

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Figure 13.15 suggests some potential benefit of oxygen but not elevated NaCN concentrations.

Figure 13.15 FRS Master Composite Oxygen and NaCN Sensitivity – 75µm

FRS MC Performance, 75 um, NaCN and Oxygen Sensitivity 1.20

1.00

0.80

0.60

Au Au Residue Grade g/t 0.40

0.20

0.00 0 5 10 15 20 25 30 Time, h

FRS MC Air 500 mg/L NaCN FRS MC Oxy 250 mg/L NaCN FRS MC Oxy 500 mg/L NaCN FRS MC Oxy 750 mg/L NaCN

Figure 13.16 is similar, but at a P80 of 90µm. The use of oxygen rather than air shows some small benefit. Importantly, the initial leach kinetics using oxygen is much faster than that using air as expected. This is advantageous if a tank is taken off-line in an operating plant, as the impact on losses will be less if oxygen is utilised. Also, the higher leaching rate means higher carbon loadings and lower elution rates for the same amount of recoverable gold. These are tangible operating advantages that also need to be considered when assessing the use of oxygen injection in the circuit.

Note FRS MC2 is a second master composite build up once the original FRS MC master composite was exhausted.

Figure 13.17 also considers the influence of grind size along with different air and oxygen

inputs. Kinetic data for the FRS MC at P80 of 150µm, 125µm, 106µm and 75µm are plotted. As anticipated, the gold recovery increases as the grind become finer. The P80 of 106µm line does not fit the sequence but is within the variability expected given the variability of the gravity gold component and the calculated head grade. The single case of oxygen injection

for the P80 of 75µm shows significantly lower residue grades that the other options and suggests that oxygen should be used when processing the FRS ore blends. A 0.05g/t leach benefit provides a nominal US$2/t revenue benefit at a gold price of US$1,200/oz, which is much higher than the cost of generating the oxygen.

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Figure 13.16 FRS Master Composite Oxygen and NaCN sensitivity – 90µm

FRS MC Performance, 90 um, NaCN and Oxygen Sensitivity 1.40

1.20

1.00

0.80

0.60 Au Au Residue Grade g/t 0.40

0.20

0.00 0 5 10 15 20 25 30 Time, h

FRS MC2 Air 300 FRS MC2 Oxy 300

Figure 13.17 FRS Master Composite Variable Sensitivity

FRS MC Performance, 500 mg/L NaCN, Grind and Oxygen Sensitivity 1.20

1.00

0.80

0.60

Au Au Residue Grade g/t 0.40

0.20

0.00 0 5 10 15 20 25 30 Time, h

FRS MC Air 150 FRS MC Air 125 FRS MC Air 106 FRS MC Air 75 FRS MC Oxy 75

Figure 13.18 presents data for the MOX MC sample. In this case a small improvement in gold recoveries can be observed if oxygen is used. However, the data is in effect inside the range of error/variation that could be expected and as such is not considered significant.

Figure 13.19 presents similar data for three WOX variability samples. The dotted lines represent those samples where air has been used compared to full lines of the same colour representing the same sample but with oxygen. This data supports the use of oxygen for these materials. This work suggests elevated oxygen has the potential to allow for reduced NaCN levels and still achieve improved leach kinetics and/or extractions.

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Figure 13.18 WOX Master Composite Oxygen and NaCN Sensitivity

WOX NaCN Sensitivity 1.20

1.00

0.80

0.60 Residue Residue grade,Au g/t 0.40

0.20

0.00 0 5 10 15 20 25 30 Time h

WOX MC, P80 : 75, NaCN mg/L 250 Oxy WOX MC, P80 : 75, NaCN mg/L 500 Oxy WOX MC, P80 : 75, NaCN mg/L 750 Oxy WOX MC, P80 : 75, NaCN mg/L 500 Air WOX TANMET MC, P80 : 75, NaCN mg/L 500 Air

Figure 13.19 WOX Master Composite Oxygen Sensitivity

WOX Oxygen Sensitivity 1.20

1.00

0.80

0.60 Residue Residue grade Au g/t

0.40

0.20

0.00 0 5 10 15 20 25 30 Time h

M1 WOX 1, P80 : 90, NaCN mg/L 250 Oxy M10 WOX 16, P80 : 90, NaCN mg/L 250 Oxy M10 WOX 17, P80 : 90, NaCN mg/L 250 Oxy M1 WOX 1, P80 : 90, NaCN mg/L 300 Air M10 WOX 16, P80 : 90, NaCN mg/L 300 Air M10 WOX 17, P80 : 90, NaCN mg/L 300 Air

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Table 13.57 presents a summary of leach data for 4 FRS composites including the kinetics over 48 hours (note that the extraction at time zero is equivalent to the gravity recovery achieved). Figure 13.20 presents the averages of the data for air and for oxygen.

Table 13.57 FRS Composite Air/Oxygen Data

Residue Oxygen Calc Head Extn 0h Extn 2h Extn 6h Extn 12h Extn 24h Extn 48h Sample ID Grade Supply (Au g/t) (%) (%) (%) (%) (%) (%) (Au g/t) M2 FRS 4 Air 1.52 0.290 7.9 60.0 71.1 78.2 80.5 81.0 M4 FRS 7 Air 1.80 0.255 16.4 36.0 76.9 78.9 82.0 85.8 M7 FRS 14 Air 2.61 0.380 16.7 57.2 78.3 81.7 84.9 85.4 M9 FRS 15 Air 4.64 0.340 48.7 59.7 83.5 87.5 91.5 92.7 M2 FRS 4 Oxygen 1.31 0.285 4.6 67.3 73.5 74.0 78.3 78.3 M4 FRS 7 Oxygen 2.28 0.245 28.1 78.2 83.0 86.5 88.3 89.2 M7 FRS 14 Oxygen 3.20 0.380 17.0 80.3 80.3 82.5 83.8 88.1 M9 FRS 15 Oxygen 5.23 0.210 54.7 91.1 91.6 92.4 97.8 96.0

Figure 13.20 FRS Composites – Averaged Data

Oxygen vs Air - FRS Composites 100.0 90.0 80.0 70.0 60.0 50.0 40.0 Extraction, Extraction, % 30.0 20.0 10.0 0.0 0 102030405060 Leach time, h

Air Oxygen

The averaged data showed that there will be a significant benefit using oxygen rather than only air during leaching. The data also shows that using oxygen will reduce the leach time to 24 hours compared to the 48 hours needed without oxygen.

The data from 68 leach tests were compared at 24 hour and 48 hour leach times. The average head grade for all these samples (which included all oxidation types) was 2.03g/t Au. The average extraction after 24 hours using air was 83.6% compared to the average gold extractions of 86.1% at 48 hours, a difference of 2.5%.

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As this data was not considered fully representative of oxidation type, an exercise was undertaken to separate master composite data based on oxidation type. Table 13.58 summarises the findings from 15 sets of samples where the data has been averaged for each oxidation type. Note that not all tests had oxygen injection and so these values are expected to be inflated with regard to the benefit of additional leach time. This exercise showed that difference in extraction between 24 and 48 hours of leach times is between 0.85% and 3.12% ,equivalent to 0.02g/t to 0.07g/t Au, depending on the head grade and source of material.

Table 13.58 Influence of Residence Time

Calc Head Residue Grade Extn 24h Extn 48h Difference Difference Sample ID (Au g/t) (Au g/t) (%) (%) (%) (Au g/t)

SOX 1.49 0.10 92.16 93.54 1.37 0.02 MOX 2.63 0.16 92.17 94.11 1.94 0.05 WOX 2.20 0.15 92.46 93.31 0.85 0.02 FRS 2.27 0.30 84.28 87.40 3.12 0.07

The benefit of additional gold recovery at 48 hours versus the additional capital and operating cost needs to be evaluated, however economic analysis shows that the use of oxygen, slightly higher NaCN concentrations and potential preferential grinding of higher SG particles in a full-scale plant will all be more economically beneficial than longer leaching times. Thus, it was decided that a 24 hour leach time would be more than adequate and consequently most of the later testwork was undertaken at 24 hour leaching times (selected design criteria). There is an opportunity to increase revenue during operations by increasing leaching times as well as decreasing or increasing the grind size depending on the blend of material treated.

13.15.2 Gravity Gold Recovery A gravity circuit in a full-scale plant is usually used to protect the downstream leaching facility from a high proportion of coarse gold. The leach rate of gold is such that in a typical CIL or CIP plant, partially leached coarse gold can report to the tailings, thus resulting in elevated gold residue grades.

The effectiveness of a gravity circuit is a function of the free gold size in the material processed and the size of such gold as it reports to the gravity unit operation. Conventional gravity circuits utilising centrifugal concentrators are effective in recovering this free gold down to around 100µm in size. After that, the efficiency decreases and higher proportions of free gold will report to the leach plant as the size decreases. The other key parameter is the mass pull of the concentrate. The higher the mass, the more chance coarse gold will report to the gravity concentrate.

Typically, metallurgical testwork overstates the gravity recovery as the testwork mass to concentrate (yield) is normally one order of magnitude greater than the full-scale plant scenario for a typical gravity/leach test.

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Section 13.8.3 presented the bulk gravity tests for SOX and MOX composites and Section 13.10.3 the bulk test results for the WOX and FRS composites from M5. A plant- scale free gold gravity recovery of <5% was suggested for SOX material and <10% for MOX material based on the bulk tests and extrapolating the data into the range of plant-scale gravity yields. Values of nominally 10% were indicated for WOX and <10% for FRS.

Figure 13.21 presents gravity recovery data along with the leach feed gold head grade as a function of the calculated head grade. Data was selected to eliminate duplication and the

test conditions were similar (nominal P80 of 90µm). There is only one SOX data point and two MOX data points. The rest of the data is predominantly WOX and FRS data. The gravity recovery data is the free gold component, that is, the gold that was collected from amalgamation of the gravity concentrate.

Figure 13.21 Select Gravity Recovery Data

Leach Feed Grade and Gravity Recovery

4.50 70.0 Lch Fd Head SOX MOX 4.00 WOX FRS 60.0 Log. (Lch Fd Head) Log. (FRS) 3.50

50.0

3.00

y = 1.2181ln(x) + 0.7785 R² = 0.6635 40.0 2.50

2.00

30.0 Gravity Recovery, % Leach Feed Hd Grade Au g/t

1.50

20.0

1.00

10.0 0.50

0.00 0.0 0.00 1.00 2.00 3.00 4.00 5.00 6.00 7.00 Calculated Hd Grade Au g/t

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As the sample masses were in the 1kg and 2kg range, the mass pull as a percentage of the feed is of the order of 3% to 6% of the new feed mass.

The non-linear relationship of the leach feed head grade to the calculated head grade is reasonably strong. This suggests a reasonably consistent gravity recovery as a function of head grade. However, data scattering occurs when the calculated head grades exceed 2g/t Au.

The range of WOX and FRS gravity recovery is similar to the MOX data in this dataset.

The bulk sample gravity testwork results for free gold have previously been described. The graphs summarising the data are repeated as Figure 13.22, Figure 13.24, Figure 13.26 and Figure 13.28 for SOX, MOX, WOX and FRS respectively. In addition, the free plus intensive leached gold has been presented as Figure 13.23, Figure 13.25, Figure 13.27 and Figure 13.29 for SOX, MOX, WOX and FRS respectively, as a function of mass pull to concentrate. The free gold plus intensively leached gold component is reflective of the full-scale plant flowsheet where the gravity concentrate is intensively leached.

Figure 13.22 SOX Free Gold Recovery

SOX TANMET MC Gravity Performance - Free Gold 30.00%

25.00%

20.00%

15.00%

y = 0.0283ln(x) + 0.2238 R² = 0.9897 10.00% %Recovery of gold toamalgam

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

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Figure 13.23 SOX Free + Intensive Leach Gold Recovery

SOX TANMET MC Gravity Performance - Includes Intensive Leach 30.00%

25.00%

20.00% y = 0.0573ln(x) + 0.4313 R² = 0.9937

15.00%

10.00%

5.00% % Recovery of gold to amalgam + Intensive Intensive ofamalgam +Recovery Leach to% gold

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

Figure 13.24 MOX Free Gold Recovery

MOX TANMET MC Gravity Performance - Free Gold 30.00%

25.00%

20.00% y = 0.0195ln(x) + 0.2359 R² = 0.9827

15.00%

10.00% % Recovery Recovery % of gold toamalgam

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

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Figure 13.25 MOX Free + Intensive Leach Gold Recovery

MOX TANMET MC Gravity Performance - Includes Intensive Leach 30.00%

25.00%

y = 0.0462ln(x) + 0.4787 R² = 0.9926 20.00%

15.00%

10.00%

5.00% % Recovery of gold toamalgam +Intensive Leach

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

Figure 13.26 WOX Free Gold Recovery

WOX MC Gravity Performance - Free Gold 30.00%

25.00%

20.00% y = 0.0155ln(x) + 0.2194 R² = 0.9979

15.00%

10.00% % Recovery of % Recovery gold to amalgam

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

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Figure 13.27 WOX Free + Intensive Leach Gold Recovery

WOX MC Gravity Performance - Includes Intensive Leach 80.00%

70.00%

60.00%

y = 0.0921ln(x) + 1.0956 R² = 0.9975 50.00%

40.00%

30.00%

20.00% % Recovery of gold to amalgam % Recovery of gold toamalgam + intensive leach 10.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

Figure 13.28 FRS Free Gold Recovery

FRS MC Gravity Performance - Free Gold 30.00%

25.00%

20.00%

y = 0.0149ln(x) + 0.1894 R² = 0.9966 15.00%

10.00% % Recovery of gold to amalgam gold to of % Recovery

5.00%

0.00% 0.00% 0.20% 0.40% 0.60% 0.80% 1.00% 1.20% 1.40% Mass pull to concentrate

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Figure 13.29 FRS Free + Intensive Leach Gold Recovery

The mass pull is still an order of magnitude higher than the full-scale plant value of around 0.03%. However, the additional gold that can be recovered by leaching is identified when the graphs are compared. The most notable increase was observed for the WOX material.

Figure 13.29 includes a blue curve and a red curve added to the graph starting at the origin. These curves represent two of the many options which may represent the shape of the curve at mass pull values of less than 0.4%. The red curve suggests that the full-scale plant recovery could be anywhere between 10% and 20% (on a 2g/t head grade), reflecting the level of error possible when extrapolating this data. Furthermore, whereas the earlier Figure 13.21 suggested that the MOX, WOX and FRS samples had similar free gold content, the domain specific curves present a wider range of variation.

With regard to the design of the process plant, the gravity gold operations can influence the carbon loadings, electrowinning of the intensive leach component and the mass flows from the gravity concentrator to the intensive leach unit. The full-scale plant has been designed such that the carbon circuit can cater for the total gold flow if the gravity circuit is not operational and so actual gravity recovery accuracy is not critical. The mass pull is fixed and so the actual recovery with regard to the physical sizing of the intensive leach unit is not relevant. The intensive leach electrowinning is driven by the actual amount of gold and so this must be considered.

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The previous figures suggest that the gravity recovery at full-scale operation will be in the order of 20%. Earlier work suggested that the more oxidised samples have lower gravity gold component, probably due to sulphide release on oxidation and reduction in the number of complex sulphide composites.

This all indicates that the range of gravity recovery values in the full-scale operation can be expected to be as follows:

. SOX: 5% to 10%. . MOX, WOX and FRS: 10% to 20% with a design consideration for electrowinning (FRS) at the top of the range.

13.15.3 Reagent Consumptions Figure 13.30, Figure 13.31, Figure 13.32 and Figure 13.33 present the reagent consumption data for a selection of gravity/leach tests for SOX, MOX, WOX and FRS, respectively. In general, conservative values have been selected for use in the process design and as support for operating costs.

The NaCN consumptions are generally low compared to similar operations. This is supported by the mineralogy of the materials which shows few cyanicides present. The low consumptions are also supported by the low NaCN concentrations required to achieve adequate gold leaching rates.

Figure 13.30 SOX Reagent Consumption

SOX NaCN and Quicklime Consumption 3.50

3.00

2.50

2.00

1.50 Consumption, kg/t

1.00

0.50

0.00 NaCN kg/t Lime kg/t Reagents

MET 3A SOX TANMET MC SOX 5-1 SOX 6-2 SOX 7-3 SOX 8-1 SOX 10-1

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Figure 13.31 MOX Reagent Consumption

MOX NaCN and Quicklime Consumption 3.50

3.00

2.50

2.00

1.50 Consumption, kg/t

1.00

0.50

0.00 NaCN kg/t Lime kg/t Reagents

M4 MOX 6 MOX 5‐3 MOX 6‐4 MOX 6‐6 MOX 10‐2 MOX TANMET MC

Figure 13.32 WOX Reagent Consumption

WOX NaCN and Quicklime Consumption (250 and 330 mg/L NaCN target) 1.2

0.8

0.6

0.4 Consumption, kg/t

0.2

0 NaCN kg/t Lime kg/t Reagents

M1 WOX 1 M5 WOX 8 M10 WOX 16 M10 WOX 17 M10 WOX 19 WOX MC M1 WOX 1 M10 WOX 16 M10 WOX 17

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Figure 13.33 FRS Reagent Consumption

FRS NaCN and Quicklime Consumption (250 and 330 mg/L NaCN target) 0.7

0.6

0.5

0.4

0.3

Consumption, 0.2 kg/t

0.1

0 NaCN kg/t Lime kg/t Reagents

FRS MC M2 FRS 5 M4 FRS 7 M5 FRS 10 M5 FRS 11 M5 FRS 9 M6 FRS 12 M6 FRS 13 FRS MC M2 FRS 4 M4 FRS 7 M6 FRS 12 M7 FRS 14 M9 FRS 15

A final NaCN concentration in the CIL tails was assumed to be 200mg/L in line with conservative conventional operating practice (often values of 100mg/L to 150mg/L are utilised). It was assumed that the tails would be discharging at 42% solids w/w and various pulp densities were applied for the various material types discharging the tails thickener. A credit of 30% of the recycled thickener overflow NaCN was applied to the consumption value. The final consumption as applicable to a full-scale plant was then estimated. These NaCN consumption values along with an allowance for quicklime based on the data are summarised in Table 13.59.

Table 13.59 Reagent Consumption Considerations

NaCN Final NaCN NaCN Quicklime Consumption CIL Tail PD Tails PD Sample Concentration Consumption Consumption Testwork % w/w %w/w mg/L Full-Scale kg/t Full-Scale kg/t kg/t

SOX 0.25 200 42 55 0.20 2.5 MOX 0.25 200 42 55 0.20 1.1 WOX 0.2 200 42 58 0.15 0.65 FRS 0.2 200 42 58 0.15 0.5

13.15.4 Silver Metallurgy Silver metallurgy has not been analysed in detail.

Silver grades have been reported in the mineralogy to be around 10% of the gold grade for coarse grains. The leaching work would suggest that the silver grade is somewhat higher and in the range of 0.5g/t based on calculated head grades. Gravity recovery of the silver is around 10%, suggesting practical plant recovery will be of the order of 5% or lower. Leach extraction is of the order of 50%.

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Part of the issue in analysing the silver behaviour is the limitation on the assay resolution. Assay head resolution is 0.3g/t and residue grade resolution is 0.15g/t. These silver assay resolutions add significant error to any determinations including the values noted in the previous paragraph.

Silver takes up sites on the carbon otherwise occupied by gold and silver requires twice the required current as gold for electrowinning on a mass to mass basis. Silver is therefore important when designing a carbon-based recovery plant including the elution and electrowinning systems, and an allowance for silver needs to be made in the design of these facilities.

13.15.5 Sulphide Influences and Leach Feed Head Grade The testwork provided some examples where certain samples reported elevated leach residue grades. Diagnostic leaching suggested such high residues were due to gold locked in aqua regia soluble minerals. Mineralogical studies identified free gold grains associated with sulphide minerals. High residue grades may therefore be a function of gold locked in sulphides especially the FRS oxidation type.

Eight FRS composites representing different areas of the M5 deposit had been subjected to leaching at both 106µm and 75µm. The sulphide assay of these composites was between 0.4% to 2.36%. Table 13.60 presents the summarised data.

Table 13.60 Influence of Sulphides

P80 Grind Au Extraction Consumption Au Grade (g/t) Size (µm) (%) (kg/t) Composite Sulphide ID Calc. Grav. 2hrs 6hrs 24hrs Tail NaCN Lime Head 106 35.0 81.0 88.6 90.5 3.88 0.37 0.27 0.63 2.36 M11 FRS30 75 30.6 66.8 81.4 92.6 3.97 0.30 0.22 0.63 2.36 106 45.5 62.1 79.4 88.9 2.03 0.23 0.19 0.3 1.58 M11 FRS31 75 52.8 76.0 90.0 94.1 2.1 0.13 0.11 0.4 1.58 106 28.3 48.4 69.8 80.7 3.39 0.66 0.21 0.35 1.9 M12 FRS32 75 29.0 60.3 78.0 79.6 2.74 0.56 0.19 0.38 1.9 106 26.0 66.7 88.5 92.1 6.27 0.50 0.14 0.39 1.8 M12 FRS33 75 62.3 74.1 91.3 97.1 10.2 0.30 0.14 0.47 1.8 106 31.4 76.8 90.9 92.4 2.71 0.21 0.14 0.28 0.98 M12 FRS34 75 46.9 74.4 93.6 96.2 3.41 0.13 0.11 0.41 0.98 106 16.7 74.7 82.1 85.2 2.16 0.32 0.11 0.3 0.86 M13 FRS35 75 18.2 71.6 86.5 87.7 2.36 0.29 0.19 0.28 0.86 106 12.1 39.5 72.3 88.6 8.11 0.93 0.11 0.3 0.92 M13 FRS36 75 19.0 46.7 80.0 91.4 11.5 1.00 0.16 0.26 0.92 106 47.3 84.0 89.8 92.7 2.82 0.21 0.08 0.35 0.44 M15 FRS38 75 50.4 87.5 92.3 93.2 2.14 0.15 0.11 0.31 0.44 106 15.3 77.1 84.6 85.4 4.73 0.69 0.11 0.49 0.4 M16 FRS39 75 18.6 80.9 83.3 88.0 4.84 0.58 0.11 0.52 0.4

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The residue grades at 106µm and 75µm were plotted as a function of the sulphide level and this data is presented as Figure 13.34. There is no obvious relationship between residue grade and sulphide head grade. Other plots of % Extraction vs Sulphide Head Grade and a series of other experimental application of data similarly showed no relationship.

Figure 13.34 Grind Sensitivity and Sulphide Relationship

Residue grade and sulphide influence

1.2

Res 106 um Res 75 um 1

0.8

0.6

0.4 Residue Grade Au Residue Grade Au g/t 0.2

0 00.511.522.5 Sulphide sulphur %

An exercise was undertaken where the residue grade was estimated as a linear function of calculated gold head grade, leach feed head grade and sulphide content. The leach feed grade is the calculated head grade less the gravity recovered content. The observations are presented in Figure 13.35, Figure 13.36, Figure 13.37 and Figure 13.38.

When the calculated head grade is multiplied by a factor (around 0.1 depending on which case is being explored) to estimate a leach residue, a line of best fit can be generated for

the P80 106µm material as shown in Figure 13.35. If a similar exercise is made by adding a residue grade component by factoring the sulphide content, the regression accuracy decreases (R2 value decreases. No graph of this sulphide influence is presented). This again suggests that the sulphide content does not have a linear relationship with residue grade.

If the residue grade is estimated as a ratio of the leach feed grade, the regression is stronger with an increased R2 value, as presented in Figure 13.36 compared to Figure 13.35.

These same relationships were repeated using the results from the 75µm tests where the leach feed head grade was shown to be a more reliable estimator of residue grade than the calculated head grade. (see Figure 13.37 and Figure 13.38).

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Figure 13.35 Grind Sensitivity, FRS Master Composite (106µm)

Residue Estimation - 106 um grind, Calc Head 1.2

1 y = 0.9938x R² = 0.6091

0.8

0.6

0.4

0.2

Estimated Estimated Grade of Residue, Au g/t 0 0 0.2 0.4 0.6 0.8 1 Actual Grade of Residue, Au g/t

Figure 13.36 Grind Sensitivity, FRS Master Composite (106µm)

Residue Estimation - 106 um grind, Leach Head 1.2

1 y = 0.9834x R² = 0.7565 0.8

0.6

0.4

0.2

Estimated Estimated Grade of Residue, Au g/t 0 00.20.40.60.81 Actual Grade of Residue, Au g/t

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Figure 13.37 Grind Sensitivity, FRS Master Composite (75µm)

Residue Estimation - 75 um grind, Calc Head 1.2 y = 0.9964x 1 R² = 0.3316

0.8

0.6

0.4

0.2

Estimated Estimated Grade of Residue,g/t Au 0 0 0.2 0.4 0.6 0.8 1 1.2 Actual Grade of Residue, Au g/t

Figure 13.38 Grind Sensitivity, FRS Master Composite (75µm)

Residue Estimation - 75 um grind, Leach Head 1.4

1.2 y = 0.9992x 1 R² = 0.7802

0.8

0.6

0.4

0.2

Estimated Estimated Grade of Residue, Au g/t 0 0 0.2 0.4 0.6 0.8 1 1.2 Actual Grade of Residue, Au g/t

The relationships, in line with the earlier mineralogy, suggest the following:

. Whilst there is a probability that sulphides contain some gold not fully liberated for leaching and may contribute to elevated residue grades at times, the behaviour and influence is not consistent. High sulphide grades do not necessarily result in high residue grades. . Gravity gold recovery is inconsistent and not necessarily related to head grade. Free gold continues to be liberated as the grind size is reduced. The gravity gold component of the SOX material is lower than the other oxidation states represented. . The residue gold grade appears to be associated with the leach feed gold head grade, particularly when compared to the actual gold head grade (calculated head grade) prior to removal of coarse/free gold. The distribution of gold and residual gold throughout the sample would be more consistent when the coarse/free gold is removed.

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Leach Recovery Algorithms A series of leach recovery algorithms were developed for input to the mining reserve estimation and subsequent financial modelling. Separate models were developed for SOX, MOX, WOX and FRS.

Various relationships were explored using the calculated gold head grade and the leach feed gold head grade as inputs. Linear, power based and logarithmic relationships were trialled to assess the most effective representation of the data.

13.16.1 Methodology

All data relevant to the proposed flowsheet of nominal P80 of 90µm grind, gravity concentration followed by a 24 hour CIL with oxygen sparging was used for each oxidation state. Some relevant data from tests where higher NaCN concentrations and some data from tests at 75µm grind size were also used. The data was analysed and various relationships explored. The influence of sulphide level was not part of this evaluation.

Evaluation of the calculated head grade and the leach feed grade versus the gold residue grade was explored in more detail. Figure 13.39 and Figure 13.40 show some of the relationships for FRS ore as an example. The trends are similar for MOX and WOX material.

Figure 13.39 Residue Grade as a Function of Calculated Head Grade - FRS

Residue Grade as a function of Calculated Head Grade - 24 hr 1.20

1.00

0.80

y = 0.1766x0.5007 R² = 0.2701 0.60 Residue Residue Grade, Au g/t 0.40

0.20

0.00 0.00 2.00 4.00 6.00 8.00 10.00 12.00 Head Grade, Au g/t

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Figure 13.40 Residue Grade as a Function of Leach Feed Grade - FRS

Residue Grade as a function of Leach Feed Grade - 24 hr 1.20

1.00 y = 0.1746x0.8022 R² = 0.503

0.80

0.60 Residue Residue Grade, Au g/t 0.40

0.20

0.00 0.00 1.00 2.00 3.00 4.00 5.00 6.00 7.00 8.00 Head Grade, Au g/t

As a result, the gold leach feed grade was used as the initial basis for development of the algorithms.

There are various models available to explain or predict the gold recovery. A line of best fit relationship such as presented in Figure 13.39 can be used, however the lower and higher grades are often not well represented by the relationship. Other methods such as fixed residue grade can be used but are not normally accurate as they ignore head grade fluctuations. Fixed recovery methods are sometimes employed but again due to gold head grade influences they are usually inaccurate and only suited to scoping level evaluations at best.

Another method is the use of different ratios of the gold residue grade as a function of the gold head grade and as a function of residence time. Residence time is included to model the influence of the rate (shape) of the leach curve which is often influenced by the head grade.

The leach kinetics were represented by the ratio of the residue gold grade as a function of residence time. This method inherently includes the influence of gold head grade variations. A family of curves was generated, an example of which is given in Figure 13.41. The curves follow a general pattern and the power relationships fitted similarly as the leach feed gold grade changes.

The analysis then models the change in curve shape as a function of the leach feed grade, which in turn allows an algorithm to be derived that uses the leach feed grade and the leach time as inputs to predict the residue ratio and subsequent residue grade itself.

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Figure 13.41 FRS Family of Curves – Residue Ratio

Residue Ratio - Lch Fd M2 FRS 5 NaCN 250 mg/L Oxy 0.60 M4 FRS 7 NaCN 250 mg/L Oxy

M5 FRS 9 NaCN 250 mg/L Oxy

M5 FRS 10 NaCN 250 mg/L Oxy

0.50 M5 FRS 11 NaCN 250 mg/L Oxy

M6 FRS 12 NaCN 250 mg/L Oxy

y = 0.3791x-0.154 M6 FRS 13 NaCN 250 mg/L Oxy R² = 0.9484

M9 FRS 15 NaCN 250 mg/L Oxy y = 0.4635x-0.227 0.40 y = 0.7024x-0.458 R² = 0.9073 R² = 0.9391 FRS MC NaCN 250 mg/L Oxy

M2 FRS 4 NaCN 300 mg/L Oxy

y = 0.2604x-0.084 R² = 0.8199 M7 FRS 14 NaCN 300 mg/L Oxy

0.30 Power (M2 FRS 5 NaCN 250 mg/L Oxy)

Power (M4 FRS 7 NaCN 250 mg/L Oxy) Residue Residue ratio

Power (M5 FRS 9 NaCN 250 mg/L Oxy)

Power (M5 FRS 10 NaCN 250 mg/L Oxy) 0.20 Power (M5 FRS 11 NaCN 250 mg/L Oxy)

y = 0.5524x-0.398 Power (M6 FRS 12 NaCN 250 mg/L Oxy) R² = 0.9308

y = 0.5771x-0.389 Power (M6 FRS 13 NaCN 250 mg/L Oxy) R² = 0.9322 y = 0.5087x-0.441 y = 0.5771x-0.425 0.10 R² = 0.9036 Power (M9 FRS 15 NaCN 250 mg/L Oxy) R² = 0.7987 y = 0.506x-0.32 R² = 0.9085 y = 0.3074x-0.476 Power (FRS MC NaCN 250 mg/L Oxy) R² = 0.8883 y = 0.4206x-0.328 Power (M2 FRS 4 NaCN 300 mg/L Oxy) R² = 0.8605

Power (M7 FRS 14 NaCN 300 mg/L Oxy) 0.00 0 5 10 15 20 25 30 Time

This algorithm allows for variation in leach times to be interpreted apart from the 24 hour case. This allows for modelling tanks off-line, the influence of extended leach times or differential throughputs if required.

This method when applied to the Sanbrado data produced relationships which more accurately predicted the gold residue grade than other methods that were evaluated.

The model can be checked by comparing the actual gold residue grade with the modelled outcome. Results of the FRS algorithm are presented in Figure 13.42. Apart from one obvious outlier, the correlation is considered acceptable.

The algorithm is based on the leach gold feed head grade, and an estimate of the leach feed head grade from a given calculated gold head grade is required.

Referring to Figure 13.21 in Section 13.16.2, a line of best fit for the FRS gravity recovery/leach feed grade can be applied to estimate the leach feed grade. This estimation adds error to the algorithm derivation as the fit is not as robust as the actual residue estimate based on leach feed grade alone. The very reason for using the leach feed grade in the first instance is to remove the “noise” associated with the variable gravity recovery.

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Figure 13.42 FRS Algorithm Actual versus Modelled Leach Feed Grade

FRS Algorithm - M5 FRS Data Leach Feed Head Grade Input 1.20

1.00

y = 0.9664x R² = 0.8166 0.80

0.60

0.40 Algorithm Algorithm Residue, t =24 h, Au g/t 0.20

0.00 0.00 0.20 0.40 0.60 0.80 1.00 1.20 Test Work Residue at t = 24 h Au g/t

Figure 13.43 presents the algorithm once the calculated head grade – leach feed grade adjustment is made. The correlation is not as strong as shown in Figure 13.42 due to the influence of variable gravity recoveries, particularly that of composite M9 FRS 15 which had a 6.38g/t Au calculated head grade but a nominal 50% gravity recovery.

Figure 13.43 FRS Algorithm Actual versus Modelled Calculated Head Grade

FRS Algorithm - M5 FRS Data Calculated Head Grade Input 0.700

0.600

y = 1.0106x 0.500 R² = 0.0368

0.400

0.300

0.200 Algorithm Residue t=24 h, Au g/t Au h, t=24 Residue Algorithm

0.100

0.000 0.00 0.10 0.20 0.30 0.40 0.50 0.60 Test Work Residue 24 h Au g/t

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Whilst such outliers exist in most of the datasets processed, they were retained in the data analysis. Such outliers do not reflect the norm, but are still valid data points in that they influence the variability of response.

Once the general form of the algorithms was defined, manipulation was required to allow for full-scale plant inefficiencies. For each of SOX, MOX, WOX and FRS, the gold extraction was estimated by applying head gold grades to the algorithms selected and estimating a gold residue grade. A soluble loss typical of a CIL operation of 0.01g/t Au in tailings thickener underflow was applied to give the predicted gold recovery. To simplify matters for mine resource and financial modelling, the percentage recovery as a function of gold head grade was calculated and equations fitted to this data for application by others.

As the SOX and MOX datasets were limited, the curves were adjusted to align with the actual leach data using the same form of equation, giving matches within 1% of the testwork. The WOX and FRS datasets were adequate to provide curves that did not require re-alignment.

Whilst the FRS algorithm provided a testwork to estimate ratio of 1:1, the WOX model ended up with a bias of 1: 1.1 as is shown in Figure 13.44. That is, the algorithm presents a 10% higher residue grade than actual. Given the limited dataset and the conservative estimate output, this was accepted as the basis for algorithm finalisation.

Figure 13.44 WOX Algorithm Actual versus Modelled Calculated Head Grade

WOX Algorithm - Residue Grade Estimate 0.300

y = 1.1035x 0.250

0.200

0.150

0.100 Algorithm Algorithm Residue, g/tAu

0.050

0.000 0.00 0.05 0.10 0.15 0.20 0.25 0.30 Test Work Residue, Au g/t

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13.16.2 Final Algorithms A separate M1 FRS zone algorithm was derived based on the M5 FRS algorithm to take into account the higher leach extractions at lower grades exhibited by the M1 testwork data.

To improve the fit of the algorithms, the final percentage based versions applied to resource development and financial modelling were broken into grade bands for the M1 and M5 FRS only. The grade bands used were 0.4g/t to 1.0g/t, 1.01g/t to 3.0g/t and 3.01g/t to 20g/t for M5 or 3.01g/t to 30g/t for M1. Recovery was fixed at 97.3% for M5 for grades over 20g/t and at 96% for M1 for grades over 30g/t.

For SOX, MOX and WOX, the algorithms were considered applicable for the grade range 0.1g/t to 5.0g/t, with maximum recoveries applied to grades in excess of 5.0g/t.

The algorithms (at various gold head grades) utilised in the final reserve estimation are as follows:

. SOX

 0.1g/t to 5.0g/t: (0.015* LN(Au g/t head) + 0.93)*100%

 5.0g/t: 95.4% . MOX

 0.1g/t to 5.0g/t: (0.020* LN(Au g/t head) + 0.915)*100%

 5.0g/t: 94.7% . WOX

 0.1g/t to 5.0g/t (0.0364* LN(Au g/t head) + 0.8843)*100%

 +5.0g/t: 94.3% . FRS M5

 0.4g/t to 1.0g/t: (0.1237 * LN(Au g/t head) + 0.8158)*100%

 1.01g/t to 3.0g/t: (0.075 * LN(Au g/t head) + 0.8158)*100%

 3.01g/t to 20g/t: (0.0396 * LN(Au g/t head) + 0.8542)*100%

 +20g/t: 97.3%. . FRS M1

 0.4g/t to 1.0g/t: (0.0589 * LN(Au g/t head) + 0.8567)*100%

 1.01g/t to 3.0g/t: (0.0388 * LN(Au g/t head) + 0.8574)*100%

 3.01g/t to 30g/t: (0.0255 * LN(Au g/t head) + 0.8721)*100%

 +30g/t: 96%. The algorithms presented provide an estimate of the recovery for each material type within the range of accuracy applicable to the feasibility study. It is appreciated that the variable gravity gold recovery/deportment as well as potential variable sulphide deportment will result in deviations between testwork data and the algorithm forecasts.

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Conclusions The Sanbrado materials tested present as free-milling providing high gold leach extractions, medium gold leaching rates with relatively low reagent consumption. The liberation size required is typical for such ores and a conventional grinding flowsheet applies. Moderate levels of free gold can be recovered using conventional centrifugal type concentrators.

The free gold component is not particularly coarse and is amenable to gravity recovery. The bulk of the non-free gold is contained in non-sulphide minerals as composite or complex particles. Some gold is associated with sulphides which can be expected to result in variable leaching rates due to un-liberated gold and possibly lock-up in final residues.

Composite samples were developed to reflect a spread of head grades, mineralogy, host lithology as well as spatial and physical location. There does not appear to be any significant difference in metallurgical performance as a function of location or lithology. However, there are differences in performance for the four metallurgical (oxidation state) domains identified, namely SOX, WOX, MOX and FRS.

The materials do not show high oxygen demands nor do they present problematic rheological characteristics. Thickening testwork was successful but required high rates of flocculant, however the flocculant used in the testwork is considered sub-optimal and there is opportunity to improve in this area.

SAG milling characterisation testwork suggests that the FRS ore is very hard/tough and energy intensive. The milling circuit design will need to consider the practicality of blending ore types and the risk profile associated with single stage SAG milling on such materials.

Further Testwork The various testwork programs have provided the necessary input to support the feasibility study. There are some areas of additional work recommended which will be included in the Definitive Feasibility Study technical report due for completion in mid-2018. These include the following:

. Additional flocculant screening and repeat dynamic thickener testing is warranted due to the opportunity to save capital and operating costs. In addition, a series of blends of material types should be tested to attempt improvement in the oxide material performance. . Carbon characterisation work. Current testwork data presents very high carbon loadings which should be confirmed as valid. . Leach kinetic testwork such as cyanide consumption optimisation and oxygen mass transfer should be undertaken to increase gold leach kinetics and gold recoveries at P80 of 90µm and 24 hour leaching time. . The M3 deposit has not been tested metallurgically. Geologically it is considered to present the same characteristics as M5 and M1, however a small number of scoping level tests would be prudent.

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. Any additional infill DC drilling should have the standard gravity/leach flowsheet testwork undertaken for completeness. . Adjustment of recovery algorithms should be undertaken as any new data comes to hand. . Materials handling testwork was originally planned, however the sample allocated was used for tailings characterisation work. Note this aspect is not considered critical given the flowsheet selected. . Investigation of metallurgical responses of select drillhole intervals used to make up composites that gave lower leach extractions. Design Criteria Table 13.61 presents a summary of proposed key design criteria based on the testwork results. Comminution data are taken from the OMC report and is reported in Section 17.

Table 13.61 Design Values Based on Testwork

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MINERAL RESOURCE ESTIMATES Introduction This section is presented in two parts, the first relates to the mineral resource updates for M1 South and M5 as at the effective date of the report and the second relates to the previous mineral resource estimates as at February 20th 2017. The previous mineral resource estimate is described in mineral reserve estimates and subsequent sections of the report are based on the February 20th 2017 mineral resource and have not been updated to reflect the updates to the M1 South and M5 mineral resources. This work is in progress and is expected to be completed in mid-2018.

Mineral Resource Estimates 13th December 2017 This Mineral Resource for the Sanbrado Gold Project has been estimated as at the effective date of the 13th December 2017. Grade estimates were updated for M1 South and M5 (Figure 14.1). Grade estimates for M1 North and M3 have not been updated and have been described in the report entitled “NI 43-101 Technical Report: - Open Pit Feasibility Study, Sanbrado Gold Project, Burkina Faso “ and dated 6th April 2017. Gold grade estimation was completed using a combination of Multiple Indicator Kriging (MIK) and Ordinary Kriging (OK). MIK grade estimates have been localised to the SMU dimension using an analogous methodology to Localised uniform Conditioning. This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralisation, and the style, geometry and tenor of mineralisation. The estimation was constrained with geological and mineralisation interpretations.

14.2.1 Database Validation The resource estimation was based on the available exploration drillhole database which was compiled in-house by WAF. The database was reviewed and validated prior to commencing the resource estimation study.

The database includes assay results from trenching, AC, RC and DC samples. Trenching data was excluded from the database for the purposes of modelling and grade estimation. The resultant database was extensively validated. Checks made to the database prior to use included the following:

. Check for overlapping sample intervals. . Check downhole surveys at 0m depth. . Review consistency of depths between different data tables. . Check for any gaps in the data. . Replace less than detection assays with half detection limit. . Replace intervals with no sample with -999. . Replace intervals with assays not received with -1000.

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Figure 14.1 Plan View of Drilling

14.2.2 Summary of Data Used in Estimate The area of the M1 South Mineral Resource was drilled using AC, RC and DC drillholes and RC drillholes with a diamond tail (DT) on a nominal 25m by 25m grid spacing or closer. A total of 217 AC holes (3,290m), 42 DC holes (12,041m), 52 DT holes (14,113m) and 115 RC holes (12,602m) were drilled by WAF between 2015 and 2017. A total of 27 RC holes (3,378m) and seven DC holes (1,199m) were drilled by Channel in 2003 and between 2010 and 2012. Hole azimuths were either 225° or 180° magnetic at declinations of between - 50° and -60°, to optimally intersect the mineralised zones. A small number of drillholes had azimuths of approximately 020° or 045° due to restricted rig access.

The area of the M5 Mineral Resource was drilled using AC, RC and DC drillholes on a nominal 50m by 25m grid spacing with infill to 25m by 25m in the far south portion. A total of 723 AC holes (23,314m), 71 DC holes (14,537m), 41 DT holes (8,739m) and 69 RC holes (7,078m) were drilled by WAF between 2013 and 2017. A total of 60 RC holes (7,296.2m) and 71 DC holes (15,439.6m) were drilled by Channel between 2010 and 2012. Hole azimuths were 120° or 300° magnetic at declinations of between -50° and -60°, to optimally intersect the mineralised zones.

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14.2.3 Interpretation and Modelling Geological Interpretation Oxidation and lithology boundaries were interpreted based on grade information and geological observations, resulting in modelled wireframes used to constrain resource estimations. Interpretation and digitising of all constraining boundaries was undertaken on cross sections at the appropriate orientation for the relevant deposit (drill line orientation). The resultant digitised boundaries were used to construct wireframe surfaces or solids defining the three-dimensional geometry of each interpreted feature.

Mineralisation Interpretation Gold mineralisation in the project area is mesothermal orogenic in origin and is entirely hosted within sheared and deformed pelitic and psammitic meta-sediments and volcanic sediments. The sedimentary sequence has been intruded by a granodiorite which is generally parallel or sub-parallel to the main shear direction. Gold occurs with quartz vein and veinlet arrays. Gold mainly occurs as free gold associated with pyrite, pyrrhotite and minor arsenopyrite disseminations and stringers. Observed alteration includes silica, sulphide, carbonate-albite and tourmaline-biotite alteration.

Mankarga 1 Mineralised domains were constructed on approximately 25m spaced cross sections orientated perpendicular to drilling using a nominal 0.3g/t Au edge cutoff to define overall shear zone mineralisation. Good continuity of the mineralised envelopes was observed at the scale of the sectional spacing with moderate grade variability typical for deposits of this type.

This interpretation was designed to capture the broad mineralisation halo that encompasses the geological vein system and was not intended to constrain individual veins or vein clusters. This interpretation is appropriate for the MIK with a change of support grade estimation technique for the major domains, and for the proposed open pit mining method.

Mineralisation at M1 South exists in a similar setting to, and may be structurally offset from, M1 North. The observed mineralisation envelope is, however, more diffuse and less well defined, possibly indicating a structurally complex setting. Mineralisation has been interpreted to occur in discrete blocks which have been faulted and subsequently rotated from the plane of the main structure (Figures 14.1 and 14.2). Additionally, extremely high grade intervals occur within the overall mineralised envelope which variably demonstrate good to poor continuity.

In addition to the broad mineralised domains that were constrained using a nominal 0.3g/t Au edge cutoff (Zone 100 to Zone 600 and Zone 900), a number of high grade mineralised domains largely internal to the broader domains were interpreted (Zone 1000 to Zone 5000, Figure 14.3). Gold grades in excess of 50g/t have been observed and grade continuity was generated at a higher cutoff, nominally 10g/t Au. However, within the higher grade intercepts the internal grades can fall below 10g/t Au. In these cases, the lower cutoff grade was lowered to enforce grade continuity. Continuity of extremely high grade zones has improved with a tighter drillhole spacing since the previous grade estimate.

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Figure 14.2 Isometric NE View of M1

Figure 14.3 M1 South Section SE 0450

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Mankarga 5 (M5) Mineralisation at M5 largely lies within a broad, steeply dipping low grade shear zone corridor that can be traced over a strike length of 3,000m (Figure 14.1). The overall strike direction of the shear zone is approximately 035° and is sub vertical to steeply west dipping (Figure 14.4). Distinctive higher grade hangingwall and footwall lodes can be identified within this broader lower grade halo.

Mineralisation interpretations were subsequently reviewed in multiple orientations and in plan and section view prior to being accepted for grade estimation and block modelling purposes. Three main mineralisation estimation domains (Zones 100, 200, 300) were defined, with a fourth domain (Zone 400), to the south (Figure 14.5). Zone 102 is a minor sub-domain located adjacent to and in the hangingwall of Zone 100. Previous resource estimations included another sub-domain (Zone 101), interpreted as a potential strike extension of Zone 100. This sub-domain has now been included in Zone 100. Zone 301 is a minor domain located in the footwall of Zone 300.

14.2.4 Data Flagging and Compositing Drillhole samples were flagged with the mineralisation and the oxidation wireframes described in the previous section. Coding was undertaken on the basis that if the individual sample centroid fell within the mineralisation wireframe boundary it was coded as within the mineralisation wireframe. Each domain was assigned a unique numerical code to allow the application of hard boundary domaining where required during grade estimation. Numerical codes as described in the previous section have been assigned to the mineralisation domains.

The drillhole database coded within each mineralisation wireframe was then composited as a means of achieving a uniform sample support. It should be noted, however, that equalising sample length is not the only criteria for standardising sample support. Factors such as angle of intersection of the sampling to mineralisation, sample type and diameters, drilling conditions, recovery, sampling/sub-sampling practices and laboratory practices all affect the ‘support’ of a sample. Exploration/mining databases which contain multiple sample types and/or sources of data provide challenges in generating composite data with equalised sample support, and uniform support is frequently difficult to achieve.

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Figure 14.4 Typical Cross Section (Section NE 350 local grid) of Mineralisation Domains at M5

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Figure 14.5 Isometric SE View of Mineralisation Domains

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The lengths of the samples were statistically assessed prior to selecting an appropriate composite length for undertaking statistical analyses, variography and grade estimation. A total of 70,421 assayed intervals were available for analysis. Summary statistics of the sample length indicate that approximately 65% of the samples were collected at 1m intervals or less, 9% were collected at 2m intervals, 15% were sampled at intervals between 1m and 2m and 11% at intervals of 3m or more including 10% at 4m intervals (mainly AC sampling).

After consideration of relevant factors relating to geological setting and mining, including likely mining selectivity and bench/flitch height, a regular 3m run length (downhole) composite was selected as the most appropriate composite interval to equalise the sample support at M5. In the case of M1, a 2m downhole length was determined to be more appropriate. Compositing was broken when the routine encountered a change in flagging (wireframe boundary) and composites with residual intervals of less than 1.5m within the wireframe were excluded from the estimation.

14.2.5 Statistical Analysis Descriptive Statistics The composites flagged as described in the previous section were used for subsequent statistical, geostatistical and grade estimation investigations. Summary descriptive statistics were generated for all domains (Table 14.1). The grade distributions are typical for gold deposits of this style and show a positive skew or near lognormal behaviour (Figures 14.6 to 14.7). The coefficient of variation is moderate to high, consistent with the presence of high grade outliers that potentially require cutting (capping) for grade estimation.

Figure 14.6 Log Histograms of Uncut Gold Grade by Domain Mankarga 1 South

0.09 0.15 0.08

0.07

0.06 0.10

0.05

0.04 Frequencies Frequencies

0.03 0.05

0.02

0.01

0.00 0.00 0.01 1 100 0.1 10 1000 auppm auppm M1 South MIK Domains Combined M1 South High Grade Domains

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Table 14.1 Summary Statistics Grouped by Domain for 2m and 3m Composites of Uncut Gold Grade (g/t)

Mankarga 1 South (2m) 100 101 200 300 301 400 500 600 900 Count 338 73 24 32 50 158 236 24 50 Minimum 0.002 0.002 0.01 0.014 0.022 0.007 0.002 0.037 0.04 Maximum 19.374 12.828 5.543 24.459 171.805 31.274 24.659 31.638 21.820 Mean 1.104 1.150 1.027 1.993 12.873 1.224 0.851 4.100 2.667 Std. Dev. 1.916 2.218 1.298 4.667 29.227 3.280 2.303 7.977 4.308 Variance 3.671 4.921 1.686 21.784 854.195 10.758 5.305 63.633 18.555 CV 1.735 1.929 1.264 2.342 2.270 2.679 2.705 1.946 1.615 Mankarga 1 South High Grade (2m) 1000 1001 1002 1003 2000 4000 5000 Count 62 52 16 9 4 8 20 Minimum 0.025 0.389 0.113 2.527 10.667 1.9 1.625 Maximum 716.415 315.334 184.330 84.143 166.755 83.468 231.568 Mean 43.290 32.595 35.031 23.048 74.243 24.549 36.273 Std. Dev. 118.181 55.367 47.082 25.772 57.295 26.640 57.575 Variance 13966.754 3065.540 2216.752 664.170 3282.750 709.681 3314.829 CV 2.730 1.699 1.344 1.118 0.772 1.085 1.587 Mankarga 5 (3m) 100 102 200 300 301 400 Count 3,161 201 1,630 674 23 83 Minimum 0.008 0.011 0.006 0.006 0.301 0.04 Maximum 72.035 17.246 9.125 7.236 2.625 7.833 Mean 1.155 0.658 0.786 0.479 0.827 0.823 Std. Dev. 2.598 1.350 0.971 0.674 0.533 1.143 Variance 6.748 1.822 0.943 0.454 0.284 1.306 CV 2.250 2.052 1.235 1.406 0.644 1.389

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Figure 14.7 Log Histograms of Uncut Gold Grade by Domain Mankarga 5

0.10

0.09 0.125 0.08

0.07 0.100

0.06 0.075 0.05

Frequencies 0.04 Frequencies 0.050 0.03

0.02 0.025 0.01

0.00 0.01 0.1 1 10 100 0.01 0.1 1 10 100 auppm auppm Zone 100 Zone 102

0.10

0.09 0.100

0.08

0.07 0.075

0.06

0.05 0.050

Frequencies 0.04 Frequencies

0.03 0.025 0.02

0.01

0.000 0.01 0.1 1 10 0.01 0.1 1 10 auppm auppm Zone 200 Zone 300

0.20 0.25

0.20 0.15

0.15 0.10

Frequencies 0.10 Frequencies

0.05 0.05

0.00 0.00 0.1 1 10 0.01 0.1 1 10 auppm auppm Zone 301 Zone 400

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Cell Declustering Analysis Visual inspection of the available datasets for each of the estimation domains indicated some clustering of the data within higher grade regions of the deposit. Data clustering often occurs when drilling campaigns selectively target higher grade regions of the deposit, resulting in an artificially high mean grade in many cases. Declustering was therefore completed to remove any effects of preferential sampling of high grade areas that may have occurred. Cell declustering was not undertaken at M1S high grade domains due to the relatively low number of samples per domain.

Cell declustering was completed with weights determined as 1/n, with “n” representing the number of data in each cell. Declustered composite statistics are presented in Table 14.2. As expected, the declustered mean grades tend to be less than the composite mean grades due to the data configuration issues discussed above.

Table 14.2 Summary Statistics Grouped by Domain for 3 Composites of Uncut, Declustered Gold Grade (g/t)

High Grade Outlier Analysis Grade datasets for the various estimation domains at the Sanbrado Gold Project are characterised by moderately high CV values, indicating that high grade values may contribute significantly to the mean grades reported for the various datasets.

The effects of the highest grade composites on the mean grade and standard deviation of the gold dataset for each of the estimation domains were investigated by compiling and reviewing statistical plots (histograms and probability plots). An upper grade cut for each dataset was chosen that coincided with a pronounced inflection or increase in the variance of the data. The appropriateness of the high grade cut was assessed by viewing the composite data in 3D to determine the clustering or otherwise of the high grades in each domain. Clustering of the highest grades in one or more particular area may indicate that the grades do not require cutting and may need to be dealt with using an alternate approach.

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In the case of the high-grade domains at M1 South where extreme grade values exist and OK has been selected as the grade estimation method a range of different top cut values was considered and their effect on the composite statistics evaluated. Composites affected by top cuts were reviewed in three dimensions to validate their location and relevance relative to the entire population. The results of this analysis are summarised in Table 14.3. Ultimately, a capping value of 250g/t Au was selected for the high grade domains at M1 South.

Table 14.3 Composite Top Cut Statistics M1 South High Grade Domains

Number of % Reduction Count Min Max Mean Std. Dev. Variance CV Samples Cut Mean Grade 716.415 37.34 83.23 6,927.01 2.23 - - 250 32.22 53.25 2,835.50 1.65 4 13.71% 169 0.025 200 30.85 48.02 2,305.64 1.56 5 17.38% 150 28.48 40.31 1,625.22 1.42 10 23.73% 100 25.03 30.84 951.38 1.23 14 32.98%

A list of the determined upper cuts applied to other deposits and domains and their impact on the mean grades of the datasets is provided in Table 14.4.

Table 14.4 Summary Statistics Grouped by Domain for 2m and 3m Composites of Top cut Gold Grade (g/t)

Mankarga 1 South (2m) 100 101 200 300 301 400 500 600 900 Count 338 73 24 32 50 158 236 24 50 Minimum 0.002 0.002 0.01 0.014 0.022 0.007 0.002 0.037 0.04 Maximum 13 12.828 5.543 24.459 90 13 13 13 13 Mean 1.086 1.150 1.027 1.648 11.237 1.083 0.786 2.746 2.491 Std. Dev. 1.761 2.218 1.298 3.892 21.571 2.291 1.755 3.904 3.650 Variance 3.102 4.921 1.686 15.148 465.302 5.247 3.079 15.240 13.323 CV 1.623 1.929 1.264 2.362 1.920 2.115 2.232 1.422 1.465 Mankarga 5 (3m) 100 102 200 300 301 400 Count 3,161 201 1,630 674 23 83 Minimum 0.008 0.011 0.006 0.006 0.301 0.04 Maximum 72.035 17.246 9.125 7.236 2.625 7.833 Mean 1.155 0.658 0.786 0.479 0.827 0.823 Std. Dev. 2.598 1.350 0.971 0.674 0.533 1.143 Variance 6.748 1.822 0.943 0.454 0.284 1.306 CV 2.250 2.052 1.235 1.406 0.644 1.389

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It should be noted that while gold grades are not cut or capped for the purposes of MIK estimation, the use of cut grades is often employed for variography analysis and the change of support process. As MIK estimates are essentially a series of OK estimates applied to the binary transformation of a series of indicator cutoff grades, high grade cutting will have no effect on the resultant MIK estimate unless the high grade cut is lower than the chosen upper indicator cutoff grade. This scenario would be considered highly sub-optimal in the context of MIK estimation. A full description of the MIK estimation method with change of support is provided in Section 14.10.

Statistics by Drill Type The drillhole database for the M5 gold deposit contains samples obtained by AC, RC and DC drilling methods. In order to assess the potential for any inherent bias between the different drilling methods, descriptive statistics were generated for samples grouped by drilling type. Only samples within the mineralised envelopes at M5 were considered. Descriptive statistics by drilling type are presented in Table 14.5. Review of these statistics indicated that there was no material assay bias between the drilling methods and that the data can be combined for grade estimation.

Table 14.5 Summary Statistics Grouped by Drillhole Type

AC RC DC Count 5,213 1184 5,103 Minimum 0.00 0.01 0.00 Maximum 81.77 22.20 51.31 Mean 0.85 0.86 0.81 Std. Dev. 2.48 1.78 2.16 Variance 6.13 3.18 4.65 CV 2.91 2.08 2.66

Multiple Indicator Kriging Cutoff Grades Based on the number of samples and gold grade distributions, it was decided to estimate parts of M1 South and M5 by MIK and the remainder by OK. Indicator Kriging cutoff grades or indicator bins were selected for each domain or domain combination to be estimated by MIK. All domains were combined for M1 South, (six domains). Three separate domains were investigated for M5. Cutoff grades were based upon population distributions and metal proportions above and below the mean composite value of the proposed cutoff bins. Conditional statistics for data within each domain to be estimated by MIK are listed in Tables 14.6 and 14.7.

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Table 14.6 Indicator Class Statistics M1

M1 South

Grade Threshold Class Mean Probability Threshold (Au g/t) (Au g/t) 0.199 0.1 0.0426 0.411 0.3 0.205 0.504 0.4 0.3516 0.571 0.5 0.4393 0.616 0.6 0.5441 0.662 0.7 0.6371 0.695 0.85 0.7768 0.731 1 0.8994 0.776 1.2 1.0954 0.800 1.35 1.2963 0.834 1.55 1.431 0.870 2 1.8204 0.893 2.3 2.158 0.916 3 2.5736 0.940 4.5 3.7587 0.962 6.5 5.5966 0.980 10 8.3483 Max Max 18.0179

Table 14.7 Indicator Class Statistics M5

Zone 100 Zone 200 Zone 300

Grade Grade Grade Probability Class Mean Probability Class Mean Probability Class Mean Threshold Threshold Threshold Threshold (Au g/t) Threshold (Au g/t) Threshold (Au g/t) (Au g/t) (Au g/t) (Au g/t) 0.140 0.2 0.1166 0.196 0.2 0.1201 0.188 0.12 0.0662 0.253 0.3 0.2509 0.311 0.3 0.2453 0.374 0.22 0.1733 0.378 0.4 0.3453 0.410 0.4 0.3444 0.504 0.3 0.2573 0.473 0.5 0.4476 0.505 0.5 0.4490 0.644 0.4 0.3412 0.546 0.6 0.5488 0.573 0.6 0.5531 0.731 0.5 0.4403 0.604 0.7 0.6494 0.641 0.7 0.6489 0.810 0.65 0.5662 0.653 0.8 0.7480 0.709 0.85 0.7745 0.859 0.85 0.7486 0.703 0.95 0.8758 0.765 1 0.9252 0.915 1.1 0.9597 0.743 1.1 1.0182 0.810 1.15 1.0730 0.955 1.5 1.3002 0.781 1.25 1.1694 0.851 1.35 1.2536 0.970 2 1.6902 0.822 1.5 1.3708 0.887 1.6 1.4741 0.982 3.2 2.6081 0.858 1.8 1.6332 0.922 1.9 1.7361 0.994 4 3.6409 0.889 2.2 2.0049 0.942 2.2 2.0427 0.188 Max 5.5642 0.916 2.7 2.4230 0.960 2.6 2.3822 0.941 3.5 3.0578 0.975 3.5 3.0277 0.962 5.0 4.0779 0.990 4.5 4.0417 0.981 7.0 5.9165 0.995 7.5 5.9238 Max Max 13.6189 Max Max 8.0776

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14.2.6 Variography Introduction Variography is used to describe the spatial variability or correlation of an attribute (gold, silver etc.). The spatial variability is traditionally measured by means of a variogram, which is generated by determining the averaged squared difference of data points at a nominated distance (h), or lag (Srivastava and Isaacs, 1989). The averaged squared difference (variogram or γ(h)) for each lag distance is plotted on a bivariate plot, where the X-axis is the lag distance and the Y-axis represents the average squared differences (γ(h)) for the nominated lag distance.

Several types of variogram calculations are employed to determine the directions of the continuity of the mineralisation:

. Traditional variograms are calculated from the raw assay values. . Log-transformed variography involves a logarithmic transformation of the assay data. . Gaussian variograms are based on the results after declustering and transformation to a normal distribution. . Pairwise-relative variograms attempt to ‘normalise’ the variogram by dividing the variogram value for each pair by its squared mean value. . Correlograms are ‘standardised’ by the variance calculated from the sample values that contribute to each lag. . Fan variography involves the graphical representation of spatial trends by calculating a range of variograms in a selected plane and contouring the variogram values. The result is a contour map of the grade continuity within the domain. The variography was calculated and modelled in the geostatistical software, Isatis. The rotations are tabulated as dip and dip direction of major, semi-major and minor axes of continuity. Modelled variograms were generally shown to have good structure and were used throughout the MIK estimation and for the change of support process.

Sanbrado Gold Project Variography Grade and indicator variography was generated to enable grade estimation by MIK and change of support analysis to be completed. In addition, Gaussian variograms were examined as part of the change of support process. Indicator thresholds for domains estimated by MIK had variograms modelled, generally with every second variogram modelled. Variograms not modelled have had their parameters interpolated based on the bounding modelled variograms. Interpreted anisotropy directions corresponded well with the modelled geology and overall geometry of the interpreted domains.

Mankarga 1 South Variography modelled for M1 South displayed moderate anisotropy between the major and semi-major axes. Gaussian variography was back transformed into normal space and modelled. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a relative nugget effect.

The interpreted major direction of continuity dips at 80º towards 040º. The semi-major and minor directions are orthogonal to this direction. The modelled Gaussian back transformed variogram plot is provided in Figure 14.8.

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Figure 14.8 Mankarga 1 South Variogram

5 N40 D-

N310 3

2 Variogram : au cut : Variogram N40 D80

0 0 50 100 150

Distance (m)

Using the raw cut data, a grade variogram was calculated for the high grade domains modelled at M1 South. The variogram as modelled is presented in Figure 14.9 and Table 14.12.

Figure 14.9 Mankarga 1 South High Grade Variogram

5000

4000 N40 D-

3000

N40 D80

2000 Variogram : au cut : au Variogram N130 1000

0 0 25 50 75 100 125

Distance (m)

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Table 14.8 presents the fitted grade and indicator variogram models for M1 South.

Mankarga 5 Zone 100 Variography modelled for Zone 100 showed good structure and displayed moderate anisotropy between the major and semi-major axes. Gaussian variography was back transformed into normal space and modelled. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a relative nugget effect (calculated by dividing the nugget variance by the sill variance) of 42%. The short-range structure (which was modelled with ranges of 22m, 13m and 4m for the major, semi-major and minor axis respectively) accounts for 65% of the non-nugget variance. The overall ranges fitted were 132m, 99m and 21m for the major, semi-major, and minor axis respectively.

The interpreted major direction of continuity lies in a plane that strikes at 045º with a sub- vertical dip and displays a steep plunge of 60º towards 225º. The semi-major and minor directions are orthogonal to this direction. The modelled Gaussian back transformed variogram plot is provided in Figure 14.10.

Figure 14.10 Zone 100 – Back-transformed Gaussian Variogram

N315 D 5

N222 D 3

Variogram : auppm : Variogram 2 N234 D

0 0 50 100 150 200

Distance (m)

Modelled indicator variograms displayed a range of relative nugget values from 22% to 47%. Table 14.9 presents the fitted grade and indicator variogram models for Zone 100.

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Table 14.8 M1 South Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 2.35 310 80 90 2.00 39 26 4 1.31 114 79 9 Indicator Variography 0.10(1) 0.0412 310 80 90 0.0725 30 30 5 0.0511 80 70 9 0.30(1) 0.0647 310 80 90 0.1031 30 30 5 0.0719 75 70 9 0.40 0.0790 310 80 90 0.0950 30 20 4 0.0760 150 70 9 0.50(2) 0.0806 310 80 90 0.0921 30 17.5 4 0.0740 145 70 9 0.60 0.0810 310 80 90 0.0880 30 15 4 0.0710 140 70 9 0.70(3) 0.0785 310 80 90 0.0834 27.5 15 4 0.0659 130 65 8.5 0.85 0.0770 310 80 90 0.0800 25 15 4 0.0620 120 60 8 1.00(4) 0.0745 310 80 90 0.0742 25 15 4 0.0558 120 60 8 1.20 0.0690 310 80 90 0.0660 25 15 4 0.0480 120 60 8 1.35(5) 0.0657 310 80 90 0.0603 25 15 4 0.0434 120 60 8 1.55 0.0590 310 80 90 0.0520 25 15 4 0.0370 120 60 8 2.00(6) 0.0521 310 80 90 0.0439 25 15 4 0.0318 120 60 8 2.30 0.0470 310 80 90 0.0380 25 15 4 0.0280 120 60 8 3.00(7) 0.0419 310 80 90 0.0328 25 15 4 0.0244 120 60 8 4.50 0.0290 310 80 90 0.0220 25 15 4 0.0165 120 60 8 6.50(8) 0.0203 310 80 90 0.0140 20 10 3 0.0108 100 50 7 10.0(8) 0.0109 310 80 90 0.0069 15 8 2 0.0053 80 40 5 Note: 1) Assumed model based on 0.40 Au g/t variogram model 2) Assumed model based on 0.40 Au g/t and 0.60 Au g/t variogram models 3) Assumed model based on 0.60 Au g/t and 0.85 Au g/t variogram models 4) Assumed model based on 0.85 Au g/t and 1.20 Au g/t variogram model 5) Assumed model based on 1.20 Au g/t and 1.55 Au g/t variogram model 6) Assumed model based on 1.55 Au g/t and 2.30 Au g/t variogram models 7) Assumed model based on 2.30 Au g/t and 4.50 Au g/t variogram models 8) Assumed model based on 4.50 Au g/t variogram model

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Table 14.9 M5 Zone 100 Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 0.417 45 80 0 0.383 22 13 4 0.199 132 99 21 Indicator Variography 0.20(1) 0.200 45 80 0 0.395 40 25 9 0.41 180 130 35 0.30(1) 0.210 45 80 0 0.390 38 22 8 0.40 175 125 32 0.40 0.235 45 80 0 0.393 35 20 7 0.372 170 120 30 0.50(2) 0.249 45 80 0 0.382 32.5 17.5 6.5 0.369 165 115 27.5 0.60 0.262 45 80 0 0.371 30 15 6 0.367 160 110 25 0.70(3) 0.281 45 80 0 0.359 30 15 5 0.360 145 100 20 0.80 0.300 45 80 0 0.348 30 15 4 0.352 130 90 15 0.95(4) 0.325 45 80 0 0.336 30 15 4 0.339 125 87.5 15 1.10 0.351 45 80 0 0.325 30 15 4 0.325 120 85 15 1.25(5) 0.387 45 80 0 0.307 30 15 3.5 0.305 100 75 13.5 1.50 0.424 45 80 0 0.290 30 15 3 0.286 80 65 12 1.80(6) 0.433 45 80 0 0.282 27.5 13.5 3 0.285 70 57.5 9.5 2.20 0.442 45 80 0 0.274 25 12 3 0.284 60 50 7 2.70(7) 0.467 45 80 0 0.264 22.5 12 3 0.269 50 40 6.5 3.50 0.492 45 80 0 0.254 20 12 3 0.254 40 30 6 5.00(8) 0.520 45 80 0 0.240 16 8 3 0.240 35 25 5 7.00(8) 0.540 45 80 0 0.230 14 7 2 0.230 30 20 4 Note: 1) Assumed model based on 0.40 Au g/t variogram model 2) Assumed model based on 0.40 Au g/t and 0.60 Au g/t variogram models 3) Assumed model based on 0.60 Au g/t and 0.80 Au g/t variogram models| 4) Assumed model based on 0.80 Au g/t and 1.15 Au g/t variogram model 5) Assumed model based on 1.15 Au g/t and 1.60 Au g/t variogram model 6) Assumed model based on 1.60 Au g/t and 2.20 Au g/t variogram models 7) Assumed model based on 2.20 Au g/t and 3.20 Au g/t variogram models 8) Assumed model based on 3.20 Au g/t variogram model

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Zone 200 Variography modelled for Zone 200 showed good structure and displayed moderate to strong anisotropy between the major and semi-major axes. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a moderate relative nugget effect of 34%. The short-range structure (which was modelled with ranges of 37m, 27m and 11m for the major, semi-major and minor axis respectively) accounts for 60% of the non-nugget variance. The overall ranges fitted to the Zone 200 variogram were 148m, 111m and 25m for the major, semi-major, and minor axis respectively.

The interpreted major direction of continuity lies in a plane that strikes at 45º with a sub- vertical dip of 80º to the west and displays a moderate plunge of 80º towards 225º. The semi-major and minor directions are orthogonal to this direction. Table 14.10 presents the fitted grade variogram and indicator variogram models for Zone 200. The Gaussian back transformed variogram plot is provided in Figure 14.11.

Figure 14.11 Zone 200 – Back-transformed Gaussian Variogram

0.9

0.8

0.7 N270 D

0.6

0.5 N223 D 0.4

0.3 Variogram : Variogram : auppm N315 D 0.2

0.1

0.0 0 50 100 150 200

Distance (m)

Zone 300 Variography showed good structure and displayed moderate to strong anisotropy between the major and semi-major axes. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a moderate relative nugget effect of 40%. The short-range structure (which was modelled with ranges of 34m, 27m and 8m for the major, semi-major and minor axis respectively) accounts for 57% of the non-nugget variance. The overall ranges fitted to the Zone 300 variogram were 143m, 94m and 19m for the major, semi-major, and minor axis respectively.

The interpreted major direction of continuity lies in a plane that strikes at 225º with a sub- vertical dip of 80º to the west and a sub-vertical plunge of 60º towards 225º. The semi- major and minor directions are orthogonal to this direction. Table 14.11 presents the fitted grade variogram and indicator variogram models for Zone 300. The Gaussian back transformed variogram plot is provided in Figure 14.12.

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Table 14.10 M5 Zone 200 Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 0.341 220 80 50 0.402 37 27 11 0.256 148 111 25 Indicator Variography 0.20(1) 0.0283 220 80 50 0.0614 34 15 10 0.0677 180 130 35 0.30(1) 0.0407 220 80 50 0.0814 32 15 10 0.0921 180 120 32 0.40 0.0500 220 80 50 0.0870 30 15 7 0.0970 180 120 25 0.50(2) 0.0621 220 80 50 0.0890 27.5 15 6.5 0.0989 160 110 22.5 0.60 0.0700 220 80 50 0.0840 25 15 6 0.0930 140 100 20 0.70(3) 0.0737 220 80 50 0.0775 23.5 15 5.5 0.0789 120 87.5 17.5 0.85 0.0750 220 80 50 0.0700 22 15 5 0.0650 100 75 15 1.00(4) 0.0663 220 80 50 0.0595 21 13.5 5.5 0.0539 100 67.5 15 1.15 0.0590 220 80 50 0.0510 20 12 6 0.0450 100 60 15 1.35(5) 0.0498 220 80 50 0.0401 20 11 5 0.0370 95 60 15 1.60 0.0400 220 80 50 0.0300 20 10 4 0.0290 90 60 15 1.90(6) 0.0314 220 80 50 0.0200 20 10 4 0.0204 82.5 55 13.5 2.20 0.0240 220 80 50 0.0130 20 10 4 0.0140 75 50 12 2.60(7) 0.0187 220 80 50 0.0099 17.5 10 4 0.0104 75 50 12 3.50 0.0095 220 80 50 0.0049 15 10 4 0.0050 75 50 12 4.50(8) 0.0049 220 80 50 0.0023 10 8 4 0.0024 35 20 12 7.50(8) 0.0026 220 80 50 0.0011 8 6 3 0.0012 30 18 10 Note: 1) Assumed model based on 0.40 Au g/t variogram model 2) Assumed model based on 0.40 Au g/t and 0.60 Au g/t variogram models 3) Assumed model based on 0.60 Au g/t and 0.80 Au g/t variogram models| 4) Assumed model based on 0.80 Au g/t and 1.15 Au g/t variogram model 5) Assumed model based on 1.15 Au g/t and 1.60 Au g/t variogram model 6) Assumed model based on 1.60 Au g/t and 2.20 Au g/t variogram models 7) Assumed model based on 2.20 Au g/t and 3.20 Au g/t variogram models 8) Assumed model based on 3.50 Au g/t variogram model

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Table 14.11 M5 Zone 300 Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 0.15 225 80 60 0.13 34 27 8 0.098 143 94 19 Indicator Variography 0.12(1) 0.0341 225 80 60 0.0559 30 20 6 0.0463 150 90 30 0.22(1) 0.0615 225 80 60 0.0911 30 17 6 0.0752 140 85 28 0.30 0.0700 225 80 60 0.0970 30 15 6 0.0820 130 80 25 0.40(2) 0.0675 225 80 60 0.0903 30 15 6 0.0773 125 70 25 0.50 0.0580 225 80 60 0.0750 30 15 6 0.0650 120 60 25 0.65(3) 0.0440 225 80 60 0.0555 27.5 15 5 0.0479 110 55 22.5 0.80 0.0350 225 80 60 0.0430 25 15 4 0.0370 100 50 20 1.10(4) 0.0245 225 80 60 0.0274 25 15 4 0.0243 95 50 20 1.50 0.0120 225 80 60 0.0122 25 15 4 0.0112 90 50 20 2.00(5) 0.0089 225 80 60 0.0082 25 15 4 0.0077 85 45 15 3.20. 0.0060 225 80 60 0.0050 25 15 4 0.0049 80 40 10 5.00(8) 0.0014 225 80 60 0.0010 20 12 3 0.0010 70 37 8 Note: 1) Assumed model based on 0.30 Au g/t variogram model 2) Assumed model based on 0.30 Au g/t and 0.50 Au g/t variogram models 3) Assumed model based on 0.50 Au g/t and 0.80 Au g/t variogram models 4) Assumed model based on 0.80 Au g/t and 1.50 Au g/t variogram model 5) Assumed model based on 1.50 Au g/t and 3.20 Au g/t variogram models 6) Assumed model based on 3.20 Au g/t variogram model

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Figure 14.12 Zone 300 – Back-transformed Gaussian Variogram

0.4

N315 D 0.3

N270 D 0.2 Variogram : auppm 0.1 N223 D

0.0 0 50 100 150 200

Distance (m)

Variography for OK Domains Variography for Domains 101 and 301 (domains estimated by OK) was assumed and was based on variography calculated for Domains 100 and 300 (the minor domains are characterised by low numbers of samples resulting in low confidence variograms). Variography for Domain 400 was calculated and modelled (Figure 14.13). Variogram models as applied to OK estimates are tabulated in Table 14.13.

Figure 14.13 Zone 400 – Back-transformed Gaussian Variogram

0.3 N135

0.2 D-90

Variogram : auppm 0.1 N45

0.0 0 25 50 75 100 125

Distance (m)

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Table 14.12 M1 Variogram Models for OK Estimates

Table 14.13 M3 Variogram Models for OK Estimates

Structure 1 Structure 2 Rotation Grade Variable or Nugget Isatis Geological Plane Convention Relative Range (m) Relative Range (m) Indicator Threshold (C0) Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor 102, Gold (Au g/t) 1.3 35 80 0 1.32 25 15 5 0.55 160 90 25 301, Gold (Au g/t) 0.0138 45 75 0 0.057 60 30 10 0.05 150 80 25 400, Gold (Au g/t) 0.133 30 80 0 0.092 45 30 4 0.111 110 50 15

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14.2.7 Block Modelling 3-D block models were created for the four separate deposit areas in the Universal Transverse Mercator (UTM) grid projection, Zone 30 North, using the World Geodetic System 1984 datum (WGS84). The parent block size was selected on the basis of the average drill spacing. Sub-blocking was applied to an appropriate dimension to ensure adequate volume representation. The models covered all the interpreted mineralisation zones and included suitable additional waste material to allow later pit optimisation studies. Block coding was completed on the basis of the block centroid, wherein a centroid falling within any wireframe was coded with the wireframe solid attribute. Rotations were applied where appropriate to ensure suitable coverage.

The main block model parameters are summarised in Table 14.14. Variables were coded into the block model to enable MIK and OK estimation and subsequent change of support and grade tonnage reporting. A visual review of the wireframe solids and the block model indicated correct flagging of the block model. Additionally, a check was made of coded volume versus wireframe volume which confirmed the above.

Table 14.14 Block Model Parameters

M1 South Northing (Y) Easting (X) RL (Z) Min. Coordinates 1,336,650m 741,600m -320m Offset 650m 450m 650m Block size (m) 25 10 10 Sub Block size (m) 2.5 2.5 2.5 Rotation (° around axis) 0° 0° -45° M5 Northing (Y) Easting (X) RL (Z) Min. Coordinates 1,335,621.257m 741,559.695m -300m Offset 3,175m 900m 630m Block size (m) 25 20 10 Sub Block size (m) 12.5 5 5 Rotation (° around axis) 0° 0° 30°

14.2.8 Dry Bulk Density Data An extensive dry bulk density database was supplied containing a total of 13,679 dry bulk density determinations widely dispersed throughout the different mineralisation domains and covering the entire project. The database was subdivided on the basis of oxidation, resulting in 411,379 determinations of fresh rock, 746 of transition material and 1,553 of oxide material. Measured dry bulk density readings were subdivided by oxidation state and mineralised domain. No bias was observed based on readings within mineralised zones or outside. Measured average dry bulk density values subdivided by oxidation state are presented in Table 14.15 below. Dry bulk densities were applied to the block models as average dry bulk densities per oxidation state as per Table 14.15.

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Table 14.15 Measured Average Dry Bulk Density Value

Measured Average Bulk Deposit Domain Lithology Density (t/m³) Strongly oxidised 2.00 Moderately oxidised 2.52 M1 South all Transition 2.77 Fresh 2.80 Strongly oxidised 2.05 Moderately oxidised 2.51 100 and 300 Transition 2.71 Fresh 2.75 Strongly oxidised 2.10 Moderately oxidised 2.63 200 and 400 Transition 2.76 Fresh 2.76 M5 Strongly oxidised 2.1 Moderately oxidised 2.57 Other Mineralised Domains Transition 2.72 Fresh 2.75 Strongly oxidised 2.04 Moderately oxidised 2.52 All Other Material Transition 2.66 Fresh 2.76

14.2.9 Grade Estimation Introduction MIK was used as the grade estimation method within the defined mineralisation wireframes at M1 and M5. OK was used as the grade estimation method for the minor domains at M5 and the high grade domains at M1. Estimation was completed in the mining package Vulcan using the GSLib geostatistical software. Change of support was applied to the MIK grade estimates to produce ‘recoverable’ gold estimates targeting a selective mining unit (SMU) of 5mE by 12.5mN by 5mRL at M5 and 5mE by 5mN by 5mRL at M1S.

The Multiple Indicator Kriging Method The MIK technique is implemented by completing a series of OK estimates of binary transformed data. Composite assay results equal to or above a nominated cutoff or threshold grade are assigned a value of 1. Composite assay results below the nominated indicator threshold grade are assigned a value of 0. The indicator estimates, with a range between 0 and 1, represent the probability that the point will exceed the indicator cutoff grade. The probability of the points exceeding a cutoff can be considered broadly equivalent to the proportion of a nominated block that will exceed the nominated cutoff grade.

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The estimation of a complete series of indicator cutoffs allows the reconstitution of the local histogram or conditional cumulative distribution function (ccdf) for the estimated point. Based on the ccdf, local or block properties, such as the block mean and proportion (tonnes) above or below a nominated cutoff grade can be investigated.

Post MIK Processing - E-Type Estimates The E-type estimate provides an estimate for the grade of the total block, which represents a realistic bulk-mining scenario. This is achieved by discretising the calculated ccdf for each block into a nominated number of intervals and interpolating between the given points with a selected function (e.g. the linear, power or hyperbolic model) or by applying intra- class mean grades. The sum of all these weighted interpolated points or mean grades enables an average whole block grade to be estimated.

An example of the determination of an E-type estimate for a block containing three indicator cutoffs is outlined below. The indicator cutoffs and calculated probabilities are shown in Table 14.16.

Table 14.16 Indicator Cutoff and Probability

Cutoff Grades Indicator Probability Indicator (Au g/t) (cumulative) minimum grade * 0 0.00 ** indicator 1 1 0.40 indicator 2 2 0.65 indicator 3 3 0.85 maximum grade * 4 1.00 **

Note: * Cutoff grades determined by the user. ** Indicator probability is assumed at the minimum and maximum cutoff.

The whole block grade can now be estimated using the following steps:

. Interval 1 (0-1g/t Au) median grade x probability/proportion attributed to the interval (0.5g/t Au x 0.40 = 0.200). . Interval 2 (1 - 2g/t Au) median grade x proportion (1.5g/t Au x 0.25 = 0.375). . Interval 3 (2 - 3g/t Au) median grade x proportion (2.5g/t Au x 0.20 = 0.500). . Interval 4 (3 - 4g/t Au) median grade x proportion (3.5g/t Au x 0.15 = 0.525). . Calculate total grade sum of calculated intervals ((0.2+0.375+0.500+0.525)/1) = 1.60g/t Au. It is also possible to calculate the proportion and grade above a nominated cutoff (e.g. 2g/t - at sample support or complete selectivity). The following steps would be undertaken to calculate the tonnes and grade at sample selectivity using a 2g/t cutoff:

. Interval 3 (2 - 3g/t Au) median grade x proportion (2.5g/t Au x 0.20 = 0.500). . Interval 4 (3 - 4g/t Au) median grade x proportion (3.5g/t Au x 0.15 = 0.525). . Calculate total grade sum of calculated intervals ((0.500+0.525)/0.35) = 2.93g/t Au with 0.35% of the block above the cutoff.

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The effect of using a non-linear model to interpolate between cutoffs is to shift the grade weighting associated with that cutoff away from the median. For the Sanbrado Gold Project, the intra-class means based on the cut composite data were used to reconstitute the ccdf and to produce block statistics.

It should be noted that the calculation of the E-type estimate and complete selectivity does not allow mine planning to the level of selectivity proposed for production. To achieve an estimate which reflects the levels of mining selectivity envisaged, a selective mining unit (SMU) correction is often applied to the calculated ccdf.

Support Correction (Selective Mining Unit Estimation) A range of techniques can be used to produce a support correction to allow for SMU emulation. The common features of the support correction are:

. Maintenance of the mean grade of the histogram (E-type mean). . Adjustment of the histogram variance by a variance adjustment factor (the ‘f’ factor). The variance adjustment factor, used to reduce the histogram or ccdf variance, can be calculated using the variogram model. The variance adjustment factor is often modified to account for the likely grade control approach or ‘information effect’.

In simplest terms, the variance adjustment factor takes into account the known relationship derived from the dispersion variance.

Total variance = variance of samples within blocks + variance between blocks.

The variance adjustment factor is calculated as the ratio of the variance between the blocks and the variance of the samples within the blocks, with a small ratio (e.g. 0.10) indicating a large adjustment of the ccdf variance and large ratio (e.g. 0.80) representing a small shift in the ccdf.

Simple support corrections that are available include the Affine and Indirect Lognormal corrections, which are both based on the permanence of distribution. The Indirect Lognormal correction was applied to the M5 MIK grade estimates. The discrete Gaussian model is often applied to global change of support studies and as such was generated on the composite dataset as a comparison.

Indirect Lognormal Correction The Indirect Lognormal correction is implemented by adjusting the quantiles (indicator cutoffs) of the ccdf with the variance adjustment factor so that the adjusted ccdf represents the statistical characteristics of the block volume of interest.

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This is implemented with the following formula:

At the completion of the quantile adjustments, grade and tonnes (which are considered a pseudo tonnage proportion of the blocks) can be estimated at a nominated cutoff grade using the methods described above for estimation of the E-type means. The indirect lognormal correction, as applied to the Sanbrado gold deposit, is the best suited of the common adjustments applied to MIK to produce selective mining estimates for positively skewed distributions.

Multiple Indicator Kriging Parameters MIK estimates were completed using the indicator variogram models (Section 14.7) and a set of ancillary parameters controlling the source and selection of composite data. The sample search parameters were defined based on the variography and the data spacing, and a series of sample search tests were performed using Isatis geostatistical software.

Mankarga 1 South At M1 south, a total of 17 indicator thresholds were estimated for the applicable domains. The sample search parameters are provided in Table 14.17. A hard domain boundary approach was used for the estimation whereby the MIK estimation was not allowed to select any composites from the adjacent high grade domains. Where required, a two-pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations (10mE by 25mN by 10mRL) were applied throughout and discretisation was applied on the basis of 3mE by 3mN by 2mRL, resulting in 18 discretisation points per block.

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Table 14.17 M1 South MIK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH Pass 1 -60/308 28/285 10/021 70 60 20 18 24 4 100, 101 Pass 2 -60/308 28/285 10/021 1000 1000 100 18 24 - Pass 1 -60/308 28/285 10/021 100 800 30 18 24 4 200 ------Pass 1 -60/014 15/312 25/050 70 60 20 18 24 4 400 Pass 2 -60/014 15/312 25/050 1000 1000 100 18 24 - Pass 1 -60/014 15/312 25/050 70 60 20 18 24 4 500 Pass 2 -60/014 15/312 25/050 1000 1000 100 18 24 - Pass 1 -60/014 15/312 25/050 1000 1000 100 18 24 4 600 ------

Mankarga 5 In the case of M5, a total of 12 indicator thresholds were estimated for Zone 300 and 17 indicator thresholds were estimated for Zones 100 and 200.

The sample search parameters are provided in Table 14.18. Domain 100 was subdivided into north and south portions to allow for a subtle change in orientation that is evident around 1,336,400mN. Hard domain boundaries were used for the estimation throughout, with the exception of Domain 100, where samples from the north portion were allowed to be selected during estimation of the south portion and vice versa. A two-pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations (20mE by 25mN by 10mRL) were applied throughout and discretisation was applied on the basis of 3mE by 3mN by 2mRL, resulting in 18 discretisation points per block.

Table 14.18 M1 South MIK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH 100 Pass 1 -40/32 50/32 0/122 150 150 30 24 32 5 north Pass 2 -40/32 50/32 0/122 1000 1000 100 12 32 - 100 Pass 1 -40/45 50/45 0/135 150 150 30 24 32 5 south Pass 2 -40/45 50/45 0/135 1000 1000 100 12 32 - Pass 1 -49/232 39/213 10/311 150 150 30 24 32 5 200 Pass 2 -49/232 39/213 10/311 1000 1000 100 12 32 - Pass 1 -59/242 29/220 10/316 150 150 30 24 32 5 300 Pass 2 -59/242 29/220 10/316 1000 1000 100 12 32 -

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Ordinary Kriging Parameters OK estimates were completed using the variogram models and a set of ancillary parameters controlling the source and selection of composite data. The sample search parameters were defined based on the variography and the data spacing, and a series of sample search tests were performed using Isatis geostatistical software.

The sample search parameters are provided in Table 14.19 to Table 14.20. Hard domain boundaries were used for the OK estimation throughout. Where necessary, a two-pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations were applied throughout and discretisation was applied on the basis of 2mE by 3mN by 2mRL, resulting in 12 discretisation points per block.

Table 14.19 Mankarga 1 South OK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH 1000 Pass 1 -80/040 0/310 10/040 200 200 30 4 8 2 1001 Pass 1 -80/040 0/310 10/040 200 200 30 4 8 2 1002 Pass 1 -80/050 0/320 10/050 150 150 30 4 8 2 1003 Pass 1 -80/050 0/320 10/050 150 100 30 4 8 2 2000 Pass 1 -80/040 0/310 10/040 100 100 30 4 8 2 3000 Pass 1 -80/050 0/320 10/050 100 100 30 4 8 2 3001 Pass 1 -80/050 0/320 10/050 100 100 30 4 8 3 4000 Pass 1 -75/050 0/320 15/050 100 100 30 4 8 2 5000 Pass 1 -70/050 0/320 20/050 100 100 50 4 8 2

Table 14.20 Mankarga 5 OK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH Pass 1 -80/040 0/310 10/040 150 50 30 8 8 2 102 Pass 2 -80/040 0/310 10/040 300 150 50 6 8 - Pass 1 -75/035 0/305 15/035 150 150 30 8 8 2 301 Pass 2 -75/035 0/305 15/035 300 300 30 6 8 - Pass 1 -90/25 0/25 0/110 150 150 30 8 8 2 400 Pass 2 -90/25 0/25 0/110 300 300 50 6 8 -

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Estimate Validation All relevant statistical information was recorded to enable validation and review of the MIK and OK estimates. The recorded information included:

. Number of samples used per block estimate. . Average distance to samples per block estimate. . Estimation flag to determine in which estimation pass a block was estimated. . Number of drillholes from which composite data were used to complete the block estimate. The estimates were reviewed visually and statistically prior to being accepted. The review included the following activities:

. Comparison of the E-type or OK estimate versus the mean of the composite dataset, including weighting where appropriate to account for data clustering. . Comparison of the reconstituted cumulative conditional distribution functions of the MIK estimated blocks versus the input composite data. . Visual checks of cross sections, long sections, and plans. Comparative estimates using different grade estimation methods were also completed to test the sensitivity of the reported model to the selected MIK and OK interpolation parameters. An insignificant amount of variation in overall grade was noted in the comparative estimations.

MIK Estimate - Change of Support Applying the modelled variography, variance adjustment factors were calculated to emulate a 5mE by 12.5mN by 5mRL SMU via the Indirect Lognormal change of support for the domains estimated via MIK. The intra-class composite mean grades (Section 14.6.5) were used to calculate whole block and SMU grades. The change of support study also included the calculation of the theoretical global change of support via the discrete Gaussian change of support model.

An ‘information effect’ factor is commonly applied to the originally derived panel-to-block variance ratios to determine the final variance adjustment ratio. The goal of incorporating information effect is to calculate results taking into account that mining takes place based on grade control information. There will still be a quantifiable error associated with this data and it is this error we want to incorporate. This is achieved in practice by running a test kriging estimation of an SMU using grade control data (the results required to incorporate this option in the change of support do not depend on the assay data so the grade control data can be hypothetical). The incorporation of the information effect is commonly found to be negligible, however can have a significant effect in some cases. In this case, the information effect factor was found to have a minor effect and has been incorporated in the calculation.

The variance adjustment ratios are provided in Table 14.21.

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Table 14.21 Variance Adjustment Ratios

Mankarga 1 South Zone All MIK Domains Variance adjustment factor (f) 0.25 Mankarga 5 Zone 100 200 300 Variance adjustment factor (f) 0.25 0.25 0.27

Resource Classification The resource categorisation was based on the robustness of the various data sources available, including:

. Geological knowledge and interpretation. . Variogram models and the ranges of the first structure in multi-structure models. . Drilling density and orientation. . Estimation quality statistics. The resource estimates for the Sanbrado Gold Project have been classified as Indicated and Inferred Mineral Resources based on the confidence levels of the key criteria as presented in Table 14.22.

Applying these confidence levels, resource classification codes were assigned to the block model using the following criteria:

. Indicated Mineral Resources

 Blocks are predominately estimation pass 1.

 Distance to nearest data of 35m or less.

 Drillhole section spacing predominantly 50m.

 Appropriate quality of estimate statistics. . Inferred Mineral Resources

 Blocks are predominantly estimation pass 2.

 Blocks not assigned as Indicated.

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Table 14.22 Confidence Levels by Key Criteria

Items Discussion Confidence Drilling Techniques AC/DC/RC - Industry Standard approach. Moderate/High Logging Standard nomenclature has been adopted. Moderate Recoveries are not recorded in entire database. Review of Drill Sample Recovery current drilling suggests RC recoveries are of acceptable Moderate/High standard. Sub-sampling Techniques and Sampling conducted by industry standard techniques. Moderate/High Sample Preparation Appropriate quality control procedures are available. They were reviewed on site and considered to be of industry Quality of Assay Data Moderate/High standard. All analyses showed acceptable assay precision and accuracy. Verification of Sampling and Sampling and assaying procedures have been assessed and Moderate/high Assaying are considered of appropriate industry standards. Survey of all collars conducted with accurate survey Location of Sampling Points equipment. Investigation of downhole survey indicates Moderate/High appropriate behaviours. Majority of regions defined on a notional 25mE x 50mN drill Data Density and Distribution Moderate spacing. Audits or Reviews Data collection assessed during site review. High Assay certificates have been verified and no issues were Database Integrity Moderate identified. Mineralisation controls are generally well understood. The mineralisation constraints overall are robust but can be Geological Interpretation relatively broad and therefore of moderate confidence. High Moderate grade mineralisation continuity at M1 South is less certain, however continues to improve with additional drilling. A combination of Ordinary Kriging and Multiple Indicator Estimation and Modelling Kriging are considered to be appropriate given the geological High Techniques setting and grade distribution. MIK is independent of cutoff grade although the mineralisation constraints were based on a notional Cutoff Grades Moderate/High 0.3g/t Au lower cutoff grade. A 0.5g/t lower cutoff grade is considered appropriate for reporting. A 5mE x 12.5mN x 5mRL SMU emulated for gold assuming a Mining Factors or bulk open pit mining scenario. Change of support for Moderate Assumptions Inferred component has higher degree of uncertainty. Testwork has shown LOM recovery greater than 90% and low Metallurgical Factors or reagent consumption. All ore types are amenable to Moderate/high Assumptions conventional CIL processing. Sufficient bulk density data exists to enable estimation of Tonnage Factors bulk density via ordinary kriging. Average densities as High (In-situ Bulk Densities) reported from the model are as per Section 10.3.

14.2.10 Resource Reporting The summary total resource for the Sanbrado Gold Project is provided in Table 14.13. The Mineral Resource estimates are predicated on an open pit mining and conventional CIL processing scenario and have been reported within optimised pit shells using a gold price of US$1650/oz and a cutoff grade of 0.3g/t Au for oxide mineralisation and 0.4g/t Au for fresh rock. Key pit optimisation input parameters are as follows:

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. Metallurgical recovery for CIL processing of 95% for oxide and transitional material, and 90% for fresh rock. . Mining costs: $1.50/t oxide; $1.90/t transitional; $2.50/t fresh. . Process costs: $9.00/t for oxide; $12.00/t for transitional and fresh. . Pit slope angles of 45° for oxide and 50° for transitional and fresh . Conceptual annual production rate of 2.0 million tonnes per annum (Mtpa).

Table 14.23 Mineral Resource Statement, Sanbrado Gold Project, 13th December 2017 Summary Grade Tonnage Report, MIK and OK estimates

Indicated Resource Inferred Resource Cutoff Grade Grade (Au g/t) Tonnes Au Oz Tonnes Au Oz (Au g/t) (Au g/t) O/P <120m 0.5 730,000 6.8 161,000 70,000 5.1 11,000 M1 South U/G >120m 3 470,000 26.4 395,000 350,000 16.1 180,000 Total Combined 1,200,000 14.4 556,000 410,000 14.4 191,000 M5 O/P 0.5 35,890,000 1.3 1,461,000 11,950,000 1.1 412,000 M1 North* O/P 0.5 780,000 1.9 49,000 660,000 1.9 41,000 M3* 0.5 170,000 2.0 11,000 260,000 1.4 12,000 Note: Mineral resources are not mineral reserves and have not demonstrated economic viability. All figures have been rounded to reflect the relative accuracy of the estimates.

Additionally, underground mining is assumed for the high grade portions of M1 South. In this case the resource has been reported at a 3g/t Au lower cut-off grade. This assumption is based on review of other similar operations both in the region and internationally.

The mineral resources for both M3 and M1 North have not been updated since the 20 February 2017 resource estimate. They are included in the tabulation below for completeness.

The effective date of this Mineral Resource is 13th December 2017. It is not anticipated that this Mineral Resource estimate will be materially affected, to any extent, by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors.

The preferred reporting cutoff grade is 0.5g/t Au, which was selected based on previous cutoff grade calculations and reporting cutoffs applied to similar gold deposits in the region. The Mineral Resource estimate has also been reported at a 1g/t Au cutoff grade to demonstrate grade tonnage relationships at higher cutoff grades.

Cautionary Note About Mineral Resources: Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. Mineral Resource estimates do not account for mineability, selectivity, mining loss and dilution. The Sanbrado mineral resource estimates include Inferred Mineral Resources that are normally considered too low confidence to have economic considerations applied to them that would enable them to be categorised as Mineral Reserves. There is no certainty that these Inferred Mineral Resources will be converted to the Measured and Indicated categories through further drilling.

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Mineral Resource Estimates 20th February 2017 This Mineral Resource for the Sanbrado Gold Project has been previously estimated as at the effective date of the 20th February 2017. Grade estimates were completed for M1 North, M1 South, M3 and M5 (Figure 14.14). Gold grade estimation was completed using a combination of Multiple Indicator Kriging (MIK) and Ordinary Kriging (OK). This estimation approach was considered appropriate based on review of a number of factors, including the quantity and spacing of available data, the interpreted controls on mineralisation, and the style, geometry and tenor of mineralisation. The estimation was constrained with geological and mineralisation interpretations.

Figure 14.14 Plan View of Drilling

14.3.1 Database Validation See Section 14.2.1 for a description of database validation procedures.

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14.3.2 Summary of Data Used in Estimate The area of the M1 Mineral Resource was drilled using AC, RC and DC drillholes and RC drillholes with a diamond tail (DT) on a nominal 25m by 25m grid spacing. A total of 360 AC holes (6,950m), 53 DC holes (11,441m) and 148 RC holes (15,320m) were drilled by WAF in 2015 and 2016. A total of 27 RC holes (3,378m) and seven DC holes (1,199m) were drilled by Channel in 2003 and between 2010 and 2012. Hole azimuths were either 225° or 180° magnetic at declinations of between -50° and -60°, to optimally intersect the mineralised zones. A small number of drillholes had azimuths of approximately 020° or 045° due to restricted rig access.

The area of the M3 Mineral Resource was drilled using AC and DC drillholes on a nominal 20m by 20m grid spacing. A total of 269 AC holes (9,008m), four DC holes (384m) and nine RC holes (962m) were drilled by WAF in 2015 and 2016. Hole azimuths were 090° or 225° at declination of -50° to optimally intersect the mineralised zones.

The area of the M5 Mineral Resource was drilled using AC, RC and DC drillholes on a nominal 50m by 25m grid spacing with infill to 25m by 25m in the far south portion. A total of 675 AC holes (22,088m), 40 DC holes (7,480m) and 31 RC holes (3,515m) were drilled by WAF in 2013 and 2014. A total of 60 RC holes (7,296.2m) and 71 DC holes (15,439.6m) were drilled by Channel between 2010 and 2012. Hole azimuths were 120° or 300° magnetic at declinations of between -50° and -60°, to optimally intersect the mineralised zones.

14.3.3 Interpretation and Modelling Geological Interpretation Oxidation and lithology boundaries were interpreted based on grade information and geological observations, resulting in modelled wireframes used to constrain resource estimations. Interpretation and digitising of all constraining boundaries was undertaken on cross sections at the appropriate orientation for the relevant deposit (drill line orientation). The resultant digitised boundaries were used to construct wireframe surfaces or solids defining the three-dimensional geometry of each interpreted feature.

Mankarga 1 At M1 North, mineralisation largely lies within a moderately dipping shear zone corridor (Zone 700) which can be traced over a strike of at least 300m (Figures 14.14 and 14.15). Overall strike direction of the shear zone is approximately 135° with a dip of approximately 60° to the northeast (Figure 14.14.16). A smaller, less well mineralised parallel structure exists to the northwest (Zone 800, Figure 14.15).

Mineralised domains were constructed on approximately 25m spaced cross sections orientated perpendicular to drilling using a nominal 0.3g/t Au edge cutoff to define overall shear zone mineralisation. Good continuity of the mineralised envelopes was observed at the scale of the sectional spacing with moderate grade variability typical for deposits of this type.

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Figure 14.15 Isometric NE View of M1

Mineralisation at M1 South exists in a similar setting to, and may be structurally offset from, M1 North. The observed mineralisation envelope is, however, more diffuse and less well defined, possibly indicating a structurally complex setting. Mineralisation has been interpreted to occur in discrete blocks which have been faulted and subsequently rotated from the plane of the main structure (Figures 14.14 and 14.15). Additionally, extremely high grade intervals occur within the overall mineralised envelope which variably demonstrate good to poor continuity.

In addition to the broad mineralised domains that were constrained using a nominal 0.3g/t Au edge cutoff (Zone 100 to Zone 600 and Zone 900), a number of high grade mineralised domains largely internal to the broader domains were interpreted (Zone 1000 to Zone 5000, Figure 14.17). Gold grades in excess of 50g/t have been observed and grade continuity was generated at a higher cutoff, nominally 10g/t Au. However, within the higher grade intercepts the internal grades can fall below 10g/t Au. In these cases, the lower cutoff grade was lowered to enforce grade continuity. Continuity of extremely high grade zones would likely improve with a tighter drillhole spacing.

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Figure 14.16 M1 North Section SW 0200 (Zone 700)

Mankarga 3 Mineralisation at M3 to date has mainly been intersected within shallower oxide material. Few drillholes have intersected the mineralised bodies at depth in fresh rock. This is largely a function of the drilling method used, the majority of which was AC. Mineralisation is separated into eastern and western zones, each consisting of a number of separate sub- vertical en-echelon mineralised lodes (Figure 14.18).

Mineralised domains were constructed on approximately 20m spaced cross sections orientated perpendicular to drilling (azimuth approx. 225°) using a nominal 0.3g/t Au edge cutoff for overall oxide mineralisation. Moderate continuity of the mineralised envelopes is observed at the scale of the sectional spacing.

Mankarga 5 Mineralisation at M5 largely lies within a broad, steeply dipping low grade shear zone corridor that can be traced over a strike length of 3,000m (Figure 14.14). The overall strike direction of the shear zone is approximately 035° and is sub vertical to steeply west dipping (Figure 14.19). Distinctive higher grade hangingwall and footwall lodes can be identified within this broader lower grade halo.

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Figure 14.17 M1 South Section SE 0450

Figure 14.18 Isometric SW View of M3

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Mineralisation interpretations were subsequently reviewed in multiple orientations and in plan and section view prior to being accepted for grade estimation and block modelling purposes. Three main mineralisation estimation domains (Zones 100, 200, 300) were defined, with a fourth domain (Zone 400), to the south. Zone 102 is a minor sub-domain located adjacent to and in the hangingwall of Zone 100. Previous resource estimations included another sub-domain (Zone 101), interpreted as a potential strike extension of Zone 100. This sub-domain has now been included in Zone 100. Zone 301 is a minor domain located in the footwall of Zone 300.

14.3.4 Data Flagging and Compositing Drillhole samples were flagged with the mineralisation and the oxidation wireframes described in the previous section. Coding was undertaken on the basis that if the individual sample centroid fell within the mineralisation wireframe boundary it was coded as within the mineralisation wireframe. Each domain was assigned a unique numerical code to allow the application of hard boundary domaining where required during grade estimation. Numerical codes as described in the previous section have been assigned to the mineralisation domains.

The lengths of the samples were statistically assessed prior to selecting an appropriate composite length for undertaking statistical analyses, variography and grade estimation. A total of 70,421 assayed intervals were available for analysis. Summary statistics of the sample length indicate that approximately 65.5% of the samples were collected at 1m intervals or less, 8.7% were collected at 2m intervals, 15.2% were sampled at intervals between 1m and 2m and 10.5% at intervals of 3m or more including 10% at 4m intervals (mainly AC sampling).

After consideration of relevant factors relating to geological setting and mining, including likely mining selectivity and bench/flitch height, a regular 3m run length (downhole) composite was selected as the most appropriate composite interval to equalise the sample support at M5. In the case of M1 and M3, a 2m downhole length was determined to be more appropriate. Compositing was broken when the routine encountered a change in flagging (wireframe boundary) and composites with residual intervals of less than 1.5m within the wireframe were excluded from the estimation.

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Figure 14.19 Typical Cross Section (Section NE 350 local grid) of Mineralisation Domains

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Figure 14.20 Isometric SE View of Mineralisation Domains

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14.3.5 Statistical Analysis Descriptive Statistics The composites flagged as described in the previous section were used for subsequent statistical, geostatistical and grade estimation investigations. Summary descriptive statistics were generated for all domains (Table 14.24). The grade distributions are typical for gold deposits of this style and show a positive skew or near lognormal behaviour (Figures 14.21 to 14.23). The coefficient of variation is moderate to high, consistent with the presence of high grade outliers that potentially require cutting (capping) for grade estimation.

Table 14.24 Summary Statistics Grouped by Domain for 2m and 3m Composites of Uncut Gold Grade (g/t)

Mankarga 1 North (2m) 700 800 Count 279 54 Minimum 0.002 0.024 Maximum 50.179 32.722 Mean 1.886 1.837 Std. Dev. 4.185 4.76 Variance 17.518 22.662 CV 2.219 2.591 Mankarga 1 South (2m) 100 200 300 400 500 600 Count 228 24 56 154 206 22 Minimum 0.004 0.01 0.005 0.007 0.002 0.037 Maximum 14.353 5.543 8.031 31.885 24.659 26.849 Mean 1.404 0.981 1.042 1.107 0.856 2.957 Std. Dev. 2.036 1.273 1.652 3.15 2.171 6.804 Variance 4.143 1.619 2.727 9.921 4.713 46.297 CV 1.45 1.297 1.586 2.846 2.538 2.301 Mankarga 1 South High Grade (2m) 1000 1001 2000 3000 4000 5000 Count 39 15 4 8 16 8 Minimum 0.18 2.902 10.667 0.512 0.022 1.9 Maximum 553.869 315.334 333.395 85.075 171.805 83.468 Mean 37.362 75.011 115.903 18.429 25.492 24.549 Std. Dev. 95.421 78.559 127.269 26.205 41.565 26.64 Variance 9,105.24 6,171.52 16,197.50 686.675 1,727.65 709.681 CV 2.554 1.047 1.098 1.422 1.631 1.085 Mankarga 3 (2m) 100 200 300 400 500 600 Count 48 82 21 35 59 54 Minimum 0.03 0.028 0.084 0.093 0.08 0.01 Maximum 5.972 5.964 3.838 7.144 11.439 5.457 Mean 1.496 0.947 1.332 1.461 2.477 1.336 Std. Dev. 1.271 0.987 1.168 1.382 2.615 1.447 Variance 1.615 0.975 1.365 1.911 6.839 2.094 CV 0.850 1.042 0.877 0.946 1.056 1.083 Mankarga 5 (3m) 100 102 200 300 301 400 Count 2604 117 1566 641 57 84 Minimum 0.007 0.016 0.008 0.006 0.11 0.014 Maximum 33.38 17.246 9.124 7.236 2.625 7.833 Mean 1.077 0.862 0.741 0.457 0.514 0.817 Std. Dev. 1.945 1.685 0.932 0.66 0.423 1.139 Variance 3.784 2.84 0.868 0.436 0.179 1.297 CV 1.806 1.955 1.258 1.445 0.822 1.393

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Figure 14.21 Log Histograms of Uncut Gold Grade by Domain Mankarga 1

0.100

0.075

0.050 Frequencies

0.025

0.000 0.01 1 100 auppm M1 North Combined

0.10

0.09 0.15

0.08

0.07

0.06 0.10

0.05

Frequencies 0.04 Frequencies

0.03 0.05

0.02

0.01

0.00 0.00 0.01 1 100 0.1 10 1000 auppm auppm M1 South MIK Domains Combined M1 South High Grade Domains

Figure 14.22 Log Histograms of Uncut Gold Grade by Domain Mankarga 3

0.125 0.15

0.100

0.10 0.075 Frequencies Frequencies 0.050 0.05

0.025

0.000 0.00 0.01 0.1 1 10 0.01 1 100 auppm auppm M3 East M3 West

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Figure 14.23 Log Histograms of Uncut Gold Grade by Domain Mankarga 5

0.100 0.20

0.075 0.15

0.050 0.10 Frequencies Frequencies

0.025 0.05

0.000 0.00 0.01 1 100 0.01 1 100 auppm auppm Zone 100 Zone 102

0.10

0.09 0.100 0.08

0.07 0.075 0.06

0.05 0.050 Frequencies Frequencies 0.04

0.03

0.02 0.025

0.01

0.00 0.000 0.01 1 100 0.01 1 100 auppm auppm Zone 200 Zone 300

0.15

0.10

0.10 Frequencies Frequencies 0.05 0.05

0.00 0.01 1 100 0.01 1 100 auppm auppm Zone 301 Zone 400

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Cell Declustering Analysis Visual inspection of the available datasets for each of the estimation domains indicated some clustering of the data within higher grade regions of the deposit. Data clustering often occurs when drilling campaigns selectively target higher grade regions of the deposit, resulting in an artificially high mean grade in many cases. Declustering was therefore completed to remove any effects of preferential sampling of high grade areas that may have occurred. Cell declustering was not undertaken at M3 due to the relatively low number of samples per domain.

Cell declustering was completed with weights determined as 1/n, with “n” representing the number of data in each cell. Declustered composite statistics are presented in Table 14.25. As expected, the declustered mean grades tend to be less than the composite mean grades due to the data configuration issues discussed above.

Table 14.25 Summary Statistics Grouped by Domain for 3m Composites of Uncut, Declustered Gold Grade (g/t)

Mankarga 1 North 700 800 Count 279 54 Minimum 0.002 0.024 Maximum 50.179 32.722 Mean 1.825 1.682 Std. Dev. 4.182 4.071 Variance 17.486 16.57 CV 2.292 2.421 Mankarga 1 South 100 200 300 400 500 600 Count 228 24 56 154 206 22 Minimum 0.004 0.01 0.005 0.007 0.002 0.037 Maximum 14.353 5.543 8.031 31.885 24.659 26.849 Mean 1.422 1.01 1.672 1.093 0.799 6.704 Std. Dev. 1.984 0.948 2.595 2.356 1.718 9.68 Variance 3.937 0.898 6.732 5.551 2.952 93.704 CV 1.395 0.938 1.552 2.155 2.15 1.444 Mankarga 1 South High Grade 1000 1001 2000 3000 4000 5000 Count 39 15 4 8 16 8 Minimum 0.18 2.902 10.667 0.512 0.022 1.9 Maximum 553.869 315.334 333.395 85.075 171.805 83.468 Mean 32.857 67.539 91.937 20.592 29.085 30.679 Std. Dev. 78.882 74.146 109.946 29.617 46.509 30.049 Variance 6,222.35 5,497.68 12,088.03 877.19 2,163.06 902.958 CV 2.401 1.098 1.196 1.438 1.599 0.979 Mankarga 5 100 102 200 300 301 400 Count 2604 78 1566 641 51 28 Minimum 0.007 0.016 0.008 0.006 0.11 0.04 Maximum 33.38 17.246 9.124 7.236 2.625 7.833 Mean 1.032 1.095 0.75 0.441 0.531 0.777 Std. Dev. 1.788 2.287 0.896 0.596 0.376 0.726 Variance 3.197 5.229 0.803 0.355 0.141 0.527 CV 1.732 2.088 1.194 1.351 0.708 0.934

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Table 14.26 Summary Statistics Grouped by Domain for 2m and 3m Composites of Uncut Gold Grade (g/t)

Mankarga 1 North 700 800 Count 279 54 Minimum 0.002 0.024 Maximum 16 16 Mean 1.648 1.487 Std. Dev. 2.692 2.775 Variance 7.248 7.699 CV 1.633 1.866 Mankarga 1 South 100 200 300 400 500 600 Count 228 24 56 154 206 22 Minimum 0.004 0.01 0.005 0.007 0.002 0.037 Maximum 13 5.543 8.031 13 13 13 Mean 1.415 1.01 1.672 1.052 0.776 4.176 Std. Dev. 1.938 0.948 2.595 2.002 1.457 5.46 Variance 3.757 0.898 6.732 4.009 2.122 29.816 CV 1.37 0.938 1.552 1.903 1.877 1.308 Mankarga 1 South High Grade 1000 1001 2000 3000 4000 5000 Count 39 15 4 8 16 8 Minimum 0.18 2.902 10.667 0.512 0.022 1.9 Maximum 200 200 200 85.075 171.805 83.468 Mean 25.542 61.111 69.889 20.592 29.085 30.679 Std. Dev. 38.861 54.886 62.413 29.617 46.509 30.049 Variance 1,510.21 3,012.43 3,895.39 877.19 2,163.06 902.958 CV 1.521 0.898 0.893 1.438 1.599 0.979 Mankarga 5 100 200 300 400 500 600 Count 2604 78 1566 641 51 28 Minimum 0.007 0.016 0.008 0.006 0.11 0.04 Maximum 14.5 4 9.124 5 2.625 7.833 Mean 1.023 0.862 0.75 0.438 0.531 0.777 Std. Dev. 1.676 0.857 0.896 0.564 0.376 0.726 Variance 2.81 0.735 0.803 0.318 0.141 0.527 CV 1.639 0.994 1.194 1.288 0.708 0.934

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Statistics by Drill Type The drillhole database for the M5 gold deposit contains samples obtained by AC, RC and DC drilling methods. In order to assess the potential for any inherent bias between the different drilling methods, descriptive statistics were generated for samples grouped by drilling type. Only samples within the mineralised envelopes at M5 were considered. Descriptive statistics by drilling type are presented in Table 14.27. Review of these statistics indicated that there was no material assay bias between the drilling methods and that the data can be combined for grade estimation.

Table 14.27 Summary Statistics Grouped by Drillhole Type

AC RC DC Count 5,213 1184 5,103 Minimum 0.00 0.01 0.00 Maximum 81.77 22.20 51.31 Mean 0.85 0.86 0.81 Std. Dev. 2.48 1.78 2.16 Variance 6.13 3.18 4.65 CV 2.91 2.08 2.66

Multiple Indicator Kriging Cutoff Grades Based on the number of samples and gold grade distributions, it was decided to estimate M1 North and parts of M1 South and M5 by MIK and the remainder by OK. Indicator Kriging cutoff grades or indicator bins were selected for each domain or domain combination to be estimated by MIK. All domains were combined for M1 South, (six domains) and for M1 North (two domains). Three separate domains were investigated for M5. Cutoff grades were based upon population distributions and metal proportions above and below the mean composite value of the proposed cutoff bins. Conditional statistics for data within each domain to be estimated by MIK are listed in Tables 14.28 and 14.29.

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Table 14.28 Indicator Class Statistics M1

M1 South M1 North

Probability Grade Threshold Class Mean Probability Grade Threshold Class Mean Threshold (Au g/t) (Au g/t) Threshold (Au g/t) (Au g/t) 0.234 0.1 0.0460 0.135 0.1 0.0537 0.438 0.3 0.2098 0.248 0.25 0.1676 0.544 0.4 0.3461 0.319 0.4 0.3431 0.618 0.5 0.4503 0.390 0.5 0.4365 0.656 0.6 0.5383 0.448 0.6 0.5375 0.703 0.7 0.6286 0.510 0.7 0.6519 0.739 0.85 0.7642 0.561 0.8 0.7511 0.775 1.0 0.8994 0.603 0.9 0.8561 0.804 1.2 1.0723 0.652 1.1 0.9869 0.831 1.35 1.2901 0.700 1.35 1.2460 0.856 1.55 1.4150 0.748 1.7 1.5368 0.885 1.9 1.7415 0.794 2.2 1.9314 0.912 2.2 2.0451 0.835 2.9 2.4104 0.937 2.6 2.4065 0.874 3.6 3.1814 0.957 5.0 3.3263 0.916 5.0 4.3028 0.978 6.5 5.6862 0.955 7.0 5.7279 0.991 10.0 7.7235 0.987 15.0 11.7836 Max Max 22.5527 Max Max 31.9606

Table 14.29 Indicator Class Statistics M5

Zone 100 Zone 200 Zone 300

Grade Grade Grade Probability Class Mean Probability Class Mean Probability Class Mean Threshold Threshold Threshold Threshold (Au g/t) Threshold (Au g/t) Threshold (Au g/t) (Au g/t) (Au g/t) (Au g/t) 0.164 0.2 0.1187 0.229 0.2 0.1175 0.204 0.12 0.0652 0.284 0.3 0.2494 0.360 0.3 0.2443 0.402 0.22 0.1680 0.415 0.4 0.3469 0.457 0.4 0.3435 0.542 0.3 0.2549 0.510 0.5 0.4492 0.551 0.5 0.4495 0.667 0.4 0.3419 0.586 0.6 0.5442 0.616 0.6 0.5534 0.754 0.5 0.4453 0.645 0.7 0.6488 0.679 0.7 0.6532 0.832 0.65 0.5688 0.692 0.8 0.7462 0.736 0.85 0.7643 0.888 0.85 0.7582 0.739 0.95 0.8740 0.785 1.0 0.9207 0.925 1.1 0.9497 0.779 1.15 1.0400 0.823 1.15 1.0669 0.956 1.5 1.2320 0.818 1.35 1.2405 0.865 1.35 1.2379 0.972 2.0 1.7296 0.855 1.6 1.4676 0.899 1.6 1.4710 0.986 3.2 2.6759 0.881 1.9 1.7502 0.934 1.9 1.7432 0.997 5.0 3.8144 0.901 2.2 2.0666 0.952 2.2 2.0241 Max Max 6.6480 0.921 2.6 2.3916 0.966 2.6 2.3707 0.938 3.2 2.8937 0.980 3.5 3.0299 0.960 4.5 3.6706 0.992 4.5 4.0715 0.972 6.0 5.3160 0.998 8.0 7.0361 Max Max 9.6450 Max Max 8.9493

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14.3.6 Variography Introduction Variography is used to describe the spatial variability or correlation of an attribute (gold, silver etc.). The spatial variability is traditionally measured by means of a variogram, which is generated by determining the averaged squared difference of data points at a nominated distance (h), or lag (Srivastava and Isaacs, 1989). The averaged squared difference (variogram or γ(h)) for each lag distance is plotted on a bivariate plot, where the X-axis is the lag distance and the Y-axis represents the average squared differences (γ(h)) for the nominated lag distance.

Several types of variogram calculations are employed to determine the directions of the continuity of the mineralisation:

Mankarga1 North Variography was calculated and modelled on the two combined domains. The variography showed good structure and moderate anisotropy between the major and semi-major axes. Gaussian variography was back transformed into normal space and modelled. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a relative nugget effect.

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The interpreted major direction of continuity dips at 40º towards 338º. The semi-major and minor directions are orthogonal to this direction. The modelled Gaussian back transformed variogram plot is provided in Figure 14.24.

Figure 14.24 Mankarga 1 North Variogram

6 N45 D-

4 N284 D

3 Variogram : au cut Variogram 2 N338 D

0 0 10 20 30 40 50 60 70 80 90 100

Distance (m)

Table 14.30 presents the fitted grade and indicator variogram models for Mankarga 1 North.

The interpreted major direction of continuity dips at 40º towards 315º. The semi-major and minor directions are orthogonal to this direction. The modelled Gaussian back transformed variogram plot is provided in Figure 14.25.

Using the raw cut data, a grade variogram was calculated for the high grade domains modelled at M1 South. The variogram as modelled is presented in Figure 14.26.

Table 14.30 presents the fitted grade and indicator variogram models for M1 South.

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Figure 14.25 Mankarga 1 South Variogram

4 N50 D-

N314 D

2 Variogram : au cut au : Variogram N337 D 1

0 0 10 20 30 40 50 60 70 80 90

Distance (m)

Figure 14.26 Mankarga 1 South High Grade Variogram

5000

4000 N115

3000

N205 D

2000

Variogram : au cut : au Variogram N25 D- 1000

0 0 10 20 30 40 50 60 70 80

Distance (m)

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Table 14.30 M1 North Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 2.41 315 60 40 3.36 40 30 4 1.15 80 50 8 Indicator Variography 0.10(1) 0.0273 315 60 40 0.0469 59 34 8 0.0349 100 65 16 0.25(1) 0.0496 315 60 40 0.0771 57 32 7 0.0569 95 60 14 0.40 0.0600 315 60 40 0.0880 55 30 6 0.0650 90 55 12 0.50(2) 0.0719 315 60 40 0.0947 47 28 6 0.0709 83 53 12 0.60 0.0800 315 60 40 0.0950 38 25 6 0.0720 75 50 12 0.70(3) 0.0863 315 60 40 0.0949 38 25 6 0.0685 75 50 12 0.80 0.0900 315 60 40 0.0920 38 25 6 0.0630 75 50 12 0.90(4) 0.0934 315 60 40 0.0884 37 25 5 0.0551 73 50 10 1.10 0.0930 315 60 40 0.0820 35 25 4 0.0460 70 50 8 1.35(5) 0.0900 315 60 40 0.0739 35 25 4 0.0403 65 45 8 1.70 0.0850 315 60 40 0.0650 35 25 3 0.0345 60 40 8 2.20(6) 0.0742 315 60 40 0.0538 33 23 3 0.0295 58 38 8 2.90 0.0640 315 60 40 0.0440 30 20 3 0.0250 55 35 8 3.60(7) 0.0524 315 60 40 0.0327 25 18 3 0.0186 53 33 8 5.00 0.0370 315 60 40 0.0210 20 15 3 0.0120 50 30 8 7.00(8) 0.0192 315 60 40 0.0102 18 12 3 0.0058 48 28 6 15.0(8) 0.0051 315 60 40 0.0025 16 10 2 0.0015 45 26 5 Note: 1) Assumed model based on 0.40 Au g/t variogram model 2) Assumed model based on 0.40 Au g/t and 0.60 Au g/t variogram models 3) Assumed model based on 0.60 Au g/t and 0.80 Au g/t variogram models 4) Assumed model based on 0.80 Au g/t and 1.10 Au g/t variogram model 5) Assumed model based on 1.10 Au g/t and 1.70 Au g/t variogram model 6) Assumed model based on 1.70 Au g/t and 2.90 Au g/t variogram models 7) Assumed model based on 2.90 Au g/t and 5.00 Au g/t variogram models 8) Assumed model based on 5.00 Au g/t variogram model

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Mankarga 3 Variography was calculated and modelled on combined domains from the east and west parts of the deposit (Figures 14.27 and 14.28). Variography showed acceptable structure and displayed moderate anisotropy between the major and semi-major axes. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a relative nugget effect.

Figure 14.27 M3 East Variogram

1.5 N325

1.0 N55 D80 Variogram : auppm Variogram 0.5 N55 D-

0.0 0 10 20 30 40 50 60 70 80 90 100

Distance (m)

Figure 14.28 M3 West Variogram

6 N70 D80

4 N340

3 Variogram : auppm : Variogram 2 N70 D-

0 0 10 20 30 40 50 60 70 80

Distance (m)

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Mankarga 5 Zone 100 Variography modelled for Zone 100 showed good structure and displayed moderate anisotropy between the major and semi-major axes. Gaussian variography was back transformed into normal space and modelled. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a relative nugget effect (calculated by dividing the nugget variance by the sill variance) of 37%. The short-range structure (which was modelled with ranges of 25m, 15m and 5m for the major, semi-major and minor axis respectively) accounts for 49% of the non-nugget variance. The overall ranges fitted were 160m, 100m and 25m for the major, semi-major, and minor axis respectively.

The interpreted major direction of continuity lies in a plane that strikes at 030º with a vertical dip and displays a moderate plunge of 40º towards 030º. The semi-major and minor directions are orthogonal to this direction. The modelled Gaussian back transformed variogram plot is provided in Figure 14.29.

Figure 14.29 Zone 100 – Back-transformed Gaussian Variogram

N305 D

2 N221 D

Variogram : auppm 1 N31 D40

0 0 50 100 150 200

Distance (m)

Modelled indicator variograms displayed a range of relative nugget values from 22% to 47%. Table 14.32 presents the fitted grade and indicator variogram models for Zone 100.

Zone 200 Variography modelled for Zone 200 showed good structure and displayed moderate to strong anisotropy between the major and semi-major axes. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a moderate relative nugget effect of 37%. The short-range structure (which was modelled with ranges of 25m, 15m and 15m for the major, semi-major and minor axis respectively) accounts for 40% of the non-nugget variance. The overall ranges fitted to the Zone 200 variogram were 130m, 80m and 20m for the major, semi-major, and minor axis respectively.

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The interpreted major direction of continuity lies in a plane that strikes at 220º with a sub- vertical dip of 80º to the west and displays a moderate plunge of 50º towards 232º. The semi-major and minor directions are orthogonal to this direction. Table 14.33 presents the fitted grade variogram and indicator variogram models for Zone 200. The Gaussian back transformed variogram plot is provided in Figure 14.30.

Figure 14.30 Zone 200 – Back-transformed Gaussian Variogram

0.8

0.7 N305 D 0.6

0.5 N29 D29 0.4

0.3 Variogram : auppm : Variogram N52 D- 0.2

0.1

0.0 0 50 100 150

Distance (m)

Zone 300 Variography showed good structure and displayed moderate to strong anisotropy between the major and semi-major axes. Two spherical models were fitted to the experimental variogram, with the variogram exhibiting a moderate relative nugget effect of 40%. The short-range structure (which was modelled with ranges of 60m, 30m and 10m for the major, semi-major and minor axis respectively) accounts for 44% of the non-nugget variance. The overall ranges fitted to the Zone 300 variogram were 150m, 80m and 25m for the major, semi-major, and minor axis respectively.

The interpreted major direction of continuity lies in a plane that strikes at 225º with a sub- vertical dip of 80º to the west and a steep plunge of 60º towards 251º. The semi-major and minor directions are orthogonal to this direction. Table 14.34 presents the fitted grade variogram and indicator variogram models for Zone 300. The Gaussian back transformed variogram plot is provided in Figure 14.31.

Variography for OK Domains Variography for Domains 101 and 301 (domains estimated by OK) was assumed and was based on variography calculated for Domains 100 and 300 (the minor domains are characterised by low numbers of samples resulting in low confidence variograms). Variography for Domain 400 was calculated and modelled (Figure 14.32). Variogram models as applied to OK estimates are tabulated in Table 14.37.

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Table 14.31 M1 South Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 2.1 320 80 60 1.465 40 10 4 0.93 70 40 8 Indicator Variography 0.10(1) 0.0421 320 80 60 0.0741 34 19 6 0.0522 80 50 12 0.30(1) 0.0663 320 80 60 0.1057 32 17 5 0.0737 75 45 10 0.40 0.0790 320 80 60 0.1040 30 15 4 0.0630 70 40 9 0.50(2) 0.0795 320 80 60 0.0967 30 15 4 0.0578 70 40 9 0.60 0.0800 320 80 60 0.0900 30 15 4 0.0530 70 40 9 0.70(3) 0.0793 320 80 60 0.0802 27.5 15 4 0.0490 65 37.5 8.5 0.85 0.0780 320 80 60 0.0710 25 15 4 0.0450 60 35 8 1.00(4) 0.0742 320 80 60 0.0632 25 15 4 0.0407 60 32.5 8 1.20 0.0690 320 80 60 0.0550 25 15 4 0.0360 60 30 8 1.35(5) 0.0650 320 80 60 0.0501 25 15 4 0.0323 57.5 30 8 1.55 0.0550 320 80 60 0.0410 25 15 4 0.0260 55 30 8 1.90(6) 0.0488 320 80 60 0.0350 25 15 4 0.0220 55 30 8 2.20 0.0420 320 80 60 0.0290 25 15 4 0.0180 55 30 8 2.60(7) 0.0381 320 80 60 0.0248 25 15 4 0.0155 52.5 30 7 5.00 0.0260 320 80 60 0.0160 25 15 4 0.0100 50 30 6 6.50(8) 0.0163 320 80 60 0.0091 23 13 3 0.0060 48 28 5 10.0(8) 0.0114 320 80 60 0.0059 21 11 2 0.0038 46 26 4 Note: 1) Assumed model based on 0.40 Au g/t variogram model 2) Assumed model based on 0.40 Au g/t and 0.60 Au g/t variogram models 3) Assumed model based on 0.60 Au g/t and 0.85 Au g/t variogram models 4) Assumed model based on 0.85 Au g/t and 1.20 Au g/t variogram model 5) Assumed model based on 1.20 Au g/t and 1.55 Au g/t variogram model 6) Assumed model based on 1.55 Au g/t and 2.20 Au g/t variogram models 7) Assumed model based on 2.20 Au g/t and 5.00 Au g/t variogram models 8) Assumed model based on 5.00 Au g/t variogram model

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Table 14.32 M5 Zone 100 Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 1.33 35 -85 140 1.32 25 15 5 0.55 160 90 25 Indicator Variography 0.20(1) 0.0302 32 90 40 0.0596 40 25 8 0.0612 140 65 24 0.30(1) 0.0458 32 90 40 0.0850 38 20 7 0.0872 130 60 22 0.40 0.0550 32 90 40 0.0950 35 15 7 0.0970 120 50 20 0.50(2) 0.0590 32 90 40 0.0953 33 15 7 0.0947 115 48 19 0.60 0.0600 32 90 40 0.0910 30 15 6 0.0880 110 45 17 0.70(3) 0.0604 32 90 40 0.0841 30 14 5 0.0795 105 45 16 0.80 0.0600 32 90 40 0.0770 30 12 4 0.0710 100 45 15 0.95(4) 0.0570 32 90 40 0.0675 28 12 4 0.0625 88 40 14 1.15 0.0530 32 90 40 0.0580 25 12 4 0.0540 75 35 12 1.35(5) 0.0464 32 90 40 0.0467 25 12 4 0.0439 70 35 11 1.60 0.0420 32 90 40 0.0390 25 12 4 0.0370 65 35 10 1.90(6) 0.0366 32 90 40 0.0321 24 12 4 0.0303 63 33 9 2.20 0.0320 32 90 40 0.0265 23 12 4 0.0250 60 30 8 2.60(7) 0.0291 32 90 40 0.0204 21 11 4 0.0185 55 25 7 3.20 0.0250 32 90 40 0.0150 18 9 4 0.0130 50 20 6 4.50(8) 0.0182 32 90 40 0.0105 16 8 3 0.0093 45 15 5 6.00(8) 0.0115 32 90 40 0.0062 14 7 2 0.0053 40 10 4 Note: 1) Assumed model based on 0.40 Au g/t variogram model 2) Assumed model based on 0.40 Au g/t and 0.60 Au g/t variogram models 3) Assumed model based on 0.60 Au g/t and 0.80 Au g/t variogram models| 4) Assumed model based on 0.80 Au g/t and 1.15 Au g/t variogram model 5) Assumed model based on 1.15 Au g/t and 1.60 Au g/t variogram model 6) Assumed model based on 1.60 Au g/t and 2.20 Au g/t variogram models 7) Assumed model based on 2.20 Au g/t and 3.20 Au g/t variogram models 8) Assumed model based on 3.20 Au g/t variogram model

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Table 14.33 M5 Zone 200 Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 1.21 232 75 50 1.62 30 20 5 0.48 160 100 20 Indicator Variography 0.20(1) 0.0317 232 75 50 0.0634 34 15 10 0.0810 180 130 35 0.30(1) 0.0437 232 75 50 0.0828 32 15 10 0.1035 170 120 32 0.40 0.0500 232 75 50 0.0870 30 15 8 0.1100 160 110 30 0.50(2) 0.0564 232 75 50 0.0859 26 15 7 0.1047 140 100 27.5 0.60 0.0600 232 75 50 0.0810 22 15 6 0.0950 120 90 25 0.70(3) 0.0642 232 75 50 0.0706 22 15 6 0.0832 115 85 22.5 0.85 0.0650 232 75 50 0.0590 22 15 6 0.0700 110 80 20 1.00(4) 0.0619 232 75 50 0.0471 21 13.5 6 0.0600 105 70 17.5 1.15 0.0580 232 75 50 0.0370 20 12 6 0.0510 100 60 15 1.35(5) 0.0477 232 75 50 0.0290 17.5 11 6 0.0404 95 55 15 1.60 0.0380 232 75 50 0.0220 15 10 6 0.0310 90 50 15 1.90(6) 0.0267 232 75 50 0.0147 15 10 5.5 0.0206 75 45 15 2.20 0.0200 232 75 50 0.0105 15 10 5 0.0147 60 40 15 2.60(7) 0.0156 232 75 50 0.0091 13.5 10 5 0.0083 50 32.5 15 3.50 0.0095 232 75 50 0.0060 12 10 5 0.0033 40 25 15 4.50(8) 0.0042 232 75 50 0.0024 10 8 4 0.0014 35 20 12 8.00(8) 0.0011 232 75 50 0.0006 8 6 3 0.0003 30 18 10 Note: 1) Assumed model based on 0.40 Au g/t variogram model 2) Assumed model based on 0.40 Au g/t and 0.60 Au g/t variogram models 3) Assumed model based on 0.60 Au g/t and 0.80 Au g/t variogram models| 4) Assumed model based on 0.80 Au g/t and 1.15 Au g/t variogram model 5) Assumed model based on 1.15 Au g/t and 1.60 Au g/t variogram model 6) Assumed model based on 1.60 Au g/t and 2.20 Au g/t variogram models 7) Assumed model based on 2.20 Au g/t and 3.20 Au g/t variogram models 8) Assumed model based on 3.20 Au g/t variogram model

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Figure 14.31 Zone 300 – Back-transformed Gaussian Variogram

0.3 N315 D

0.2 N219 D Variogram : auppm 0.1 N242 D

0.0 0 50 100 150

Distance (m)

Figure 14.32 Zone 400 – Back-transformed Gaussian Variogram

0.3 N135

0.2 D-90

Variogram : auppm Variogram 0.1 N45

0.0 0 25 50 75 100 125

Distance (m)

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Table 14.34 M5 Zone 300 Variogram Models

Structure 1 Structure 2 Rotation Grade Variable Nugget Isatis Geological Plane Convention or Indicator Relative Range (m) Relative Range (m) (C0) Threshold Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor Back-transformed Gaussian Variography Gold (Au g/t) 0.14 225 80 60 0.15 60 30 10 0.05 150 80 25 Indicator Variography 0.12(1) 0.0355 225 80 60 0.0582 30 15 6 0.0483 150 90 30 0.22(1) 0.0635 225 80 60 0.0940 30 15 6 0.0776 140 85 28 0.30 0.0700 225 80 60 0.0970 30 15 6 0.0800 130 80 25 0.40(2) 0.0633 225 80 60 0.0822 30 15 6 0.0695 125 70 24 0.50 0.0550 225 80 60 0.0670 30 15 6 0.0580 120 60 23 0.65(3) 0.0443 225 80 60 0.0493 27.5 15 5 0.0444 110 55 21.5 0.80 0.0335 225 80 60 0.0340 25 15 4 0.0320 100 50 20 1.10(4) 0.0237 225 80 60 0.0235 25 15 4 0.0218 95 47.5 17.5 1.50 0.0150 225 80 60 0.0145 25 15 4 0.0132 90 45 15 2.00(5) 0.0101 225 80 60 0.0095 25 15 4 0.0084 85 42.5 12.5 3.20. 0.0050 225 80 60 0.0046 25 15 4 0.0040 80 40 10 5.00(8) 0.0012 225 80 60 0.0010 20 10 3 0.0008 70 37 8 Note: 1) Assumed model based on 0.30 Au g/t variogram model 2) Assumed model based on 0.30 Au g/t and 0.50 Au g/t variogram models 3) Assumed model based on 0.50 Au g/t and 0.80 Au g/t variogram models 4) Assumed model based on 0.80 Au g/t and 1.50 Au g/t variogram model 5) Assumed model based on 1.50 Au g/t and 3.20 Au g/t variogram models 6) Assumed model based on 3.20 Au g/t variogram model

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Table 14.35 M1 Variogram Models for OK Estimates

Table 14.36 M3 Variogram Models for OK Estimates

Structure 1 Structure 2 Rotation Grade Variable or Nugget Isatis Geological Plane Convention Relative Range (m) Relative Range (m) Indicator Threshold (C0) Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor 100 to 300, Gold (Au g/t) 0.5 330 80 0 0.4 14 14 4 0.52 60 45 10 400 to 600, Gold (Au g/t) 0.5 17 90 0 0.4 14 14 4 0.52 60 45 10

Table 14.37 M3 Variogram Models for OK Estimates

Structure 2 Rotation Structure 1 Grade Variable or Nugget Isatis Geological Plane Convention Relative Range (m) Relative Range (m) Indicator Threshold (C0) Sill 1 Sill 2 Major Semi Major Minor (C1) Major Semi Major Minor (C2) Major Semi Major Minor 102, Gold (Au g/t) 1.3 35 80 0 1.32 25 15 5 0.55 160 90 25 301, Gold (Au g/t) 0.0138 45 75 0 0.057 60 30 10 0.05 150 80 25 400, Gold (Au g/t) 0.133 30 80 0 0.092 45 30 4 0.111 110 50 15

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14.3.7 Block Modelling 3-D block models were created for the four separate deposit areas in the Universal Transverse Mercator (UTM) grid projection, Zone 30 North, using the World Geodetic System 1984 datum (WGS84). The parent block size was selected on the basis of the average drill spacing. Sub-blocking was applied to an appropriate dimension to ensure adequate volume representation. The models covered all the interpreted mineralisation zones and included suitable additional waste material to allow later pit optimisation studies. Block coding was completed on the basis of the block centroid, wherein a centroid falling within any wireframe was coded with the wireframe solid attribute. Rotations were applied where appropriate to ensure suitable coverage.

The main block model parameters are summarised in Table 14.38. Variables were coded into the block model to enable MIK and OK estimation and subsequent change of support and grade tonnage reporting. A visual review of the wireframe solids and the block model indicated correct flagging of the block model. Additionally, a check was made of coded volume versus wireframe volume which confirmed the above.

Table 14.38 Block Model Parameters

M1 South Northing (Y) Easting (X) RL (Z) Min. Coordinates 1,336,650m 741,600m -20m Offset 650m 450m 350m Block size (m) 25 10 10 Sub Block size (m) 2.5 1 1 Rotation (° around axis) 0° 0° -45° M1 North Northing (Y) Easting (X) RL (Z) Min. Coordinates 1,337,400m 741,400m 80m Offset 450m 250m 250m Block size (m) 25 10 10 Sub Block size (m) 2.5 1 1 Rotation (° around axis) 0° 0° -45° M3 Northing (Y) Easting (X) RL (Z) Min. Coordinates 1,336,950m 740,600m 200m Offset 4550m 450m 120m Block size (m) 25 10 10 Sub Block size (m) 5 2.5 2.5 Rotation (° around axis) 0° 0° 0° M5 Northing (Y) Easting (X) RL (Z) Min. Coordinates 1,335,621.257m 741,559.695m -100m Offset 3,175m 900m 430m Block size (m) 25 20 10 Sub Block size (m) 12.5 5 5 Rotation (° around axis) 0° 0° 30°

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14.3.8 Dry Bulk Density Data An extensive dry bulk density database was supplied containing a total of 6,648 dry bulk density determinations widely dispersed throughout the different mineralisation domains. The database was subdivided on the basis of oxidation, resulting in 4,952 determinations of fresh rock, 593 of transition material and 1,103 of oxide material. Measured average dry bulk density values subdivided by oxidation state are presented in Table 14.39 below. Dry bulk densities were applied to the block models as average dry bulk densities per oxidation state as per Table 14.39.

Table 14.39 Measured Average Dry Bulk Density Value

Measured Average Bulk Deposit Domain Lithology Density (t/m³) Strongly oxidised 1.81 Moderately oxidised 2.62 M1 North all Transition 2.79 Fresh 2.81 Strongly oxidised 2.00 Moderately oxidised 2.57 M1 South all Transition 2.78 Fresh 2.79 Strongly oxidised 1.90 Moderately oxidised 2.53 M3 all Transition 2.61 Fresh 2.75 Strongly oxidised 2.01 Moderately oxidised 2.46 100 Transition 2.71 Fresh 2.75 M5 Strongly oxidised 2.10 Moderately oxidised 2.51 All other material Transition 2.71 Fresh 2.75

14.3.9 Grade Estimation Introduction MIK was used as the grade estimation method within the defined mineralisation wireframes at M1 and M5. OK was used as the grade estimation method for the minor domains at M5, the high grade domains at M1 and for all domains at M3. Estimation was completed in the mining package Vulcan using the GSLib geostatistical software. Change of support was applied to the MIK grade estimates to produce ‘recoverable’ gold estimates targeting a selective mining unit (SMU) of 5mE by 12.5mN by 5mRL.

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The Multiple Indicator Kriging Method See Section 14.2.9 for a description of the MIK method applied.

Post MIK Processing - E-Type Estimates See section 14.2.9 for a description of the post MIK processing applied

Support Correction (Selective Mining Unit Estimation) See section 14.2.9 for a description of the Support Correction method applied

Indirect Lognormal Correction See section 14.2.9 for a description of the indirect lognormal correction applied

Mankarga 1 North At M1 North, a total of 17 indicator thresholds were estimated for the applicable domains. The sample search parameters are provided in Table 14.40. A hard domain boundary approach was used for the estimation whereby the MIK estimation was only allowed to select composites from the applicable domain. Where required, a two-pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations (10mE by 25mN by 10mRL) were applied throughout and discretisation was applied on the basis of 3mE by 3mN by 2mRL, resulting in 18 discretisation points per block.

Table 14.40 M1 North MIK Sample Search Criteria

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Mankarga 1 South At M1 south, a total of 17 indicator thresholds were estimated for the applicable domains. The sample search parameters are provided in Table 14.41. A semi-hard domain boundary approach was used for the estimation whereby the MIK estimation was allowed to select one composite from the adjacent high grade domain where applicable. Where required, a two-pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations (10mE by 25mN by 10mRL) were applied throughout and discretisation was applied on the basis of 3mE by 3mN by 2mRL, resulting in 18 discretisation points per block.

Table 14.41 M1 South MIK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH Pass 1 -60/308 28/285 10/021 70 60 20 18 24 4 100 Pass 2 -60/308 28/285 10/021 1000 1000 100 18 24 - Pass 1 -60/308 28/285 10/021 100 800 30 18 24 4 200 ------Pass 1 -60/308 28/285 10/021 70 60 20 18 24 - 300 Pass 2 -60/308 28/285 10/021 1000 1000 100 18 24 - Pass 1 -60/014 15/312 25/050 70 60 20 18 24 4 400 Pass 2 -60/014 15/312 25/050 1000 1000 100 18 24 - Pass 1 -60/014 15/312 25/050 70 60 20 18 24 4 500 Pass 2 -60/014 15/312 25/050 1000 1000 100 18 24 - Pass 1 -60/014 15/312 25/050 1000 1000 100 18 24 4 600 ------

Mankarga 5 In the case of M5, a total of 12 indicator thresholds were estimated for Zone 300 and 17 indicator thresholds were estimated for Zones 100 and 200.

The sample search parameters are provided in Table 14.42. Domain 100 was subdivided into north and south portions to allow for a subtle change in orientation that is evident around 1,336,400mN. Hard domain boundaries were used for the estimation throughout, with the exception of Domain 100, where samples from the north portion were allowed to be selected during estimation of the south portion and vice versa. A two-pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations (20mE by 25mN by 10mRL) were applied throughout and discretisation was applied on the basis of 3mE by 3mN by 2mRL, resulting in 18 discretisation points per block.

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Table 14.42 M1 South MIK Sample Search Criteria

The sample search parameters are provided in Table 14.43 to Table 14.45. Hard domain boundaries were used for the OK estimation throughout, with the exception of M1 South where a semi-hard boundary was applied allowing the OK estimation to select one composite from the adjacent low grade domain where applicable. Where necessary, a two- pass estimation strategy was applied to each domain, applying a progressively expanded and less restrictive sample search to the successive estimation pass, and only considering blocks not previously assigned an estimate. Parent cell estimations (5mE by 12.5mN by 5mRL) were applied throughout and discretisation was applied on the basis of 2mE by 3mN by 2mRL, resulting in 12 discretisation points per block.

Table 14.43 Mankarga 1 South OK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH 1000 Pass 1 -60/308 28/285 10/021 100 100 30 4 8 3 1001 Pass 1 -60/308 28/285 10/021 100 100 30 3 8 - 2000 Pass 1 -60/308 28/285 10/021 100 100 30 4 8 3 3000 Pass 1 -60/308 28/285 10/021 100 100 30 4 8 3 4000 Pass 1 -60/014 15/312 25/050 100 100 30 4 8 3 5000 Pass 1 -60/014 15/312 25/050 100 100 50 4 8 3

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Table 14.44 Mankarga 3 OK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH 100 Pass 1 0/330 80/240 10/060 120 120 30 6 8 3 200 Pass 1 0/325 50/235 40/55 100 100 30 6 8 3 300 Pass 1 0/330 80/240 10/060 120 120 30 6 8 3 400 Pass 1 0/017 90/000 0/107 120 120 30 6 8 3 500 Pass 1 0/345 90/000 0/075 120 120 30 6 8 3 600 Pass 1 0/335 90/000 0/065 120 120 30 6 8 3

Table 14.45 Mankarga 5 OK Sample Search Criteria

Sample Search Orientation Sample Search Distance Numbers of Samples (dip/dip direction˚) (m) Domain Pass Semi Semi Max per Major Minor Major Minor Min. Max. Major Major DH Pass 1 0/32 -85/302 5/302 150 50 30 8 12 3 102 Pass 2 0/32 -85/302 5/302 300 150 50 6 12 - Pass 1 0/32 -85/302 5/302 150 150 30 8 12 3 301 Pass 2 0/32 -85/302 5/302 300 300 30 6 12 - Pass 1 0/32 -85/302 5/302 150 150 30 8 12 4 400 Pass 2 0/32 -85/302 5/302 300 300 50 6 12 -

Estimate Validation All relevant statistical information was recorded to enable validation and review of the MIK and OK estimates. The recorded information included:

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The variance adjustment ratios are provided in Table 14.46.

Table 14.46 Variance Adjustment Ratios (5mE x 12.5mN x 5mRL SMU)

Mankarga 1 North Zone All MIK Domains Variance adjustment factor (f) 0.1 Mankarga 1 South Zone All MIK Domains Variance adjustment factor (f) 0.11 Mankarga 5 Zone 100 200 300 Variance adjustment factor (f) 0.25 0.25 0.27

Resource Classification The resource categorisation was based on the robustness of the various data sources available, including:

. Geological knowledge and interpretation. . Variogram models and the ranges of the first structure in multi-structure models. . Drilling density and orientation. . Estimation quality statistics.

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The resource estimates for the Sanbrado Gold Project have been classified as Indicated and Inferred Mineral Resources based on the confidence levels of the key criteria as presented in Table 14.47.

Table 14.47 Confidence Levels by Key Criteria.

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Applying these confidence levels, resource classification codes were assigned to the block model using the following criteria:

. Indicated Mineral Resources

 Blocks are predominately estimation pass 1.

 Distance to nearest data of 35m or less.

 Drillhole section spacing predominantly 50m.

 Appropriate quality of estimate statistics. . Inferred Mineral Resources

 Blocks are predominantly estimation pass 2.

 Blocks not assigned as Indicated.

14.3.10 Resource Reporting The summary total resource for the Sanbrado Gold Project is provided in Table 14.48. The Mineral Resource estimates are predicated on an open pit mining and conventional CIL processing scenario and have been reported within optimised pit shells using a gold price of US$1650/oz and a cutoff grade of 0.3g/t Au for oxide mineralisation and 0.4g/t Au for fresh rock. Key pit optimisation input parameters are as follows:

Table 14.48 Mineral Resource Statement, Sanbrado Gold Project, 20 February 2017 Summary Grade Tonnage Report, MIK and OK estimates, (SMU dimension 5mE x 12.5mN x 5mRL)

Indicated Resource Inferred Resoure Cutoff

Grade Grade (Au g/t) Tonnes Au Oz Tonnes Au Oz (g/t) (g/t) 0.5 27,660,000 1.2 1,049,000 17,360,000 1.3 712,000 M5 1 11,100,000 1.9 670,000 7,810,000 2.0 495,000 0.5 960,000 7.2 224,000 270,000 8.6 74,000 M1 South 1 610,000 11.0 215,000 210,000 10.8 73,000 0.5 780,000 1.9 49,000 660,000 1.9 41,000 M1 North 1 610,000 2.3 45,000 480,000 2.3 36,000 0.5 170,000 2.0 11,000 260,000 1.4 12,000 M3 1 130,000 2.3 10,000 180,000 1.7 10,000 0.5 29,570,000 1.4 1,332,000 18,550,000 1.4 839,000 Totals 1 12,453,737 2.3 940,000 8,680,000 2.2 614,000

Note: Mineral resources are not mineral reserves and have not demonstrated economic viability. All figures have been rounded to reflect the relative accuracy of the estimates.

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The effective date of this Mineral Resource is 20 February 2017. It is not anticipated that this Mineral Resource estimate will be materially affected, to any extent, by any known environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors.

Cautionary Note About Mineral Resources: Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. Mineral Resource estimates do not account for mineability, selectivity, mining loss and dilution. The Mankarga mineral resource estimates include Inferred Mineral Resources that are normally considered too low confidence to have economic considerations applied to them that would enable them to be categorised as Mineral Reserves. There is no certainty that these Inferred Mineral Resources will be converted to the Measured and Indicated categories through further drilling.

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MINERAL RESERVE ESTIMATES Introduction Mineral Reserves for the Sanbrado Gold Project have been estimated in accordance with NI 43-101 guidelines, and have excluded the use of Inferred Mineral Resources. Inferred Mineral Resources are too low confidence to be converted to Mineral Reserves. Mineral Reserves and associated project economics have not been updated in this report. Information from the February 20, 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018.

Mineral Reserve Mineral Reserves were estimated based on the resources contained within pit designs including an allowance for mining dilution. Only Indicated Mineral Resources were used in the pit optimisation and design process. Consequently the Mineral Reserve comprises only Probable Reserves.

Mineral Reserves contained within the pit designs are summarised in Table 15.1. These Mineral Reserves were used in the preparation of production schedules and cashflow analyses.

Table 15.1 Mineral Reserve by Category

Tonnes Gold Grade Contained Gold Mining Inventory (Mt) (g/t) (koz) Proven 0.0 0.0 0 Probable 16.8 1.7 894 Mineral Reserve 16.8 1.7 894

The Mineral Reserves are reported with an effective date of 6 April 2017, and by definition have taken into account environmental, permitting, legal title, taxation, socio-economic, marketing and political factors and constraints, as discussed in Section 20 and Section 21 of this report. The Mineral Reserves are adequate to support mine planning.

Pit Optimisation Whittle 4D software was used to determine the optimum shell upon which the pit design was based. It is important to note that the capital costs associated with construction of the project are ignored in the optimisation process. Optimisations were run for the M1 North, M1 South, M3 and M5 deposits.

In order to produce a range of ‘nested’ pit shells, the optimisations were run over a wide range of gold prices from US$500/oz to US$1800/oz. A gold price of US$1200/oz was used to analyse the cashflows indicated by the pit optimisations.

Table 15.2 shows the optimisation results for the selected pit shells. Details for each deposit are shown in Figures 15.1 to 15.4. Each of the optimisation results show an undiscounted cashflow and two discounted cashflows based on best and worst case scenarios. The best case scenario simulates the mining of each individual nested shell completely before the next nested shell is mined. This is not generally feasible in practice given that the minimum mining width required to mine a stage is greater than the actual

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width of the nested shell. The so-called worst case scenario simulates the mining of an entire bench of a pit before the next bench is mined.

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Table 15.2 Selected Optimum Pit Shells

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Figure 15.1 Mankarga 1 North Deposit, Summary Optimisation Results

Figure 15.2 Mankarga 1 South Deposit, Summary Optimisation Results

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Figure 15.3 Mankarga 3 Deposit, Summary Optimisation Results

Figure 15.4 Mankarga 5 Deposit, Summary Optimisation Results

The optimisation results are presented in context of sensitivities, risks, contained ounces, mine life and total project size. The pit shell that generated the maximum undiscounted cashflow for each deposit was selected as the optimum pit shell. It is often the case that a lower revenue factor would be selected, based on the discounted shell values, however in this case it can be seen from Figures 15.1 – 15.4 that the value graphs are flat around the maximum undiscounted shell, which indicates that the selection is robust and relatively insensitive to modest changes in economic factors.

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15.3.1 Pit Optimisation Assumptions Slopes Pit slope angles used in the M5 pit optimisation are shown in Table 15.3. Pit slope angles used for M1 and M3 are shown in Table 15.4.

Table 15.3 Mankarga 5 Pit Slope Angles

Wall Level Wall Design Parameters Face Height 10m Surface to Base of Inter-ramp slope Overall slope angle Face Angle 55⁰ Strong Oxidation angle 39.8⁰ incl. ramps 33⁰ Berm Width 5m Face Height 20m Top of Moderate Inter-ramp slope Overall slope angle Eastern Wall Oxidation to Base of Face Angle 60⁰ angle 47.2⁰ incl. ramps 38.4⁰ Oxidation Berm Width 7m Face Height 20m Below Base of Inter-ramp slope Overall slope angle Face Angle 65⁰ Oxidation angle 50.8⁰ incl. ramps 41⁰ Berm Width 7m Face Height 10m Surface to Base of Inter-ramp slope Overall slope angle Face Angle 55⁰ Strong Oxidation angle 39.8⁰ incl. ramps 33⁰ Berm Width 5m Face Height 20m Top of Moderate Western and Inter-ramp slope Overall slope angle Oxidation to Base of Face Angle 60⁰ End Walls angle 47.2⁰ incl. ramps 38.4⁰ Oxidation Berm Width 7m Face Height 20m Below Base of Inter-ramp slope Overall slope angle Face Angle 70⁰ Oxidation angle 54.5⁰ incl. ramps 43.7⁰ Berm Width 7m

Table 15.4 Mankarga 1 and Mankarga 3 Pit Slope Angles

Wall Level Wall Design Parameters Face Height 10m Surface to Base of Inter-ramp slope Overall slope angle Face Angle 50⁰ Moderate Oxidation angle 37⁰ incl. ramps 30⁰ Western Berm Width 5m Wall Face Height 20m Below Base of Inter-ramp slope Overall slope angle Face Angle 65⁰ Moderate Oxidation angle 51⁰ incl. ramps 38⁰ Berm Width 7m Face Height 10m Surface to Base of Inter-ramp slope Overall slope angle Face Angle 50⁰ Moderate Oxidation angle 37⁰ incl. ramps 30⁰ Eastern and Berm Width 5m End Walls Face Height 20m Below Base of Inter-ramp slope Overall slope angle Face Angle 70⁰ Moderate Oxidation angle 54.5⁰ incl. ramps 42⁰ Berm Width 7m

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Mining Costs Mining costs were source by quotation from mining contractors experienced in West Africa. Tables 15.5 and 15.6 outline the average contract mining costs for the M5 and the M1 and M3 deposits respectively.

Table 15.5 Mankarga 5 Average Contract Mining Costs

Excavate, Load & Haul Management Bench Drill & Blast Waste Ore Fee ($/bcm) ($/t) Crest Toe ($/bcm) 295 285 $2.36 $3.23 $0.00 $0.60 285 275 $2.64 $2.85 $0.00 $0.60 275 265 $2.68 $3.47 $0.72 $0.60 265 255 $3.08 $3.44 $0.72 $0.60 255 245 $3.39 $3.69 $1.49 $0.60 245 235 $3.63 $3.63 $1.49 $0.60 235 225 $3.76 $3.86 $1.49 $0.60 225 215 $3.75 $4.20 $2.27 $0.60 215 205 $3.76 $4.12 $2.27 $0.60 205 195 $3.74 $4.08 $2.27 $0.60 195 185 $4.04 $4.14 $2.27 $0.60 185 175 $4.03 $4.28 $2.27 $0.60 175 165 $3.97 $4.44 $2.27 $0.60 165 155 $3.99 $4.77 $2.27 $0.60 155 145 $4.35 $4.69 $2.27 $0.60 145 135 $4.39 $4.75 $2.27 $0.60 135 125 $4.41 $4.76 $2.27 $0.60 125 115 $3.65 $4.66 $2.27 $0.60

Table 15.6 Mankarga 1 & Mankarga 3 Average Contract Mining Costs

Excavate, Load & Haul Management Bench Drill & Blast Waste Ore Fee ($/bcm) ($/t) Crest Toe ($/bcm) 305 295 $2.70 $3.86 $0.00 $0.60 295 285 $2.54 $3.89 $0.72 $0.60 285 275 $2.54 $3.94 $1.49 $0.60 275 265 $2.53 $3.97 $1.49 $0.60 265 255 $2.60 $4.02 $1.49 $0.60 255 245 $2.82 $4.24 $2.27 $0.60 245 235 $2.89 $4.30 $2.27 $0.60 235 225 $2.91 $4.48 $2.27 $0.60 225 215 $3.01 $4.58 $2.27 $0.60 215 205 $3.04 $4.61 $2.27 $0.60 205 195 $3.12 $4.64 $2.27 $0.60 195 185 $3.25 $4.68 $2.27 $0.60 185 175 $3.67 $5.20 $2.27 $0.60 175 165 $3.37 $4.92 $2.27 $0.60 165 155 $3.40 $5.06 $2.27 $0.60 155 145 $3.46 $5.13 $2.27 $0.60 145 135 $3.52 $5.10 $2.27 $0.60

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The average contract mining costs determined from the pit optimisations were as follows:

. M1 North: US$2.14/tonne mined . M1 South: US$2.40/tonne mined . M3: US$2.00/tonne mined . M5: US$2.25/tonne mined

Dilution and Mining Losses The Mineral Resource estimates used a combination of MIK and OK techniques to estimate the gold grade. The different estimation methods required different approaches to estimate mining dilution and ore losses. MIK grade estimates provide a recoverable resource estimate that takes mining selectivity into account, and no further mining dilution or losses need be applied.

OK was used to estimate the gold grade in the M3 deposit and a high grade zone in the M1 deposit. For the high grade zone in M1, the mining approach would attempt to minimise ore losses at the expense of some extra dilution, and a 20% dilution with no ore losses were applied. For the M3 deposit, the resource model was regularised to a “selective mining unit” size of 5mE by 5mN by 5mRL. The regularisation of the block model results in the calculation of weighted average gold grades for the total block volume, which are essentially diluted grades. Ore losses will occur where a block contains a small proportion of mineralised material and the resultant weighted average block grade falls below the cutoff grade.

Process and Fixed Costs The process throughput rate varies for differing ore types, resulting in differing unit rates for the fixed costs for each ore type. Table 15.7 summarises the fixed cost and variable process cost by ore type. The fixed costs are comprised of process and administration labour, laboratory costs, mobile equipment, general and administration costs, and owners mining supervision costs.

Table 15.7 Process and Fixed Costs

Process Fixed Process Total Ore Type Throughput Costs Variable Cost Cost (Mt/yr) (US$/t ore) (US$/t ore) (US$/t ore) Strongly Oxidised 2.50 5.58 5.72 11.30 Moderately Oxidised 2.50 5.58 7.57 13.14 Transitional 1.44 9.70 9.76 19.46 Sulphide 1.35 10.29 10.12 20.42

Other Costs Other costs included in the optimisation include the following:

. Grade control: US$0.40/t ore . Dewatering: US$0.25/t ore . Crusher Feed: US$0.65/t ore

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Process Recovery Metallurgical testwork has indicated that the recovery of gold is grade dependant. This means that the quantity of gold recovered from a block is a function of the head grade and the ore type. The recovery of any block is given by the formulae shown in Table 15.8.

Table 15.8 Process Recovery

Process Recovery Ore Type 0.1 - 5 g/t Au >5g/t Au (%) (%) Strongly Oxidised 0.015 * LN(Au g/t head grade) + 0.93 95.4% Moderately Oxidised 0.020 * LN(Au g/t head grade) + 0.915 94.7% Transitional 0.0364 * LN(Au g/t head grade) + 0.8843 94.3% Sulphide 0.0481 * LN(Au g/t head grade) + 0.8134 89.1%

As the sulphide ore at M5 and M1 had a greater grade range, additional algorithms were developed to model the testwork results. These are shown in Table 15.9.

Table 15.9 Process Recovery for M1 and M5 Sulphide Ore

M5 Sulphide Recoveries

Grade Range Process Recovery (%) 0.4 to 1 g/t 0.1237 * LN(Au g/t head) + 0.8158 1 to 3 g/t 0.075 * LN(Au g/t head) + 0.8158 3 to 20 g/t 0.0396 * LN(Au g/t head) + 0.8542 > 20g/t 97.3% M1 Sulphide Recoveries

Grade Range Process Recovery (%) 0.4 to 1 g/t 0.0589 * LN(Au g/t head) + 0.8567 1 to 3 g/t 0.0388 * LN(Au g/t head) + 0.8574 3 to 30 g/t 0.0255 * LN(Au g/t head) + 0.8721 > 30g/t 96.0%

Royalties and Refining A Government Royalty of 4% and an additional 1% community development levy were applied to the optimisation process.

A US$5.00/oz refining charge was allowed for in the optimisation.

Metal Price A gold price of $1,200 per Au ounce was assumed for pit optimisation.

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Cutoff Grades Because the recovery is based on a grade dependant formula there is no in-situ cutoff grade applied. The cutoff grade is determined by the cost of processing divided by the realisable sale value of 1 gram of recovered gold. The average recoveries by deposit and ore type as calculated from the pit optimisations are shown in Table 15.10. The average cutoff grades calculated from the recovery data are shown in Table 15.11.

Table 15.10 Process Recovery for M1 and M5 Sulphide Ore

M1 M3 M5 Strongly Oxidised 94% 94% 93% Moderately Oxidised 93% 93% 92% Transitional 92% 91% 90% Sulphide 94% 83% 86%

Table 15.11 Average Calculated Cutoff Grade (g/t Au)

M1 M3 M5 Strongly Oxidised 0.37 0.37 0.37 Moderately Oxidised 0.43 0.43 0.43 Transitional 0.62 0.63 0.63 Sulphide 0.63 0.72 0.70

Pit Design Final pit designs were prepared for each deposit to enable practical and efficient access to each bench. The designs were based on the selected optimised shells as shown in Table 15.2 and geotechnical design criteria described in Tables 15.3 and 15.4. As shown in Table 15.12, the final designs reconcile well with the optimised pit shells.

Table 15.12 Comparison of Optimised Pit Shells and Pit Designs

Processed Ore Total Waste Strip Au Grade Cont. Au (Mt) (Mt) Ratio (Mt) (g/t) (koz) Total Shell Inventory 104.1 87.7 5.4 16.4 1.70 897 Total Design Reserve 101.0 84.2 5.1 16.8 1.65 894 Variance -2.9% -4.0% 2.9% -3.1% -0.4%

The main M5 pit is 2km long and averages 300m wide and 170m deep at the southern end. The pit has been designed so the southern higher grade portion can be mined independently of the northern portion of the pit. Both the northern and southern pit will be mined in two stages (an initial starter pit and then a cutback to final limits) in order to target higher grade earlier in the schedule and defer waste movement until later in the mine life. The initial and final pit outlines are shown in Figure 15.5.

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Figure 15.5 M5 Pit Design showing Initial Pit Stages and Waste Rock Dump

The M1 deposit will be mined in two pits, a north and a south pit. The M1 South pit is approximately 540m long, 360m wide and 180m deep. The M1 North pit is 350m long, 240m wide and 90m deep. The M3 deposit has two small, predominately oxide pits less than 40m deep. The pit outlines for M1 and M3 are shown in Figure 15.6.

Figure 15.6 M1 and M3 Pit and Waste Rock Dump Designs

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The final pit inventories are given in Table 15.13.

Table 15.13 Pit Inventories

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MINING METHODS The majority of the defined mineral resources are within 200m depth from the surface and of a lode style mineralisation. The material to be excavated will be predominantly free dig from surface with blasting required deeper in the oxidation profile. Given these conditions, conventional open pit mining techniques using drill and blast with material movement by hydraulic excavator and trucks will be employed. The project scale suits 120t to 250t class excavators in a backhoe configuration matched to 95t class mine haul trucks.

Mineral Reserves and associated project economics have not been updated in this report. Information from the February 20, 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018.

Mine Operations Mining activities will be undertaken by an experienced contractor, with WAF retaining responsibility for technical services comprising mine planning, production scheduling, grade control, surveying and supervision and management of contract mining operations.

Production schedules were completed with the primary aim of providing the highest value ore to the mill as early as possible in order to maximise the value to the project. To achieve this aim, the higher grade and higher strip ratio material from the M1 South pit and the southern part of M5 are mined first. This results in a total material movement 25Mtpa to 30Mtpa for the first two years of production, reducing to 13Mtpa to 14Mtpa for the remainder of the mine life. The nominated mining fleet of 1 x 250t class excavator and 3 x 120t class excavators with matching truck fleet (95t class) has sufficient capacity for these production rates. The fleet will be reduced in size from Year 3 onwards.

16.1.1 Drilling & Blasting Four different competencies of rock will be mined, comprising both waste and crusher feed. These consist of strongly oxidised (SOX), moderately oxidised (MOX), transition or weakly oxidised (WOX), and fresh rock types (FRS). The oxide rock types are the least competent rock with fresh rock being the most competent.

Based on the geotechnical investigations and observations of drillhole core, it has been assumed that the SOX rock will not require blasting; MOX requires some light blasting and the transition and sulphide rocks are harder, requiring higher powder factors to achieve efficient loading operations. Blasts will be engineered to ensure minimum displacement of the ore so as to minimise dilution and throw.

16.1.2 Loading & Hauling The load and haul will be initially carried out by 1 x 250t class excavator and 3 x 120t class excavators and matching truck fleet (95t class). When the production rate reduces in Year 3, the 250t class excavator and associated truck fleet will leave site. Loading operations will also be supported by a 7m³ capacity wheel loader.

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16.1.3 Grade Control In order to define the boundaries between high grade, low grade and waste material, it is necessary to drill RC holes on a close spacing and assay the mineralised intercepts. It is proposed that rows of RC holes be drilled with a 6m spacing across strike and an 8m spacing along strike. This data would be combined with the exploration data to define ore block boundaries and grade. Hole direction would be normal to the strike of mineralisation with a declination of -60°. Samples will be collected for assay at 1m intervals. Grade control results will be reconciled with the block model to determine the accuracy of the initial resource estimation process.

16.1.4 Auxiliary Equipment The nominated ancillary fleet will consist of the following:

. Cat D9R Dozer x 5 . Cat 330E Excav/break x 1 . Cat 16M Grader x 2 . Cat 777D WT Watertruck x 2 . Lighting Plants x 10

16.1.5 Pit Dewatering Although the region is fairly arid, it is expected that some groundwater and seasonal rain will be encountered. The M5 pit lies within a surface drainage channel with a small catchment area. The channel will be dammed upstream to prevent ingress to the pit and the water used for processing and dust suppression. Skid or trailer-mounted centrifugal pumps will be required to remove water from the pit sump locations. An allowance has been included in the operating and capital costs for a pit dewatering system to pump water from pit sumps, although further groundwater studies are being undertaken to provide an estimate of groundwater inflows.

16.1.6 Waste Rock Storage Waste rock dumps have been designed with 15m batter heights with angles of 20°. The batters are separated by 15m wide berms. The overall slope angle of the waste dump faces is 15°. The waste dump designs are shown in Figures 15.5 and 15.6.

The M5 waste rock dumps will be developed adjacent to the access ramps on the eastern side of the pit. Total capacity of the design is 23Mm³. Given the waste volumes required for the TSF embankments and ROM pad construction, the waste dump provides excess capacity for the M5 pit design. The M5 waste dump has a maximum height of 55m and is 1,800m long by 400m wide.

The M1 waste dump will be placed directly to the northeast of the M1 South and M1 North pits and will have a capacity of 17Mm³.

Over 50% of waste material mined will be moderately or strongly weathered, and should be benign from an acid mine drainage perspective. Of the balance, 15% is from the transition (WOX) zone whilst the remaining 35% is fresh rock. Further work is underway to determine the acid generating potential of rock types. Construction of the dump will be scheduled to encapsulate any acid generating rock within non-acid generating or acid neutralising rock.

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Production Schedule As mentioned previously, the primary aim of the production schedule is to provide the highest value ore to the mill as early as possible in order to maximise the value to the project. This has been achieved by prioritising the highest grade pits and pit stages and stockpiling low grade oxide ore in the first three years of the schedule. Oxide ore from the M5 pit between the cutoff grade and 0.8g/t Au is stockpiled during the first 3 years of production to be treated in the final years of the mine plan.

The schedule is developed to satisfy physical and practical constraints, including: a sustainable production profile, achievable vertical advance rates and practical use of low grade stockpiling. The resultant mine production schedule is shown in Table 16.1.

Table 16.1 Mine Production Schedule

Pre- Total Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Prod

It should be noted that the oxide ore types can be milled at a faster rate than the more competent transitional and fresh ores. Over the life of the project, the weighted average processing rate is 2Mtpa. The ore grade and throughput rate were balanced in the schedule to achieve the greatest possible gold production for a given period.

In order to provide the highest value ore to the process plant in the earlier years of the schedule, the higher grade and higher strip ratio material from the M1 South pit and the southern part of M5 are mined. This results in a total material movement of 31Mt in the first production year and 24Mt in the second production year, reducing to 13Mtpa to14Mtpa for the remainder of the mine life. The nominated mining fleet of 1 x 250t class excavator and 3 x 120t class excavators and matching truck fleet (95t class) has sufficient capacity for these production rates. The fleet will be reduced in size from Year 3 onwards.

Rather than reducing the mining rate to match the milling rate from Year 4 onwards, the pits will be mined at a constant rate and excess ore will be stockpiled. This has been done to minimise the fixed mining costs over the life of the project. It has been estimated that compressing the mining schedule saves in excess of US$10M in mining costs.

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RECOVERY METHODS Additional test work is in progress and will be updated in the Definitive Feasibility Study technical report due for completion in mid-2018. Information from the February 20, 2017 Technical Report has been restated for completeness. The Sanbrado Gold Project process plant will have a design throughput of 2.0Mtpa. The plant will be located at an elevation of 290masl approximately 2km from the M1 South open pit and 1km from the M5 open pit.

The process flow diagrams (PFD) for the Sanbrado Gold Project Bankable Feasibility Study (BFS) have been developed from the process design criteria (PDC) prepared from metallurgical testwork. The proposed plant design is simple, but robust, and broadly comprises the following:

. Primary Crushing. . Crushed Ore Stockpile. . Grinding and Classification. . Gravity Recovery. . Leaching and Adsorption. . Elution. . Electrowinning. . Smelting. The design of the comminution circuit was undertaken by Orway Minerals Consultants (OMC) in Perth, Australia in consultation with Mintrex. OMC was requested by Mintrex to model a Single-Stage SAG milling circuit as well as a SAG and Ball Mill circuit as the basis for the design. The single-stage SAG milling circuit was selected due to its lower capital cost and its flexibility in allowing possible future expansion.

The BFS flowsheet and plant equipment selections are based on the results of the BFS metallurgical testwork program (see Section 13). The simplified BFS flowsheet is shown in Figure 17.1.

Process Design Philosophy The plant will be designed to achieve the required throughput, as stated in the PDC.

The crushing circuit will be designed with a throughput of 450tph and availability of 68.5%, on a 24 hours per day operation. Crushed product will report to an open stockpile, which has a total capacity of 15,000 tonnes and a nominal live capacity with free running material of 3,000 tonnes (note: with some ore types there will be zero live capacity).

A buried apron feeder installed in a reclaim tunnel will reclaim ore and directly feed the milling circuit via the mill feed conveyor. An emergency reclaim feeder is also installed in the reclaim tunnel to provide feed to the mill when reclaiming dead ore from the stockpile with a front end loader (FEL).

The milling circuit is designed for a nominal throughput of 250tph. It will operate at 91.3% availability, and achieve a design grind product size of 80% passing (P80) 90 micron.

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Figure 17.1

Sanbrado Process Plant Flow Diagram

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The CIL circuit will consist of seven adsorption tanks, treating the cyclone overflow. The CIL circuit is designed for a design leach feed grade of 2.8g/t Au.

The metal recovery and refining will consist of an elution circuit, electrowinning cells and smelting.

A tailings storage facility (TSF) for tailings deposition to a paddock type storage dam located 1.5km from the plant site.

Water, which will be required for a wide range of services, will be sourced primarily from the Gibgo River, located 6.5km southwest of the plant site. The river water will be pumped to a water storage facility (WSF) located approximately 800m from the plant, adjacent to the TSF. For further information relating to the TSF and WSF, refer to Section 18.

Comminution Circuit Design The preliminary design of the comminution circuit was undertaken by OMC to a BFS level.

The design throughput rate is 2.0Mtpa (250tph) of ore over nine years production with a P80 product transfer size of 90 micron from the comminution circuit.

17.2.1 Blending of Process Plant Feed The feed sources are abbreviated as Fresh (FRS), Weak Oxidation (WOX), Mild Oxidation (MOX) and Strong Oxidation (SOX). The proposed feed blend will be 50% FRS/WOX and 50% (SOX/MOX), which is the design basis. The mining schedule indicates that the design basis will be suitable and adequate for the first four years of production with an increase of fresh material from Years 5 onwards. From Years 5 to 9, the feed deviates from the 50/50 design blend with higher design specific energies and reduced throughputs evident during those years (refer Figure 17.2).

Figure 17.2 Mining Schedule Feed Blend

Mining extraction rates will be greater than the mill design throughput rate, meaning ore will be stockpiled by source, weathering and grade type, and blended to optimise mill throughput,

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feed grade, and metallurgical recovery. Mill feed blending will occur by feeding different ore types stockpiled at the run of mine (ROM) pad into the crusher circuit.

During periods of high crusher utilisation, ore will be crushed and stockpiled to maintain mill feed availability during crusher shutdowns.

17.2.2 Comminution Circuit Design Basis Operating Design Criteria Table 17.1 summarises the operating criteria adopted for the design of the comminution circuit.

Table 17.1 Comminution Circuit Operating Design Criteria

Criteria Units Value Annual Throughput Mtpa 2.0 Crushing Availability % 74.9 Crushing Circuit Operating Hours h/yr 6,557 Crushing Circuit Throughput dtph 305 Crushing Circuit Throughput (Design) dtph 450 Grinding Circuit Availability % 91.3 Grinding Circuit Operating Hours h/yr 8,000 Grinding Circuit Throughput dtph 250

Grinding Circuit Transfer Size P80 µm 90

Ore Specific Design Criteria The ore specific design criteria were developed based on the metallurgical testwork results and the design approach adopted by OMC. The key comminution design data and results are contained in Table 17.2.

Table 17.2 Ore Specific Comminution Parameters

85th Percentile 85th Percentile 50/50 Feed Criteria Units SOX/MOX FRS/WOX Blend UCS# MPa - 57.55 - Bond Crushing Work Index (CWi) kWh/t 4.1 10.8 7.4 Bond Rod Mill Work Index (RWi)* kWh/t 8.7 22.2 15.4 Bond Ball Mill Work Index (BWi) kWh/t 8.7 21 13.1 Abrasion Index (Ai) 0.098 0.351 0.224 Specific Gravity (SG) 2.69 2.72 2.70 A x b 150 32.9 47.2 Total Specific Energy Requirements kWh/t 10.1 25.6 17.8 Notes: #UCS data only available for fresh samples from M1 and M5. *no RWI data available for SOX/MOX, assumed to be similar to BWi.

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17.2.3 Comminution Circuit Selection The comminution circuit selected is a single stage SAG mill, fed with primary crushed ore. The SAG Mill has an installed motor size of 6.2 megawatts. The SAG mill specifications as determined by OMC are presented in Table 17.3.

Table 17.3 SAG Mill Specifications

Parameter Units SAG Mill Number of Mills 1 Mill Diameter (Inside Shell) m 7.30 Effective Grinding Length m 6.50 L:D Ratio 0.89 Discharge Configuration Grate Open Area % 8 Backing Rubber mm 6 New Liner Thickness mm 100 Grinding Media Size mm 125 %NC 75 VSD Mill Speed Max rpm 11.9 Max %Nc 78 Ball Charge - Maximum %Vol 15 Total Charge - Maximum %Vol 35 Mill Motor Size at Pinion kW 5,688 Installed Mill Motor Power kW 6,200

As more competent material is mined during the latter years of the operation, secondary grinding might be required in order to maintain the 90 micron product size.

Process Plant Description The plant layouts were developed in 3D and general arrangement drawings were produced for each area from the 3D model together with the overall plant site layout including positioning of the crushing plant, ore stockpile, feed conveyers, SAG mill, leach tanks, gold room, reagents storage and preparation areas, and infrastructure buildings.

Equipment selections have been completed for all major process plant mechanical equipment on the basis of the PDC.

17.3.1 Primary Crushing Haul trucks operating directly from the open pit will deliver ROM ore to the ROM pad and stored on the ROM pad in separate stockpiles of varying ore types and grades to facilitate blending of the feed into the crushing plant. The estimated maximum particle size of ore on the ROM pad will be 600mm in any dimension. Any oversized rock will be placed to one side and reduced to minus 600mm on the ROM pad.

The primary crushing plant provides single stage crushing to feed the SAG mill. The primary crushing plant includes a 150 tonne capacity ROM bin (10-BN-01), an apron feeder (10-FE- 01), a primary jaw crusher (10-CR-01), a stockpile feed conveyor (10-CV-01), a rock breaker (10-RB-01) and associated electrical equipment, steelwork and plate work.

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The primary crusher will be installed on a concrete foundation with the ROM bin, apron feeder and rock breaker, located adjacent to a 12m high concrete retaining wall against the ROM pad. Walkways and stairs will provide full operational and maintenance access throughout the primary crusher building.

The ROM bin will be fed blended ore from the ROM stockpiles using a FEL (CAT980 or equivalent). The ROM bin will be sized to accommodate direct tipping from 100 tonne sized haul trucks (CAT 777 or equivalent) to the crushing plant if required. The ROM bin will be lined with replaceable steel wear-resistant liners. Feeding of the ROM bin will be controlled by a ‘dump – no dump’ traffic signal mounted adjacent to the ROM bin. The traffic signal will be controlled by a radar level sensor mounted above the ROM bin.

The ROM bin discharge will be controlled by a 1,500mm wide apron feeder, which will direct feed into the primary jaw crusher. The primary jaw crusher will be a single toggle jaw crusher (C150 equivalent) with a feed opening of 1,200mm by 1,400mm that accepts nominal minus 600mm rocks. The rock breaker mounted adjacent to the jaw crusher will break any oversized rocks that lodge in the crusher. The crusher jaw plates will be adjusted to a 125mm closed side setting and will reduce the rock to a nominal minus 140mm, although there will be some rocks larger than this that pass through the crusher. The material flow path will be conservatively sized to accommodate rocks to up to 350mm in size.

The crushed product from the primary jaw crusher will discharge onto the 1,500mm wide stockpile feed conveyor. A weightometer (10-WE-01) installed on the stockpile feed conveyor will provide the tonnage of crushed ore passing through the circuit and onto the stockpile. Dust control will be achieved using the dust suppression system (10-DC-01) with high pressure water sprays, installed within the ROM bin.

17.3.2 Crushed Ore Stockpile Primary crushed ore discharges 10-CV-01 to a 15,000 tonne total capacity conical stockpile. The stockpile is sized to provide a maximum 12 hours live feed capacity to the grinding circuit at the design throughput of 250tph. The total capacity of the stockpile is equivalent to 60 hours of feed at the design throughput.

A 1,200mm wide by 7,000mm long reclaim apron feeder (20 FE 02) installed in a concrete reclaim vault will reclaim crushed ore from under the stockpile and feed it onto a 1,200mm wide by 120m long mill feed conveyor (20-CV-02). The mill feed conveyor will provide a nominal 250dtph instantaneous feed rate to the mill feed chute. A weightometer (20-WE- 02) installed on the pebble crusher feed conveyor (20-CV 03) will monitor and control the reclaim apron feeder variable speed drive, which in turn will control the feed rate to the nominated operator set point.

An emergency reclaim apron feeder (20-FE-03) is positioned towards the exit of the reclaim tunnel in the same concrete reclaim vault adjacent to 20-FE-02, but outside the edge of the stockpile. This will allow the FEL which feeds the crusher to also feed the emergency reclaim feeder by loading the feed chute (20-CH-08) with ore from the dead parts of the stockpile when there is no live ore in the stockpile or when 20-FE-02 is under maintenance. The emergency feeder will also be used for the addition of grinding media to the SAG mill.

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Lime is added directly onto 20-CV-03 from the lime handling system (20-LS-01) which is positioned after the emergency reclaim apron feeder (20-FE-03). The lime system consists of a 100 tonne capacity silo and will include a pneumatic bin activator, discharge isolation slide gate, rotary valve feeder, level instrumentation, dust collector, free standing structure and access platforms, and stairs. Bulk bags of powdered lime will be split and emptied into a transfer hopper, from which the lime will be pneumatically conveyed into the storage silo.

17.3.3 Grinding and Classification Circuit Primary crushed ore will be fed via 20-CV-02 to the SAG mill from the crushed ore stockpile. A 7.3m diameter by 6.5m long effective grinding length SAG mill (20-ML-01) is proposed for the primary grinding duty. The SAG mill will operate with a duty ball charge of 10% (125mm balls at a maximum 15% ball charge) with an expected pinion power draw of 5.7 megawatt. The SAG mill will have an installed motor power of 6.2 megawatt. A variable speed drive (VSD) will be installed on the mill to vary the mill speed to adjust for changes in ore characteristics.

Mill slurry will discharge from the mill to a mill discharge vibrating screen (20-SC-01), installed to separate pebbles from the discharge, control the particle size of the slurry reporting to the classification circuit and to adequately rinse the oversize ore, prior to it reporting to the pebble crushing circuit. The undersize slurry from the discharge screen will fall into the mill discharge hopper (20-HO 01).

Mill discharge pumps (20-PP-01/02) will be installed in a duty/standby arrangement, each having separate suction lines from the mill discharge hopper. Pneumatically actuated suction inlet knife gate valves, pneumatically actuated knife gate dump valves on the pump suction pipework and pneumatically actuated knife gate valves on the discharge pipework will facilitate pump operations and maintenance of the off duty pump when the system is operating the duty pump.

The slurry in the mill discharge hopper will be pumped to a 12 pack classifying cyclone cluster (20 CY 01). The cluster will have 10 operating cyclones and two standby cyclones.

The cyclones will classify the slurry feed into an overflow product with a P80 of 90 micron, which will be directed to the leaching circuit, and a coarse cyclone underflow product. From the cyclone underflow, 60% will be directed to the gravity circuit, while the remaining 30% will be directed to the cyclone underflow boil box (20-FD-01) and then back to the mill feed chute (20-CH-12).

New feed from the mill feed conveyor will be added to the recirculating loads in the mill feed chute (20-CH-12). The recirculating load has been designed to a nominal 190% of new mill feed. Proportional controllers will provide density control in the circuit by varying water addition to either the mill feed chute or discharge hopper in fixed proportion to the SAG mill feed rate.

The oversized material from the mill discharge screen (20-SC-01) will be directed either to a scats bunker or to the pebble crusher circuit. The pebble crusher feed conveyor (20 CV- 03) will deliver pebbles to a 90 kilowatt pebble crusher (20 CR-02) at a design rate of 47tph with a design feed size of 70mm. The crusher will operate at a close setting of 11mm to 14mm. The product from the pebble crusher will return via the pebble crusher discharge conveyer (20-CV-04) to the mill feed conveyor (20-CV02). The crusher station will have an overflow to allow excess pebbles to recirculate to the SAG mill if the crusher is overloaded

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A magnet (20-MA-01) on the pebble crusher feed conveyor and a metal detector (20 MD 01) further along the same conveyor will protect the pebble crusher from tramp metal. Any tramp metal detected will be diverted to bypass the crusher by a shuttle (20-SH-01) located at the end of the pebble crusher feed conveyor.

A fixed tower crane (30-CN-01) will be installed in the CIL area. The tower crane can also be used for cyclone pack maintenance activities. Major maintenance activities around the SAG mill and discharge pumps will be undertaken with a mobile hydraulic crane. Platforms and stairs will provide full operational and maintenance access.

17.3.4 Gravity Recovery Circuit A proportion of the cyclone underflow will be directed onto a distribution box (20-FD-02) which will distribute the feed between two scalping screens (20-SC-02/03) located below the classifying cyclones. The oversized product from the two scalping screens will be sent to the cyclone underflow boil box (20-FD-01), while the undersized product will be directed to two centrifugal gravity concentrators (20-KC-01/02) which will act as gravity roughers. It is expected that the concentrators will remove 67kg/h of concentrate into a rougher concentrate storage tank, which is part of the intensive leach reactor (50-LR-01), to be located in the gold room.

The intensive leach reactor is a batch process, intensively leaching the gravity concentrate to dissolve the contained gold into solution. The pregnant liquor from the reactor will be pumped to the electrowinning module tank (50-TK-17) for electrowinning in a dedicated electrowinning cell (50-EW-03). Solution from the electrowinning cell will be recirculated back to the electrowinning module tank for further leaching. From the leach reactor, final barren solution will be discharged to the mill discharge hopper (20-HO 01). The gold sludge collected from the electrowinning cell will be refined to produce the final gold doré bar.

The gold collected from the gravity circuit electrowinning cell will be smelted separately using the calcine oven (50-OV-01) and smelting furnace (50-FU-01). This will allow for separate metallurgical accounting of the gravity circuit. The final gold doré bars will be stored in the gold room safe (50-SF-01) ready for delivery to the market.

17.3.5 Carbon in Leach Circuit The classified slurry from the cyclone overflow will be directed to a pair of trash screens (30-SC-04/05) from the trash screen distribution box (30-FD-03). There will be capability to operate either one or both screens. Oversize material from the trash screens will typically be returned to the SAG mill feed chute. The trash screen overflow chutes will be fitted with twin outlets to allow periodic dumping of trash to a trash bin positioned at ground level.

Trash screen underflow will be collected into a distribution box that will allow the slurry to be directed to either the first or the second CIL tank (30-TK-01/02). Slurry will only be directed to the second tank in the case that the first tank is offline for maintenance.

The CIL train will comprise seven tanks (30-TK-01 to 07) with nominal dimensions of 12m diameter and 14m height, each with a nominal capacity of 1,500m³, providing a slurry residence time of 27 hours with a slurry density of 45% solids by weight.

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Each CIL tank will be fitted with pumped inter-stage tank screens (30-SC-06 to 12). Carbon will be held in all tanks except the first tank where the inter-stage screen (30-SC-06) will act as a safety screen to prevent oversize material entering downstream tanks in the event of cyclone roping and a trash screen overflow or failure.

All tanks will be equipped with hollow shaft agitators to facilitate oxygen injection through the shafts. Only the first three tanks will be sparged with oxygen. The first three tanks will also be fitted with a nozzle to facilitate oxygen injection through the side of the tanks.

All CIL tanks will be equipped with recessed impeller type carbon transfer pumps (30-PP-04 to 10), which will advance the carbon between tanks and remove carbon from the circuit.

The recessed impeller pump in CIL tank 1 will be used to pump slurry back to the trash screens to prevent trash material entering proceeding tanks in the event of cyclone roping. The recessed impeller pump in tank 2 will be used to pump slurry to the carbon recovery screen (30-SC-13) for the loaded carbon to be removed from the circuit. The recessed impeller pump (30-PP-06) in tank 3 will normally provide for carbon transfer between tank 3 and tank 2, however in the event that tank 2 is offline for maintenance, it will be used to recover loaded carbon to the carbon recovery screen. The loaded carbon will undergo an acid wash before proceeding to the elution circuit.

A vibrating carbon safety screen (30-SC-14) will be located adjacent to tank 6. This screen will collect any carbon that escapes from tank 7 (or tank 6 in the event that tank 7 is off-line) in a disposal drum for reintroduction to the circuit manually. The undersize product from the carbon safety screen will be gravity fed to a tailings thickener (30-TH-01). The thickener will separate the slurry into densified tailings and water. The tailings will be pumped to the TSF and the water will be gravity fed to the process water tank.

The tower crane will facilitate removal of the inter-stage screens for maintenance and cleaning. The tanks will be constructed on concrete ring beams within a concrete bunded containment structure. The CIL bund will be fitted with a sump pump (30-PP-10) which will collect any spillage within the bund and direct it back to the trash screen distributor box (30- FD-03). The bunded structure around the tanks will not be designed to comply with dangerous goods regulations, because none of the process fluids are classified as dangerous goods and it is not normal practice to contain them beyond normal operational spillage. Minor spillages will be contained by the bund system. In the highly unlikely event of a large spill, containment will be via the drainage system of the plant for collection in the contaminated water pond adjacent to the plant. Solids will be recovered by mechanical means (e.g. FEL).

17.3.6 Elution, Electrowinning and Smelting Acid wash and rinse cycles will be performed in the 10m³ capacity rubber lined acid wash hopper (50-TK-08) located beneath the loaded carbon recovery screen. Following the rinse cycle, the carbon in the acid wash hopper will be discharged into the elution column (50- PV-01) through an actuated ball valve. The elution column will have a volumetric capacity of 10m³ and be capable of holding four tonnes of carbon.

The strip solution will be injected with sodium hydroxide and sodium cyanide and then be preheated by the in-line elution heater (50-HE-01) to reach a solution temperature of 130°C. The hot strip solution will then be introduced to the bottom of the elution column.

After approximately one bed-volume of caustic cyanide solution has been passed through the elution column to pre-soak the carbon, a further five bed-volumes of hot rinse water will be

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passed through the column. A further one bed-volume of cold rinse water will be passed through the column after the hot rinse water to cool down the carbon. The first 3.5 bed- volumes of pre-soak and hot rinse water will be returned via the duty/standby eluate filters (50-FL-01/02) to either one of the two pregnant solution tanks (50-TK-11/12) via a recovery heat exchanger to recover heat to the strip solution from the eluate. The last 3.5 bed- volumes of hot and cold rinse water will be directed to the intermediate solution tank (50- TK-10) which will supply the feed water to the first half of the next strip.

Elution of the gold from the carbon is expected to take about six hours. Pregnant solution will be collected into one of two pregnant solution tanks. The pregnant solution tanks will have a common pregnant solution pump (50-PP-15) which will feed to the electrowinning cells (50 EW-01 or 50-EW-02). The barren solution from the electrowinning cycle will be returned to CIL tank 1 (30-TK-01) using a barren solution pump (50-PP-16).

At the completion of the elution cycle, barren carbon will be pumped from the elution column (50-PV-01) to the regeneration kiln carbon feed hopper (50-HO-03). The hopper is located on top of the regeneration kiln (50-KN-01) above CIL tank 7. From this hopper the carbon will be either regenerated in the kiln or discharged directly into CIL tank 6 under gravity depending on the carbon activity level. Prior to regeneration, the barren carbon will be de- watered over a carbon dewatering screen positioned above the storage hopper. The rotary kiln feed chute will drain any residual and interstitial water from the carbon prior to it entering the kiln. Kiln off-gases will be used to dry the carbon before it enters the kiln. At the end of the regeneration process, the regenerated carbon will discharge back into CIL tank 7.

The gold sludge from the gravity and the elution circuit electrowinning cell cathodes will be washed in the cathode wash box (50-WB-01), a manual process. The resultant sludge will then be transferred via the cell sludge trolley (50-TR-01), to the calcine oven (50-OV- 01) to remove the steel wool cathodes through oxidation. The product from the calcine oven will be direct smelted using fluxes in an LPG fired smelting furnace (50-FU-01) to produce the final gold product doré bars, which after weighing using a Sartorius Balance (50-WE-04), will be stored in the gold safe (50-SF-01). The gold sludge from the gravity circuit will be refined separately from that of the elution circuit to allow separate accurate metallurgical accounting of the gravity and CIL circuits.

17.3.7 Tailings Disposal Underflow from the carbon safety screen will gravitate to a 30m diameter high-rate tailings thickener (30-TH-01) to produce a tailings underflow density of 55% solids. Thickener underflow will be gravity fed to the tailings discharge hopper (30-TK-26) from which it will be pumped to the TSF by two tailings discharge pumps (30-PP-11/12), installed in a duty/standby arrangement. Each pump will be equipped with manual dump valves on the pump inlets and pneumatically controlled inlet and discharge knife gate valves.

The tailings pipeline to the TSF will be installed above ground except for locations where road crossings necessitate burial of some sections. The tailings pipeline has been sized as a 315OD PN12.5 HDPE pipe. Leaks in the tailings line will be detected by comparison between two flow meters, one located at the plant and the other located at the TSF. The tailings and decant return pipelines will be laid in a fully bunded and lined trench between the process plant and the TSF.

Reagents

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17.4.1 Lime Quicklime will be delivered to site in shipping containers containing 20 tonnes of quicklime in one tonne bulk bags which will be stored in the reagent store and in shipping containers.

The lime handling system (20-HL-01) will consist of the following items:

. A bag unloading system including;

 Bag splitter and transfer hopper; and,

 A pneumatic conveying system to transfer lime from the hopper to the storage silo. . A 100 tonne lime silo. . Silo pneumatic activation system to fluidise the lime. . Rotary valve to deliver lime onto the mill feed conveyor belt, with a VVF drive controlled by a proportional controller with a set point related to the SAG mill feed conveyor rate. . A volume equivalent to 5.2 days use is stored per batch. The various mixing/storage/dosing facilities provide a short term buffer for operating such that there is at least one days storage available under most conditions. This allows reagent management to be undertaken on day shift only.

17.4.2 Cyanide Cyanide will be delivered to site in 20 tonne shipping containers in one tonne bulk bags which will be stored in the reagent store.

Cyanide will be mixed with raw water to create a 30.5%w/w solution in the cyanide mixing system, which will comprise the following items:

. A hoist (60-HT-08), which will lift the bags directly onto the bag splitter. . A bag splitter (60-BS-01). . A 10m³ mixing tank (60-TK-15). . A mixing agitator (60-AG-09), which will mix the cyanide and the water to create a homogenous solution. The mixed solution will be transferred by a cyanide transfer pump (60-PP-23) to a separate 22m³ cyanide storage tank (60-TK-16), where duty/standby cyanide recirculating pumps (60 PP 19/20) will circulate the cyanide solution through the plant ring main with a constant pressure bypass return to the tank. In addition, a cyanide dosing pump (60-PP-21) will deliver cyanide from the ring main to the elution circuit in a controlled manner. The cyanide mixing and storage tank will be contained within a concrete bund with a collection sump to recover spillage. The sump pump, (60 PP 22), will recover any minor spillage and deliver it to the trash screen distributor box (30-FD-03).

Sodium cyanide and sodium hydroxide will be mixed to provide a solution at a strength suitable for direct dosing to the process facilities.

The sodium cyanide mixing facility will include a dedicated mixing and storage tank. Once a mix is undertaken the solution is pumped to the storage tank which holds a buffer volume of nominally 3.6 days. This arrangement also allows additional mixes to be undertaken during operations without upsetting the dosed reagent concentration.

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The various mixing/storage/dosing facilities provide a short term buffer for operating such that there is at least one day storage available under most conditions. This allows reagent management to be undertaken on day shift only.

17.4.3 Caustic Soda/Sodium Hydroxide Caustic soda will be delivered in 25kg bags to site and stored in the reagent store in one tonne pallet lots.

The caustic soda will be mixed with raw water in a skid mounted caustic mixing system to create a solution with 50%w/w concentration. The mixing system will consist of the following items:

. A bag splitter (60-BS-02). . A 1m³ mixing tank (60-TK-14). . An agitator (60-AG-08), which will mix the caustic soda and the water to create a homogenous solution. The mixing system will be located in the same containment bund as the cyanide mixing and storage tanks. A caustic dosing pump (60-PP-23) will draw the solution from the mixing tank and deliver it to the elution circuit.

The sodium hydroxide facility is arranged such that the mixing and storage is contained in one tank. As the sodium hydroxide use is intermittent, mixing will be undertaken in those periods where there is no demand, thus allowing the mixed reagent to homogenise prior to use. A volume equivalent to 1.7 days use is mixed and stored per batch.

The various mixing/storage/dosing facilities provide a short term buffer for operating such that there is at least one days storage available under most conditions. This allows reagent management to be undertaken on day shift only.

17.4.4 Hydrochloric Acid Concentrated liquid hydrochloric acid (32%w/w) will be supplied in 1,185kg intermediate bulk containers (IBC) and delivered to site in shipping containers of 23.7 tonnes capacity. The acid will be transferred from the IBC’s to the acid wash hopper by an acid dosing pump (60-PP-25) for a carbon acid wash cycle, comprising injection into a water stream pumped from the water tank (60-TK-13) to create a diluted 3%w/w hydrochloric acid solution.

The concrete containment bund which surrounds the acid preparation area will comply with the dangerous goods statutory requirements, and be protected with a coating to prevent acid damage to the concrete.

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Hydrochloric acid is received in 1m³ IBC tanks at the appropriate dose strength. The IBC vessels are connected to a common suction manifold so that the dosing pump always has a reserve of acid to draw from. Each IBC provides around two days of reagent and so management of the IBC levels is required to ensure there is always one vessel with adequate volume for intermittent dosing to the acid wash facility. A volume equivalent to three days use is available.

17.4.5 Activated Carbon Activated carbon in 500kg bulk bags will be transported to the site by road in 22 tonne sea containers. The carbon will be stored in the containers or under tarpaulins to protect it from the weather. When required, carbon will be hoisted up to the top of CIL tank 7 and broken directly into the tank.

17.4.6 Oxygen Oxygen gas will be manufactured on site using a dedicated pressure swing adsorption plant (70-PS-01) and distributed to the leach tank agitators in the CIL area.

17.4.7 Flocculant Flocculant will be delivered to site in one tonne bulk bags. A forklift/telehandler will lift bulk bags onto the bag splitter for filling of the flocculant loading hopper. Flocculant will be mixed to a 0.05%w/v concentration and transferred to a storage tank. From the storage tank, flocculant dosing pumps will dose liquid flocculant to the tailings thickener via a static mixer to further dilute the flocculant concentration to the required dosage strength of 0.025%w/v.

Control Systems The Plant Control Systems (PCS) will be a network of Process Logic Controllers (PLC) sitting beneath a Supervisory Control and Data Acquisition (SCADA) network layer. The PLC’s will perform the necessary controls and interlocking, whilst the SCADA terminals will monitor the PLC’s and provide an interface for operator interaction.

The PLC’s and SCADA terminals will communicate via a plant wide Ethernet network, the backbone of which will be dedicated single mode fibre optic cables. For short distances, Cat 6 Ethernet cable will be installed.

GE Fanuc RX3i PLC’s and Citect SCADA are proposed. This combination has worked very well on past projects and has proven to be very reliable. Deviations from this to an alternative PLC and SCADA system would need to be investigated during the final design stage for reliability and cost.

Field instrumentation and drive status signals will interface to the plant control system by means of hard-wired signals. Vendor packages may be connected to the SCADA network via a communications link, where appropriate.

The PCS equipment installed within each area will function autonomously, such that a failure of the PCS in one plant area will not affect the other areas.

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The control philosophy of the plant will provide an appropriate level of automatic start up and shut down of various plant areas which will aid the plant operator in performing his tasks. Automatic interlocking, sequence control and analogue control will be implemented by the PCS equipment where required. Safety interlocks will be hard-wired.

PID loop controllers will be programmed into the PCS and be accessible via the SCADA terminals in the control rooms.

The PCS will provide detailed information including:

. Plant status monitoring. . Fault annunciation and logging. . Drive and systems diagnostics. . Trending for all analogue process parameters. The PCS will be powered by uninterrupted power supply (UPS) equipment, providing bumpless, fully synchronised power for thirty minutes after total power failure.

PLC’s will be installed in the main plant motor control centre (MCC).

Vendor panels may contain PLC’s depending on the complexity of control provided. Where possible, vendors will be asked to comply with the site standard PLC’s, to minimise spare holdings.

SCADA terminals will be installed in the following locations:

. CIL Control Room x 2 (above CIL deck). . Crusher Control Room. . Desorption Control Panel. . Electrical Supervisor’s Office. The SCADA system will be configured so that only Wet Plant drives can be controlled from the Main Control Room, only Crusher Drives from the Crusher Control Room and only the Desorption Sequence from the Desorption Control Panel. In situations where SCADA terminals have failed, it will be possible to bypass this by the user access level.

Password protected user accounts will be set-up in the SCADA to limit access to certain control functions. All functions required for day-to-day running of the plant will be made available at the operator level. Changing of set-points and PID parameters will be allowed at the Supervisor level (e.g. Plant Manager/Metallurgist/Plant Shift Supervisor). Complete control and development access will be allowed at the Administrator level (e.g. Electrical Supervisor).

Two SCADA terminals will exist in the Main Control Room and provide the redundancy so that should one terminal fail then the Wet Plant can still be operated from the other terminal.

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The Desorption terminal will be installed in a stand-alone, metal cabinet with a perspex window for viewing of the monitor. The panel will also include a pull-out draw for the keyboard so that it can be drawn out when required. The panel will be located within the desorption area, most likely beside the Electrowinning Rectifiers. Operators will be able to monitor and control the desorption sequence locally, from this control panel, avoiding constant trips to the Main Control Room.

The SCADA terminal in the Electrical Supervisor’s office will contain the necessary licencing for future on-site development of the SCADA application. Application updates of all other SCADA terminals will be possible from the supervisor’s terminal.

Electrical Reticulation Power will be accepted at the terminals of the plant high voltage feeder housed in the Plant Main Substation. This will house the Plant Main 11 kilovolt distribution board. Power distribution within the plant area and vicinity will be at 11 kilovolts and 415 volts. Power consumption for each general plant area will be metered as indicated on the plant single line diagram. Power metering will generally take place at the 11 kilovolt switchboard and at MCC incomers.

11 kilovolt power distribution cables will generally be underground within the plant area, while all other plant cabling will be in above-ground cable ladder attached to buildings and structural steelwork. Overhead power lines will not be installed in the immediate plant area to avoid interference with the movement of mobile equipment.

Substation buildings will be of the demountable/transportable type and be fully air- conditioned to maintain the internal air temperature at a maximum of 25°C. Equipment in substations will be designed for continuous operation at rated output in a substation ambient temperature of 40°C maximum, 5°C minimum. Substation buildings will house the MCC, DB’s and VSD’s for that area and have sufficient space to allow the extension of switchboards as appropriate. Each substation building will incorporate a personnel access door and a two- leaf equipment door. The doors will be fitted with a panic release device. All substation building doors will open outwards.

In addition to the electrical and instrumentation equipment, each substation will be equipped with an internal light and small power system, emergency lighting, safety notices, fire detection system and fire extinguishers. Fire detection systems will be limited to smoke detectors and a Vesda system wired to a fire panel within each building. Local annunciators will be installed on the outside of the building. Fire suppression systems have not been allowed for nor has the painting of cables with fire retardant paint.

The substation buildings will be mounted on supports 1.5m high to facilitate cable entry into the MCC’s from the base. Transformers associated with plant substations will be located in outdoor compounds adjacent to substation buildings.

Substation buildings have been allowed for in the following areas:

. Crushing Area LV Substation. . Wet Plant Area LV Substation (this may be one large building or split into two smaller buildings depending on final plant layout). . SAG Mill HV Substation.

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All transformers on the plant site will be pad mounted with compound fencing and underground earthing. They will include cables boxes on the HV and LV terminations. The following transformers have been allowed within the process plant:

. Crushing Area Transformer (pad mount). . Wet Plant Area Transformer x 2 (pad mount). . SAG Mill Motor Transformer (pad mount). . Plant Buildings Transformer (kiosk). . Mining Contractor Transformer (kiosk). High voltage switchgear will be supplied for the SAG mill so that isolations of the drive can be performed under the control of the site maintenance personnel without relying on the power station operator or requiring access to the power station switchboard.

The switchgear will be indoor, metal clad switchgear with a vacuum or SF6 circuit breaker on a withdrawable truck, enclosed to IP41. A Multilin 469 electronic protection relay has been allowed for protection of the SAG mill motor.

Motor current indication will be provided where specified, either as a panel mounted ammeter on the motor starter door, or as a current input to the PCS. Motors requiring control system current indication will require a current transducer with a 4-20mA dc output incorporated into the motor starter.

The following MCCs will be supplied within the plant site:

. Crushing Area MCC (indoor, c/w PLC). . Wet Plant MCC x 2 (indoor, c/w PLC). Electronic VSD panels will be either floor mounted or wall mounted panels, depending on size. Motors driven by VSD’s will be provided with thermistor protection.

All VSD’s will be capable of having their speed regulated by the PCS. However, when the associated drive control is selected to “local” mode, it will be possible for local speed setting to take place at the VSD. VSD’s have been allowed as indicated in the Maximum Demand Calculation.

Services

17.7.1 Compressed Air Plant air and instrument air will be supplied from two air compressors (70-CU-01/02) operating in duty/standby mode. The plant air compressors are fitted with inlet air filters and will discharge to the plant air receiver (70-AR-02). Instrument air will be produced from a bleed off the discharge header of the plant air receiver through a dedicated instrument air dryer (70-DR-01) which will feed a dedicated instrument air receiver (70-AR- 03) supplying upstream users. A low pressure alarm on the instrument air receiver will close the plant air header supply valve ensuring the instrument air system is not starved of air supply. Another bleed off the plant air receiver header will feed the primary crushing air receiver (10-AR-01) and supply plant air to the primary crushing area. Plant air will be piped from the plant air receiver throughout the process plant to provide compressed air to the various downstream users, which include general service points. Instrument air will be piped throughout the process plant to feed automated valves and instruments.

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17.7.2 Process Water Process water will be stored in and delivered to the plant from the process water tank (70- TK-22). The sources of the process water inputs will be from the TSF decant return water and from the tailings thickener overflow.

The raw water tank (70-TK-21) will provide an overflow to the process water tank but not vice versa. Decant water will be returned from the TSF by the duty/standby decant return water pumps (40-PP-53/54). Process water will be delivered from the process tank to various end users in the process plant by the duty/standby process water pumps (70-PP- 31/32). The capacity of the process water tank will be 400m³ which represents approximately 1hr storage at normal flows.

17.7.3 Raw Water Raw water will be sourced from a small impoundment on the Gibgo River some 6.5km from the plant site. Duty/standby submersible catchment pumps will seasonally harvest water from the lowest point in the impoundment and transfer it to the lined raw water pond located adjacent to the process plant. Duty/standby raw water transfer pumps will be provided for this duty. The raw water storage pond capacity will be 2.1 megalitres. This capacity has been sized for the annual water consumption requirements, based upon a 1:10 year dry season to accommodate the highly seasonal and unpredictable river flows at the project location.

Raw water will be transferred from the raw water pond to the plant raw water tank (70-TK-21) as required. The capacity of the raw water tank will be 620m³ which represents approximately four hours storage at normal flows and will also provide two hours fire water supply capacity. Raw water will be transferred from the raw water tank to various end users in the process plant by duty/standby raw water pumps (70-PP-28/29).

17.7.4 Potable Water Potable water will be produced from local water bores. Due to the quality of this water it will require filtration and chlorination treatment to ensure it complies with Word Health Organisation potable water standards.

The only potable water to be used in the treatment plant processes will be in the stripping plant and that water will be first stored in the strip water tank. To prevent contamination of the potable water supply, there will be no potable service points or direct connection of potable water to process equipment.

Potable water will be required for all safety shower and eye-wash stations around the process plant and in all ablutions, offices and workshops, and drink fountains.

17.7.5 Sewage Sewage from the process plant will be delivered by the sewage pumping system (70-PP- 60/61) to the camp sewage treatment plant (90-RO-02), where it will treat sewage from both the plant and the camp sites. The treated water will then be used to irrigate the gardens and lawns around the camp.

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PROJECT INFRASTRUCTURE Existing Infrastructure and Services There is limited existing infrastructure and no existing services that are suitable to support the Sanbrado Gold Project. The majority of existing infrastructure is to support the local subsistence and small-scale agricultural practices. Further work is in progress and will be updated in the Definitive Feasibility Study technical report due for completion in mid-2018.

Site Access

18.2.1 Upgrade of Existing Gravel Road from Zempasgo The existing gravel road from Zempasgo requires remedial and upgrade works for permanent use as the main access road to the Project. The majority of the upgrade works have been completed by WAF with only approximately 2km of works remaining. The upgrade work is primarily importing additional fill material to allow re-profiling of the road surface. In addition, there are several floodway crossings (refer Figure 18.1) that require upgrading.

Figure 18.1 Existing Gravel Road from Zempasgo

18.2.2 Site Access Road The gravel site access road will be constructed from a T-junction off the existing gravel road from Zempasgo, to the southeast of the WSF. The site access road will pass the accommodation camp to service the process plant, a distance of approximately 1.5km.

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A separate 1km access road will be constructed to service the mining contractor’s area. This road will intersect the site access road approximately 250m south of the process plant entrance.

18.2.3 Mine Access and Haulage Roads The mining contractor shall be responsible for the construction of the mining access and haulage roads. A secondary access road will be constructed between the mining contractor’s area and the ROM pad.

A main haulage road will head northeast from the ROM pad to the M5 pit. Another road will connect the M1 North and M1 South pits to the M5 haul road. The M3 pit access road will intersect the M1 access road immediately south of the M1 North pit. Approximately 6.5km of mining haul roads will be constructed.

Site Development

18.3.1 Process Plant Ore will be transported from the open pit mining operations and placed in stockpiles on the ROM pad located approximately 400m from the southern boundary of the M5 pit. The ROM pad is located adjacent to the process plant so that ore can be fed directly from the ROM stockpiles to the primary crusher via a FEL. Figure 18.2 presents the Sanbrado Gold Project Layout.

Figure 18.2 Sanbrado Gold Project Layout

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The proposed process plant location (refer Figure 18.3) is approximately 500m to the south of the M5 pit. The process plant and specific infrastructure will be located within a high security area with controlled access points.

Figure 18.3 Proposed Process Plant Site

18.3.2 Mining Contractor’s Area The mining contractor area will be located to the southwest of the ROM pad within a general security perimeter fence. Power, raw and potable water services will be provided to the mining contractor area at a designated battery limit at the boundary of the area. The provision of reticulated services within this area will be the responsibility of the mining contractor. The mining contractor will construct their own facilities within this area which will include, but not be limited to, workshops, warehouse, change room, ablution blocks and offices.

18.3.3 Tailings Storage Facility Knight Piésold Consulting (Knight Piésold) have undertaken the BFS investigation, preliminary design and cost studies for the TSF and related infrastructure, the WSF, water abstraction from the Gibgo River and the site surface water management. The ‘Tailings and Water Management Definitive Feasibility Study Report’ prepared by Knight Piésold.

The TSF will be located 1.5km northeast of the process plant, situated outside the 500m blast zone around the M5 pit.

The proposed TSF is designed with a storage capacity of 25Mt and consists of two cells of equal 12.5Mt capacity. The current mine plan requires 16.8Mt tailings storage capacity. It is proposed that one TSF cell with approximately 1 year of tailings capacity is constructed prior to commencement of ore processing. The construction of the second cell will be deferred until the dry season of the first year’s operation. The TSF cells will then be raised and operated concurrently.

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The TSF will comprise a four-sided paddock storage facility formed by a multi-zoned earthfill embankment and a HDPE geomembrane basin liner. Tailings within the two TSF cells will be raised at similar rates. Raises will be completed using full downstream embankment construction (i.e., no embankments constructed on the tailings surface).

ROM waste rock arising from the M5 open pit will be used as the primary embankment fill during construction of the TSF subsequent to Stage 1. Based on the mine schedule, the volumes of pre-strip and mine waste material are sufficient for the construction of the Stage 1 TSF Cell 1 and the WSF, the Year +1 construction of TSF Cell 2 and all subsequent embankment raises.

The TSF basin area will be cleared, grubbed and, where present, the topsoil stripped. There are numerous outcrops of granite throughout the area. It is proposed that, where practicable, some of the granite boulders are reshaped or removed and the remaining outcrops of intact granite covered with locally borrowed soil to provide a suitable subgrade for the HDPE liner.

In accordance with Burkina Faso requirements, both cells will be provided with HDPE liner. The ground is moderately permeable and the HDPE liner will act as the primary seepage barrier to prevent significant seepage losses and reduce the site water demand. The Stage 1 cells will be individually HDPE lined as separate basins. At Stage 2 the liner will be joined to form a single basin. This approach allows the volume of the central dividing embankment and the area of HDPE liner to be reduced.

The TSF design incorporates an underdrainage system to reduce the pressure head acting on the HDPE liner, reduce seepage, increase water recovery, increase tailings densities, and improve the geotechnical stability of the embankments. The underdrainage system will comprise a network of finger and collector drains. The underdrainage system drains by gravity to a collection sump located at the lowest point in the TSF basin from which solution collected in the sump will be pumped back into the TSF.

The deposition of tailings into the TSF will be sub-aerial from a HDPE pipeline located around the perimeter of each cell. Deposition will occur from multiple spigots inserted along the tailings distribution line which will be located on the upstream crest of each of the cells. The deposition locations will be moved progressively along the distribution line as required to maintain even deposition of tailings from the TSF perimeter.

Supernatant water will be removed from the TSF by submersible pumps located in concrete decant towers surrounded by rockfill in the centre of the facility. Access to the decant towers will be from an earthfill causeway. Solution recovered from the decant system will be directly pumped back to the plant process water tank for re-use in the process circuit.

An emergency spillway between the two cells will be installed at all times during TSF operation and an overflow spillway will be connected to the adjacent surface water management system. The emergency spillway will be constructed at the southern end of Cell 1 to protect the main embankment from overtopping in the event of an extreme storm event exceeding the storage capacity. The preliminary TSF design provides a 500mm freeboard above the tailings and below the emergency spillway sufficient to allow the supernatant capacity and the storm surge capacity from a 1:100 year, 72 hour duration storm event.

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18.3.4 Water Storage Facility The WSF will be located immediately to the southwest of the TSF, in the area of a small creek line that intermittently flows to the southeast towards the Gibgo River.

The WSF is designed to store up to 15Mm³ of water at the maximum operating level. The WSF will be primarily supplied with water from the Gibgo River over a 4 month period during the wet season. No natural catchments feed into the WSF. The design details are similar to those of the TSF with the exception of the underdrainage system which is not required. The HDPE liner will limit seepage losses.

The design of the water dam comprises a zoned earth embankment with a 1.5mm thick smooth HDPE liner installed over the compacted soil subgrade over the full WSF basin and upstream embankment area.

Water will be discharged to and abstracted from the WSF via two abstraction towers located adjacent to one another at the southeast corner of the facility. The pumps and pipelines associated with the water discharge and abstraction will be permanent installations.

18.3.5 Gibgo River Water Abstraction Process water will be sourced from the Gibgo River, which flows in a north to south direction approximately 6km east of the site (Figure 18.4). The river is ephemeral, with flows that are suitable for harvesting expected to occur over 4 months of the year during the wet season.

Figure 18.4 Proposed Gibgo River Water Abstraction Area

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Preliminary monitoring of the Gibgo River indicated that water flows available for abstraction occur as intermittent peaks. Four stream monitoring stations have been installed, upstream and downstream of the abstraction point, which confirmed that the river flows are highly responsive to local rainstorm events and that there will be extended periods when there is little or no flow through the wet season. Continuous additional detailed monitoring is proposed to verify modelled flow against actual runoff conditions.

A low embankment will be constructed to attenuate a portion of the flow and facilitate a continuous steady pumping abstraction rate. The on-river storage has been designed to allow a longer harvesting period, reduce the peak pumping duty and mitigate the risk associated with harvesting from an intermittent stream.

Current estimates indicate that a water harvesting dam (WHD) constructed on the river with a storage capacity of approximately 200,000m³ will provide sufficient attenuation of flows for the required 1.5Mm³/yr of water to be harvested from the system, pumping continuously at 140L/s.

The dam will comprise a low permeability Zone A embankment with Zone E and geotextile erosion protection to both the upstream and downstream faces of the dam and a broad shallow spillway. Water will be abstracted from two concrete towers located towards the western end of the dam which are surrounded by coarse rock fill.

The catchment of the Gibgo River at the proposed abstraction location is estimated to be 440km². There were government plans to construct the Koankin Dam upstream of the abstraction point which would reduce the catchment available for harvesting in the short term to approximately 100km² (it is noted that Koankin Dam is expected to overflow annually after first fill).

Under average climatic conditions and assuming the first fill of the Koankin Dam is taking place, the remaining catchment is expected to just meet the project water demand. To account for drier than average conditions and the reduced water reporting to the abstraction point during the Koankin Dam first fill, additional water sources would need to be identified to reduce project risk once a decision to construct Koankin Dam was made.

Administration and Plant Buildings The administrative, maintenance and support functions will be housed in blockwork constructed buildings and steel-frame construction sheds. The following sections describe each of the administration and plant buildings.

18.4.1 Main Entrance Gatehouse and First Aid The security gatehouse and first aid building will be located at the mine entrance. The security office will house a security reception area and the security manager’s office. The first aid area will house the nurse and the doctor within the low security area. One end of the building will act as a garage for the ambulance. A parking lot will be located at this building for site visitors.

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18.4.2 Administration Building The main administration office will be located within the low security area, with an approximate floor area of 440m². The administration building will be a block construction built on concrete slab and comprise a meeting room, male and female ablutions and kitchen, as well as offices for management, mine and process plant technical services and administrative personnel. The administration office will be fitted throughout with split- system air-conditioners and reticulated power from a UPS to service computers and additional electrical equipment. A small parking lot for visitors and administration and process plant light vehicles will be located at the front of the administration building.

18.4.3 Laboratory and Core Shed The laboratory building will be located within the low security area and will have an internal area of approximately 250m³. The laboratory building will accommodate the following:

. Laboratory supply stores. . Environmental laboratory. . Metallurgical laboratory and sample preparation area. . Laboratory offices. The building will be air-conditioned and the offices will have windows. A large area beyond the metallurgical and environmental laboratories will house four sea containers for sample and supply storage.

The core shed will be located adjacent to the laboratory building within the low security area. The core shed will receive exploration drill cores and grade control samples and provide storage for the samples. The core shed will utilise steel-framed construction with the exterior clad in metal sheeting. The internal area of the shed will be approximately 370m².

18.4.4 Crib Rooms Two crib rooms will be incorporated in the plant and administration building areas to provide shift meals to site personnel. The 32 seat mess (approximately 105m² internal floor area) will be located inside the high security area adjacent to the security building. The 48 seat mess (approximately 195m² internal floor area) will be located in the low security area adjacent to the main administration building. Both buildings will have verandas attached to them. Ventilation will be provided by large fly-screened windows and ceiling fans to circulate the air.

Separate washing facilities will be provided outside the mess buildings. All meals are expected to be prepared at the village outside the high security area and transported into the high security mess at meal times.

18.4.5 Security, Laundry and Change Room The security, laundry and change room building will be located at the entrance to the high security area. This building will have a guard house, in/out one way turnstiles, a laundry room, and male and female change rooms, as well as an ablution section that will only be accessible from the high security area. The building will have an internal area of approximately 225m².

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18.4.6 Plant Workshop The plant workshop and storage facilities will be located within the process plant high security area. The workshop will comprise three bays with a total floor area of approximately 365m² and a 70m² apron immediately outside the workshop entrance. One of the three bays will be designated as a welding area and will be separated from the mechanical and electrical bays by a clad internal dividing wall.

18.4.7 Warehouse and Reagent Storage The warehouse and reagent stores will be located within the high security area. Delivery vehicles will report to the security office in the high security area for inspection before and after deliveries have been made.

The plant warehouse will have an approximate floor area of 360m². The internal storage within the warehouse will be comprised of high-rack shelving and open areas. The warehouse will have an eave height in excess of 6m to allow crane and forklift access. The warehouse will have a 30m² office space annexed to it for warehouse and supply staff.

Two reagent storage buildings are proposed, each with an approximate internal floor area of 180m². The buildings will be located to the northwest of the mill and wet plant. The buildings will be fully enclosed to provide weather protection for reagent bulk bags. The buildings will have internal storage racks installed and an eave height in excess of 6m to allow forklift operation within the building.

18.4.8 MCC Building (Switchroom Buildings) Electrical switchrooms will be located near the crushing building and the SAG mill. The layout of these switchrooms will be finalised during detailed engineering. There will be one 5m by 5m switchroom for the crusher plant and two 14m by 5m switchrooms for the wet plant.

18.4.9 Process Control Rooms The process control room will be a prefabricated building with dimensions of 8.4m by 3.3m. The building will be lined, air-conditioned and fitted with windows along both long sides, through which the operators will be able to view the mill on one side and the CIL circuits on the other. The control room will include a titration room. The control room will be accessed from the mill feed tower with a second means of egress from the CIL tank stair tower.

The crusher control room will be located atop the primary crushing building overlooking the ROM bin and primary crusher. The control room will be a single room building for crushing plant operators. The control room will be lined, air conditioned and fitted with lighting, power, data and windows. A communications link will be established between the crushing control room and the process control room in the wet plant.

18.4.10 Plant Office The plant office will provide approximately 70m² of internal area located within the process plant high security area. The building will contain office areas for the process maintenance department including the maintenance superintendent, plant foremen, maintenance supervisors, maintenance planner and the mill clerk. The building also provides an area for the cleaning office/supply area. The plant office will be located adjacent to the plant workshop.

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18.4.11 Gold Room Building The gold room will be a steel clad building with approximately 170m² of floor area. The building will house the leach reactor, calcine oven, electrowinning cells, smelting furnace, safe (enclosed within a concrete vault), and associated equipment as described in Section 17. A supervisor workstation will be installed in the gold room, equipped with a telephone and data connection.

The building will not be air-conditioned, but will be designed to allow natural ventilation in addition to the mechanical exhaust ventilation to be installed. Extraction fans will be installed above the thermal equipment to remove hot exhaust. The walls will include louvered sections to allow through-flow ventilation from exhaust fans to keep fumes, temperature, and humidity down. The roof will have a ridge vent. The entire building fabric (louvres and sheeting) will have steel mesh under it for security.

A secure area with inner and outer doors will ensure that the gold room remains sealed during bullion transfer to the transport vehicle. All operations within the gold room will be subject to full time closed circuit television (CCTV) surveillance with security alarms provided to the security coordinator.

Operational Water Supply

18.5.1 Raw Water Raw water will be supplied to the plant from the WSF located northeast of the process plant. Water will be transferred from the WSF via an 800m, 355OD HDPE PN10 pipeline fed by two water storage dam transfer pumps. A 620m³ capacity tank, situated adjacent to the tailings thickener, will supply the process plant with raw and fire water. The tank capacity provides for over 12 hours of raw water supply for the process plant. The bottom portion of this tank is a dedicated fire water source connected to the emergency diesel fire water pump. The fire water reserve in the raw water tank is sufficient for four hours of supply according to Australian Standard AS2419.

Raw water will also be provided to the mining contractor’s area by a 110OD HDPE PN10 pipeline from the plant raw water tank to a water storage tank within the area (provided by the mining contractor).

18.5.2 Potable Water Potable water will be produced by two dewatering bores located to the west of the M5 pit. Water will be delivered from the bores to a tank located at the accommodation camp, from which water will be fed to a 170m³ per day water treatment plant. The treatment plant will utilise filtration and chlorination to treat the water. Potable water for use in the process plant will be pumped from the camp to a 160m³ tank for use in the process plant and for distribution to the mining contractor’s area.

18.5.3 Water Management The process plant operators at the wet plant control room will control the water delivery from the WSF to the plant raw water tank. A telemetry system will be installed to provide reliable control from the plant control room.

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18.5.4 Sewage One sewage treatment system, located at the camp site, will be installed to service the plant buildings and the accommodation camp. Sewage from the plant and the contractor’s area will be pumped to the treatment facility at the camp via a pump station fitted with macerating sewage pumps. All sewage water will be treated before the treated effluent is pumped to the TSF.

Power Supply Heavy fuel oil (HFO) and grid power options were evaluated. While grid power is available from a HV powerline 13.5km south of the Project, current fuel prices support the HFO option. Therefore, a HFO power station will be constructed at the process plant by an independent power provider (IPP) under a build-own-operate (BOO) agreement. An 11 kilovolt feeder will be taken from the IPP distribution board to the process plant. The point of supply will be at the outgoing terminals of the supply feeder from the HFO power station which is where the tariff metering will be installed.

The power station will be fitted with three duty HFO generator engines and one standby generator to maintain supply in the event of a shutdown of one of the engines.

The power station will utilise a dedicated bulk HFO storage facility located adjacent to the power house.

The estimated power demand for the Project is provided in Table 18.1.

Table 18.1 Estimated Power Demand

Average Demand Annual Energy Consumption

(MWh) (GWh) Process Plant Area 10 – Crushing 0.18 1.6 High Voltage 0.64 5.6 Area 20 – Milling & Classification Low Voltage 6.15 49.2 Area 30 – Leaching & Adsorption 0.53 4.7 Area 40 – Tailings Management 0.04 0.4 Area 50 – Metal Recovery & Refining 0.09 0.8 Area 60 – Reagents 0.01 0.1 Area 70 – Services 0.62 5.3 Area 80 – Infrastructure 0.15 1.3 Camp 0.28 2.5 Mining Contractor 0.14 1.2 Total 8.42 69.0

11 kilovolt aerial transmission lines will be constructed from the substation to the TSF, the WSF, the Gibgo River abstraction dam, the accommodation camp and the mining contractor’s area.

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Mining Contractors Infrastructure An area west of the processing plant has been demarcated as the mining contractor area. The mining contractor will provide its own workshop, store facilities, and offices. The mining contractor will provide its own washdown area and waste oil management facility, which will be located within the mining contractor’s area. The washdown slab will incorporate a silt and oil trap, and an oil separator will remove any contaminant oil from the waste water before it is recycled at the washdown facility, with excess water used for dust suppression. The mining contractor will manage the safe removal of waste oil by using approved suppliers of waste oils, as required by law.

The treatment and disposal of sewage from the contractor’s area will be through the sewage treatment facility located at the camp. A macerating pump station will be installed in the mining contractor’s area to transfer the sewage via an overland pipe.

The explosive materials will be stored in a magazine located in a remote area to the west of the plant to provide separation from infrastructure and personnel. The magazine will be secured within a high security fenced compound and surrounded by embankments. Section 18.9.3 details the security to be established within this compound.

Communications There is no significant or reliable telecommunication infrastructure in the immediate mine site area. Telecommunications will be established by satellite link which will include voice, email and internet traffic for the process plant, camp and main office.

A conventional UHF radio system with hand held radios and chargers will be provided for site coverage. Radio communications will be via separate channels for mining and process plant. There will be a separate dedicated emergency channel.

Security Project security will be provided by a specialist local security provider and will consist of the following:

. Mining lease access control. . Read in/read out access control. . Fencing (double and single layer) and electronic security gates. . Electronic surveillance including CCTV. . Physical and visual barriers. . Lighting. . Patrols.

18.9.1 Process Plant The perimeter of the process plant will be surrounded by double layer high security fencing, with a 5m wide track between the outer and inner fences to allow for patrol access. The security fence will be 1.8m steel ring lock fence with razor wire atop the support posts.

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Access to the process plant will involve passing through two controlled access points. The first will be the electronic access gates to enter the administration area which is considered the low security zone beyond the first layer of perimeter fencing. Entry to the high security process plant zone will be through a second set of access gates. Personnel will only be allowed to enter the process plant through the security building and undergoing a metal detector scan.

Security guards will be stationed at both access points, within the process plant, warehouse and administration areas, and within perimeter lookout points. Continuous perimeter patrols will be implemented.

The electronic security and surveillance system will consist of a combination of various access control points, intruder detection devices, and a CCTV system consisting of approximately fifteen cameras located across the site. Some of the remote cameras and access control locations will be interlinked by a line-of-sight wireless network connection with a common receiver located appropriately to operate within line-of-site protocols.

18.9.2 Accommodation Camp The accommodation camp will be enclosed within a perimeter security fence with a vehicle track allowing patrol access around the entire perimeter inside the fence. A single controlled access point will provide entry to the camp with all vehicles and personnel entering camp entering via the security office.

Security guards will be stationed at the access point, in the security office, in perimeter lookout positions and on continuous perimeter patrols.

18.9.3 Infrastructure and Mining Contractor’s Area The TSF, WSF and Gibgo River abstraction dam will all be enclosed in high security fencing. Security patrols will routinely inspect these sites and access will be restricted by read- in/read-out access points.

The mining contractor will be responsible for erecting its own security perimeter fencing and establishing and operating its own security and surveillance systems. A gatehouse will control access to the area and the mine by means of a boom gate on the road outside the area. Heavy vehicles will enter through a 20m wide gate, which will be normally closed and opened as required by the gatekeeper at the beginning and end of shift, whilst light vehicles and visitors will be given access by the gatekeeper through a second smaller gate.

The explosives magazine will be within a fenced, high security compound with security guards permanently positioned there. Entry will be controlled via read-in/read-out access points. Electronic surveillance will be installed within the magazine so that monitoring can be done from the security building at the process plant.

Accommodation Camp The accommodation camp will be constructed prior to the commencement of process plant construction in order to accommodate construction personnel.

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The accommodation camp will be located adjacent to the site access road to the southeast of the process plant. A turn-off from the site access road will provide entry to the camp from a single gated and controlled access point. A car park will be constructed adjacent to the camp and security office. A perimeter access road will be constructed along the western perimeter of the camp to allow vehicle access from the entry gate to the small carpark at the southern end of the accommodation camp. A service track will be constructed along the eastern perimeter of the camp and a single layer security fence will surround the perimeter of the camp.

The accommodation camp and supporting facilities will be designed for an accommodation capacity of 172 people. The camp will be comprised of the following accommodation units:

. Four managers accommodation units (four person capacity). . Four 12 man accommodation dorms. . Three 36 man accommodation dorms. The camp will be supported by the following facilities/buildings:

. Kitchen/diner building. . Laundry building. . Recreation building. . Dry storage building. . Security office. Services will be provided from the process plant to the service compound located within the camp compound. As discussed in Section 18.5, power will be provided from an 11 kilovolt overhead powerline. Raw water will be provided to the camp raw water tank. As discussed in Section 18.4.2, potable water will be produced at the camp from a production dewatering bore and stored in a 180m³ tank for consumption in the camp. Potable water will then be distributed to the plant and mining contractor’s area via pipelines.

Project Implementation The project implementation strategy describes the organisation and philosophy that is necessary for the effective design, engineering, construction and commissioning of the process plant and associated infrastructure and services, together with the detailed schedules for each phase of the project development up to plant operation at rated capacity.

A goal for the execution phase of the project is the attainment of the best safety record possible. To accomplish this, all contractors and involved personnel will be required to adhere to defined safety objectives and standards developed by WAF and their consultants. These will include all appropriate safety requirements of Burkina Faso and will adopt as a minimum the safety standards required of any similar project if it was undertaken in Australia.

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The project implementation philosophy is to utilise an Engineering, Procurement and Construction Management (EPCM) approach where possible, with the builder’s risk remaining with WAF. That risk will be minimised by packaging many of the project components, particularly infrastructure, into fixed cost (lump sum) design and construct packages.

18.11.1 Project Management Plan The project delivery will be managed by WAF Managing Director in the role of Project Sponsor. The project delivery will be managed in two distinct parts comprising (i) Mining and (ii) Process Plant and Infrastructure. WAF will appoint Client Representatives for each part. This section deals with the organisational structure and details for the delivery of the Process Plant and Infrastructure only. Details of the implementation for the Mining part are given in Section 16. Mining development will be overseen and managed by the WAF Mining Manager.

WAF will appoint a Process Manager who will be responsible for all aspects of the process plant and infrastructure delivery. All project managers appointed by consultants and contractors will report directly to the Process Manager. The overall organisation chart is presented in Figure 18.5.

Figure 18.5 Project Organisation Chart

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Administration WAF will implement overall project administrative controls internally within their corporate office located in Perth, Australia. To support the Perth corporate office, WAF will establish a commercial and administrative office in Ouagadougou to handle project administration functions such as accounting, finance, personnel, supply, logistics and government relations functions.

The Managing Director will establish work authorisation reporting structures to keep abreast of project progress and enable corrective and preventative actions as necessary.

WAF will raise company orders for work in accordance with the Procurement Management Plan. The value of those orders will be recorded into the project accounting system in Perth together with the value budgeted for the order and the variance.

Each month, using the outputs from the accounting system and knowledge gained from reports from contractors and other project personnel, the EPCM Project Manager will produce a project cost report to the Process Manager which will identify the expenditure to date, the anticipated expenditure to complete the project, the budget variance to complete the project and the cashflow forecast (for the process plant and infrastructure).

In addition, the Project Manager will collate weekly short-form reports and more comprehensive monthly reports of project performance against schedule and budget for the Process Manager. The monthly reports will provide all stakeholders with adequate information regarding the cost and time performance of the project.

The EPCM consultant will undertake the basic project administrative and implementation tasks for the plant and infrastructure development. Overall project administration and control will be managed by WAF corporate administration in Perth.

WAF will establish administration, safety, occupational health and personnel policies for the project implementation and the same policies and procedures will be further modified and used for the operational phase of the project.

Payment of employees and contractors will be arranged by WAF through either the Perth or Ouagadougou offices. Following review and approval, the company will draw down sufficient funds against the project loan facility on a monthly basis.

Engineering The EPCM design engineers will provide design drawings, specifications and procurement documents for the new plant and infrastructure.

The plant design and documentation process will commence six months before the site works commence. The design engineers will provide ongoing support to the construction team throughout the construction period and will provide mechanical and electrical assistance during commissioning.

The major infrastructure components of the project will be awarded to specialist EPCM groups for the detailed design, procurement, documentation and QA/QC during construction of the HFO power supply, the TSF, WSF and access roads.

The Process Manager will co-ordinate the design consultants.

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Procurement Procurement of major capital expenditure items will be based upon recommendations received from the EPCM consultant. They will prepare the documentation, call for prices and tenders, prepare tender evaluations, negotiate prices with contractors and make recommendations to WAF in the form of drafted contracts and purchase requisitions.

Limited contracts will allow specialist contractors to work in their area of expertise and reduce cost through higher efficiencies. Wherever possible, lump sum contracts will be sought. It is anticipated that the six major site contracts will be Bulk Earthworks, Civil Construction, Structural, Mechanical and Piping Installation, Electrical and Instrumentation, Camp and Infrastructure Buildings, and 11kV Power Lines.

Minor procurement at the construction site will be performed by a clerk based upon approved requisitions. WAF will establish authorisation limits to ensure that expenditure is controlled.

All orders raised for the project will be part of a special sequence of orders to distinguish them from other company orders and will be typed directly into the project accounting system. WAF will establish a procedure to ensure that a separate register of orders is kept, referenced by order number, supplier and order description. These steps will ensure that the orders can be easily referenced in the future for plant maintenance and other purposes.

Items of major expenditure will incorporate contract terms and conditions appropriate to the level of expenditure. Minor expenditure will rely upon the standard conditions of purchase for company orders. Expenditure greater than US$500,000 will generally be documented as a standard form contract. Australian Standard contracts will be used for contracts awarded in Australia and FIDIC, DIN or ISO standards will be used for contracts awarded in Burkina Faso.

Construction Management The EPCM Project Manager will appoint subordinates to assist him with management of the construction sites. The personnel required will be:

. Construction Manager . Emergency Services and Occupational Health and Safety Manager. . Procurement and Logistics Manager. . Project Engineer(s). . Construction Supervisor (earthworks, civil, structural/mechanical, piping, electrical, mill installation). . Project Administration Clerk. The Project Manager will be the Superintendent of the contracts and will spend about 50% of his time on site. Once a substantial amount of construction work is underway the Construction Manager will be the Superintendent’s Representative on site full time, supported by the Project Manager on the opposite roster.

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Mining mobile equipment will be mobilised early in the construction phase and made available to the project. In particular, the mobile cranes, telehandler, integrated tool carrier, elevated work platform and light vehicles will be required by the construction management team.

18.11.2 Work Breakdown Structure The work will be divided into the following broad disciplines for the purpose of awarding contracts and placing orders:

. Bulk Earthworks . Civil Construction. . Transport and Logistics. . Structural Steel and Platework Supply. . Structural, Mechanical and Piping Installation (SMP). . SAG Mill Supply. . Mechanical Equipment Purchases. . Bolted Tank Supply. . Electrical and Instrumentation Purchases. . Electrical and Instrumentation Installation. . Major Pipelines. . Powerlines. . Communications. . Camp and Plant Infrastructure. . Power Generation. . Bulk Diesel Fuel Storage. . Bulk HFO Storage. . Design and Documentation. . Construction Management. . Security. . Commissioning. . Owner’s Costs. The work will be identified as either work that can be scoped for a firm price with the Contractor taking most of the risk away from the Principal (either paying for the worst case up front or having the risk present as contract variations), or work that cannot be scoped for a firm price without the Principal retaining the risk.

The work packages will be let as time and materials, firm price, or unit rate contracts or as a combination of all three. Where practical, lump sum firm pricing (with adequate provision for variations) is preferred.

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18.11.3 Project Schedule Construction is expected to take 19 months to complete from award of the process plant EPCM contract. In order to meet this schedule, the following design, procurement and construction activities have been identified as early works activities to be completed prior to a final investment decision by WAF:

. Surface Water Management and Sediment Control Structure Construction. . Accommodation Camp Procurement. . Accommodation Camp Site Earthworks Design. . Stage 1 Accommodation Camp Construction. . Site Layout and Process Plant Site Earthworks Design. . Bulk Earthworks Contract Tendering and Evaluation. . SAG Mill Specification, Tendering and Evaluation. The construction of the surface water management and sediment control structures is to be completed prior to the start of the 2017 wet season to allow for catchment of seasonal rains to provide a construction water supply to the project.

Stage 1 construction of the accommodation camp prior to the commencement of full site construction will ensure that sufficient accommodation will be available for construction personnel.

The remaining early works activities will enable site construction to commence as soon as possible after project commitment and remove the SAG mill supply from the project schedule’s critical path.

18.11.4 Operational Readiness Mobilisation of operations personnel will occur progressively over a 15 month period prior to production. Operations readiness will commence early in the project to initially accommodate the needs of open pit mining (commencing five months prior to production). This will incorporate infrastructure and the processing plant as the construction program progresses.

The Operations team will manage the following:

. Open pit mining contract. . Recruitment of operations and maintenance staff. . Pre-production training to operators and maintenance personnel. . Process consumables and reagents contracts. . Operational support contracts. . Operating and equipment care strategies.

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Management of the site will transition to the Operations team as personnel mobilise to site. Construction will become an activity within an operating mine site. This will require close monitoring and detailed definition of roles and responsibilities. The objective is to have the Operations team managing the site prior to commencement of ore feed to the process plant.

Operations

18.12.1 Introduction This operational implementation strategy covers the organisation structures, recruitment and employment policies and the accommodation and support facilities which are to be implemented.

The Resident Manager will be responsible for overall site operations and report to the WAF Chief Operating Officer. The total operating manpower for the project will be 636, including all staff and contractors. It is expected about 168 of that number will be WAF employees and the remaining 468 will be Contractors including the mining contractor staff, security personnel, medical staff and accommodation/catering contract staff. Where possible, the workforce will be drawn from the local area proximal to the project, however, the population of Nedogo village and surrounding region is not large enough to support an adequate workforce for the operation. It is anticipated that up to 50% of personnel will be drawn from the local area and more than 75% of all personnel will be Burkina Faso Nationals. Some personnel will be sourced internationally and operate on a fly-in, fly-out (FIFO) basis. The intended approach to nationalisation of the workforce will be to attempt reduction of start-up expatriate (international) numbers by 10% each year.

Expatriate staff requirements will be advertised and sourced from the broader international mining community, primarily from South Africa, Europe, Canada or Australia.

18.12.2 Transition from Construction to Operational Start-Up Prior to the commissioning and ramp-up phases, at the conclusion of the implementation phase, detailed operational readiness planning will be implemented by the pre-production team to ensure that a structured and orderly transition occurs from construction to operations. This transition will include the timely completion of construction, hands-on operations training and recognition of operation learning curves required to achieve full scale production. This transition will require close coordination and open communications between the construction, commissioning, and operations groups.

18.12.3 Site Operations Roster The processing operations will run 24 hours per day, seven days per week, on 12 hour shifts. Mining operations will be conducted to match the shift roster of the mining contractor, which is expected to also be 12 hour shifts on a continuous seven day per week roster.

The working rosters of staff will vary between local and expatriate staff and between operational departments. The operations strategy has been based on the following working rosters:

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. WAF Employees:

 Expatriate Staff: Six week on, three week off (fly-in/fly-out).

 Burkina Staff & Workers (Site): 10 day on, four day off.

 Burkina Staff & Workers (Ouagadougou): Five day on, two day off. . Mining Contractor:

 Expatriate Staff: Eight week on, four week off (fly-in/fly-out).

 Burkina Workers: 12 day on, two day off (10 hour shifts).

18.12.4 Operations Organisation Structure and Manning The operations strategy is based on the use of directly employed personnel in full time positions, except for mining, security, laboratory, power station and catering operations which will be undertaken by contract personnel. Operational support functions such as bullion transport, shutdown maintenance services, access road maintenance and freight services will be provided by service contractors. An Operational Manning Summary is given in Table 18.2.

Table 18.2 Operational Manning Summary

Classification WAF Personnel Contractor Personnel Management 2 0 Mining 28 241 Processing 39 12 (Laboratory) Maintenance 41 40 Administration (Site) 24 83 Caterer E & OHS 13 0 Administration (Ouagadougou) 16 0 Security 5 86 Power Station and Fuel Provider - 6 Total 168 468

It is anticipated that four departments in addition to service contractors will be required to support the management team namely Mining, Processing, Maintenance and Administration (which will include environmental, health and safety and security).

Employee Classifications The organisation chart and salary costing have been based on a three element classification system as shown in Table 18.3.

Table 18.3 Labour Classification System

Code Description E Expatriate Labour Source L Local S Staff Labour Type W Wage (non-staff) 1 Highest Level Labour Level 9 Lowest Level

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Using the classification system, a job described as LS4 would be a local staff employee with a moderate seniority level.

This classification system has been used by a number of mining companies developing projects in West Africa.

General Management Structure All operations at the mine site will be managed by the Resident Manager. The proposed senior management structure is shown in Figure 18.6.

Figure 18.6 General Management Organisation Chart

Mining The Mining Department will have three sections (Operations, Technical Services and Geology) under the control of the Mining Manager. An Organisation Chart is shown in Figure 18.7. Mining Department positions will be filled by a mixture of Expatriate and National employees. The total WAF workforce in the mining area is forecast at 28 persons.

WAF will require that the Mining Contractor operations are headed by a senior expatriate supervisor at an experienced level of at least ES3.

Processing The Processing Department will come under the control of the Processing Manager. The Processing Department will be located in a plant office, within the confines of the high security area. The Processing Department will be responsible for the following:

. All process operations from the primary crusher to the gold room including ROM pad stockpile management. . TSF and WSF operations and monitoring. . The laboratory. . Power reticulation. . Maintenance of the plant including purchasing of all process plant maintenance and operations requirements. The total workforce in the Processing area is forecast at 39 persons. A Processing organisation chart is given in Figure 18.8

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Figure 18.7 Mining Organisation Chart

Figure 18.8 Process Organisation Chart

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Maintenance The Maintenance functions will come under the control of the Processing Manager.

The Maintenance section will be responsible for all maintenance planning, scheduling and implementation of both mechanical and electrical maintenance work. The purchasing and warehouse management will be included within this area given that the most significant client for the warehouse will be the process plant. The Commercial Department will undertake QAQC including regular reviews to ensure that stock levels remain within acceptable norms and stock controls are being applied.

The total workforce of the Maintenance area is forecast at 41 persons. These positions will be filled by a mixture of Expatriate and National employees. A Maintenance organisation chart is given in Figure 18.9.

Figure 18.9 Maintenance Organisation Chart

Commercial and Administration The Commercial and Administration functions will be managed by the Administration Manager on site with governance review by the General Manager – Ouagadougou.

The Commercial and Administration Department will be responsible for the following functions:

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. Financial management reporting. . Accounts payable and account receivable. . Payroll and human resources. . Taxation and other regulatory (financial) obligations. . Administration functions. . Contracts and legal services. . Insurance. . Statutory reporting. . Government relations. The total workforce of the Administration area is forecast at 40 people. These positions will be filled by a mixture of Expatriate and National employees distributed between Ouagadougou and site.

A Commercial and Administration organisation chart is shown in Figure 18.10.

Sustainability The Sustainability department staff will manage the following:

. Environmental monitoring. . Rehabilitation. . Community relations. . Occupational health and safety (including training). . Nursing and First Aid. . Emergency Response functions. The environmental function will cover rehabilitation, compliance monitoring, reporting and ecology. It will be responsible for the implementation and ongoing performance of the Environmental Management Plan (EMP). The environment team will work to rehabilitate waste dumps and exhausted mine areas to achieve the clean sustainable return of the active mining areas to socially responsible land usage as jointly agreed with the long term land users.

The community relations function will manage community liaison and communicating the planning of the sustainable development of the mine within the community through joint initiatives that will see benefit and enrichment to the social fabric through understanding and open dialogue with the mine. Specifically the mine will assist with social capacity building programs that will focus on improved financial capacity, health, education, hygiene, and infrastructure.

Dedicated Safety Officers will monitor safety aspects of the entire operation and conduct safety training courses on a regular basis.

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Figure 18.10 Commercial and Administration Organisation Chart

This department will also be responsible for coordinating the emergency response team which will be drawn from the shift crews as necessary.

The total workforce of the sustainability department has been estimated at 13 persons. An organisation chart is shown in Figure 18.11.

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Figure 18.11 Sustainability Organisation Chart

Security Security functions on site will be managed by a security contractor. The Security department staff will manage all aspects of mine-site security function, inclusive of the contractor, and will be controlled by a senior employee.

The WAF security management team will consist of five members. The security contractor will report to the WAF security co-ordinator.

The security department will approve security clearances for all employees.

A Security organisation chart is shown in Figure 18.12.

Figure 18.12 Security Organisation Chart

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18.12.5 Workforce Make-Up In the first year of operation, the total project manning level is expected to be 636 inclusive of WAF and Contractor staff as detailed in Table 18.4.

Table 18.4 Total Project Manning Level Year 1

Burkina Faso Burkina Faso Company/Contractor Expatriate Salaried Wage West African Resources 31 50 87 Laboratory - 4 8 Maintenance - - 40 Mining 37 40 164 Catering 3 9 71 Security 2 13 71 Fuel Storage Facility - 6 - Total 73 122 441

The breakdown of WAF employees by expatriate, local staff and local non-staff on a departmental basis is summarised in Tables 18.5 to 18.7, which shows the personnel distribution for the LOM.

Table 18.5 Personnel Breakdown Year 1 Onwards

Expatriate Expatriate – Burkina Faso Burkina Faso- Department Total (Aus/UK/RSA) Other Africa Salaried Wage Management 1 - 1 - 2 Mining 3 6 8 11 27 Process 3 5 3 28 39 Maintenance 3 3 4 31 41 Commercial (Ouagadougou) 2 1 9 4 16 Commercial (Site) 1 - 10 13 24 E&OHS 1 1 11 - 13 Security - 1 4 - 5 Total 14 17 50 87 168

Table 18.6 Personnel Breakdown Year 3 Onwards

Expatriate Expatriate – Burkina Faso Burkina Faso- Department Total (Aus/UK/RSA) Other Africa Salaried Wage Management 1 - 1 - 2 Mining 2 7 8 11 27 Process 2 6 3 28 39 Maintenance 1 5 4 31 41 Commercial (Ouagadougou) 1 2 9 4 16 Commercial (Site) 1 - 10 13 24 E&OHS - 2 11 - 13 Security - 1 4 - 5 Total 8 23 50 87 168

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Table 18.7 Personnel Breakdown Year 6 Onwards

Expatriate Expatriate – Burkina Faso Burkina Faso- Department Total (Aus/UK/RSA) Other Africa Salaried Wage Management - - 1 - 1 Mining - 1 4 2 7 Process 1 6 3 28 38 Maintenance - 6 4 31 41 Commercial (Ouagadougou) 1 2 9 4 16 Commercial (Site) - 1 10 13 24 E&OHS - 2 11 - 13 Security - 1 4 - 5 Total 2 19 46 78 145

18.12.6 Contractors Mining Contractor All mining operations will be carried out by a suitably experienced open pit mining contractor. The contractor will also be responsible for the mining-related construction activities, including ROM pad and haul road construction and maintenance during operations.

The contractor is expected to supply its own maintenance equipment and facilities. Scheduled maintenance and repair of equipment will occur in the mine contractor’s workshops. Power and water will be supplied to the contractor’s workshop area. In all other aspects, it is expected that the contractor will be self-sufficient.

The mining contractor shall be responsible for the following:

. Drilling and blasting to produce minus 600mm crusher feed. . Loading and hauling. . Supply of ancillary equipment. . Equipment maintenance. . Ore rehandling. . Haul road construction and maintenance. . Waste dump construction. . Short term mine planning. . Procurement of mining supplies. . General Administration of mining activities. Advantages of contract mining include the following:

. A reduction in initial capital expenditure required. . Increased operational flexibility and the ability to change mining parameters (ramp up and ramp down mining rates). . Improved access to trained personnel, specialised services and equipment. . Allow WAF personnel to focus on gold recoveries and core business activities.

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ROM Stockpile Management ROM stockpile management will be shared by the Mining and Processing Departments. The Mining Department will engage the mining contractor to haul ore to and stockpile ore on the ROM pad. Feeding of the primary crusher will be done by the Processing Department.

Power Management Power for the project will be provided from the HFO power station. The power station will be a BOO facility supplying the power to the project on a cost per kWh unit basis. The power station operator will supply all operational and maintenance staff, consumables (excluding HFO) and spares.

Communications Management Central communications will be located at the administration offices. General mine and plant communications will be delivered by UHF mobile radio, local and repeater channels. Mobile radios are proposed in all vehicles and portable radios for mining, milling and maintenance staff.

A telephone network based on a dedicated bandwidth, broadband, voice and data package will be implemented on site and will service both the plant and mining administration requirements.

Communications protocols and requirements will be developed by WAF as part of the occupational health and safety management systems.

Accommodation and Catering Accommodation and catering will be provided by a suitably experienced catering contractor on a negotiated man-day rate, inclusive of all accommodation and catering costs including maintenance and camp operating personnel.

18.12.7 Occupational Health and Safety In addition to complying with requirements of the various statutory Acts and Regulations which govern workplace health and safety in Burkina Faso (and where there are gaps, Western Australian statutes will provide a minimum requirement), the Sanbrado Gold Project will develop and implement a site-wide occupational health and safety (OHS) management system to govern site operations. The key driver behind the development and implementation of these OHS management systems is the commitment to providing a safe, healthy working environment for all staff, contractors and visitors to the project.

The following objectives will form the core of the OHS management systems developed by WAF prior to commencement of operations:

. Ensure compliance with statuary regulations and maintain constant awareness of new and changing regulations. . Aim to eliminate or control safety and health hazards in the working environment to achieve the highest possible standards for OHS in the mining industry. . Develop and maintain a workplace culture that fosters behaviour among employees which reinforces that each employee is responsible for their own health and safety as well as that of their fellow employees. . Maintain the highest possible levels of staff morale and wellbeing. . Ensure effective communication of the OHS management system policies, standards and procedures.

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Some of the key features of the OHS management system include, but not limited to the following:

. Ensuring prospective employee suitability through a structured recruitment process. . Provision of relevant inductions and training to all personnel. . Promotion of fitness for work through the implementation of random or ‘for cause’ drug and alcohol testing. . Development of a suitable fatigue management policy. . Implementation of employee welfare programs to assist employees and their families. . Implementation of risk assessment procedures including job hazard analyses and safe work practices. . Active onsite promotion of various health and wellbeing topics and programs such as physical health and preventing heat stress and heat stroke. . Development of an emergency response team and regular emergency training – both onsite and through external specialist training organisations . Implementation of crisis management strategies, including live exercises including both site and Perth office personnel. . Standard requirements for a suitable number of occupation health level first aid personnel being onsite at any time. . Ensuring contracting personnel engaged in onsite work adhere to the OHS management systems and maintain the required safety standards. . Consulting with personnel at all levels regarding OHS issues and engaging them to initiate continual improvement within the systems.

18.12.8 Pre-Employment Procedures and Recruitment WAF will implement a structured recruitment procedure to ensure that prospective employees are suitable for employment. Some of the key stages and requirements involved in the procedure are as follows:

. Proof of formal qualifications are to be provided. . Reference checks are to be provided. . Pre-employment medical examinations are to be completed which include drug and alcohol screening as well as fitness for work assessments. . Pre-employment Gold Stealing Detection Unit (GDSU) police clearances are to be obtained for Australian personnel and equivalent for other nationals.

Personnel Inductions All new personnel will be inducted into the organisation in a structured, logical and coordinated manner to ensure they are aware of both their own and WAF’s responsibilities. The induction program will be developed as part of the OHS management systems. The induction program for all personnel will include a general site induction and workplace specific induction(s).

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Specific workplace inductions shall be performed in the following areas:

. Administration. . Active (Open Pit) Mining Areas. . Process Plant. . Tailings Storage Facility. . Remote Facilities. . Exploration. Recruiting Sequence Recruitment will commence with the early employment of senior local supervisors in the areas of environment, safety and community relations. These personnel will overlap with the construction phase of the project.

It is expected that the Resident Manager will be employed first during the first six months after formal approval of the project, and he will be responsible for employing the first level managers. These managers will be involved in determining exact manning levels in their departments and will oversee the recruiting process for their own departments.

The pre-production labour recruitment and training costs are shown in Section 21.

Recruitment Guidelines WAF has a recruiting policy implemented in Burkina Faso and intends to establish a recruitment committee within the local village groups in the region to ensure fair distribution of local labour.

Recruitment for the project will be carried out in accordance with the recruitment guidelines outlined in the following sections.

Internal Recruitment Jobs approved by management for filling shall be advertised by notices throughout the company.

External Recruitment The company intends to recruit employees who meet the criteria and qualifications and have suitable experience for the position. Existing employees who are suitably qualified will be given priority to fill a vacant position. Priority will also be given to local residents and Burkina citizens.

Selection Selection will be based on the principle of merit, in terms of the requirements of the position and the company’s objectives. The selection process will include the following:

. Aptitude testing. . Review of past experience. . Academic qualifications. . Personal interview. . Reference checking. . Medical examination.

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Employment Agreement A job description must accompany an Individual Employment Agreement in all cases, indicating position (job) title, roles, responsibilities and immediate reporting.

Prior to an Individual Employment Agreement being issued, all pre-qualifications, including psychological, aptitude, technical and practical testing, interviewing and pre-employment medical examination will have been completed and successfully passed in accordance with company standards.

Prior to an Individual Employment Agreement being issued, an Approval to Offer form must be obtained and signed off by senior management.

The Individual Employment Agreement will be in writing and appropriate to the classification of employment. The Individual Employment Agreement will be signed by the department manager.

Disclosure of the terms and conditions of Individual Employment Agreement will be strictly prohibited and may result in disciplinary action up to termination.

Probation Period During the initial three month probation period, the Company or the employee may unilaterally terminate employment without the obligation to provide a reason.

The HR section will send a notification letter to the employee’s supervisor at least three (3) weeks prior to the completion of the probationary period providing a Standard Evaluation Form of the probationary period.

A review of suitability for employment will be conducted at least fourteen days prior to the completion of the probationary period. The employee’s supervisor will be responsible for conducting a formal assessment to determine the individual’s suitability as a permanent employee.

Expatriate Replacement The Company will have the objective, where practical and not detrimental to operational performance, to replace some expatriate positions with local employees over the LOM.

This objective and its implementation will be assessed over time as the operation reaches a steady and performance optimal state. Initial intentions are to reduce the expatriate workforce by 10% each year.

18.12.9 Training and Development The majority of employees will be drawn from the project area. They will generally be unskilled and living in the vicinity of the village of Nedogo. Training programs will begin at a fundamental level, be delivered in French, and follow instructional methods that are known to be effective with the Burkinabe learning style.

Skills Training New employees in both the mine and process departments will be provided job specific training including a skills test. This training will be conducted at on-site training facilities and in the actual job location. Instructors will be Burkina Faso Nationals, vendor-supplied trainers, or WAF training personnel.

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Operator Training A substantial recruiting and training effort will be required to ensure that qualified operators are available for mining, processing and support operations. Supervisory roles within the operations departments will be filled by skilled expatriate workers who, along with vendor trainers, will supervise the development of operating personnel and the structuring of training to ensure the quality of operating personnel.

Operator trainees in the mine operations department will be required to attend mechanical equipment appreciation training. This general program emphasises basic mechanical principles involved in the operation of heavy equipment including a brief review of hydraulic and pneumatic systems, equipment components, and health and safety considerations.

The training programs will be complimented with on the job training provided by the skilled expatriate supervisors and foremen.

General Training In addition to training aimed at specific skills as described above, general training will also be carried out for various functions where applicable, including the following:

. Safety. . Environment. . Computer and Database. . Languages. . Supervisor and Management.

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MARKET STUDIES AND CONTRACTS No Market Studies were carried out for this study. The final product of the Sanbrado Gold Project will be gold doré bars. These can be sold in the current market at prevailing global gold prices.

No material contracts have been entered into as of the date of this report. Construction and mining contracts will be negotiated in the near future.

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ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT Environmental Studies Done and Relevant Environmental Issues An Environmental and Social Impact Assessment (ESIA) was prepared for WAF during 2015 and 2016 for the Tanlouka Heap Leach Gold Project (INGRID 2016). This ESIA and associated baseline surveys predicted that there will be generally minor impacts on local flora and fauna, air quality and visual amenity. The most significant impacts identified in the ESIA were noise and vibration levels during mine construction and operations, and socio- economic issues associated with the quarantining of approximately 749ha of agriculture and pasture lands (253ha are in the mining permit area) and the associated need to resettle approximately 220 households and approximately 674 residents.

The Burkina Faso Ministry of the Environment, Green Economy and Climate Change approved the ESIA for the Tanlouka Gold Project during December 2016. The final stage of the mine permitting process is the granting of a Mining Licence from the Burkina Faso Ministry of Mines, which occurred on 20 March 2017.

WAF has since commenced a CIL Feasibility Study which includes update and amendment of the approved Tanlouka Gold Project ESIA to reflect the larger Sanbrado Gold Project. Modelling the impacts of mining activity on local air quality, noise and vibration levels will be carried out and there will be further field work to identify the additional environmental and socio-economic impacts associated with the proposed water pipeline, road and dam site at the River Gibgo.

The Sanbrado Gold Project ESIA will also determine whether or not the updated project will reduce some overall impacts, including a reduction in the number of households subject to relocation due to changes in buffer zone requirements. The re-submitted ESIA will also identify community benefits associated with infrastructure remaining at cessation of mining such as the Water Storage Facility, Gibgo River Dam and water pipeline, the camp and electrical power line.

The key environmental and social risk associated with the project is water availability. Water availability for crop production and grazing is dependent on seasonal flows. Changes to those flows will be managed to accommodate the needs of the community whilst satisfying the water demand of the operation. The cumulative impact of this and other existing and proposed projects will need to be considered in the revised ESIA as the CIL based gold extraction process has a higher water demand than the presently approved project.

Existing baseline information suggests that current background levels of air borne particulates and some water quality parameters, related to specific chemical elements, exceed current Burkina and international standards. Dust (PM10) levels recorded during the dry Harmattan season were recorded at levels up to 10 times the Burkina standard of 40mg/m³. Water quality samples identified sites with cyanide and arsenic levels due to artisanal mining higher than the Burkina standards for surface water quality and higher than IFC standards for iron, dissolved solids and oils and greases. Compliance limits associated with these factors will need to be negotiated to permit exceedance of current regulated limits and make some allowance for any predicted inputs above those limits.

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The local community has a relatively low standard of living with significant challenges educating and retaining younger members of the community who could provide the energy and knowledge to assist in the improvement of socio-economic conditions. Local infrastructure is not well developed and the proposed mine will aid the upgrading of transport, power and water infrastructure in the area. It will also offer skills development and indirect business opportunities which can be sustained beyond the life of the mine. The mine will also provide a stimulus for educated and trained young people to remain in the community.

Control measures will be identified for all potentially negative environmental and social impacts to mitigate negative effects and enhance positive effects of the project. Monitoring and management plans will be developed to address key environmental issues such as air quality, surface and groundwater, noise and vibration and erosion and sediment control. A similar management plan is under development for the identified socio-economic impacts of the project.

Waste Management Hazardous and non-hazardous solid and liquid wastes generated at the mine site will be managed in accordance with local regulations. A landfill will be constructed onsite and the management of solid wastes will be addressed by an approved contractor in accordance with the Waste Management Plan.

A modular sewage treatment plant will be installed onsite and operated and maintained during the life of the mine. Both landfill and sewage treatment plant will minimise odour impacts. Upon cessation of mining, the sewage treatment plant will be removed and the landfill will be capped. Any hazardous waste component will be delineated in line with the provisions of the Mine Closure Plan.

Sediment Control The site is located within two key drainage catchments. The northeastern half of the site tends to drain to the northeast and the southwestern half to the southeast. The planned open pits are located in the base of a drainage catchment which discharges to the northeast. Surface water diversion channels will be required around all open pits.

A sediment control structure (SCS) is located at the site boundary of each discharge point for sediment control and compliance monitoring. Water quality in the SCS reservoirs will be monitored, and the water may be pumped back to the plant for use as process make-up water or for dust suppression, or released, depending on suitability of test results.

Within the site, water will also be harvested from structures such as the WRDs, TSFs and diversion bunds around the pits, and stored in settling ponds and diversion dams. This water will be classed as clean, contact or contaminated water and treated and rehandled accordingly. Contaminated water will be returned directly to the processing circuit.

Post-Closure Management

20.4.1 Closure and Rehabilitation At the end of the TSF and WSF operation, the external embankments will have an average downstream slope of 1:3.5. The profiles will be stable under both normal and seismic loading conditions, provide a stable drainage system, and allow for revegetation.

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The closure objectives include the following:

. Ensure a physically stable landform is achieved at decommissioning. . Decommission the facility at a maximum final height of not more than 50m, and/or in a manner which is physically compatible with the height of adjacent mine domains. . Develop appropriately designed and engineered seepage and surface water runoff barriers and intervention systems. . Manage any potential acid rock drainage (ARD) arising from tailings leachate so that it is captured by containment barriers or intervention systems constructed as part of the TSF. . Establish a low permeability soil cover over the TSF sufficient to promote development of a stable final landform profile, plant growth and a sustainable vegetation cover. . Integrate the final rehabilitated TSF landform into the surrounding landscape, with particular regard to blending constructed surfaces with the existing local topography. Once the final embankment height has been reached, slurry deposition into the TSF has ceased and an appropriate level of tailings consolidation has been achieved, decommissioning and reclamation can commence.

The principal aspects of TSF closure are:

. The protection of any exposed HDPE geomembrane liner with a soil layer of suitable thickness. . Perimeter slopes to be reshaped to an overall grade no steeper than 1:3.5 that ensures long-term stability and is safe for machinery to operate on during rehabilitation and future maintenance works. . Covering the top surface with a low permeability soil cover and topsoil layer and shaping this cover system to shed water in a controlled manner towards the engineered spillways that connect with the surrounding surface water management system. . The finished surfaces will be shallow contour ripped and seeded with locally adapted shrubs and grasses. . The TSF operational spillway will be upgraded on closure using geochemically neutral or non-acid forming material to provide the final closure spillway. The final closure spillway will be sized to discharge a Probable Maximum Precipitation (PMP) storm in a controlled manner from the facility to the surface water management system. The underdrainage systems will cease operation. The final soil cover profile for the tailings surface subsequent to decommissioning will be confirmed during operations based on operational tailings geochemistry test results.

In consideration of the anticipated chemistry of the TSF material on closure and the proximity of waste materials to watercourses that are used for domestic and agricultural use, a very low infiltration cover system (>90% reduction of net infiltration) is proposed for the TSF. Reducing infiltration into the TSF will reduce the potential for ARD and leaching of cyanide by minimising the quantities of leachate produced.

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A low permeability layer with a saturated hydraulic conductivity of less than or equal to 1 × 10-9m/s would be required to construct a cover system capable of reducing net infiltration to greater than 90% of the infiltration rate to be expected for a bare waste rock surface.

The main elements of the cover will be:

. Low permeability soil (1m thick) or synthetic liner to prevent ingress of water. . Protective fill layer (0.6m to 1.5m thick depending on liner system adopted) which may also act as ‘store and release’ capping. . Topsoil to allow revegetation (up to 100mm thick). Studies into the most appropriate cover system will be undertaken during the operational phase of the mine including availability of suitable fill materials. Once in place, the cover must not be removed and the planned land uses will prohibit activities which may result in its exposure, directly or indirectly.

The construction of the closure cover system will provide a landform suitable for revegetation with grasses and small native plants. Seeding and planting across the finished surface will provide long-term erosion protection and sustainability of the re-profiled landform.

20.4.2 Site Monitoring The monitoring of the mine area is vital to ensure that the environment impact is managed effectively and that negative impacts are minimised. Regular sampling and assessment of groundwater, surface water, air quality and soil resources will be undertaken throughout the life of the mine and conducted post-closure at defined intervals sufficient to satisfy legal requirements and the relevant conditions of project approval. Specific monitoring and management plans will be developed and implemented to address each of these areas.

In addition, a monitoring program will be developed for the TSF and WSF to identify any potential problems that may arise during operations. The monitoring program will comprise a combination of the following:

. Deep and shallow monitoring bores for deep seepage monitoring. . Surface water monitoring in the downstream sediment control structures for shallow seepage. . Standpipe piezometer and vibrating wire piezometers installed in the embankments to monitor pore water pressures within the embankments to ensure that stability is not compromised. . Settlement pins will be installed at regular intervals along the embankment crests in order to monitor embankment movements. As the Closure Plan is developed during the mine life, a set of completion criteria for rehabilitation, which are consistent with overall site closure objectives, will be determined and agreed with the regulator and relevant stakeholders. Through long-term monitoring of the site, it will be demonstrated that the development of rehabilitated areas is consistent with the completion criteria.

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The Monitoring Plan will include the following as a minimum:

. Physical stability monitoring. . Chemical stability. . Water Quality. . Reclamation. . Air quality. . Conditions at selected receptors. . Noise quality. . Environmental impacts and anticipated mitigation, management measures and associated monitoring. . Expected maintenance requirements. . Monitoring of staff transition from mining to other employment. . Monitoring of community initiatives. . Monitoring of community health and safety. . Monitoring of socio-economic activities. . Land resettlement, use and management. The Company expects that post-closure monitoring requirements will diminish over time, as closure and reclamation activities progress. The Company will conduct regular data reviews to identify trends in air, noise and water quality monitoring data and to monitor the success of reclamation activities. Proposal to change the post-closure monitoring requirements will made to the Government authorities and commenced on receipt of approval from those authorities.

Permit Requirements, Status of Permit Applications and Bond Requirements

20.5.1 Introduction The legislative and regulatory framework in Burkina Faso sets out the requirements for developing a mining project proposal and receiving project approvals. It also governs the construction, operation and closure of the project.

The Burkina Faso Code of the Environment stipulates that submission of an ESIA is mandatory for any mine whose production capacity is greater than 100tpd. The Burkina Faso ESIA development, review and approval procedure includes the following steps (from INGRID 2016):

. Preparation of a Project Description and draft ESIA Terms of Reference (proponent). . Review, approval and finalisation of the ESIA Terms of Reference (Burkina Faso Ministry of Environment, Green Economy and Climate Change (MEEVC)). . ESIA development (proponent). . Public consultation and information sessions (proponent). . Submission of the ESIA to MEEVC for review (MEEVC).

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. Review and preliminary analysis of the ESIA (National Office of Environmental Assessments (BUNEE)). . Consideration of the ESIA (Technical Committee on Environmental Assessments (COTEVE)). . Public inquiry (MEEVC). . Report of the Public Inquiry (MEEVC). . Decision on the approval of the project (MEEVC). . Granting of a Mining Licence, if the project is approved (Ministry of Mines).

20.5.2 Legislation The legal framework in Burkina Faso for environmental and social management of mining activities is based on the following legislation:

. The Constitution of Burkina Faso. . The Mining Code. . The Environment Code. . The Water Management Orientation Law. . The Forest Code. . The Local Authorities Code. . The Public Health Code. . The Labour Code. . Land Reorganisation. . Rural Land Tenure Law. The two codes that are most relevant to mining activities are the Mining Code and the Environment Code as outlined below.

20.5.3 Mining Code The Mining Code 2016 states that activities governed by the Code shall be conducted in a manner which shall ensure proper preservation and management of the environment and rehabilitation of mined and operated sites, in accordance with applicable standards, terms, conditions and regulations.

The Code provides for the creation of various funds, including a mining fund for local development, a fund for the protection and rehabilitation of the environment, and a fund for mining research. The rehabilitation and closure fund will be financed through a mandatory annual contribution from mining companies that will be determined based on the project environmental impact assessment.

Other relevant requirements of the Mining Code include complying with and protecting human rights and the rights of local communities affected by mining activities and promoting local businesses and the hiring of national and local workers.

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20.5.4 Environmental Code The Environment Code 2013 (established by Act 006-2013/AN of April 2, 2013) sets out public participation and information, prevention, precaution, “user-pays”, “polluter-pays”, sustainability and subsidiary principles as foundations for environmental management. It grants local populations the right to use natural and genetic resources, as well as share the benefits arising from exploitation of the resources.

Regarding mining activities, the Code provides in Article 37 that all mining operations/activities shall be performed so as to avoid damage or nuisance to the environment. The Code also establishes specific environmental requirements for classified facilities, including the following:

. Liability to inspections (Article 39). . Obligation to have an operational service in charge of environmental issues (Article 40) (Class 1 facilities only). . Obligation to produce annual environmental management reports (Article 41). . Obligation to pay an environmental licence and annual fee for inspections (Article 44). . Liability to a closure order in case of danger or serious inconveniences to the surrounding environment (Article 46). The rules for the application of these articles are established by decree as follows:

. Identification of classified facilities (Article 42). . Assignment of environmental inspectors (Article 39, Para. 3). . Opening and operating conditions for classified facilities (Article 43). . Amounts and distribution of licence and annual fee (Article 45).

20.5.5 Regulations All legislation is supported by implementing regulations. The following Burkina Faso regulatory instruments are relevant to the project:

. Order No. 2006/025/MECV/CAB of May 19, 2006, on the creation, assignment, composition and functioning of the Technical Committee on Environmental Assessment (COTEVE). . Decree No. 2007-853/PRES/PM/MCE/MECV/MATD of December 26, 2007, on specific environmental regulations for mining activities in Burkina Faso. . Decree No. 2006-590/PRES/PM/MAHRH/MECV/MRA of December 6, 2006, on the protection of aquatic ecosystems. . Decree No. 2006-588/PRES/PM/MAHRH/MECV/MPAD/MFB/MS of December 6, 2006, determining the perimeters of protection of water bodies and streams. . Decree No. 2005-515/PRES/PM/MAHRH of October 6, 2005, on authorisation and reporting of facilities, activities and works and undertakings. . Decree No. 2005-191/PRES/PM/MAHRH of April 4, 2005, on priority uses and governmental power regarding water control and distribution in times of shortage. . Decree No. 2005-192/PRES/PM/MAHRH/MFB of April 4, 2005, on procedures for preparation, approval, implementation and monitoring of land use and water management plans.

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. Decree No. 2005-193/PRES/PM/MAHRH/MFB of April 4, 2005, on procedures for determining boundaries of public domain water. . Decree No. 2001-342/PRES/PM/MEE of July 17, 2001, on the scope, content and procedure of the Environmental Impact Assessment and Environmental Impact Statement. . Decree No. 2001-185/PRES/PM/MEE of May 7, 2001, on setting standards for discharges of pollutants into the air, water and soil. . Decree No. 2001-731/PRES/PM/MJDH of December 28, 2001 (JO 2002 No. 05), on the adoption of the action plan and policy for promotion and protection of human rights. . Decree No. 2001-251/PRES/PM/MS of May 30, 2001 (JO 2001 No. 25), on the adoption of documents entitled "Strategic Framework for Fight Against HIV/AIDS (2001-2005)" and "Action Plan Against HIV/AIDS in Burkina Faso (2001)". . Decree No. 2001-624/PRES/MEF (JO 2001 No. 06) on the adoption of the National Population Policy. . Decree No. 99-198/PRES/PM/MFPDI of June 14, 1999 (JO 1999 No. 26), on the establishment of management bodies of good governance. . Decree No. 98-322/PRES/PM/MEE/CEC/MEM/MS/MATS/METSS/MEF of July 28, 1998, on the requirements for opening and operating dangerous, unhealthy and unsuitable facilities. . Decree No. 98/365/PRES/PM/MEE of September 10, 1998, on water policy and strategy. . Decree 97-054/PRES/PM/MEF of February 6, 1997, on Agrarian and Land Reform in Burkina Faso. . Enforcement Decree of the Mining Code. . Enforcement Decree of the Pastoral Law. . Order No. 68-7/PRES/J of February 21, 1968, on the Code of Criminal Procedure.

20.5.6 Environmental Protection Policies and Strategies Since the early 1990s, Burkina Faso has developed a number of natural resource management policies and strategies. The 1995 Mining Policy Statement stresses the importance of the private sector as a driver of economic development and the mining sector as an important factor in ensuring the country’s growth. The following environmental policies and strategies are relevant to mining activities:

. Strategy for Accelerated Growth and Sustainable Development (SCADD) – includes the objective of ensuring a sustainable environment. . Government’s Action Plan for Emergence and Sustainable Development (PAGEDD) – integrates the government’s priorities in gender, demography, environment, capacity building and economic intelligence. . National Environmental Policy (PNE) – includes providing for the rational management of natural resources and the assurance of a decent life in a better environment. . Environment Plan for Sustainable Development (PEDD) – includes encouraging stakeholders to consider the environment in their policies, strategies and actions for better preservation of the environment. . National Action Program for Adaptation to Climate Variability and Change (PANA) – includes ensuring a sustainable environment.

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20.5.7 Rural and Land Development Policies and Strategies The following rural and land development policies and strategies are relevant to mining activities:

. Rural Development Strategy (SDR) – includes the efficient management of natural resources, and the empowerment of populations in development. . National Policy on Land Tenure Security (PNSFMR) - aims to provide all rural stakeholders with equitable access to land, investment guarantee and effective management of land disputes, in order to reduce poverty, consolidate social peace and achieve sustainable development.

20.5.8 International Agreements, Protocols and Conventions Burkina Faso has signed and ratified several international agreements, protocols and conventions regarding environmental protection that are relevant to mining activities, including:

. Convention for the Protection of the World Cultural and Natural Heritage (Paris, 1972), ratified on April 2, 1987. . Convention on Wetlands of International Importance especially as Waterfowl Habitat (Ramsar, 1971), ratified on October 27, 1990, and the protocol for amendment (Paris, 1982). . Convention on Biological Diversity (1992), ratified on September 2, 1993. . Basel Convention on the Control of Transboundary Movements of Hazardous Wastes and their Disposal (Basel, 1989), ratified on November 4, 1999. . Rotterdam Convention on the Prior Informed Consent (PIC) Procedure for Certain Hazardous Chemicals and Pesticides in International Trade, ratified on March 14, 2002. . Stockholm Convention on Persistent Organic Pollutants (2001), ratified on December 31, 2004. . United Nations Framework Convention on Climate Change (1992), ratified on September 2, 1993, and agreement of the Kyoto Protocol (1997), ratified on March 31, 2005. . Vienna Convention for the Protection of the Ozone Layer (1985). . Montreal Protocol on Substances that Deplete the Ozone Layer (1987). . Convention on the Ban of the Import into Africa and the Control of Transboundary Movement and Management of Hazardous Wastes within Africa (Bamako). Additional conventions pertain to integrated water resource management, notably the Convention on the Status of the Volta River and the Establishment of the Volta Basin Authority. This Convention governs cross-border issues with, amongst others, Ghana and emphasises the following:

. Sustainable and equitable use of water resources. . Prohibition of causing significant harm to other states in the use of watercourses. . Prior notification for planned measures regarding resource management. . Emergency notification for any resource-related issue. As WAF develops management plans to mitigate environmental and social effects of the Sanbrado Gold Project, the plans will account for the requirements of the Codes and

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Regulations and reflect the policies and initiatives taken by the Burkina authorities to improve the environmental and socio-economic conditions in the local area.

Social and Community Related Requirements In line with Burkina Faso ESIA requirements, the potentially affected populations and their representatives have been consulted during the ESIA process for the Tanlouka Gold Project and will continue to be consulted as part of the Sanbrado Gold Project.

For the Tanlouka Gold Project, consultation focused on the four villages potentially affected by the proposed operating license of the project. These villages are Sanbrado, Pousghin, Manessé and Mankarga with a total population of 2,084 persons grouped into 250 households (see Table 20.1). Consultation was carried out at the household level during the ESIA process. However, not all of the 2,084 persons identified during the consultation process need to relocate. The Tanlouka Gold Project ESIA indicates that this smaller relocation subset is 122 households consisting of 674 persons (INGRID 2016).

Table 20.1 Communities Affected by the Project Operating License Boundaries

Number of Households Total Population Villages Neighbourhoods / Hamlets Workforce (%) Workforce (%) Silmiougou 36 14.40 368 17.66 Pousghin 21 8.40 174 8.35 Pousghin Roulghin 50 20.00 375 17.99 Noesse 29 11.60 223 10.70 Sanbrado Sanbrado 78 31.20 635 30.47 Manesse Pilaka 25 10.00 174 8.35 Mankarga Peulh Village 11 4.40 135 6.48 Total 7 250 100.00 2084 100.00

In addition to biophysical impacts, potential socio-economic impacts identified during the consultation process include the following issues (INGRID 2016):

. Reduced access and quarantine of approximately 749haof agricultural and pasture lands of which 253 ha are in the mining permit area. . The loss of homes on approximately 20 concessions including 122 households requiring the relocation of approximately 674 people. . The social impacts of resettlement including community disturbance, land use conflicts and integration issues. . Development of an equitable formula used to determine the amount of financial compensation received by those to be resettled. . Increased safety risks for those travelling on the access road or working at the mine site. . A potential influx of people to the area seeking access to the economic opportunities and improved supply of social services which may bring an increased risk of spread of communicable diseases and crime.

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. Changes in the health of the community arising from changes in ambient conditions associated with air quality, noise and vibration as well as exposure that may arise from unplanned access to the mine site. . Archaeological, culture and heritage issues related to the probable displacement of grave site(s) and a cemetery. . Visual amenity issues related to incremental light spill during mining as well as the change in view scape resulting from pit construction and associated new landforms constructed for waste rock and tailings management. Modelling the impacts of mining activity on local air quality, noise and vibration levels will be carried out as part of the Sanbrado Gold Project ESIA and there will be further field work to identify the additional environmental and socio-economic impacts associated with the proposed water pipeline, road and dam site at the Gibgo River.

The Sanbrado Gold Project ESIA will also determine whether or not the updated project will reduce some overall project impacts, including the number of households affected and subject to relocation because of changes in buffer zone requirements.

Recognised positive socio-economic effects of the project include local employment opportunities, improved local infrastructure and medical services as well as regional and national economic benefits. The re-submitted ESIA will also identify community benefits associated with infrastructure remaining at cessation of mining such as the Water Storage Facility, Gibgo River Dam and water pipeline, the camp and electrical power line.

Specific mitigation measures and management plans will be developed to address the identified environmental and social impacts of the project. From the socio-economic perspective, this will include a re-submitted Resettlement Action Plan as well as individual management plans for community natural resource management, integrated community awareness and archaeology, cultural and heritage. Other components of the mitigation and management approach are a stakeholder engagement plan, grievance management mechanism, community development program, an employment strategy, a sexually transmitted infections education policy and an accommodation code of conduct for the camp.

Mine Closure

20.7.1 General Approach WAF’s short-term closure and reclamation objectives (during construction and operations) can be summarised as follows:

. Minimise disturbance footprint as far as practicable. . Remove all available topsoil and subsoil material within the Project disturbance footprint and use immediately or store sustainably for future use; . Progressively reclaim disturbed areas. . Reduce the risk and impact of aeolian processes, water erosion, sediment transportation, surface and groundwater contamination and control of non-native species including weeds, feral animals and livestock. . Stabilise constructed landforms including re-profiled top surfaces and slopes. . Restore drainage patterns and flow volumes where required. . Cover ground to prevent soil drifting and generation of dust.

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The long-term closure and reclamation objectives are summarised as follows:

. Reclaim the land to a condition where long-term environmental degradation does not occur. . Reduce care and maintenance requirements. . Restore natural drainage patterns and flow volumes as far as practicable to the original conditions if required. . For areas which cannot be restored to the original conditions, rehabilitate the areas to develop landforms that are physically and chemically stable in the long-term and, where possible, are in keeping with the prevailing local topography; . Prevent physical or chemical pollutants from entering and subsequently degrading the downstream environment, including surface and ground waters. . Retain site infrastructure that may provide community benefits after closure to the mine, such as buildings, roads, water and power infrastructure. . Develop the site to achieve the long-term land use goals developed in consultation with the community and government. . Reclaim/rehabilitate the land to a condition where safety and environmental risks associated with the mine are low and enable the natural rehabilitation and/or reintroduction of a biologically diverse, stable environment where local communities can use the site without inheriting significant future liability. . Reduce adverse socio-economic impacts and provide positive socio-economic benefits. . Agree success/completion criteria with relevant stakeholders, monitor achievements against those criteria and report results to the stakeholders. . Ensure at all stages of operations that there are adequate and available funds to implement the closure plan. The development of a Preliminary Closure Plan allows integration with operational design concepts and ensures that the design and operation of the site are compatible with the closure plan. A Detailed Closure Plan will be developed after the Project has been commissioned. The Detailed Closure Plan will be reviewed periodically.

In order to progress the Preliminary Closure Plan to a detailed design stage, a number of consultations and studies will be undertaken during the mine operation to refine the plan. These are likely to include:

. Consultation with the local population to determine preferred final land use options including preservation of structures and infrastructure that can provide a resource for the socio-economic benefit of the community. . Consultation with government to establish final land use parameters. . Rehabilitation trials to assess the most appropriate revegetation strategies for each key area on the site. . Ongoing investigation and review of available rehabilitation materials and new technologies to improve rehabilitation outcomes. A closure land-use plan will be developed for the site in consultation with the relevant community and stakeholder groups. On closure, the open pit is planned to remain as an open void. The remaining areas will be rehabilitated to one of four land-use categories, namely scrub/grazing, agriculture, pond/wetland and landfill.

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Progressive rehabilitation will be undertaken, where practicable. All equipment and infrastructure will be removed from site, unless there is a beneficial use to a stakeholder or community. Such equipment and infrastructure includes buildings, water storages or roads which, if retained, would incur no liability to WAF.

Due to the potential for long-term environmental pollution, constructed landforms, such as the waste rock dumps (WRDs) and the TSF, will be covered with low permeability capping systems to reduce infiltration.

Surface water run-off and leachate from constructed landforms, such as the WRDs, will be controlled via a surface water management system that separates clean water from potentially contaminated water, provides settling areas to reduce the suspended solids load prior to discharge. The surface water management system would also allow treatment of water through the use of passive wetland systems to ensure that necessary standards are complied with before release downstream from the site.

Some materials arising from clearance of the site will have a commercial use and value. These may include generators, pumps and pre-fabricated buildings. Where practicable, these materials will be removed from site and either sold or reused elsewhere. Any materials which cannot be handled in this manner are considered wastes for the purposes of this closure plan.

If it is practicable to reuse a waste stream either on site or for a different application by a stakeholder group, then this will be adopted at no liability to WAF. Potential options include using old pipework for fencing or corrugated metal sheeting as roofing on community buildings.

Where waste materials cannot be reused, attempts will be made to recycle those materials through reputable off-site merchants and processing facilities. Types of waste that could be recycled include crushed concrete which can be used as aggregate and steelwork that can be sold as scrap metal.

If the waste cannot be reused or recycled then it will be disposed of in a suitable landfill facility. Three landfills will be constructed on site. One facility will accept domestic waste, the second will contain inert/non-hazardous waste and the third will be a hazardous waste landfill probably to be included within the TSF.

20.7.2 Closure Costs Closure costs will be accounted for in line with corporate accounting practices. WAF will estimate mine closure costs throughout the operational life of the Project and will accrue mine closure cost provisions on an annual basis. This will ensure that the accrued closure provision meets the costs of a planned closure event. In the event of temporary and/or unplanned mine closure, WAF will develop a care and maintenance regime agreed to by the applicable regulatory authorities.

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20.7.3 Cost Provisions for Mine Closure WAF’s approach to the estimation and management of costs related to the final closure of the Project at the end of the operational mine life comprises the following:

. Mine closure costs will be refined as mine plans are developed. A comprehensive closure cost estimate will be provided when the project design is finalised. This estimate will be revised throughout mine life and will be provided to the relevant government departments. . A provision for mine closure costs on WAF’s balance sheet to be determined in accordance with International Financial Reporting Standards including appropriate actuarial assumptions. . A commitment to review, monitor and update mine closure costs on a yearly basis (Annual Mine Closure Assessment). The Annual Mine Closure Assessment will provide updated estimates of date of closure (Mine Closure Date) and cost of closure (Mine Closure Cost) as well as an assessment of the date by which WAF will start setting aside cash contributions against the Mine Closure Cost (Closure Contribution Date). . The Closure Contribution Date will be the earlier of (i) 2 years before the Mine Closure Date or (ii) the date on which WAF forecasts aggregate post-closure finance cashflows from such date to the Mine Closure Date falls below twice the Mine Closure Costs. From the Closure Contribution Date, WAF will commit to making cash contributions a priority over shareholder dividend payments, ensuring that a dedicated mine closure reserve account is fully funded by the Mine Closure Date.

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CAPITAL AND OPERATING COSTS Operating Cost Estimate Project economics have not been updated in this report. Information from the February 20, 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018. Operating cost estimates for the Sanbrado Gold Project have been prepared to allow for the modelling of the proposed processing schedule, taking into account the variability of the ore feed from the different pit domains and the proposed blended mill feed. Four operating cases have been used to estimate the corresponding operating costs as follows:

. 2Mtpa throughput - 100% SOX Ore . 2Mtpa throughput - 100% MOX ore . 2Mtpa throughput - 100% WOX ore . 2Mtpa throughput - 100% FRS ore Ore characteristics vary between domains, in some cases significantly. The financial model for the project is constructed to reflect the proposed processing route and blending strategy LOM and the processing costs have been estimated to allow them to be aligned to both the mining and milling schedules.

The operating cost estimate covers the administration, owners mining, processing and maintenance costs. The operating costs have been estimated from a variety of sources including the following:

. First principle estimates. . Consumable consumption rates as provided by metallurgical testwork results. . Grinding power and media consumptions as provided by OMC. . Quotations for the supply of consumables and services. . Advice provided by WAF. . Published and unpublished costs for operations of a similar scale and location. All costs are in United States Dollars (USD) and reflect an estimate accuracy of ±15% at Quarter 1 2017 (Q12017) with a 90% level of confidence.

The following sections describe the basis of the operating cost estimates, details of assumptions and allowances made, and provide high level summaries of the estimate results.

21.1.1 Qualifications and Exclusions The following qualifications and exclusions apply to the operating cost estimates:

. ROM pad costs, including rehandling costs, are included within the mining costs. . A net amount of $4.4 million has been included in the sustaining capital costs for mine closure. This amount is calculated as the difference between the mine closure costs and the salvage value of the equipment and facilities. . Ongoing environmental costs are included in the General and Administration costs. . Corporate overheads/head office costs are excluded from this report.

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. Labour rates, consumables prices, diesel prices and HFO prices include applicable Burkina Faso on-costs and taxes. Other applicable taxes, duties and levies have been separately estimated by WAF as described in Section 22. . Project financing or interest charges are excluded. . Escalation is excluded. . Ongoing exploration costs are excluded. . The refining cost allowance of $2/oz has been estimated by WAF based on the costs incurred by a similar sized gold project located in Burkina Faso. Note: Subsequent to the completion of the pit optimisation, in which refining costs of $5/oz were used, WAF has consulted with other mining companies in Burkina Faso and has revised refining costs to closer reflect the current market rate.

21.1.2 Mining Costs Mine operating costs have been proposed by African Mining Services (AMS). A Request for Quotation (RFQ) document was compiled and forwarded to AMS. AMS is the subsidiary of an Australian based company with operations in West Africa. The RFQ included details on the production schedule, material types, mining elevations and pit and dump designs. The RFQ stipulated all the terms and conditions of tendering.

Owner’s mining costs relate to the salaries and wages of supervisory and support personnel employed by WAF to ensure efficient operations. This cost has been included in the General and Administration costs.

An allowance of $500,000/year has been made for dewatering. This relates to control and extraction of ground water and rain water runoff throughout the LOM. It covers the operating cost of pumps, consumables and labour.

Grade control cost relates to the cost of RC drilling and assaying for orebody delineation and grade estimation.

The average costs over the life of mine are shown in Table 21.1.

Table 21.1 Average LOM Operating Cost Summary

Mining Costs Rate US$ Million $/T Ore $/oz Drill and Blast $0.58/t $49.4 $2.93 $61 Load and Haul Waste $1.52/t waste $128.1 $7.61 $158

Contract Mining Load and Haul Ore $1.89/t Ore $31.8 $1.89 $39 Costs Crusher Feed/Stockpile Rehandle $0.72/t / $1.20/t $15.6 $0.93 $19 Contract Mining Fixed Cost $723,901/month $44.6 $2.65 $55 Total $269.4 $16.00 $333 Mining Supervision Included in General and Administration Costs

Owner’s Mining Grade Control $0.37/t ore $6.23 $0.37 $8 Cost De-watering $500,000/yr $2.58 $0.15 $3 Total $8.81 $0.52 $11

The total LOM mining costs of $278.2 million comprises pre-production mining costs of $11.9 million and total mining costs during the production period of $266.3 million.

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21.1.3 Processing and Maintenance Costs The process plant will treat a 50% SOX/MOX, 50% WOX/FRS material blend over the first 4 to 5 years at an treatment rate of 2.0Mtpa with nominal gold production of 107,300 troy ounces per annum.

The operating cost estimates take into account the planned crushing plant utilisation of 6,000 hours per annum (67.5%) and wet plant utilisation of 8,000 hours per annum (91.3%). The utilisation factors allow for plant downtime for scheduled and unscheduled maintenance activities.

A summary level breakdown of processing costs for each ore type is presented in Table 21.2.

Table 21.2 Processing and Maintenance Costs

Year 1+ Year 3+ Year 6+ Processing Costs $/t $/yr $/t $/yr $/t $/yr Labour $1.97 $3.94M $1.66 $3.31M $1.40 $2.80M Mobile Equipment $0.39 $0.77M $0.39 $0.77M $0.39 $0.77M Power $2.61 $5.21M $2.61 $5.21M $2.61 $5.21M SOX Consumables $2.29 $4.59M $2.29 $4.59M $2.29 $4.59M Maintenance $0.87 $1.73M $0.87 $1.73M $0.87 $1.73M Laboratory $0.45 $0.90M $0.45 $0.90M $0.45 $0.90M Labour $1.97 $3.94M $1.66 $3.31M $1.40 $2.80M Mobile Equipment $0.39 $0.77M $0.39 $0.77M $0.39 $0.77M Power $4.22 $8.44M $4.22 $8.44M $4.22 $8.44M MOX Consumables $2.18 $4.35M $2.18 $4.35M $2.18 $4.35M Maintenance $1.04 $2.08M $1.04 $2.08M $1.04 $2.08M Laboratory $0.45 $0.90M $0.45 $0.90M $0.45 $0.90M Labour $1.97 $3.94M $1.66 $3.31M $1.40 $2.80M Mobile Equipment $0.39 $0.77M $0.39 $0.77M $0.39 $0.77M Power $5.87 $11.73M $5.87 $11.73M $5.87 $11.73M WOX Consumables $2.44 $4.88M $2.44 $4.88M $2.44 $4.88M Maintenance $1.15 $2.31M $1.15 $2.31M $1.15 $2.31M Laboratory $0.45 $0.90M $0.45 $0.90M $0.45 $0.90M Labour $1.97 $3.94M $1.66 $3.31M $1.40 $2.80M Mobile Equipment $0.39 $0.77M $0.39 $0.77M $0.39 $0.77M Power $6.14 $12.28M $6.14 $12.28M $6.14 $12.28M Fresh Consumables $2.52 $5.03M $2.52 $5.03M $2.52 $5.03M Maintenance $1.15 $2.31M $1.15 $2.31M $1.15 $2.31M Laboratory $0.45 $0.90M $0.45 $0.90M $0.45 $0.90M

21.1.4 Labour Labour is a fixed yearly cost, independent of the processing blend and is therefore consistent across all operating scenarios. The biennial labour costs, reflecting the transition of expatriate to local staff are presented in Table 21.3.

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Table 21.3 Annual Labour Cost

No. No. Total Cost Unit Cost Department Personnel Expatriates ($/yr) ($/t)

Year 1 Onwards Management 2 1 $0.5M $0.25 Mining 28 9 $1.85M $0.93 Process 39 8 $2.06M $1.03 Process Maintenance 41 6 $1.88M $0.94 Commercial & Administration (Ouagadougou) 16 3 $1.17M $0.59 Commercial & Administration (Site) 24 1 $0.77M $0.39 E&OHS 13 2 $0.83M $0.41 Security 5 1 $0.23M $0.12 Total 168 31 $9.30M $4.65 Year 3 Onwards Management 2 1 $0.5M $0.25 Mining 28 9 $1.63M $0.81 Process 39 8 $1.84M $0.92 Process Maintenance 41 6 $1.48M $0.74 Commercial & Administration (Ouagadougou) 16 3 $0.97M $0.49 Commercial & Administration (Site) 24 1 $0.77M $0.39 E&OHS 13 2 $0.60M $0.30 Security 5 1 $0.23M $0.12 Total 168 31 $8.02M $4.01 Year 6 Onwards Management 1 - $0.03M $0.01 Mining 7 1 $0.20M $0.10 Process 38 7 $1.54M $0.77 Process Maintenance 41 6 $1.25M $0.63 Commercial & Administration (Ouagadougou) 16 3 $0.97M $0.49 Commercial & Administration (Site) 24 1 $0.57M $0.29 E&OHS 13 2 $0.60M $0.30 Security 5 1 $0.23M $0.12 Total 145 21 $5.40M $2.70

The labour costs associated with the owner’s mining team are excluded from the processing cost estimate and are instead accounted for as part of the mining cost estimate.

The manning levels and salaries/wages have been developed by WAF. The labour estimates have been benchmarked against similar projects in West Africa. The estimated manning represents the likely availability of skilled workers and the operating environment in Burkina Faso.

Expatriate labour costs are inclusive of Burkina government income taxes which are to be paid by WAF, social security and medical insurance costs, training levies and travel costs. Accommodation costs for personnel at the site accommodation camp are included in the general and administration costs. Labour costs for local Burkinabe staff and workers are inclusive of social security, medical insurance and training levy costs. For local staff that will not be accommodated in the camp, instead residing in the nearby villages, a 20% (of base salary/wage) housing allowance is included in the labour cost as well as the cost of casual meals whilst onsite.

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As discussed in the operations strategy, the manning is based on a tiered classification system to classify expatriate staff, local staff and local workers according to their employment classification and skill level. The classification system is detailed in Table 21.4.

Table 21.4 Labour Classification System

Code Description E Expatriate Labour Source L Local S Staff Labour Type W Wage (non-staff) 1 Highest Level Labour Level 9 Lowest Level

21.1.5 Mobile Equipment Mobile equipment costs are a fixed yearly cost, independent of the processing blend and are therefore consistent across all operating scenarios.

The numbers of mobile vehicles and plant have been estimated by Mintrex based on previous operational experience. The study assumes that all plant and equipment (with the exception of the FEL) are owned and operated by WAF. The purchase costs are included in the capital cost estimate whilst the operating cost covers all costs involved with the running, operation and maintenance of the vehicles. The cost of fuel is included in the estimate on the basis of the annual fuel consumption for each vehicle type. Fuel consumptions have been estimated by Mintrex based on assumed vehicle performance specifications and usage.

The operating cost of the FEL only covers the operation of the loader at the crushed ore stockpile rehandling ore to the emergency reclaim hopper. The operating cost of the FEL whilst feeding the primary crusher on the ROM pad is included in the mining cost estimate. The rehandle cost of $0.72/t has been provided by WAF.

21.1.6 Power Power supply will be from the HFO power station, provided by a independent power provider (IPP) under a build-own-operate (BOO) agreement. The ongoing operating costs of the power station and the supply costs of the power itself, inclusive of the HFO supply costs, are combined into a single unit power rate of $0.1924/kWh. This rate is comprised of the following:

. WAF has advised that the cost of HFO will be $0.6491/kg. . HFO consumption for power generation has guaranteed by an IPP at 207g/kWh (±5%). . The HFO cost for power generation based on the unit supply cost and average consumption rate is $0.1344/kWh. . The power station BOO charge has been provided by an IPP at $0.058/kWh.

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Plant loadings to determine the overall power demands were estimated by Mintrex in consultation with CardnoBEC (CBEC). The installed power summaries were developed on the basis of the BFS equipment selections and subsequent equipment powers as detailed in the Mechanical Equipment List (refer Section 17). The operating factors were determined based on the plant utilisation percentages outlined in the Process Design Criteria. The load factors were estimated by Mintrex and CBEC based on operating experience and historical power consumption data. A summary of the estimate power costs for each operating case are presented in Table 21.5.

Table 21.5 Power Costs

Annual Cost Unit Cost Consumption Operating Case ($/yr) ($/t) (kWh/t) SOX $5.21M $2.61 14.43 MOX $8.44M $4.22 23.37 WOX $11.73M $5.87 32.50 Fresh $12.28M $6.14 34.00

21.1.7 Consumables Reagent consumption rates are based on laboratory testwork result summaries provided by ALS, which provides the estimates for the varying reagent consumptions of the different ore types. Reagent costs have been based on budget pricing obtained from international and West African reagents suppliers.

Crushing and comminution consumables (equipment liners and grinding media) have been estimated by OMC based on circuit modelling and simulations. Budget pricing for these consumables have been obtained from the respective equipment suppliers (for the equipment liners) and from international and West African grinding media suppliers. Allowances have been made for crushing and comminution lubricants and general consumables based on operation experience.

The consumable costs differ according to the ore type. The variance in consumable costs across the four operating cases is summarised in Table 21.6.

Table 21.6 Consumable Costs

SOX MOX WOX Fresh Consumables $/t $/yr $/t $/yr $/t $/yr $/t $/yr Crushing $0.04 $0.02M $0.04 $0.02M $0.08 $0.04M $0.08 $0.04M Grinding $0.50 $0.25M $1.45 $0.73M $2.66 $1.33M $2.93 $1.47M Leaching & Elution $3.77 $1.89M $2.59 $1.29M $1.87 $0.93M $1.75 $0.87M Smelting $0.13 $0.07M $0.13 $0.07M $0.13 $0.07M $0.13 $0.07M Services $0.06 $0.03M $0.06 $0.03M $0.06 $0.03M $0.06 $0.03M Fuel $0.09 $0.04M $0.09 $0.04M $0.09 $0.04M $0.09 $0.04M Total $4.59 $2.29M $4.35 $2.18M $4.88 $2.44M $5.03 $2.52M

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21.1.8 Maintenance Cost Mechanical spares (excluding capital insurance spares), wear parts, contract labour and mill reline costs are accounted for as part of the maintenance cost. Maintenance costs have been estimated on a “by discipline” basis in which the maintenance costs are calculated as a percentage of the installed capital cost per discipline (mechanical equipment, structural steel, infrastructure etc.).

The maintenance costs for each ore type have been calculated as a percentage of the total maintenance cost estimate. The percentages applied to each ore type are provided in Table 21.7.

Table 21.7 Maintenance Costs

Maintenance Factor Total Cost Unit Cost Ore Type (%) ($/yr) ($/t) SOX 75% $1.73M $0.87 MOX 90% $2.08M $1.04 WOX 100% $2.31M $1.15 Fresh 100% $2.31M $1.15

Maintenance costs associated with the mining contractor’s fleets and facilities are excluded from the processing operating costs and are covered in the mining costs.

21.1.9 Laboratory Full onsite laboratory services (wet and dry laboratories) will be provided by a laboratory services provider. All maintenance, labour and consumables associated with the operation of the laboratory will be provided by the contractor and included in the services charge. The laboratory costing has been based on a quotation from a previous project in West Africa.

21.1.10 Water Supply Water will be sourced from the Gibgo River and pumped from the abstraction dam to the WSF located adjacent to the process plant. The operation and maintenance costs of the water abstraction and transport facilities and equipment are included in the processing operating costs as part of the power and maintenance costs.

21.1.11 General and Administration Costs General and administration (G&A) costs cover the fixed operational overheads of the project. G&A costs have been estimated on an annualised basis and are based upon WAF’s operational experience, benchmarking of G&A costs with similar operations and third-party pricing provided where applicable. The G&A costs are fixed costs, independent of process plant throughput.

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A summary of the fixed G&A costs are provided in Table 21.8.

Table 21.8 Estimated General & Administration Cost

Annual Cost ($/yr) Unit Cost ($/t) General & Administration Year 1 Year 3 Year 6 Year 1 Year 3 Year 6 Staff $3.50M $3.08M $2.40M $1.75 $1.54 $1.20 Ouagadougou Office $0.15M $0.08 Mine Site Office(s) $0.29M $0.15 Insurances $0.41M $0.21 Financial $0.11M $0.06 Government Charges $0.30M $0.15 Consultants $0.24M $0.12 Personnel $0.31M $0.15 Contracts $1.51M $1.50M $1.25M $0.76 $0.75 $0.63 Environmental $0.08M $0.04 General $0.56M $0.28 Total $7.46M $7.03M $6.11M $3.73 $3.51 $3.05

Staff Staff costs account for the cost of WAF’s administration, commercial, sustainability and security labour. The cost of this labour has been estimated as per the remaining labour estimate (refer Section 21.1.4).

Ouagadougou Office The Ouagadougou office cost estimate covers all costs associated with operating and maintaining the commercial operations located in Ouagadougou. These costs have been provided by WAF on the basis of in-country advice. Rental costs cover the cost of renting the commercial office space, whilst the communications network will operate on the back of the site communications system to allow communication between the two sites and WAF’s head office in Western Australia.

Additional costs in this area allow for IT infrastructure maintenance, replacements and minor upgrades, postage and light freight services, office stationary and other miscellaneous administration and commercial expenditure.

Mine Site Office(s) The site office cost estimate covers all costs associated with running and maintaining the onsite administration operations. The communications system makes allowance for the use of voice-over internet protocol (VOIP) phone systems and the utilisation of a dedicated communication server.

Additional costs in this area allow for IT infrastructure maintenance, replacements and minor upgrades, postage and light freight services, office stationery and other miscellaneous administration and commercial expenditure.

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Insurances General workers liability insurances (medical, disability and death) have been included in the labour cost estimate as part of the salary overhead costs. The following additional insurances have been allowed for as part of the G&A costs:

. Industrial Special Risks – Business Interruption. . Public and Products Liability. . Marine Transit (General). . Bullion. . Mobile Plant and Motor Vehicles. Quotations for these insurance policies have been provided by WAF based on actual quotations from insurance brokers.

Financial Allowances have been advised by WAF for specific financial fees and charges that are not included in the financial model.

Government Charges WAF has provided cost estimates for government charges associated with the operating licenses and the maintenance of exploration and mining tenement rights.

Consultants Allowances have been advised by WAF to cover the cost of utilising consultants to complete specific accounting, human resources and metallurgy activities.

21.1.12 Personnel The personnel cost estimate has been advised by WAF to cover the cost of recruiting, relocating (where required), obtaining visas and training staff and workers.

Contracts The contracts cost estimate covers the following specific service contracts to be provided to the project:

. Accommodation Camp. . Security. . Clinic. . Ouagadougou Hotels and Overflow Accommodation. . Ouagadougou Accommodation. The accommodation camp services will be provided by an accommodation and catering service contractor. WAF has provided a budget quotation for the provision of these services. The accommodation cost is based on a man day rate of $23.18 which has been developed from the quoted rates.

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Security services will be provided by an independent security services contractor which will operate under the management of the WAF security department. The cost estimate for the provision of security services has been based on a quotation from a previous project in Burkina Faso.

Medical clinic services will be provided by an independent medical services provider. WAF has provided a quotation for the provision of clinic services on which the cost estimate has been based. The clinic provider will be responsible for all costs associated with the operation of the clinic, including the provision of labour, equipment and supplies.

Environmental Allowances have been made for environmental testing and monitoring, along with nursery expenses to support the environmental rehabilitation plan to be implemented by WAF.

Allowances for environmental technicians and monitoring personnel have been included in the administration labour component of the G&A costs.

General The general cost estimate covers the costs of the administration departments’ light vehicles and the cost of the camp power supply. The light vehicle cost has been estimated as per the processing mobile equipment and vehicles (refer Section 21.1.5). The camp power supply cost has been estimated at the same rate as the process plant power cost estimate (refer Section 21.1.6).

21.1.13 Sustaining Capital A sustaining capital estimate has been developed by WAF in consultation with Mintrex and Knight Piésold, separate to the operating cost estimate. Sustaining capital has been estimated by WAF from first principles. General annual allowances have been allocated for the process plant and accommodation camp sustaining capital expenditures. The sustaining capital cost estimate is summarised in Table 21.9.

Table 21.9 Sustaining Capital Estimate

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Mining The mining sustaining capital estimate comprises three elements:

. Contractor mobilisation and site establishment. . Capitalised pre-strip costs. . Mining haul road construction.

Tailings Storage Facility The TSF sustaining capital estimate comprises three elements:

. Construction of a second TSF cell in Year 1. . Ongoing LOM costs for cell embankment raises. . TSF closure and rehabilitation in Year 9. Tailings Discharge The tailings discharge sustaining capital estimate covers the extension of the tailings discharge pipeline to allow tailings deposition in the second TSF cell constructed during Year 1. Second stage tailings discharge pumps, an extension to the existing tailings discharge pipeline and the deposition spigot piping around the embankment of the second cell are included in this sustaining capital.

Access Road The existing site access road will require a small diversion around the toe of the second TSF cell embankment when it is constructed.

Surface Water Management As site development continues during year 1 with the construction of the second TSF cell, additional mining haulage roads and open pit developments, additional sustaining capital will be required for the construction of diversion and sediment control structures.

Water Storage Facility The WSF sustaining capital estimate covers the closure of the facility at the conclusion of operations.

Mobile Equipment & Vehicles A number of replacement light vehicles have been allocated on an ongoing basis over the LOM to replace site operations vehicles as necessary.

Process Plant The process plant estimate comprises annual allowances for sustaining capital expenditure.

Camp and Administration The camp and administration estimate comprises annual allowances for sustaining capital expenditure.

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21.1.14 Owner’s Mining Costs The owners mining costs are those costs associated with management and technical supervision and support of the mining operation. The total annual cost is $1.98 million, including labour costs of $1.63 million and light vehicle running costs, software licences, and other consumable costs of $0.35 million.

21.1.15 Summary The average All-in Sustaining Cost for the project during the production period is estimated to be $759/oz, as summarised in Table 21.10.

Table 21.10 LOM Operating Costs

Operating Costs US$/t ore (processed) US$/Oz (produced) Mining 15.82 329 Processing 10.77 224 G & A 4.92 102 Cash Operating Cost 31.51 655 Royalties 2.89 60 Refining 0.10 2 Total Cash Cost 34.50 717 Sustaining Capital 2.03 42 All-in Sustaining Cash Cost 36.53 759

G&A is comprised of the costs summarised in Section 21.1.11 plus the community development fund contributions, duties, levies and surface taxes described in Section 22. The average community development fund contributions, duties, levies and surface taxes for the project during the production period are $0.81/t ore (processed) and $17/oz (produced).

Capital Cost Estimate The purpose of the capital cost estimate is to provide current costs suitable for use in assessing the economics of the project and to provide the initial control of capital expenditure. The estimated project capital cost is $124M.

The capital cost estimate is based upon an EPCM approach whereby the owner assumes the builder’s risk. As a result, the capital cost estimate does not include a builder’s margin.

The capital cost estimate has been prepared as a bankable level feasibility study and is presented in United States dollars (USD) to an accuracy level of ±15%, with a 90% level of confidence, as at Quarter 1 2017 (Q1 2017). The capital cost estimate is presented in both an area breakdown structure and a work breakdown structure to facilitate its future use as both a budget control document and source document for the future asset register of the operation.

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Table 21.11 summarises the capital cost estimate for the project, including contingency.

Table 21.11 Capital Cost Summary

21.2.1 Estimate Basis Mintrex has prepared the capital cost estimate for the process plant and associated infrastructure. Cost estimates for the TSF, WSF, Gibgo River abstraction dam and the sedimentation and surface water controls have been prepared by Knight Piésold (KP). Electrical and instrumentation capital costs for the process plant and infrastructure have been prepared by CBEC. Power supply capital costs have been prepared by ECG Engineering (ECG).

21.2.2 Estimate Assumptions and Clarifications The following assumptions and clarifications apply to the cost estimates:

. The capital estimate is based on an EPCM implementation strategy and the overall contracting strategy described in Section 18. It is assumed that the EPCM Engineer is based in Western Australia for the Engineering and Procurement phase. . Capital costs include the preliminary BFS bulk earthworks design and plant layout, both of which are subject to the completion of a detailed site survey. . Limited geotechnical data has been provided, and the preliminary earthworks design is based on best engineering knowledge and assumptions from geotechnical work completed in the same area of Burkina Faso. . Limited materials handling testwork has been undertaken with preliminary plant design based on best engineering knowledge and previous experience. . Mining will be performed by an experienced and competent mining contractor. . It is assumed that the cost of mining-related bulk earthworks, such as the construction of the ROM ore stockpile area, primary crushed ore stockpile and the mining haulage roads are the responsibility of the mining contractor and hence form part of the mining pre-strip costs. . It is assumed that sufficient material is available for local borrow stockpiles around the M5 pit for use as subgrade material and structural fill for bulk earthworks construction. . It is assumed that the power generation facility will be supplied by a build-own-operate (BOO) agreement with an independent power provider (IPP) with only the exclusions to the BOO agreement to be costed as part of the capital cost estimate. . The estimate is based on current wage rates and the expected site safety regulations and work practices. . It has been assumed that international air services are operating and sufficient seats on flights are available when required to meet the program schedule.

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. It has been assumed that sufficient manpower resources are available in Burkina Faso to undertake the project in the timescale envisaged. . It has been assumed that voice communications around the plant during the construction period will be by radio. . It has been assumed that vehicles for use by the EPCM team will be leased, with the lease to include both the vehicle and a driver. . It has been assumed that expatriate work permits for the construction management workforce are available from the Burkina Faso government.

21.2.3 Exclusions No allowance has been made in the cost estimates for:

. Escalation of prices. . Financing costs or interest. . Import duty for capital items. . Government approvals and special permits. . Currency exchange rate variations. . VAT (it is expected not to apply). . Special provision of process guarantees or performance warranties beyond the normal vendor obligations. . Owner’s sunk costs prior to project implementation approval. . Expatriate construction personnel taxation and employment law compliance costs. . Inclement weather delays. . Pre-production costs incurred by the mining contractor. . Working capital.

21.2.4 Estimate Detail The following section provides a description of the estimation methods used to develop the capital cost estimate.

Plant Bulk Earthworks A preliminary site bulk earthworks and drainage design was completed by Mintrex. From the preliminary design, a detailed bill of quantities was prepared for use in the estimate.

Unit rates applied to the bulk earthworks quantities were adopted from KP’s estimate for the construction of the TSF and WSF (refer Section 21.1.10). These rates were benchmarked against rates provided in responses to budget pricing requests from earthworks contractors operating in Burkina Faso.

Concrete Supply and Construction Concrete quantities were established using a combination of the general arrangement drawings and the 3D model prepared for the study, previous similar detailed designs and Mintrex’s database of quantities from previous projects.

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Concrete rates were based on detailed responses to requests for budget pricing provided by three civil contractors operating in Burkina Faso. WBHO’s pricing was adopted for the BFS as it was the median price received.

Allowances for concrete reinforcement have been included in the budget pricing received and quantities specified by Mintrex based on previous project experience. An additional allowance has been added to each of the bids to allow for concrete embedment’s and HD bolts as a ratio between the total embedment mass required and the total mass of concrete.

Structural Steel – Supply Structural steel quantities including light, medium, heavy gauge steel, grating, guarding, handrails and stair treads were established using a combination of general arrangement drawings and the 3D model prepared for this study, previous similar detailed designs and Mintrex’s database of quantities from previous projects.

Detail, supply and fabrication rates were based on responses to requests for budget pricing received from fabricators in Malaysia, South Africa and Spain. The rates for structural steel adopted are based on pricing from Horta Coslada in Spain.

As this study has assumed international fabrication of structural steel, an allowance has been included for expediting and quality assurance. This allowance has been developed on the basis of Mintrex’s past experience with international steel fabrication. The allowance covers the placement of an expediter in country at the fabricators works, including the associated flight and accommodation costs.

Shop detailing of structural steel has been estimated separately by Mintrex.

Platework – Supply Platework quantities, including shop fabricated tanks, bisalloy liners and rubber liners, were established using a combination of the general arrangement drawings and the 3D model prepared for the study, previous similar detailed designs and Mintrex’s database of quantities from previous projects.

Detail, supply and fabrication rates were based on responses to requests for budget pricing received from fabricators in Malaysia, South Africa and Spain. The rates for platework adopted are based on pricing from Horta Coslada in Spain.

Site erected tankage is based on the assembly of bolted tanks. Multiple budget quotations were received for the supply of bolted tanks for site assembly.

Shop detailing of structural steel has been estimated separately by Mintrex.

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Mechanical Equipment – Supply The major mechanical equipment items were sized to prepare the mechanical equipment list for the BFS, reflecting the process design criteria and the plant throughput. Equipment data sheets were prepared for all major equipment and these were issued to equipment vendors for budget pricing. Multiple pricing was requested for major items of mechanical equipment with minor items either being priced by a single vendor or from Mintrex’s database. Major equipment pricing was updated in Q42016.

Structural, Mechanical and Piping (SMP) Installation A Request for Quotation (RFQ) was prepared and issued to four contractors experienced in SMP installations in Burkina Faso. The RFQ scope was for the complete SMP installation for the process plant including the erection of the CIL tanks, supply and installation of major pipelines for tailings, decant, water supply and sewage transfer, erection of all principal supplied structural steel, platework and mechanical equipment. Contractor supply included all craneage and equipment to complete the works other than a defined list supplied by the Principal.

The study has adopted a combination of the budget pricing and estimates received from Group 5 and estimates based on Mintrex’s previous project experience. The following summarises what has been adopted for the BFS capital estimate:

. Preliminary and General costs as provided by Group 5, adjusted by Mintrex to reflect project duration. . SMP installation rates as provided by Group 5. . Structural steel installation man-hours as provided by Group 5. . Mechanical equipment installation man-hours are a combination of man-hours provided by Group 5 and estimates by Mintrex based on Australian labour with a productivity multiplier of 2.5 applied. . Platework man-hours as provided by Group 5. . Bolted tanks installation man hours as provided by Group 5. The total estimated installation hours provided by Group 5 were benchmarked against Mintrex’s total estimated installation hours and the estimates were comparable.

Piping – Supply and Install The supply and installation estimate for process plant piping was based on factors from historical project costs. These factors were based on a percentage of the mechanical equipment supply price and calculated for each area of the plant. An overall factor of 40% of mechanical equipment supply has been adopted.

Major pipeline costs for the tailings discharge, decant return, water supply and sewage transfer duties have been estimated separately.

Electrical and Instrumentation The electrical and instrumentation capital cost estimate has been prepared by CBEC based on the preliminary process plant design provided by Mintrex.

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Freight Freight has been estimated on the following basis:

. Structural Steel: cost per tonne . Platework: cost per tonne . Mechanical equipment: percentage of supply cost. . Electrical and instrumentation: percentage of supply cost. . Piping: percentage of piping supply cost. Freight costs for structural steel and platework have been estimated based on pricing obtained from logistics contractors for the shipment of 40’ shipping containers from Spain to Lome and the inland transport from Lomé in Togo to site and an assumed total tonnage of steel/plate per container.

The freight percentage to be applied to mechanical equipment, electrical and instrumentation and piping has been estimated based on costs provided by Antrak Logistics for a similar sized project completed in Senegal.

Tailings Storage Facility and Water Abstraction & Storage Facilities KP provided a detailed Bill of Quantities (BOQ) as well as design and supervision estimates for the construction of the TSF and associated decant water system. The capital estimate utilises rates from KP’s database of previous projects. The rates have been benchmarked against current budget pricing received against the BOQ from two earthworks contractors operating in Burkina Faso.

KP recommended constructing the TSF in three stages over the life of the project. For the purpose of this study, the cost of Stage 1 is included in the capital cost estimate and the remaining stages will be included as sustaining capital.

The primary water supply will be sourced from the Gibgo River abstraction dam via a WSF located adjacent to the TSF. KP provided a detailed BOQ as well as design and supervision estimates for the construction of the WSF. As per the TSF BOQ, the unit rates were benchmarked against current budget pricing received from contractors.

Plant Buildings, Camp and Equipment An RFQ was prepared and issued to suitably experience international contractors from Ghana, South Africa and Turkey for pricing. The RFQ included a scope of work, general arrangements and floor plans of all buildings.

DeSimone Group Ltd pricing was adopted for the construction of block buildings for the plant and accommodation camp. Where specific exclusions from DeSimone’s scope existed, allowances have been made by Mintrex on the basis of historical pricing obtained from previous projects.

Camp communications were estimated as per the project infrastructure communications.

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Project Infrastructure Access Roads Three access roads are included in capital cost estimate:

. Main access road from the existing unsealed road to the accommodation camp entrance. . Plant access road from the accommodation camp entrance to the plant site access control point. . Mining contractor’s area access road from the contractor’s area to a T-junction with the plant access road. The access roads formed part of the preliminary surface earthworks design completed by KP. A detailed and priced BOQ was provided for the access road construction and as per the other KP estimates, the unit rates were benchmarked against pricing from earthworks contractors operating in Burkina Faso.

Communications A detailed scope of work was provided to an international communications provider by CBEC, to cater for the required voice and data communication infrastructure. The cost estimate is based on the budget pricing received in response to the scope of work.

Major Pipelines Major pipeline costs for the tailings discharge, decant return, water supply and sewage transfer duties have been estimated separately based on budget pricing from contractors operating in Burkina Faso for the complete supply and installation of the pipelines. Additional allowances have been added to the supply costs to cover exclusions to the supply and installation.

Security Plant site security has been estimated based on pricing received from a security system provider in South Africa. The plant site security estimate covers the cost of the supply and installation of CCTV and access control systems.

Dual layer (inner/outer) high security fencing has been allowed for around the perimeter of the plant site along with single layer high security fencing around the perimeter of the high voltage substation and the administration area. Pricing per linear metre of security fencing has been obtained from local contractors.

Power Supply The power supply options for the project were investigated by ECG and included capital cost estimates for HFO power generation and grid power. The capital cost estimate for the HFO power station covers only the exclusions to the BOO contract, including the construction of a bulk HFO storage facility. The supplied power will be distributed to the entire mine site area via a reticulated 11kilovolt HV transmission system. All capital costs associated with the electrical reticulation downstream of the power station are included in CBEC’s capital cost estimate.

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Plant Mobile Equipment The number and type of vehicles/plant in the mobile equipment fleet have been nominated by WAF. All plant and equipment is assumed to be owned and operated by WAF with the exception of the FEL which will be owned and operated by the mining contractor. Budget pricing has been obtained by WAF based on a plant and equipment list provided by Mintrex.

Engineering, Procurement and Construction Management (EPCM) The EPCM estimate was established from first principles based on an assessment of man-hours required per EPCM discipline; in accordance with the approach outlined in the Project Implementation Strategy (Section 18). The EPCM man-hours associated with process engineering, design engineering, drafting, procurement and expediting were developed using the mechanical, electrical and instrumentation lists, process plant description, preliminary general arrangement drawings and preliminary 3D model developed for this study.

The project and construction management is estimated as a time-based EPCM cost established on the basis of previous project experience and the implementation schedule that has been developed for this study (refer Section 18).

Current Mintrex and CBEC rates were applied to the estimated EPCM man-hours to develop the overall EPCM cost estimate.

The estimate includes EPCM overhead costs such as site vehicles, flights, accommodation, IT hardware and miscellaneous expenses. The overheads were estimated as follows:

. EPCM site vehicle hire cost was estimated by Mintrex based on the number of vehicles required for the EPCM team during the construction time at $1,500 per vehicle per month. It is assumed that a number of operations light vehicles will be purchased early for use by the EPCM team. The monthly cost covers the operation of the vehicles only. . Accommodation costs have been included for the onsite EPCM team based on an estimated number of man-days required onsite at $23.50/man day as advised by WAF. . Flight and travel costs have been included for the EPCM team based on the estimated number of flights required from Australia to Burkina Faso. The flight cost which has been adopted is $9,750 return business class. . An allowance has been made for the purchase of IT infrastructure for the EPCM site offices. . Miscellaneous expenses estimated at $20,000 per month for the duration of the site construction.

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Temporary Construction Facilities The construction overheads estimate for the owner’s components of the project was derived as follows:

. Temporary construction facilities cost was estimated by Mintrex based on 0.35% of pre- contingency process plant direct cost. . Temporary construction equipment cost was estimated by Mintrex based on 0.35% of pre-contingency process plant direct cost. . An allowance has been made for the purchase/lease of diesel generators to provide temporary construction power to contractor facilities and the EPCM office. Additional costs for the provision of emergency generators for the process plant and camp are included in the electrical sections of the capital cost estimate. Capital Spares Budget pricing was received from all major equipment vendors and a detailed list of insurance spares was prepared. Commissioning and consumable spares were estimated by Mintrex as 2.5% and 5.0% of total mechanical equipment supply cost respectively.

First Fills Quantities for first fill consumables were assembled from first principles using the process design criteria. First fills allow for the initial SAG mill ball charge and the initial carbon circuit stocks along with 3 months of stock for consumables and reagents based on the design consumption rates.

Additional allowances have been included for first fills of equipment lubricants and initial stocking of the plant warehouse for miscellaneous items other than mechanical equipment spares.

Construction Overheads Construction overheads estimates have been included for each major horizontal site contract. The construction overheads have been based on the estimates provided by contractors as part of their budget pricing. Mintrex has adjusted the estimates overheads where applicable to account for variations in the work under contract and/or the contract duration.

In general, the construction overheads include the following cost items:

. Mobilisation and demobilisation of contractors and sub-contractors. . Site establishment and site facilities. . Travel costs for contractors and sub-contractors.

Owner’s Costs Owner’s costs (with the exception of pre-production labour costs) have been estimated on the basis of advice from WAF.

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The pre-production labour costs have been developed based on an operational team manning forecast developed by WAF. This forecast aligns with the implementation strategy (refer Section 18) in terms of the recruitment of key operational personnel during the construction and commissioning phases. The salaries and overhead costs for operational personnel are those adopted by the operating cost estimate (refer Section 21) which have been provided by WAF.

21.2.5 Contingency The purpose of the contingency is to make specific provision for uncertain elements of cost within the project scope and thereby reduce the risk of cost overrun to a pre-determined acceptable level. Contingencies do not include allowances for scope changes, escalation or exchange rate fluctuations.

Contingency reflects the measure of the level of uncertainties related to the scope of work and associated assumptions and omissions. It is an integral part of an estimate and has been applied to all parts of the estimate, i.e. direct costs and indirect costs. Mintrex has recommended the contingency rates shown in Table 21.12:

Table 21.12 Capital Cost Estimate Contingencies

Area / Item Contingency Earthworks 20.0% Supply 5.0% Mechanical Equipment Install 7.5% Supply 10.0% Platework Install 10.0% Supply 10.0% Structural Steel Install 10.0% Civil 15.0% Supply 12.5% Piping Install 12.5% Supply 15.0% Electrical & Instrumentation Install 15.0% Tailings Storage Facility 10.0% Plant Buildings 10.0% Project Infrastructure 10.0% Site Access Roads 10.0% Plant Vehicles & Mobile Equipment 10.0%

21.2.6 Exchange Rates The capital cost estimate is presented in USD. Where cost estimates have been included in foreign currencies, the exchange rates shown in Table 21.13 have been applied.

As discussed in Section 21.2.3, no allowances have been made in the capital estimate for variations of these foreign exchange rates.

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Table 21.13 Exchange Rates (Q1 2017)

United States Dollars Currency (USD) Australian Dollars (AUD) 0.740 Euro (EUR) 1.050 Canadian Dollar (CAD) 0.742 South African Rand (ZAR) 0.071 British Pound (BDP) 1.250 Central African Franc (XOF) 0.002

21.2.7 Statistical Analysis The capital cost estimate has been completed to a 90% level of confidence meaning there is a 10% likelihood that the capital costs for the project will be less than or exceed the estimated $124M (standard deviation probability calculation). The calculated estimates have been statistically evaluated to provide an estimate of accuracy of the overall estimate.

A capital overrun facility allowance of $5.4M in addition to the $12M of contingency is suggested by the analysis.

The statistical analysis done is provided as part of the capital cost estimate. The capital cost estimate by discipline is detailed in Table 21.14.

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Table 21.14 Sanbrado Gold Project Capital Costs by Discipline

Installation Hours Material Installation Freight Grand Subtotal Contingency Concrete E&I SMP Total Costs Costs Costs Total Process Plant Costs Construction Overheads - - - - $ 304,985 $ 5,780,138 $ - $ 6,085,122 $ 548,801 $ 6,633,924 Bulk Earthworks - - - - $ 699,975 $ - $ - $ 699,975 $ 139,995 $ 839,970 EPCM - - - - $ 10,541,055 $ - $ - $ 10,541,055 $ 1,054,106 $ 11,595,161 Primary Crushing 9,791 7,890 28,017 45,699 $ 2,622,311 $ 614,343 $ 376,806 $ 3,613,460 $ 338,678 $ 3,952,137 Milling & Classification 32,627 22,223 109,618 164,469 $ 13,480,014 $ 2,640,139 $ 2,196,574 $ 18,316,726 $ 1,558,498 $ 19,875,225 Leaching & Adsorption 9,189 10,191 53,315 72,695 $ 6,201,862 $ 1,085,554 $ 1,009,929 $ 8,297,345 $ 773,590 $ 9,070,935 TSF & Decant Return 88 5,993 7,021 13,102 $ 256,974 $ 131,088 $ 30,810 $ 418,872 $ 52,076 $ 470,948 Metal Recovery & Refining 2,538 6,281 18,164 26,984 $ 1,710,314 $ 446,905 $ 249,964 $ 2,407,183 $ 211,051 $ 2,618,235 Reagents 1,254 3,496 6,138 10,888 $ 459,986 $ 164,196 $ 67,257 $ 691,439 $ 66,249 $ 757,689 Services 3,969 9,600 12,370 25,939 $ 1,583,246 $ 400,011 $ 193,026 $ 2,176,284 $ 225,921 $ 2,402,205 Total Process Plant Costs 59,456 65,674 234,644 359,774 $ 37,860,722 $ 11,262,373 $ 4,124,367 $ 53,247,461 $ 4,968,966 $ 58,216,427 as % of Total Plant Costs 17% 18% 65% 65% 19% 7% 91% 8.5% Infrastructure Tailings Storage Facility - - 25,377 25,377 $ 11,594,740 $ 748,621 $ 74,287 $ 12,417,648 $ 1,241,765 $ 13,659,413 Process Plant Infrastructure 11,742 - - 11,742 $ 7,043,526 $ 2,313,791 $ 359,547 $ 9,716,865 $ 1,163,468 $ 10,880,333 Camp 4,603 2,187 - 6,790 $ 6,356,138 $ 135,249 $ 92,346 $ 6,583,732 $ 722,707 $ 7,306,439 Power Supply - - - - $ 2,209,464 $ - $ - $ 2,209,464 $ 220,946 $ 2,430,410 Plant Vehicles & Mobile Equipment - - - - $ 2,203,200 $ - $ - $ 2,203,200 $ 220,320 $ 2,423,520 Other Costs Temporary Construction Facilities - - - - $ 572,732 $ - $ - $ 572,732 $ 57,273 $ 630,005 Capital Spares - - - - $ 2,843,189 $ - $ 463,411 $ 3,306,600 $ 330,660 $ 3,637,260 First Fills - - - - $ 1,306,105 $ - $ - $ 1,306,105 $ - $ 1,306,105 Owners Costs Construction Insurance - - - - $ 425,000 $ - $ - $ 425,000 $ 63,750, $ 488,750 Compensation - - - - $ 2,808,966 $ - $ - $ 2,808,966 $ 421,345 $ 3,230,311 Pre-Production Labour - - - - $ 4,674,046 $ - $ - $ 4,674,046 $ 701,107 $ 5,375,153 Pre-Production Expenses - - - - $ 7,524,709 $ - $ - $ 7,524,709 $ 1,128,706 $ 8,653,415 Mining Pre-Production - - - -$-$-$-$-$-$- Owners General - - - - $ 300,000 $ - $ - $ 300,000 $ 45,000 $ 345,000 Owners Project Management - - - - $ 3,835,450 $ - $ - $ 3,835,450 $ 575,318 $ 4,410,768 Business Systems - - - - $ 820,000 $ - $ - $ 820,000 $ 123,000 $ 943,000 Training - - - - $ 200,000 $ - $ - $ 200,000 $ 30,000 $ 230,000 Working Capital - - - -$-$-$-$-$-$- Total Indirect Costs 16,344 2,187 25,377 43,908 $ 54,717,265 $ 3,197,661 $ 989,591 $ 58,904,517 $ 7,045,365 $ 65,949,882 as % of Total Indirect Costs 83% 5% 1.5% 89% 10.7% Total Project Costs 75,800 67,861 260,021 403,683 $ 92,577,987 $ 14,460,034 $ 5,113,957 $112,151,978 $ 12,014,332 $124,166,310 as of % of Grand Total 19% 17% 64% 75% 12% 4% 90% 10%

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ECONOMIC ANALYSIS Summary Project economics have not been updated in this report. Information from the February 20, 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid-2018. The Sanbrado Gold Project has been evaluated on a discounted cashflow basis. The results of the analysis show the Sanbrado Gold Project to be economically very robust. The present value of the net cashflow with a 5% discount rate (PVNCF5%) is $143 million on a pre-tax, project basis, using a base gold price of $1,200/oz. Project post-tax PVNCF5% at a $1,200 gold price is $100 million on an all equity basis. Project internal rates of return (IRR) are respectively 27% pre-tax and 21% post-tax.

The government of Burkina Faso is entitled to a 10% free-carried interest in the mine operating entity. The economic analysis assumes that WAF will provide all development funding via loans to the mine operating entity, which will be paid back with interest from future gold sales. On this basis, WAF’s 90% interest in the project is expected to provide a post-tax PVNCF5% of $91 million and a post-tax IRR of 20% at a gold price of $1,200/oz.

Project payback period at a gold price of $1,200/oz is expected to be 2.1 years on a pre- tax basis and 2.3 years on a post-tax, all equity basis. Payback period is defined as the time after process plant start-up that is required to recover the initial expenditures incurred developing the project. This is the point in time when the cumulative undiscounted net cashflow is zero.

The cashflow analysis has been prepared on a constant first quarter calendar 2017 US dollar basis. No inflation or escalation of revenue or costs has been incorporated.

Mine Production and Mill Feed The mine production schedule presented in Section 16 is summarised into an annual mine production and mill feed schedule as shown in Table 22.1. Life of mine mill feed totals 16.8Mt at an average grade of 1.7g/t Au. This feed is comprised of 1.3Mt at an average grade of 5.7g/t from the M1 pits, 0.1Mt at an average grade of 1.8g/t from the M3 deposit and 15.3Mt at an average grade of 1.3g/t gold from the M5 deposit. The average mill feed grade during the first three years of operations is 2.7g/t Au.

Mining commences five months prior to the commencement of mill operations and continues for the first five years of processing plant operations. The ore is milled over an 8.75 year period. It is assumed that ore will be either directly fed to the crusher or stockpiled adjacent to the crusher and rehandled to the plant as required to meet the processing schedule. Table 22.1 shows that significant run of mine and low grade ore stockpiles are maintained throughout the mine life. These stockpiles represent a significant asset and a buffer between the mine and the processing plant.

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The ore processing rate is a function of the testwork results summarised in Section 13 of this report and the process flowsheet design described in Section 17. A lower plant throughput rate is assumed during the initial two month ramp up period.

Table 22.1 Sanbrado Gold Project Mine Production and Mill Feed Schedule

Year Year Year Year Year Year Year Year Year Year Units Total -1 1 2 3 4 5 6 7 8 9

Mining Ore Mined Mt 16.8 0.4 2.7 2.7 2.3 4.3 4.5 Ore Grade g/t Au 1.7 1.7 2.0 2.4 2.1 1.2 1.2 Processing Ore Milled Mt 16.8 2.1 2.0 1.6 1.9 2.0 2.0 2.0 1.9 1.3 Head Grade g/t Au 1.7 2.3 2.9 2.8 1.5 1.1 1.1 1.1 1.1 0.9 Recovery % 90.7 93.1 92.6 91.8 88.8 89.1 88.7 88.6 86.8 88.9 Gold Recovered Koz 810.2 145.4 166.7 137.9 81.3 64.3 62.4 62.4 55.6 34.2 Run of Mine Ore Stockpile, End of Year Run of Mine Ore Mt 0.3 0.5 0.7 1.3 3.7 6.2 4.1 2.1 0.7 Grade g/t Au 1.9 1.8 2.1 1.2 1.0 1.1 1.2 1.2 1.3 Contained Gold Koz 17.8 27.8 46.9 49.3 117.6 224.1 153.7 83.4 26.6 Low Grade Ore Stockpile, End of Year Low Grade Ore Mt 0.1 0.5 1.0 1.1 1.1 1.1 1.1 1.1 0.7 Grade g/t Au 0.6 0.6 0.6 0.6 0.6 0.6 0.6 0.6 0.6 Contained Gold Koz 1.0 8.3 18.6 19.2 19.2 19.2 19.2 19.2 11.9

Differences between the total over life of mine, and the sum of the years, is due to rounding.

Table 22.1 includes annual estimates of recovered gold, which are derived from the overall (i.e. gravity plus cyanidation) process recovery algorithms described in Section 13 of this report. The average overall process recovery over life of mine is estimated to be 90.7%. Total recovered gold over life of mine is estimated to be 810.2koz, for an average of 92.6kozs per year over the 8.75 year ore processing period. The average production rate during the first three years of operations is estimated to be 150koz per year.

Gold Price Assumption A base case gold price of $1,200/oz was utilised to evaluate the project. This price is similar to the prevailing spot gold price at the effective date of this report, and the three year trailing average metal price (i.e., $1,222/oz for the period from March 2014 to February 2017, inclusive) which is often used to evaluate mining projects.

The gold price has been relatively stable over the past four years as shown in Figure 22.1. Despite this, the impact of a range of higher and lower gold prices on project economic results is addressed as part of project sensitivity analysis.

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Figure 22.1 Historical Gold Price

Data Source: Perth Mint

Cost Estimates All cost estimates are in US dollar currency as of the first quarter of calendar 2017.

22.4.1 Capital and Operating Costs Capital cost estimates are presented in Section 21 of this report. Initial capital is estimated at $124.2 million and sustaining capital is estimated at $34.1 million over life of mine.

Pre-production mining cost estimates are presented in Section 21 of this report. The estimated pre-production mining cost of $6.9 million is calculated as a total cost of $11.9 million minus $5 million which is carried by the mining contractor for a period of twelve months and paid out of the first year of project cashflows.

Pre-production general and administrative costs of $1.8 million comprise an estimate of surface taxes, duties and levies payable during the pre-production period. Taxes, duties and levies are presented in Section 22.5 of this report.

Table 22.2 includes an estimate of the operating costs during the production period.

Operating cost estimates are presented in Section 21 of this report. In addition, WAF has informed Dorado that an allowance of $2/oz of recovered gold has been included for refining. The general and administrative costs include surface taxes, duties and levies and contributions to a local development fund, which are described in Section 22.5 of this report.

The capital and operating cost estimates are based on a diesel fuel price of $1.05/L and a heavy fuel oil price of $0.6491/kg.

Table 22.2

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Sanbrado Gold Project Operating Costs During the Production Period

LOM Cost LOM Cost / Ore Tonne LOM Cost / Ounces Item (US$M) (US$/t) Produced (US$/oz) Mining 266 15.82 329 Processing and Maintenance 181 10.77 224 General & Administrative 83 4.92 102 Refining 2 0.10 2 Total 532 31.62 657

The difference between the calculated LOM Cost / Ore Tonne and the Total is due to rounding.

22.4.2 Closure and Salvage Value Mine closure is presented in Section 20 of this report and closure costs are presented in Section 21. WAF has informed Dorado that a net amount of $4.4 million for mine closure has been included in the sustaining capital costs. This amount is calculated as the difference between the mine closure costs and the salvage value of the equipment and facilities. This cost is incurred in the final production year.

WAF has informed Dorado that under Burkina Faso legislation an environmental fund needs to be established by the fourth year of operation, and contributions made to the fund after that point to ensure that closure funds are available at the end of the mine life.

22.4.3 Working Capital WAF has informed Dorado that approximately two weeks of gold production will be contained within mill circuits and in doré inventory on site or in transit to the refinery. These delays in the receipt of gold revenue contribute to project working capital requirements.

Working capital is also required to maintain an operating supplies inventory. As presented in Section 21 of this report, approximately three months of operating supplies will be purchased in advance of commencement of operations and stored on site.

WAF has informed Dorado that accounts payable, equivalent to one month of capital and operating costs, partially off-sets these working capital requirements.

Taxes and Royalties A number of taxes and royalties are included in the economic evaluation. WAF provided Dorado with advice on Burkina Faso mining legislation and taxation and worked together with Dorado in the modelling of the tax treatment.

22.5.1 Surface Taxes The government of Burkina Faso assesses annual surface taxes on the Mining Permit which are estimated to be $312,500 per annum.

22.5.2 Government Royalty The government of Burkina Faso assesses a flat rate 5% gross revenue royalty on gold projects.

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22.5.3 Local Development Fund Under Burkina Faso legislation a local development fund needs to be established, to which 1% of gross turnover is contributed.

22.5.4 Duties and Levies The government of Burkina Faso assesses a customs duty of 5% and other levies totalling 2.5% on imported goods. During the pre-production period, WAF is exempt from the customs duty but the other levies are applicable. After start up, all imported goods are assessed at a total of 7.5% duties and levies.

22.5.5 Value Added Tax Burkina Faso has a VAT rate of 18%. The VAT is refunded, with the exception of VAT on fuel. In this study, the base case diesel fuel price of $1.05/L includes non-refundable VAT. HFO is exempt from VAT.

A detailed estimate of VAT for each non-fuel item has not been completed for this study. For the purposes of cashflow forecasting, it is assumed that VAT is applicable to 68% of the initial and sustaining capital costs and 77% of mining, processing and general and administrative costs and is refunded eleven months after it is charged.

22.5.6 Corporate Income Tax A corporate tax rate of 27.5% is applicable to gold mining projects in Burkina Faso. Deductions from income for the purpose of estimating income subject to tax include the items described below.

Depreciation Dorado understands that there are a large number of asset classes with varying depreciation rates. WAF provided the following guidance as an approximation of depreciation.

All facilities are depreciated using the units-of-production depreciation method. Depreciation commences once the facilities are placed into service and the mine and mill are operating. Using the units-of-production approach, equipment and facilities are fully depreciated over the mine life.

Carry Forward Income Tax Losses Dorado understands that sunk exploration and other eligible project costs can be carried forward and deducted from income. WAF estimates that its eligible sunk project costs total $15 million.

Mine operating losses can also be carried forward and deducted from income in future years.

Other Deductions Other deductions from income for the purposes of estimating income subject to tax include management fees and interest expenses, which are discussed further in Section 22.6.

22.5.7 Withholding Tax The government of Burkina Faso assesses withholding taxes of 6.25% on dividends and 12.5% on interest income.

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Government Carried Interest Under the mining code of Burkina Faso, the government is entitled to a 10% interest in the project upon formal award of an exploitation permit. Based on guidance provided by WAF, the planned mechanism to develop and operate the mine and incorporate the government’s interest has been modelled as follows:

. An operating company has been formed, with WAF holding 90% of the shares and the government of Burkina Faso holding 10% of the shares. WAF, as managing shareholder, will receive an annual management fee of $2 million per annum. . WAF’s sunk costs and funds provided to develop the mine will be booked as a loan to the operating company, to be repaid with interest out of available cashflow. . The loan to the operating company will be repaid before any distributions are made to the two shareholders of the operating company. . Following repayment of the loan, the operating company free cashflow will be distributed to the two shareholders in the form of dividends, with 10% of the dividends going to the government of Burkina Faso and 90% to WAF. . Dividends and interest received by WAF will be subject to Burkina Faso withholding taxes. Economic Results Project economics have not been updated in this report, and expected to improve significantly once further mineral resource estimate and feasibility work has been completed. Information from the February 20, 2017 Technical Report has been restated for completeness. Further work is in progress and is expected to be completed in mid- 2018. Base case economic results as summarised in Table 22.3 are favourable for the Sanbrado Gold Project. The pre-tax project PVNCF5% is $143 million at the base gold price of $1,200/oz. Project, post-tax, all equity basis PVNCF5% at $1,200 gold is $100 million. IRRs are 27% pre-tax and 21% post-tax.

The government of Burkina Faso is entitled to a 10% interest in the project. At a gold price of $1,200/oz, WAF’s 90% interest is expected to provide a PVNCF5% of $91 million on a post- tax basis and an IRR of 20%, also on a post-tax basis. This assumes WAF provides all funds to develop the mine in the form of loans to the operating company that are repaid, with interest, from gold sales.

At a gold price of $1,200/oz, the project post-tax, all equity basis payback period is expected to be 2.3 years.

Detailed cashflow estimates by year are presented in Table 22.4.

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Table 22.3 Sanbrado Gold Project Base Case Economic Results Summary

Units LOM Total

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Table 22.4 Sanbrado Gold Project Base Case Cashflow Details

Units Units Total Year -2 Year -1 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Mill Feed Ore Milled Mt 16.8 2.1 2.0 1.6 1.9 2.0 2.0 2.0 1.9 1.3 Head Grade g/t Au 1.7 2.3 2.9 2.8 1.5 1.1 1.1 1.1 1.1 0.9 Recovery % 90.7 93.1 92.691.8 88.8 89.1 88.7 88.6 86.8 88.9 Gold Recovered and Payable Koz 810.2 145.4 166.7 137.9 81.3 64.3 62.4 62.4 55.6 34.2 Revenue Gold Price $/oz 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 1,200 Gross Revenue $M 972 175 200 165 98 77 75 75 67 41 Operating Costs Mining Pre-production $M -12 -7 -5 Mining $M -266 -74 -63-40 -40 -41 -2 -2 -2 -2 Processing $M -181 -21 -22-20 -21 -22 -21 -21 -20 -14 General and Administrative Pre-production $M -2 -1 -1 General and Administrative $M -83 -12 -12 -11 -10 -10 -7 -7 -7 -6 Refining $M -2 0 00 0 0 0 0 0 0 Royalties $M -49 -9 -10-8 -5 -4 -4 -4 -3 -2 Capital Costs Initial Capital $M -124 -74 -51 Sustaining Capital $M -34 -9 -3 -4 -3 -4 -3 -3 0 -4 Working Capital Change in Working Capital $M 0 9 -4 -1 -1 -2 2 1 -4 0 0 0 VAT VAT Paid net of Refunds $M 0 -8 1 -7 2 4 0 -1 6 0 1 3 Environmental Rehabilitation Fund Release from (Contribution to) Environmental Rehabilitation Fund $M 0 0 0 0 -1 -1 -1 -1 -1 3 Project Net Cashflow, pre-tax $M 219 -74 -61 36 90 84 19 -4 38 37 33 19 Project PVNCF5%, pre-tax $M 143 Project IRR, pre-tax % pa 27 Project Payback Period, pre-tax Years 2.1 Project Net Cashflow, Post-Tax, All Equity Basis Project Net Cashflow, pre-tax $M 219 -74 -61 36 90 84 19 -4 38 37 33 19 Income Tax, all equity basis $M -56 0 0 -2 -18 -17 -2 0 -4 -7 -6 -1 Project Net Cashflow, post-tax, all equity basis $M 163 -74 -61 35 72 67 17 -4 34 30 28 18 Project PVNCF5%, post-tax, all equity basis $M 100 Project IRR, post-tax, all equity basis % pa 21 Project Payback Period, post-tax, all equity basis Years 2.3 WAF Net Cashflow, Post-Tax Project Net Cashflow, pre-tax, from above $M 219 -74 -61 36 90 84 19 -4 38 37 33 19 Income Tax, Based on Financing Provided by WAF $M -42 0 0 0 -11 -16 -1 0 -3 -6 -5 -1 Withholding Taxes on Interest and Dividends $M -11 0 -1 -2 -10 0 0 0 -2 -2 -2 Dividends to Government (Government 10% carried interest) $M -10 0 0 0 0 0 0 0 0 -3 -3 -4 WAF Net Cashflow, post-tax $M 156 -74 -62 35 79 68 18 -4 8 25 26 38 WAF PVNCF5%, post-tax $M 91 WAF IRR, post-tax % pa 20 WAF Payback Period, post-tax Years 2.3 Differences between the total cashflow over life of mine and the sum of the cashflows in each year is due to rounding. Differences between (i) Project Net Cashflow, pre-tax, Project Net Cashflow, post-tax, WAF Net Cashflow, post-tax, and (ii) the calculation of Project Net Cashflow, pre-tax, Project Net Cashflow, post-tax, WAF Net Cashflow, post-tax based on the cash inflows and cash outflows in Table 22.4, is due to rounding. Differences between (i) PVNCF5% for Project Net Cashflow, pre-tax, Project Net Cashflow, post-tax and WAF Net Cashflow, post-tax, and (ii) the calculation of PVNCF5% based on Project Net Cashflow, pre-tax, Project Net Cashflow, post-tax and WAF Net Cashflow, post-tax based on the cashflows in Table 22.4 are due to a) rounding, and b) the difference between the periodicity of the cashflows in Table 22.4 and the cashflows (monthly for the pre-production period and the first year of operations, quarterly for production years two and three and annual thereafter) which were used in the calculation of PVNCF5%.

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Sensitivity Analysis The PVNCF5% and IRR sensitivity to changes in gold price is shown in Table 22.5. At the base case gold price of $1,200/oz, the project post-tax, all equity basis PVNCF5% is $100 million and the project, post-tax, all equity basis IRR is 21%. If the gold price increases to $1,300/oz, the project post-tax, all equity basis PVNCF5% rises to $143 million and the project, post-tax, all equity basis IRR rises to 27%. Conversely, a reduction in the gold price to $1,100/oz results in a drop in the project post-tax, all equity basis PVNCF5% to $56 million and a reduction in the project, post-tax, all equity basis IRR to 14%. At a gold price of $1,100/oz the project post-tax, all equity basis payback period increases to 2.6 years.

Table 22.5 Sanbrado Gold Project Project Economics Sensitivity to Gold Price

Gold Price $/oz 1,100 1,200 1,300

Project Net Cashflow, Pre-Tax Project Net Cashflow, Pre-Tax $M 143 219 295 PVNCF5% $M 83 143 202 IRR % pa 18 27 35 Payback Years 2.5 2.1 1.7 Project Net Cashflow, Post-Tax, All Equity Basis Project Net Cashflow, pre-tax $M 108 163 218 PVNCF5% $M 56 100 143 IRR % pa 14 21 27 Payback Years 2.6 2.3 2.1 WAF Net Cashflow, Post-Tax Project Net Cashflow, pre-tax $M 113 156 202 PVNCF5% $M 58 91 125 IRR % pa 15 20 25 Payback Years 2.6 2.3 2.1

The sensitivity of the project, post-tax, all equity basis PVNCF5% and IRR to ±20% changes in the key operating parameters of revenue, capital costs and operating costs are shown in Figure 22.2 and Figure 22.3.

The sensitivity results reflect a change in one parameter at a time, assuming the other parameters are unchanged.

A review of Figure 22.2 and Figure 22.3 indicates that, like most mining projects, the post- tax PVNCF5% and IRR are most sensitive to changes in revenue parameters. These parameters include gold price, process plant head grade, and, process plant recovery. Note, however, that the base case process plant recovery is 90.7%, and cannot exceed 100%.

The project post-tax, all equity basis PVNCF5% and IRR are both more sensitive to changes in operating costs than to capital costs. This is attributed to the fact that total base case operating costs (mining, processing, refining and general and administrative costs) over life of mine are about 3.6 times the total capital costs (initial capital and sustaining capital).

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Figure 22.2 Project, Post-Tax, All Equity Basis PVNCF5% ($M) Sensitivity to Key Input Parameters

Figure 22.3 Project, Post-Tax, All Equity Basis IRR (% pa) Sensitivity to Key Input Parameters

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 357 West African Resources Limited

ADJACENT PROPERTIES Introduction The region is principally prospective for orogenic gold deposits, which typically exhibit a strong relationship with regional arrays of major shear zones, similar to that found elsewhere in late Proterozoic Birimian terrains of West Africa and other greenstone belts around the world.

Properties adjacent to the Sanbrado Gold Project include WAF’s Mogtedo project, Orezone’s Bombore Project and B2 Gold’s Kiaka and Toega projects.

West African Resources: Exploration Projects The Mogtedo prospect is located approximately 20km north northwest of Sanbrado, in the southern part of the Zam Permit. Gold mineralisation at Mogtedo is associated with a network of shear hosted veins and veinlets within a northeast trending shear corridor. Drilling has been carried out over an area of approximately 1,800m of strike, and 150m width, within which mineralisation has been found to be discontinuous.

The Meguet and Fatmatenga prospects are located in the northeastern corner of the Zam permit near the town of Meguet, some 24km northwest of Zorgho, and are centrally situated within the Boulsa Project. The Meguet and Fatmatenga prospects have been tested with various drilling programs and warrant further drilling.

Orezone Gold Corporation: Bomboré Project Orezone Gold Corporation’s Bomboré gold deposit is located immediately adjacent to the southwestern portion of the Sanbrado Gold Project. The Bombore deposit is a large tonnage, low grade gold deposit for which a hybrid heap leach-CIL process method has been proposed.

B2 Gold The B2 Gold has two project sites near the Sanbrado Gold Project: Kiaka is an advanced exploration project located 46km south of Sanbrado, and the more recently discovered Toega project is 14km southwest of Sanbrado.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 358 West African Resources Limited

OTHER RELEVANT DATA AND INFORMATION To the authors’ knowledge, there is no other relevant data or information related to the Sanbrado property.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 359 West African Resources Limited

INTERPRETATION AND CONCLUSIONS The key conclusions are:

. The Sanbrado Gold Project hosts a substantial gold deposit that has strong potential for development. . Exploration drilling, sampling and assaying by West African Resources and its predecessor Channel Resources has been carried out to applicable industry standards and the data derived is appropriate for resource estimation. . It is the Qualified Person’s opinion that the Sanbrado Gold Project resource documented in this report meets NI 43-101 standards and JORC Code guidelines. . The Sanbrado Gold Project Resource Estimate data spacing, quality of data, and current confidence in the geological understanding of the deposit is sufficient to imply or infer continuity of mineralisation and grade. Additional infill drilling is needed to improve confidence in the Inferred Mineral Resources to a level needed for detailed economic assessment (i.e. to define Indicated Mineral Resources). . The Qualified Person understands that presently no major environmental, permitting, legal, taxation, socio-economic, marketing or political factors have been identified which would materially affect the resource estimate. . Metallurgical testwork returned to date indicates that gold is amenable to recovery by conventional CIL processing techniques. Life of mine recoveries average greater than 90%. . Mining studies have determined an open pit project is robust, but would benefit from optimisation and inclusion of underground mining methods. It is proposed that mining will take place using conventional open-pit and underground mining operations. . The Sanbrado Gold Project resource is largely open at depth, and further drilling is warranted to test strike and depth extensions. . The results of the analysis of the Sanbrado Gold Project, as it is currently envisaged, are very positive. The Sanbrado Gold Project has been evaluated on a discounted cashflow basis. The present value of the net cashflow with a 5% discount rate (PVNCF5%) is $143 million on a pre-tax, project basis, using a base gold price of $1,200/oz. Project post-tax PVNCF5% at a $1,200 gold price is $100 million on an all equity basis. Project internal rates of return (IRR) are respectively 27% pre-tax and 21% post-tax. Additional work is in progress and will be updated in the Definitive Feasibility Study technical report due for completion in mid-2018.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 360 West African Resources Limited

RECOMMENDATIONS The following activities are recommended:

. Infill and extensional drilling at M5 and M1 South to upgrade the Mineral Resource Inferred category falling within the proposed open pits to Indicated, enabling the estimation of updated Mineral Resources and Mineral Reserves. . Investigate potential for underground development of M1 South and M5 deposits. . Continuation of environmental monitoring and advancement of permitting of statutory requirements to enable a smooth transition to production. . Implementation of Resettlement Action Plan. Table 26.1 provides recommendations and associated costs.

Table 22.5 Recommendations and Associated Costs

Item Cost Infill and extensional Drilling Program $US8M Underground Study $0.5M Environmental and Permitting $0.5M Resettlement Action Plan $2M

Note: Costs are indicative and will be refined as the project progresses.

NI 43-101 Technical Report: - Sanbrado Gold Project, Burkina Faso P a g e | 361 West African Resources Limited

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