Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the Nico Gold-Coba

Home , Cobalt extraction, Stoping

TECHNICAL REPORT AND UPDATED MINERAL RESERVE ESTIMATE AND FRONT-END ENGINEERING & DESIGN (FEED) STUDY ON THE NICO GOLD-COBALT-BISMUTH-COPPER DEPOSIT MAZENOD LAKE AREA, NORTHWEST TERRITORIES, CANADA

FOR

FORTUNE MINERALS LIMITED

P&E Mining Consultants Inc. NI-43-101 & 43-101F1 TECHNICAL REPORT No. 247

Eugene Puritch, P.Eng.: P&E Mining Consultants Inc., Wayne Ewert, P.Geo.: P&E Mining Consultants Inc., Tracy Armstrong, P.Geo.: P&E Mining Consultants Inc., Fred Brown, CPG, PrSciNat.: P&E Mining Consultants Inc., David Orava, P.Eng: P&E Mining Consultants Inc., James L. Pearson, P.Eng.: P&E Mining Consultants Inc., Tim Hayes, P.Eng.: Jacobs Minerals Canada Inc. Alex Duggan P.Eng.: Jacobs Minerals Canada Inc. Graham Holmes, P.Eng.: Jacobs Minerals Canada Inc. Diogenes Uceda, P.Eng.: Jacobs Minerals Canada Inc. Wade Sumners, P.Ag., P.Biol.: MDH Engineered Solutions. Dan Mackie, P.Eng.: Dan Mackie and Associates. Marc Rougier, P.Eng.: Golder Associates Ltd. Ken Bocking, P.Eng.: Golder Associates Ltd. Alex Mezei, P.Eng.: SGS Mineral Services. Bill Horne P.Eng.: EBA Engineering Consultants Ltd.

Effective Date: July 2, 2012 Signing Date: August 16, 2012

TABLE OF CONTENTS

1.0 SUMMARY ...... 1 1.1 INTRODUCTION ...... 1 1.2 STUDY PARTICIPANTS AND RESPONSIBILITIES ...... 1 1.3 PERSONAL SITE INSPECTIONS ...... 2 1.4 PROPERTY DESCRIPTION AND LOCATION ...... 2 1.5 GEOLOGY AND MINERALIZATION ...... 2 1.6 EXPLORATION STATUS ...... 3 1.7 MINERAL RESOURCES ...... 4 1.8 MINERAL RESERVES ...... 5 1.8.1 Underground Reserve ...... 5 1.8.2 Open Pit Reserve...... 6 1.8.3 Total Reserve ...... 6 1.9 MINING ...... 6 1.9.1 Mine Production Schedule ...... 6 1.9.2 Underground Mine Development ...... 7 1.9.3 Underground Stope Development...... 7 1.9.4 Underground Stoping ...... 8 1.9.5 Capital Costs – Underground Mine ...... 9 1.9.5.1 Pre-production Underground Capital Costs ...... 9 1.9.5.2 Underground Sustaining Capital Costs ...... 10 1.9.6 Underground Mine Operating Cost ...... 10 1.10 ENVIRONMENTAL ...... 10 1.10.1 SMPP Water Management ...... 11 1.10.2 Permitting ...... 12 1.11 TAILINGS ...... 14 1.11.1 NICO Mine Tailings and Waste Co-disposal Facility ...... 14 1.11.2 SMPP Waste and Tailings Disposal ...... 15 1.12 WATER SUPPLY ...... 15 1.12.1 NICO Site Water Management ...... 15 1.13 MARKETING...... 16 1.13.1 Cobalt ...... 16 1.13.2 Bismuth ...... 16 1.14 PROCESS PLANT ...... 16 1.15 PROJECT EXECUTION ...... 17 1.16 CAPITAL AND OPERATING EXPENSES...... 17 1.17 PROJECT ECONOMIC ANALYSIS ...... 18 1.17.1 Economic Analysis ...... 18 1.18 CONCLUSIONS AND RECOMMENDATIONS ...... 19 1.18.1 Interpretation and Conclusions ...... 19 1.18.2 Recommendations ...... 21 1.18.2.1 Bismuth Processing Plant (BiPP)...... 23 1.18.2.2 Cobalt Sulphate Option Recommendation ...... 23 1.18.2.3 Mining ...... 23 1.18.2.4 Marketing ...... 23 2.0 INTRODUCTION ...... 24 2.1 SOURCES OF INFORMATION AND STUDY PARTICIPANTS ...... 24 2.2 QUALIFIED PERSONS AND RESPONSIBILITIES ...... 24 2.3 INDEPENDENT SITE INSPECTIONS ...... 25

2.4 SOURCES OF INFORMATION ...... 26 2.5 UNITS AND CURRENCY ...... 26 2.6 GLOSSARY AND ABBREVIATION OF TERMS ...... 26 3.0 RELIANCE ON OTHER EXPERTS ...... 29 4.0 PROPERTY DESCRIPTION AND LOCATION ...... 32 4.1 LOCATION ...... 32 4.2 PROPERTY DESCRIPTION AND TENURE ...... 32 4.2.1 Northwest Territories Claims and Mining Leases ...... 34 4.2.2 Surface Rights and Land Agreements ...... 35 4.2.3 Saskatchewan Metals Processing Plant ...... 36 5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 37 5.1 ACCESSIBILITY ...... 37 5.2 CLIMATE ...... 38 5.3 LOCAL RESOURCES AND INFRASTRUCTURE ...... 38 5.4 PHYSIOGRAPHY ...... 38 6.0 HISTORY ...... 40 6.1 PREVIOUS MINERAL RESOURCE AND RESERVE ESTIMATES ...... 42 6.2 MICON MINERAL RESOURCE AND RESERVE ESTIMATE, 2007 ...... 43 7.0 GEOLOGICAL SETTING AND MINERALIZATION ...... 45 7.1 REGIONAL GEOLOGY ...... 45 7.2 PROPERTY GEOLOGY ...... 47 7.3 MINERALIZATION ...... 49 8.0 DEPOSIT TYPES ...... 52 8.1 IRON OXIDE COPPER-GOLD (IOCG) DEPOSITS ...... 52 9.0 EXPLORATION...... 54 10.0 DRILLING ...... 55 11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY ...... 58 11.1 CORE LOGGING ...... 58 11.2 CORE SAMPLING ...... 58 11.3 SAMPLE PREPARATION ...... 59 11.4 ANALYSIS ...... 59 11.4.1 Method Au-AA23/AU-AA24 ...... 59 11.4.2 Method Au-GRA21/Au-GRA21 ...... 60 11.4.3 Method ME-AA46 (includes As-AA46 and Bi-AA46) ...... 60 11.4.4 Method ME-AA62 (includes Cu-AA62a and Co-AA62) ...... 61 11.5 OTHER METHODS ...... 62 11.5.1 Specific Gravity Measurements ...... 62 12.0 DATA VERIFICATION ...... 63 12.1 SITE VISIT AND INDEPENDENT SAMPLING ...... 63 12.2 QUALITY ASSURANCE/QUALITY CONTROL ...... 64 12.2.1 Fortune Au-Co Standards ...... 64 12.2.2 Fortune Minerals Blank ...... 65 13.0 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 66 13.1 NICO SUMMARY DESCRIPTION ...... 66 13.2 ORE CHARACTERISTICS ...... 66 13.2.1 Ore Grade Metal Content ...... 66 13.2.2 Mineralogy ...... 67 13.2.3 Physical Properties ...... 68 13.2.4 Work Indices ...... 68

13.2.5 Materials Handling...... 69 13.2.6 Mineralization ...... 70 13.3 SULPHIDES ...... 70 13.4 LABORATORY BENCH-SCALE AND PILOT PLANT TESTING ...... 71 13.4.1 Bulk and Bismuth Flotation ...... 71 13.4.2 2006 Hydromet Mini Pilot Plant ...... 71 13.4.3 2007 Flotation Pilot Plant ...... 72 13.5 2010 FLOTATION PILOT PLANT ...... 78 13.5.1 Sample Description, Source Blending and Feed Grades ...... 78 13.6 FLOTATION DESIGN CRITERIA ...... 82 13.7 PROCESS HYDROMETALLURGICAL PLANT ...... 82 13.7.1 Concentrate Characteristics ...... 83 13.7.2 Laboratory Bench-Scale and Pilot Plant Testing ...... 84 13.7.2.1 Laboratory Testing ...... 84 13.7.3 Pressure Oxidative Autoclave Corrosion ...... 92 14.0 MINERAL RESOURCE ESTIMATES ...... 94 14.1 INTRODUCTION ...... 94 14.2 PREVIOUS MINERAL RESOURCE ESTIMATES ...... 95 14.3 SAMPLE DATABASE ...... 95 14.4 DOMAIN MODELING ...... 96 14.5 COMPOSITING ...... 96 14.6 TREATMENT OF EXTREME VALUES...... 97 14.7 CONTINUITY ANALYSIS ...... 98 14.8 BLOCK MODELS...... 98 14.9 ESTIMATION & CLASSIFICATION ...... 99 14.10 MINERAL RESOURCE ESTIMATE ...... 99 14.11 MINERAL RESOURCE VALIDATION...... 100 15.0 MINERAL RESERVE ESTIMATES...... 102 15.1 MINERAL RESERVES CRITERIA ...... 102 15.2 NET SMELTER RETURN (NSR) ...... 102 15.3 UNDERGROUND RESERVE ...... 103 15.4 OPEN PIT RESERVE ...... 103 15.5 TOTAL RESERVE ...... 104 15.6 RESPONSIBILITY FOR ESTIMATION ...... 104 16.0 MINING METHODS ...... 105 16.1 OPEN PIT ...... 105 16.1.1 Conventional Open Pit Operation ...... 105 16.1.2 Open Pit Production Schedule ...... 105 16.1.3 Open Pit Operation ...... 107 16.1.4 Open Pit Labour ...... 107 16.2 UNDERGROUND MINING ...... 110 16.2.1 Underground Mine Design and Stope Layout ...... 116 16.2.2 Underground Production Schedule ...... 121 16.2.2.1 Mine Development...... 121 16.2.2.2 Stope Development ...... 121 16.2.2.3 Stoping ...... 121 16.2.3 Ventilation...... 123 16.2.4 Hydrology ...... 124 16.2.5 Manpower ...... 124 16.2.5.1 Underground Operations and Maintenance Labour ...... 124

16.2.6 Equipment ...... 125 16.2.7 Electrical Power ...... 127 16.2.8 Stockpile ...... 128 16.3 GEOTECHNICAL ...... 128 16.3.1 Mine Plan Summary ...... 128 16.3.2 Geology ...... 130 16.3.3 Discontinuous Permafrost Conditions ...... 130 16.3.4 In Situ Stresses ...... 131 16.3.5 Geotechnical Drillhole Locations ...... 131 16.3.6 Engineering Geology ...... 133 16.3.7 Open Pit Design ...... 135 16.3.7.1 Volcanics...... 136 16.3.8 Slope Design Definitions ...... 137 16.3.9 Kinematic Assessment ...... 138 16.3.10 Slope Design Recommendations ...... 139 16.3.11 Pit and Underground Seepage Estimates ...... 140 16.3.12 Underground Mine ...... 141 16.3.12.1 Semi-Empirical Open Stope Stability Analyses ...... 141 16.3.12.2 2005 Stope Stability Assessment ...... 143 16.3.12.3 2010 Stope Stability Assessment ...... 144 16.3.13 Stope Backfilling ...... 144 16.3.14 2011 Planned Stopes ...... 145 16.4 MISCELLANEOUS UNDERGROUND DRAWINGS ...... 149 16.4.1 Underground Development Drawings ...... 149 17.0 RECOVERY METHODS ...... 158 17.1 NICO CONCENTRATOR (DWG-0000-F-001) ...... 158 17.2 NICO PLANT UTILITIES ...... 161 17.3 SMPP PROCESS DESCRIPTION (DWG-0000-F-001/002) ...... 162 17.3.1 Bulk Concentrate Regrind and Bismuth Flotation ...... 163 17.3.2 Pressure Oxidation ...... 167 17.3.3 Solution Neutralization and Iron-Arsenic Removal ...... 168 17.3.4 Copper Cementation and Precipitation ...... 169 17.3.5 Cobalt Precipitation ...... 169 17.3.6 Cobalt Dissolution and Ion Exchange (IX) Purification ...... 170 17.3.7 Cobalt Electrowinning ...... 172 17.3.8 Cobalt Sulphate Heptahydrate Option ...... 175 17.3.9 Cobalt Residue Cyanidation and Merrill Crowe ...... 175 17.3.10 Tailings Handling and Cyanide Destruction ...... 176 17.4 SMPP PLANT UTILITIES...... 177 17.4.1 Water Supply ...... 177 17.4.2 Process Ancillaries and Services ...... 178 17.5 EXECUTION PLAN ...... 182 17.5.1 NICO Concentrator Overview ...... 182 17.5.2 NICO Construction Schedule Presentation ...... 183 17.5.3 Saskatchewan Metallurgical Processing Plant (SMPP) Overview ...... 185 17.5.4 SMPP Construction Schedule Presentation ...... 185 18.0 PROJECT INFRASTRUCTURE ...... 187 18.1 NICO MINE SITE AND CONCENTRATOR ...... 187 18.1.1 Access Road ...... 187 18.1.2 Access Road Geotechnical ...... 188

18.1.3 Plant Site Geotechnical ...... 189 18.1.4 Power Plant ...... 191 18.1.5 Fuel Depot ...... 191 18.1.6 Process Facility ...... 191 18.1.7 Accommodation Complex ...... 191 18.1.8 Corridors / Utilidors ...... 191 18.1.9 Water Supply ...... 194 18.1.10 Co-Disposal Facility ...... 194 18.1.10.1 Tailings and Waste Rock Co-Disposal Facility ...... 194 18.1.11 Water Treatment ...... 197 18.1.12 Plant Mobile Equipment...... 197 18.2 SASKATCHEWAN METALLURGICAL PROCESSING PLANT (SMPP) ....198 18.2.1 Plant Site and Facility ...... 198 18.2.2 Process Facility ...... 198 18.2.3 Administrative, Laboratory and Warehouse Building ...... 201 18.2.4 Railway Access ...... 201 18.2.5 Oxygen Plant ...... 201 18.2.6 Utilities ...... 201 18.2.7 Site Drainage ...... 202 18.2.8 Plant Mobile Equipment ...... 202 19.0 MARKET STUDIES AND CONTRACTS ...... 203 19.1 COBALT...... 203 19.1.1 Uses ...... 203 19.1.2 Supply ...... 203 19.1.3 Demand ...... 204 19.1.4 Pricing ...... 204 19.2 BISMUTH...... 206 19.2.1 Uses ...... 206 19.2.2 Supply ...... 206 19.2.3 Demand ...... 206 19.2.4 Pricing ...... 207 19.3 PRECIOUS AND BASE METALS ...... 208 19.4 CASH FLOW COMMODITY PRICES ...... 209 20.0 ENVIRONMENTAL STUDIES PERMITTING AND SOCIAL OR COMMUNITY IMPACT ...... 210 20.1 NICO MINE SITE ...... 210 20.1.1 Environmental Study Results ...... 210 20.1.1.1 Caribou ...... 210 20.1.1.2 Water Quality ...... 211 20.1.2 Waste and Tailings Disposal, Site Monitoring and Water Management ...... 214 20.1.2.1 Co-disposal Facility ...... 214 20.1.2.2 Water Management ...... 215 20.1.3 Project Permitting ...... 217 20.1.3.1 NICO Project and the Impact Assessment Process ...... 217 20.1.3.2 Reclamation Bond Requirements ...... 218 20.1.4 Social and Community Requirements and Considerations ...... 218 20.1.4.1 Corporate Commitments and Agreements ...... 219 20.1.5 Mine Closure Requirements and Costs ...... 220 20.2 SMPP ...... 221

20.2.1 Waste and Tailings Disposal ...... 221 20.2.1.1 Reclamation ...... 222 20.2.2 Waste Water Injection Well ...... 223 20.2.2.1 Site Monitoring ...... 224 20.2.2.2 Post-Decommissioning Monitoring ...... 224 20.2.3 Storage Ponds...... 225 20.2.4 Anticipated Decommissioning Costs ...... 226 20.2.5 Potential Social or Community Related Requirements ...... 226 20.2.5.1 Geotechnical Foundation Investigation ...... 227 20.2.5.2 Ground Water Supply ...... 227 20.2.5.3 Waste Characterization ...... 228 21.0 CAPITAL AND OPERATING COSTS ...... 229 21.1 CAPITAL COST SUMMARY ...... 229 21.1.1 Process Plants Capital Cost Summary ...... 229 21.1.2 Capital Cost Summary – Cobalt Sulphate Option ...... 230 21.1.3 Open Pit Equipment Capital Costs ...... 231 21.1.4 Underground Capital Costs ...... 233 21.1.4.1 Pre-production Underground Capital Costs ...... 233 21.1.4.2 Underground Sustaining Capital Costs ...... 233 21.2 OPERATING COST SUMMARY ...... 233 21.2.1 Concentrator and Southern Metallurgical Processing Plant Operating Cost ...... 233 21.2.1.1 NICO Operating Cost ...... 234 21.2.2 Open Pit Operating Costs...... 235 21.2.3 Underground Operating Cost ...... 237 21.2.3.1 Fortune‟s Labour Costs ...... 237 22.0 ECONOMIC ANALYSIS ...... 240 22.1 COBALT METAL OPTION ...... 240 22.2 COBALT SULPHATE OPTION ...... 240 22.3 ECONOMIC ANALYSIS ...... 243 22.4 PROJECTED OPERATING CASH COSTS...... 251 22.5 SENSITIVITY ANALYSIS ...... 251 23.0 ADJACENT PROPERTIES ...... 253 24.0 OTHER RELEVANT DATA AND INFORMATION ...... 254 24.1 OPTION TO PRODUCE COBALT SULPHATE ...... 254 24.2 SCOPE OF WORK ...... 254 24.3 PROCESS DESCRIPTION (OPTION B) ...... 255 24.4 COBALT SULPHATE HEPTAHYDRATE GRADE ...... 259 24.5 CAPITAL COST SUMMARY ...... 259 24.6 OPERATING COST SUMMARY ...... 260 25.0 INTERPRETATION AND CONCLUSIONS ...... 261 25.1 MINING ...... 261 25.2 GEOLOGY ...... 261 25.3 MINING METHODS ...... 262 25.4 NICO SITE - TAILINGS AND WASTE ROCK CO-DISPOSAL FACILITY ...... 263 25.5 NICO SITE – ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT ...... 263 25.5.1 Environmental Study Results ...... 263 25.5.2 Water Management ...... 263

25.5.3 Closure ...... 264 25.5.4 Project Permitting ...... 264 25.5.5 Social and Community Requirements and Considerations ...... 264 25.5.6 Mine Closure Requirements and Costs ...... 264 25.5.7 Metallurgy and Processing ...... 265 25.5.8 Project Execution ...... 265 25.5.9 Project Economic Analysis ...... 265 25.6 MARKETING...... 266 25.6.1 Cobalt ...... 266 25.6.2 Bismuth ...... 266 26.0 RECOMMENDATIONS ...... 267 26.1 GEOTECHNICAL ...... 267 26.2 METALLURGY AND PROCESSING ...... 267 26.2.1 NICO Concentrator – Detailed Engineering Execution...... 267 26.2.2 NICO Process Opportunities – Future Works ...... 267 26.2.3 SMPP - Detailed Engineering Execution ...... 269 26.2.4 SMPP Process Opportunities - Future Works ...... 269 26.2.5 Bismuth Processing Plant ...... 270 26.2.6 Cobalt Sulphate Option Recommendation ...... 271 26.3 ENVIRONMENTAL-- PERMITTING, TAILINGS AND CLOSURE ...... 271 26.4 MARKETING...... 271 27.0 REFERENCES ...... 272 28.0 CERTIFICATES ...... 279

LIST OF TABLES

Table 1.1 NICO Estimated Open Pit and Underground Mineral Resources ...... 5 Table 1.2 Underground Mineral Reserve Estimates ...... 5 Table 1.3 Open Pit Reserves ...... 6 Table 1.4 Total Reserves ...... 6 Table 1.5 Mine Development Schedule Summary ...... 7 Table 1.6 Stope Development Schedule Summary ...... 8 Table 1.7 Production Blast Hole Blasting, Mucking and Haulage Schedule Summary ...... 8 Table 1.8 Summary of U/G Stope Mining Sequence ...... 9 Table 1.9 Summary of 2016 Underground Pre-production Capital Costs ...... 10 Table 1.10 Summary of Sustaining Capital Costs ...... 10 Table 1.11 Average Metal Production ...... 17 Table 1.12 Projected Operating Cash Costs for the Cobalt Metal Option ...... 17 Table 1.13 Summary of OverallCapital Costs ...... 18 Table 1.14 Summary of Operating Costs (LOM) Financial Model ...... 18 Table 1.15 NICO Economics ...... 19 Table 4.1 NICO Property Mining Leases, Effective January 2012 ...... 34 Table 6.1 NICO Mineral Resources Summary, 2007 ...... 43 Table 6.2 Mineral Reserve Estimate, Micon 2007 ...... 44 Table 10.1 Highlights of Drill Intercepts from 2010 Drill Program ...... 56 Table 11.1 ALS Chemex Method Au-AA23/Au-AA24 Summary ...... 60 Table 11.2 ALS Chemex Method Au-GRA21/Au-GRA22 Summary...... 60 Table 11.3 ALS Chemex Method ME-AA46 Summary ...... 61 Table 11.4 ALS Chemex Method ME-AA62 Summary ...... 61 Table 13.1 Stage and Overall Plant Recoveries ...... 66 Table 13.2 NICO Run-of-Mine Metal Head Grades Comparison ...... 67 Table 13.3 Bulk Concentrate Metal Grades ...... 67 Table 13.4 Compressive Strength Results ...... 68 Table 13.5 Plant Design ROM Ore Composition...... 70 Table 13.6 Plant Design ROM Ore Mineralogy Composition ...... 70 Table 13.7 Summary of Composite Head Grades ...... 73 Table 13.8 Rod Mill and Ball Mill Bond Work Indices ...... 73 Table 13.9 Summary of Average Head Grades and Circuit Performance of Composite P-1 ...... 74 Table 13.10 Summary of Average Head Grades and Circuit Performance of Composite P-2 ...... 74 Table 13.11 Summary and Comparison of Crushing, Grinding and Flotation Design Parameters ...... 74 Table 13.12 Solid / Liquid Design Criteria from Pocock Testwork ...... 75 Table 13.13 Flow Sheet Option 4+6 Input Parameters ...... 77 Table 13.14 Flow Sheet Option 4 Stage Recoveries ...... 77 Table 13.15 Flow Sheet Option 6 Stage Recoveries ...... 78 Table 13.16 Summary of Composite Head Grades ...... 79 Table 13.17 Rod Mill and Ball Mill Bond Work Indices ...... 79 Table 13.18 Flow Sheets Options 4 & 6 Pilot Plant Results for Composite P3 ...... 79 Table 13.19 Flow Sheets Options 4 & 6 Pilot Plant Results for Composite P5 ...... 80 Table 13.20 Gravity Recoverable Gold Results for Composite P3 ...... 81 Table 13.21 Gravity Recoverable Gold Results for Composite P4 ...... 81 Table 13.22 NICO Plant Stage Recoveries and Resulting Grades ...... 82

Table 13.23 Stage and Overall Plant Recoveries ...... 83 Table 14.1 Historical mineral resource estimate dated November 2004 ...... 95 Table 14.2 NICO Drilling Database Records ...... 95 Table 14.3 Modeled Domains ...... 96 Table 14.4 NICO Uncapped Composite Statistics by Domain ...... 97 Table 14.5 NICO Capping And Threshold Values ...... 98 Table 14.6 NICO Experimental Semi-Variograms ...... 98 Table 14.7 NICO Block Model Setup ...... 98 Table 14.8 Economic Parameters ...... 99 Table 14.9 NSR Metal Unit Values ...... 100 Table 14.10 NICO Estimated Mineral Resources ...... 100 Table 14.11 Validation Statistics for Capped Composites and Block Estimates ...... 101 Table 15.1 Resource to Reserve Conversion Parameters ...... 102 Table 15.2 Underground Reserves ...... 103 Table 15.3 Open Pit Reserves ...... 104 Table 15.4 Total Reserves ...... 104 Table 16.1 Open Pit Production Schedule...... 106 Table 16.2 Production Drilling Patterns ...... 107 Table 16.3 Open Pit Equipment Schedule...... 108 Table 16.4 Open Pit Personnel ...... 109 Table 16.5 Summary of U/G Stope Mining Tonnages ...... 111 Table 16.6 Typical 5 m x 5 m Development Round Parameters ...... 116 Table 16.7 Typical 4.5 m x 4 m Development Round Parameters ...... 117 Table 16.8 Mine Development Schedule Summary ...... 121 Table 16.9 Stope Development Schedule Summary ...... 121 Table 16.10 Production Blast Hole Blasting, Mucking and Haulage Schedule Summary ..... 122 Table 16.11 Summary of U/G Stope Mining Sequence ...... 122 Table 16.12 Ventilation Requirements for Underground Diesel Equipment ...... 124 Table 16.13 Projected Number of Underground Operations and Maintenance Personnel ..... 125 Table 16.14 Projected Number of Underground Fortune Minerals Personnel...... 125 Table 16.15 Summary of Contractor Equipment ...... 126 Table 16.16 Summary of Fortune Mineral‟s Equipment for the Underground ...... 127 Table 16.17 Estimated Underground Electrical Power Requirements ...... 127 Table 16.18 Geotechnical Units ...... 134 Table 16.19 Peak Orientations of Discontinuity Populations in the Meta-sedimentary Rocks ...... 136 Table 16.20 Peak Orientations of Discontinuity Populations in the Volcanic Cap Rock (Potassium Feldspar Altered Rhyolite) ...... 137 Table 16.21 NICO Open Pit Slope Design Recommendations ...... 140 Table 16.22 Preliminary Values of Maximum Hydraulic Radius (HR) for Unsupported Walls ...... 143 Table 16.23 Revised Values of Maximum Hydraulic Radius (HR) for Unsupported Walls .. 144 Table 16.24 Example of Slurry and Lean Mix Concrete Design ...... 145 Table 16.25 Hydraulic Radius Values for 2011 Planned Stopes – Back Dimensions ...... 147 Table 16.26 Hydraulic Radius Values for 2011 Planned Stopes – Sidewall Dimensions ...... 148 Table 18.1 Plant Mobile Equipment for NICO ...... 198 Table 18.2 Plant Mobile Equipment for SMPP ...... 202 Table 20.1 Process/Brine Solution Characteristics ...... 223 Table 21.1 Capital Cost Summary of the Process Plants ...... 229 Table 21.2 Capital Cost Comparison ...... 231

Table 21.3 Open Pit Equipment Capital and Sustaining Capital Expenditures ...... 232 Table 21.4 Summary of 2016 Underground Pre-production Capital Costs ...... 233 Table 21.5 Summary of Sustaining Capital Costs ...... 233 Table 21.6 Annual Concentrator and Southern Metallurgical Processing Plant Operating Cost Summary ...... 234 Table 21.7 Annual NICO Process Plant Operating Cost Summary ...... 234 Table 21.8 Annual SMPP Process Plant Operating Cost Summary ...... 234 Table 21.9 Operating Cost Summary ...... 235 Table 21.10 Open Pit Operating Costs ...... 236 Table 21.11 Labour Rates For Underground Fortune Personnel ...... 238 Table 21.12 Total Estimated Labour Cost ...... 239 Table 22.1 Results of the economic analysis for the cobalt metal option ...... 240 Table 22.2 Results of the economic analysis for the cobalt sulphate option ...... 241 Table 22.3 Economic Parameters and Key Assumptions ...... 242 Table 22.4 Cobalt Metal, Base Case Metal Prices-Exchange Rate Cashflow Model ...... 244 Table 22.5 Projected Operating Cash Costs for the Cobalt Metal Option ...... 251 Table 22.6 Projected Operating Cash Costs for the Cobalt Sulphate Option ...... 251 Table 22.7 Sensitivity Analysis (Cobalt Metal Option, Base Case Metal Prices-Exchange Rate) ...... 252 Table 24.1 Target Product Specification ...... 259 Table 24.2 Capital Cost Comparison ...... 260 Table 24.3 Operating Cost Summary ...... 260 Table 25.1 Underground Reserves ...... 261 Table 25.2 Open Pit Reserves ...... 262 Table 25.3 Total Reserves ...... 262

LIST OF FIGURES

Figure 4.1 Location Map of the NICO Property ...... 32 Figure 4.2 NICO Mining Leases ...... 33 Figure 5.1 Access Map for the NICO Property ...... 37 Figure 7.1 Regional Geology Map ...... 46 Figure 7.2 Property Geology Map ...... 48 Figure 7.3 Typical Cross Section through the Bowl Zone (at 20+50W) ...... 50 Figure 8.1 IOCG Deposit Subtypes ...... 53 Figure 12.1 NICO Deposit Site Visit Sample Results for Bismuth ...... 63 Figure 12.2 NICO Deposit Site Visit Sample Results for Cobalt...... 63 Figure 12.3 NICO Deposit Site Visit Sample Results for Copper ...... 64 Figure 16.1 NICO - As Built Underground Workings ...... 112 Figure 16.2 NICO - As Built Underground Workings ...... 113 Figure 16.3 NICO - As Built Underground Workings ...... 114 Figure 16.4 Longitudinal Projection of the Proposed Underground Workings ...... 115 Figure 16.5 Typical 5.0m High by 5.0m Wide Ramp Development Heading ...... 118 Figure 16.6 Typical 5.0m High by 5.0m Wide Drift Development Heading ...... 119 Figure 16.7 Typical 4.5m High by 4.0m Wide Cross Cut Heading ...... 120 Figure 16.8 Schematic Representation of the Phase Pit Shells and the 2011 Planned Underground Stopes ...... 129 Figure 16.9 Open Pit and Mine Waste Geotechnical Investigations Borehole Location Plan ...... 132 Figure 16.10 NICO Deposit Structural Fabric based on Surface Mapping and Oriented Core ...... 136 Figure 16.11 Peak Orientations of Discontinuity Populations in the Volcanic Cap Rock (Potassium Feldspar Altered Rhyolite) ...... 137 Figure 16.12 Schematic Representation of Bench Face Angle (“BFA”) and Inter-Ramp Angle (“IRA”) and Overall Slope Angle (“OSA”) ...... 138 Figure 16.13 Example of kinematic analyses for the FW wall (slope dip direction of 30°), considering the rock mass fabric for the meta-sedimentary rocks...... 139 Figure 16.14 Application of the Mathews/Potvin open stope stability graph to the back of a stope located at a depth of 250 m...... 143 Figure 16.15 Phase 1B and 2011 Planned Underground Stopes ...... 146 Figure 16.16 Phase 2 and Planned Underground Stopes ...... 146 Figure 16.17 Phase 3 and 2011 Planned Underground Stopes ...... 147 Figure 16.18 Existing Fresh Air Raise ...... 149 Figure 16.19 Plan Showing 195 – 215 Level Development ...... 150 Figure 16.20 Plan Showing 170 – 195 Level Development ...... 151 Figure 16.21 Plan Showing 161 Level Development ...... 152 Figure 16.22 Plan Showing 135 - 141 Level Development ...... 153 Figure 16.23 Plan Showing 95 - 116 Level Development ...... 154 Figure 16.24 Entrance View of Ramp Portal ...... 155 Figure 16.25 Underground Explosive Detonator Magazine GA ...... 156 Figure 16.26 Surface Mine Air Heaters GA ...... 157 Figure 17.1 Project Schedule NICO Concentrator ...... 184 Figure 17.2 Project Schedule SMPP ...... 186 Figure 18.1 Plant Site Borehole Locations ...... 190 Figure 18.2 NICO Site Layout ...... 192

Figure 18.3 Plant Layout ...... 193 Figure 18.4 General Arrangement Plan of NICO Co-Disposal Facility ...... 194 Figure 18.5 Typical Cross-Section of the Co-Disposal Facility Perimeter Dyke ...... 195 Figure 18.6 Typical Layered Co-disposal Scheme ...... 196 Figure 18.7 Typical Cross-Section of Seepage Collection Ponds Dams ...... 197 Figure 18.8 SMPP Site Layout ...... 199 Figure 18.9 SMPP Plant Layout ...... 200 Figure 19.1 Application of Industrial Usage ...... 205 Figure 19.2 Global Electric Vehicle Battery Sales ...... 205 Figure 19.3 Proportion of World Cobalt Production (%) ...... 205 Figure 19.4 Applications and Uses of Bismuth ...... 207 Figure 19.5 World Bismuth Reserves ...... 208 Figure 19.6 Historical and Forecast Gold Price...... 208 Figure 22.1 Sensitivity Analysis (Cobalt Metal Option, Base Case Metal Prices-Exchange Rate) ...... 252 Figure 24.1 Crystalization Block Flow Diagram (Option A) ...... 257 Figure 24.2 Crystalization Block Flow Diagram (Option B) ...... 258

1.0 SUMMARY

1.1 INTRODUCTION

Fortune Minerals Limited (“Fortune”) began a program of exploration for iron oxide-hosted copper gold deposits (IOCG or Hydrothermal Iron Oxide-Hosted Replacement deposits”) in the Great Bear magmatic zone in the 1990‟s. This led to the identification of the Lou Lake area as a prospective location and to the staking of the NICO claims which Fortune has been actively exploring since 1994 Significant Gold-Cobalt-Bismuth-Copper mineralization in a number of different zones on the property have been discovered, including the “Bowl Zone” in 1995, which hosts the currently known mineral reserves on the property.

After the initial discovery of surface mineralization in the western part of the Bowl Zone, a number of drilling campaigns, resource estimates and studies were carried out as described below, each campaign and study building on a more complete database than the previous one. Exploration work on the property or offsite studies have been conducted continuously since the Bowl Zone discovery. This zone is the principal mineralized zone of interest on the property and was the subject of an initial feasibility study by Micon International Limited (“Micon”) 2007 and subsequently by additional advanced studies summarised in this report.

1.2 STUDY PARTICIPANTS AND RESPONSIBILITIES

Fortune engaged Jacobs Mineral Canada Inc., formerly Aker Solutions a division of Aker Solutions Canada Inc. (“Jacobs”), to produce Front End Engineering Design (“FEED”) reports regarding the development of the NICO Concentrator and related Saskatchewan Metals Processing Plant (“SMPP”). The scope included mineral and metallurgical plant design, project infrastructure, and development of process capital and operating costs.

The following firms were engaged by Fortune to provide specific project services listed below with inputs to the FEED and Technical reports related to their scope:

 P&E Mining Consultants Inc. (“P&E”) updated the geological model and prepared the updated mineral reserve estimates, as well as the mine plan, mining fleet selection, mine operating and capital costs, and production scheduling.  Golder Associates Ltd. (“Golder”) completed geotechnical engineering, environmental baseline studies and modelling, the design of the water treatment and mine rock and thickened tails co-disposal facilities in the NT, and they are the lead consultants for the environmental assessment (“EA”) process in the NT.  EBA Engineering Consultants Ltd. (“EBA”) completed the NICO site infrastructure geotechnical work and design of the NICO access road.  MDH Engineered Solutions Corp. (“MDH”) was responsible for the SMPP environmental baseline studies and site geotechnical work, the design of the process residue storage facility and is the lead consultant for the EA review process in Saskatchewan.  SGS Lakefield Research Limited (“SGS Lakefield”) completed the metallurgical test work and contributed to the SMPP design engineering  Dan Mackie Associates (“DMA”) designed the bismuth process building and equipment  Skybeco Inc., (“Skybeco”) conducted a metals price and marketing study

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A number of additional engineering consultants contributed to specific parts of the grinding, flotation and hydrometallurgical circuit designs and the metal marketing analysis.

All of the work was conducted to an accuracy of +/- 15% variance, except for the cobalt sulphate solvent extraction option, which was conducted with an overall accuracy of -10% to +25%.

The studies outlined in this report, have been prepared pursuant to the regulations and guidelines of NI 43-101FI.

1.3 PERSONAL SITE INSPECTIONS

The following site visits were carried out by the Qualified Persons:

 Marc Rougier of Golder visited the NICO property in September 2003 for three weeks and in September 2008 for three days as part of investigations to address the mining geotechnical and physical hydrogeology aspects of the project.  Ken Bocking of Golder visited the NICO property from 06 July 2010 to 07 July 2010 in order to examine the site of the Co-disposal Facility (CDF) at the NICO mine site.  Eugene Puritch of P&E Mining Consultants Inc., visited the NICO mine site on July 10 and 11, 2004 and on April 24, 2012. At that time a verification sampling program was undertaken.  Graham Holmes of Jacobs Minerals Canada visited the NICO mine site on April 24, 2012.  Alex Duggan of Jacobs visited the NICO mine and SMPP sites during the period of April 17 2012 through April 24, 2012.  Wade Sumners of MDH visited the SMPP site on June 9, August 11, and September 24, 2010 to conduct a biological assessment for the EIA.

1.4 PROPERTY DESCRIPTION AND LOCATION

The NICO project is found in National Topographic System (“NTS”) quadrant 85 N/10 at 63º, 33‟ N and 116º, 45‟ W in Canada‟s Northwest Territories (“NWT”). The property is approximately 160 km by air to the northwest of Yellowknife, which is located on the north shore of Great Slave Lake.

The NICO Property consists of 10 mining leases covering approximately 5,140 ha. As of August 9, 2007, Fortune holds 100 % of the interest in the NICO Property after it purchased the minority interest previously held by Candou Industries Inc.

As of the effective date of this report the 10 mining leases that comprise the NICO Property are in good standing and can remain so until at least 2023.

1.5 GEOLOGY AND MINERALIZATION

The mineralization at NICO is hosted in brecciated clastic sedimentary rocks of the Treasure Island Group that were previously thought to be part of the nearby Snare Group. The deposit occurs near the unconformity with overlying felsic volcanic rocks of the Faber Group. The sedimentary and volcanic rocks are both intruded by feldspar +/- quartz +/- amphibole porphyritic felsic dykes that broadly coeval with and related to the Faber Group volcanics. P&E Mining Consultants Inc., Report No. 247 Page 2 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Ore is contained in a series of 40º north dipping stacked stratabound lenses of ironstone. The main mineralized lenses are referred to as the Upper, Middle and Lower Zones, which are up to 1.5 km in length, 550 m in width (down dip) and 70 m in thickness (across dip).

The host sedimentary rocks in the vicinity of the ore zones have been extensively hydrothermally altered to biotite amphibole ironstones and schists and biotite-amphibole-magnetite ironstones and schists, an alteration assemblage representing extensive addition of iron and potassium to the rocks. Ore is associated with the approximate 5% disseminated and fracture filling sulphide fraction. The minerals of economic interest are primarily native gold, a solid solution series between cobaltian arsenopyrite and cobaltite, bismuthinite, native bismuth and chalcopyrite.

The overlying and capping Faber Group volcanics have also been extensively hydrothermally altered on a nearly regional scale. They have experienced the emplacement of significant amounts of potassium and the associated development of microcline giving them a distinct and ubiquitous orange-pink hue.

Mineral resource estimates were previously prepared by Mumin in 1997 and 1998b, SNC Lavalin in 1999, and Strathcona Mineral Services (Strathcona) in 2000 (presented in a scoping study in 2002), as well as an updated in-house estimate by Goad and Puritch in 2002. Two of these estimates were accompanied by scoping studies and preliminary economic evaluations. The study results were generally encouraging and each identified further drilling and/or mineralogical and metallurgical studies. In 2004, Micon was engaged to prepare an updated mineral resource estimate for the NICO deposit. Micon was then subsequently engaged, together with other engineering companies, to prepare a full feasibility study for the NICO project that was completed in February 2007 and was summarized in a technical report that was filed on the SEDAR website (www.Sedar.com).

1.6 EXPLORATION STATUS

Since the 2007 Micon feasibility study and technical report considerable additional work has been conducted on the NICO project. This includes completion of an underground test mining program that had been initiated prior to the feasibility study. The program was comprised of approximately two kilometres of underground ramp development work and crosscuts through two levels of the Middle Ore Zone. A ventilation and emergency egress raise was also driven to the surface.

Approximately 200 tonnes of NICO ores were composited in two bulk samples from the underground test mining program emulating the grades of the underground ores and the open pit ores in the NICO mineral reserves. The samples were shipped to SGS Lakefield in Lakefield, Ontario for pilot plant processing in 2007. This pilot plant and proved the process flow sheet, verified the production of cobalt and bismuth cathode products, and resulted in higher flotation recoveries than had been used in the previous Micon feasibility study. The economic impact of this metallurgical test work, together with updated metal price and currency exchange rate assumptions were used to by Micon to prepare an updated financial model and was summarized in a news release, dated May 8, 2008.

In 2009, Fortune made the decision to move the downstream processing of NICO concentrates to higher value metal products from the mine site to a property near Langham, Saskatchewan, approximately 26 kilometres northwest of the city of Saskatoon. This decision was made to P&E Mining Consultants Inc., Report No. 247 Page 3 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. mitigate the impact of higher diesel generated power costs at the NICO site after the Northwest Territories Power Corporation indicated that there would not be power available from the Snare electrical grid. The Saskatchewan site also had the additional benefit of being close to other services, process reagent sources, a skilled pool of engineers and chemical plant technologists and tax incentives that were offered by the Saskatchewan Government.

Between 2010 and 2012, an additional 30 tonne pilot plant was conducted on NICO ores at SGS Lakefield in Lakefield, Ontario. This pilot plant was conducted to test the impact of grade variability of ores during flotation, prove the production of a bismuth ingot product as well as a cobalt sulphate heptahydrate product, and also to prove gold and cobalt recoveries from blending the bismuth process residue with the cobalt concentrate autoclave feed in order to simplify the gold recovery circuit and mitigate the potential for gold recovery losses from refractory ores. The pilot plant was successful in achieving these objectives.

In 2010, an additional 38 holes were drilled into the NICO deposit to test for extensions to the deposit and extend the gold zones that were locally open for potential expansion. This program was successful in extending the overall strike length of the deposit, as well as expanding and better defining the deposit limits, including its gold-rich, high grade core and the crosscutting post mineral felsic intrusions. The infill drilling program was also successful in capturing resource blocks stranded from the main portion of the deposit in previous estimates.

Fortune has completed several phases of diamond drilling, totalling 326 holes, between 1996 and 2010 in the general vicinity of the NICO deposit. Most of the holes fall within the interpreted mineralized extents of the three tabular zones. Of the total holes drilled 299 were utilized for resource estimation. These holes are located between sections 1400 NW and 2800 NW. To date, gold-cobalt-bismuth-copper mineralization at the NICO deposit has been intersected over a strike length of over approximately 1,500 m and the deposit is now essentially closed off by drilling.

1.7 MINERAL RESOURCES

The mineral resource was constrained with a geological model prepared by Fortune and reviewed by the estimator. Three mineralization domains and five lithological domains were modeled. Mineralization domains were defined by continuous mineralization and assay intervals equal to or greater than a calculated NSR value of $40.00.

Within the domains a block model was reported by estimating a net smelter return (NSR) value for each block using parameters provided from the extensive metallurgical test work and mining scoping studies completed on the deposit. At the time of resource estimation it was anticipated that the upper portions of the deposit (to approximately 75 m below surface) would be mined by open pit methods and the remainder would be mined from underground in order to provide early access to the gold-rich mineralization at the core of the deposit.

All open pit mineral resources are reported against a $46.00 NSR cut-off, as constrained within the optimized pit shell. Underground mineral resources are reported outside the optimized pit shell against an $80 NSR cut-off. The effective date of this estimate is 30 November 2011. P&E Mining Consultants Inc., Report No. 247 Page 4 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

The Mineral Resource Estimate is shown in Table 1.1 below.

TABLE 1.1 (1)(2)(3) NICO ESTIMATED OPEN PIT AND UNDERGROUND MINERAL RESOURCES NSR Cut-off Tonnes x Area Class Au g/t Bi % Co % $CDN/t 1,000 Measured 18,911 1.05 0.15 0.12 Open Indicated 10,983 1.19 0.14 0.12 $46 Pit M+I total 29,894 1.10 0.15 0.12 Inferred 2 0.30 0.07 0.08 Measured 231 2.29 0.06 0.15 Indicated 764 1.72 0.07 0.16 U/G $80 M+I total 995 1.85 0.07 0.16 Inferred 31 0.65 0.11 0.25 (1) Mineral resources are defined within an optimized pit shell that incorporates project metal recoveries, estimated operating costs and metals price assumptions. (2) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, marketing, or other relevant issues. The mineral resources in this news release were estimated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council. (3) The quantity and grade of reported Inferred resources are uncertain in nature and there has been insufficient exploration to define these Inferred resources as an Indicated or Measured mineral resource and it is uncertain if further exploration will result in upgrading them to an Indicated or Measured mineral resource category.

1.8 MINERAL RESERVES

1.8.1 Underground Reserve

The underground mineral reserves were re-estimated using a revised mine plan to extract high gold grade ore for the purposes of the present post-feasibility mining update study. The underground mineral reserves estimate used in the present study is presented in Table 1.2.

TABLE 1.2 (1) UNDERGROUND MINERAL RESERVE ESTIMATES Classification Tonnes Au (g/t) Co (%) Bi (%) Cu (%) Proven 282,100 4.93 0.14 0.27 0.03 Probable 93,900 5.60 0.11 0.19 0.01 Total 376,000 5.09 0.13 0.25 0.02 (1) Mine recovery and dilution are included in these quantities with metal grades (2) All of the material designated as Reserves in the underground was derived from Measured and Indicated Resources that have been demonstrated to be economic as result of the current study and are therefore designated as Proven and Probable Reserves using the Canadian Institute of Mining, Metallurgical and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council on December 11, 2005 (the CIM Standards)

The existing exploration decline, completed in 2006, will be utilized for access to the underground stope mining areas. The proposed underground mine plan is based on a P&E Mining Consultants Inc., Report No. 247 Page 5 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. development program that includes: the extension of the existing 5 m x 5 m exploration decline to the 95 m level. There are a total of 21 blasthole stopes, life-of-underground-mine.

1.8.2 Open Pit Reserve

The open pit reserve estimate is based on dilution and extraction to bench defined mining outlines. The open pit mineral reserve estimate is shown in Table 1.3 and has been determined based on selection of blocks that are above the marginal economic NSR cut-off of C$48.07 per tonne. TABLE 1.3 (1)(2) OPEN PIT RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 20,513 0.94 0.11 0.15 Probable 12,099 1.05 0.11 0.13

Total 32,612 0.98 0.11 0.14 (1) Mine recovery and dilution are included in these quantities along with metal grades (2) All of the material designated as Reserves in the underground portion was derived from Measured and Indicated Resources that have been demonstrated to be economic as result of the current study and are therefore designated as Proven and Probable Reserves using the Canadian Institute of Mining, Metallurgical and Petroleum (CIM) Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.

1.8.3 Total Reserve

The total combined underground and open pit reserves are presented below in Table 1.4 and are based on dilution and extraction to defined mining outlines.

TABLE 1.4 (1)(2) TOTAL RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 20,795 0.99 0.11 0.15 Probable 12,193 1.09 0.11 0.13

Total 32,988 1.02 0.11 0.14 (1) Mine recovery and dilution are included in these quantities along with metal grades (2) All of the material designated as Reserves in the underground portion was derived from Measured and Indicated Resources that have been demonstrated to be economic as result of the current study and are therefore designated as Proven and Probable Reserves using the Canadian Institute of Mining, Metallurgical and Petroleum (CIM) Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.

1.9 MINING

1.9.1 Mine Production Schedule

The mine production schedule assumes that open pit waste stripping will commence in July 2015 (month 19), open pit ore production in September 2015 (month 21 ), underground dewatering and rehabilitation of existing workings in May 2016 (month 29), underground ore and waste development starting in June 2016 (month 30) and underground stoping starting in July 2016

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(month 31). Underground ore production will end in January 2017 (month 37). Open pit ore production will end in 2035 (year 22nd). The mine schedule is based on:

 The open pit mining operation will operate from September 2015 (month 21) to 2035 (year 22) and produce a total of 32,611,500 t grading 0.98 g/t Au, 0.14% Bi, 0.11% Co and 0.04% Cu.  The underground mining operation will operate from May 2016 (month 29) to January 2017 (month 37) and produce a total of 376,000 t grading 5.09 g/t Au, 0.25% Bi, 0.13% Co and 0.02% Cu.  The underground schedule is based on working 7 days per week.  During the 8 months of underground production an average 1,544 tpd of underground ore will be supplemented with 3,106 tpd open pit ore.  After the completion of the underground mining program, the open pit is scheduled to produce 4,650 tpd ore.

1.9.2 Underground Mine Development

An underground mining contractor will mobilize and setup on site in May 2016. It is assumed that mine dewatering and rehabilitation of the existing underground infrastructure will be completed during May 2016, by this contractor. Two trackless development crews are scheduled to start working during the first week of June 2016. Development Crew 1 (C1) will start with mine development on the 195 - 215 level area. Development Crew 2 (C2) will start with mine development on the 170 – 195 level area. Initially the contractor development crews will advance at a rate of 6.5 metres per day, single heading. Once established the contractor will advance at a rate of 8.0 metres per day, double heading. A summary of the level access and drift development schedule is presented in Table 1.5.

TABLE 1.5 MINE DEVELOPMENT SCHEDULE SUMMARY Level Crew Start Date Finish Date 195-215 C1 June 1, 2016 Aug 2, 2016 170-195 C2 June 1, 2016 Sept 3, 2012 161 C1 July 28, 2016 Sept 5, 2016 135-141 C1 Aug 12. 2012 Oct 30, 2016 116-95 C2 Aug 24, 2012 Nov 26, 2016

1.9.3 Underground Stope Development

Once the access and footwall drifts have been completed to the first accessible stopes No. 21 and No. 1 on the 195-215 and 170-195 levels, development Crews 1 and 2 will proceed with stope development to all stopes as they become accessible. Development crews will excavate undercut cross-cuts, undercut slashes, slot raises and complete stope drilling in those stopes. A summary of the stope development schedule is presented in Table 1.6.

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TABLE 1.6 STOPE DEVELOPMENT SCHEDULE SUMMARY Level Crew Start Date Finish Date 195-215 C1 June 9, 2016 Sept 2, 2016 170-195 C2 June 11, 2016 Oct 25, 2016 161 C1 Aug 6, 2016 Nov 4, 2016 135-141 C1 Aug 17, 2016 Dec 17, 2016 116-95 C2 Sept 13, 2016 Jan 20, 2016

1.9.4 Underground Stoping

Stoping includes blasthole blasting, and mucking and truck haulage to surface. There will be one stope blasting crew, and an average 3 scooptram and 2.4 haulage truck drivers per day. A schedule summary of production blast hole blasting, mucking and truck haulage to surface is summarized in Table 1.7.

TABLE 1.7 PRODUCTION BLAST HOLE BLASTING, MUCKING AND HAULAGE SCHEDULE SUMMARY Level Start Date Finish Date 195-215 July 28, 2016 Sept 7, 2016 170-195 Sept 5, 2016 Oct 27, 2016 161 Nov 4, 2016 Nov 12, 2016 135-141 Nov 13, 2016 Dec 21, 2016 116-95 Dec 20, 2016 Jan 26, 2016

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TABLE 1.8 SUMMARY OF U/G STOPE MINING SEQUENCE Level Stope Tonnes Start Date Finish Date 195 - 215 Stope 21 9,253 July 28, 2016 Aug 2, 2016 195 - 215 Stope 3 7,112 Aug 8, 2016 Aug 21, 2016 195 - 215 Stope 20 9,424 Aug 25, 2016 Aug 30, 2016 195 - 215 Stope 19 9,210 Sept 2, 2016 Sept 7, 2016

170 - 195 Stope 2 3,588 Sept 5, 2016 Sept 7, 2016 170 - 195 Stope 1 13,740 Sept 13, 2016 Sept 21, 2016 170 - 195 Stope 4 10,137 Sept 28, 2016 Oct 3, 2016 170 - 195 Stope 18 15,140 Oct 10, 2016 Oct 18, 2016 170 - 195 Stope 17 8,207 Oct 18, 2016 Oct 23, 2016 170 - 195 Stope 16 6,456 Oct 24, 2016 Oct 27, 2016

161 - 161 Stope 15 14,735 Nov 4, 2016 Nov 12, 2016

135 - 141 Stope 5 11,558 Nov 13, 2016 Nov 20, 2016 135 - 141 Stope 14 10,601 Nov 22, 2016 Nov 28, 2016 135 - 141 Stope 13 10,393 Nov 30, 2016 Dec 6, 2016 135 - 141 Stope 6 11,695 Dec 6, 2016 Dec 9, 2016 135 - 141 Stope 11 8,472 Dec 17, 2016 Dec 22, 2016

116 - 95 Stope 9 2,971 Dec 20, 2016 Dec 22, 2016 116 - 95 Stope 8 6,968 Dec 25, 2016 Dec 29, 2016 116 - 95 Stope 7 5,994 Jan 2, 2017 Jan 5, 2017 116 - 95 Stope 12 10,474 Jan 11, 2017 Jan 17, 2017 116 - 95 Stope 10 12,788 Jan 19, 2017 Jan 26, 2017

Total 198,917

1.9.5 Capital Costs – Underground Mine

The underground mining manpower and equipment requirements are based on contractor estimates for contractor supplied underground services and P&E estimates for Fortunes‟ supplied underground services. The site already has a ditching excavator, explosive and detonator magazines, and a propane-fired mine air heater.

1.9.5.1 Pre-production Underground Capital Costs

All underground pre-production costs will be capitalized. The underground pre-production period will start early May 2016 and end on May 31, 2016. A summary of 2016 pre-production capital costs is presented in Table 1.9.

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TABLE 1.9 SUMMARY OF 2016 UNDERGROUND PRE-PRODUCTION CAPITAL COSTS Description Estimated Cost ($) Contractor Mobilization and Dewatering 1,356,248 Diesel Fuel 50,381 Fortunes Indirect Labour 22,362 Fortunes Support Equipment 574,426 Electric Power 171,205

Total Estimated Cost 2,174,623

1.9.5.2 Underground Sustaining Capital Costs

A summary of sustaining capital purchases in 2016 and 2017 is presented in Table 1.10.

TABLE 1.10 SUMMARY OF SUSTAINING CAPITAL COSTS Description / Year 2016 ($) 2017 ($) Total ($) Contract Mining 31,135,493 1,932,360 33,067,853 Diesel Fuel 1,100,838 49,373 1,150,210 Fortunes Indirect Labour 557,545 91,760 649,306 Fortunes Support Equipment 140,270 27,579 167,849 Electric Power 1,737,018 305,214 2,042,232 Sample Preparation & Assaying 86,518 7,483 94,001

Total Estimated Cost 34,757,683 2,413,768 37,171,451

1.9.6 Underground Mine Operating Cost

All underground operating costs are capitalized.

1.10 ENVIRONMENTAL

A Developer‟s Assessment Report (“DAR”) was submitted to the Mackenzie Valley Environmental Impact Review Board in May 2011. The following are high level summaries of the impact predictions to the biophysical environments from the NICO Project, which were presented in the DAR:

Changes to the biophysical environment from the NICO Project are not predicted to result in significant adverse residual impacts to valued ecosystem components. Consequently, the NICO Project is not predicted to have significant adverse impacts on traditional and non-traditional land use practices.

The active mine area will be small (approximately 485 hectares (“ha”)), with limited changes made to the natural flow of water. The NICO Project will have a minimal effect on water quantity, air, soils, vegetation, and wildlife and fish health. Closure, caribou and water quality have been identified as the most important concerns related to the environment by the communities. People should not be able to observe a change in the availability of wildlife due to

P&E Mining Consultants Inc., Report No. 247 Page 10 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. effects of the NICO Project, relative to current natural changes in population size. Changes in water, soils, and plants caused by the NICO Project in the small area at and near the mine site will not affect the health of wildlife, or the health of people that eat wildlife.

Fortune has prepared an Environmental Impact Statement (“EIS”) to obtain environmental approval to construct the SMPP to meet global demands of high-value metal cathode products. On 19 July 2010, Fortune submitted an Environmental Project Proposal for the proposed SMPP to the Saskatchewan Ministry of Environment (“MOE”). The MOE determined that an EIS was required for the project. MOE released draft Project Specific Guidelines (PSGs) for the development on 21 January 2011. The PSGs were finalized on 31 March 2011 after a review and input from the public. The EIS document addresses the items identified in the PSGs. Numerous options and alternatives for the proposed SMPP project have been considered by Fortune. These include alternatives for the selection of the site, processing facility and site layout, water supply, water and residue storage, process solution disposal, and the metallurgical processes. Fortune is confident that the proposed plans to construct the SMPP have considered all viable options to ensure that potential environmental impacts are avoided and/or minimized. Fortune is committed to preventing or reducing adverse environmental effects associated with the project, wherever possible. General mitigation measures planned for this project include the following:

 Obtaining all required permits and approvals prior to construction;  Using best practice, environmentally sound construction methods (i.e. minimal clearing, salvaging topsoil, etc.);  Construction of appropriate containment systems (i.e. berms, perimeter ditches, ditches and ponds with engineered liners, etc.);  Use of best available technologies to reduce water, power, and energy use;  Use of best available technologies to reduce air emissions; and  Development and implementation of a site Environmental Monitoring Program.

These measures are intended to provide effective long-term containment and mitigation of environmental impact. Residual effects, following the mitigation measures, are expected to the terrain, air emissions, soils, vegetation, surface water runoff, land use, and socio-economics; however, these effects have an environmental consequence rating of low to minor. Impacts beyond the site boundaries are expected to be minimal. A summary of the predicted environmental effects is provided in the EIS and continual environmental monitoring during the facility operation and closure will ensure the appropriate mitigation measures are applied. Considering the potential impacts of the SMPP on the surrounding environment, there are no predicted cumulative effects from the proposed project. The area immediately surrounding the proposed development is dominated by agricultural activities and the processing of metal concentrates has a negligible effect on this surrounding land use.

1.10.1 SMPP Water Management

Fresh, Fire and Potable Water

The groundwater well system supplies fresh water for the process plant.

The fresh water tank has a capacity of 760 m³, equivalent to 3.5 hours of usage.

Flotation Process Water

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Flotation process water is collected from overflows from the cobalt concentrate and bismuth concentrate thickeners. A nominal fresh water flow of 4.2 m³/h is added as make-up. The flotation process water tank has a capacity of 90 m³, equivalent to 1.7 hours usage.

Hydrometallurgical Process Water

Hydrometallurgical process water is collected from agitator seal cooling water return and treated RO water, with 21.3 m³/h of fresh water make-up. The treated RO water flow rate is 31 m³/h.

Demineralized Water and Steam Boiler

The demineralised water treatment plant produces 10 m³/h of demineralized water to be pumped to boilers and various other users in the plant.

1.10.2 Permitting

The NICO Project is regulated by the Wek‟èezhìı Land and Water Board (“WLWB”) under the Mackenzie Valley Resource Management Act (“MVRMA”). The Tłįchǫ Government and the WLWB regulate the use of settlement and Crown land and water in their respective settlement areas.

The Mackenzie Valley Review Board (“MVRB”) is established under the authority of the MVRMA to review the potential environmental effects of developments proposed within the Mackenzie Valley area of the NWT. There are 3 stages in the environmental assessment process (“EAP”) in the Mackenzie Valley. The MVRB provides the following description of the stages:

 Preliminary Screening  Environmental Assessment  Regulatory Phase

Progress under the EAP has led to significant milestones which to date have included: submission of the DAR, [the Northwest Territories analogue to an Environmental Impact Statement], on 20 May 2011, and a Technical Session held in Yellowknife, NWT from 7 to 9 February 2012. Public Hearings are tentatively scheduled for the week of August 27-31, 2012.

The NICO Project is anticipated to have significant positive impacts on the economics of the Tłįchǫ communities, and both positive and negative (but not significant) impacts on the social and cultural environments. Following the public hearing noted above the MVRB will make a recommendation to the Federal Minister and, assuming the recommendation is affirmative, the file will be returned to the WLWB for the regulatory stage of permitting. The regulatory phase involves the development and issuance of a water license and land use permit that defines the specific conditions under which the mine must be constructed and operated. Once these authorizations are in place, other permits, licenses, and authorizations can be obtained.

Fortune and the Tłįchǫ Government have signed a Co-operative Relationship Agreement for the NICO Project. This agreement establishes the framework and path forward for further negotiations, defines primary liaison officials, and sets the communication protocol for the two parties. The Tłįchǫ Government and Fortune have also signed an Environmental Assessment Funding Agreement to support the Tłįchǫ Government with their review of the NICO Project DAR. P&E Mining Consultants Inc., Report No. 247 Page 12 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Fortune has also agreed to fund a Traditional Knowledge (“TK”) Study that will focus on providing traditional knowledge and land use practice information for the environmental review of the NICO Project. This study, which will be carried out by the Tłįchǫ Government, will contribute to the environmental assessment process.

As previously stated, the EIS for the SMPP has been submitted to the MOE, who act as the project coordinator for an interdepartmental and intergovernmental (i.e. federal government) technical review of the EIS. A preliminary review is complete and preliminary Technical Review Comments (“TRC”) has been issued to Fortune. Fortune is in the process of addressing the preliminary TRCs and will submit an addendum document with the EIS for the creation of final TRCs by MOE. It generally takes 60 days to complete a review of the revised EIS and creation of final TRCs. If further revisions are required, the review process may require an additional 30 to 45 days to review the revised EIS. Once MOE is satisfied the items have been properly addressed, it will issue a final TRC document and release the EIS and addendum documents for a public review for a minimum of 30 days.

Additional permit, license, approval, and notification requirements that are anticipated to be required for the SMPP, once ministerial approval is received, include, but may not be limited to:

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Aeronautical Obstruction Clearance Closure Notice Notice of Work Close to Railways Commencement of Work Notice Air Emissions Permits – Permit to Construct Facility Closure Approval (during operation) – Alter, or Extend Fuel Burning Equipment/ Decommissioning and Reclamation Plan Permit to Operate Fuel Burning Equipment Approval Aquatic Habitat Protection Permit Gas Permit Groundwater Investigation Permit (received for Boiler Licenses EIS) Business name registration Electrical Permits Pipeline Work License – license to construct, Overweight/Over Dimension Permit alter, operate, or abandon a pipeline Pollutant Control Facility Permits – Approval to Construct, Alter or Extend a Pollutant Control Facility/Approval to Operate a Plumbing Permit Pollutant Control Facility/Approval to Decommission and Reclaim a Pollutant Control Facility Release from Site (after successful completion Sign Permit (off premise) of the decommissioning and reclamation plan) Storage Facility (Hazardous Goods) Permits/Licence – Approval to construct a Hazardous Substance or Waste Dangerous Goods Storage facility/Approval to operate a Vendor‟s License/Consumer Permit Hazardous Substance or Waste Dangerous Goods Storage facility/ Approval to decommission a Hazardous Substance or Waste Dangerous Goods Storage facility Wastewater Disposal Well Permit Water Rights License Building Permit – Construction or Alternation of a Building/Authorization of Construction Discretionary Use Approval (may be part of and Approval of Fire Prevention/Protection the rural municipality zoning bylaw) Systems Overweight Permit (possibly during Road Haul Agreement (municipality) construction)

Each permit/license may have a different regulatory agency responsible for issuing a permit/license and the application submittal time and/or regulatory agency processing time will vary for each permit/license.

1.11 TAILINGS

1.11.1 NICO Mine Tailings and Waste Co-disposal Facility

The mining process will generate a total of 29.9 Mt of tailings and 96.9 Mt of mine rock. Both these waste streams will be disposed together in a facility referred to as Co-disposal Facility (“CDF”).

The CDF will be contained by a Perimeter Dyke comprising a prism of mine rock at least 25 m thick. The Perimeter Dyke will be raised continually in 5 m lifts using the upstream construction

P&E Mining Consultants Inc., Report No. 247 Page 14 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. method. Inside the Perimeter Dyke, the CDF will comprise a “layer cake” of alternating layers of mine rock and tailings about 5 m thick. The Perimeter Dyke will be free draining but it will retain tailings particles. Five Seepage Collection Ponds (“SCPs”) will be constructed downstream of the CDF at topographically low areas (Figure 18.3) to intercept any tailings water that may seep through the Perimeter Dyke. Water collected in the SCPs will be pumped to the Process Plant for re-use.

1.11.2 SMPP Waste and Tailings Disposal

Each SMPP cell will have a dual containment liner and a leak detection system. The primary liner will be a composite liner consisting of a geomembrane placed directly over approximately 0.45 m of compacted soil. Leak detection is provided by a geocomposite material installed beneath the primary liner. Secondary containment is provided by approximately 0.2 m of a compacted soil liner under the geocomposite material. There is approximately 9 m to 18 m of low conductivity till between the base of the PRSF and the Dalmeny Aquifer, providing a high level of secondary containment for the process residue.

A leachate collection system was also provided for each cell to collect any fluid (i.e. leachate, precipitation, snowmelt, etc.) that accumulates when the cell is open.

Perimeter ditches around the PRSF facility and a runoff collection pond dedicated to the PRSF will collect any runoff once the cells are capped, prior to the establishment of vegetation. This collected runoff may be directed to the process water storage pond, for use in the facility, or monitoring may indicate its suitability to be released to the environment

1.12 WATER SUPPLY

1.12.1 NICO Site Water Management

The major components of the water management system of the Project will comprise: Lou Lake, the CDF Reclaim Pond, five SCPs, a Surge Pond, Open Pit sumps, a Process Plant Runoff Pond, sewage treatment plant (“STP”) and an Effluent Treatment Facility (“ETF”).

During Operation: Lou Lake will be the source of fresh water for Process Plant use, dust control, and potable water. The Surge Pond will temporarily store contact water pumped back from the SCPs and the Reclaim Pond. Water will be pumped from the Surge Pond either to the Process Plant for reuse or to the Effluent Treatment Facility (ETF) for treatment. Treated water from the P&E Mining Consultants Inc., Report No. 247 Page 15 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

ETF will be pumped through a diffuser directly into Peanut Lake. Water balance analysis indicates that the average flow discharged into Peanut Lake will be a relatively small flow of about 290,000 m3/year.

1.13 MARKETING

1.13.1 Cobalt

Fortune has engaged Skybeco. to prepare cobalt and bismuth market assessments and pricing forecasts. According to Skybeco, the long-term outlook for cobalt is generally positive. Despite the numerous projects previously expected to bring large cobalt supply quantities on-stream, many of these have not materialized as projected.

It is anticipated that 2012 will see a small surplus of refined Co supply over demand, but it is expected that this situation will switch to a small deficit in 2013. Demand will again exceed supply in 2014 and 2015, though supply of refined 99.8% cobalt cathode will increase as Chinese refiners ramp up production to keep up with battery demand in Asia.

It is anticipated that cobalt market prices will experience primarily sideways movement while markets adjust to the financial tension in Europe and a slightly slower growth rate in China over the mid-year months. Sometime approaching late third quarter 2012, it is anticipated prices will begin to buoy upwards as cutbacks of marginal producers start to make themselves felt. Towards end 2012 it is forecast pricing of high grade cobalt metal to be in the $16/lb range. While fluctuations will always be a part of market prices, longer term pricing beyond 2012 is expected to move within the range of US$18/lb to US$22/lb. Cobalt sulphate heptahydrate, the cobalt product used broadly for the manufacture of lithium ion and nickel metal hydride batteries as well as other chemicals, trades at an average 21% premium over 99.8% cobalt metal.

1.13.2 Bismuth

According to Skybeco, bismuth supply is expected to remain stable during 2012 and 2013; many of the smaller mines in China will be forced to discontinue operations as a result of environmental clampdowns by the government and/or the unfavourable economic operating conditions prevailing at this time.

Demand growth for bismuth is expected to be in the 8% to 10% range annually over the coming 3 years. Bismuth demand is forecasted to enjoy healthy growth due to its role as a non-toxic substitute for lead, especially in such applications as free cutting (machining) steels and coppers/brasses.

With continued healthy bismuth demand growth and supply being kept in check, prices for 99.995% bismuth ingot are forecasted to increase by US$1.50 in 2013 reaching $$14.00/lb, and then hitting $15.50/lb by the end of 2014. Longer term pricing may see even higher pricing if growth momentum in lead-substitution applications is speeded along by regulations/legislation.

1.14 PROCESS PLANT

The NICO plant is a conventional concentrator that will be located in the NWT with an ore production rate of 1.7 Mt/a. The process includes crushing, screening, material handling, grinding, gravity concentration, flotation, concentrate thickening, packaging, tailings handling P&E Mining Consultants Inc., Report No. 247 Page 16 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. and transportation. The concentrate will be transported by road and rail to the SMPP for metal separation and further refining.

The metallurgical processing plant in Saskatchewan includes multi-step polymetallic extraction processes after re-grinding the bulk concentrate and secondary flotation to produce separate gold-bearing cobalt and bismuth concentrates. These are subjected tochloride leach electrorecovery, autoclave pressure oxidation, cobalt precipitation, copper re-leach and cementation, cobalt dissolution, ion exchange with precipitation of zinc and nickel, cobalt electrowinning and/or crystallization of cobalt sulphate heptahydrate, autoclave residue cyanidation, and gold recovery and refining. The unwanted solids of the process are sent to an onsite process residue storage facility. The brine stream is subject to reverse osmosis before injection into a saline aquifer. The cobalt cathode, electrowon bismuth, copper cement, gold doré, nickel and zinc carbonate products are packaged and shipped to various end users

1.15 PROJECT EXECUTION

The NICO plant will be built in 12 months after completion of the all-weather road (“AWR”). The project will require delivery of materials on site at least 6 months prior to the opening of the AWR. The SMPP plant will be constructed to coincide with production of ore concentrate from NICO.

1.16 CAPITAL AND OPERATING EXPENSES

The following table shows the projected average annual metal production for each of NICO‟s component commodities.

TABLE 1.11 AVERAGE METAL PRODUCTION Gold Cobalt Bismuth Copper

(oz) (lbs) (tonnes) (lbs) (tonnes) (lbs) (tonnes) Average 40,500 3,473,586 1,576 3,681,824 1670 559,397 254 Annual LOM Total 800,091 69,471,715 31,512 73,636,474 33,401 11,187,946 5,079

The cash cost net of by-product credits for gold, cobalt and bismuth were determined for several of the metal price cases and are shown in Table 1.12.

TABLE 1.12 PROJECTED OPERATING CASH COSTS FOR THE COBALT METAL OPTION Cash Cost Metal Price – Cash Cost Net of By-Product Credits Equivalent Oz Au Exchange Rate Gold Gold Cobalt Bismuth Cases $US/equivalent oz US$/oz $US/lb $US/lb Base Case 831.30 (356.70) (0.81) (8.63) 3-Year Trailing 859.94 (77.23) 1.98 (5.79) Average Current 990.44 142.52 (1.07) (4.83) Escalated 943.87 (551.70) (4.58) (13.05)

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The underground mining fleet is assumed to be provided by contracted service and the cost of the equipment is built into the operating costs for the underground part of the mine. It is expected that the open pit mine fleet will be leased. Payback is approximately 6.8 years for the Base Case cobalt metal option.

The overall capital cost investment for both sites is summarized in Table 1.13.

TABLE 1.13 SUMMARY OF OVERALLCAPITAL COSTS Site Location Projected Capital Cost NICO Site C$M 210.1 SMPP Site C$M 230.3 Total Direct / Indirect & contingencies C$M 440.5 Sustaining Capital LOM C$M 113.5

The life of mine (“LOM”) average operating costs for the NICO project are shown in Table 1.14.

TABLE 1.14 SUMMARY OF OPERATING COSTS (LOM) FINANCIAL MODEL Activity Unit Costs Open Pit Mining Including Stripping C$ 8.67/tonne of ore processed Underground Mining C$ 99.34/tonne of ore processed Processing (includes milling, transportation & refining) C$ 43.91/tonne of ore processed General and Administrative Costs / Shared Services / Camp C$ 8.36/tonne of ore processed

Total Costs LOM C$ 61.97/tonne of ore processed

1.17 PROJECT ECONOMIC ANALYSIS

1.17.1 Economic Analysis

The internal rate of return (“IRR”), and 7% and 5% discounted net present value (“NPV”) for the NICO project have been determined for a variety of commodity price and U.S. dollar (“$US”) to Canadian dollar (“C$”) exchange rate cases, and also for two separate cobalt product options.

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TABLE 1.15 NICO ECONOMICS Cobalt Metal Option Cobalt Sulphate Option Metal Price Pre-Tax After Tax Pre-Tax After Tax & $M $M $M $M $M $M $M $M Exchange IRR IRR IRR IRR NPV NPV NPV NPV NPV NPV NPV NPV Rate Case % % % % (7%) (5%) (7%) (5%) (7%) (5%) (7%) (5%) Base Case 10.8 164.5 293.2 9.6 101.0 207.1 14.0 308.5 466.0 12.4 212.6 338.7 Prices 3-yr Trailing Average 7.4 17.1 114.6 6.6 (15.3) 69.0 10.5 146.8 270.0 9.3 86.7 188.4 Prices Current 7.1 2.1 99.7 6.2 (30.6) 53.4 9.6 109.5 228.2 8.5 57.6 156.8 Prices Escalated 13.9 315.2 477.8 12.3 214.9 344.7 17.1 467.1 660.1 15.2 332.4 483.7 Prices Optimistic 18.3 539.5 749.8 16.3 387.5 551.3 21.6 707.0 951.1 19.3 514.5 702.3 Prices Note: Base Case Price assumptions are US$1,450/troy ounce (“oz”) for gold, US$20/pound (“lb”) for cobalt, US$11/lb for bismuth and US$3.50/lb for copper at an exchange rate of US$ 0.95 = C$ 1. The 3-year Trailing Average Prices Case are as at May 31, 2012 and are US$1,359.94/oz for gold, US$18.53/lb for cobalt, US$9.83/lb for bismuth and US$3.51/lb for copper and an exchange rate of US$ 0.98 = C$ 1. The Current Price Case uses prices as at May 31, 2012 and are US$1,558.00/oz for gold, US$15.23/lb for cobalt, US$10.55/lb for bismuth and US$3.40/lb for copper and an exchange rate of US$ 0.97 = C$ 1. The Escalated Price Case uses metal price assumptions of US$1,800.00/oz for gold, US$22.50/lb for cobalt, US$12.50/lb for bismuth and US$4.00/lb for copper and an exchange rate of US$ 1 = C$ 1. For the Optimistic Price Case uses US$2,000.00/oz for gold, US$25.00/lb for cobalt, US$15.00/lb for bismuth and US$4.50/lb for copper at an exchange rate of US$ 1 = C$ 1. The estimated costs for the cobalt metal option were developed to an accuracy of ±15%. The estimated costs for the cobalt sulphate solvent extraction option were developed to an overall accuracy of -10% to +25% with cost estimates for the cobalt sulphate circuit having scoping level accuracy (e.g. ±40%).

1.18 CONCLUSIONS AND RECOMMENDATIONS

The following conclusions and recommendations regarding the NICO project were derived from the present engineering and design study.

1.18.1 Interpretation and Conclusions

 The NICO deposit currently being developed by Fortune is a new cobalt-gold- bismuth-bearing IOCG deposit located 160 km northwest of Yellowknife in the NWT. A 2011 mineral Resource Estimate, prepared by P&E concluded that the deposit shows good continuity of mineralization and consequently the NICO deposit resources were entirely in the Measured and Indicated category. The resource estimate is suitable for use in an economic evaluation of a mining operation including Front-End Engineering and Design (Feed) level studies as presented in this report.  The pit slope design investigations indicate an engineering geology model comprised of strong competent rock masses. Results indicate that bench geometries with an inter-ramp slope angles of 50˚  Underground geometries were based on the same engineering geology model as the open pit. Recommended maximum hydraulic radius (“HR”) for unsupported walls ranges from 4.5 m (Hanging wall stability) to 6.6 m (sidewall stability). Systematic ground support, such as cable bolts, is not required. P&E Mining Consultants Inc., Report No. 247 Page 19 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

 Pit/underground workings interactions were assessed with respect to induced stresses on the pillars as the pit deepens. To mitigate the hazard of open pit mining above unfilled workings, void filling of stopes beneath the pit floor working areas, walls or ramps will be accomplished from the open pit, as part of safe mining practices..  The mining process will generate a total of 29.9 Mt of tailings and 96.9 Mt of mine rock at the NICO site. Both these waste streams will be disposed together in a facility referred to as the CDF. The CDF will be entirely located within the valley of the Grid Ponds and will be contained on all sides by a Perimeter Dyke.  Runoff and bleed water from the tailings deposition will be “reclaimed” back to the process plant for reuse. The major components of the water management system of the Project will comprise: Lou Lake, the CDF Reclaim Pond, five SCPs, a Surge Pond, Open Pit sumps, a Process Plant Runoff Pond, sewage treatment plant (STP) and an ETF.  A DAR was submitted to the Mackenzie Valley Environmental Impact Review Board in May 2011. Changes to the biophysical environment from the NICO Project are not predicted to result in significant adverse residual impacts to valued ecosystem components. Consequently, the NICO Project is not predicted to have significant adverse impacts on traditional and non-traditional land use practices.  For closure, a soil cover will be placed over the entire CDF facility. Seepage water will be routed through Wetland Treatment Systems into NICO Lake. The ETF will be maintained on site for 10 years as a backup to the Wetland Treatment Systems.  The NICO Project is anticipated to have significant positive impacts on the economics of the Tłįchǫ communities, and both positive and negative (but not significant) impacts on the social and cultural environments. The NICO Project is a small development compared to other mines in the NWT, but it will contribute to the overall labour, financial, physical, human, and social resources of both the NWT and more specifically the nearby communities.  The FEED study completed for the NICO Project is comprehensive covering all the main Project elements of mineral processing at the concentrator and hydromet processing at SMPP. The following is concluded:

 Available data is sufficient to meet the requirements of a FEED study.  The project is technically viable, subject to further works to be carried out in the detailed engineering.  The process plant is designed for approximately a 20 year Life of Mine at a production rate of approximately 1.7 Mt/a.

 The most crucial element for project execution will be the construction of the all- weather road to allow year round access and the implementation of standard northern construction methodology.  The projected economic outcome of the project has been demonstrated using a discounted after-tax cashflow model  Potential significant risks and uncertainties that could reasonably be expected to offset the projected economic outcome include:

 Cobalt product Option: the technical and economic viability of the cobalt metal option has been demonstrated. However based on the results of the

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economic analysis the cobalt sulphate option offers a higher IRR and NPV in comparison to the cobalt metal option and appears promising.  It is assumed that an all-weather access road to the NICO site will be built before the commencement of the main construction and available over the life of the project. The project schedule and operating cost estimates include allowances for envisaged delays / lost time due winter weather conditions. The cost estimates do not however include costs for a winter road or additional indirect operating costs due to schedule extending delays.  There is a possibility that additional higher NSR value ore, in relation to typical NSR values for ROM pit ore, could be extracted from the underground mine during the early phase of the project.

1.18.2 Recommendations

Geotechnical

 It is recommended that a geotechnical assessment of the exposed development ramp and drifts be carried out as part of any remediation work following dewatering and updated as additional drifts are advanced/excavated in order to confirm the rock mass fabric and parameters used in these geotechnical studies.  The open pit and underground geotechnical analyses have been carried out based solely on results of previous investigation drilling and summaries, completed prior to 2005. It is recommended that these numerical analyses be re-visited as soon as the presently flooded underground workings have been pumped out and a geotechnical assessment of the exposed development and test stope completed and rock mass parameters confirmed.

Metallurgical and Processing

 Initiate and complete remaining process testwork and investigations.  Confirm concentrate feed rate and specification.  Complete a HAZOP review to assess the FEED design from the operations viewpoint.

It is recommended that the following items be further reviewed and discussed in the next round of detailed engineering studies:

 An opportunity exists for the replacement of the existing primary Jaw crusher with either a Sandvik hybrid crusher, or by engineered blasting. It is believed that the hybrid crusher option, as offered by Sandvik, could further reduce capital costs. However, since the hybrid crushing is relatively new to hard rock mining, sufficient due diligence and review must be performed.  There is also the opportunity to reduce the cost of the primary crusher by employing engineered blasting technique at the mine to decrease the size of the blasted ROM ore. More investigation is required at the next phase of the Project to determine if this technique is suitable.  Knelson modelled the potential gravity gold recovery of a conceptual stream that took on the characteristic of a blend of 50/50 P3 and P4 composite. The results obtained from the model should be considered very preliminary at this stage and P&E Mining Consultants Inc., Report No. 247 Page 21 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

further bench-scale work with representative drill core could be considered, in order to provide a more definitive prediction of the overall gold recovery benefit.  The 2007 pilot plant conducted at SGS Lakefield showed that the feed size to flotation can have an impact on bulk flotation performance. Further testwork is recommended to determine the optimal feed size to bulk flotation.  It is recommended that a sample of the bulk concentrate intermediate product is provided to the vendor for confirmatory testing for filter selection and sizing.  The recoveries in the second pilot plant were lower than the results from the FLEET model (first pilot plant). Additional tests are recommended to verify earlier pilot recovery results and regrind mill sizes.  A list of opportunities and recommendations to be further reviewed and discussed in detail engineering are briefly summarized below.

Regrind and Flotation

 A further review in DE should be conducted to optimize the size, quantity and configuration of the two 355-kW SMDs proposed for re-grinding.

Ion Exchange – Continuous Versus Fixed Bed

 Further testwork is recommended to assess continuous IX columns as previous testwork was on fixed bed IX columns. CIX columns could be more cost effective.

Design of Cobalt Electrowinning Circuit

 Further testwork is recommended to confirm adequate deposition of cobalt pucks at higher current density and higher cobalt electrolyte concentration.

Cyanide Destruction

Cyanide destruction has been specified by Fortune to complete a testwork program to evaluate cyanide destruction options from the SMPP plant. Cyanco Canada will obtain the autoclave residue and perform cyanidation testwork on the material.

 The most suitable cyanide destruction method was not determined by the release of this report destruction testwork by Cyanco Canada is in-progress and should be reviewed to determine the best methodology.

Other Opportunities and Recommendations

Further review is recommended on potential layout changes, vendor packages and other process engineering changes as provided in the list of recommendations to the client in the FEED report. Similarly, a complete review is recommended on the list of risk and opportunities provided to the client in the FEED report.

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Cobalt Sulphate Option

A scoping level study as described in Item 24 Other Relevant Data has been completed with provisional results from flowsheet modelling provided by SGS. When the final results and report have been received the flowsheet can be optimized and the changes implemented.

1.18.2.1 Bismuth Processing Plant (BiPP)

A review the BiPP design and equipment list was conducted to provide an interim report that identifies potential cost saving if certain process optimizations can be achieved. The following test and reviews are proposed:

 Leach Temperature Optimization testwork  Electrowinning testwork to optimize the cathode/anode current density.  Density optimization of the bismuth concentrate should be implemented if a supply of bismuth concentrate becomes available.  Further work should be conducted to determine if the iron levels in anolyte are acceptable for direct injection to the saline aquifer.  The electrical design philosophy should be reviewed for some potential cost savings with alternate suppliers, components and reducing some redundant processors and network.  In the detailed engineering a control estimate should be produced that will update the CAPEX (and OPEX) estimates to current costing

1.18.2.2 Cobalt Sulphate Option Recommendation

It is recommended that Fortune carry out the further engineering for the cobalt sulphate circuit along with supporting metallurgical testwork to improve the overall accuracy of the cobalt sulphate option cost estimates to bring the overall accuracy of the cobalt sulphate option cost estimate to ±15% accuracy and that the economic analysis be re-run before deciding to proceed with this option.

1.18.2.3 Mining

The underground mine will be developed and operated by a contractor. This underground work will have to be well coordinated with the open pit operation to attain overall anticipated tonnage and grade results.

1.18.2.4 Marketing

The long-term outlook for cobalt is generally positive. While fluctuations will always be a part of market prices, longer term cobalt pricing beyond 2012 is expected to move within the range of US$18/lb to US$22/lb. Bismuth supply is expected to remain stable during 2012 and 2013. Demand growth for bismuth is expected to be in the 8% to 10% range annually over the coming 3 years. Bismuth prices are forecasted to increase by US$1.50 in 2013 reaching $$14.00/lb, and then hitting $15.50/lb by the end of 2014.

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2.0 INTRODUCTION

2.1 SOURCES OF INFORMATION AND STUDY PARTICIPANTS

At the request of Mr. Robin E. Goad, President and CEO of Fortune, Jacobs acted as the consultant to prepare FEED studies for the development of the NICO Concentrator and related SMPP, with the following firms engaged by Fortune to provide specific project services as detailed below with inputs to the FEED and Technical reports related to their scope:

 Golder – responsible for geotechnical and hydrological parameters relevant to mine or pit design; plans for waste and tailings disposal; site monitoring; water management during operations and post mine closure; mine closure (remediation and reclamation) requirements and costs; environmental studies; and permitting requirements and status. They are the lead consultants for the environmental assessment (EA) process in the NWT.  EBA – responsible for design of the NICO access road and for geotechnical studies of infrastructure, excluding mine waste disposal facilities.  P&E – responsible for mine planning, production scheduling, equipment selection and reserve estimation.  DMA – responsible for metallurgical studies and process plant design for the Bismuth facilities at the SMPP.  SGS Lakefield – responsible for metallurgical testing and mineral processing research  MDH – responsible for the SMPP environmental baseline studies and site geotechnical work, the design of the process residue storage facility and is the lead consultant for the EA review process in Saskatchewan  Skybeco –responsible for metal market and product contract studies.

A number of additional engineering consultants contributed to specific parts of the grinding, flotation and hydrometallurgical circuit designs and the metal marketing analysis. All of the work was conducted to an accuracy of +/- 15%, except for the cobalt sulphate solvent extraction option, which was conducted with an overall accuracy of -10% to +25%.

2.2 QUALIFIED PERSONS AND RESPONSIBILITIES

The following Qualified Persons (“QP”) authored various sections or portions thereof of this technical report as detailed below. Some of the sections were co-authored by multiple QP and these sections are therefore repeated against the names of more than one QP:

 Tim Hayes P.Eng., employed by Jacobs as Project Engineer. Prepared technical contributions for infrastructure engineering and design of the Concentrator and Metallurgical Processing Plant.  Alex Duggan P.Eng., employed by Jacobs as Manager Estimating and Planning  Graham Holmes P.Eng employed by Jacobs as a Process Specialist.  Diogenes Uceda, P.Eng., employed by Jacobs as a Senior Process Engineer.  Wade Sumners P.Ag., P.Biol., employed by MDH as Senior Biologist.  Dan Mackie P.Eng., Vice President, principal and Senior Mechanical Engineer of DMA, provided project management, technical coordination, engineering and design of the BiPP for the SMPP.

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 Marc Rougier, P.Eng., employed by Golder as Senior Geological Engineer, Mining and Rock Engineering Division, is responsible for Item 16 (a) relating to geotechnical and hydrological parameters relevant to mine or pit design.  Ken Bocking, P.Eng., employed by Golder as Principal, Mine Waste & Environment Division, is responsible for: Item 20 (b) relating to plans for waste and tailings disposal, site monitoring and water management during operations and post mine closure, and Item 20 (d) relating to mine closure (remediation and reclamation) requirements and costs.  Alex Mezei P.Eng., employed by SGS Lakefield as Director, Engineering Technology Services.  Wayne D. Ewert P.Geo., Exec. VP of P&E responsible for Geology and overall report compilation.  Eugene Puritch, P.Eng., President of P&E responsible for Resource and Reserve Estimation and Pit Design.  Tracy Armstrong P.Geo., Senior Associate Geologist with P&E responsible for Geological Quality Assurance/Quality Control.  Fred H. Brown CPG, Pr.Sci.Nat.., Senior Associate Geologist with P&E responsible for Mineral Resource Estimate  David Orava, P.Eng., Senior Associate Engineer with P&E responsible for Open Pit Cost Estimating.  James L. Pearson, P.Eng., Senior Associate Engineer with P&E responsible for Underground Mine design and Costing and Reserve Estimate.  Bill Horne P.Eng., VP Environmental and Engineering Management for EBA Engineering Consultants Inc. (“EBA”), Provided input to Geotechnical studies on select Project Infrastructure items.

2.3 INDEPENDENT SITE INSPECTIONS

The following visits to either the SMPP or NICO Mine sites were carried out by the following Qualified Persons:

 Rod Lecher of International Quest Engineering visited the NICO property and the SMPP site from May 2008 to October 2009.  Marc Rougier of Golder visited the NICO property in September 2003 for three weeks and in September 2008 for three days as part of investigations to address the mining geotechnical and physical hydrogeology aspects of the project.  Ken Bocking of Golder visited the NICO property from 06 July 2010 to 07 July 2010 in order to examine the site of the Co-disposal Facility (CDF) at the NICO mine site.  Eugene Puritch of P&E, visited the NICO mine site on April 24, 2012. At that time a verification sampling program was undertaken.  Graham Holmes of Jacobs visited the NICO mine site on April 24, 2012.  Alex Duggan of Jacobs visited the NICO mine and SMPP sites during the period of April 17, 2012 through April 24 2012.  Wade Sumners of MDH visited the SMPP site on June 9, August 11, and September 24, 2010 to conduct a biological assessment for the EIA.

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2.4 SOURCES OF INFORMATION

This report is based in part, on internal Fortune technical reports and maps, published government reports, company letters and memoranda, and public information as listed in the "References” Section 27.0 at the conclusion of this report. Several sections from reports authored by other consultants have been directly quoted in this report, and are so indicated in the appropriate sections.

Specification data and related information used in engineering designs and scheduling work was obtained from select equipment and service providers as noted in the text and References section 27;

P&E has not conducted detailed land status evaluations, and has relied upon previous qualified reports, public documents and statements by Fortune regarding property status and legal title to the project.

The authors are pleased to acknowledge the helpful cooperation of Fortune‟s management and field staff, all of whom made any and all data requested available and responded openly and helpfully to all questions, queries and requests for material.

2.5 UNITS AND CURRENCY

All currency amounts are stated in Canadian or US dollars, as specified, with costs typically expressed in Canadian dollars (“$CAD”) and commodity prices in US dollars (“US”).Quantities are generally stated in SI units, the Canadian and international practice, including metric tons (“tonnes‟) or (“t”), kilograms (“kg”) and grams (“g”) for weight, kilometres (“km”) or metres (“m”) for distance, hectares (“ha”) for area, weight percent (“%”) for bismuth (“Bi”) and cobalt (“Co”) grades and grams per metric tonne (“g/t”) for gold grades (“g/t Au”). Precious metal grades may be expressed in parts per billion (“ppb”) or parts per million (“ppm”) and their quantities may also be reported in troy ounces (“ounces”) or (“oz”), a common practice in the mining industry.

Historical resource estimate tons and grade may be presented in the units short tons and troy ounces per short ton (“oz/ton”). Bismuth and cobalt prices are generally expressed on a per pound basis, the North American practice.

2.6 GLOSSARY AND ABBREVIATION OF TERMS

In this document, in addition to the definitions contained heretofore and hereinafter, unless the context otherwise requires, the following terms have the meanings set forth below.

“$” and “CAD$” means the currency of Canada. “AA” is an acronym for Atomic Absorption, a technique used to measure metal content subsequent to fire assay. “AANDC” means Aboriginal Affairs and Northern Development Canada “AGAT” means AGAT Labs in Mississauga, Ontario “asl” means above sea level. “Au” means gold. “AWR” means all-weather road “Azi” means azimuth. P&E Mining Consultants Inc., Report No. 247 Page 26 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

“BAMS” means biotite-amphibole magnetite schist “BFA” means Bench Face Angle “Bi” means Bismuth “CCME” means Canadian Council of Ministers of the Environment “CCRP” means Conceptual Closure and Reclamation Plan Angle “CDF” means Co-disposal Facility “CIM” means the “Canadian Institute of Mining, Metallurgy and Petroleum.” “CLER” means Electro Recovery “Co” means Cobalt “Cu” means Copper “DAR” means Developer‟s Assessment Report “DDH” means diamond drillhole. “DMA” means Dan Mackie Associates “DFS” means Definitive Feasibility Study “DSA” means dimensionally stable anodes “E” means east. “EBA” means EBA Engineering Consultants Ltd “el” means elevation level. “EIS” means Environmental Impact Statement “ETF” means Effluent Treatment Facility “FEED” means Front End Engineering Design study “Fortune” means Fortune Minerals Limited “g/t” means grams per tonne. “g/t Au” means grams of gold per tonne of rock “Golder” means Golder Associates Ltd. “GBMZ” means Great Bear magmatic zone “ha” means Hectare. “ICRP” means Interim Closure and Reclamation Plan “IOCG” means hydrothermal iron oxide copper-gold “IRA” means Inter-Ramp Angle “IX” means adsorption Ion Exchange “IRR” means Internal Rate of Return. “Jacobs” means Jacobs Mineral Canada Inc., formerly Aker Solutions a division of Aker Solutions Canada Inc. “kg” means kilogram. “km” means kilometre equal to 1,000 metres or approx. 0.62 statute miles. “LOM” means life of mine “m” means metric distance measurement equivalent to approximately 3.27 feet “M” means million. “Ma” means millions of years. “MDH” means MDH Engineered Solutions Corp. “MIBC” means methyl isobutyl carbinol “Micon” means Micon International Limited “MOE” means the Saskatchewan Ministry of Environment “Mt” means millions of tonnes. “MVEIRB” means Mackenzie Valley Environmental Impact Review Board “MVRMA” means Mackenzie Valley Resource Management Act “MVRB” means

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“N” means north. “NE” means northeast. “NI 43-101” means Canadian Securities Administrators National Instrument 43-101. “NN” means Nearest Neighbour “NPV” means net present value “NTS” means National Topographic System. “NW” means northwest. “NWT” means the Northwest Territories, Canada “OP” means open-pit “OSA” means Overall Slope Angle “oz/T” means Troy ounces per ton. “P&E means P&E Mining Consultants Inc. “PA/PEA” means a Preliminary Assessment or Preliminary Economic Assessment study. “PAX” means potassium amyl xanthate “Pocock” means Pocock Industrial Inc. “PRSF” means process residue storage facility “PSG” means Project Specific Guidelines “ppm” means parts per million. “Property” means. NICO Gold Cobalt-Bismuth-Copper deposit “QMS” means Quality Management System “QA”/”QC” means quality assurance/quality control “REE” means Rare Earth Elements “RO” means reverse osmosis “S” means south. “SAG” means semi autogenous grinding “SCPs” means Seepage Collection Ponds “SE” means southeast. “SEDAR” means the System for Electronic Document Analysis and Retrieval. “SG” means specific gravity “SGS” means SGS Lakefield Research Limited “Skybeco” means Skybeco Inc. “SMPP” means Saskatchewan Metals Processing Plant “SSWQOs” means Site Specific Water Quality Objectives “STP” means sewage treatment plant “SW means southwest. “t” means metric tonne equivalent to 1,000 kilograms or approximately 2,204.62 pounds “T” means Short Ton (standard measurement), equivalent to 2,000 pounds. “t/a” means tonnes per year. “tpd” means tonnes per day “TRC” means Technical Review Comments “UG” means underground “US$” means the currency of the United States. “UTM” means Universal Transverse Mercator. “VPSA” means vacuum pressure swing “W” means west. “WLWB” means the Wek‟èezhìı Land and Water Board “WQG” means water quality management “XPS” Xstrata Process Support Laboratories P&E Mining Consultants Inc., Report No. 247 Page 28 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

3.0 RELIANCE ON OTHER EXPERTS

The authors of this report state that they are QP‟s for those areas as identified in the appropriate “Certificate of Qualified Person” attached to this report. The authors have relied upon, and believe there is a reasonable basis for this reliance, the following experts and reports, who/which have contributed information regarding legal, land tenure, corporate structure, permitting, environmental and other issues in portions of this Technical Report in the Sections as noted below.

Section 4.0: Property and Description: Although copies of the abstracts for the mineral claims and licences, along with copies of the permits and work contracts were reviewed, an independent verification of land title and tenure was not performed. P&E has not verified the legality of any underlying agreement(s) that may exist concerning the licences or other agreement(s) between third parties.

The authors have relied, and believe that they have a reasonable basis to rely upon the land manager of Fortune who has contributed portions of the tenure, and legal status of the NICO project.

Section 19: The authors of this report have relied upon Mr. Marc Von Schwerin of Skybeco, for all information regarding metal marketing and product contracts.

P&E has reviewed the project database and used all reasonable diligence in checking, and verifying the data that was used in preparation of the NICO project mineral resource estimate. Ultimately, however, P&E has relied upon the authenticity of the data presented to it by Fortune.

A draft copy of the report has been reviewed for factual errors by Fortune. Any changes made as a result of these reviews did not involve any alteration to the basic conclusions presented by the authors. Hence, the statements and opinions expressed in this document are given in good faith and in the belief that such statements and opinions are neither false nor misleading at the date of this report.

P&E has relied upon:

 Procon Mining and Tunnelling Ltd, the underground mining contractor that developed the existing decline for the underground development and underground mining unit costs used in this study  Mr. Marc Von Schwerin of Skybeco in regard to the projected metal prices used in the discounted cashflow model presented in the economic analysis  Orica Canada Inc., of Stony Plain, Alberta in regard to preliminary open pit blasting patterns

The after-tax cashflow model utilized in the economic analysis presented in this Technical Report was developed by Fortune under the direction of Julian Kemp, Vice President Finance and CFO of Fortune. The cashflow model encompasses the development, operation and closure of the NICO open pit and underground mines, the NICO concentrator, bulk concentrate shipping, the SMPP facility and associated infrastructure. Fortune populated its cashflow model using information obtained from its consultants and other information provided by Fortune staff. Fortune ran its cashflow model to produce the economic analysis results presented in the present Technical Report. P&E have relied upon Mr. Kemp for the after tax cashflow model including P&E Mining Consultants Inc., Report No. 247 Page 29 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. the methodology utilized to estimate as examples net revenues from metal sales, depreciation and amortization, project financing / interest / leasing costs, available tax pools and taxation rates; and the use of the most recent approved capital and operating costs in the cashflow analysis. P&E had also assessed the suitability of Fortune Mineral‟s cashflow model and are of the opinion that Fortune Minerals‟ cashflow model is suitable for use in the present Technical Study. As examples:

 David Orava and James L. Pearson of P&E reviewed selected aspects of Fortune‟s cashflow model. David Orava of P&E reviewed and assessed the general approach that Fortune Minerals utilized to estimate revenues; handle and distribute estimated capital and operating costs, capital cost contingencies in the cashflow model; estimate and handle depreciation and taxes; deal with sunk costs; deal with salvage values; and estimate Project NPV, IRR and payback – this was done remotely and in face to face meetings with Julian Kemp and other Fortune staff. P&E also sampled selected aspects of the cashflow model such as the methodology used in the cashflow model to convert ROM gold grade expressed in grams/tonne to recovered troy ounces. David Orava and James L. Pearson of P&E also checked the open pit and underground mining capital and operating costs, and mine production schedule including scheduled ore tonnes and grades, and waste rock tonnages used in the cashflow model.  P&E utilized grade variable gold recoveries provided by Jacobs to estimate the recovered troy ounces of gold. Eugene Puritch of P&E checked the recovered troy ounces of gold utilized in the cashflow model.  James L. Pearson of P&E relied on an underground contract mine development and mining quotation dated 16 May 2012 received from Procon Mining and Tunnelling Ltd. The quotation provided the basis of the underground mining cost inputs used in Fortune Minerals‟ cashflow model.  The metal price projections used in the cashflow model were selected for use in the cashflow model by Fortune Minerals. The information base included but was not limited to trailing gold price information provided by Eugene Puritch of P&E; projected cobalt and bismuth price forecasting by Marc von Schwerin of Skybeco, and cobalt and market reviews by Merchant Research and Consulting Ltd. David Orava of P&E discussed Skybeco‟s metal price projections with the report‟s author by telephone on Friday, June 29, 2012 and confirmed that Skybeco‟s projections are suitable for use in the present study.  The P&E QPs for this Technical Report are not familiar with the proposed Nico concentrator and SMPP metallurgical processes and as such could not review the process recoveries and metal recoveries for cobalt, bismuth and copper utilized / reported in the cashflow model. Similarly, the P&E QPs for this Technical Report were not familiar with the capital and operating costs that Fortune Minerals received from its‟ geotechnical and environmental consultants such as Golder. David Orava, James L. Pearson and Eugene Puritch of P&E proactively participated in teleconferences with Fortune Minerals, Jacobs, and Golder that were held to check and confirm that Fortune had incorporated relevant metal recoveries, capital and operating costs into Fortune‟s cash flow model.  David Orava of P&E also re-estimated the results of the cashflow analysis (i.e. the NPV and IRR estimates) for the base case (cobalt metal) scenario by running a simplified cashflow model in parallel with Fortune Minerals‟ detailed cashflow model. The P&E simplified cashflow model utilized the revenues and costs reported in Fortune‟s cashflow and re-estimated the pre-tax and after-tax P&E Mining Consultants Inc., Report No. 247 Page 30 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

cashflows. P&E‟s basic cashflow model served as a check on the estimated pre- tax and after-tax cashflows and provided NPV and IRR estimates similar to the results of Fortune‟s cashflow analysis.

Metallurgical testwork programs were arranged by Fortune and completed by reputable facilities such as SGS Lakefield and others, Jacobs has relied upon the testwork results to develop their NICO process design.

To provide information shown in this 43-101 for the SMPP, MDH has relied on Rick Schryer at Fortune to supply the following items of information:

 Design limits for the PRSF and other on-site storage ponds;  Waste water volumes for the deep well injection disposal;  Process/brine solution characteristics for the deep well injection disposal;  Strategy for the anticipated decommissioning costs;  Plans for community engagement prior to the EIS approval and once the facility is operational;  Site layout to complete foundation design;  Desire to use groundwater as the SMPP water supply; and  Supply of process residue material from pilot plant operations.

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4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 LOCATION

The NICO property is located in the Mazenod Lake area in the NWT, approximately 160 km northwest of Yellowknife, 22 km west of the Snare hydro complex and 85 km north of the community of Behchokö. It is approximately centred at Latitude 63°33' N and Longitude 116°45' W, as shown on the NTS map sheet 85N/10.

Figure 4.1 Location Map of the NICO Property

4.2 PROPERTY DESCRIPTION AND TENURE

The NICO Property consists of 10 mining leases covering approximately 5,140 ha (Figure 4.2) held 100 % by Fortune.

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Figure 4.2 NICO Mining Leases

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As of August 9, 2007, Fortune holds 100 % of the interest in the NICO Property after it purchased the minority interest held by Candou Industries Inc. (“Candou”). Candou‟s interest in the Property was 9.2 % as at March 30, 2007 but was reduced to 8.0 % during the last quarter for non-participation in programs. Pursuant to the agreement, Fortune issued to Candou 1,000,000 common shares of Fortune and 100,000 common share purchase warrants exercisable at $3.00 per share for a term of five years from the date of issuance. Fortune also transferred its interest in and receivable from Formosa Environmental Aggregates Ltd to Candou (Fortune News Release, August 9, 2007).

The NICO Property originally comprised 12 contiguous exploration claims covering 6,622 ha. The claims, NICO 1 through to NICO 12, were staked in 1992, 1994 and 1995. Fortune had completed and filed sufficient assessment work to hold the claims in good standing until their 10 year anniversary date, at which time claims had to be taken to lease or abandoned. Two of the original claims were taken to lease in 2002 and eight others brought to lease in 2004, after two, (NICO 6 and 10), were allowed to expire. In order to bring claims to lease status they had to be surveyed. Accordingly, the leased claims have been surveyed and had monuments erected at their corners.

As of the effective date of this report, the 10 mining leases that comprise the NICO Property are in good standing (Table 4.1). An annual rental fee of $1,077 is due in July 2012 for leases 4237 and 4238 (NICO 2 and 1) and $11,623 in fees is due in October 2012 for the remaining NICO mining leases, totalling $12,700 in fees for 2012 to keep the leases in good standing.

TABLE 4.1 NICO PROPERTY MINING LEASES, EFFECTIVE JANUARY 2012 Annual Lease Mining Claim Area Lease Expiry Status Rental Fee Issue Date Lease No. Name (ha) Date ($1/acre) 4237 NICO 2 334 Lease $824 17/07/2002 17/07/2023 4238 NICO 1 102 Lease $253 17/07/2002 17/07/2023 4677 NICO 3 181.3 Lease $448 04/10/2004 04/10/2025 4678 NICO 4 323.34 Lease $799 04/10/2004 04/10/2025 4679 NICO 5 104.41 Lease $258 04/10/2004 04/10/2025 4680 NICO 7 698.08 Lease $1,725 04/10/2004 04/10/2025 4681 NICO 8 1,111.67 Lease $2,747 04/10/2004 04/10/2025 4682 NICO 9 827.18 Lease $2,044 04/10/2004 04/10/2025 4683 NICO 11 370.29 Lease $915 04/10/2004 04/10/2025 4684 NICO 12 1,087.39 Lease $2,687 04/10/2004 04/10/2025

Total 5,139.66 $12,700

4.2.1 Northwest Territories Claims and Mining Leases

The following subsection is taken from the Northwest Territories and Nunavut Mining Regulations Act, April 1, 2008.

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Once a mineral claim has been staked, the claim holder has 60 days to submit the following information to the Mining Recorder's Office:

 A completed Form 3, signed by the person or company in whose name the claim was staked;  A 1:50,000 sketch of the claim which must show any recognizable landmarks, where the claim is in relation to other claims, and the position of all posts;  The fee of $0.10 per acre; and  Any authorization from the surface rights holder (if applicable).

Before a mineral claim can be leased it must be surveyed by a Canada Land Surveyor. Adjoining claim holders must be notified and the surveyor or claim holder must supply proof that adjoining claim holders have been notified.

Application can made to lease a mineral claim if representation work of at least $10 / acre on the claim has been completed. There are fees of $25 / claim plus the first year rental of $1 / acre contained in the surveyed claim. The lease will be a term of 21 years with the annual rental payable yearly in advance. The lease is subject to renewal of further 21 year periods.

A lease is required if the intention is to sell or otherwise dispose of minerals or ore with a gross value of more than $100,000 in one year. A mineral claim will expire at the end of 10 years if the holder does not apply for a lease and the holder will be unable to re-stake the area for a period of one year.

As of the effective date of this Report, Fortune has maintained all of the required permits for exploration and related activities on the NICO property and land use, water use and quarrying permits appropriate to exploration activities have been continually renewed as required. The expiry date of the current land use permit is March 25, 2014.

4.2.2 Surface Rights and Land Agreements

Fortune reports that the NICO claims are located in an area on which a land claim agreement has been completed between the Tłîchǫ (First Nation) Government and the governments of Canada and became effective on August 5, 2005. This agreement established approximately 39,000 km2 of fee simple lands where the surface and subsurface rights are owned by the Tłîchǫ except for the “excluded” parcels listed in Part 2 of the Appendix to Chapter 18 of the Tłîchǫ Agreement. The Tłîchǫ Agreement also established Tłîchǫ self-government. The NICO claims were staked between 1992 and 1995, prior to this agreement. Fortune has advised P&E that the mineral rights conveyed by ownership of the claims are grandfathered with respect to this land claim. The surface and sub-surface rights and power line corridors outside of the excluded parcels are owned by the Tłîchǫ.

In November, 2011 Fortune and the Tłîchǫ Government signed a Co-operative Relationship Agreement for the NICO gold-cobalt-bismuth-copper project in the NWT. This agreement, which is similar to a Memorandum of Understanding, establishes the framework and path forward for further negotiations, defines primary liaison officials, and sets the communication protocol for the two parties. The agreement states that the Tłîchǫ Government and Fortune “wish to develop a co-operative relationship through which they will attempt to reach mutually beneficial agreement on matters affecting their respective interests.” (Fortune News Release, November 8, 2011) P&E Mining Consultants Inc., Report No. 247 Page 35 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

The Tłîchǫ Government and Fortune have also signed an Environmental Assessment Funding Agreement to support the Tłîchǫ Government with their review of the NICO Project‟s DAR. The environmental assessment is currently at the Information Request stage where stakeholders, including the Tłîchǫ Government, submit questions on the DAR. Fortune has received the first set of information requests from the Tłîchǫ Government for which responses are currently being prepared. The Tłîchǫ Government has also formed the Kwe Beh Working Group which manages the Tłîchǫ Government‟s involvement in the regulatory process of environmental assessment for the NICO project (Fortune News Release, November 8, 2011).

Fortune has been collecting baseline environmental data at the NICO site, along access routes and the surrounding area since 1998 as part of the first phase in the EA process. Studies have been conducted primarily by Golder and, as of March 30, 2009, were essentially complete. Ongoing environmental monitoring is being carried out. The environmental data was used in the formal EA process to outline baseline conditions, assess impacts, and develop mitigation strategies (Fortune News Release, March 30, 2009).

The MVRB is conducting an EA of the NICO project after it was referred to the Board by Aboriginal Affairs and Northern Development Canada in February, 2009. The Terms of Reference (“TOR”) for the EA were received later that year and the Company submitted its DAR in May, 2011 to begin the process of addressing the issues that had been identified. Fortune notes that the MVRB has completed its conformity check - no deficiencies were identified and the EA has advanced to the next phase.

4.2.3 Saskatchewan Metals Processing Plant

In order, to begin the EA process on the SMPP portion of the NICO Project Fortune held the last of its formal information sessions in the communities of Dalmeny and Langham, Saskatchewan on February 7 and 8, 2011 respectively. Subjects covered included the design and operation parameters of the plant, plans for residue storage, baseline environmental programs, and potential impacts and mitigation plans to minimize or eliminate impacts at the site and surrounding region. This follows the issuance of draft PSG‟s by the Environmental Assessment Branch of the MOE on January 21, 2011. The final PSG‟s outline the requirements of the provincial EA process and identified key issues that need to be addressed in the EIS (Fortune News Release dated February 17, 2011). An EA is now underway in Saskatchewan to permit the refinery. (Fortune News Release dated August 23, 2011). Fortune plans to maintain open communication with residents and elected officials from the relevant communities and surrounding region throughout the environmental assessment process and subsequent operations.

The Company has also filed its EIS with the Saskatchewan Government (Fortune news release, dated May 24, 2011).

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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 ACCESSIBILITY

Access to the Property, during the summer months, is via charter float plane or helicopter from Yellowknife and takes approximately one hour. In the winter, access is available by charter aircraft on skis or via 95 km of winter road maintained by the NWT Government (Figure 5.1). A short spur from this winter road was constructed by Fortune in 1996 to access Lou Lake and the NICO Property. NICO is located approximately 80 km north of the community of Behchokö.

Figure 5.1 Access Map for the NICO Property

Fortune is working with the federal, NWT and Tłîchǫ governments to engineer and construct an all-weather road to the communities and mine site. The existing winter road extending north from Behchokö on Highway 3 between Yellowknife and Edmonton, Alberta to the communities P&E Mining Consultants Inc., Report No. 247 Page 37 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. of Whatì and Gamètì is the subject of a government plan for realignment to an overland route. The construction of the all-weather road will provide year-round access to the community of Whatì and to the Snare River hydro power facility and would be the most likely route for construction and operating supplies to the property. Preliminary engineering and environmental scoping studies and preliminary consultations with Tłîchǫ have now been completed The NWT and federal governments committed $18 million of additional resources toward completing the next phase of the road, which includes additional engineering and environmental work, community consultation, re-alignment of the road to an all-land route, construction of permanent bridged water crossings and the laying of road bed on portions of the route. Approximately 25 km of all-season road has already been built by the community of Whatì and was to be extended by the community in 2009. Fortune is in discussions with the GNWT, federal and Tłîchǫ governments with respect to joint financing the upgrade of other segments of the route to an all-weather standard so that it can be used for mine operations. The next phase of this road initiative received significant support from Tłîchǫ citizens during the consultation process and the Tłîchǫ Government has approved the engineering and environmental studies now being tendered by the NWT Department of Transportation (Fortune News Release, August 5, 2009).

Access to Yellowknife from the south is available via an all-weather highway which connects the city of Edmonton and passes through Behchokö. The airport at Yellowknife offers multiple daily commercial flights to Calgary, Edmonton, Winnipeg and Ottawa by Canadian North and First Air. Yellowknife is the major access point to most of the smaller communities in the NWT.

5.2 CLIMATE

The climate is continental to subarctic, with short warm summers and long cold winters. Average summer high temperatures are in the 15°C range, while average winter temperatures range from - 15°C to -30°C, with a minimum of -45°C. Snow fall is moderate and the overall operating conditions do not present any unusual difficulties that would not have been encountered previously in the many former and ongoing mining operations in northern Canada.

5.3 LOCAL RESOURCES AND INFRASTRUCTURE

Yellowknife is currently a regional support and logistics for much of the mining activity in the NWT including the Ekati, Diavik and Snap Lake diamond mines. A pool of labour and support industries familiar with mining is locally available.

A semi-permanent camp with dining facilities consisting of trailers and all-wood office building are constructed on the east shore of Lou Lake. A permitted fuel depot and a steel-sided shop and maintenance building are also located on the property. These facilities were hauled in on the winter road.

During the spring and summer of 2006 and summer of 2007, Fortune drove an underground ramp down into the NICO deposit and collected a bulk sample from the core of the mineralization. The surface infrastructure, adjacent to the decline, includes the camp and eight double-walled fuel storage tanks.

5.4 PHYSIOGRAPHY

The physiography of the NICO Property is moderately rugged, with steep slopes and local relief ranging up to 200 m. Absolute elevation ranges from 150 to 350 m asl. Positive land forms are P&E Mining Consultants Inc., Report No. 247 Page 38 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. characterized by abundant outcrop and sparse vegetation of jack pine and alder, while lowland areas are covered by lakes, muskeg and glacial drift with black spruce, jack pine, poplar, birch, alder, grass, moss and lichen. The NICO deposit is located on the northern slope of a bowl- shaped depression known as the „Bowl Zone‟.

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6.0 HISTORY

The following section is summarized and updated from the previous technical report on the Property by Thalenhort and Farquharson (2002), Hennessey and Puritch (2004) and Hennessey et al (2007).

Early indications of mineralization in the NICO area were obtained by local prospectors in the 1930‟s and then by New Athona Mines Ltd. which explored the CAB claim group from 1968 to 1970. The CAB claim group was staked to cover two cobalt-bismuth-copper-arsenide showings. Exploration by New Athona included geological mapping, electromagnetic and magnetic geophysical surveys, trenching and approximately 4,636 ft (1,413 m) of diamond drilling in 21 holes. This work led to a “drill-indicated resource” totalling 214,540 tons, averaging „3.24 pounds of bismuth with lesser values in cobalt, copper and gold‟. Chemical analysis of a bulk sample of massive arsenopyrite yielded 2.36 % Co, 0.63 % Bi, 22.46 % Fe, 16.01 % S, 40.84 % As, 0.18 oz/ton Ag and 0.14 oz/ton Au (Bryan 1981, 1982). It appears that New Athona investigated the near-surface mineralization in the volcanic cover rocks of the NICO deposit.

In 1977 and 1978, Eldorado Nuclear Limited conducted exploration for uranium in the area (Thomas and Olson 1978), and Noranda Exploration followed up the New Athona work from 1978 to 1989 as part of a larger exploration effort that also included work on the Sue-Dianne copper deposit located approximately 25 km north-northwest of NICO.

Fortune acquired the NICO Project, comprising 12 staked mineral claims, during 1992-1994. The Bowl Zone, the principal mineralized occurrence at NICO, was discovered by Fortune in 1994 as the result of geologic surface work. The program was based on the concept that the general area represented a geological target comparable to the Olympic Dam deposit in Australia. The Bowl Zone was drilled systematically for the first time in 1996, with additional drilling in 1997 and 1998.

A. H. Mumin (1997, 1998a, 1998b), a consulting geologist retained by Fortune, prepared mineral resource estimates in 1997 and 1998, based on the available drill hole data at each time. The second estimate (Mumin 1998b) included the results of initial metallurgical test work at Lakefield (Lakefield 1997a, 1997b, 1998). These, together with preliminary geotechnical and environmental investigations by Golder (1998), were used in a scoping study by Kilborn SNC- Lavalin (“Kilborn”) (1998) which evaluated the economic merits of the project. The 1998 drilling program was not incorporated into the Kilborn scoping study.

The 1998 Kilborn scoping study evaluated three production rates over a range of cobalt prices from $US 5/lb to $US 30/lb and was based on total mineral resources of 88.6 Mt with average cobalt, gold and bismuth grades of 0.07 % Co, 0.54 g/t Au and 0.08 % Bi. Of this total, 50 Mt with average grades of 0.10 % Co, 0.92 g/t Au and 0.11 % Bi were determined to be the mineable portion of the resources for an open-pit operation with a waste to ore stripping ratio of 2.7. The study envisaged on-site flotation of ore to produce a sulphide concentrate followed by an acid-leach pressure-oxidation process and the recovery of metallic cobalt using solvent extraction-electrowinning (“SX/EW”).

After the 1998 drilling program, SNC-Lavalin Engineers & Constructors (“SNC-Lavalin”), the successor company to Kilborn was retained to validate and verify the drill hole data base, including the assay and specific gravity data, and to provide a geologic interpretation of the Bowl Zone and its mineralization. The second SNC-Lavalin study was comprehensive. Using all of the P&E Mining Consultants Inc., Report No. 247 Page 40 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. available drill hole data in 1999, SNC-Lavalin also produced an estimate of the mineral resources for the Bowl Zone at a declared pre-feasibility level of accuracy. The unclassified mineral resources were estimated to be 39.6 Mt with average grades of 0.08 %Co, 0.41 g/t Au and 0.10 % Bi at a cut-off grade of 0.06 % Co (SNC-Lavalin 1999).

In September, 1999 Fortune retained Strathcona Mineral Services Limited (“Strathcona”) to update the earlier SNC-Lavalin scoping study and to conduct new mineral resource estimates based on the geologic model that had been developed by Fortune and validated by SNC-Lavalin. Strathcona agreed with most of the SNC-Lavalin conclusions pertaining to the database and basic geological interpretation but prepared a new estimate of the mineral resources based on a more constrained geological model. Strathcona‟s study was also based on additional metallurgical test work, open pit optimization studies and a preliminary economic evaluation of the project that excluded on-site processing beyond the flotation concentrate stage.

Strathcona (2000) made a number of recommendations including conducting additional field programs and studies that were conducted in 2000 and 2001, and the results were summarized in Thalenhort and Farquharson (2002). They included:

An additional 6,300 m of in-fill drilling (33 holes) was done on the NICO property and the surface geology was remapped and tied into the drill grid. An open pit mineral resource using a cobalt price of $US 7.50 per pound was estimated at 34 Mt grading 0.08 % Co, 0.12 % Bi, and 0.4 g/t Au and a waste-to-ore ratio of 1:7.

Electron microprobe analyses revealed that the cobalt is contained in an arsenopyrite (FeAsS)- cobaltite (CoAsS) solid solution series (122 grains analyzed; Lakefield 2001). This observation had an impact on metallurgy and the related metal production, because as the proposed bulk sulphide flotation was essentially arsenopyrite flotation, the arsenic to cobalt ratio was determined the cobalt grade of the cobalt concentrate.

A simple and low-cost flow sheet was further developed in metallurgical test work at a laboratory scale for the production of a bulk sulphide flotation concentrate followed by separation into cobalt and bismuth concentrates. The cobalt concentrate would have a cobalt content of 2.0% to 7.0%, dependent on the local arsenic-to-cobalt ratio of the sulphides in the deposit. Cobalt recovery was predicted at 85% and not sensitive to ore grade or to the arsenic-to- cobalt ratio.

Flotation test work was conducted to determine the efficiencies of producing separate bismuth and cobalt concentrates on site with a view to determining the contribution of bismuth to ore value. A bismuth sulphide concentrate grading 45% Bi and recovering 55% of the bismuth appeared possible but needed further confirmation on a larger scale. It became apparent that both the bismuth and cobalt concentrates carry recoverable and payable amounts of gold that increase with depth in the deposit.

The existing autoclave at the Con Mine in Yellowknife was identified as a potential treatment facility for the NICO cobalt concentrate. Subsequently, it was determined that this option was not valid.

Both the SNC-Lavalin (1999) and Strathcona (Thalenhort and Farquharson 2002) mineral resource estimates were based solely on relatively large scale open pit mining methods. In October, 2002, Eugene Puritch, P.Eng., was retained by Fortune to assist in an in-house update P&E Mining Consultants Inc., Report No. 247 Page 41 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. of the mineral resource estimate which envisaged a combination of underground and open pit mining methods.

In December 2002, Micon was initially engaged to provide an independent review of this work. However, it was realized that some additional drilling would be required to fully evaluate the new proposal. In 2003, a 4,720 m drill program consisting of 33 diamond drill holes was completed, which focused mainly at the extremities of the then proposed East and West pits as well as two holes drilled to test unexplored geophysical anomalies. Drilling in the West pit area better defined the extreme up-plunge and up-dip extents of mineralization and provided data for arsenic / cobalt modelling. Drilling in the Central part of the deposit better defined, and increased sample density, within the higher grade gold-rich core and allowed better modelling of the felsic dyke waste rocks. Drilling in the East pit area further extended the eastern limits of known mineralization and extended the strike length of potentially mineable resources (Hennessey and Puritch 2004).

In 2003, Golder conducted geotechnical engineering, hydrogeology, environmental, and archaeological surveys for feasibility assessment of the NICO deposit (Golder, 2003a, 2003b, 2004a and 2004b). The geotechnical information serves as a complement to data previously collected by Golder (1999) to investigate ultimate slope conditions for a proposed single, large open pit. The environmental surveys provide additional information to the earlier Golder (1998) study, which included a baseline environmental summary and an aquatics survey. On September 29, 2003, a 1:18,000 scale aerial photographic survey was flown by Eagle Mapping Ltd. over a 7 km by 10 km area of the NICO claims in order to establish a digital topographic map at 1:2,000 scale with 2 m contour intervals. Subsequent to the mineral resource estimate described in Hennessey and Puritch (2004), the decision was made to proceed with a bankable feasibility study. In conjunction with this, further environmental, geotechnical and metallurgical work was undertaken. None of this affected the block model produced in 2004 but resulted in a mineral reserve estimate being determined from that model.

In late 2006, following completion of the 2004 mineral resource estimate, three more infill diamond drill holes were drilled, totalling 517.85 m. All three holes were drilled into areas of insufficient drill hole density and all intersected Au-Co-Bi mineralization. Notable results include hole NICO-06-286 which intersected 1.36 g/t Au and 0.16 % Co over 24.05 m (Fortune News Release, December 6, 2006). The intercept extended the high grade mineralization another 50 m in the central part of the deposit where previously it was thought to have terminated. All of the drill hole intercepts approximated the true width of mineralization. The three holes were completed late in the 2007 feasibility study process and were not used in the mineral resource or mineral reserve estimates presented in Hennessey and Puritch (2004) and Hennessey et al. (2007) technical reports.

The exploration program in 2006 and 2007 included the extraction of an underground bulk sample and compositing of two samples, totalling 200 t for future use in the 2007 pilot plant testing program (Fortune News Release, December 4, 2007, Hennessey et al. 2007). For detailed information on the 2006-2007 bulk sampling the reader is referred to Section 15 of this Report.

6.1 PREVIOUS MINERAL RESOURCE AND RESERVE ESTIMATES

Mineral Resource Estimates were previously prepared by Mumin in 1997 and 1998, SNC- Lavalin in 1999 and Strathcona in 2000 (Thalenhort and Farquharson 2002), as well as an updated, in-house estimate prepared in 2002. Two of these estimates were accompanied by P&E Mining Consultants Inc., Report No. 247 Page 42 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. scoping studies and preliminary economic evaluations. The most recent previous mineral resource and reserve estimate is outlined below.

It needs to be emphasized that the Mineral Resource Estimates noted below in section 6.1 of this report are historical in nature, and as such, are based on prior data and reports prepared by previous operators. The Mineral Resources have not been verified by P&E and the Mineral Resources therefore, cannot be treated as NI 43-101 defined resources verified by a qualified person. The historical Mineral Resource Estimates should not be relied upon, and there can be no assurance that any of the Mineral Resources, in whole or in part, will ever become economically viable.

6.2 MICON MINERAL RESOURCE AND RESERVE ESTIMATE, 2007

The additional information from the 2003 drill campaign was added to the drill hole data base for use in the calculation of the mineral resource estimate by Micon prepared in 2004 (Hennessey and Puritch 2004).

At metal prices of US$10/lb Co, US$375/oz au, US$3.25/lb Bi and a CDN$/US$0.72 exchange rate, Micon used an NSR cut-off grade of CDN$50/t for reporting underground mineralization and CDN$20/t for reporting the open pit mineralization of the NICO project (Table 6.1). The block model was reported using these cut-offs to produce the mineral resources, which were first reported in Hennessey and Puritch (2004).

TABLE 6.1 NICO MINERAL RESOURCES SUMMARY, 2007 Measured & Indicated NSR Cut-off Mining Area Au Bi Co As/Co NSR $CDN/t Tonnes (g/t) (%) (%) Ratio ($CDN/t) Open Pit $20 8,231,000 0.48 0.136 0.131 10.0 34.33 Underground $50 5,123,000 3.44 0.210 0.160 6.3 79.40

Total 13,354,000 1.62 0.164 0.142 8.6 51.62 Note: It should be noted that although all mineral resource estimates quoted above are considered by Micon as NI 43-101 compliant, that the data has not been originally sourced or verified by P&E.

The 2007 mineral reserve estimate used different cost, commodity price and exchange rate assumptions than were used in the 2004 mineral resource in order to determine NSR values (Hennessey et al. 2007). Mineral reserves for the open pit and underground mining operation were determined based upon operating costs estimated for the annual production rate of 1,460,000 t of ore, metallurgical recovery values determined from testing, metal prices of $US15.00/lb Co, $US500/oz Au, $US4.00/lb Bi, and a $CDN/$US0.844 exchange rate. Mining cut-off limits were determined as NSR values based on the above parameters.

The mineral reserves resulting from the Micon bankable feasibility study presented in Hennessey et al. (2007) are reported in Table 6.2 below.

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TABLE 6.2 MINERAL RESERVE ESTIMATE, MICON 2007 Classification Tonnes Au (g/t) Bi (%) Co (%) Open Pit Mineral Reserves (Cut-off $32.21/t NSR) Proven Mineral Reserve 7,058,000 1.142 0.160 0.114 Probable Mineral Reserve 13,555,000 0.698 0.158 0.131

Total Mineral Reserve 20,613,000 0.850 0.159 0.125

Underground Mineral Reserves (Cut-off $77.13/t NSR) Proven Mineral Reserve 231,000 5.318 0.126 0.133 Probable Mineral Reserve 973,000 5.006 0.200 0.147

Total Mineral Reserve 1,204,000 5.066 0.186 0.144

Total Mineral Reserves Proven Mineral Reserve 7,289,000 1.274 0.159 0.115 Probable Mineral Reserve 14,528,000 0.987 0.161 0.132

Total Mineral Reserve 21,817,000 1.083 0.160 0.126 Note: It should be noted that although all mineral reserve estimates quoted above are considered by Micon as NI 43-101 compliant, that the data has not been originally sourced or verified by P&E.

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7.0 GEOLOGICAL SETTING AND MINERALIZATION

The following section is primarily sourced from the previous technical report on the Property by Hennessey et al. (2007) and scientific papers on the Property by Goad et al. (2000a, 2000b).

7.1 REGIONAL GEOLOGY

The NICO deposit occurs in the southern part of the Proterozoic Bear Structural Province which is further subdivided into the Wopmay Orogen and the Amundsen Basin (Fraser et al. 1972) (Figure 7.1). The Great Bear magmatic zone (“GBMZ”), which consists of a Paleoproterozoic belt calc-alkaline volcanic and plutonic rocks (1,880 to 1,840 Ma) exposed from Great Slave Lake in the south to Great Bear Lake in the north, forms the central tectonic zone of the Wopmay Orogen. The GBMZ formed during eastward subduction of an oceanic plate beneath the Slave craton and the accreted volcanoplutonic Hottah Terrane and now occupies the suture zone between them (Hildebrand et al. 1987). The GBMZ along with the Hottah Terrane, are juxtaposed against peralkaline intrusions and sedimentary rocks of the Coronation Margin along the 10 km wide mylonitic crustal suture, the Wopmay Fault Zone (Goad et al. 2000b). Subaerial volcanics of the GBMZ overlap both the Hottah Terrane and Coronation margin and are intruded by plutonic rocks of a similar age (Hildebrand et al. 1987). The Wopmay fault is the boundary between the Slave craton and the Hottah Terrane and also marks the eastern limit of the GBMZ which covers the Hottah Terrane for a distance of approximately 90 km westward.

The GBMZ is approximately 800 km long and 100 km wide and consist of low titanium oxide and high alumina calc-alkaline volcano-plutonic rocks. Felsic to intermediate rocks of the 1.87 to 1.84 Ga Faber Group, predominate in the southern part of the GBMZ, consist of rhyodacite ignimbrites and associated flows, tuffs, breccias and volcaniclastics and are bordered by granodiorite to monzogranite plutons and intruded by coeval rapakivi granite and feldspar porphyritic plugs (Goad et al. 2000a).

The Coronation margin is comprised of the Snare, Akaitcho and Epworth Groups which formed as continental margin shelf and slope sediments as a result of rifting along the margin of the Slave craton (Hoffman 1973, 1980). The Snare Group consists of arenite, dolomite, siltstone and shale. A tectonic shift to eastward subduction beneath the Hottah Terrane was proposed by Goad et al. (2000a) to account for the formation of the GBMZ and the three stages of deformation and metamorphism recognized in the Coronation Margin. The final deformation postdates plutono- volcanic activity within the GBMZ and gave rise to conjugate transcurrent faults with subordinate normal and reverse faults (Hildebrand and Bowring 1984).

The southernmost Faber Group unconformably overlies metasedimentary basement rocks that were initially considered to be Snare Group, but are now re-classified as the Treasure Island Group by the Geological Survey of Canada. They are preserved in supracrustal keels separating dominal granite plutons. The unconformity at the base of the volcanic rocks is strongly potassium and iron metasomatized and commonly brecciated, whereas contacts between granitic plutons and metasediments are mylonitic. The „dome and keel‟ basement and mylonitic detachment faults are indicative of post-collisional extended rift terrains and are domed due to the diapiric rise of late A-type granites (Goad et al. 2000a). Goad et al. (2000a) proposed that iron concentrations may originate from the generation of dry potassic melts in response to crustal thinning and doming during rifting, with iron enrichment derived from melting of iron-rich crust possibly iron-rich members of the Treasure Island Group.

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Figure 7.1 Regional Geology Map

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7.2 PROPERTY GEOLOGY

The NICO deposit is hosted in iron- and potassium-altered, brecciated basement sedimentary rocks of the Treasure Island Group at and beneath the unconformity with the volcanic Faber Group rocks.

NICO‟s Co-Au-Bi-Cu mineralization is located within locally altered biotite-amphibole magnetite schist (“BAMS”) of the Treasure Island Group (Figure 7.2). The metasediments of the Treasure Island Group are the oldest rocks in the area and consist of dominantly subarkosic wacke, arenite and minor siltstone and carbonate and are unconformably overlain by a north- dipping succession of rhyolite to rhyodacite tuffs, flows and minor volcaniclastics of the Faber Group. These rocks are bound between the GBMZ granite to the southwest and monzogranite of the Marian River Batholith to the northeast (Goad et al. 2000a). Treasure Island Group sedimentary rocks are strongly hornfelsed marginal to the GBMZ granites.

Some of the widely developed intermediate to felsic volcanics of the Faber Group are interpreted as high-level sill-type intrusive rocks rather than extrusive rocks due to their stratigraphic location below, rather than above, the metasediments (Hennessey et al. 2007). The basal, potassium-feldspar altered rhyolite yields an approximate U-Pb date of 1851 +18/-16 Ma (Gandhi et al. 1996). A felsic intrusive, compositionally identical to the Faber Group volcanic unit except for 10-15 % mm-sized plagioclase crystals, appears to postdate both the Treasure Island and Faber Groups (Hennessey et al. 2007).

Breccias are common, particularly in the Treasure Island Group immediately below the unconformity. The breccias have been interpreted variously as fragmentals and as hydrothermal diatreme breccias formed in a near-surface environment (Goad et al. 2000b). The breccias are spatially related to zones of polymetallic sulphide mineralization, and capped by massive potassium-feldspar altered rhyolite (felsite) along the unconformity. For the most part, the breccias are composed of clasts of the Treasure Island Group with lesser felsite clasts in a matrix of iron oxides, biotite, amphibole, chlorite and potassium feldspar (Goad et al. 2000a). The breccias contain minor sulphide concentrations despite their proximity to sulphide bearing mineralized lenses with similar hydrothermal alteration. The breccias have been interpreted by Goad et al. (2000a) as maar-facies breccias suggesting volcanism was initiated by near surface diatreme activity. The proximity of these breccias to mineralization and the presence of iron oxide and potassium metasomatism suggest that formation of diatreme and maar breccia was coeval with sulphide mineralization (Goad et al. 2000a).

The Treasure Island and Faber Groups are cross-cut by a quartz-feldspar porphyry dyke parallel to the strike of the Treasure Island Group and a younger feldspar-amphibole ± quartz porphyry dyke, emplaced parallel to the strike of the Faber Group at Lou Lake. These dykes, along with dykes discovered in drill core, are considered feeders to the overlying volcanics (Goad et al. 2000a).

Late northeast striking (40°) transverse faults transect the Snare and Faber Groups and adjacent intrusives. These faults merge into the Wopmay Fault and are thought to be related splays (Goad et al 2000a). Major, regional faults trend 70°. Large-scale quartz veins have been emplaced locally along both fault directions. A thick, sporadically mineralized and persistent quartz vein, similar to the vein hosting the Rayrock uranium deposit to the south of the Property, transverses east-northeast through Champion‟s mining lease 4678 (NICO 4).

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Figure 7.2 Property Geology Map

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7.3 MINERALIZATION

The following section draws heavily from the previous technical report on the Property by Thalenhort and Farquharson (2002) and Hennessey et al. (2007).

NICO‟s Au-Co-Bi-Cu mineralization is intimately associated with a regional scale metasomatic event and a later more restricted sulphide mineralizing event. Both the Treasure Island Group metasediments and the overlying Faber Group volcanics were subjected to intense regional potassium and iron metasomatism. In the NICO area, the combination of iron and potassium metasomatism has resulted in the creation of an assemblage of amphibole-biotite ± magnetite altered metasedimentary rocks which extend along strike for approximately 2 km. The deposit is hosted within a 200 m thick package of northwest striking and northeast dipping amphibole- biotite ± magnetite ironstone and schist and amphibole-biotite altered subarkosic wacke. The latter is considered to be the protolith to the ironstone/schist and becomes increasingly more abundant in the upper part of the hanging wall. The dominant amphiboles are iron-rich.

Sulphide mineralization is disseminated and makes up between 3 to 10 % of the mineralized rocks. The sulphides are predominantly aligned along the foliation planes. Only small native gold grains have been observed, and are mainly associated with sulphides but also with silicate minerals such as feldspar (Thalenhort and Farquharson 2002). The sulphides consist primarily of cobaltite/cobaltian arsenopyrite, bismuthinite and chalcopyrite.

Walker (1999) suggested the following mineralizing stages based on detailed mineralogical studies:

 Magnetite is the earliest oxide  Pyrrhotite and pyrite and the earliest sulphides  Introduction of gold / bismuth telluride, chalcopyrite and native bismuth. Bismuthinite is interpreted to occur from stages 3 through to 6  Precipitation of cobaltite  Introduction of colbaltian arsenopyrite  Introduction of massive arsenopyrite (with minor cobalt) clots and veins which appear to overprint all previous sulphides  Formation of hematite

The occurrence of fracture-filled native gold in both stage 4 cobaltite and stage 6 arsenopyrite is thought to be due to some post-stage 3 crystallization of gold.

Sulphide mineralization of economic significance is younger than the iron metasomatism and is restricted to certain domains within the biotite-amphibole-magnetite schist (“BAMS”) and amphibole-altered wacke. These sulphide-bearing domains are relatively more restricted in volume and together comprise the Bowl Zone. Figure 7.3 illustrates a cross-section through the Bowl Zone.

Figure 7.3 also illustrates the extent of cobalt and bismuth mineralization within the BAMS. Gold mineralization forms a central „bulls-eye‟ to the deposit within the cobalt-bismuth core of the magnetite mineralization and is confined largely to the middle and lower zones. Two correlatable horizons of less altered subarkosic wacke partially demarcate the lower / middle and middle/upper zone boundaries (Thalenhort and Farquharson 2002). The upper zone is much smaller and the sulphide and Au-Co-Bi mineralization is weaker. P&E Mining Consultants Inc., Report No. 247 Page 49 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Figure 7.3 Typical Cross Section through the Bowl Zone (at 20+50W)

F1 F5

UPPER ZONE

MIDDLE ZONE

Cobalt (Co, grey) 00224S annotated above 0.5% Bismuth (Bi, magenta) annotated above 0.5%

97074 LOWER 00225S ZONE 97075 metres Gold (Au) annotated 0 25 98185S above 5g/tonne 98193

A minor amount of mineralization occurs in the overlying felsic extrusive/intrusive Faber Group units, and is hosted by east-west striking, sub-vertical, vein-like structures. The arsenopyrite in these near vertical zones was most likely introduced during Walker‟s (1999) stage 6 mineralizing event. Sulphide intersections also occur in the crosscutting quartz-feldspar and feldspar- amphibole ± quartz porphyry dykes. P&E Mining Consultants Inc., Report No. 247 Page 50 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Many of the late stage felsic intrusions cutting the deposit (seen as the orange unit in Figure 7.3) have „removed‟ Co-Bi-Au mineralization and left blocks of relatively unmineralized rock in the centre of the deposit.

Petrographic evidence shows that the majority of cobalt in the Bowl Zone is contained in the arsenopyrite-cobaltite solid solution series, the rest in cobaltite proper and minor amounts in the minerals modderite and cobaltian loellingite (Goad et al 2000b, SGS Lakefield 2001). NICO‟s cobalt content is noted to be dependent on the relative abundances of both cobaltite and danaite in the ore and in the concentrate, and on the danaite composition, the latter being the most important factor because of the dominance of danaite in the NICO deposit (Thalenhort and Farquharson 2002). Walker (1999) reported that the third phase of mineralization started with cobaltite, continuing with cobalt-rich varieties of danaite and ending with cobalt-poor danaite or arsenopyrite. An arsenic/cobalt ratio zonation in the NICO deposit has been recognised with a systematic decrease from high As-Co ratios in the upper part of the deposit to lower ratios at greater depths (Thalenhort and Farquharson 2002).

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8.0 DEPOSIT TYPES

The NICO Au-Co-Bi-Cu deposit can be classified as a hydrothermal iron oxide copper-gold (“IOCG”) deposit. The NICO deposit and Fortune‟s Sue-Dianne Cu-Ag deposit, located 25 km to the northwest of NICO, are the only known significant IOCG deposits currently defined in Canada. The GBMZ hosts the NICO and Sue-Dianne as well as IOCG prospects, showings, occurrences and alteration zones in or adjacent to Andean-type, 1.85-1.88 Ga, calc-alkaline, basaltic to rhyolitic caldera-fill complexes and stratovolcanoes, diatremes, and coeval felsic to intermediate epizonal plutons (Corriveau 2007).

8.1 IRON OXIDE COPPER-GOLD (IOCG) DEPOSITS

The IOCG deposit type encompasses a wide spectrum of sulphide-deficient low-Ti magnetite and / or hematite ore bodies of hydrothermal origin where breccias, veins, disseminations, and massive lenses with polymetallic enrichments. These hydrothermal deposits are associated with large-scale continental A- to I-type granitic suites with intermediate and mafic facies, alkaline- carbonatite stocks, crustal-scale fault zones, regional sodic-calcic alteration, focused potassic and iron oxide alteration, and coincident aeromagnetic and gravity highs. The deposits typically have more than 20% iron oxides (Corriveau 2007).

Although Phanerozoic deposits exist, the most important IOCG deposits are Early to Middle Proterozoic in age (Goad et al. 2000a). The deposits are situated in anorogenic cratonic settings with rifting, and are usually located on major structural lineaments which were likely extensional and/or transcurrent faults related to extensional rifting. Deposits are proximal and located preferentially in the roof zones of megacrystic syenogranite intrusions, which may display unusual myrmekitic, granophyric and rapakivi textures (Goad et al. 2000a). Rocks that host IOCG deposits formed in regionally oxidized settings through which fluids could flow and/or react.

Because of the diversity of mineralization, there has been debate as to whether IOCG deposits form a single deposit type or whether they are iron oxide-rich variants of other deposit types.

IOCG subtypes have been proposed but due to the diversity of iron oxide Cu-Au, U, Ag, REE, Bi, Co deposits there is the possibility of many potential subtypes. Corriveau‟s (2007) review on Canadian IOCG deposit types utilized Gandhi‟s (2004) classification of six subtypes defined for the World Minerals Geoscience Database Project (Figure 8.1).

The NICO deposit has been classified as both a Cloncurry subtype (Goad et al. 2000a) and an Olympic Dam subtype (Gandhi 2004).

Regionally, cratonic rift basins are characterized commonly by positive Bouguer gravity and total field magnetic trends. The associated granites and related volcanic rocks are usually rich in K and U and will generate positive radiometric anomalies if exposed at surface. The intersection of regional-scale, structural lineaments related to rifting with plutonism, is important in localizing deposits; these lineaments can be detected as linear-magnetic and very-low frequency electromagnetic anomalies. Iron-rich alteration dominated by magnetite are characterized by strongly positive magnetic anomalies, those dominated by hematite can be identified by a relatively low-intensity anomaly (Goad et al. 2000a).

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Figure 8.1 IOCG Deposit Subtypes

(Source: Corriveau, 2007 after Gandhi, 2004)

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9.0 EXPLORATION

Since 2008, the majority of the work done on the property was focused on development, mainly the permitting and financing of the NICO mine site. Metallurgical test work has also been conducted which is elaborated upon in Section 13. The only exploration work carried out on the Property has been diamond drilling which is summarized in Section 10.

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10.0 DRILLING

In the summer of 2010, Fortune began a 38-hole drill program at the NICO site. The main objectives of the program were to extend the known mineral reserves for the deposit and to test for extensions to the known deposit where it was locally open for expansion near the surface, at the deposit ends and also at depth. In particular, the gold-rich central core of the deposit was open for possible extension to depth and also between some broad spaced drill hole intersections. A summary of significant intersections, along with high grade intervals over 2 m long and over 2 g/t Au, are presented in Table 10.1.

P&E is not aware of any drilling, sampling, or recovery factors that could materially impact the accuracy and reliability of the results.

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TABLE 10.1 HIGHLIGHTS OF DRILL INTERCEPTS FROM 2010 DRILL PROGRAM True Borehole ID Grid East (m) Grid North (m) Az Dip(o) From (m) To (m) Interval (m) Width Au (g/t) Bi (%) Co (%) Cu (%) (m) NICO 10-300 24+00W 346 N 200 -56 52.00 85.00 33.00 32.41 - 0.28 0.07 0.01 111.71 163.00 51.29 50.37 2.22 0.04 0.11 0.08 including 121.00 148.00 27.00 26.51 3.99 0.04 0.12 0.08 and 122.00 124.00 2.00 1.96 7.51 0.14 0.15 0.25 and 129.00 131.00 3.00 2.95 15.59 0.05 0.46 0.11 and 135.00 148.00 13.00 12.77 4.93 0.03 0.16 0.06 and 136.00 139.00 3.00 2.95 15.59 0.05 0.46 0.11

NICO 10-299 19+00W 352 N 200 -90 228 232 4.00 2.82 4.68 0.01 - - including 228.00 230.00 2.00 1.41 10.60 0.02 - -

NICO 10-298 19+50W 343 N 200 -90 32.82 33.82 1.00 0.71 0.12 - 0.12 - NICO 10-297 18+00W 206 N 200 -70 96.00 97.00 1.00 0.91 0.34 0.01 0.11 0.01 100.00 101.00 1.00 0.91 0.12 0.01 0.11 0.01

NICO 10-296 17+00W 120 N 200 -73 25.44 27.50 2.06 1.82 0.10 - 0.14 - 32.55 33.55 1.0 0.88 0.89 - 0.29 -

39.74 43.65 3.91 3.45 0.65 0.01 0.10 0.01

NICO 10-295 17+00W 208 N 200 -65 98.09 99.10 1.01 0.95 0.23 - 0.12 0.01 NICO 10-294 17+50W 205 N 200 -53 64.84 72.98 8.14 8.06 1.53 0.01 0.07 0.01 65.86 67.88 2.02 1.99 4.50 0.02 0.14 0.01

117.43 118.43 1.00 0.99 0.50 - 0.69 0.03

NICO 10-293 18+50W 191 N 200 -45 102.24 122.35 20.11 20.11 0.01 0.54 0.31 - NICO 10-292 25+50W 358 N 200 -45 67.00 71.00 4.00 4.00 0.01 0.34 0.06 0.01 85.00 91.00 6.00 6.00 - 0.15 0.08 0.02

95.00 108.00 13.00 13.00 0.01 0.37 0.38 0.01

137.50 148.00 10.50 10.50 - 0.17 0.11 0.01

NICO 10-291 25+50W 358 N 200 -70 160.18 168.20 8.02 8.02 0.03 0.12 0.04 0.01 NICO 10-301 26+50W 413 N 200 -45 85.03 89.04 4.01 4.01 - 0.09 0.18 0.01 121.03 126.05 5.02 5.02 0.01 0.08 0.15 -

NICO 10-302 26+50W 367 N 200 -45 91.51 103.56 12.05 12.05 - 0.11 0.12 0.01 NICO 10-303 27+50W 355 N 200 -45 44.00 46.00 2.00 2.00 0.01 0.07 0.25 - 78.00 82.00 4.00 4.00 - 0.11 0.13 0.01

95.00 97.00 2.00 2.00 - 0.10 0.15 0.01

NICO 10-305 27+50W 272 N 200 -45 56.00 58.00 2.00 2.00 0.25 0.11 0.10 0.02 NICO 10-309 26+50W 189 N 200 -45 34.64 40.57 5.93 5.93 1.01 0.01 0.19 0.02 NICO 10-312 13+50W 48 N 200 -45 118.47 124.84 6.37 6.37 0.02 0.01 0.13 - NICO 10-316 16+50W 20 N 200 -45 35.61 37.17 1.56 1.56 0.18 - 0.22 0.02 NICO 10-317 14+50W 131 N 200 -55 84.00 86.00 2.00 1.97 0.27 - 0.14 0.01 150.54 159.52 8.98 8.84 1.93 0.21 - -

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TABLE 10.1 HIGHLIGHTS OF DRILL INTERCEPTS FROM 2010 DRILL PROGRAM True Borehole ID Grid East (m) Grid North (m) Az Dip(o) From (m) To (m) Interval (m) Width Au (g/t) Bi (%) Co (%) Cu (%) (m) including 150.54 157.55 7.01 6.90 2.39 0.18 - - NICO 10-318 15+00W 115 N 200 -50 111.00 115.00 4.00 3.98 0.24 - 0.11 - NICO 10-320 21+50W 135 N 200 -45 56.15 57.00 0.85 0.85 0.01 0.09 0.17 0.01 98.00 98.94 0.94 0.94 0.97 0.02 0.16 0.01 NICO 10-321 24+50W 137 N 200 -45 17.00 22.00 5.00 5.00 1.27 0.05 0.14 0.01 32.00 35.00 3.00 3.00 1.24 0.01 0.10 - 38.00 41.00 3.00 3.00 1.60 0.01 0.08 - 46.00 49.00 3.00 3.00 0.59 0.01 0.12 0.01 NICO 10-322 25+50W 170 N 200 -45 22.00 24.00 2.00 2.00 0.12 0.05 0.12 - 45.00 47.00 2.00 2.00 0.77 0.02 0.16 - NICO 10-323 21+50W 419 N 200 -65 19.00 22.19 3.19 3.00 0.21 0.04 0.15 0.24 46.74 48.10 1.36 1.28 0.06 0.01 0.21 0.04 59.27 59.97 0.70 0.66 0.40 0.02 0.16 0.78 151.00 152.00 1.00 0.94 1.03 - 0.01 0.15 164.70 165.70 1.00 0.94 0.56 - 0.15 0.01 168.50 170.50 2.00 1.88 0.20 - 0.12 0.01 175.20 179.00 3.80 3.57 0.20 - 0.09 - 225.67 226.50 0.83 0.78 2.72 0.02 - - 234.80 236.80 2.00 1.88 0.26 - 0.22 0.01 260.70 263.62 2.92 2.74 0.02 0.26 0.21 0.02 NICO 10-324 20+00W 443 N 200 -60 23.00 25.00 2.00 1.93 0.19 0.04 0.11 0.31 71.06 78.93 7.87 7.60 0.14 0.10 0.09 0.11 146.62 150.00 3.38 3.26 11.59 0.16 0.37 0.14 222.00 223.00 1.00 0.97 2.65 0.06 0.09 - 268.00 270.00 2.00 1.93 0.03 0.10 0.13 - NICO 10-325 21+00W 443 N 200 -65 10.40 11.40 1.00 0.94 2.55 0.40 0.06 0.42 25.44 26.74 1.30 1.22 0.64 0.03 0.13 0.26 189.00 197.00 8.00 7.52 4.74 0.16 0.01 - 226.00 227.00 1.00 0.94 2.75 0.58 - - 260.00 274.62 14.62 13.74 0.33 0.01 0.22 - NICO 10-326 20+50W 433 N 200 -65 141.72 142.80 1.08 1.01 0.21 0.12 0.19 - 148.00 149.00 1.00 0.94 1.64 0.03 0.01 0.06 NICO 10-327 22+50W 400 N 200 -65 15.00 16.00 1.00 0.94 0.14 0.05 0.10 0.27 34.00 36.00 2.00 1.88 0.40 0.03 0.15 0.05 73.08 74.57 1.49 1.40 0.34 0.01 0.29 0.01 86.00 92.92 2.92 2.74 0.25 0.19 0.20 - 133.00 138.00 5.00 4.70 1.86 - 0.03 0.03 166.00 171.00 5.00 4.70 4.84 0.01 0.04 0.01 192.00 194.00 2.00 1.88 0.25 - 0.37 0.03 202.00 204.50 2.50 2.35 9.21 0.02 0.01 0.01 231.30 233.57 2.27 2.13 0.56 0.08 0.17 0.01 P&E Mining Consultants Inc., Report No. 247 Page 57 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

11.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY

The following section draws heavily upon material presented in the 2007 technical report by Hennessey et al. (2007) for the Micon feasibility study; A consistent sampling, analytical and security methodology as summarized below has been employed by Fortune throughout their various drilling programs on the NICO project

11.1 CORE LOGGING

Prior to sampling, all core was logged at the Lou Lake camp located on the NICO claims. The following items are checked or recorded by the logging geologists:

 Check blocking of all core  Convert feet measurements to metres  Consolidate core and line up fractures and joints  Note recovery and rock quality designation  Describe rock strength of all lithologies  Test all core with a magnet to determine areas of absent, weak, moderate or strong magnetism  Describe the principal lithologies present and their locations  Lay out sample intervals (generally 1-m samples used but with minor variances allowing for lithological breaks and missing core). Earlier programs had used longer sampling intervals, generally 2 to 3 m and, very occasionally, up to 6-m or more in length for samples that were unlikely to contain significant metal enrichment, but which required analytical verification  Measure and describe the weak to strong magnetic areas by sample  Describe the sulphides, structure and hematite alteration present in each sample interval

Since the spring of 2000, all lithologic logs have contained a coded percentage of each principal lithology within each sample interval. Sample intervals were also categorized according to the amount of sulphide mineralization within the interval.

The logging was performed, or in some cases overseen, by Fortune geologists, principally Kathryn Neale, Miroslav Sidor and Derek Mulligan, all of whom were/are long term consultants to Fortune and familiar with most phases of the project.

11.2 CORE SAMPLING

Sampling of the drill core was conducted by a technician supervised by the respective logging geologists. All sampling was performed in separate shed adjacent to the logging facility at the Lou Lake camp.

All drill core samples were split for sampling and one half was assayed. Most of the core was sampled, except for some of the post-mineralization dykes. Samples that were logged as weakly, moderately or strongly mineralized core were split using a diamond blade saw. Trace mineralized core was split with a conventional guillotine type, knife blade splitter, in order to minimize costs.

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Once split, samples were bagged individually in heavy plastic bags and then placed, in numerically ordered groups, into large „rice bags‟ for shipment. Samples were shipped to Yellowknife by float plane in summer or pickup truck in winter, where they were palletized and shipped by transport truck to the assay laboratory.

Intervals which were expected to contain negligible grade (trace mineralization) were shipped as separate 1 m samples with instructions to composite the pulps from groups of consecutive samples for analysis. If gold grades greater than 300 ppb, or cobalt grades greater than 0.05 % were returned, separate pulps for each sample were analyzed. No compositing was performed if any noticeable amount of arsenical sulphide was present.

Once sampled, the core boxes and remaining core were taken to an outdoor core storage area where they were stacked on large timbers in piles of approximately 10 boxes. The timbers were used to get the boxes well off of the ground to promote ventilation and to prevent rot.

11.3 SAMPLE PREPARATION

All sample preparation and primary assaying of drill core from the 1996 to 2000, 2003, 2006 and 2010 programs was performed at ALS Chemex Canada Limited in North Vancouver (“ALS Chemex”). ALS Chemex is an ISO accredited laboratory with its quality assurance system in place at its laboratories complying with the requirements of the international standards ISO 9001:2000 and ISO 17025:1999 http://www.alsglobal.com/Mineral/ALSContent.aspx?key=66).

Fortune employed ALS Chemex‟s CRU-31 crushing, SPL-21 splitting and PUL-31 pulverizing protocols (together called the PREP-31 package) for its preparation of samples from NICO. For the PREP-31 protocol a sample is dried and the entire sample is crushed to better than 70% passing a 2 mm (Tyler 10 mesh) screen. A sub sample of up to 250 g is taken with a Jones-type riffle splitter and pulverized to better than 85% passing a 75 micron (Tyler 200 mesh) screen using a puck and bowl pulverizer.

11.4 ANALYSIS

ALS Chemex reports that Fortune regularly receives its analyses by the Au-AA23, As-AA46, Cu-AA62a, Co-AA62, and Bi-AA46 analytical methods for gold, arsenic copper, cobalt and bismuth respectively. Over limits for gold are rerun by Au-GRA21. The information on methodology employed in these methods, as presented by ALS Chemex, is summarized in the sections below. The Cu-AA62a method is performed in essentially the same manner as the Cu AA62 method (as outlined in the ME-AA62 description below) however it has a lower detection limit which is achieved through the instrument curve set up for this method.

11.4.1 Method Au-AA23/AU-AA24

Sample Decomposition Fire Assay Fusion Analytical Method Atomic Absorption Spectroscopy (“AAS”)

A prepared sample pulp is fused with a mixture of lead oxide, sodium carbonate, borax, silica and other reagents as required, inquarted with 6 mg of gold-free silver and then cupelled to yield a precious metal bead.

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The bead is digested in 0.5 ml of dilute nitric acid in a microwave oven, 0.5 ml concentrated hydrochloric acid is then added and the bead is further digested in the microwave at a lower power setting. The digested solution is cooled, diluted to a total volume of 4 ml with demineralized water, and analyzed by atomic absorption spectroscopy against matrix-matched standards. The Au-AA23/Au-AA24 methods are summarized below in Table 11.1.

TABLE 11.1 ALS CHEMEX METHOD AU-AA23/AU-AA24 SUMMARY Lower Upper ALS Chemex Sample Element Reporting Reporting Units Method Code Weight (g) Limit Limit Au-AA23 Gold 30 0.005 10.0 ppm Au-AA24 Gold 50 0.005 10.0 ppm

11.4.2 Method Au-GRA21/Au-GRA21

Sample Decomposition Fire Assay Fusion Analytical Method Gravimetric

A prepared sample pulp is fused with a mixture of lead oxide, sodium carbonate, borax, silica and other reagents in order to produce a lead button. The lead button containing the precious metals is cupelled to remove the lead. The remaining gold and silver bead is parted in dilute nitric acid, annealed and weighed as gold. Silver, if requested, is then determined by the difference in weights. The Au-GRA21 and Au-GRA 22 methods are summarized below in Table 11.2.

TABLE 11.2 ALS CHEMEX METHOD AU-GRA21/AU-GRA22 SUMMARY Lower Upper ALS Chemex Sample Element Reporting Reporting Units Method Code Weight (g) Limit Limit Ag-GRA21 Silver 30 5 10,000 ppm Ag-GRA22 Silver 50 5 10,000 ppm Au-GRA21 Gold 30 0.05 1000 ppm Au-GRA22 Gold 50 0.05 1000 ppm

11.4.3 Method ME-AA46 (includes As-AA46 and Bi-AA46)

Sample Decomposition Aqua Regia Digestion Analytical Method AAS

A prepared sample pulp (0.4 to 2.00 g) is digested with concentrated nitric acid for one half hour. After cooling, hydrochloric acid is added to produce aqua regia and the mixture is then digested for an additional hour and a half. An ionization suppressant is added if molybdenum is to be measured. The resulting solution is diluted to volume (100 or 250 ml) with demineralized water, mixed and then analyzed by atomic absorption spectrometry against matrix-matched standards. The ME-AA46 method is summarized below in Table 11.3.

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TABLE 11.3 ALS CHEMEX METHOD ME-AA46 SUMMARY ALS Chemex Method Code Element Detection Limit Upper Limit Units As-AA46 Arsenic 0.01 30 % Bi-AA46 Bismuth 0.001 30 % Cd-AA46 Cadmium 0.001 10 % Co-AA46 Cobalt 0.01 50 % Cu-AA46 Copper 0.01 50 % Fe-AA46 Iron 0.01 30 % Pb-AA46 Lead 0.01 30 % Mo-AA46 Molybdenum 0.001 10 % Mn-AA46 Manganese 0.01 50 % Ni-AA46 Nickel 0.01 50 % Ag-AA46 Silver 1 1,500 ppm Zn-AA46 Zinc 0.01 30 %

11.4.4 Method ME-AA62 (includes Cu-AA62a and Co-AA62)

Sample Decomposition HNO3-HClO4-HF-HCl digestion Analytical Method AAS

A prepared sample pulp (0.2 to 2.0 g) is digested with nitric, perchloric, and hydrofluoric acids, and then evaporated to dryness. Hydrochloric acid is added for further digestion, and the sample is again taken to dryness. The residue is dissolved in nitric and hydrochloric acids and transferred to a volumetric flask (100 or 250 ml). The resulting solution is diluted to volume with demineralized water, mixed and then analyzed by atomic absorption spectrometry against matrix-matched standards. The ME-AA62 method is summarized below in Table 11.4.

TABLE 11.4 ALS CHEMEX METHOD ME-AA62 SUMMARY ALS Chemex Method Code Element Lower Reporting Limit Upper Reporting Limit Units Ag-AA62 Silver 1 1000 ppm Al-AA62 Aluminium* 0.01 50 % Ca-AA62 Calcium* 0.05 50 % Cd-AA62 Cadmium 0.0001 10 % Co-AA62 Cobalt 0.001 30 % Cu-AA62 Copper 0.01 50 % Fe-AA62 Iron 0.01 30 % K-AA62 Potassium* 0.01 30 % Li-AA62 Lithium 0.01 50 % Mg-AA62 Magnesium* 0.01 50 % Mn-AA62 Manganese* 0.01 50 % Mo-AA62 Molybdenum 0.001 10 % Na-AA62 Sodium* 0.001 30 % Ni-AA62 Nickel 0.01 50 % Pb-AA62 Lead 0.01 30 % Sr-AA62 Strontium 0.01 20 % V-AA62 Vanadium 0.01 30 % Zn-AA62 Zinc 0.01 30 % *Elements reported as oxide

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11.5 OTHER METHODS

Fortune reports that arsenic values were only determined from samples collected in 1996, 2000, 2003, 2006 and 2010. For the 1996 sampling, a geochemical method with an upper limit of 10,000 ppm (or 1 %) was used whereas in 2000, 2003, 2006 and 2010, the arsenic values were determined in the manner described above.

11.5.1 Specific Gravity Measurements

A large number of specific gravity (“SG”) measurements based on rock type have been taken at the NICO project and utilized in the mineral resource estimation process.

It is the author‟s opinion that the sample preparation, security and analytical procedures are satisfactory.

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12.0 DATA VERIFICATION

12.1 SITE VISIT AND INDEPENDENT SAMPLING

The NICO deposit was visited by Mr. Eugene Puritch, P. Eng. Who is a QP for the purposes of NI 43-101. Mr. Puritch visited the site on July 10 and 11, 2004 and on April 24, 2012, and conducted a detailed site review including recording positions of drill hole collars, examination of the core logging facilities and practises as well as the collection of core samples for independent data verification. A total of six samples from two diamond drill holes were collected and analyzed for bismuth, cobalt, copper and gold. See Figure 12.1, to Figure 12.3.

Figure 12.1 NICO Deposit Site Visit Sample Results for Bismuth

Figure 12.2 NICO Deposit Site Visit Sample Results for Cobalt

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Figure 12.3 NICO Deposit Site Visit Sample Results for Copper

Samples were collected by quarter cutting the remaining half core in the box. They were given a unique sample number, placed in a bag and once all samples were collected they were placed into a larger bag and taken by Mr. Puritch to the offices of P&E in Brampton, ON. From there the samples were sent by courier to AGAT Labs in Mississauga, ON for analysis.

 AGAT Labs has developed and implemented at each of its locations a Quality Management System (“QMS”) designed to ensure the production of consistently reliable data. The system covers all laboratory activities and takes into consideration the requirements of ISO standards.  AGAT maintains ISO registrations and accreditations. ISO registration and accreditation provides independent verification that a QMS is in operation at the location in question. Most AGAT laboratories are registered or are pending registration to ISO 9001:2000.  Samples were analyzed for bismuth, cobalt and copper using a 4-acid digest and ICP-ICP/MS finish. Gold was analysed using fire assay, ICP/MS. Neither the original NICO analysis nor the P&E analyses returned above detection limit for gold, therefore the graph is not included in this section.

12.2 QUALITY ASSURANCE/QUALITY CONTROL

The quality assurance/quality control (“QA”/“QC”) practices for the 2010 NICO Infilling Program consisted of the insertion of alternating blanks and standards approximately every ten samples. In total, two (2) different blanks and four (4) different standards were used. Fortune monitored the QC on a real-time basis and the following paragraphs are taken from the Fortune report. The author of this section has independently verified all QC data, as reported below.

12.2.1 Fortune Au-Co Standards

Three different standards were prepared by SGS Minerals for previous drilling programs for Fortune using material procured from the NICO site itself and prepared by SGS Lakefield. For the purposes of this document, they have been named S1 (0.77 ppm Au; 0.066% Co), S2 (0.70 ppm Au; 0.51% Co), and S3 (1.47 ppm Au; 0.19% Co). P&E Mining Consultants Inc., Report No. 247 Page 64 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

For S1 and S2 the results generally fell within two standard deviations of the mean. The values that were found slightly beyond two standard deviations did not occur within mineralized zones. One S1 standard returned a value significantly different from acceptable values and did not seem to match those values of any other QC sample used. Seven samples surrounding the failed standard were rerun and the new results were used in the master database.

All four S3 standards were within two standard deviations.

12.2.2 Fortune Minerals Blank

Blank material was also prepared for previous drilling programs for Fortune using material procured from the NICO site and prepared by SGS Lakefield. All samples were within two standard deviations, apart from one sample which did not lie within a mineralized zone.

CDN Resource Laboratories Ltd. Au-Cu Standard (CDN-CGS-20)

A gold-copper standard (CDN-CGS-20) was prepared for Fortune by CDN Resource Laboratories. For the purposes of this document and report, it was named S4 (7.75+0.47 ppm Au; 3.36+0.17% Cu). The results generally fell within two standard deviations of the mean, and those values which were found slightly beyond two standard deviations did not occur within mineralized zones.

CDN Laboratories Blank

A blank (CDN-BL-7) was prepared for Fortune by CDN Resource Laboratories. While most values fell within two standard deviations, two values reported slightly high, and another failed.

One of the slightly elevated blanks falls within a mineralized zone, and indicates slight carry over contamination. There is no impact to the database. The other slightly elevated blank did not lie within a mineralized zone.

Blank G225200 was significantly elevated in gold, with a value of 0.06 g/t Au. It did not lie within a mineralized zone, and therefore no action was taken.

The authors state that the data were robust and satisfactory for use in a resource estimate.

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13.0 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 NICO SUMMARY DESCRIPTION

The following sections describe briefly the NICO concentrator process, as well as the testwork program that was carried out previously for the development of both the NICO and SMPP mineral processing circuits.

The mineral processing plant has a throughput of 1 695 060-t/a ROM ore, which is the equivalent of 193.5 t/h at 100% availability. With an operating availability design criteria of 90%, the plant is designed for processing 215 t/h of throughput.

The majority of the design criteria are based on the information from the pilot testwork performed at SGS Lakefield in 2008 and in 2010 to 2012, the locked cycle and FLEET testwork performed by SGS Lakefield in 2009, and supplied by Fortune. Each criterion has been referenced to provide the source of the information. The design criteria and recovery results from the bankable feasibility study (“BFS”) completed by Micon in 2007 is also used as a comparative tool for cross-checking results. The overall cobalt, bismuth, gold and copper recoveries at the NICO concentrator plant for the FEED study are 90.9%, 82.1%, 72.6% and 89.1% respectively. The individual stage recovery criteria recoveries are summarised in Table 13.1.

TABLE 13.1 STAGE AND OVERALL PLANT RECOVERIES Stage Recovery Overall Recovery

Co Bi Au Cu Co Bi Au Cu Concentrator Plant Feed 100 100 100 100 100 100 100 100 Gravity Conc. - - 10.2 - - - 7.8 - Bulk Flot Conc. 90.5 80.8 68.3 88.2 90.5 80.8 63.0 88.2 Secondary Flot Conc. 8.9 13.7 16.1 13.0 +0.5 +1.3 +1.8 +0.9 Bulk and Secondary Flot Tails 9.0 17.9 27.4 10.9

Overall Recovery 90.9 82.1 72.6 89.1

13.2 ORE CHARACTERISTICS

13.2.1 Ore Grade Metal Content

The annual head grade of the ore and grade variation through the mine life was obtained from the NICO Production Schedule - Open Pit and Underground, dated December 14, 2009. The production schedule projects the quarterly average grade of Bi, Co, Au, Cu, as well as the As / Co ratio expected for the first 15 years of the Project.

The averaged metal head grades for the first 15 years of the Project, the maximum head grades, and the grades used as the plant design, are given in Table 13.2.

The As / Co ratio is known to be a significant parameter in relationship to the cobalt content in this ore, because arsenic occurs naturally with cobalt in cobaltite and in arsenopyrite. The As /

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Co ratio varies from a high of 11.6 in the early few years of the mining plan, to a low of 3.6, with the average at 6.64. An As / Co ratio of 7.0 was selected for plant design.

TABLE 13.2 NICO RUN-OF-MINE METAL HEAD GRADES COMPARISON Bismuth Cobalt Gold Copper Grade Arsenic Grade

Grade (%) Grade (%) Grade (g/t) (%) (%) Average 0.165 0.127 0.961 0.039 0.845 Maximum 0.311 0.219 3.819 0.091 1.491 Plant Design - Selected 0.190 0.145 2.386 0.034 1.002

Based on these head grades and the recovery from the bulk and secondary flotation circuits at the NICO concentrator, the average bulk concentrate grades are summarized in Table 13.3. The As / Co ratio of 7.0 remains unchanged in the bulk concentrate.

TABLE 13.3 BULK CONCENTRATE METAL GRADES Bismuth Cobalt Gold Copper Arsenic Grade Grade Grade Grade Grade (%) (%) (g/t) (%) (%) Plant Design Grades 4.33 3.66 44.28 0.84 25.61

13.2.2 Mineralogy

The mineralogical structure of the NICO ore determines the metal extraction processes used. Most of the NICO ore consists of non-opague minerals, magnetite and arsenopyrite with minor to trace amounts of pyrite chalcopyrite, sphalerite, pyrrhotite, native bismuth, bismuthinite, cobaltite, hematite, goethite and a Bi-Cu sulphosalt. Non-opague minerals are mainly amphibole, feldspars, biotite, quartz and possibly pyroxene.

An economically important metal in the NICO ore is cobalt. The principal cobalt carrier is arsenopyrite (FeAsS). Cobalt replaces some of the iron in the mineral lattice. The arsenopyrite is relatively coarse grained and liberated at a K80 of 72 microns. The cobalt content of the arsenopyrite is low in the ore closer to the surface and rises with increasing depth.

Bismuth is present as native bismuth and bismuthinite. The bismuth minerals are ≤ 20 microns in size.

The gold grade is higher when underground mining is in operation and lower with open pit mining operation. The grain size ranges from 1 to 50 microns. Gold recovery is feed grade dependent and decreases with lower feed grades. The recovery used in the FEED design is more realistic when the high grade gold is mined but it is expected to decrease when the lower grade gold is mined.

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13.2.3 Physical Properties

Bulk Density

The specific gravity of the dry solids in the ore material is approximately 3.3. Including the 6.5 wt% entrained moisture in the ore, the SG is 3.39. The estimated bulk density of the unconsolidated, wet ore material is 2.4 t/m3.

Abrasion

The abrasion index for the ore has been measured twice, 0.3299 g and 0.4173 g were the results.

Compressive Strength

In 2005, Golder performed Unconfined Compressive Strength tests on a variety of samples. Table 13.4 presents a summary of the results.

TABLE 13.4 COMPRESSIVE STRENGTH RESULTS Rock Type UCS (MPa) Dyke Felsic 204 Dyke Quartz Feldspar Porphyry 129 Footwall Meta-Siltstone 119 Ore Zone Meta Greywacke 127 Ore Zone Black Rock 75 Hanging Wall Rhyolite 238

In August 2010, a series of ore zone samples was tested at Queens University with values ranging from a low of 55.7 MPa to a high of 212 MPa and averaging 123.3 MPa.

13.2.4 Work Indices

Crushing Work Index

The crusher work Index was not measured.

JK Dropweight

The JKTech Dropweight test was performed on two samples and both gave results of Axb of 22 which is a very hard ore.

MacPherson Test

One MacPherson grindability test was done in 1999. The result indicated a power consumption in the semi-autogenous grinding (“SAG”) mill of 23.3 kWh/t, which is indicative of a very hard ore.

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SAG Design

A series of 10 SAG Design tests were run on different ore composites. The results ranged from a SAG pinion energy requirement of 12.58 kWh/t to 29.64 kWh/t, with an average of 22.33 kWh/t. The design point recommended at the 80th percentile was 29.5 kWh/t, which is again indicative of a very hard ore.

Rod Mill Work Index

A rod mill work index was measured in 1999 at 17.8 kWh/t. Two further tests were done in 2004 and yielded results of 17.5 kWh/t and 20.2 kWh/t. The pilot plant run in 2008 and 2010 results were 19.0 kWh/t and 20.6 kWh/t, respectively.

Ball Mill Work Index

In 1999, a Bond ball mill test was performed which gave a figure of 16.1. In 2004, several more tests were done, giving results between 10.5 and 13.6. In 2005, a further test was performed, giving a result of 13.0. In the 2008 and 2010 pilot plant work, the results were 13.0 and 14.3, respectively.

13.2.5 Materials Handling

Wall Friction (mass flow angle)

Testwork performed by Jenike & Johanson (“J&J”) indicated that, after 24 hours at rest, for both moisture contents tested (3.4% and 7.5%) the maximum recommended wall angle was 10 degrees from the vertical for flows in chutes and bins.

Valley / Slope Angles

Testwork performed by J&J to determine the minimum chute angles necessary for flow indicated that with mild steel plate and 7.5% moisture an angle of 52 degrees from the horizontal was required. Mention was also made that the ore is somewhat sensitive to impact pressure and that, therefore, the transfer heights should be minimised.

Bulk Ore Angle of Repose

The design bulk ore angle of repose was 37%.

ROM Ore Size

The ROM ore has a top size of 500 mm and a particle size of 80% passing (P80) of 300 mm.

Moisture

For the FEED study, the design moisture content of the ore is 6.5%.

Bulk Concentrate Density

The bulk density of the concentrate filter cake was assumed to be 2.1 t/m3. P&E Mining Consultants Inc., Report No. 247 Page 69 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

13.2.6 Mineralization

Whole Rock and Trace Metal Analysis

Table 13.5 and Table 13.6 summarizes the whole rock and mineralogical composition for the ROM ore used for the FEED design. The numbers in the table represent a blend of the open-pit (OP) and underground (UG) ore.

TABLE 13.5 PLANT DESIGN ROM ORE COMPOSITION Blend (OP + UG) Co % 0.145 Au g/t 2.386 Bi % 0.190 Cu % 0.034 S % 0.55 As % 1.002 Ni % 0.003 Zn % 0.002 Fe % 19.4

TABLE 13.6 PLANT DESIGN ROM ORE MINERALOGY COMPOSITION Mineral Species ROM Ore1 (vol %) CoAsS Cobaltite Trace Au Elemental Gold Au2Bi Bi Native Bismuth 0.2 Bi2S3 Bismuthinite 0.2 Fe3O4 Magnetite 10.0 Fe2O3 Hematite 1-2 FeOOH Goethite < 0.2 FeAsS Arsenopyrite 7.0 FeS2 Pyrite < 0.1 FeS Pyrrhotite < 0.1 CuFeS2 Chalcopyrite < 1.0 ZnS Sphalerite < 0.2

SGS Lakefield mineralogical examination of a Co-Bi Met Comp Head Sample from the NICO Deposit Report LR10226-001, April 27, 2001.

13.3 SULPHIDES

The ROM is a sulphide ore. Sulphide sulphur is mostly associated with arsenopyrite (FeAsS), pyrite (FeS2), cobaltite (CoAsS), and chalcopyrite (CuFeS2).

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Arsenic Cobalt Ratio

As discussed previously, an As / Co ratio of 7.0 was selected for plant design.

Iron to Arsenic Ratio

To effectively precipitate arsenic in the cobalt autoclave and subsequent iron arsenic precipitation steps, the iron in the cobalt stream needs to exceed the arsenic to form a stable FeAs compound scorodite.

13.4 LABORATORY BENCH-SCALE AND PILOT PLANT TESTING

13.4.1 Bulk and Bismuth Flotation

The first series of tests were carried out in collaboration with SNC-Lavalin in 1997 to 1998. Flotation was aimed at the production of low-grade concentrates. A second test program was completed by September, 2001. In this series, the target was production of saleable cobalt and bismuth concentrates, but the market for such concentrates was limited. The development program was modified to include a hydrometallurgical process was added for the treatment of these concentrates into readily marketable products.

A number of bench flotation tests have been undertaken for the NICO ore since initiation in 1997, and a number of flotation parameters have been evaluated. The basic bulk circuit has been established with a larger rougher to pull both bismuth and cobalt concentrates into a bulk rougher concentrate. This bulk rougher is cleaned once and the cleaner scavenger concentrate is returned to the bulk rougher. The cleaner circuit has been tested with a single stage and two stage cleaners. PAX (potassium amyl xanthate) collector and MIBC frother are the only reagents required to produce the bulk cleaner concentrate. Both flotations are set at a natural operating pH of 8.2.

At the time of the laboratory testing, the bulk cleaner tailings were planned to be discarded to tailings, if ores of low-gold content are treated. For ores of higher-gold content, bulk cleaner tailings are directed to the hydrometallurgical section for cyanidation.

The cleaned bulk concentrate is reground to a fineness of P80 at 14 microns. Cobalt and bismuth separation is achieved by depressing the arsenopyrite with cyanide, while allowing unhindered flotation of bismuth. The floated material from the bismuth rougher is cleaned twice to produce the bismuth concentrate. Tails from bismuth flotation circuit becomes the cobalt concentrate product.

It was determined that roughly two-thirds of the recovered gold reports to the bismuth concentrate and one-third to the cobalt concentrate.

(Note: The bulk concentrate regrind and cobalt / bismuth separation is not a part of the NICO flow sheet. They will be carried out in the SMPP hydrometallurgical plant.)

13.4.2 2006 Hydromet Mini Pilot Plant

In 2004 to 2005, a more detailed testwork program was conducted using concentrate samples produced during the flotation cycle testwork and concentrates derived from two bulk samples. P&E Mining Consultants Inc., Report No. 247 Page 71 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

This testwork included both a bench scale and a mini-pilot scale component. The purpose of the bench testwork is to verify the 2003 results and establish solution chemistry and reaction rates in advance of the pilot scale continuous test. A single batch pressure oxidation test and a number of semi-continuous tests were proposed. The extractions are consistent with previous results, with cobalt extraction at 97% and gold extraction ranging from 95.2% to 97.7%.

Mini-pilot scale pressure oxidation was conducted on two bulk concentrates recovered from two ore composites. The first composite was largely derived from assay reject samples which were exposed to slight oxidation, and responded relatively poorly to flotation. The second composite was produced from fresh drill core. The cobalt extractions from the two tests are coherent at 95% to 96%. A significant difference in gold extraction was noted with the samples, with the former sample at only 88%, while the latter of about 95%.

The testwork performed during the 2001 to 2006 period became the basis for Micon‟s BFS study in 2007. The data and results used in the Micon study are referenced and compared periodically in the FEED report to the FEED results obtained.

13.4.3 2007 Flotation Pilot Plant

In October 2007, Fortune Minerals provided approximately 176 t of ore from the NICO deposit to SGS for an extensive laboratory and pilot plant test program that included operations from grinding through to final products.

A summary of test reports, are given below.

 A Pilot Plant Investigation into Cobalt, Bismuth, and Gold Recovery from the NICO Deposit, Project 11747-001, January 16, 2009 (DRAFT)  Flocculant Screening, Gravity Sedimentation and Pulp Rheology for SGS Lakefield Research - Fortune Minerals, December 2007  Flocculant Screening, Gravity Sedimentation, Pulp Rheology, Pressure Filtration and Vacuum Filtration Studies for SGS Lakefield Research – Fortune Minerals, May 2008

Sample Description, Source Blending and Feed Grades

The ore was received in 255 supersacks which were subsequently separated into 12 composite samples representing both underground ore samples (UG) and open-pit ore samples (OP). Composites Nos. 1 through 9 were blended to form sample No. P-1, which represents a 2(OP):1(UG) blend of ore that would be representative of the first 2 years of operation. Composite Nos. 10 to 12 were blended to form sample No. P-2, which represents the remaining years of open-pit operation, including some of the variability that may be encountered. Smaller subsamples, Nos. L-1 and L-2, were produced for the accompanying bench-scale programs. The head grades for the composites are shown in Table 13.7.

The beneficiation pilot plant was carried out in December 2007. The purpose of the pilot plant was to produce bulk quantities of concentrate for hydrometallurgical testing, to confirm the concentrator flow sheet, to provide engineering design criteria confirmations for Micon‟s BFS, to test the variability of the ore on plant production, to evaluate the impact of recycled water on plant production and to perform environmental impact testing on all discharge streams. The hydrometallurgical test program extended throughout 2008 and 2009. P&E Mining Consultants Inc., Report No. 247 Page 72 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

TABLE 13.7 SUMMARY OF COMPOSITE HEAD GRADES Assays Ore Type Comp ID Co, % Bi, % Au, g/t As, % S, % Underground P-1 0.10 0.18 2.72 0.47 0.56 Open-Pit P-2 0.11 0.25 0.90 0.46 0.60 *2007 Beneficiation Pilot Plant

Grinding Data

Both P-1 and P-2 composites were subjected to rod mill and ball mill grindability tests.

The rod mill values were considered high and the ball mill values were average. The results are shown below in Table 13.8.

TABLE 13.8 ROD MILL AND BALL MILL BOND WORK INDICES Rod Mill Work Index Ball Mill Work Index Composite ID kWh/t Hardness Percentile kWh/t Hardness Percentile P-1 19.1 88 14.2 47 P-2 20.1 93 14.8 53

Flotation Circuit Results and Scale up Factors for FEED Design

The pilot plant flow sheet consisted of a rod mill / ball mill combination to grind the ore, followed by bulk flotation (rougher, cleaner and cleaner / scavenger), regrind, and bismuth flotation (rougher, rougher / scavenger and three stages of cleaning) to produce a bismuth concentrate and a cobalt concentrate. The initial intent was to grind the ore from a crushed product at 80% passing minus 9 mm to a final grind fineness of 80% minus 70 microns. The crushing setting was changed to 80% passing minus 5 mm, because of the limited capacity of the rod mill. From this crushed product size, it was possible to maintain a good throughput rate; but the ball mill was too large, so the final grind was 80% passing minus 57 microns.

The metal recoveries and concentrate grades from the pilot plant were better than laboratory testwork results. However, it is expected that an actual primary grind size at 80% passing minus 57 microns would result in better metal recoveries at the rougher stage. The regrind fineness prior to bismuth flotation was as the laboratory flotation testwork at 80% minus 14 microns.

The results of the flotation circuit for the underground and open-pit composites including averaged head grades are presented in Table 13.9 and Table 13.10 respectively.

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TABLE 13.9 SUMMARY OF AVERAGE HEAD GRADES AND CIRCUIT PERFORMANCE OF COMPOSITE P-1 Grade (%, g/t) Recovery (%) Product Bi Co Au Bi Co Au Head 0.18 0.10 2.71 100.0 100.0 100.0 Bulk Concentrate 1.26 0.76 17.15 87.4 94.2 77.5 Bulk Cleaner Concentrate 5.37 3.40 69.16 79.8 90.0 67.1 Bi Concentrate 45.1 0.83 492 71.6 2.3 51 Co Concentrate 0.61 3.70 18.5 8.1 87.6 16

TABLE 13.10 SUMMARY OF AVERAGE HEAD GRADES AND CIRCUIT PERFORMANCE OF COMPOSITE P-2 Grade (%, g/t) Recovery (%) Product Bi Co Au Bi Co Au Head 0.26 0.11 0.91 100.0 100.0 100.0 Bulk Concentrate 1.81 0.86 5.03 86.8 94.5 69.1 Bulk Cleaner Concentrate 6.82 3.38 17.33 79.5 90.3 57.8 Bi Concentrate 48.9 0.46 110 72.2 1.6 46.2 Co Concentrate 0.77 3.71 3.91 8.1 89.1 11.7

A summary and comparison of design data from the Micon BFS study, the 2007 SGS Lakefield pilot plant, and 2008 design criteria used for the FEED study is given in Table 13.11.

TABLE 13.11 SUMMARY AND COMPARISON OF CRUSHING, GRINDING AND FLOTATION DESIGN PARAMETERS Parameter Units Micon Design 2007 Pilot Plant FEED Design Crushing (P80) mm 12 9 (target), 5 (actual) 15 Primary Grind (P80) µm 72 70 (target), 57 (actual) 74 Regrind (P80) µm 14 14 15 SAG Mill Work Index(1) kWh/t 17.5 - - Bond Rod Mill Work Index kWh/t 17.5 19.1 / 20.1 20.2 Bond Ball Mill Work Index kWh/t 10.5 – 12.1 14.2 / 14.8 13.6 Regrind Mill Work Index kWh/t 8 8 Rougher Flotation minutes 35 45 – 60 45 Bulk Cleaner / Cleaner-Scavenger minutes 10 total 8 / 6.2 10 / 8 8 per stage Bismuth Rougher minutes 6.9 7 (4 stages total) Bismuth Scavenger minutes 6.5 7 Bismuth Cleaner (3 stages) minutes 3.1 / 3.5 / 10.8 3.5 (per stage) (1) SAG milling was found to be not suitable for the NICO ore, and was not considered in the FEED study.

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Flotation reagents used were MIBC, PAX and NaCN. Dosing requirements were optimised for each circuit:

 Bulk flotation rougher, >250-g/t PAX (includes PAX added to tertiary grind), 50- g/t MIBC  Bulk flotation cleaner and scavenger, 30-g/t PAX  Bismuth flotation rougher, 17-g/t NaCN, 7-g/t PAX, 0.5-g/t MIBC  Bismuth flotation cleaner scavenger, 10-g/t NaCN, 1-g/t PAX  Bismuth flotation 2nd cleaner, 1-g/t NaCN.

Settling, Thickening and Filtration Testing

As part of the beneficiation pilot plant, Pocock Industrial Inc. (Pocock) was retained to perform thickener settling tests on the bismuth and cobalt concentrates. A summary of design criteria developed by Pocock, as it compares to the Micon‟s design, is shown in Table 13.12. Bulk concentrate was not tested for settling rate. However, since both the bismuth and cobalt concentrates have the same settling rate of 0.13 m3/t/d, this rate was used for the bulk concentrate thickening design. For the FEED design, a high-rate thickener was specified for the bulk flotation tails and conventional thickeners for the bulk concentrate.

TABLE 13.12 SOLID / LIQUID DESIGN CRITERIA FROM POCOCK TESTWORK Pocock / SGS Lakefield Results and FEED Design Criteria Micon Design Specific Flocculant Underflow Loading/Settling Rate Addition Rate Solids Bulk Flotation 3.7 m3/m2·h 35 g/t 65-70 wt% 55 wt% UF Tails Cleaner Tails 0.17 m2/MTPD* 80 g/t 50-55 wt% Bismuth 0.64 m2/MTPD, 0.13 m2/MTPD 30 g/t 50-58 wt% Concentrate 55 wt% UF Cobalt 0.16 m2/MTPD, 0.13 m2/MTPD 50 g/t 60-65 wt% Concentrate 55 wt% UF Bulk 0.13 m2/MTPD 80 g/t 60 wt% Concentrate *MTPD: Metric Tons per Day

Environmental Samples Collected

As previously mentioned, a sample of all discharge streams was collected for environmental impact testing.

Hydromet Separation from NICO Site

It was decided in 2009 that the concentrator and hydrometallurgical processes should be separated. In this new scenario, the NICO concentrator in the NWT will include crushing, grinding, bulk flotation and an additional regrind circuit with secondary bulk flotation to produce a bulk concentrate. The combined bulk concentrate generated will be transported by rail into the SMPP facility in Saskatchewan. The SMPP will include a regrind circuit, cobalt-bismuth

P&E Mining Consultants Inc., Report No. 247 Page 75 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. separation flotation and the entire hydrometallurgical process for cobalt, bismuth, gold and copper recovery.

The addition of the regrind and secondary bulk flotation circuit at the NICO site was undertaken to address gold lost to the cleaner tails in the previous flow sheet that would be subsequently recovered by cyanidation.

Previous flotation testwork showed that about 10% of the gold contained in the NICO ore is typically reported to the bulk cleaner tailing stream. As a result of the separation of the concentrator and hydrometallurgical circuits, and the change of location for the hydrometallurgical plant, it was believed that the capital and operating costs for a dedicated gold circuit to recover gold from the cleaner tails at NICO would be prohibitive. Since about 8% of the tailings by mass would be cleaner tailings, this large volume of material cannot be transported into the SMPP. Therefore, there is a need to demonstrate an increase gold recovery by improving deportment of gold from the first stage bulk cleaner tailing into the bulk concentrate, such that a cyanidation circuit to treat for gold in the NICO concentrator is not required.

Kinetics and FLEET Modeling

More work was carried out at SGS Lakefield in September 2009 to optimize the bulk flotation circuit and to increase product recoveries, in particular for gold contained in the bulk cleaner tailing stream. The FLEET flotation simulator was used to evaluate different circuit performance in terms of their valuable metal recovery and final concentrate product quality and to determine the best circuit configuration for optimizing the flotation circuit. The scope of the flotation kinetics study included:

 Derivation of flotation kinetic parameters from the results of the standard flotation tests conducted on No. P-1 flotation composite  Benchmark simulation of the LCTs  Evaluation of supplied flotation circuit configurations.

The FEED design is based on Options 4 and 6 from the September 2009 testwork, which includes a bulk second cleaner, as well as a secondary flotation circuit on the product of the regrind situated on the bulk cleaner-scavenger tails. This secondary flotation circuit includes a rougher stage with a cleaner and a cleaner-scavenger stage. The concentrate from the secondary cleaner is combined with the bulk second cleaner concentrate as the bulk concentrate product. The concentrate from the secondary cleaner-scavenger is fed back into the bulk first cleaner stage. The input parameters for Option 4+6 are given in Table 13.13.

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TABLE 13.13 FLOW SHEET OPTION 4+6 INPUT PARAMETERS Cell Volume (m3) No. Cells Froth Recovery (%) Entrainment Bulk Rougher 50 7 30 0.40 Bulk Cleaner 10 4 55 0.56 Bulk Cleaner-Scavenger 10 3 55 0.56 Bulk Second Cleaner 10 2 55 0.56 Secondary Rougher 10 5 45 0.61 Secondary Cleaner 5 2 45 0.61 Secondary Cleaner-Scavenger 5 2 45 0.61

The stage recoveries obtained independently for Options 4 and 6 by using the input parameters above are provided in Table 13.14 and Table 13.15. It should be noted that the FLEET simulation work is based on a small sample of material through a LCT, and with flotation feed material at a slightly different P80 grind size. Therefore, it is possible the recoveries stated in the FLEET model below might not completely reflect the actual achievable recoveries in plant operation.

TABLE 13.14 FLOW SHEET OPTION 4 STAGE RECOVERIES Stage Recovery Mass Recovery Bi Co Au Target Model Target Model Target Model Target Model Bulk Rougher 12.1 12.1 88.0 87.0 92.0 94.2 78 80.1 Bulk Cleaner 27.6 23.5 - 88.3 - 93.1 - 83.2 Bulk Cleaner-Scavenger 14.3 8.3 - 67.2 - 67.2 - 64.8 Bulk Cleaner + Clnr- - - 97.0 96.8 96.0 99.1 90.0 97.6 Scav. Bulk Second Cleaner - 77.6 - 96.8 - 98.1 - 95.3 Rougher Circuit - 12.1 - 87.0 - 80.1 - 80.1 Cleaner Circuit - 20.6 - 95.7 - 97.3 - 93.0

Overall 3.7 2.5 85.0 83.3 89.0 91.6 70 74.5

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TABLE 13.15 FLOW SHEET OPTION 6 STAGE RECOVERIES Stage Recovery Mass Recovery Bi Co Au Target Model Target Model Target Model Target Model Bulk Rougher 12.1 12.1 88.0 87.0 92.0 94.2 78 80.1 Bulk Cleaner 27.6 24.2 - 90.1 - 94.2 - 85.5 Bulk Cleaner- 14.3 9.3 - 69.6 - 65.6 - 67.4 Scavenger Bulk Cleaner + Clnr- - - 97.0 96.8 96.0 97.9 90 94.7 Scav. Secondary Rougher - 12.6 - 55.2 - 56.4 - 58 Secondary Cleaner - 15.5 - 76.6 - 75.4 - 80.6 Secondary Clnr-Scav. - 10.7 - 58.1 - 54.8 - 57.4 Stage 1 Recovery - 3.1 - 84.2 - 92.2 - 75.9 Stage 1 Cleaner - 26.0 - 96.8 - 97.9 - 94.7 Recovery Stage 2 Recovery - 1.9 - 42.3 - 42.5 - 46.7 Stage 2 Cleaner - 15.5 - 76.6 - 75.4 - 80.6 Recovery

Overall 3.7 3.3 85.0 85.5 89.0 93.1 70 78.0

The addition of Option 6 is not likely to have much impact on the overall recovery of cobalt as most will have already reported to the bulk flotation concentrate. However, an increase in the overall recovery of bismuth and gold is expected.

13.5 2010 FLOTATION PILOT PLANT

13.5.1 Sample Description, Source Blending and Feed Grades

Another beneficiation pilot plant was carried out at SGS Lakefield in October 2010. Some laboratory work was also completed in support of the pilot plant. The purpose of the pilot plant was to gain confidence in the results obtained from the FLEET model on the option 4+6 flow sheet. The bismuth separation circuit was included to assess the overall implications on metallurgical performance.

Approximately 52 t of NICO ore material from the 2006 and 2007 underground test mining program was shipped in three separate shipments to SGS Lakefield for pilot plant testing. The first two shipments were designated as composite P-3. The last shipment was composite P-4. About 10 t from each of the composite P-3 and P-4 were taken out and blended to produce composite P-5. The head grades are reported in Table 13.16.

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TABLE 13.16 SUMMARY OF COMPOSITE HEAD GRADES Assays Comp ID Co, % Bi, % Au, g/t Cu, % As, % S, % P-3 0.021 0.25 2.86 0.017 0.076 0.28 P-4 0.11 0.021 2.04 0.022 0.29 0.47 P-5 0.07 0.14 2.45 0.020 0.18 0.38 *2010 Beneficiation Pilot Plant

Grinding Data

P-5 composite was subjected to rod mill and ball mill grindability tests. The results are shown below in Table 13.17.

TABLE 13.17 ROD MILL AND BALL MILL BOND WORK INDICES Composite ID Rod Mill Work Index Ball Mill Work Index kWh/t Hardness Percentile kWh/t Hardness Percentile P-5 20.6 95 14.3 48 P-5 HPGR Product 14.2 47

Flotation Circuit

The pilot plant flow sheet consisted of a rod mill/ball combination to grind the ore, followed by bulk flotation (rougher, cleaner, cleaner / scavenger, second cleaner), regrind, secondary flotation (rougher, cleaner and cleaner / scavenger), bismuth regrind, and bismuth flotation (rougher, rougher / scavenger and three stages of cleaning) to produce a bismuth concentrate and a cobalt concentrate.

The average results for composite P3 and P5 are summarized in Table 13.18 and Table 13.19. Composite P-4 was contaminated during the test run and therefore no results were obtained from this sample.

TABLE 13.18 FLOW SHEETS OPTIONS 4 & 6 PILOT PLANT RESULTS FOR COMPOSITE P3 Mass Assay (Adj.) (%, g/t) Distribution (%) Shift ID Product (wt %) Bi Co S Au Bi Co S Au PP Feed 100.0 0.23 0.023 0.32 2.40 100.0 100.0 100.0 100.0 Bulk Ro Conc 19.1 1.00 0.11 1.57 8.48 83.3 88.8 94.0 67.7 Average Bulk Conc 3.2 5.45 0.60 8.84 41.5 76.1 84.1 88.6 55.6 PP-08A Scav 1st Cl Conc 1.5 0.26 0.017 0.42 4.67 1.8 1.2 2.0 3.0 to PP- Bi Circuit Feed 3.8 4.73 0.51 7.59 37.4 78.8 85.1 90.7 59.7 08B Bi Ro Scav Tails 3.6 0.98 0.53 6.96 13.5 15.2 82.1 77.4 20.0 (Co Conc) Bi Conc 0.3 54.9 0.26 15.9 357 63.7 3.0 13.3 39.7

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TABLE 13.19 FLOW SHEETS OPTIONS 4 & 6 PILOT PLANT RESULTS FOR COMPOSITE P5 Mass Assay (Adj.) (%, g/t) Distribution (%) Shift ID Product (wt %) Bi Co S Au Cu As Bi Co S Au Cu As PP Feed 100.0 0.14 0.076 0.45 2.56 0.021 0.24 100.0 100.0 100.0 100 100 100 Bulk Ro 20.4 0.59 0.33 2.10 9.8 0.09 1.09 84.9 89.4 94.3 78.1 91.0 93.9 Conc Bulk 3.6 3.02 1.73 10.8 49.0 0.50 5.72 78.3 82.9 86.5 69.8 87.0 88.8 Conc Scav 1st Average 0.8 0.30 0.12 1.23 7.20 0.02 0.16 1.6 1.2 2.1 2.2 1.0 0.6 Cl Conc PP-06 to Bi Circuit PP-07B 4.4 2.55 1.45 9.1 41.7 0.40 4.61 79.9 84.1 88.6 71.9 88.0 89.5 Feed Bi Ro Scav 4.1 0.54 1.45 7.67 15.0 0.17 4.66 15.5 77.7 68.7 23.8 34.8 83.5 Tails (Co Conc) Bi Conc 0.3 26.0 1.40 26.0 353 3.19 4.02 64.4 6.4 20.0 48.1 53.2 6.0

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Gravity Recoverable Gold

In October 2010, Fortune retained SGS Lakefield to perform gravity recoverable gold (GRG) tests to determine the GRG contained in the ore. At the time of the report, only two sets of testwork result were available to perform design engineering. This work is based on the P3 and P4 composite, which respectively represent a high-bismuth and low-cobalt ore sample, and a- high cobalt and low-bismuth ore sample.

The GRG results for composite P3 and P4 are summarized in Table 13.20 and Table 13.21.

TABLE 13.20 GRAVITY RECOVERABLE GOLD RESULTS FOR COMPOSITE P3 Mass Assay Units Distribution Grind Size Product Au grams % Au (%) (g/t) 581 P80 = Stage 1 Conc. 91.3 0.46 19.6 1,792 2.6 µm Sampled Tails 249 1.26 3.23 803 1.15 214 P80 = Stage 2 Conc. 94.2 0.48 22.1 2,082 3.0 µm Sampled Tails 241 1.22 2.95 710 1.02 54 P80 = Stage 3 Conc. 121.4 0.62 29.7 3,612 5.18 µm Final Tails 18,943 96.0 3.20 60,679 87.1 Totals (Head) 19,739 100 3.53 69,677 100 Knelson Conc. 307 1.56 24.4 7,485 10.7

TABLE 13.21 GRAVITY RECOVERABLE GOLD RESULTS FOR COMPOSITE P4 Mass Assay Units Distribution Grind Size Product Au grams % Au (%) (g/t) 575 P80 = Stage 1 Conc. 96.3 0.48 22.2 2,134 4.5 µm Sampled Tails 166 0.82 2.13 352 0.74 248 P80 = Stage 2 Conc. 97.7 0.48 25.9 2,532 5.3 µm Sampled Tails 257 1.27 1.69 435 0.91 60 P80 = Stage 3 Conc. 105.8 0.52 38.5 4,069 8.53 µm Final Tails 19,480 96.4 1.96 38,195 80.0 Totals (Head) 20,202 100 2.36 47,717 100 Knelson Conc. 300 1.48 29.1 8,735 18.3

Filtration Testing

A combined bulk and secondary flotation concentrate sample was sent for pressure filter testing. The sample was conditioned to produce a feed pulp with 60% solids using flocculant Hychem

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AF303. The lowest moisture recorded was 17.8 %. A suggestion was made to investigate vacuum filtration to produce a thinner filter cake with possibly less moisture.

13.6 FLOTATION DESIGN CRITERIA

Engineering design criteria for the bulk flotation circuit was selected based on correlations developed from pilot plant data between mass pull and ROM grades (Bi ROM for Bi, and As ROM for Co), correlations between bulk concentrate grades and ROM grades, as well as the data resulted from the Option 4+6 flow sheet of the FLEET testwork.

A summary of the design parameters used in the FEED study is shown below in Table 13.22.

TABLE 13.22 NICO PLANT STAGE RECOVERIES AND RESULTING GRADES Stage Recovery (%) Grade Co Bi Au Cu Co Bi Au Cu (%) (%) (g/t) (%) Plant Feed 100 100 100 100 0.14 0.19 2.386 0.034 Gravity Concentrator 7.8 62036 Bulk Rougher Conc.1 95.6 88.2 77.7 93.8 1.07 1.31 12.56 0.25 Bulk Cleaner Conc.1 90.9 85.8 80.9 94.0 1.02 1.20 11.04 0.23 Bulk Cleaner/Scav. Conc.1 54.7 44.7 41.8 0.0 0.81 1.29 15.81 0.00

Bulk Second Cleaner Conc. Secondary Rougher Conc.2 79.5 82.9 84.6 68.1 0.50 1.30 18.20 0.13 Secondary Cleaner Conc.2 11.2 16.5 19.0 19.2 0.36 1.38 22.32 0.17 Secondary Cleaner/Scav. Conc.2 29.1 35.8 35.7 34.2 1.42 4.29 58.21 0.41 (1) SGS Lakefield 2007 pilot plant composite P1 for cobalt, bismuth and gold. SGS 2010 Pilot Plant composite P5 for copper. (2) SGS Lakefield 2010 Pilot Plant composite P5.

13.7 PROCESS HYDROMETALLURGICAL PLANT

The Process Design Criteria were developed for the handling of bulk concentrate in bulk bags from the NICO concentrator located in the NWT to the SMPP in Langham, Saskatchewan. The metals processing facility is developed for processing bulk concentrate of 8 wt% moisture at an annual throughput of 67,188 wet metric tonnes. The daily design rate is approximately 217 wet tonnes per day, equivalent to the average daily production rate of the concentrator (incorporating availability). The majority of the design criteria were based on information from the bench scale and pilot testwork performed at SGS Lakefield from 2007 to 2012, and supplied by Fortune or its contractors. Each criterion was referenced to provide the source of the information. The overall gold and cobalt recoveries at the SMPP plant are 94.7%, and 92.9%, respectively. Combined with the NICO concentrator, the overall recoveries are 73% for gold and 84.4% for cobalt. Individual stage recovery criteria and the plant overall recovery are summarized in Table 13.23. Bismuth recoveries for the CLER (Chloride Leach Electro Recovery) process will be provided by DMA.

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TABLE 13.23 STAGE AND OVERALL PLANT RECOVERIES Stage Recovery Overall Recovery Co Bi Au Cu Co Bi Au Cu NICO Concentrator Plant Feed 100 100 100 100 100 100 100 100 Bulk Concentrate 90.9(2) 82.1(2) 72.6(2) 89.1(2) 90.9(2) 82.1(2) 72.6(2) 89.1(2) SMPP Facility Bulk Concentrate Feed 100 100 100 100 100 100 100 100 Bismuth Flotation Concentrate 2.2 88.9 78.7 60.4(2) 2.2 88.9 78.7 60.4(2) Bismuth Flotation Tails (Co Concentrate) 97.8 11.1 21.3 39.6(2) 100 11.1 21.3 39.6(2) Co Concentrate, Bi Residue Feed 100 100 71.2(2)

Pressure Oxidation(2) 95.0 100(2) 76.0 95.0 100.0 54.1

Solution Neutralization, Co Precipitation 98.5(2) 85.4 93.6 46.2 and Cu Cementation(2) Purification & Ion Exchange 99.3 92.9

Co Electrowinning 100 92.9

Bi Concentrate Feed 88.9

Bismuth CLER(5) 98.86 87.9

Gold: Cyanidation Feed 100.0

Gold Cyanidation(5) 95(2) 95.0(2)

Merrill Crowe 99.7 94.7

Gold Refinery 100 94.7

SMPP Recovery 92.9 87.9 94.7 46.2

Overall (NICO + SMPP) Recovery 84.4 72.2 68.8 41.2

(1) The recoveries represent the results from SGS pilot plant testing between 2007 and 2011. The process facilities were designed using the process conditions demonstrated by the pilot plants. The recoveries in the table represent the performance of the process plant with similar ore feed conditions. (2) The current process design and the METSIM model are based on testwork results up until 2009. In the next phase, the METSIM model should be updated with the recovery values shown in the Table 13.23 from the testwork conducted in 2010 and 2011. The latest recovery values, although slightly different from METSIM are not anticipated to affect the design and equipment selection significantly. These are the changes: a. Autoclave residue cyanidation gold recovery was revised from 89.5% to 95% b. Au recovery in the autoclave was revised to be based on 100% c. Recoveries of Cu in NICO and SMPP were revised. d. Co, Bi, Au and Cu recoveries at NICO were revised from FLEET modelling values to pilot test results e. Co recovery in solution neutralization, precipitation was revised from 96.1% to 98.5% (3) The table does not represent the anticipated recovery of the ore body. The bulk metals, cobalt and bismuth have somewhat consistent recovery rates over the expected mine life/ore grade profiles and the process facility is designed to accommodate these varying feed grades. Gold however does have a noticeable trend with lower feed grades having lower flotation recoveries and conversely higher feed grades will have a higher recovery. Gold recovery for the plant design was based on a blend of ore samples that will not necessarily match the ore fed to the plant. The anticipated gold grade and recovery fluctuations will not affect the plant equipment sizing or costs. (4) Chloride Leach and Electro-Recovery (CLER) circuit for the recovery of Bi is provided by the Client based on report provided by EHA/DMA on Bi processes and recovery.

13.7.1 Concentrate Characteristics

The overall design operating availability of the SMPP facility is 85%. Based on this availability, the design bulk concentrate rate is 217 wet tonnes per day, or 9 wet tonnes per hour.

The estimated bulk density of the unconsolidated, wet filter cake bulk concentrate is 2.4 t/m³. With a safety / scale-up factor of 110% as proposed by Metso, the predicted specific energy requirement for SMD is 21.78 kWh/t.

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The bulk concentrate is a sulphide ore. The total ore contains 18.23% sulphur in the form of sulphate, elemental and sulphide sulphur. There is 0.40% sulphate sulphur, 0.05% elemental sulphur and 17.78% sulphide sulphur.

13.7.2 Laboratory Bench-Scale and Pilot Plant Testing

Since 1997, a significant amount of testwork on the NICO material has been conducted. Preliminary laboratory bench-scale work was conducted at SGS through 2003, resulting in the establishment of a basic flow sheet. A more definitive testwork program, known as the „mini- pilot‟, was conducted at SGS Lakefield and EHA Engineering in 2004 and 2005 for the preparation of the Micon BFS. An extended pilot plant test program also was performed from 2006 to 2009 at SGS Lakefield and then subsequently, in 2010 to 2012.

13.7.2.1 Laboratory Testing

Bulk and Bismuth Flotation

A number of bench flotation tests have been undertaken for the NICO ore since 1997, and flotation parameters have been determined.

Bulk rougher concentrate floated from the rougher stage is cleaned in one stage of water cleaning. The floated material from the bismuth rougher is cleaned twice to produce the bismuth concentrate. Tails from bismuth flotation circuit becomes the cobalt concentrate product.

Cobalt Hydrometallurgy

A process development program was undertaken from 1997 to 2003, to define a suitable flow sheet for the treatment and recovery of the NICO concentrates. A batch pressure oxidation test carried out at 180°C yielded a cobalt extraction of up to 97%. After a solid / liquid separation step, the copper and cobalt in solution were treated in two solvent extraction circuits. Extracted cobalt ultimately was recovered as a high-grade cobalt carbonate. Gold was recovered by conventional cyanidation of the autoclave residue.

2004 to 2005 Mini-Pilot

In 2004 to 2005, a more detailed testwork program was conducted using concentrate samples produced during the flotation cycle testwork and concentrates derived from two bulk samples. This testwork included both a bench-scale and a mini-pilot scale component. The extractions were consistent with previous results, with cobalt extraction at 97% and gold extraction ranging from 95.2% to 97.7%.

2007 to 2009 Pilot Plant

In October 2007, FML provided approximately 176 t of ore from the NICO deposit to SGS for an extensive laboratory and pilot plant test program that included operations from grinding through to final products.

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2010 to 2012 Testwork

Batch and continuous pressure oxidation test work was conducted on combined Bismuth leached residue and Cobalt concentrate feed. The cobalt and copper recoveries were found to be approximately 95% and 76% respectively and consistent with the previous test work. Following this, cyanidation tests were conducted on the autoclave residue to recover gold.

Bismuth cathode was also smelted in an induction furnace and proved that a bismuth ingot at 99.99% purity can be produced. In addition, pilot scale work was completed to demonstrate the production of a cobalt sulphate heptahydrate by solvent extraction of cobalt pregnant solution to remove metal impurities, followed by crystallization.

Flotation Circuit

A beneficiation pilot plant was carried out at SGS Lakefield in December 2007 to produce bulk quantities of concentrate for hydrometallurgical testing, to confirm the concentrator flow sheet, and to provide engineering design criteria confirmations for the definitive feasibility study.

More work was carried out at SGS Lakefield in September 2009 to optimize the bulk flotation circuit and to increase product recoveries, in particular for gold contained in the bulk cleaner tailing stream. The FLEET flotation simulator was used to evaluate different circuit performance, in terms of their valuable metal recovery and final concentrate product quality, and to determine the best circuit configuration for optimizing the flotation circuit.

Cobalt Hydrometallurgical Circuit

The purpose of the hydrometallurgical test program was to finalize proposed flow sheet and confirm assumptions made during the Micon BFS feasibility study. The bench-scale program was used to determine operating conditions for piloting. As it was not possible to pilot the complete hydrometallurgical process at the same time, the continuous pilot plant was operated in sequential sections:

 Co POX autoclave, Fe / As precipitation, Cu cementation  Co carbonate precipitation  Co carbonate releach, impurity removal (with IX), Co EW

Pocock was retained to perform thickening, vacuum and pressure filtration tests on the autoclave discharge material. The filtration rate of the material was found to be significantly lower than the rate used by Micon in the previous study, for both vacuum and pressure filtration. Vacuum filtration is considered impractical because of the slow rate. A pressure filter was selected over a filter press because of its shorter cycle time. The filtration rate used for design is estimated by Vendor A, based on Pocock`s pressure filtration data and Vendor A‟s experience with the material.

Solution Collection

The tank is used primarily for collecting various repulped precipitate streams recycled from the downstream process. The highly acidic autoclave discharge slurry, which is combined with these precipitate streams, also will leach valuable metals into solution. Since this leaching step involves the recycling of numerous streams which are generated further downstream in the P&E Mining Consultants Inc., Report No. 247 Page 85 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. process, including this leaching step as a part of a sequential pilot testwork program was not possible.

Fe/As Precipitation

SGS Lakefield testwork indicated that good Fe / As removal can be achieved with neutralization to pH 4.6, compared to pH 5.0 which was used in the Micon study. In this case, settling characteristics are better with 25 wt% underflow densities compared to 15 wt% in the residue thickener. The lower operating pH also may have contributed to the lower residence time of 5 hours compared to 8 hours in the Micon study.

The Fe / As residue is releached with sulphuric acid to solubilize co-precipitated Co and Cu. The solution from the releach is treated further with the addition of Fe powder to remove Cu. Both of these steps were tested during the continuous pilot plant.

Filtration testwork on this releached Fe / As residue material also was carried out by Pocock. This material was very difficult to filter, because of the gypsum and scorodite content. An automated pressure filter was selected for this application, because of its shorter cycle and technical time comparing to both a vacuum filter and a filter press.

Copper Polishing Precipitation

The pregnant leach solution (“PLS”) from the Fe / As precipitation is carried forward to copper precipitation to polish and remove residual copper. The carbonate precipitate is filtered and recycled back to the collection tank before Fe / As precipitation, where it can be redissolved and metal values reclaimed.

An alternative method to remove copper from solution is the use of a copper resin by ion exchange. The resin selected for this application is Lanxess TP-207. The PLS is allowed to pass through ion exchange (“IX”) columns, until a breakthrough of copper is detected in the discharged raffinate. At that point, the lead IX column is taken off-line to undergo the elution cycle. Since nickel and cobalt will load onto the resin in addition to copper, a two-step elution cycle must be applied to strip the nickel-cobalt and copper separately.

To compare the two alternatives above, the raffinate produced from each of these methods was sent to the cobalt precipitation circuit downstream. Both solutions were injected with Na2CO3 to produce cobalt carbonate cakes. The copper precipitation circuit would serve as a copper polishing step following the Fe / As precipitation, and a copper IX column would be a downstream copper guard in the dissolution-IX-EW circuit.

Cobalt Purification and Precipitation

Cobalt precipitation is completed in two stages: the first stage to produce a high-purity precipitate for further processing, and the second stage to scavenge cobalt for recycle.

Stage 1 cobalt precipitation cake was thickened at SGS Lakefield. This sample then was sent to TEMA Systems to perform a solid-liquid separation test using a centrifuge.

Similar to the copper precipitation circuit, the amount of cobalt carbonate produced during the pilot plant in Stage 2 cobalt precipitation was so small that proper filtration testing could not be P&E Mining Consultants Inc., Report No. 247 Page 86 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. carried out. Instead, the precipitates were removed with an on-stream cartridge filter. Based on consultation with EHA and Jacobs‟ experience with this type of material, a conservative filtration rate was selected for the design for this filter application.

Cobalt Dissolution, Ion Exchange (“IX”) and Electro-winning (“EW”)

The precipitate from cobalt Stage 1 precipitation is filtered and leached in cobalt spent electrolyte. The solution then is passed through ion exchange and onto EW, where cobalt is plated. The Micon study allowed for five fixed bed columns for nickel ion removal, an intermittent IX circuit for zinc, and a cobalt electrowinning circuit that produces sheet cathodes. It was determined that producing a cobalt cathode of the highest purity (LME Cobalt Specification) is critical, and hence an extensive IX circuit is required to ensure all impurities are removed before EW. The pilot testwork performed at SGS Lakefield combined the dissolution, IX and EW in a closed circuit. The EW spent electrolyte was recycled to the cobalt carbonate dissolution step, and then the PLS passed through the zinc and nickel IX before discharging to EW as the strong electrolyte to plate cobalt.

For ion exchange, various resins were tested during the pilot campaign to determine which resins responded best to remove zinc, copper and nickel in the presence of cobalt. Since only a small amount of cobalt carbonate was available for dissolution to produce PLS feed for this IX testwork and the fact that the impurities levels in the solution are relatively low (in the order of ppm), limited loading and elution cycles of the IX resins were carried out before all the feed was exhausted. It was remarked by SGS Lakefield in their pilot plant report that some of the IX testwork results are limited and preliminary, and elution conditions were not optimized. In spite of the limitations, operation targets were achieved toward the end of the pilot testwork, and parameters for fixed bed IX were determined. The chosen resins for the removal of zinc, copper and nickel in the IX circuit are Lanxess VP-OC-1026, Lanxess TP-207 and Dow XUS-43578 respectively.

Through a trade-off study between fixed bed IX and continuous IX (CIX), it was determined that CIX would be advantageous for this flow sheet, because of the reduction in water, reagents, acid, etc. CIX is included in the design. Parameters for CIX, as supplied from the CIX Vendors based on their previous experience, and confirmatory tests will be carried out in the future.

Both the Micon study as well as the SGS Lakefield pilot testwork used a cobalt tenor in dissolution of 40 g/L Co the design cobalt tenor was increased to 54.2 g/L. This is due to the use of an anode bag in the cobalt EW design to allow / achieve a larger cobalt cellhouse delta. Additionally, both Micon and SGS Lakefield produced cobalt in the form of brittle sheet cathode, which is difficult to handle and can shatter easily. Due to the natural self-stripping characteristics of cobalt sheets from the stainless steel cathodes, it was decided that circular shaped cobalt disks known as „pucks‟, with a diameter of 2.5 cm, be electrowon.

Problems also were encountered during the EW pilot runs, such as excessive levels of chlorides, low current efficiency, lead contamination and other impurities in the cobalt cathodes, etc. In the early phase of the test program, large volumes of black sludge accumulated at the bottom of the electrolytic cell. Analysis of the sludge material showed a high cobalt content, indicating probably that there were insufficient manganese ions (Mn²+) present in the electrolyte solution for proper cobalt electrowinning. This issue was resolved by dosing manganese into the electrolyte solution.

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Of the cobalt cathode that was produced in the first batch, it was found that the quality specification was not met, especially with respect to zinc. It is suspected that the IX resins utilized during that period operated unsatisfactorily in removing impurities, and were the primary cause of the unsuccessful cobalt electrolysis. As many of the problems were resolved during the test program, the purity of the cathodes improved with each successive EW cycle. Ultimately toward the end of the test program, almost all the cathodes met and in many cases exceeded the standard LME specifications for cobalt cathodes.

Cobalt Sulphate Heptahydrate Circuit

Fortune has conducted extensive marketing of its NICO project in Asia. This work determined that there is a significant demand for a cobalt sulphate hetahydrate product used to make high performance lithium ion and nickel metal hydride rechargeable batteries and that this product typically receives a premium price for the contained cobalt that averages about 21%. A testwork campaign was initiated at SGS Lakefield, including pilot plant production of more than 10kg of cobalt sulphate heptahydrate in March and April, 2012. Fortune already had 2,000 litres of cobalt pregnant solution that had been processed by acid pressure leach in an earlier autoclave pilot plant test and was used to conduct the current test. Notably, most of the flow sheet that Fortune has already piloted and engineered to produce a 99.8% cobalt cathode product is also used to produce a high purity cobalt sulphate heptahydrate crystal product. However, after removing metal impurities from the cobalt solution using sequential neutralization, solvent extraction is used to reject metal impurities, leaving a pure cobalt sulphate solution, which is then evaporated and subjected to a three-stage crystallization process instead of electro-winning metal. Approximately 10 kg of high quality cobalt sulphate heptahydrate product containing in the range of 19.3 to 20.5% was produced in the test to prove the process flow sheet and product quality. The test result data allow for commercial product design to consistently return in excess of 20.5% Co in the Hepta hydrate. Production of cobalt sulphate was added as a sensitivity option in the FEED study by Jacobs and other engineering companies.

Bismuth Circuit

The original Micon study flow sheet did not include a circuit for leaching and recovering bismuth. The bismuth concentrate material is cyanided to recover the gold content, filtered and packaged for sale to a smelter. After the completion of the Micon study, SGS Lakefield was retained to perform various leaching tests on the bismuth concentrate. Bench scale countercurrent ferric chloride leaching and cementation was shown to be successful for recovery of the bismuth into a saleable product.

During the pilot plant test at SGS Lakefield, pressure oxidation and brine leaching of the bismuth concentrate was suggested as an alternative to ferric chloride leaching. Bench testwork showed that at a pressure oxidative environment of 180°C for 1 hour, followed by a solid-liquid separation and then 2 hours of brine leach, bismuth recovery can reach as high as 99.0%. One major advantage of pressure oxidation is the possible recovery of cobalt and copper from bismuth concentrate, which are not recoverable through the ferric chloride leaching route. The bismuth autoclave discharge solution containing the cobalt and copper would be bled into the cobalt circuit by combining with the solution from the cobalt autoclave. Pilot testwork was performed on this proposed flow sheet. The high sulphur content in the bismuth concentrate was proven to be problematic. Significant scaling of element sulphur occurred, causing numerous plugging problems and shutdowns of the pilot autoclave. The two-stage countercurrent brine leach testwork was challenging to operate as well. Further, cyanide destruction of the bismuth P&E Mining Consultants Inc., Report No. 247 Page 88 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. autoclave discharge residue also proved to be difficult. Due to all the challenges encountered with the bismuth concentrate, the strategy of pressure oxidation and brine leach was abandoned.

SGS Lakefield then conducted conventional thickening and filtration test work on bismuth leach residue into 2009. Thickening test work was conducted at a temperature of 60°C using Magnafloc 455 as a flocculant. Two series of filtration tests were conducted by SGS Lakefield on bismuth leach residue. The first series was conducted at room temperature and excluded washing; these data are useful only to generally support selected form times. The second series included washing and was conducted at 60°C. Wash efficiency was determined by following iron analyses. Bismuth tends to precipitate on dilution and the initial wash requires chloride and acid. Bench scale electrowinning tests on pregnant solution prepared by brine leaching and ferric chloride leaching was conducted to develop basic EW conditions and provide proof of concept. Observations, which were oriented toward production of a solid metal cathode product, were used to guide the next phase of test work.

Current efficiency in the test work was generally satisfactory, with efficiency falling substantially as the solutions became depleted in bismuth. Leach solutions obtained by single and two-stage leaching of concentrate were subjected to electrowinning at two temperatures. The single stage leach solution contained residual ferric iron, which resulted in unsatisfactory current efficiency. This confirmed the need for two-stage leaching. Good quality cathode assaying 99%+ Bi was obtained in both tests. A temperature of 50°C and a target feed concentration of 100 g/L was adopted for further work. It was clear from the previous test work that ferric iron could not be tolerated in the catholyte, and that separation of catholyte and anolyte with a membrane would allow the iron oxidation reaction to complete the circuit and regenerate ferric iron for re-use in leaching. In the absence of sufficient ferrous iron the generation of chlorine or oxygen can occur at the anode.

Two types of membranes were evaluated: an anionic (quaternary amine) exchange membrane (Ionac® MA-7500) and a low permeability filter cloth; and MA-7500 was selected for further studies. Subsequent test work evaluated a number of physical membranes of various permeabilities; because of the low relative flow rate through the EW cell under design conditions, none of these latter membranes proved suitable and the essentially impermeable MA- 7500 material was adopted for design purposes.

Three anode materials were evaluated and compared on the basis of iron oxidation efficiency; titanium, platinized titanium, and graphite. No chlorine generation was evident in any test.

Later in 2009, an initial series of locked-cycle tests to demonstrate the viability of the Chloride Leach and Electro Recovery (“CLER”) process flow sheet for extracting Bi from the NICO flotation concentrate were carried out. Counter-current leach cycles, initially using synthetic fresh ferric chloride solution, were conducted to generate solution for continuous EW tests. Leaching was conducted at a chloride concentration of 120 g/ L NaCl and a temperature of 95 – 100 °C with slurry density and solution concentrations controlled to yield a 3:1 molar ratio of ferric iron to bismuth in solution in Stage 2 feed. The two stage leach is necessary for control of iron oxidation. The purpose of the first stage is primarily to convert all ferric iron to ferrous using the reducing capacity of the concentrate; the degree of bismuth leaching in this stage is not important. A smooth deposit of Bi was not plated and a powder product was occasionally formed.

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The spent solution from the scavenger cell was subjected to iron/arsenic removal test work using hydrated lime. Minor arsenic leached in Stage 2 yielded a solution containing +/- 100mg As/L. During Stage 1 leaching, the arsenic concentration in solution was typically reduced to <10 mg/L. The low arsenic concentration of Stage 1 results in an effluent solution from the scavenger cell that is very low in arsenic and no difficulty is expected in reducing the concentration to environmental standards. The solution also contains iron in the ferric state and preliminary tests indicated that no additional oxidant would be required, although some peroxide was added during the latter half of the test.

Additional CLER and continuous EW work was conducted in early 2010 to evaluate the production of a powder bismuth product, alternate electrodes, and membrane materials. Subsequently, melting tests were undertaken on cathode powder product. The primary objective for these tests was to demonstrate satisfactory production of a bismuth powder which would avoid uncertainties regarding the physical character of solid cathode metal. Production of a powder would also simplify the metal recovery operation and allow operation at higher current density with a consequent reduction in the capital cost.

Tests of alternate (physical) membranes were unsuccessful and the original MA-7500 IX membrane was adopted. Two types of dimensionally stable anodes (DSA) supplied by De Nora Tech were evaluated. DSA anodes consist of titanium substrate with a proprietary mixed metal oxide solution of a precious metal – in this case Palladium (Ti-Pd). A total of 10 EW-Leach cycles were conducted with variable cathode material, current density, and recirculation rates.

Average current efficiency for Bi EW of approximately 95.6% was obtained which was not affected by current density or the material used as cathode. Based on the highest anodic current density evaluated, a design current density of 200 A/m2 was selected. Further test work may allow an increase in this criterion.

The results indicated that the anode metallurgical (i.e. oxidation) efficiency, on average, exceeded the anodic current efficiency whereas they are expected to be equal. This effect was confirmed in a series of smaller tests and was shown to be related to the DSA anodes used which appeared to have a direct or indirect catalytic effect on the oxidation of iron.

Following the primary EW test work, the effect of current density on scavenger cell operation (treating the primary EW bleed stream) was investigated.

In all cases the evolution of H2 (g) in the cell was high; particularly in Test 1. The constant evolution of H2 (g) caused floating of the bismuth metal being produced ultimately forming a layer covering all the cathodic compartment area. This situation was less problematic during the late stages of Test 2 but still quite significant. A current density of less than 100 A/m2 was selected for design of the scavenger cells.

The oxidation of iron was in all cases 100% with the significant excess of power producing chlorine and/or oxygen at the anode.

Melting tests of the resulting powders from both the EW and scavenger cells confirmed a high purity bismuth ingot would be produced and that 99.99% pure was achievable.

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Gold Circuit

As previously mentioned, the gold circuit was not tested at the continuous pilot plant scale. Rather, bottle roll tests were conducted on samples taken from the pilot plant. Cyanidation tests were completed on bismuth concentrate samples, as well as on cobalt POX residue combined with cleaner tails.

Bismuth Concentrate Gold Circuit

Initial testwork on bismuth concentrate showed inconsistent gold recoveries, which most probably was due to a passivation effect from flotation reagents. A few treatment strategies were tested, such as high shear mixing and pretreating with iron powder to imitate regrinding with steel media instead of ceramic. The passivation effect eventually was overcome by adding a heat treatment stage, where the concentrate is kept at 60°C for four hours before cyanidation. Cyanidation residence time was increased to 72 hours, compared to 48 hours in the Micon study.

Cobalt Residue Gold Circuit

In the Micon study, the autoclave discharge residue reported to cobalt residue gold cyanidation. This residue was combined with the cleaner tails, at a 50 / 50 blend ratio, to recover gold from both materials. The overall gold extraction was calculated based on individual recoveries from the two streams.

Cleaner tails remain at the NICO concentrator and are not available for gold recovery; hence, only the autoclave discharge residue is treated for cyanidation. Bottle roll tests were done to verify gold recoveries and extraction.

After the completion of the pilot plant testwork, a trade-off study compared carbon-in-pulp and Merrill Crowe for the recovery of gold in cyanide solution. It was decided from the result of the trade-off study that Merrill Crowe should be used. The PLS solution first will be deaerated in a column before being contacted with zinc powder to precipitate gold. A design addition rate of 30:1 Zn to Au molar ratio is used, to reach a gold precipitation recovery of 99.7%. The gold precipitates are recovered from the barren solution by filtration.

Combined Residue Gold Circuit

In 2011, it was decided to feed the bismuth leached residue and the cobalt concentrate to the autoclave POX and treat the autoclave residue product through cyanidation and Merrill Crowe circuit. The cyanidation test work on the autoclave residue seemed to confirm the previous test work results on gold recovery from cyanidation of cobalt residue at around 95. This process route is an alternative to the bismuth residue cyanidation and the gold electrowinning circuits and allows the recovery of cobalt and copper from bismuth leach residue, which will otherwise be lost to tails.

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Cyanide Destruction (Tailings Treatment)

Cyanide destruction for both the Micon study and the SGS Lakefield lab work were based on the INCO SO2 / air process. Small-scale continuous testwork was completed on a blend of cyanided residues from cobalt POX, cleaner tails, and bismuth in the approximate ratios expected to be seen in the plant (3.1, 12.7, 0.2 kg respectively).

Because of the replacement of CIP with Merrill Crowe, the stream requiring cyanide destruction is a solution stream, free of solids. The INCO SO2 / air process, which typically is used on slurry stream, may not be the most suitable process for liquid cyanide destruction. Additional cyanide detoxification testwork was arranged to determine the most effective destruction strategy.

Cyanide detoxification with hydrogen peroxide was allowed for, using standard industrial practices and consumption as the design criteria. Allowance was made to add, in an agitated tank, 2 to 8 g of hydrogen peroxide per gram of cyanide ion (CN¯) to reach a target residual total cyanide level of 1 mg/L at discharge.

13.7.3 Pressure Oxidative Autoclave Corrosion

Corrosion Test

The material of construction for the pressure oxidation autoclave was a major focus for the SMPP plant design. The high maintenance associated with a conventional brick-lined pressure oxidation autoclave prompted an investigation into alternative materials. SJC Materials Engineering Ltd. was retained in 2008 to perform corrosion testing for a range of alloys to determine the best material for this application. The alloys tested included three super duplex stainless steel materials (2507, Zeron 100 and Ferrallium), two nickel-based alloys (Hastelloy C22 and Inconel 686), and Titanium Grade 2. The original solution produced from the pilot plant test program at SGS was used for this corrosion test, but with a spiked acid and chlorides concentration to simulate potential operational upset conditions in the SMPP autoclave. The corrosion testing results showed significant corrosion for all tested materials in both the vapour and the liquid conditions, with the exception of titanium which performed well in both phases.

Based on this corrosion test, three different materials for the autoclave were proposed for economic comparison: titanium shell, titanium clad carbon steel shell, and acid brick-lined carbon steel shell. For the two titanium options, different grades of titanium also were considered for further improvements in crevice corrosion resistance. Grades 2, 12, 16 and 26 were considered for the titanium shell option, while the softer Grades 1, 17 and 27 were considered for the titanium-clad option. It was determined from the comparison that the titanium clad autoclave with carbon steel shell is the most economical option.

Erosion and Crevice Corrosion Test on Titanium

Based on the success with titanium in the SMPP autoclave environment, further testwork was performed in 2009 and 2010 at Xstrata Process Support Laboratories (XPS) to determine the effect of erosion and crevice corrosion on various grades of titanium. Six different grades of titanium, namely Grades 2, 3, 7, 12, 16 and 26, were investigated. These grades were chosen because they are more readily available, and are softer, which makes them suitable for fabrication of a titanium clad autoclave. To determine the effect of crevice corrosion, each of the six titanium coupons were mounted on specially-designed autoclave agitator blades, submerged P&E Mining Consultants Inc., Report No. 247 Page 92 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. in the original pilot plant solution with spiked acid and chlorides for four weeks. To understand the effect of erosion, the test autoclave also was loaded with 15 wt% slurry, which represents the slurry density at the back end of the autoclave. A second round of tests with 30 wt% solids was performed later to simulate the conditions inside the first autoclave compartment. Both sets of test results confirmed negligible general corrosion, and no development of crevice corrosion on all grades of titanium tested; hence, all are suitable for autoclave cladding.

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14.0 MINERAL RESOURCE ESTIMATES

14.1 INTRODUCTION

The mineral resource estimate presented herein has been prepared following the guidelines of the Canadian Securities Administrators‟ National Instrument 43-101 and Form 43-101F1 and in conformity with generally accepted “CIM Estimation of Mineral Resource and Mineral Reserves Best Practices” guidelines. Mineral resources have been classified in accordance with the “CIM Standards on Mineral Resources and Reserves: Definition and Guidelines” as follows:

 Inferred Mineral Resource: “An „Inferred Mineral Resource‟ is that part of a mineral resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes.”  Indicated Mineral Resource: “An „Indicated Mineral Resource‟ is that part of a mineral resource for which quantity, grade or quality, densities, shape and physical characteristics, can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough for geological and grade continuity to be reasonably assumed.”  Measured Mineral Resource: “A „Measured Mineral Resource‟ is that part of a mineral resource for which quantity, grade or quality, densities, shape, and physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough to confirm both geological and grade continuity.”

Mineral resources are not mineral reserves and do not have demonstrated economic viability. There is no guarantee that all or any part of the mineral resource will be converted into a mineral reserve. Confidence in the estimate of Inferred Mineral Resources is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure.

All mineral resource estimation work reported herein was carried out by Fred H. Brown, CPG, Pr.Sci.Nat., and Eugene Puritch, P.Eng., independent Qualified Persons in terms of NI 43-101. This mineral resource estimate is based on information and data supplied by Fortune.

Mineral resource modeling and estimation were carried out using the commercially available GEMS Gemcom v5.23 and Snowden Supervisor v7.10.11 software programs.

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14.2 PREVIOUS MINERAL RESOURCE ESTIMATES

A previous mineral resource estimate (Table 14.1) was released in November 20041 , and subsequently used in a bankable feasibility study in 20072 . These estimates have now been superseded by the current mineral resource estimate.

TABLE 14.1 HISTORICAL MINERAL RESOURCE ESTIMATE DATED NOVEMBER 2004 NSR Cut- Tonnes NSR Area Class Au g/t Bi % Co % off $CDN/t x 1000 $CDN/t Measured 2,718 0.46 0.155 0.120 $32.76 Open $20 Indicated 5,513 0.49 0.126 0.137 $35.11 Pit M+I 8,231 0.48 0.136 0.131 $34.33 Measured 1,382 3.97 0.192 0.129 $78.17 U/G $50 Indicated 3,741 3.25 0.223 0.170 $79.86 M+I 5,123 3.44 0.210 0.160 $79.40

14.3 SAMPLE DATABASE

Sample data were provided in the form of Excel spreadsheets and Access databases. The supplied databases contain 25,055 assay records from 325 drillholes (Table 14.2). As implemented by P&E, 299 drillhole records were used for this mineral resource estimate, and each drillhole record consists of collar, survey and assay data. Assay data fields consist of the drillhole ID, downhole interval distances, sample number, and individual element grades. Sampling data for twenty-two trenches were also included; however, trench data were used for visual domaining only and not for estimation. All data are in metric units, and have been converted to the NAD83 coordinate system.

TABLE 14.2 NICO DRILLING DATABASE RECORDS Phase Number Metres 2009 288 55,541.37 2010 37 6,112.00

Total 325 61,653.37

Industry standard validation checks were completed on the supplied database, and minor corrections made where necessary. P&E typically validates a mineral resource database by checking for inconsistencies in naming conventions or analytical units, duplicate entries, interval, length or distance values less than or equal to zero, blank or zero-value assay results, out-of- sequence intervals, intervals or distances greater than the reported drill hole length, inappropriate collar locations, and missing interval and coordinate fields. No significant discrepancies with the supplied data were noted. P&E believes the database to be suitable for mineral resource estimation.

1 Micon 2004. An updated mineral resource estimate for the NICO cobalt-gold-bismuth deposit, Mazenod Lake District, Northwest Territories, Canada. 2 Micon 2007. Technical report on the bankable feasibility study for the NICO cobalt-gold-bismuth deposit, Mazenod Lake District, Northwest Territories, Canada. P&E Mining Consultants Inc., Report No. 247 Page 95 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

14.4 DOMAIN MODELING

Three mineralization domains and five lithological domains were modeled (Table 14.3). Domain models were generated from successive polylines oriented perpendicular to the trend of the mineralization and spaced every five metres along the strike of the deposit. Mineralization domains were defined by continuous mineralization and assay intervals equal to or greater than a calculated NSR value of $40.00.

Where necessary a small number of lower value assays have been included in order to preserve continuity. All polyline vertices were snapped directly to drillhole assay intervals, in order to generate a true three-dimensional representation of the extent of the mineralization. A topographic surface of unknown resolution was also supplied, and all domain wireframes were clipped to this surface. The resulting mineralization domains were used as hard boundaries during estimation, and for rock coding, statistical analysis and compositing limits

A total of 5,966 bulk density measurements were provided by Fortune, with an average bulk density of 3.18 t/m3 (Table 14.3). Bulk density values were back-tagged to lithological domains and used to determine average bulk density values for mineral resource estimation.

TABLE 14.3 MODELED DOMAINS Domain Rock Code Average Bulk Density (t/m3) Lower Mineralization Domain 10 3.25 Middle Mineralization Domain 20 3.21 Upper Mineralization Domain 30 2.94 QAP Dykes 100 3.05 FAP Dykes 200 3.05 Volcanics 300 3.03 Wackes 400 3.12 Siltstone 500 3.37

14.5 COMPOSITING

Drillhole assay samples display bimodal sampling lengths, averaging 1.00 m and 3.00 m in length respectively. In order to provide equal sample support length-weighted 3.00 m composites were calculated for all elements within the defined mineralization domains. A small number of unsampled intervals were assigned a nominal grade of 0.0001 for compositing purposes. The compositing process started at the first point of intersection between the drillhole and the domain intersected, and halted upon exit from the domain wireframe. Composites that were less than 1.00 m in length were discarded so as to not introduce a short sample bias into the estimation process. The wireframes that represent the interpreted mineralization domains were also used to back-tag a rock code field into the composite workspace. The composite data were then exported to Gemcom extraction files for grade estimation, and summary composite statistics were calculated by domain for each commodity (Table 14.4).

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TABLE 14.4 NICO UNCAPPED COMPOSITE STATISTICS BY DOMAIN Au Composites Statistic Domain 10 Domain 20 Domain 30 Total Samples 1837 1515 523 3875 Minimum 0.0001 0.0001 0.0001 0.0001 Maximum 43.45 47.17 8.63 47.17 Mean 0.84 0.77 0.15 0.72 Standard Dev 2.83 2.52 0.47 2.52 CV 3.39 3.29 3.12 3.52 Bi Composites Statistic Domain 10 Domain 20 Domain 30 Total Samples 1837 1515 523 3875 Minimum 0.0001 0.0001 0.0001 0.0001 Maximum 2.73 1.41 1.03 2.73 Mean 0.12 0.11 0.04 0.10 Standard Dev 0.20 0.15 0.10 0.17 CV 1.65 1.38 2.30 1.63 Co Composites Statistic Domain 10 Domain 20 Domain 30 Total Samples 1837 1515 523 3875 Minimum 0.0001 0.0001 0.0001 0.0001 Maximum 1.02 1.12 0.91 1.12 Mean 0.11 0.08 0.04 0.09 Standard Dev 0.13 0.09 0.08 0.11 CV 1.16 1.22 1.84 1.28 Cu Composites Statistic Domain 10 Domain 20 Domain 30 Total Samples 1837 1515 523 3875 Minimum 0.0001 0.0001 0.0001 0.0001 Maximum 0.88 0.59 1.17 1.17 Mean 0.03 0.03 0.05 0.03 Standard Dev 0.07 0.06 0.10 0.07 CV 2.14 1.96 2.13 2.11

14.6 TREATMENT OF EXTREME VALUES

The presence of high-grade outliers was evaluated from grade-capping curves and by the review of histograms and log-probability graphs of the grouped composite grade data. For Au, estimation was done using Indicator Kriging, and a capping threshold for Au was therefore derived from the high-grade bin. For Bi the capping threshold selected was equivalent to the 99.87 percentile, and the Co and Cu capping thresholds were set to the same percentile. Composite values were capped to the selected threshold value prior to estimation (Table 14.5).

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TABLE 14.5 NICO CAPPING AND THRESHOLD VALUES Commodity Threshold Number Capped Au 24.00 g/t 8 (0.2%) Bi 1.40 % 5 (0.1%) Co 0.94 % 4 (0.1%) Cu 0.71 % 5 (0.1%)

14.7 CONTINUITY ANALYSIS

For Bi, Co and Cu, directional experimental semi-variograms were modeled from uncapped composite data using a normal-scores transformation (Table 14.6). The downhole variogram was viewed at a 3.00 m lag spacing (equivalent to the composite length) to assess the nugget variance contribution. Nugget and standardized spherical models were used to model the experimental semi-variograms in normal-score transformed space. Semi-variogram model ranges were then checked and iteratively refined for each model relative to the overall nugget variance, and back- transformed variance contributions were calculated for grade interpolation. Continuity ellipsoids based on the semi-variogram models were then generated for each variable and used to define the appropriate search strategy.

For Au, directional indicator semi-variograms were modeled from uncapped composite data based on a 1.50 g/t indicator threshold (Table 14.6). Experimental indicator semi-variograms for each of the three principle directions were calculated, and in general the strike and dip directions were aligned with observed mineralization trends, with the cross-strike direction being variable.

TABLE 14.6 NICO EXPERIMENTAL SEMI-VARIOGRAMS Commodity Strike Dip Cross Semi-Variogram Au 80 m 80 m 10 m 0.4 + Sph(0.40, 80) Bi 120 m 40 m 40 m 0.14 + Sph(0.86, 140) Co 120 m 40 m 40 m 0.37 + Sph(0.63, 120) Cu 120 m 40 m 40 m 0.16 + Sph(0.84, 120)

14.8 BLOCK MODELS

A block model was established across the property based on a 5 m x 5 m x 5 m block size (Table 14.7). The block model consists of separate sub-models for estimated grades, indicator kriging probabilities, associated rock codes, percent, density and classification attributes, and a calculated NSR value. A percent block model was used to accurately represent the volumes and tonnages that were contained within the respective mineralization domains.

TABLE 14.7 NICO BLOCK MODEL SETUP Axis Minimum Maximum Size (m) Number X 510,907.50 513,307.50 5 480 Y 7,046,793.59 7,048,293.59 5 300 Z -100.00 400.00 5 100 Rotation -20.112° P&E Mining Consultants Inc., Report No. 247 Page 98 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

14.9 ESTIMATION & CLASSIFICATION

The mineral resource estimate was constrained by wireframes that form hard boundaries between the respective composite data files. Individual block grades were used to calculate a NSR block grade model.

Block estimates for Au were calculated using non-linear Indicator Kriging of capped composite grades. Based on the defined 1.50 g/t indicator semi-variogram, for each block a high-grade probability, high grade estimate and low-grade estimate were calculated and then combined into a single block estimate. Bi, Co and Cu were estimated using linear Ordinary Kriging of capped composite grades.

A three-pass series of expanding search ellipses with varying minimum sample requirements was used for sample selection, estimation and classification:

During the first pass, five to twelve composite values from two or more drillholes within a search ellipsoid corresponding to 45% of the defined Bi variogram ranges were required for estimation. All block grades estimated during the first pass were classified as Measured.

During the second pass, blocks not populated during the first pass were estimated. Five to twelve composite values from two or more drillholes within a search ellipsoid corresponding to 100% of the defined Bi variogram ranges were required for estimation. All block grades estimated during the second pass were classified as Indicated.

During the third pass, blocks not populated during the first or second pass were estimated. Three to twelve composite values from one or more drillholes within a search ellipsoid corresponding to about 300% of the defined range were required for estimation. All block grades estimated during the third pass were classified as Inferred.

14.10 MINERAL RESOURCE ESTIMATE

In order to ensure that the reported mineral resources meet the CIM requirement for “reasonable prospects for economic extraction”, a conceptual floating-cone optimized pit shell was developed based on all available mineral resources (Measured, Indicated and Inferred), using the economic parameters listed in Table 14.8. The results from the optimized pit-shell are used solely for the purpose of reporting mineral resources that have reasonable prospects for economic extraction. The gold price used is based on the 36-month trailing average as of November 2011. Copper was not included in the mineral resource estimate.

TABLE 14.8 ECONOMIC PARAMETERS Gold Price USD 1,250.00 / oz Bismuth Price USD 10.00 / lb Cobalt Price USD 20.00 / lb Mining Cost CAD 2.50 / t Underground Mining Cost CAD 32.00 / t Processing Cost + G&A CAD 46.00 / t Exchange Rate 0.95 Pit Wall Slope Angle 45°

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TABLE 14.9 NSR METAL UNIT VALUES Commodity Recovery % NSR $C / tonne 0.3 g/t >= Au < 0.5 g/t 56 23.31 0.5 g/t >= Au < 1.5 g/t 60 24.98 1.5 g/t >= Au < 2.5 g/t 70 29.14 2.5 g/t >= Au < 3.5 g/t 74 30.80 3.5 g/t >= Au < 4.5 g/t 79 32.89 Au >= 4.5 g/t 84 34.97 Co % 83 385.22 Bi % 70 162.44

All open pit mineral resources are reported against a $46.00 NSR cut-off, as constrained within the optimized pit shell. Underground mineral resources are reported outside the optimized pit shell against an $80 NSR cut-off (Table 14.10). The effective date of this estimate is November 30, 2011.

TABLE 14.10 (1)(2)(3) NICO ESTIMATED MINERAL RESOURCES Area NSR Cut-off $CDN/t Class Tonnes x 1000 Au g/t Bi % Co % Measured 18,911 1.05 0.15 0.12 Open Indicated 10,983 1.19 0.14 0.12 $46 Pit M+I total 29,894 1.10 0.15 0.12 Inferred 2 0.30 0.07 0.08 Measured 231 2.29 0.06 0.15 Indicated 764 1.72 0.07 0.16 U/G $80 M+I total 995 1.85 0.07 0.16 Inferred 31 0.65 0.11 0.25 (1) Mineral resources are defined within an optimized pit shell that incorporates project metal recoveries, estimated operating costs and metals price assumptions. (2) Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, marketing, or other relevant issues. The mineral resources in this news release were estimated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council. (3) The quantity and grade of reported Inferred resources are uncertain in nature and there has been insufficient exploration to define these Inferred resources as an Indicated or Measured mineral resource and it is uncertain if further exploration will result in upgrading them to an Indicated or Measured mineral resource category.

14.11 MINERAL RESOURCE VALIDATION

The block model was validated visually by the inspection of successive section lines in order to confirm that the block model correctly reflects the distribution of high-grade and low-grade samples.

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An additional validation check was completed by comparing model block grade estimates to the average grade of capped composites used for estimation (Table 14.11). P&E believes the mineral resource estimate to be reasonable and robust.

TABLE 14.11 VALIDATION STATISTICS FOR CAPPED COMPOSITES AND BLOCK ESTIMATES Commodity Capped Composite Average Block Estimate Au g/t 0.693 0.709 Bi % 0.104 0.101 Co % 0.089 0.089 Cu % 0.034 0.034

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15.0 MINERAL RESERVE ESTIMATES

15.1 MINERAL RESERVES CRITERIA

Mineral reserves for the open pit and underground mining operation have been determined based upon a mine plan, mine dilution, mine extraction and operating costs estimated for an annual production rate of 1,698,400 t, based on 4,650tpd, 365.25 days /yr. Table 15.1 presents the parameters that have been used to determine both the open pit and underground resources and reserves.

TABLE 15.1 RESOURCE TO RESERVE CONVERSION PARAMETERS Smelter/Refining/ Price US$ Metallurgical Smelter NSR (C$/%/t Commodity Transportation Charges (US$/oz or /lb) Recovery (%) Payables (%) or C$/g/t) (C$/Oz; C$/lb) Gold < 0.5 1,200 63% 100% 0 22.74 g/t Gold – 0.5- 1,200 67% 100% 0 24.37 1.5 g/t Gold – 1.5- 1,200 72% 100% 0 28.43 2.5 g/t Gold – 2.5- 1,200 77% 100% 0 30.05 3.5 g/t Gold – 3.5- 1,200 81% 100% 0 32.08 4.5 g/t Gold > 4.5 1,200 87% 100% 0 34.11 g/t Cobalt 17 83% 100% 0 327.44 Bismuth 9 70% 100% 0 146.20 Copper 2.36 60% 100% 0 32.86

Exchange Rate CDN$ 1.00 = USD$ 0.95

Operating Costs U/G Ore $104.64 per tonne mined Mining per tonne Processing $41.82 processed per tonne G&A $6.25 processed

15.2 NET SMELTER RETURN (NSR)

Given the polymetallic nature of the NICO deposit, the Net Smelter Return (“NSR”) has been used to determine the value of each mining block within the resource model. Mining cut-off limits are determined as NSR values based on the above parameters. Block NSR net values that exceed the cut-off value are classified as resources that can be converted to reserves. The NSR value per tonne for each block in the model has been determined on the basis of parameters that are used in a NSR calculation such as anticipated concentrate recoveries, smelter payables, smelter treatment charges, applicable refining charges, royalties, metal prices, metal price escalators and the $CAD/$US exchange rate.

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For currency conversion an exchange rate of US$0.95/CAD$ has been used. This exchange ratio is based on the May 31, 2012 three year trailing average based on the published average monthly Bank of Canada exchange rates. In terms of metal prices, they have also been estimated based on a three tear trailing average that has been adjusted downwards for certain metals for the purpose of rounding to reasonable values.

Smelter treatment, refining and transportation charges have been included in the $41.82/t process operating costs. Metal price escalation has been negated given Fortune‟s intention to construct and operate a smelter facility in the general area of Saskatoon, Saskatchewan. This facility, denoted as the SMPP, will receive bulk concentrate from the NICO mine that will undergo additional grinding and flotation to produce separate gold-bearing cobalt and bismuth concentrates, followed respectively by pressure acid leach and acid leach, hydrometallurgical processing, and subsequent electro-winning to gold doré, 99.8% cobalt cathode or 20.9% cobalt sulphate heptahydrate, 99.99% bismuth ingot and a copper cement by-product. Bulk concentrate would be transported from the mine site by truck and then rail to the SMPP.

15.3 UNDERGROUND RESERVE

The underground reserve estimate is based on dilution and extraction to planned stope mining outlines. The underground mineral reserve estimate is shown in Table 15.2 and has been determined based on targeting high gold grade values.

TABLE 15.2 (1)(2) UNDERGROUND RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 282 4.93 0.14 0.27 Probable 94 5.60 0.11 0.19

Total 376 5.09 0.13 0.25 (1) Mine recovery and dilution are included in these quantities along with metal grades (2) All of the material designated as Reserves in the underground portion was derived from Measured and Indicated Resources that have been demonstrated to be economic as result of the current study and are therefore designated as Proven and Probable Reserves using the Canadian Institute of Mining, Metallurgical and Petroleum (CIM) Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.

15.4 OPEN PIT RESERVE

The open pit reserve estimate is based on dilution and extraction to bench defined mining outlines. The open pit mineral reserve estimate is shown in Table 15.3 and has been determined based on selection of blocks that are above the marginal economic NSR cut-off of C$48.07 per tonne.

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TABLE 15.3 (1)(2) OPEN PIT RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 20,513 0.94 0.11 0.15 Probable 12,099 1.05 0.11 0.13

15.5 TOTAL RESERVE

The total combined underground and open pit reserves are presented below in Table 15.4 and are based on dilution and extraction to defined mining outlines.

TABLE 15.4 (1)(2) TOTAL RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 20,795 0.99 0.11 0.15 Probable 12,193 1.09 0.11 0.13

15.6 RESPONSIBILITY FOR ESTIMATION

The mineral reserve estimate presented in this report was prepared by Eugene Puritch, P.Eng. and James L Pearson P.Eng. of P&E Mining Consultants Inc. using process costs and recoveries provided by Fortune in conjunction with Jacobs, and other metallurgical consultants. Mining geotechnical parameters were provided by Golder.

Eugene Puritch, P.Eng. and the Associates of P&E mentioned herein have no relationship with Fortune other than that of independent consultants.

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16.0 MINING METHODS

16.1 OPEN PIT

16.1.1 Conventional Open Pit Operation

The open pit will be a conventional mining operation utilizing proven drilling and blasting and loading and haulage equipment and technologies. The pit will be developed and operated by the mine owner-operator using its own labour and equipment with assistance from explosive and mine equipment suppliers. The explosive supplier will provide the explosives and blasting agents and accessories to the operation, and the pit equipment will be maintained by the supplier under a repair and maintenance agreement. The mine owner-operator will provide technical services to the mine and procure mine consumables such as diesel fuel, lubricants and tires.

16.1.2 Open Pit Production Schedule

As indicated in the open pit production schedule shown in Table 16.1:

 The mine will be readied for production in months 19 and 20. Approximately 1.65 Mt of waste rock and some shallow overburden are scheduled to be stripped by the end of month 20.  Ore production and stockpiling is scheduled to commence in month 21. The plant will commence processing open pit ore in month 22.

The underground mine will operate concurrently with the open pit operation commencing in month 30. It is expected that the underground operation will be conducted over an 8 month time line.

The open pit mine is scheduled to produce a total of 32.61 Mt of ore over its operating life. The open pit mill feed will include 31.98 Mt of ore that will be shipped directly from the pit to the mill, and 0.62 Mt of ore that will be stockpiled and later reclaimed and processed. A total of 97.80 Mt of waste will be produced over the 19.5 year pit life.

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TABLE 16.1 OPEN PIT PRODUCTION SCHEDULE Months Item 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 Tonnes of ore 0 0 31,733 120,718 120,718 59,568 226,578 120,120 108,619 122,151 156,969 144,921 60,909 76,384 79,029 Tonnes of low 0 0 2,259 11,407 11,407 14,673 58,694 24,456 14,096 23,469 26,571 28,163 14,000 16,825 25,237 grade waste Tonnes of waste 650,000 1,000,000 1,016,008 917,875 917,875 975,758 764,729 905,424 927,285 904,380 866,460 876,916 975,091 956,791 945,734 Tonnes reclaimed from the ore 0 0 0 0 0 81,966 0 21,414 32,915 19,383 0 0 46,455 12,784 0 stockpile

Total tonnes 650,000 1,000,000 1,050,000 1,050,000 1,050,000 1,049,999 1,050,001 1,050,000 1,050,000 1,050,000 1,050,000 1,050,000 1,050,000 1,050,000 1,050,000 mined

Months Year 4 Year 5 Years Item 34 35 36 Q1 Q2 Q3 Q4 Q5 Q6 Q7 Q8 6 7 8 Tonnes of ore 76,029 157,568 160,315 366,800 498,847 424,602 341,949 332,576 424,604 424,604 424,604 1,855,376 1,698,408 1,698,408 Tonnes of low 25,237 42,062 30,397 94,672 123,004 95,720 146,160 128,908 109,449 85,329 51,259 263,658 268,807 265,594 grade waste Tonnes of waste 948,734 850,370 859,288 2,688,529 2,528,148 2,629,678 2,661,891 2,688,515 2,615,947 2,640,067 2,674,137 5,080,965 5,232,785 5,235,998 Tonnes reclaimed from the ore 0 0 0 27,872 0 0 82,654 92,027 0 0 0 0 0 0 stockpile

Total tonnes mined 1,050,000 1,050,000 1,050,000 3,150,001 3,149,999 3,150,000 3,150,000 3,149,999 3,150,000 3,150,000 3,150,000 7,199,999 7,200,000 7,200,000

Years Item 9 10 11 12 13 14 15 16 17 18 19 20 21 22 Tonnes of ore 1,698,408 1,698,408 1,702,517 1,698,408 1,698,408 1,698,408 1,698,408 1,698,408 1,698,408 1,698,408 1,698,408 1,698,408 1,698,408 214,995 Tonnes of low grade 185,229 206,087 265,638 303,113 311,600 339,840 297,289 311,738 267,046 197,915 231,483 420,723 144,008 1,251 waste Tonnes of waste 5,316,363 3,595,505 3,531,845 3,498,479 3,489,992 3,461,752 2,504,303 2,489,854 2,534,546 2,603,677 1,645,478 2,617,352 2,028,798 71,776 Tonnes reclaimed from the ore 0 0 0 0 0 0 0 0 0 0 0 0 0 204,107 stockpile

Total tonnes mined 7,200,000 5,500,000 5,500,000 5,500,000 5,500,000 5,500,000 4,500,000 4,500,000 4,500,000 4,500,000 3,575,369 4,736,483 3,871,214 288,022

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16.1.3 Open Pit Operation

The open pit will use conventional open pit mining equipment and technologies.

Blasthole Drilling and Blasting

The proposed blasthole drilling patterns for ore and waste are shown in Table 16.2. The 127 mm (5 inch) diameter blastholes will be drilled using diesel-powered track-mounted down-the-hole drills. The blastholes will be loaded with a blended emulsion explosive, primed, and stemmed using crushed rock and drilling cuttings. The projected powder factors are shown in Table 16.2. It is expected that the mine will progressively improve and eventually optimize its drilling pattern and blasting program. The explosives, blasting agents and blasting accessories will be supplied by a licensed explosive supplier. The supplier will set-up its explosive and detonator storage magazines and operating facility in secure locations on the mine property.

TABLE 16.2 PRODUCTION DRILLING PATTERNS Item Ore Waste Bench height (m) 10 10 Blasthole diameter (mm) 127 127 Burden (m) 3.5 3.5 Spacing (m) 3.5 4.5 Collar (m) 2.5 2.5 Subdrill (m) 1 1 Explosive type 70/30 70/30 Rock density (t/m3) 3.227 3.069 Explosive density (g/cm3) 1.15 1.15 Powder factor (kg/t) 0.31 0.26 Tonnes per metre drilled (t/m)(1) 35.2 43.1 (1) Includes a 2% allowance for blasthole re-drilling and clean-out.

Loading and Haulage

The blasted ore will be excavated and hauled to the ore crusher at the mill or stockpiled in the vicinity of the ore crusher. The blasted waste rock will be excavated and hauled to the co- disposal facility. The proposed main loading and haulage equipment includes:

 One 9.5 m3 capacity diesel-powered hydraulic Komatsu PC2000 type backhoe arrangement excavator. It is scheduled to be in service to the end of year 9.  An 11 m3 capacity Komatsu WA900 type wheel loader is scheduled to come into service in month 21. A second unit is scheduled to come into service in year 10.  Eight 91 t capacity Komatsu HD785 type haul trucks will be procured over the mine life.

See section 18 for information on the open pit shop, warehouse, mine office and dry.

16.1.4 Open Pit Labour

The projected number of open pit operations and maintenance personnel are shown in Table 16.4.

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TABLE 16.3 OPEN PIT EQUIPMENT SCHEDULE Month/Year 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 Yr 4 Q1 Yr 4 Q2 Yr 4 Q3 Yr 4 Q4 Production drill - Sandvik D25KS type 1 1 Hydraulic backhoe - Komatsu PC2000 type 1 Wheel loader - Komatsu WA900 type 1 Haul truck - Komatsu HD785 type 3 1 1 Bulldozer - Komatsu D375 type with ripper 1 Grader - Komatsu GD655 type 1 Water truck / sander 1 Blast Hole Loader (modify existing IT unit) 1 Wheel loader - WA600 type 1 IT carrier - Komatsu WA250PZ type 1 Tire manipulator attachment 1 Boom truck 1 Mechanic service truck 1 Welding truck 1 Fuel and lube truck 1 Pick-up trucks (open pit department) 1 2 1 Shop equipment Lot Lot Portable lighting 1 1 Pit dewatering pumps and pipelines Lot Pit communication radios Lot Lot

Yr 5 Yr 5 Yr 5 Yr 5 Month/Year 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 LOM Q1 Q2 Q3 Q4 Production drill - Sandvik D25KS type 1 3 Hydraulic backhoe - Komatsu PC2000 type 1 Wheel loader - Komatsu WA900 type 1 2 Haul truck - Komatsu HD785 type 3 8 Bulldozer - Komatsu D375 type with ripper 1 2 Grader - Komatsu GD655 type 1 2 Water truck / sander 1 1 Blast Hole Loader R 1 Wheel loader - WA600 type 1 2 IT carrier - Komatsu WA250PZ 1 2 Tire manipulator attachment R 1 Boom truck 1 2 Mechanic service truck 1 2 Welding truck 1 2 Fuel and lube truck 1 2 Pick-up trucks (open pit department) 2 2 1 2 11 Shop equipment Lot Portable lighting 1 1 1 5 Pit dewatering pumps and pipelines Lot Lot Lot Lot Lot Pit communication radios Lot Lot Lot Note: „R‟ indicates that an equipment rebuild cost allowance has been included in the sustaining capital costs.

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TABLE 16.4 OPEN PIT PERSONNEL Months Year 4 Position 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 Q1 Q2 Q3 Pit Superintendent 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Pit Supervisor 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 Geologist 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Mining Engineer 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Blast Technician 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 Technician 2 2 2 2 2 2 2 2 2 2 2 2 4 4 4 4 4 4 4 4 4 Surveyor 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Production Driller 4 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 Explosive Supplier Personnel 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 7 Blasthole sampler / stemmer 1 1 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2

Wheel Loader Operator 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 Excavator Operator 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 Haul Truck Driver 12 18 20 18 20 18 18 20 18 20 18 20 20 18 18 18 18 18 20 18 20 Ancillary Equipment Operator 2 4 4 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8 8

Maintenance contract mechanics 11 17 18 19 20 19 19 20 19 20 19 20 20 19 19 19 19 19 20 19 20 Tire Technician 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2

Total pit personnel(1) 55 77 81 84 87 84 84 87 84 87 84 87 89 86 86 86 86 86 89 86 89

Yr 4 Year 5 Years Position Q4 Q1 Q2 Q3 Q4 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 Pit Superintendent 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Pit Supervisor 4 4 4 4 4 4 4 4 4 3 3 3 3 3 3 3 3 3 3 3 3 3 Geologist 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Mining engineer 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 Blast Technician 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 Technician 4 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 Surveyor 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1

Production Driller 8 8 8 8 8 4 4 4 4 4 4 4 4 4 4 4 4 4 2 4 4 4 Explosive Supplier Personnel 7 7 7 7 7 7 7 7 7 5 5 5 5 5 5 5 5 5 5 4 4 4 Blasthole sampler / stemmer 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2

Wheel Loader Operator 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 Excavator Operator 4 4 4 4 4 4 4 4 4 Haul Truck Driver 18 20 20 20 18 10 10 10 12 10 8 8 10 10 8 8 8 10 8 10 8 8 Ancillary Equipment Operator 8 8 8 8 8 8 8 8 8 4 4 4 4 4 4 4 4 4 4 4 4 4

Maintenance contract mechanics 19 20 20 20 19 15 15 15 16 11 10 10 11 11 10 10 10 11 10 11 10 10 Tire Technician 2 2 2 2 2 2 2 2 2 2 2 2 2 2 2 1 1 1 1 1 1 1

Total pit personnel(1) 86 87 87 87 84 68 68 68 71 53 50 50 53 53 50 49 49 52 47 51 48 48 (1) The above table shows the projected number of pit personnel on the open pit payroll.

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16.2 UNDERGROUND MINING

The NICO Underground Project will be mined using conventional blast hole stoping underground mining methods, by a contractor. The mine production schedule assumes that sustained ore production will commence in July 2016. Development ore production will begin in June 2016. The underground mine plans are based on the following.

 During the underground pre-production period an underground mine contractor will mobilize and set-up its‟ equipment and facilities on surface and dewater and rehabilitate the existing underground workings. During this underground pre- production period no development or production ore will be mined.  Underground mining will be via retreat transverse and longitudinal blast hole open stoping methods, generally mined from the top down, without backfill. Main mining Levels will be at the 95, 116, 135, 141, 161, 170, 195 and 215 m levels.  The underground mining operation will average 1,544 tpd ore for a period of eight months for total production of 376,000 t grading 5.09 g/t Au, 0.25% Bi, 0.13% Co and 0.02% Cu.  The open pit operation will supply, on average, 3,106 tpd ore to the mill during the underground production period.  Prior to, and after, the completion of the underground mining program, the open pit will produce 4,650 tpd ore.  Total life-of-mine underground and open pit reserves, to be mined, is 32,987,500 t grading 1.02 g/t Au, 0.14% Bi, 0.11% Co and 0.04% Cu. In addition a total of 68,500 t of underground waste, 92,325,100 t of open pit waste and 5,484,500 t of sub economic open pit waste will be mined.

The underground mining method will be retreat transverse and longitudinal blast hole open stoping, using uppers, generally mined from the top down, without backfill. There are a total of 21 blasthole stopes. A summary of the stope mining sequence, location, names and ore production tonnages is presented in Table 16.5. The existing exploration decline, completed in 2006, will be utilized for access to the stope mining areas. Please refer to the as-built Figure 16.1, through Figure 16.3, on the following pages, showing the existing underground development.

The proposed mine plan is based on an underground development program that includes: the extension of the 5m x 5m exploration decline to the 95, 116, 135, 141, 161, 170, 195 and 215 m levels. Please refer to Figure 16.4, on the following pages, showing a longitudinal projection of the underground levels and stopes. Underground infrastructure includes the installation of an escapeway in the existing fresh air raise between the 95 m level and surface, a portable lunchroom/refuge stations, a main sump on the access ramp to the 105 level, a portable electrical substations and safety bays. Secondary sumps, and powder and cap magazines will be installed in unused muck bays. Most stope access drifts and cross-cuts will be driven in ore. The underground mine equipment maintenance shop will be located on surface.

Stopes will be left unfilled. As the open pit advances towards the underground, stopes will be backfilled with broken ore from the pit floor via drop raises or as they are exposed, and the pit will advance through the ore body, recovering pillars.

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TABLE 16.5 SUMMARY OF U/G STOPE MINING TONNAGES Level Stope Tonnes Start Date Finish Date 195 - 215 Stope 21 9,253 July 28, 2016 Aug 2, 2016 195 - 215 Stope 3 7,112 Aug 8, 2016 Aug 21, 2016 195 - 215 Stope 20 9,424 Aug 25, 2016 Aug 30, 2016 195 - 215 Stope 19 9,210 Sept 2, 2016 Sept 7, 2016

161 - 161 Stope 15 14,735 Nov 4, 2016 Nov 12, 2016

Total 198,917

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Figure 16.1 NICO - As Built Underground Workings

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Figure 16.2 NICO - As Built Underground Workings

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Figure 16.3 NICO - As Built Underground Workings

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Figure 16.4 Longitudinal Projection of the Proposed Underground Workings

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16.2.1 Underground Mine Design and Stope Layout

The plan is to minimize required development, and maximize this required development in ore. Access to the stopes will be via 5 m high x 5 m wide level access ramps, generally driven in waste, and 5 m high x 5 m wide level footwall drifts, 4.5 m high x 4 m wide extraction cross-cuts and 4.5 m high x 12 m wide under-cuts, generally driven in ore. Please refer to Figure 16.5, through Figure 16.7 for typical ramp, drift and cross-cut cross-sections. Typical parameters for a 5 m high x 5 m wide ramp heading is presented in Table 16.6.

TABLE 16.6 TYPICAL 5 M X 5 M DEVELOPMENT ROUND PARAMETERS Description Value Width, (m) 5.0 Height, (m) 5.0 Holes Drilled 70 Holes Reamed 4 Holes Loaded 66 Hole Diameter, (mm) 45 Drilled Length, (m) 3.7 Total Drilled Length, (m) 256 Load Length per Hole, (m) 3.4 Break Length, (m) 3.5 Tonnes Broken, (tonnes) 280 Powder Factor (kg/t) 1.15 Rebar Per Round 26 Length of Rebar, (m) 1.5 Rounds per day 1.9 Average Daily Advance - Single Heading (m) 6.5 Shift Duration, (hours) 10 Shifts per Day 2

Typical parameters for a 4.5 m high x 4 m wide extraction cross-cut heading is presented in Table 16.7.

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TABLE 16.7 TYPICAL 4.5 M X 4 M DEVELOPMENT ROUND PARAMETERS Description Value Width, (m) 4.0 Height, (m) 4.5 Holes Drilled 53 Holes Reamed 4 Holes Loaded 49 Hole Diameter, (mm) 45 Drilled Length, (m) 3.7 Total Drilled Length, (m) 194 Load Length per Hole, (m) 3.4 Break Length, (m) 3.5 Tonnes Broken, (tonnes) 198 Powder Factor (kg/t) 1.20 Rebar Per Round 21 Length of Rebar, (m) 1.5 Rounds per day 1.9 Average Daily Advance - Single Heading (m) 6.5 Shift Duration, (hours) 10 Shifts per Day 2

Ore pillars will be left in place between stopes, for ground support. Stoping includes drilling and blasting the slot raises and production drilling and blasting using hole drills, utilizing 51 mm diameter blast holes. All stope production drilling and blasting will be with up holes.

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Figure 16.5 Typical 5.0m High by 5.0m Wide Ramp Development Heading

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Figure 16.6 Typical 5.0m High by 5.0m Wide Drift Development Heading

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Figure 16.7 Typical 4.5m High by 4.0m Wide Cross Cut Heading

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16.2.2 Underground Production Schedule

16.2.2.1 Mine Development

TABLE 16.8 MINE DEVELOPMENT SCHEDULE SUMMARY Level Crew Start Date Finish Date 195-215 C1 June 1, 2016 Aug 2, 2016 170-195 C2 June 1, 2016 Sept 3, 2012 161 C1 July 28, 2016 Sept 5, 2016 135-141 C1 Aug 12. 2012 Oct 30, 2016 116-95 C2 Aug 24, 2012 Nov 26, 2016

16.2.2.2 Stope Development

Once the access and footwall drifts in waste have been completed to the first accessible stopes No. 21 and No. 1 on the 195-215 and 170-195 levels, development Crews 1 and 2 will proceed with stope development in to all stopes as they become accessible. Development crews will excavate undercut cross-cuts, undercut slashes, slot raises and complete stope drilling in these stopes. A summary of the stope development schedule is presented in Table 16.9.

TABLE 16.9 STOPE DEVELOPMENT SCHEDULE SUMMARY Level Crew Start Date Finish Date 195-215 C1 June 9, 2016 Sept 2, 2016 170-195 C2 June 11, 2016 Oct 25, 2016 161 C1 Aug 6, 2016 Nov 4, 2016 135-141 C1 Aug 17, 2016 Dec 17, 2016 116-95 C2 Sept 13, 2016 Jan 20, 2016

16.2.2.3 Stoping

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TABLE 16.10 PRODUCTION BLAST HOLE BLASTING, MUCKING AND HAULAGE SCHEDULE SUMMARY Level Start Date Finish Date 195-215 July 28, 2016 Sept 7, 2016 170-195 Sept 5, 2016 Oct 27, 2016 161 Nov 4, 2016 Nov 12, 2016 135-141 Nov 13, 2016 Dec 22, 2016 116-95 Dec 20, 2016 Jan 26, 2017

TABLE 16.11 SUMMARY OF U/G STOPE MINING SEQUENCE Level Stope Tonnes Start Date Finish Date 195 - 215 Stope 21 9,253 July 28, 2016 Aug 2, 2016 195 - 215 Stope 3 7,112 Aug 8, 2016 Aug 21, 2016 195 - 215 Stope 20 9,424 Aug 25, 2016 Aug 30, 2016 195 - 215 Stope 19 9,210 Sept 2, 2016 Sept 7, 2016

161 - 161 Stope 15 14,735 Nov 4, 2016 Nov 12, 2016

Total 198,917

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16.2.3 Ventilation

Ventilation flows are designed to sweep the mine from the centrally located Fresh Air Raise (FAR) to the extremities of the mine. The ventilation system is designed to operate as follows:

 The ventilation system will provide fresh air via the FAR to the bottom of the mine, the air then up-casts the main decline and returns to surface through auxiliary ventilation systems. Auxiliary fans and ducting directs fresh air to active working areas. The air then flows back to the main decline and returns to surface up the decline.  It is planned to use blast-hole open stoping that generally retreats from the furthest extremities on the levels to the FAR and decline complex. Adequate ducting and secondary ventilation fans will direct fresh air from the decline to the working areas.  The underground ventilation system is required to provide airflow volumes and an airflow distribution that will provide an acceptable atmosphere within the working environment for all underground workers. For this project, the ventilation system is designed to control diesel exhaust emissions concentrations in the workplace and by default, dust and blasting fumes.

The following specific design criteria and assumptions were adopted for the project:

 The system will be designed to provide at least 100 CFM/BHP (brake horsepower) of diesel equipment operating underground. Where it is planned to use combinations of equipment simultaneously in an area, the ventilation volumes will be designed to support total sum of operating horsepower for major development or production equipment.  Not all diesel equipment will operate simultaneously underground.  FAR (Intake complete with fan installation) and ramp is completed to the 135 mL.  Fresh (heated in winter) air will downcast in the FAR to the bottom of the existing decline. Ventilation air then upcasts through the decline to surface.  During installation, obstructions, restrictions and diameter reductions in the auxiliary ducting will be minimized.  Diesel equipment will be maintained to control and minimize exhaust emissions to Canadian and Territorial mining industry standards.

An equipment distribution list was developed, listing the diesel equipment, the location in the mine where the equipment is to operate, and the utilization time on a monthly basis. From the equipment distribution list, minimum required air volumes were determined for diesel exhaust emission control. The basic assumption is that Table 16.12 provides a summary of the diesel equipment, brake horsepower and utilization factors planned for the underground workings. The required minimum fresh air volume is based on providing 100 cfm/bhp for the total operating diesel horsepower (≈ 228 kCFM), plus an allowance for inactive mining areas and an estimated leakage factor (20%). The minimum volume required is therefore estimated to be approximately 324 kCFM (≈153 m3/s).

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TABLE 16.12 VENTILATION REQUIREMENTS FOR UNDERGROUND DIESEL EQUIPMENT HP for Engine Installed Overall Unit Qty Ventilation Ventilation Power Power Utilization (HP) (m3/ (HP) (HP) (%) (HP) (CFM) s) Scooptram 6.1 m3 LHD 2.5 325 813 87.5% 711 71,094 34 U/G Truck 50t Haul 3.5 575 2,013 60% 1,208 120,750 57 Truck Longhole ITH Drill 2.0 173 346 10% 35 3,460 2 Development Jumbo - 2 2.0 149 298 10% 30 2,980 1 Boom Getman Anfo Loader 1.0 173 173 20% 35 3,460 2 Getman Scissor Lift 2.0 173 346 25% 87 8,650 4 Getman Boom Truck 1.0 173 173 25% 43 4,325 2 Toromont Cat Grader 1.0 135 135 25% 34 3,375 2 M135H Machanic's Vehicle 1.0 128 128 25% 32 3,200 2 Electrician's Vehicle 1.0 128 128 25% 32 3,200 2 Staff Toyota 1.0 128 128 25% 32 3,200 2

Subtotal 18.0 2,260 4,680 2,277 227,694 107 Allow 42,500 cfm (20 m3/s) to ventilate approximately four (4) inactive mining areas 42,500 20 Minimum air required (Allow for 20% leakage and short-circuiting) 324,233 153 Say 324,300 153

Ventilation air is taken from the FAR and supplied with auxiliary fans and temporary ducting to numerous development ends in the mine.

16.2.4 Hydrology

Based on the underground exploration development program that was completed in 2006 by Procon, P&E expects the ground water inflow to be negligible at an estimated 50 m3 / day (9.2 USGPM).

16.2.5 Manpower

16.2.5.1 Underground Operations and Maintenance Labour

All of NICO‟s underground labour force will be contractor‟s personnel. The number of workers required is based on the total amount of underground development, construction and production work required, and labour productivities. A summary of the total number of contractor personnel man shifts required for the Project is presented in Table 16.13.

In summary there is an estimated 14,600 underground operations and maintenance man shifts worked, life-of-mine. The daily maximum camp requirement is for 74 workers.

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TABLE 16.13 PROJECTED NUMBER OF UNDERGROUND OPERATIONS AND MAINTENANCE PERSONNEL Man May Jun Jul Aug Sept Oct Nov Dec Jan Feb Description days 2016 2016 2016 2016 2016 2016 2016 2016 2017 2017 Super 405 1.80 1.50 1.50 1.50 1.50 1.50 1.50 1.50 1.20 1.00 Shiftboss 792 3.60 2.90 2.90 2.90 2.90 2.90 2.90 2.90 1.90 1.00 Rotation 710 1.60 2.90 2.90 2.90 2.90 2.90 2.90 2.90 1.40 Super Safty Super 355 0.80 1.50 1.50 1.50 1.50 1.50 1.50 1.50 0.70 Blasting 355 0.80 1.50 1.50 1.50 1.50 1.50 1.50 1.50 0.70 Super Proj 355 0.80 1.50 1.50 1.50 1.50 1.50 1.50 1.50 0.70 Engineer Surveyor 720 1.90 2.90 2.90 2.90 2.90 2.90 2.90 2.90 1.40 Lead 395 1.40 1.50 1.50 1.50 1.50 1.50 1.50 1.50 1.20 1.00 Mechanic Mechanic 2,942 8.70 11.60 11.60 11.60 11.60 11.60 11.60 11.60 6.30 2.00 Welder 355 0.80 1.50 1.50 1.50 1.50 1.50 1.50 1.50 0.70 Electrician 792 3.60 2.90 2.90 2.90 2.90 2.90 2.90 2.90 1.90 1.00 Surf Eqpt 385 1.10 1.50 1.50 1.50 1.50 1.50 1.50 1.50 1.20 1.00 Oper Clerk 355 0.80 1.50 1.50 1.50 1.50 1.50 1.50 1.50 0.70 Miner 4,004 13.20 13.80 21.80 28.70 23.80 16.40 8.30 3.80 1.80 L/Hole 303 0.20 1.00 1.60 2.00 2.30 1.60 1.00 0.30 Driller L/Hole 196 0.10 0.40 0.70 1.20 1.10 1.40 1.10 0.30 Loader Scoop Oper 787 0.40 1.20 2.70 4.60 4.50 5.90 5.00 1.50 Truck Oper 636 0.60 1.90 3.30 4.80 4.30 3.30 1.70 0.80 0.20

Total 41.50 51.60 62.90 73.70 71.10 62.80 54.10 46.90 24.10 7.00 Man days 14,581 871.5 1,570.6 1,914.5 2,243.2 2,164.1 1,911.5 1,646.7 1,427.5 733.5 98.0 Workdays 278.5 21.0 30.4 30.4 30.4 30.4 30.4 30.4 30.4 30.4 14.0

A summary of Fortune‟s underground personnel is presented in Table 16.14.

TABLE 16.14 PROJECTED NUMBER OF UNDERGROUND FORTUNE MINERALS PERSONNEL Description Number Required Clerk 2.0 Mine Engineer 1.0 Drafting Technician 1.0 Mine Geologist 1.0 Geological Technician S1 1.0 Geological Technician S2 1.0 Total Fortune Minerals 7.0

16.2.6 Equipment

A summary of the contractor equipment is presented in Table 16.15.

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TABLE 16.15 SUMMARY OF CONTRACTOR EQUIPMENT Description Quantities Container 8' x 40' 4 Container Shop 1 Office Trlr 12 x 40' 1 Resscue Sta. 8' x 20' 2 Mine Rescue Eqp‟t 1 lot Fire Fighting Eqp‟t 1 lot Bit Grinder 1 Office Eqp‟t 1 lot First Aid Eqp‟t 1 lot Shop Tools & Eqp‟t 1lot Steam Cleaner 1 Survey Eqp‟t 1 lot Vent/Gas Testing Equp‟t 1 lot Oxy-Acetylene Eqp‟t 1 Crewcab 4x4 3 Toyota Jeep 2 Serv. Tractor 2 U/G Grader 1 Air Heter-5.5M BTU 1 Pump-13 HP Flygt 5 Pump-58 HP Flygt 4 Pump-Water Supply 1 Fan-75 HP.Elec. 5 Fan-100 HP.Elec. 3 Scoop 3.5 C.Y. 1 Scoop 8.0 C.Y. 3 Truck-30T Volvo 6x6 1 U/G Trucks - Toro 40D 4 Jumbo – Tamrock Hyd L/H 2 Jumbo 2B Elec/Hyd 2 Hammers-Hyd. 6 Jackleg 8 Stoper 8 Dry Mix S/Crete Pump 1 600kva Sub/Trans. 3 Survey Gear 5 ANFO Loader-Hand 2 ANFO Loader-120 -POT 2 Mobile ANFO Ldr 1 Scissor Lift 2

A summary of Fortune support equipment for the underground operation is presented in Table 16.16.

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TABLE 16.16 SUMMARY OF FORTUNE MINERAL’S EQUIPMENT FOR THE UNDERGROUND Description Quantities Cap Lamps 7 Safety Gear 7 Mine ERT - In G&A 1 Recondition & Install Surface Ventilation Fans & Heater 1 Surface Explosives and Detonators Magazines Lot Compressors 3 Standby Diesel Generator – 0.8MW 1 Diesel Fuel Storage at Portal 1 Welder 1 Dry, Shift Boss & Mine Management Area Lot Mine, Engineering & Geology Offices Equipping 1 Ore Sample Assaying Equipment Lot NICO Management Underground Vehicles 1 Ambulance 1

16.2.7 Electrical Power

The required electrical power for underground mining will be generated using an existing diesel generating set, to be located near the portal.

The portal substation is also located close to the underground ramp entrance. The power from portal substation to the underground substations will be fed via two redundant mine power cables. The power to various underground power centers will be distributed at 4.16kV via mine power cables. A secondary electrical substation will be located at the 105 level to power the main dewatering pumps and portable substation required in this general area.

Three mine underground portable substations rated at 600 kVA will be used to feed the area ventilation and working machinery loads.

It is estimated that the power consumption for the underground will be 1.3 MW normal running load with peak demand load of 2.3 MW, with various splits as per the following Table 16.17.

TABLE 16.17 ESTIMATED UNDERGROUND ELECTRICAL POWER REQUIREMENTS kW Description. of Load Normal Running Load Peak Demand Load Ventilation 623 804 Mine Dewatering 139 315 Mine Equipment 551 1,194

Total 1,313 2,313

If for some reason power from the main substation is not available either due to the loss of transmission line, the generator maintenance or generator forced outage, the power to the

P&E Mining Consultants Inc., Report No. 247 Page 127 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. underground essential loads will be supplied from the standby diesel generator located at the portal substation. The total underground requirement on standby is an estimated 0.8 MW.

16.2.8 Stockpile

In general development and production ore will be hauled from underground directly to the primary mill crusher.

16.3 GEOTECHNICAL

Fortune retained Golder to provide geotechnical support for the open pit and underground designs, NICO Project, a cobalt-gold-bismuth project in the NWT. Golder has provided ongoing support since 2003.

Through 2010 and to November 2011 Fortune and P & E have been carrying out detailed design work to optimize location, number and size of the proposed underground workings relative to the ultimate pit shell and its intermediate phases, which had been finalized previously, in 2009. The updates in this document pertain specifically to the geotechnical analyses to support this mine design work. The most recent underground mine plan iteration envisions an approximate mine life of 1 year, and considers a deeper open pit.

16.3.1 Mine Plan Summary

Mining of the NICO ore zone will commence by extracting the richer ore zones using underground mining techniques during the third and fourth years of operations, followed by advancing an open pit to year 22. Starting about mid-way through the mine life the pit will intersect the existing underground workings.

The open pit is planned to be developed in three phases, identified as Phases 1 to 3, with Phase 1 sub-divided into Phases 1A and 1B, as illustrated on Figure 16.8. As observed, Phase 2 will contain both Phases 1A and 1B pit shells and will extend further to the southeast, whereas Phase 3 is a push-back from Phase 2 towards the northwest and progressing deeper forming the ultimate pit depth.

Review of the slope geometries indicates that the maximum inter-ramp slope angles conform to the recommendations provided in the Golder 2004 Technical Memorandum on slope design. Review of the pit shell Phases 1 to 3 indicates maximum inter-ramp slope angles of 50˚.

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Figure 16.8 Schematic Representation of the Phase Pit Shells and the 2011 Planned Underground Stopes

Figure 16.8 also includes the underground stopes that are planned to be mined using upper blastholes and without backfill. As the open pit advances towards these open stopes they will be backfilled with broken rock from the pit floor via drop raises or as they are exposed.

The engineering geology investigations include: i) outcrop mapping by Fortune in 1998-2000, ii) geotechnical logging and rock strength testing of exploration core in 1998, iii) geotechnical logging of exploration core in 2003, iv) geotechnical logging and core orientation of holes oriented for slope design in 2003, and v) additional laboratory rock strength testing from definition drilling core obtained in 2010.

The pit slope designs indicated competent rock masses for which overall slope stability would not be a control on slope design for the anticipated slope heights. Consequently, pit slope designs were based on kinematic assessments, with assumed achievable bench geometries controlling the inter-ramp angle.

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16.3.2 Geology

The NICO deposit is located in the southern part of the Great Bear magmatic zone, comprising Paleoproterozoic volcanic and plutonic rocks exposed from the Great Slave Lake in the south to Great Bear Lake in the north (Golder, 2004).

The oldest exposed rocks in the area are meta-sedimentary rocks. The NICO deposit is hosted in Treasure Island Group sedimentary rocks that comprise a 3 to 5 km wide succession of siltstone, impure dolomite, subarkosic wacke and arenite (Goad et al., 1998).

NICO mine will be developed within the Black Rock Ironstone formation. This is a formation of meta-sedimentary rock with mineralization consisting of several closely stacked, stratabound, irregular, mineralized bearing lenses in altered siltstone and subarkosic wacke units. The formation dips at approximately 50° towards 030° (North-North East). The major mineralized zones to be mined in this formation consist of two sub-parallel lenses, approximately 40 m apart.

16.3.3 Discontinuous Permafrost Conditions

Regional

The NICO Project is located within the zone of discontinuous permafrost (Brown, et al, 2001). Within the discontinuous zone, permafrost occurs together with unfrozen ground. The permafrost zones are quite patchy depending on certain types of terrain, mainly peat lands. The annual thaw may reach a depth of 3 m and lenses of permafrost may form or disappear, depending on local short-term influences.

Local

EBA (EBA, 2005b) described the permafrost conditions specific to the area located south of the current proposed plant site location and west of the current proposed polishing pond location as follows (area referred as “South Basin” at that time):

 “The project site is located within a zone of discontinuous permafrost. Ice-rich organic soils were encountered overlying ice-rich perennially frozen glaciolacustrine and lacustrine clays and silts. Preliminary ground temperatures measured after thermistor installation suggests that the valley slopes of the proposed tailings dam abutments may be unfrozen.  Valley bottoms and depressions are underlain by frozen deposits of glacial and post-glacial origin, and also contain meta-sedimentary rocks. The predominant vegetation along the valley floor consists of a thick Sphagnum moss cover with relatively dense stands of black spruce forest, with some stunted trees leaning in random directions as a result the cyclical frost heave that occurs in fine-grained perennially frozen soils.  Unfrozen igneous rocks, predominantly volcanic in origin, comprise the surrounding hills and ridges, which overlie and intrude the meta-sedimentary rocks. The valley slopes, bedrock hills and ridges are sparsely vegetated with lodgepole pine, white spruce, white birch, and aspen. The hills and ridges exhibit well developed cryoplanation terraces (i.e. step-like features formed on exposed bedrock slopes due to intense frost wedging from snow banks) on their sides.

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Unconsolidated rock debris, boulders, and soil are found on the terraces and lower portions of the bedrock slopes.”

16.3.4 In Situ Stresses

There are no in-situ stress measurements for the NICO project. The in situ stress state was assumed to be represented by the following relationships (common regional properties within the Canadian Shield):

Azimuth: 025° Azimuth: 115° Vertical

These gravitational in situ stresses were based off a datum (surface) elevation of 305 m. The major in-situ stress is assumed to be horizontal and acting along azimuth 025°, approximately normal to the strike of the orebody. The intermediate horizontal in-situ stress acts along azimuth 115°. The minor in-situ stress is equal to the vertical stress and increases with depth.

16.3.5 Geotechnical Drillhole Locations

The following is a brief summary of subsurface investigations at the NICO site, including selected drillholes for features other than the open pit and underground investigations, with notes on thermistor and piezometer installations, for completeness. The locations of these boreholes are shown on Figure 16.9:

 Golder 2003: Pit and underground geotechnical inclined diamond drilling investigations with stand pipe and thermistor completions.  EBA 2004, Golder 2006 and Golder 2010: Shallow geotechnical investigations to determine overburden and top of bedrock conditions. Standpipe, vibrating wire piezometer and thermistor completions as noted.  Fortune 2010: Condemnation drill holes, geotechnically logged, Co-Disposal facility area.  Fortune 2010: Definition drillhole, geotechnically sampled for laboratory strength testing. This hole is sub-vertical and located adjacent to 03-281.

Additional 2010 drill holes, completed by EBA in the area labeled Plant Site, for the investigation of foundation conditions are not shown.

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Figure 16.9 Open Pit and Mine Waste Geotechnical Investigations Borehole Location Plan

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16.3.6 Engineering Geology

The NICO site is located below the tree line in hilly, woody and muskeg terrain. Outcrop exposures on hills are common. Discontinuous permafrost has been found in low-lying and wetland areas.

The mineralized rock (black rock altered ironstone), the footwall rock (metamorphosed siltstones), hanging wall/cap rock (potassium-feldspar altered rhyolite) and feldspar dykes are all strong to very strong rocks with good to excellent RQD.

Both the black rock and the meta-siltstone have a low hydraulic conductivity, ranging from less than 10-8 m/s near top of rock to less than 10-9 m/s with depth. The dykes and volcanic cap rocks have higher hydraulic conductivity, attributed to the greater degree of fracturing observed in these rocks.

Static groundwater levels within the proposed open pit and underground mine workings footprint follow topography. On the hills it is within 20 m of ground surface. At lower elevations it discharges, such as in the “Bowl Zone” valley on the northwest end of the proposed site. This discharge feeds the Grid Stream and the Grid Ponds, and is naturally elevated in some metals, such as arsenic, due to the geochemistry of the bedrock.

The main rock type to be exposed in ultimate pit slopes (hanging wall and footwall) and in stopes will be the black rock ironstone. Review of surface mapping and oriented core data suggests that there is a consistent structural fabric within this unit, characterized by moderate to widely spaced joints dipping parallel to the 40° - 45° northeast dipping siltstone / black rock ironstone contact.

The meta-sedimentary rocks and the black rock ironstone unit are banded, and the peak orientation of discontinuities associated with the banding (foliation joints) is slightly steeper, dip 50°, than the contact dip between the footwall siltstone and the black rock ironstone. The spacing with which these banding / strata sub-parallel joints occur is moderate to wide (e.g. 0.5 m to 1.0m) in the oriented core; at drillhole 03-282, about 50 features with this orientation were intercepted over a borehole length of over about 210 metres. Similarly, at drillhole 03-283, about 190 features were intercepted over about 270 metres. The structural fabric within the siltstone and greywacke meta-sedimentary rocks is consistent with that in the black rock ironstone.

The structural fabric within the volcanic (altered rhyolite) cap rock, which will have limited exposures on upper benches of the ultimate pit slopes (no more than 20 m vertically) is distinct from the underlying meta-sedimentary rocks. The volcanic rocks are blocky, with sub-vertical and sub-horizontal joint sets. They lack the footwall parallel bedding / foliation set prevalent in the meta-sedimentary rocks.

The four main geotechnical rock types at the Fortune site are described in Golder, 2011 and Table 16.18.

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TABLE 16.18 GEOTECHNICAL UNITS Typical Rock Type (From Oldest to Typical Rock Significant Exposures in Proposed Number of Youngest) Strength Mine Workings Joint Sets Non-mineralized. R5 Host rock for bulk sample portal and Footwall Siltstone 2 to 3 Very strong rock decline (underground ramp). Light grey, fresh, fine grained Footwall on a limited number of Bedded smooth to Average UCS = underground panels. Dip 40-50 ° to the NE rough, planar 97.5 ± 39 MPa Limited exposure on proposed Ultimate Slopes Mineralized (including sulphides and magnetite). Black Rock Ironstone R3-R4 (meta-sedimentary rock) Medium strong to 2 to 3 Significantly heavier than other rock Dip 40-50 ° to the NE strong rock types. Arsenopyrite banding Banded smooth to prevalent. Note that this unit is also referred Average UCS1 = rough, planar Footwall and hanging wall for most to as Schist in previous reports. 109 ± 43.8 MPa ultimate pit slopes. Footwall, hanging wall, roof and backs of most underground panels. Non-mineralized. Greywacke R5-R6 2 to 3 Overlies the black rock altered (meta-sedimentary rock) Very strong to ironstone (schist) Dip 40-50 ° to the NE Extremely strong smooth and Also occurs as un-mineralized zones Bedded rock planar within the ironstone Mineralized (but not economic). Volcanics Overlies the greywacke as a R5 Rhyolite that is highly potassium discontinuous cap. Very strong rock 3-4 feldspar altered, which colours the Will be exposed on some open pit smooth to rock red. slopes, mainly upper benches of the Average UCS = rough, planar These rocks form the hanging hanging wall 147 ± 73 MPa wall Observed to be highly fractured at surface. R5-R6 Sub-vertical intrusions that cross-cut Extremely strong 3-4 the host rocks. Dykes rock smooth to Potential exposures both on proposed

rough, planar ultimate pits and underground Average UCS = openings 155 ± 46 MPa *(after Golder, 2005a) (1) The average UCS strength for the Black Rock Ironstone used in analyses prior to 2011 was 75 MPa, approximately 30 MPa lower, but based on very limited sample set. In late 2010, additional specimens were selected from 2010 definition drillholes, namely 10-291 for strength testing, The revised typical rock strength has been used in subsequent 2011 analyses.

Note that while the black rock ironstone is termed a schist in exploration and geological logs, in geotechnical terms the unit does not have a strong, closely spaced foliation that behaves as a preferred plane of weakness. Rather the unit is massive, with distinctive bands of arsenopyrite mineralization within the host black rock meta-sedimentary rock. Field observations and review of the failures from uniaxial strength testing indicate that the banding alignment (foliation) is not a preferred plane of weakness. The data on Table 16.8 is updated based on additional strength testing of the Black Rock Ironstone carried out in 2010 on specimens from in-pit definition drill hole 10-291, located adjacent to geotechnical hole 03-281.

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16.3.7 Open Pit Design

Pit slope design recommendations were provided for the Micon BFS in the Golder (2004) technical memorandum. The next paragraphs summarize the structural fabric assessment and rock slope stability analysis results and recommended bench configurations that were provided to the project.

Structural Data

Surface fault traces were mapped by Fortune geologists from the exposed outcrops and their locations were presented on Figures 1 and 2 of Golder‟s November 2004 Technical Memorandum. Faults are typically healed to slightly broken at surface. Fault zones interpreted at depth are based on displaced geological or mineralization features. While some zones of broken core / low RQD are reported, these do not necessarily correspond to the locations of inferred faults in the borehole records. Most faults are considered healed based on this information.

Structural fabric data were obtained from nine oriented core holes (six drillholes logged by Golder-trained Fortune geologists, three logged by Golder in 2003) and from surface mapping carried out by Fortune in 1996-1997 and 1997-1998. The results of the detailed review of lower hemisphere, equal area projections of discontinuity populations obtained from both oriented core and surface mapping are summarized below.

Meta-Sedimentary Rocks

The three (3) oriented core holes and surface mapping data indicate similar and consistent discontinuity population sets within the siltstone, the black rock altered ironstone and greywacke meta-sedimentary rocks, as shown in Table 16.19.

Peak orientations that parallel the footwall strike or the faults mapped by Fortune (Golder 2004 Technical Memorandum) and are considered most likely to be potentially continuous at the bench to multi-bench scale are labeled F1 to F3. The remaining peaks, considered more likely to be discontinuous at the bench to multi-bench scale are labeled J1 to J4. Potentially continuous discontinuities are of most concern with respect to control of multi-bench scale slope stability.

Orientation Bias

All core orientation data were obtained from geotechnically logged exploration or in-fill definition drillholes with southwest azimuths and, as such, has potentially not intercepted possible discontinuity sets dipping to the southwest (parallel to the drillholes). Those discontinuities, if they exist and were continuous and persistent, would control the slope configurations on the hanging wall of the open pit or could combine to form underground wedges. This possible footwall perpendicular “cross-joint” set is labeled J4 for purpose of these discussions.

The surface mapping data collected by Fortune in 1996 and 1997 agrees with the oriented core data, in that it captures the same set distributions as the oriented core, but also suggests, by the limited number of poles intercepted, that the J4 set is poorly developed and discontinuous. Additional surface mapping collected by Fortune in 1998 generated a similar distribution of peak orientations.

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Figure 16.10 NICO Deposit Structural Fabric based on Surface Mapping and Oriented Core

TABLE 16.19 PEAK ORIENTATIONS OF DISCONTINUITY POPULATIONS IN THE META-SEDIMENTARY ROCKS* Set Dip / Dip Direction % / Joint Set Description ID (°) Number F1 50 / 026 parallel to sub-parallel to footwall 5.8 / 16 F2 78 / 023 sub-vertical sets striking parallel to the footwall, 2.2 / 6 Sub-vertical set striking perpendicular to the F3 60 / 081 2.2 / 6 footwall. set dipping into footwall (based on surface J4 55 / 210 1.5 / 2 mapping only ) J3 15 / 207 Prevalent sub-horizontal set 3.4 / 10 J2 79 / 204 Minor joint set 2.2 / 6 J1 88 / 156 Minor joint set 2.2 / 6 *(after Golder, 2005) Notes: (1) Sets in bold, F1 to F3, are considered most likely to be potentially continuous. (2) Because of the uniformity of oriented core hole azimuths and consistency of data in core holes and surface mapping, the peak orientations above were selected from a single borehole, 03-282, which was considered representative of the meta-sedimentary rock structural fabric for purposes of this assessment. Set J4, based on surface mapping, was added for completeness. (3) %/ Number generally refers to the pole concentration at the peak orientation of the set based on a 1% counting circle on a lower hemisphere equal area stereonet. For Set J4, this is based on a surface mapping database of 87 poles from 1997. For the remaining sets, the percent and number are based on O3-282 data, which had 284 poles.

16.3.7.1 Volcanics

The altered rhyolite cap rock and the intrusive dykes have a different structural fabric than the metasediments, as shown in figure 16.11. Jointing in these rocks is characterized by steep sub- vertical sets and orthogonal, sub-horizontal sets, and the absence of the footwall parallel discontinuity set (i.e., F1 set) observed in the metasediments.

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Figure 16.11 Peak Orientations of Discontinuity Populations in the Volcanic Cap Rock (Potassium Feldspar Altered Rhyolite)

TABLE 16.20 PEAK ORIENTATIONS OF DISCONTINUITY POPULATIONS IN THE VOLCANIC CAP ROCK (POTASSIUM FELDSPAR ALTERED RHYOLITE)* Set ID Dip / Dip Direction (°) Set Description % / Number R1 16 / 105 Peak orientations of 8.0 / 4 R2 37 / 083 lat Sets (dips < 38 degrees) 8.0 / 4 Steep set with strike sub-parallel to R3 75 / 220 3.4 / 2 Footwall R4 76 / 349 5.8 / 3 Sets R4 and R5 are orthogonal. R5 84 / 071 4.6 / 3 *(after Golder, 2005) Notes: (1) Sets in bold, R1 to R3, are considered most likely to be potentially continuous based on prevalence and / or fault and contact trends. (2) Because of the uniformity of oriented core hole azimuths and consistency of data in coreholes and surface mapping, the peak orientations above were selected from a single borehole, 03-282, which was considered representative of the cap rock structural fabric for purposes of this assessment (3) %/ Number refers to the pole concentration at the peak orientation of the set based on a 1% counting circle on a lower hemisphere equal area stereonet.

16.3.8 Slope Design Definitions

A pit slope has three major components: bench configuration, inter-ramp slope and overall slope, as illustrated on Figure 16.12. The bench configuration is defined by vertical bench separation (or bench height), catch berm width (or berm width) and bench face angle (or batter). The inter- ramp slope is formed by a series of uninterrupted benches and the overall slope is formed by a series of inter-ramp slopes separated by haul roads.

The inter-ramp angle (“IRA”) corresponds to the angle subtended by a line joining the toes of the benches on the wall and the horizontal. The overall slope angle corresponds to the angle formed by the line joining the toe of the lowest bench with the pit crest and the horizontal. Therefore, the P&E Mining Consultants Inc., Report No. 247 Page 137 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. incorporation of ramps onto a wall will result in a slope that has a shallower overall slope angle than the inter-ramp angle.

Figure 16.12 Schematic Representation of Bench Face Angle (“BFA”) and Inter-Ramp Angle (“IRA”) and Overall Slope Angle (“OSA”)

16.3.9 Kinematic Assessment

The kinematic assessments indicate that bench configurations on some wall orientations will be controlled by potential planes and wedges involving structures assumed to be potentially continuous because they parallel the regional interpreted fault trends. For purposes of this assessment, a shear strength of phi=30° and cohesion of c=0 was assigned to all discontinuities. This strength is considered conservative given the variability of joint surface character. Joint surfaces showed minimum surface alteration and can be described as fresh to slightly stained.

Results of the kinematic assessment are presented in the Golder 2004 Technical Memorandum and repeated in the Golder 2011 report. Figure 16.13 exemplifies the kinematic analysis for the footwall slopes considering the rock mass fabric for the meta-sedimentary rocks.

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Figure 16.13 Example of kinematic analyses for the FW wall (slope dip direction of 30°), considering the rock mass fabric for the meta-sedimentary rocks.

16.3.10 Slope Design Recommendations

The following are the slope design recommendations for ultimate pit slopes in rock, presented by wall dip direction for the three design sectors.

Bench geometries were developed assuming adequate dewatering of the rock slopes will have been achieved due to exposure and blasting. Horizontal drains (e.g., 50 mm to 76 mm diameter, open holes, drilled at approximately 10 upwards and to a about 50 m depth into the walls) may be required in localized areas if persistent seepage is noted during pit operations or if adequate pore water depressurization is not achieved. If necessary, these holes can be lined with 25 mm perforated/slotted PVC pipe to maintain open drillholes and free draining conditions.

Recognizing the extreme cold temperature at the NICO site, water flowing from these drainholes should be manifolded into pipes to the sumps, in order to prevent face freeze-up in the winter months. The drainholes will also help reduce icing up of the slope face in the winter, which could lead to ice-falls.

Implementation of proper controlled blasting techniques on all benches will improve the probability of success in achieving the recommended bench scale and inter-ramp design parameters.

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TABLE 16.21 NICO OPEN PIT SLOPE DESIGN RECOMMENDATIONS* Bench Maximum Minimum Maximum Slope Dip Face (or Rock Type Vertical Bench Berm Width Inter-ramp Direction Batter) Separation (m) (m) Angle (o) Angle (o) Meta- sedimentary 15 75 8.5 50°(1)(2) 020° to 030° Rock Footwall Volcanic Cap 15 75 8.5 50° Rock3 Meta- 15 75 8.0 51°(4)(5) to sedimentary 200° to 210° 20 75 9.0 54° (4)(5) Rock Hangingwall Volcanic Cap 15 75 8.0 51° Rock(3) Endwalls All rock types 15 m 75 8.5 50° *(after Golder, 2004) Notes: (1) Actual Inter-ramp and overall slopes on the footwall slope will most often be less than 50 degrees, controlled by the local dip of the stratabound mineralization zones (see cross-sections) and placement of ramps. (2) Bench face angle controlled by set potential for planar failures involving set F2, dip 78°, inter-ramp slope angle controlled by set F1 (foliation), mean dip 50°. (3) Some slopes will expose significant amounts of volcanic rocks on upper benches. While the kinematics indicate that the structural fabric in the volcanic rocks is more favourable and that steeper slopes could be achieved, surface exposures are blocky and broken, and ravelling can be expected on excavated slopes. For this reason a steeper design for slopes in volcanics is not presented. Initial operating experience with volcanic slopes will determine whether ravelling will require modified blasting practices or shallower inter- ramp angle. (4) Bench face angle on the hanging wall will controlled by potential for planar failures involving set J2 (dip 79°). Inter-ramp angle on the hanging wall controlled by potential for planar instability involving set J4 (dip 55°). Should the F2-J2 wedge be prevalent (plunge 51°) the inter-ramp angle will require flattening from 54° to 51°. (5) The hanging wall slope design is considered aggressive. It is recommended that in order to optimize the hanging wall design, ramps be placed on the footwall, which will be mined at the flatter angles conforming to the dip of the stratabound mineralization zones. (6) Potential for toppling failure, particularly on the hanging wall, is not considered a control on bench design given the moderate to wide spacing of joints. Localized toppling instabilities may still occur. Should toppling failure be problematic, a mid-bench catch-berm may be required. (7) Potential wedge F1-F3, plunge 50° will control slope design on southeast dipping end walls. Northwest dipping end walls have been assigned the same recommended configuration for consistency. (8) Where the open pit slope is greater than 90m to 120m high without being crossed by a ramp or wider berm, it is recommended that extra wide (allow 12m to 15m) geotechnical bench be placed on the slope as a conservative measure. This bench is intended to provide additional catchment against potential rock fall hazard. The width has not been designed, the 12m to 15m is recommended based on experience.

16.3.11 Pit and Underground Seepage Estimates

Access ramp (decline) dewatering and underground mining development will commence prior to or in conjunction with open pit operations, which will significantly dewater the open pit zone. Groundwater seepage into the open pit is expected to be very low, as a result. Combined pit and underground seepage was estimated to be on the order of 100 m3/day throughout the life of mine. Actual seepage will be verified as excavation and initial mining occurs, starting with observations made during the dewatering of the bulk sample decline. The estimates are based on P&E Mining Consultants Inc., Report No. 247 Page 140 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. in-situ measurements of hydraulic conductivity, observed seepage into the decline and closed form and numerical modeling. These investigations are summarized in Golder (2010c).

16.3.12 Underground Mine

The stope design and the interaction of the open pit and underground stopes were discussed in the Golder (2005) and Golder (2010a and b) technical memoranda and through exchange of emails with the P&E project team.

Through these interactions, the planning for the underground mine has changed. Recognizing that the major mineralized zones consisted of two sub-parallel lenses, which were approximately 40 m apart, the initial mine planning considered the excavation of a series of sub-parallel transverse stopes that would be located near the hanging wall (labeled as the MZ stopes) and in the footwall (labeled as LZ stopes). The stopes would not be backfilled during the underground mining and would be separated by sill and rib pillars, which would be recovered later on during open pit mining.

The plan continues to backfill stopes with broken rock from the pit floor via drop raises or as they become exposed from the progression of open pit mining.

Subsequently in 2011, the number of stopes has been substantially reduced, now accounting for only 376,000 tonnes. The current plan is to mine selected longitudinal and transverse stopes and minimize their potential impacts on future open pit mining.

Since these stopes will not be backfilled during underground mining, they were designed to be mined with dimensions that would ensure stability while they remained open.

The following sub-sections present the steps taken for stope design for the NICO project.

16.3.12.1 Semi-Empirical Open Stope Stability Analyses

An assessment of potential stope dimensions has been carried out using the Mathews/Potvin stability method (Mathews et al., 1981; Potvin, 1988) for open stope stability. Input requirements for the method include assessment of rock mass quality for the various walls of the stope, assessment of the stress concentration within the excavated stope walls and assessment of the orientation of typical joint sets within the rock mass.

The following assumptions have been applied to this assessment (Golder 2005 Technical Memorandum):

 The ore body has a strike length of approximately 1250 m (based on current ore body interpretation) and dips at approximately 50°. A median ore body width of 25 m in the transverse direction was considered as a number of stopes will be developed in this direction.  Consideration was given to physical limitations such as equipment size and drilling accuracy (drillhole deviance). As a result the excavation dimensions considered initially a strike span of 15 m to provide enough room for equipment movement and a sublevel height of 25 m to minimize drillhole deviance.

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 Rock mass qualities are based on the summary results from the 1998 and 2003 core logging and strength testing exercises. The metasediments identified at the deposit will comprise the back, sidewalls, and hanging wall of all stopes.  The quality of rock masses can be classified as fair (3  Q‟  10) for the -1 std case and good (10  Q‟  40) for the mean and +1 std case. As the major controlling factor within the ratings for the rock mass classification is the strength of the rock unit, the metasediments can be generally simplified as strong (50MPa < UCS < 100 MPa) with an assumed uniaxial compressive strength of 75 MPa, to take into account some of the variation in the rock strength (note that the average UCS = 109 ± 43.8 MPa).  Rock mass for the Black Rock Ironstone was assessed as having on average good quality (10 ≤ Q‟ ≤ 40) with an average Q‟ of 14.9 (Golder, 2005). Considering a standard deviation of ±1, the Q‟ values equate to an approximate GSI range of 64 to 72, with an average of 68.  The assumed in-situ stress conditions are provided in Section 16.3.4. Three- dimensional numerical modeling was carried out using the MAP3D© program to estimate the induced stresses at the mid points of the back, side walls, and the hanging wall. For the stability assessment the stopes were grouped as being located at depths above and below a depth of 150 m, which was used for general reference.  Critical discontinuity set orientations were extracted from the summary stereonets produced from the oriented core and surface mapping program as summarized in Table 16.19 and Table 16.20.

In the Mathews/Potvin stability graph, rock quality (reflected by the modified stability number N‟) is related to stope geometry (reflected by the hydraulic radius, HR). Several curves are included in this graph to divide the chart into different zones, referred to as stable, transitional and caving zones for unsupported and supported stope walls, as depicted on Figure 16.14.

The stope hydraulic radius (HR) corresponds to the area divided by the perimeter of the exposed stope surface analyzed, as follows:

Where w and h refer to the width and height of the stope wall.

The N‟ stability number is assessed as follows:

N‟ = Q‟ x A x B x C

Where Q‟ is the rock mass quality, and A, B, and C are adjustment factors based on the stress conditions, orientation of structures, and most likely mode of failure respectively.

The recommended HR radius for stope design was based on the intersection of the stability number (N‟) with the stable zone / unsupported transition curve that corresponds to the maximum sized unsupported opening considered “stable” (i.e. minimum dilution and potentially no backfilling requirements). This approach to stope design was recommended for design, recognizing that stope development would be undertaken using upper blastholes and without the possibility of installing ground support on the back or sidewalls. P&E Mining Consultants Inc., Report No. 247 Page 142 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

The midpoint within the unsupported transition zone of the Mathews chart corresponds to the average maximum sized opening that lies in the transition between no support required and support required. The HR for this mid-point has also been presented for comparison purposes.

Figure 16.14 Application of the Mathews/Potvin open stope stability graph to the back of a stope located at a depth of 250 m.

16.3.12.2 2005 Stope Stability Assessment

A preliminary stope stability assessment was carried out using the Mathews/Potvin Stability Graph. The results were presented in the Golder (2005) Technical Memorandum. This initial assessment estimated the following:

TABLE 16.22 PRELIMINARY VALUES OF MAXIMUM HYDRAULIC RADIUS (HR) FOR UNSUPPORTED WALLS Depth (M) Hr Back (m) Hr for the Hw (m) Hr for the Sidewalls (m) 100 4.3 7.5 6.8 250 3.9 7.5 5.4

Based on these hydraulic radius, HR values and considering that the stopes would not be supported, for the Scoping Level Study, it was recommended that the stopes have a strike length of 12 m, stope height of 25 m and stope length (in the transverse direction) of 30 m to 35 m, creating HR values for the back of 4.3 to 4.5 m, HR for the HW of 4.1 m, and HR for the P&E Mining Consultants Inc., Report No. 247 Page 143 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. sidewalls of 6.8 m to 7.3 m. In addition, 12 m wide sill and rib pillars were recommended to be used between the LZ (footwall) and MZ (hanging wall) series of stopes. The 2005 analyses considered only a single opening and did not evaluate the interaction/redistribution of stresses that would result from multiple stope openings and pit advancement.

16.3.12.3 2010 Stope Stability Assessment

For the updates to the 2005 work, additional stability analyses were carried out to assess the induced stresses generated as a result of the interaction of the underground and open pit excavation. Three-dimensional numerical modeling was carried out using the boundary element code MAP3D©. Based on the numerical results, the HR were modified from those presented in the Scoping Study.

Values of hydraulic radius were calculated for the geometries of the stopes provided by P & E. The calculated HR values were then compared to those shown in Table 16.23.

The stope geometries that initially exceeded the recommended HR values were highlighted and re-designed in order not to exceed the „maximum stable unsupported‟ value of HR.

TABLE 16.23 REVISED VALUES OF MAXIMUM HYDRAULIC RADIUS (HR) FOR UNSUPPORTED WALLS HYDRAULIC RADIUS 1 MAXIMUM AVERAGE Calculated Rock Type Q' A B C STABLE UNSUPPORTED N' UNSUPPORTED TRANSITION (m) (m) Depth < 150 m BACK STABILITY Metasediments (-1 std.) 9.3 0.90 0.5 2 8.4 5.1 6.8 Metasediments Stable (mean) 14.9 0.90 0.5 2.0 13.4 6.1 7.9 Metasediments Stable (+1 std.) 24.0 0.90 0.5 2.0 21.6 7.3 9.2 HW STABILITY Metasediments (-1 std.) 9.3 1.00 0.2 4 7.45 4.9 6.5 Metasediments Stable (mean) 14.9 1.00 0.2 4 11.94 5.9 7.6 Metasediments Stable (+1 std.) 24.0 1.00 0.2 4 19.16 7.0 8.9 SIDEWALL STABILITY Metasediments (-1 std.) 9.3 0.55 0.5 4.0 10.2 5.5 7.2 Metasediments Stable (mean) 14.9 0.55 0.5 4.0 16.4 6.6 8.4 Metasediments Stable (+1 std.) 24.0 0.55 0.5 4.0 26.3 7.9 9.8 Depth > 150 m BACK STABILITY Metasediments (-1 std.) 9.3 0.45 0.5 2.0 4.2 4.0 5.4 Metasediments Stable (mean) 14.9 0.45 0.5 2.0 6.7 4.7 6.3 Metasediments Stable (+1 std.) 24.0 0.45 0.5 2.0 10.8 5.6 7.3 HW STABILITY Metasediments (-1 std.) 9.3 0.50 0.2 4.0 3.7 3.8 5.2 Metasediments Stable (mean) 14.9 0.50 0.2 4.0 6.0 4.5 6.1 Metasediments Stable (+1 std.) 24.0 0.50 0.2 4.0 9.6 5.4 7.1 SIDEWALL STABILITY Metasediments (-1 std.) 9.3 0.30 0.5 4.0 5.6 4.4 5.9 Metasediments Stable (mean) 14.9 0.30 0.5 4.0 9.0 5.3 6.9 Metasediments Stable (+1 std.) 24.0 0.30 0.5 4.0 14.4 6.3 8.1 1 Recommended range of HR values are highlighted in grey.

16.3.13 Stope Backfilling

Various options of backfilling for the underground stopes were discussed in detail during design meetings. Stopes beneath the direct excavation of the pit are required to be backfilled as the pit floor approaches them, in order to provide a secure working floor. The mine plan is that this void

P&E Mining Consultants Inc., Report No. 247 Page 144 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. filling will be accomplished from the open pit, since the stopes will be left empty, i.e., there will be no backfilling during the underground mine.

It is anticipated that as the footprint of the pit floor will be bigger than that of the underground stopes sufficient backfilling material should be available. To supplement backfilling, if required, low grade waste rock from surface operations can be used.

Stopes located immediately underneath the planned pit ramp must be fully backfilled prior to advancing the ramp. These stopes will likely require drop raises to be drilled from the surface for placement of muck, and the stopes would be tightly filled prior to advancing the ramp over them. Fill utilized for tight filling the tops of these stopes could consist of a slurry muck (or cemented rock fill) or lean mix concrete. Such mixes have been used for stope backfill in other similar projects utilizing a mix design as shown below in Table 16.24.

TABLE 16.24 EXAMPLE OF SLURRY AND LEAN MIX CONCRETE DESIGN Coarse Fine 28-Day Type of Water Cement Slump Aggregate Aggregate Compressive Mix (kg/m3) (kg/m3) (mm) (kg/m3) (kg/m3) Strength (MPa) Slurry 114 132 1650 412 (20%) 40 ± 10 15 Muck Lean Mix 170 ± 210 100 1040 1130 1 Concrete 20

16.3.14 2011 Planned Stopes

Most of the design issues discussions presented in Section 16.8 resulted in a re-evaluation by Fortune and P&E on the mine plan and design philosophy. A decision was made in 2011 to:

 Use stope dimensions that would minimize the potential for dilution;  Not carry out any backfilling during underground operations; and  Minimize as much as practically possible the interaction of the stopes with the open pit.

Figure 16.15 through Figure 16.17 show the revised 2011 stopes and open pit phases 1B to 3, respectively. There will be minimum interaction of the stopes and the Phase 1B pit shell (Figure 16.15). In places where the stopes will intercept the wall, the berm was considered wide enough to accommodate any local sloughing from the stope. By the time that Phases 2 and 3 are excavated, almost all the stopes would be located inside the pit as depicted in Figure 16.16 and Figure 16.17.

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Figure 16.15 Phase 1B and 2011 Planned Underground Stopes

Figure 16.16 Phase 2 and Planned Underground Stopes

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Figure 16.17 Phase 3 and 2011 Planned Underground Stopes

Based on the 2010 planned stope dimensions, HR values were calculated for the back and walls, as shown in Table 16.25 and Table 16.26, respectively. These HR values were then compared to the target HR ones presented in Table 16.22. As observed, the planned stope dimensions (i.e., back and hanging walls) are within the recommended HR values and systematic ground support (e.g., cable-bolts) is not anticipated to be required.

TABLE 16.25 HYDRAULIC RADIUS VALUES FOR 2011 PLANNED STOPES – BACK DIMENSIONS Back Hydraulic Depth Strike Width Area Perimeter Target Stope # Tonnes Radius, (m) Length (m) (m2) (m) HR HR (m) (m) Stope 1 16,637 124 35 15 525 100 5.3 6.1 Stope 2 7,222 124 15 15 225 60 3.8 6.1 Stope 3 14,459 90 30 15 450 90 5.0 6.1 Stope 4 21,166 110 45 15 675 120 5.6 6.1 Stope 5 17260 170 25 15 375 80 4.7 4.7 Stope 6 17,517 150 25 15 375 80 4.7 4.7 Stope 7 9,035 185 20 15 300 70 4.3 4.7 Stope 8 14,040 >150 45 10 450 110 4.1 4.7 Stope 9 5,909 >150 20 10 200 60 3.3 4.7 Stope 10 17,347 >150 25 15 375 80 4.7 4.7 Stope 11 9,575 >150 15 15 225 60 3.8 4.7 Stope 12 16,149 >150 25 15 375 80 4.7 4.7 Stope 13 16,498 >150 25 15 375 80 4.7 4.7 Stope 14 16,173 >150 25 15 375 80 4.7 4.7 Stope 15 22,765 <150 35 15 525 100 5.3 6.1 Stope 16 13,938 <150 24 12 288 72 4.0 6.1 (RL190S2) P&E Mining Consultants Inc., Report No. 247 Page 147 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

TABLE 16.25 HYDRAULIC RADIUS VALUES FOR 2011 PLANNED STOPES – BACK DIMENSIONS Back Hydraulic Depth Strike Width Area Perimeter Target Stope # Tonnes Radius, (m) Length (m) (m2) (m) HR HR (m) (m) Stope 17 13,382 <150 20 12 240 64 3.8 6.1 (RL190S1) Stope 18 20,800 <150 25 15 375 80 4.7 6.1 Stope 19 15,133 <150 25 15 375 80 4.7 6.1 Stope 20 15,271 <150 25 15 375 80 4.7 6.1 Stope 21 14,674 <150 25 15 375 80 4.7 6.1

TABLE 16.26 HYDRAULIC RADIUS VALUES FOR 2011 PLANNED STOPES – SIDEWALL DIMENSIONS Stope Hydraulic Depth Width Area Perimeter Target Stope # Tonnes Height Radius, (m) (m) (m2) (m) HR (m) HR (m) Stope 1 16,637 124 11 35 385 92 4.2 5.9 Stope 2 7,222 124 10 15 150 50 3.0 5.9 Stope 3 14,459 90 12 30 360 84 4.3 5.9 Stope 4 21,166 110 15 45 675 120 5.6 5.9 Stope 5 17260 170 17 25 425 84 5.1 5.3 Stope 6 17,517 150 17 25 425 84 5.1 5.3 Stope 7 9,035 185 15 20 300 70 4.3 5.3 Stope 8 14,040 >150 11 45 495 112 4.4 5.3 Stope 9 5,909 >150 13 20 260 66 3.9 5.3 Stope 10 17,347 >150 18 25 450 86 5.2 5.3 Stope 11 9,575 >150 18 15 270 66 4.1 5.3 Stope 12 16,149 >150 18 25 450 86 5.2 5.3 Stope 13 16,498 >150 18 25 450 86 5.2 5.3 Stope 14 16,173 >150 18 25 450 86 5.2 5.3 Stope 15 22,765 <150 18 35 630 106 5.9 5.9 Stope 16 13,938 <150 20 12 240 64 3.8 5.9 (RL190S2) Stope 17 13,382 <150 20 12 240 64 3.8 5.9 (RL190S1) Stope 18 20,800 <150 22 25 550 94 5.9 5.9 Stope 19 15,133 <150 16 25 400 82 4.9 5.9 Stope 20 15,271 <150 16 25 400 82 4.9 5.9 Stope 21 14,674 <150 16 25 400 82 4.9 5.9

Recognizing that the interactions of the 2011 stopes with the phase pit shells have reduced considerably, as illustrated in Figure 16.15 to Figure 16.17, it was decided that it would not be necessary to do additional three-dimensional numerical modeling.

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16.4 MISCELLANEOUS UNDERGROUND DRAWINGS

16.4.1 Underground Development Drawings

Figure 16.18 Existing Fresh Air Raise

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Figure 16.19 Plan Showing 195 – 215 Level Development

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Figure 16.20 Plan Showing 170 – 195 Level Development

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Figure 16.21 Plan Showing 161 Level Development

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Figure 16.22 Plan Showing 135 - 141 Level Development

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Figure 16.23 Plan Showing 95 - 116 Level Development

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Figure 16.24 Entrance View of Ramp Portal

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Figure 16.25 Underground Explosive Detonator Magazine GA

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Figure 16.26 Surface Mine Air Heaters GA

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17.0 RECOVERY METHODS

17.1 NICO CONCENTRATOR (DWG-0000-F-001)

The mineral processing plant described below is for the treatment of a bismuth-cobalt-gold sulphide ore, mined from underground and open pit, at a design throughput rate of 5160 t/d. The mineral processing plant at the NICO site will produce a bulk concentrate, which will be packaged and shipped to the Saskatchewan Metals Processing Plant (SMPP) for further processing.

ROM ore will be delivered by haul trucks from underground (for the first 8 months) and the open-pit mine. The ore will usually be dumped directly into the dump pocket of the primary crusher, but can also be dumped on the ROM receiving pad and reclaimed by a front-end loader.

The dump hopper has a capacity of 180 t and has only one access side.

A heavy-duty, hydraulic rock-breaker is provided to break up oversize boulders. ROM ore discharges from the dump hopper onto an apron feeder and is fed into a jaw crusher.

The primary jaw crusher is capable of crushing large ROM ore up to 500 mm in diameter. The jaw crusher is driven directly by a 250-kW motor through an extension drive shaft. The crusher will operate at a nominal open side setting of 150 mm to produce a product with a P80 of approximately 120 mm, at an average throughput rate of 516 t/h. The jaw crusher discharges the crushed ore onto the primary crushed ore conveyor.

The primary crushed ore conveyor discharges crushed ore onto the vibrating screen No. 01 ahead of the secondary cone crusher. This scalping screen is a double-deck vibrating screen with openings of 50 mm on the top deck and 16 mm on the bottom deck. The screen undersize is final crushed product and discharges onto the fine ore to stockpile conveyor. This conveyor transfers the screen undersize product to the fine ore stockpile. The screen oversize from both decks feeds the secondary cone crusher by gravity via a feed bin. The secondary crusher is a standard 2.135- m cone crusher driven by a 261-kW electric motor. The secondary crusher discharges onto the secondary crushed ore conveyor. This conveyor transfers crushed ore to the secondary crushed ore to TT2 conveyor. Then, it discharges onto the secondary crushed ore from TT2 conveyor.

Secondary crushed ore from TT2 conveyor feeds the tertiary cone crusher feed bin. From there, the material is extracted by two vibrating feeders and each feeder discharges onto a vibrating screen ahead of a tertiary cone crusher. The screens are double-decked vibrating screen with openings of 25 mm on the top deck and 16 mm on the bottom deck. The screen oversize from both decks feed the tertiary cone crusher by gravity. Each of the tertiary cone crushers is a short head 2.135-m cone crusher driven by a 261-kW electric motor. The screen undersize is final crushed product and discharges onto the fine ore to stockpile conveyor.

The fine ore stockpile provides crushed ore surge capacity so that the process plant can be supplied with a continuous source of feedstock. The total live capacity of the fine ore stockpile is 3000 t, representing about 14 hours of operation at the design throughput rate of 215 t/h.

The fine ore reclaim system consists of three fine ore belt feeders. The feeders deliver the fine crushed ore to the ball mill via the ball mill feed conveyor.

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A 16.5-ft by 24-ft long (5.03 m by 7.32 m) ball mill, driven by a 3200-kW motor is operated in closed circuit with a cluster of cyclones from which the overflows become flotation feed and the majority of the underflow returns by gravity to the ball mill and the rest to the gravity screen. The discharge from the mill passes through a trommel, mounted on the mill discharge flanges, to remove tramp metal. The discharge underflow product from the ball mill trommel is directed by gravity via a launder into the cyclone feed pumpbox. From this pumpbox, the slurry is pumped to the grinding cyclone cluster.

The cyclone cluster is fed by one of two rubber-lined horizontal centrifugal slurry pumps. The cyclone feed pumps are driven by 336-kW motors.

The cyclone cluster consists of nine 381-mm-dia cyclones with eight units operating under nominal flow conditions. The underflow is collected in a rubber-lined launder from which a portion flows by gravity to the gold gravity circuit and the rest to the ball mill feed spout. The cyclone overflow products (80% passing 74 μm) flows by gravity to the bulk flotation feed pumpbox.

The gravity screen is a single deck screen with 2-mm aperture. The oversize from the gravity screen returns by gravity to the cyclone feed pumpbox. The undersize from the gravity screen flows by gravity to the gravity concentrator. Gravity concentrator tails return by gravity to the cyclone feed pumpbox and the gravity concentrator concentrate is pumped to the bulk concentrate thickener by a vertical tank pump.

Slurry from the bulk flotation feed pumpbox is pumped to the first bulk flotation cell. Flotation reagents that will be added to the feed box of the first flotation tank include Potassium amyl xanthate (“PAX”) collector and methyl isobutyl carbinol (“MIBC”) frother. PAX and MIBC can be also added at the fourth cell of the bulk rougher flotation cell bank.

Bulk rougher flotation consists of a bank of five 70 m3 circular tank-type flotation cells. Each flotation cell has an agitator. The flotation cells are stepped for gravity flow. The first rougher flotation cell is a single cell, while the last four are grouped in pairs. Bulk rougher concentrate froths are collected from one side of the flotation cells and flow by gravity into pipe launders leading to the bulk rougher concentrate vertical tank pump sump.

The bulk rougher concentrate froth is pumped to the first bulk cleaner cell using a vertical tank pump.

Tails from the bulk rougher flotation are sent to the tailings thickener.

The bulk cleaner and bulk cleaner scavenger section consists of a bank of seven 10 m3 circular tank-type flotation cells. The flotation bank is configured such that bulk cleaner flotation is carried out in the first four cells, while bulk cleaner-scavenger flotation in the final three cells. The cells are equipped with agitators. All flotation cells are single cells and stepped for gravity flow. PAX collector is added to the first cleaner and first cleaner-scavenger cells.

Bulk cleaner concentrate and bulk cleaner scavenger froths are collected from one side of the flotation cells and flow by gravity into pipe launders leading to respective bulk cleaner and cleaner scavenger concentrate vertical tank pumps.

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The bulk second cleaner flotation section was added based on the results of the locked cycle and FLEET testwork performed by SGS in 2009. The design is based on Option 4 of the SGS Lakefield FLEET report.

The bulk cleaner concentrate froth is pumped into a bank of two 10 m3 circular tank-type flotation cells for bulk second cleaner flotation. The two cells are stepped for gravity flow, and are equipped with agitators. PAX collector can be added to the first cell. The bulk second cleaner concentrate is collected from the side of the flotation cells, and flows by gravity into a pipe launder leading to a vertical tank pump. The concentrate froth is then pumped to the bulk concentrate thickener. The bulk second cleaner tails flows into a pumpbox by gravity, and is pumped by one of two horizontal centrifugal pumps as feed into the bulk cleaner.

The bulk cleaner-scavenger concentrate froth is pumped back to the first bulk cleaner cell feed using a vertical tank pump. The bulk cleaner scavenger tails are pumped to the cleaner tails regrind circuit using one of two horizontal centrifugal pumps.

The cleaner tails regrind circuit consists of a cyclone cluster ahead of two regrind mills operating in closed circuit.

The tails from the bulk cleaner-scavenger flotation are collected in the cleaner tails regrind cyclone feed pumpbox. This slurry is then pumped to the regrind cyclone cluster by one of two rubber-lined horizontal centrifugal pumps. The cyclone cluster consists of five 254 mm dia. cyclones of, with three of them operating and two on standby. Cyclone underflow is collected in the underflow launder which is equipped with two outlets. Each outlet has an isolation knife gate valve and is piped to a regrind mill. This enables the cyclone underflow to be fed directly into one of the two regrind mills utilizing the gravity head available. The cyclone overflow product, with a particle size of 80% passing 20 microns, is collected in the overflow launder and then flows by gravity to the secondary rougher flotation feed pumpbox.

The regrind mills are vertical Stirred Media Detritor (“SMD-185”) type, driven by a 186-kW motor. The mill is equipped with a rubber liner for the body and media retention screens.

The secondary rougher consists of a bank of five 10 m3 circular tank-type flotation cells. The cells are equipped with agitators. The flotation cells are stepped for gravity flow. The first flotation cell is a single cell, while the last four are grouped in pairs of two. PAX collector and MIBC frother are added to the first secondary rougher cell. MIBC frother can also be added to the fourth secondary rougher cell. Secondary rougher concentrate froths are collected from one side of the flotation cells and flow by gravity into pipe launders leading to the secondary rougher concentrate vertical tank pump sump.

The secondary rougher concentrate froth flows is pumped to secondary cleaner flotation using a vertical tank pump.

The last secondary rougher cell discharges into the secondary flotation tails pumpbox, where it is combined with the secondary cleaner scavenger tails, before being pumped to the flotation tails pumpbox.

The secondary cleaner and cleaner-scavenger flotation section was added based on the results of the locked cycle and FLEET testwork performed by SGS in 2009. The design is based on Option 6 of the SGS FLEET report and the 2010 SGS pilot plant report. P&E Mining Consultants Inc., Report No. 247 Page 160 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

The secondary cleaner and secondary cleaner scavenger section consists of a bank of four 5 m3 circular tank-type flotation cells. The flotation bank is configured such that secondary cleaner flotation is carried out in the first two cells, while cleaner-scavenger flotation in the final two cells. The cells are equipped with agitators. All flotation cells are single cells and stepped for gravity flow. PAX collector is added to the first cleaner and first cleaner-scavenger cells.

Secondary cleaner concentrate and secondary cleaner scavenger froths are collected from one side of the flotation cells and flow by gravity into pipe launders leading to respective secondary cleaner and cleaner scavenger concentrate vertical tank pumps.

The secondary cleaner concentrate froth is pumped to the bulk concentrate thickener using a vertical tank pump.

The secondary cleaner-scavenger concentrate froth is pumped back to the cleaner tails regrind cyclone feed pumpbox using a vertical tank pump. The secondary cleaner scavenger tails are discharged into the secondary flotation tails pumpbox.

The final bulk concentrate from bulk second cleaner and secondary cleaner flotation cell is pumped to the 5.8 m-diameter bulk concentrate thickener. The thickener is of conventional design with a bridge-mounted drive mechanism. Thickener overflow is first directed by gravity into a thickener overflow tank and then pumped to the concentrator process water tank. Thickener underflow is pumped by a horizontal centrifugal pump into the bulk concentrate thickener underflow tank.

The thickened bulk concentrate is then pumped by one of two horizontal centrifugal pumps to feed the bulk concentrate pressure filter. The pressure filter is a fully-automatic vertical recessed plate type filter. The discharged cake from the pressure filter will contain less than 8% moisture. The filtrate collected in the filtrate pumpbox is pumped to the concentrator process water tank. The filter cake is discharged onto the concentrate loadout conveyor.

The concentrate loadout conveyor is a reversible conveyor. Under normal operation, the belt conveyor conveys forward to feed the concentrate packaging system. The packaging system is of dual-bag design, and a flop gate within the system is used to divert the feed from one bulk bag to another. The reversible conveyor is stopped automatically when both bulk bags are filled to approximately 1.5 m3 capacity and no empty bag is available. A roller conveyor and forklift move skidded bulk bags to product storage, until a trailer is ready at the loading area.

Bulk rougher tails, secondary rougher tails and secondary cleaner-scavenger tails from the flotation circuit are pumped to the 12.0 m diameter tailings thickener. The thickener is of high- rate paste thickener design with a bridge-mounted drive mechanism. The thickener rake can be raised by means of an automatic motorised lifting device, if the torque exceeds a pre-set value. Thickener underflow is pumped by one of two positive displacement pumps to the co-disposal area. Thickener overflow is first directed by gravity into a thickener overflow tank and then pumped to the concentrator process water tank.

17.2 NICO PLANT UTILITIES

The service building will house four air compressors, with two operating and two on standby, to provide compressed plant air, instrumentation air and air for the pressure filter. P&E Mining Consultants Inc., Report No. 247 Page 161 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Two pairs of flotation air blowers are also provided in the service building, for supplying air to the flotation circuit.

The fresh water supply for the NICO concentrator users and the potable water treatment plant will be from Lou Lake.

The fresh water tank is located in close vicinity to the concentrator, and has a capacity of 70 m3. Water is distributed from the tank to the various fresh water users.

Fresh water is pumped to a boiler package to generate hot water for general uses around the plant. There is one boiler in the NICO plant. The main plant and camp heating is provided by waste heat recovery from the diesel power plant.

Fresh water for making potable water is first treated in a 2.9 m3/h potable water treatment plant before storing in the potable water supply tank. Potable water is supplied to all end users by potable water supply pumps.

The concentrator process water comprises of the overflows from the tailings thickener and bulk concentrate thickener, bulk concentrate filtrates, as well as pumped reclaim water from the surge pond.

PAX, MIBC and flocculant will be delivered to the plant and prepared as required.

17.3 SMPP PROCESS DESCRIPTION (DWG-0000-F-001/002)

The SMPP described in the following sections is for the treatment of the filtered bulk flotation concentrate containing gold, cobalt, bismuth and copper from the NICO mine and concentrator in the NWT.

The SMPP will treat approximately 217 wet tonnes of bulk concentrate per day, at an overall plant availability of 85%. This is equivalent to the average daily production rate of the concentrator (incorporating availability).

The bulk concentrate first will be treated to produce a separate bismuth concentrate and cobalt concentrate via a regrind to 14 microns and then a secondary flotation circuit.

The bismuth concentrate will be treated in the CLER circuit to produce bismuth cathode product. The CLER circuit bismuth residue and the cobalt concentrate will undergo pressure hydrometallurgical treatment and multiple purification steps ultimately to meet the target 99.8% cobalt specification cobalt cathode „puck‟. The term „puck‟ was chosen specifically to describe the round cathode product produced by the facility. Alternatively, cobalt pregnant solution can be processed by solvent extraction to remove metal impurities, followed by a three-stage crystallization process to produce a cobalt sulphate heptahydrate product with 20.9% cobalt content.

The residue from the pressure treatment, which contains significant amount of gold, will be leached with cyanide. The gold in solution will be precipitated with zinc powder in a Merrill Crowe circuit, and then refined to produce gold doré.

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The main process and operation units of the SMPP facility are described below. An overall simplified flow diagram for the entire plant, including the mineral processing and the hydrometallurgical facilities, is shown in drawings 0000-F-001/002.

Feed Preparation

Bulk concentrate arrives at the SMPP facility in 3.1 t FIBC bulk bags loaded on railcars. A 15 t gantry crane lifts the bulk bags from the railcars and into the storage area, from where each bag is moved to the feed preparation area by forklift.

Facility throughput is 9 t/h of bulk concentrate or the equivalent of a bulk bag approximately every 21 minutes. Slurry samples will be withdrawn for an on-stream analyzer (OSA).

17.3.1 Bulk Concentrate Regrind and Bismuth Flotation

Bulk Concentrate Regrind

The bulk concentrate regrind circuit consists of a cyclone cluster and a pair of regrind mills operating in closed circuit.

Two vertical Stirred Media Detritors (SMD) grinding mills operate in parallel, each driven by a 355 kW motor and using 3 mm Colorado sand for grinding media.

The repulped bulk concentrate slurry is collected in the regrind cyclone feed pumpbox, along with the recycled bismuth first cleaner tails. Sodium cyanide (NaCN) is added to the slurry for depression of cobaltite and arsenopyrite in the downstream flotation circuit. The slurry is adjusted to 28% solids by adding flotation water, then pumped to the regrind cyclone cluster.

Cyclone underflow feeds directly to the regrind mills by gravity. The SMD mills discharge to the cyclone feed pumpbox. Cyclone overflow, with a particle size of P80 15 microns is collected in the bismuth rougher feed pumpbox.

Bismuth Rougher Flotation

The rougher flotation stage spans six cells and the scavenger flotation stage is the last three. PAX collector and NaCN depressant are added to both rougher and scavenger cell banks, while MIBC frother is added only to the rougher cell bank.

The bismuth rougher concentrate is pumped to the bismuth first cleaner section by a froth pump.

The bismuth rougher-scavenger concentrate is pumped to the regrind circuit with a froth pump.

The last bismuth rougher-scavenger cell discharges through a gravity sampler and is pumped to the cobalt concentrate thickener using horizontal centrifugal pumps.

Bismuth Cleaner Flotation

Bismuth first cleaner flotation consists of seven conventional cells with cascading flow and individual paddles. NaCN depressant can be added to the bismuth first cleaner cell. Bismuth first cleaner concentrate froth is pumped to the bismuth second cleaner section by a froth pump. P&E Mining Consultants Inc., Report No. 247 Page 163 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Bismuth first cleaner tails are pumped back to the regrind cyclone feed pumpbox with horizontal centrifugal pumps.

Bismuth second cleaner flotation consists of eight conventional cells. NaCN depressant can be added to the bismuth second cleaner cell. Bismuth second cleaner concentrate is pumped to the bismuth third cleaner. Bismuth second cleaner tails flow by gravity to the bismuth first cleaner.

Bismuth third cleaner flotation consists of two conventional flotation cells with cascading flow. Bismuth third cleaner concentrate is pumped to the OSA section using a froth pump and then to the bismuth concentrate thickener of the CLER circuit via the analyzer return pump. Bismuth third cleaner tails flow by gravity to the bismuth second cleaner.

All three flotation cell banks are connected to spray water for froth cleaning, control and dilution.

Cobalt Concentrate Thickening

Cobalt concentrate (bismuth rougher-scavenger tails) is pumped to the cobalt concentrate thickener to obtain an underflow slurry of 60 wt% solids with a flocculant dosage of 50 g/t feed.

Flotation Area Services

A forced air system, consisting of low-pressure blowers and distribution piping, supplies air to each flotation cell.

The regrind and bismuth flotation area is divided into three bundled areas: regrind, flotation and cobalt thickener. Three floor sump pumps return wash water and spills back to their respective dedicated process circuits.

The regrind and bismuth flotation area is equipped with a 5 t overhead crane.

On-stream Analyzer System

An OSA system is provided in the flotation area for near real-time flotation product grade information.

The element assays include cobalt, bismuth, copper, iron and arsenic. In combination with intermediate stream grade information, on-line calculations of recovery and slurry solid percentage will be used to stabilize and optimize the circuits.

Bismuth Processing Plant

Bismuth concentrate, received as slurry from the Bismuth Flotation circuit contains about 46% bismuth as bismuth sulphide and elemental bismuth. The slurry is subjected to two-stage leaching at elevated temperature in a concentrated ferric/ferrous chloride solution with intermediate liquid/solid separation. Leached residue is filtered and washed to remove values and chlorides and returned to the hydromet plant for recovery of contained gold. Pregnant solution from the first leach stage is clarified, partially evaporated for control of the solution balance, and electrowon to recover metallic bismuth as a powder product. The bismuth powder is filtered, dried, briquetted and melted to ingot. A bleed stream is directed to a scavenger EW cell to P&E Mining Consultants Inc., Report No. 247 Page 164 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. recover residual bismuth and then treated with lime to precipitate iron and other impurities. A washed gypsum cake is suitable for disposal, and the effluent solution is directed to the treatment facility for injection into the saline aquifer.

Leach and Thickening

The two stage leach is necessary for control of iron oxidation because the pregnant solution returning to EW must have iron in the reduced state to maximize current efficiency. The purpose of the first stage is primarily to convert all ferric iron to ferrous using the reducing capacity of the concentrate. The second leach stage, after liquid/solid separation, is conducted in recycled EW anolyte containing iron, substantially in the ferric state. Given sufficient ferric iron, leaching of bismuth is effectively quantitative, with about 1% of the metal remaining in the residue solids.

The leach is conducted at elevated temperature at relatively low slurry density. The incremental heat required to maintain design temperature is partially supplied by heats of reaction and partially by indirect heating with steam from the hydromet plant boilers. During this process, minor bismuth is leached and some elemental sulphur is formed. Arsenic, present in the solution advanced from the second-stage leach, is reduced to a very low concentration by precipitation. The leached slurry is thickened in a conventional thickener and the underflow is pumped to the second-stage leach. Thickener overflow is directed to pre-coat pressure filters for clarification prior to EW. The overall leach process is controlled to yield a controlled pregnant solution concentration of bismuth. This requires balancing of the concentrate feed rate with solution flows and EW production. Short term control is assisted by the buffering capability of a relatively large solution inventory.

First stage thickener underflow is directed to the second stage leach, maintaining leach temperature and slurry density. During the second stage leach substantially all of the bismuth is solubilized and the associated sulphide is converted to elemental sulphur. The leach solution is comprised of anolyte from the electrowinning circuit containing ferric iron and a nominal design level of chloride plus any required make-up sodium chloride and sulphuric acid to offset solution losses to the bleed stream. Combined Stage 1 and Stage 2 leach extraction is approximately 99%.

The leach discharge is thickened in a conventional thickener. Thickener underflow is cooled in a double-pipe heat exchanger and filtered on a horizontal belt filter. The filter cake is washed in three or four stages to remove soluble bismuth and contained chlorides. The residue filter cake contains recoverable gold and is returned to the hydromet plant for cyanidation.

Filtrate from the residue filter is recycled to the second-stage thickener and wash filtrate is directed to the scavenger EW circuit for recovery of contained bismuth. The second-stage thickener overflow is pumped to first-stage leaching and used to re-pulp concentrate feed filter cake.

Electrowinning

EW feed solution joins re-circulated catholyte and is distributed to thirteen of sixteen EW cells. Each cell contains 30 copper cathodes and 32 DSA anodes which are separated by IX membranes. At the design amperage, bismuth plates as a non-adhering powder at the cathodes. The cells incorporate design features to allow for continuous withdrawal of powder which is recovered in a pressure filter. The solution accompanying the powder product is re-circulated to EW feed. P&E Mining Consultants Inc., Report No. 247 Page 165 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Catholyte level is maintained with a slightly positive head to the anolyte to maintain a pressure differential to ensure no backflow through the membrane. In the anolyte compartment, ferrous iron is converted to ferric to complete the circuit and regenerate oxidant required in the leach circuit. The anolyte is re-circulated to second stage leaching as lixiviant.

A bleed of catholyte controls iron levels in the circuit, which joins wash filtrate from the residue filter as feed to EW scavenger cells. The scavenger cell circuit mirrors the production cell circuit and recovered bismuth powder is combined with product from the primary EW circuit. Anolyte from the scavenger cell is directed to the iron removal circuit.

A consequence continuous production of bismuth powder, there is little handling of the anodes and cathodes. In test, there was no evidence of fouling, consuming of the plates or other issues that would require maintenance. It is therefore expected that service requirements will be restricted to scheduled cleaning of cell components.

Iron Removal

Anolyte from the scavenger EW cell comprises the main exit stream from the circuit, required to complete the water balance and to stabilize iron levels in the circuit. Iron is precipitated and acid neutralized with lime in a series of tanks. Iron is substantially in the ferric state and an oxidant is unlikely to be required. Precipitated iron and gypsum, along with other minor impurities is separated from solution in a thickener followed by a horizontal belt filter. The filter cake is washed with hot water to remove chlorides from the cake. The filter cake at about 0.14 t/d solids is discharged to a hopper and periodically removed to permanent residue storage. Thickener overflow, containing primarily sodium chloride and sulphate, is directed to the effluent facility for deep-well injection to the saline aquifer.

Product Recovery

Bismuth powder is separated from re-circulated catholyte in a pressure filter equipped for washing with a chloride solution. Periodically the filter is dumped to a hopper, slurried with diluted chloride and filtered on a horizontal belt filter where soluble chlorides are removed by washing with fresh water. Filtrate from the belt filter, after polishing in a cartridge filter to recover entrained bismuth, joins other effluent from the iron removal circuit for injection into the saline aquifer. Product filter cake is dried in a tunnel oven and briquetted, without binders, for melting with a nitrogen blanket provided to prevent oxidation of the bismuth. The briquettes are melted in a low-temperature induction furnace with caustic as a cover to prevent oxidation and to collect impurities. There may be a requirement to add a sulphur or sulphide flux for additional impurity removal. The furnace operation serves to upgrade the bismuth purity from about 99.5% as electrowon to 99.99%. Bismuth is poured in 12.5 kg ingots as a final product. Dross/slag from the furnace will contain some recoverable bismuth. The preferred method for slag treatment, which will be finalized in operation, may include dissolution of caustic in water for use as a base within the overall hydromet facility and/or recovery of residual bismuth by periodic re-melting or re-introduction to the leach circuit.

Bismuth Leached Residue Caustic Leach

The well-washed leached residue from the bismuth ferric chloride leaching circuit can be further leached in a caustic solution to recover by-products such as tellurium and selenium. The leached P&E Mining Consultants Inc., Report No. 247 Page 166 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. slurry is filtered and washed thoroughly. The filter cake is repulped in an agitated tank until it reaches a pumpable slurry density, at which point it is pumped to the bismuth residue cyanidation circuit.

The design for this bismuth leached residue caustic leach was not incorporated, as testwork was not been complete at the time of the report.

17.3.2 Pressure Oxidation

Cobalt Concentrate Emergency Storage

Thickened cobalt concentrate slurry is normally pumped continuously into the autoclave surge tank. , When the autoclave surge tank is full, cobalt concentrate slurry is diverted into the cobalt concentrate emergency tank with a 75 kW agitator. A storage capacity of 90 hours is available.

Pressure Oxidation, Thickening, Clarification and Filtration

Cobalt concentrate passes a trash screen into the agitated autoclave surge tank. Oversize rocks are returned to the grinding circuit, while foreign matter is removed to a trash bin and disposed. Repulped bismuth leach residue is pumped in to the agitated autoclave surge tank. The process stream undergoes dilution and pre-acidification during the 4 h retention of the surge tank.

Pre-acidification dissolves and removes carbonates contained in the ore. A surfactant is added at a rate of 6 kg per tonne of feed. based on pilot plant work. Actual operation will see lower consumption.

Cobalt concentrate and repulped bismuth leach residue are fed to the POX autoclave via an in-line basket strainer and a high pressure feed pump. The 16.6 m long, five-compartment autoclave provides a nominal 60 minutes of retention time, sufficient to leach up to 95% of the cobalt, 76% of the copper and 87% of the sulphides. The autoclave operates at 180°C and 2,100 kPa pressure. A vacuum pressure swing adsorption (“VPSA”) oxygen plant constructed on-site delivers 120 t/d pure oxygen at 93% concentration. The leaching process is highly exothermic, and cooling water from an autoclave quench water tank is pumped into each autoclave compartment for temperature control. Superheated steam at 180°C can be injected directly into the autoclave for start-up and maintenance.

A nuclear level gauge and a redundant scintillator in the last compartment, attenuate a let-down valve and choke valve to control discharge from the autoclave. Quench water is pumped into the slurry line before the choke valve to prevent flashing from occurring until the slurry reaches the flash tank. The leached slurry is reduced to a pressure of 180 kPa and a temperature of around 115°C in a flash tank. Slurry from the flash tank gravity flows to the cobalt residue thickener.

Autoclave off-gases are vented to the flash tank also. In the winter months, exhaust gases are directed into the autoclave heat recovery heat exchanger. The captured heat is transferred into a water/glycol loop for use in the thawing shed and to various parts of the plant for general building and tank heating. Residual gases from the autoclave heat recovery heat exchanger are directed to the scrubber, which uses cobalt barren solution as the make-up scrubbing agent. The cleaned gases are discharged to the atmosphere. The scrubber discharge is sent to the cobalt residue thickener. During periods when heat recovery is not required, gases from the flash tank will vent directly to scrubber, bypassing the heat exchanger. P&E Mining Consultants Inc., Report No. 247 Page 167 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Flash tank discharge slurry, scrubber discharge, and recovered fines from the subsequent clarifier are fed to a cobalt residue thickener. The rubber-lined conventional thickener thickens to 60 wt% solids with a 30 g/t flocculent dosage. A lamella clarifier provides further recovery of fine solids. The overflow solution from the clarifier, as cobalt PLS, is stored in a surge tank of approximately 1.6 hours residence time, and is pumped to the iron-arsenic precipitation step.

The cobalt residue thickener underflow is first pumped by horizontal centrifugal pumps to a cobalt residue filter surge tank, from where it is fed to a horizontal pressure belt filter for further dewatering. A 3:1 displacement wash ratio is required to ensure efficient washing. The collected filtrate from the filter is recycled back to the lamella clarifier feed by a horizontal centrifugal pump. The cobalt residue filter cake is discharged onto a filter cake discharge conveyor. The material is conveyed to a feed repulp mixer and then to an agitated tank containing a mixture of cyanide-rich solution and milk of lime. This cyanide rich solution includes the gold barren solution from the Merrill Crowe circuit. Milk of lime is added to the tank to maintain a pH of 10.5 to suppress the generation of toxic hydrogen cyanide gas. The filter cake is repulped to 45 wt% solids and the slurry is pumped to the cobalt residue cyanidation circuit for the recovery of gold.

17.3.3 Solution Neutralization and Iron-Arsenic Removal

Cobalt Iron-Arsenic Precipitation

Clarified cobalt PLS is first combined with several recycled streams containing metal precipitates in the solution collection tank. This collection tank allows the highly acidic PLS solution, at approximately 30 g/L H2SO4, to redissolve the valuable metal precipitates, specifically cobalt, as soluble sulphates.

Lime is added to the first, third and fifth precipitation tanks to neutralize acid and gradually bring the solution pH level up to 4.6 through five hours of residence time. Oxygen-enriched air is introduced to the precipitation tanks to provide the oxidative environment necessary to convert ferrous iron (Fe²+) to ferric (Fe³+) and precipitate. Hydrogen peroxide addition is also provided at the last Fe-As precipitation tank, as an extra dose of oxidant.

Two important criteria must be satisfied for the majority of arsenic in solution to precipitate with the iron as stable scorodite (FeAsO4·2H2O). The first is a pH of 4.6, which can be achieved easily by the addition of lime. The second is the availability of iron to arsenic, since by stoichiometry the minimum Fe:As ratio is 1:1.

The remaining iron in solution reacts with lime and precipitates as iron hydroxides to reach an iron tenor of 1 ppm or less. Approximately 80% of the copper, 2% of the cobalt and other minor metals also will precipitate as hydroxides. With the neutralization of acid with lime, a significant amount of gypsum also will be formed.

Slurry from the final precipitation tank overflows to a conventional thickener via a launder. Flocculant at 60 g/t of solids is added to the launder. A portion of the thickener underflow at 25 wt% solids is recycled back to the first Fe-As precipitation tank as precipitation seed, while the remainder is pumped downstream to the copper releach tank. Thickener overflow proceeds to further copper removal.

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Copper Releach and Iron-Arsenic Filtration

Acidic eluate solution from the stripping of resin in the downstream copper IX circuit, and make- up sulphuric acid are added to the thickened precipitates in two agitated copper releach tanks installed in series to reach a pH of 2. Cobalt barren solution is also added to adjust the slurry to approximately 20 wt% solids. The releach step is carried out for a total of 30 minutes. At this pH, the acid causes 85% of the copper and 60% of the cobalt to releach back into solution, while most iron and arsenic, as well as gypsum, remains as solids in the precipitates.

17.3.4 Copper Cementation and Precipitation

Copper Cementation

Two tanks in series with a combined residence time of 1 hour are provided in copper cementation. Fine iron powder is added to the solution, which reacts with the soluble copper at 50°C to precipitate elemental copper at a pH of 2.5. The discharged slurry overflows to a feed tank and then is pumped to a filter press to recover the copper solids. The solid cake undergoes hot water washes before being discharged from the filter. The filter press filtrate, which is barren of copper but contains soluble cobalt, is recycled back to the solution collection tank. The copper cement product is packaged into 1 t bulk bags, for transport off-site.

Copper Precipitation

Iron-arsenic thickener overflow solution is treated in a copper precipitation circuit to further remove any residual soluble copper. Copper precipitation is performed at 50°C in two agitated tanks with a total residence time of 50 minutes. Sodium carbonate is added continuously to precipitate the copper as copper carbonate and to maintain the pulp pH at 6.5. Precipitation discharge slurry overflows to an agitated filter feed tank before being pumped to a filter press, where the precipitated carbonate is separated from the liquor. No wash is applied on the filter cake, as it is repulped with clarified autoclave discharge solution and recycled back to the solution collection tank for solids releaching. Filtrate from precipitation, now completely barren of copper but rich in cobalt, is pumped forward to the cobalt precipitation stage.

17.3.5 Cobalt Precipitation

Cobalt Precipitation – Stage 1

Cobalt pregnant solution is stored in a 1 h capacity surge tank before advancing to a two-stage precipitation circuit for cobalt recovery.

Stage 1 cobalt precipitation includes four agitated precipitation tanks arranged in series, providing a total residence time of 2 h. Steam can be sparged into the first and second tanks to maintain the operating temperature above 50°C. Sodium carbonate is used in the first stage to precipitate cobalt at a pH of 7.4.

This first stage of cobalt precipitation recovers approximately 97% of the total cobalt in solution. In addition to cobalt, other metals such as copper, zinc and nickel are precipitated as carbonates at this pH.

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Underflow from the cobalt precipitation thickener, at 31 wt% solids, is partially recycled as precipitation seed to the head end of the precipitation circuit. The remaining underflow is pumped to a feed tank before being pumped to a centrifuge. Washed centrifuge cake is discharged and stored in the cobalt carbonate bin. From this surge bin, cobalt carbonate cake is continuously conveyed by a screw conveyor to the cobalt dissolution tanks in the ion exchange (IX) purification circuit

Cobalt Precipitation – Stage 2

In this precipitation stage, any soluble cobalt remaining in solution is precipitated with sodium hydroxide to form cobalt hydroxide. Stage 2 cobalt precipitation is carried out in three agitated tanks in series at a pH of 9.3.

At this pH, other metals such as nickel, manganese and magnesium also are precipitated as hydroxides. Steam can be sparged to the first tank to maintain the operating temperature above 50°C. After a residence time of 1.25 hours, the slurry overflows into a filter feed tank and from where it is pumped to a pressure candle filter. The cobalt hydroxide cake is repulped with clarified autoclave discharge solution in an agitated repulp tank, and is recycled back to the solution collection tank in the solution neutralization circuit for solids releaching. The filtered solution from the candle filter, which is depleted of any valuable metals, is collected in a cobalt barren solution tank. This solution is reused for filter cake washing by various pressure filters in the plant, solids repulping, lime slaking, and gas scrubbing.

17.3.6 Cobalt Dissolution and Ion Exchange (IX) Purification

Cobalt Dissolution

Cobalt carbonate from Stage 1 cobalt precipitation is screw conveyed to the first of two cobalt dissolution tanks in series. The tanks provide a total residence time of 2 hours. Steam can be injected by sparger to maintain the operating temperature above 60°C. Spent electrolyte (anolyte) solution from cobalt electrowinning is pumped to this tank to continuously redissolve the cobalt as cobalt sulphate to a tenor of 107 g/L of cobalt. Sulphuric acid is added to adjust and maintain the operating pH at 4.0. In addition to the cobalt, metals such as copper, zinc and nickel (as carbonates) are dissolved as metal sulphates. Discharged slurry from the dissolution tanks flows by gravity to a pumpbox, from where it is pumped to a cobalt dissolution filter feed tank and ultimately to one of two pressure candle filters installed in parallel.

Since a very pure cobalt solution is required for the electrowinning of high-grade cobalt to meet the LME Cobalt Specification, contaminants must be removed from the cobalt solution before EW. Provision for zinc, copper and nickel removal from the electrolyte is also included within the circuit using ion exchange technology.

Filtrate from the candle filters is collected in the cobalt dissolution filtrate surge tank that provides a residence time of 30 minutes. Filtrate is pumped to the downstream zinc ion exchange circuit. Cake from the filter is discharged via a chute to the cobalt dissolution residue repulp tank, where it is repulped with a cobalt solution or various IX backwash waters. The slurry is pumped back to the solution collection tank in the iron-arsenic precipitation stage.

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Zinc Ion-Exchange, Elution and Precipitation

Cobalt solution from the cobalt dissolution filtrate surge tank is pumped through a cartridge filter to remove any entrained fine solids before entering a continuous zinc ion exchange unit. Two cartridge filters are installed in parallel, one operating and one stand-by, so that the process is not interrupted during a cartridge change. In the continuous zinc IX unit, Lanxess VP-OC-1026 resin is used in 30 small IX columns to remove dissolved zinc ions at an operating temperature of 50°C. Solution is fed continuously to the resin columns in an upflow direction.

The target zinc content in the raffinate solution leaving the zinc IX unit is at or below 2 ppm. When the raffinate solution zinc concentration exceeds this target, known as a breakthrough, it indicates that some of the lead resin columns are loaded with metal ions and require elution. There are three steps during the elution operation. The first is a slow rinse step where acidified water at a pH of 2 is pumped through the column to remove 90% of the cobalt that has adsorbed onto the resin. This cobalt-rich solution is recycled and used to repulp cobalt dissolution residue in a repulp tank. The second step is a zinc elution step using 150 g/L sulphuric acid solution. The eluate solution, with a 4:1 cobalt-to-zinc ratio, is recycled back to the zinc IX eluant tank where fresh sulphuric acid is added to restore acid concentration to 150 g/L. This recirculation allows for build-up of the zinc tenor in solution. The final step is a fast rinse of the resin with acidified water at a pH 2.

The zinc IX eluate is sent to a surge tank before being pumped to an acid purification unit to recover sulphuric acid. The unit employs ion exchange resins to adsorb acids in the feed stream and uses demineralized water to desorb the acids. The recovered acid is collected in a surge tank before being pumped back to the zinc IX eluant tank for reuse.

The zinc containing eluate is sent to one of two zinc precipitation tanks in parallel, each with 12 hours of residence time. Sodium carbonate is used to precipitate zinc at a pH of 9.5. Zinc precipitation is carried out in batches.

Reacted slurry from the zinc precipitation tanks is pumped to a zinc precipitation pressure candle filter. Filtrate returns to the SMPP process water tank. Filter cake discharges via a chute into a zinc carbonate storage bin of 1 t capacity, from where it is packaged into bulk bags.

Raffinate solution leaving the zinc IX unit passes through two carbon columns in a lead / lag configuration. Activated carbon acts as an adsorber for the D2EHPA in solution. The concentration of D2EHPA in the stream is low; and thus the carbon columns are designed as fixed bed columns that do not require elution. A fresh carbon batch is charged in when the existing batch is exhausted. The cleaned zinc raffinate leaving the carbon column is collected in a storage tank, and is pumped through a plate-and-frame type heat exchanger where solution is heated by steam to 65°C.

Copper / Nickel Ion Exchange, Elution and Precipitation

Heated zinc raffinate is pumped into one of three duplex copper guard IX units in parallel. Each duplex unit operates in a lead / lag configuration for maximum copper removal and resin capacity. Lewatit MonoPlus TP207 resin is used in the copper columns to extract any copper in the zinc raffinate before it is sent to downstream nickel IX. These copper guard columns are required to prevent poisoning of the nickel IX resin by copper. When a breakthrough of copper is

P&E Mining Consultants Inc., Report No. 247 Page 171 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. detected in the copper IX discharge solution, the lead vessel of the duplex unit undergoes a regeneration cycle.

From the copper guard IX units, heated zinc raffinate is fed to the nickel IX unit. In this unit, Dow XUS-43578 resin is used in 30 small columns to remove dissolved nickel ions from solution at an operating temperature of between 60 and 65°C. Solution is fed to the resin columns continuously in a downflow direction. The target nickel content of the raffinate solution leaving the nickel IX unit is at or below 120 ppm. When the nickel concentration of the raffinate solution exceeds this target, the elution cycle is activated. There are five consecutive elution / rinse steps.

Nickel IX eluate is sent to a collection tank before being pumped to an APU to recover sulphuric acid. The unit operates in the same manner as the zinc APU unit. The recovered acid is collected in a surge tank before being pumped back to the nickel IX eluate tank for reuse. The nickel containing eluate is sent to one of two agitated nickel precipitation tanks in parallel, each with a residence time of 12 hours. Sodium carbonate is used to precipitate nickel at a pH of 9.5. Nickel precipitation is carried out in batches.

Reacted slurry from the nickel precipitation tank is pumped to a nickel precipitation pressure candle filter. Filtrate returns to the SMPP process water tank. Filter cake discharges via a chute into a nickel carbonate storage bin of 1 t capacity, from where it is packaged into bulk bags.

Nickel IX raffinate is collected in a cobalt strong electrolyte solution tank, where it is pumped to the cobalt electrowinning circuit. The solution now contains approximately 106 g/L of cobalt, 120 ppm of nickel and less than 2 ppm of zinc.

17.3.7 Cobalt Electrowinning

Cobalt Electrowinning Cells

Electrowinning has a metal deposition capacity of 6 t/d of cobalt.

In the EW cells, the electrolysis process reduces cobalt ions to cobalt metal, which deposits on the cathodes, and oxidizes the oxygen component of water to gaseous oxygen at the lead anodes. Hydrogen ions dissociate from the water, resulting in a net increase in the sulphuric acid concentration.

The anodes are enclosed in porous bags. Electrolyte is removed by suction tubes in the bags to maintain a differential head across the bag, which reduces the escaping of the “anolyte” in the bag to the bulk “catholyte”. The bag material and porosity also are chosen specifically to nearly eliminate the transfer of sulphuric acid, or more specifically the hydrogen ions formed at the anode, from reaching the cathode.

A trim heater using low pressure steam heats the spent electrolyte recycle to pre-set EW temperature between 60 and 75°C. The nickel IX raffinate solution, at 106 g/L cobalt is mixed with the heated spent electrolyte recycle at 1:5 ratio in the cobalt electrowinning circulating tank. Sodium carbonate is added to this strong electrolyte solution to adjust the pH to around 3.5. Two sets of spare reagent mixing tanks with agitators and bin vents are included in this area. They are provided for preparation of EW reagents, such as pH buffer or smoothing agents, if additional reagents are found to be required in operations for this area. P&E Mining Consultants Inc., Report No. 247 Page 172 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Each electrowinning cell contains 39 cathodes and 40 anodes. Cathodes are 316L stainless steel blanks with a 1.0 m by 1.0 m wetted area masked with dielectric heat cured epoxy paint. The masking will leave approximately 43% of the cathode blank wetted area exposed as 25 mm diameter circular areas to produce cobalt deposits as “pucks”. The nominal current density is 450 A/m² at a current efficiency of 95%. The deposition period is approximately 4 to 6 days. The nominal intercathode spacing is 160 mm. Fresh feed electrolyte is delivered to a manifold around the floor of each cell, which feeds into the interelectrode spaces.

The anodes are lead-antimony (6-10%) rolled sheet, 6 mm thick. In the cell, anodes are encased in individual membrane bags. The anode bags are suspended from a cap across the anode width over a frame, which separates the bag from the anode face. The cap is close fit over the anode suspension bar and is effectively gas tight. Anolyte and catholyte handling are carried out separately.

The anodic reaction will generate a large volume of oxygen, as well as sulphuric acid mist. At the cell surface, the gases are collected by the caps on the anode bags fitted to the top of the anodes. Each anode cap is piped to a gas collection manifold embedded in the EW cell wall. Anolyte, equal in volume to the incoming strong electrolyte, also is withdrawn from the same collection manifold by a dip tube extending below the bulk electrolyte level in the cell, to maintain a differential pressure of approximately 40 mm between the interior level and exterior level across the bag membrane. Using an exhaust fan, the gas and the anolyte solution are transported to a collection manifold, and then to a knock-out pot, followed by a wet scrubber, via a large bore pipe. The collected anolyte solution in the knock-out pot, with a target cobalt tenor of 45 g/L, is pumped back to the cobalt dissolution area. The gas is washed and scrubbed clean in the acid mist scrubber, and then discharged to the atmosphere. Sodium lauryl sulphate (SLS) may be used in the EW cells to further minimize mist generation.

The tankhouse has 16 cells arranged in one tier. The cells are fed hydraulically in parallel by centrifugal pumps. Electrically, the cells are arranged in series using Walker multiple connections. The circulation tank is divided into two compartments, separated by an underflowing baffle. The first compartment is fed with strong electrolyte and additives as described previously, as well as with the back mixing of catholyte from the second compartment. The cells are fed from the first compartment at a specific flow rate of 0.50 L/min·m² of overall wetted cathode area. Catholyte leaving the cells via the cell overflow box downcomers is collected in manifolds and returned to the second compartment of the circulation tank.

Power for the tankhouse is provided by one 17 KS, 72 VDC output rectifier.

Manganese sulphate is used for the maintenance of the lead anodes within the circuit. The manganese forms MnO2 on the anode, which provides a protective coating and prevents anode corrosion. It is suggested from test work to maintain a manganese concentration of 1.5 g/L in the electrolyte.

Cobalt Cathode Harvest and Wash

Cathode blanks are lifted in sets of 1/3 (every third cathode) from the cells and transported to a dip wash tank using the pendant-controlled overhead bridge crane.

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The stainless steel cathodes with cobalt cathode „pucks‟ are placed first in a hot water tank sparged with air. The harvested cathodes remain in this water bath, while the crane retrieves another one-third cell load of cathodes from an adjacent cell. These cathodes are lowered and placed adjacent to the cathodes already in the tank. The crane repeats the harvesting for a third time from another cell, and places the cathodes in the water bath. When the third set of cathodes is submerged, the crane moves to the first set of cathodes and transfers them to the hammering / stripping station.

The hot water tank is provided with a 316L stainless steel basket resting above the air sparge piping, to collect dislodged cobalt pucks.

Cobalt Puck Removal

The stripping unit removes the deposited cobalt pucks from the cathode blank using a vibrating jig. The tankhouse crane places the cathode load into the jig. Pneumatic actuated hammers strike the suspension bars and the transmitted vibration dislodges the deposited pucks, which fall onto a feeder, followed by a conveyor that transfers them to a “vibrofinishing” tank. A grid under the vibrating jig separates any pucks that may have clumped together. Burrs are removed from the pucks as swarf in this tank by washing with hot water.

The pucks are fed to a heated screen, where oversize and fines are separated from on-spec product. Oversize and fines products are fed directly into drums. Mini drums are placed on a platform scale, automatically filled to 250 kg with product, and then adequately sealed for shipping.

Product Storage and Shipping

Filled drums are transferred by forklift to the packaging area in the service complex building. They are packaged in 1 t lots and banded or shrink-wrapped onto shipping pallets.

Cathode Blank Pre-treatment and Refurbishment

Periodically, the cathode blanks will require refurbishment. It is expected this will be done every 4 to 5 weeks. The remaining epoxy coating is removed using ultra-high pressure water jet cutting technology. Residual water from the epoxy-free cathode blanks is evaporated in an electrically powered dryer. The dried cathode blanks then are sent to a grit blasting station, where the entire active surface of the cathode blank is blasted with 60 to 80 grit abrasive blast. After the grit blast, dielectric epoxy coating is applied to both sides of the cathode blanks. Cathode blanks then are sent to a curing oven.

Anode Blank Refurbishment

Anode blanks also will require periodic refurbishment. The entire process is performed manually. The operator uses a jib crane to pick up a spent anode blank, and sprays the blank with water to remove any residual electrolyte. Cutting open the anode bag releases anolyte or residue. The blank then is hooked onto a tilt table where it is disassembled. The old anode bag is removed and disposed of appropriately. After rinsing, the anode is reassembled with a fresh bag. The bagged anode then is moved to a test tank and tested for leaks before being returned to the electrowinning cells.

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17.3.8 Cobalt Sulphate Heptahydrate Option

Cobalt pregnant solution that has been processed by acid pressure leach in the autoclave and after the iron and arsenic removal and copper precipitation step is processed by SX using Cyanex 272 to remove metal impurities through sequential stripping, leaving a pure cobalt sulphate solution. This solution is then evaporated and subjected to a three-stage crystallization process to produce a 20.9% cobalt sulphate heptahydrate product, which is then filtered and loaded into drums.

17.3.9 Cobalt Residue Cyanidation and Merrill Crowe

Cobalt Residue Cyanidation

Cobalt autoclave discharge residue, having been repulped in the cobalt residue cyanidation feed tank, is pumped to the first of six covered agitated cobalt residue cyanidation tanks arranged in series. Sodium cyanide and lime are added continuously to the tank to reach 1 g/L of cyanide in the solution and a pH of approximately 10.5, with a pulp density of 45% solids. Oxygen gas and / or air also are fed to the tank continuously for the cyanidation reaction (Elsner‟s equation).

The discharge from the final cyanidation tank overflows to the cobalt cyanidation discharge pumpbox, where 2.5 g/m³ of diatomaceous earth bodyfeed is mixed. This slurry then is pumped to a vacuum belt filter. A 1.5 times displacement countercurrent wash, first with gold barren solution, followed by process water, is applied to the filter to displace entrained pregnant liquor in the filter cake. The final rinse of the cake with process water ensures that most of the cyanide solution is washed from the cake. The washed filter cake discharges onto the cobalt residue discharge conveyor, which transfers the tailings to the tails residue storage bin along with the cobalt residue cyanidation tails. The gold pregnant solution from cyanidation is pumped to the Merrill Crowe circuit.

Merrill Crowe Circuit

Gold pregnant solution from the autoclave residue cyanidation circuit is pumped to the Merrill Crowe circuit for gold recovery.

The gold pregnant solution is passed through a clarifying filter to remove any suspended solids from the solution. Diatomaceous earth precoat is applied to the filter. The filtered solution flows through a deaeration column to reduce the oxygen content in the solution to less than 0.5 ppm. The filtered and deaerated pregnant solution then is pumped to the first of two conical bottomed precipitation tanks in series. Zinc dust and lead nitrate are added via screw feeders to the first precipitation tank. Zinc addition is based on a zinc-to-gold molar ratio of 30:1. Lead nitrate is added as 10 wt% of the zinc addition. Gold is precipitated from the cyanide solution via a cementation reaction on the surface of the metallic zinc particles. If present, silver and copper also will cement out.

Bodyfeed is added at 2.5 g/m³ of precipitated slurry. A centrifugal pump sends the slurry to a recessed plate type filter. Precoat also can be added to the filter. The filtrate, depleted of gold, is collected in the gold barren solution tank and distributed to various locations within the cyanidation circuit, mainly for cake repulp or as filter wash solution. Any excess of this gold barren solution is sent to cyanide destruction before discharge to the environment. Gold precipitates from the filter are sent to the drying area of the gold refinery. P&E Mining Consultants Inc., Report No. 247 Page 175 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Gold Refining

The gold refinery receives filter cake from the Merrill Crowe circuit. This filter cake is dried in an oven, mixed with weighed amounts of flux and placed into an induction crucible furnace. Molten gold is poured into moulds to form gold bars. Oven and furnace off-gases are processed through a wet scrubber before discharging into the atmosphere. The slag generated from the furnace is collected.

17.3.10 Tailings Handling and Cyanide Destruction

Tailings Handling

Cobalt Fe-As filter cake and cobalt cyanidation filter cake are combined on the cobalt residue discharge conveyor, and conveyed to the tails residue storage bin. The storage bin has a capacity of 20 m³, or approximately 2 hours of tails generation. The combined cake is discharged from the bin via a rotary valve onto a truck, to be hauled to a residue storage area.

Cobalt precipitation thickener overflow, excess cobalt barren solution, intermittent surface sumpage and other collected solutions from the EW building and tails residue areas are collected and pumped to outdoor ponds. The cobalt precipitation thickener overflow is pumped to a heat exchanger to recover heat before it is sent to the process water pond.

Three ponds collect process solutions. The process water pond is the largest, with a capacity of approximately 5,000 m³.

The cobalt solution pond has a capacity of 300 m³ and acts as emergency storage for EW tank house contents. The brine solution pond, with a capacity of 125 m³, stores injection well water during injection well maintenance. In case of emergency when the contents of the cyanidation circuits need to be dumped and stored, these streams are pumped to the brine solution pond.

Cooled water is pumped from the process water pond to the water treatment system. This system is comprised of media filters and cartridge filters for removal of suspended solids, an antiscalant system for injecting scale inhibitor, a chemical injection system for injecting hydrochloric acid for pH control, and one RO train. The permeate from the RO train is pumped directly back to the hydromet process water tank. The rejected brine is treated further in the effluent treatment tanks.

Sodium hydroxide is used to precipitate any soluble metals in the RO reject / brine in two effluent treatment tanks in series. Tank discharge is pumped to the sediments thickener via a feed launder where flocculant is added. Thickener overflow is collected in an overflow tank where it combines with the cyanide destruction tank discharge, boiler area sump discharge and brine from the bismuth CLER circuit, and is pumped to the underground saline aquifer.

Cyanide Destruction

Excess gold barren solution containing cyanide requires treatment before being discharged. The solution is pumped to the agitated cyanide destruction tank, which has a residence time of 3 hours. Hydrogen peroxide is added as an oxidizing agent. Cyanide is destroyed and is reduced to a total cyanide target of 1 ppm. Discharge from cyanide destruction is pumped to the sediments thickener overflow tank. P&E Mining Consultants Inc., Report No. 247 Page 176 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

17.4 SMPP PLANT UTILITIES

Plant and Instrument Air

Seven air compressor packages in total are provided in the main building, with five operating and two stand-by. They provide compressed plant air, instrumentation air, pressure filter air and pressure filter membrane air. Each plant and instrumentation air compressor package has a capacity of 1,000 m³/h (at 20°C and 101.3 kPa) and a delivery pressure of 790 kPa (absolute). Each of the two pressure filter air compressors has a capacity of 6,050 m³/h (at 20°C and 101.3 kPa), while the pressure filter make-up air compressor has a capacity of 1,250 m³/h (at 20°C and 101.3 kPa). Both the pressure filter air and pressure filter make-up air compressors have a delivery pressure of 790 kPa (absolute). The pressure filter membrane air packages have a capacity of 340 m³/h (at 20°C and 101.3 kPa) and a delivery pressure of 1,500 kPa (absolute).

Flotation Air

Two flotation air blowers, one operating and one stand-by, also are provided near the flotation cells. Each flotation air blower has a capacity of 5,400 m³/h (at 20°C and 101.3°kPa) and a delivery pressure of 10 kPa (gauge).

Fuels

Natural gas is supplied to the process plant via pipeline. Natural gas is used mainly in the plant for steam and heat generation. A regulation / metering station will be supplied by the utility company. Each piece of equipment in the plant will have the appropriate safety devices for regulation and / or metering.

17.4.1 Water Supply

Fresh, Fire and Potable Water

The groundwater well system supplies fresh water for the process plant.

The fresh water tank is located indoors in the main building. It has a capacity of 760 m³, equivalent to 3.5 hours of usage.

A monorail hoist is provided for pump maintenance in the water area.

Flotation Process Water

Hydrometallurgical Process Water

Hydrometallurgical process water is collected from agitator seal cooling water return and treated RO water, with 21.3 m³/h of fresh water make-up. The treated RO water flow rate is 31 m³/h. The hydromet process water tank has a capacity of 200 m³, equivalent to 3.75 hours usage. P&E Mining Consultants Inc., Report No. 247 Page 177 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Hose and Gland Water Distribution

Hose water is distributed from the fresh water tank and the flotation process water tank to individual hose stations in the hydrometallurgical plant and in the regrind and flotation circuits, respectively.

Demineralized Water and Steam Boiler

The demineralised water treatment plant produces 10 m³/h of demineralized water to be pumped to boilers and various other users in the plant.

There are two steam boilers in the SMPP plant. The auxiliary equipment for the boilers includes a condensate storage tank, deaerator feed pumps, deaerator, boiler feed pumps, combustion air fans, blowdown tank and a stack. The boiler generates steam at 1,000 kPa and 180°C. A pressure reducing station reduces the pressure to 220 kPa, generating superheated steam. This steam is used for transferring heat to various solutions in heat exchangers, maintaining the temperature of various process solutions in tanks, as well as for start-up of the autoclave. Boiler blow down is collected in the sump and is pumped to the sediments thickener overflow tank in the effluent treatment area.

17.4.2 Process Ancillaries and Services

Sewage Holding System

The sewage storage system consists of a septic tank to store domestic waste, installed to the west of the service building. It has a capacity of 12.8 m³, equivalent to the storage volume required for 30 shift and 25 day employees over a three-day period. The waste is collected by vacuum trucks on a regular basis.

Materials, Consumables and Reagents

The reagent storage area stores all the dry bagged reagents. Reagent consumptions in this section are based on the average head grade.

Oxygen

The oxygen plant is installed to the west of the process plant. The VPSA oxygen plant has a capacity of 120 t/d of oxygen. The system requires a source of very clean air and water; instrumentation air and fresh water will be provided.

The oxygen produced, at a minimum purity of 93%, is used in the cobalt autoclave for pressure oxidation, in cobalt Fe-As precipitation, and in the autoclave residue cyanidation circuits.

Argon

Argon is provided as a blanket in the final stages of bismuth production to prevent oxidation. An area for bottle storage is allocated, however it may be beneficial to generate the argon in the oxygen plant.

P&E Mining Consultants Inc., Report No. 247 Page 178 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Nitrogen

Nitrogen is provided as a blanket in the final stages of bismuth production to prevent oxidation. An area for bottle storage is allocated, however it may be beneficial to generate the nitrogen in the oxygen plant.

Iron Powder

Iron powder is used to precipitate copper from the solution in the copper cementation process. It is delivered by trucks or rail cars in 1 t bulk bags. The iron powder bag is moved by a hoist and monorail system and off-loaded in to a storage bin. The bin is equipped with a bin vent, bag breaker and a rotary feeder which discharges the solid powder to a screw conveyor. The screw conveyor transfers the iron powder to the copper cementation tank.

Lime Slaking System

The lime slaking plant supplies milk of lime for pH control at various points, including Fe-As precipitation, autoclave residue cyanidation and the bismuth CLER circuit, by means of a lime loop distribution system. The package consists of a lime silo, a pneumatic conveyor, a lime slaker mill, a wet scrubber, a cyclone and associated equipment.

Cobalt barren solution from Stage 2 cobalt precipitation is used for lime slaking. Slaking mill discharge collects in a pumpbox and is pumped to a cyclone. Cyclone underflow returns to the mill by a recycle pump. Cyclone overflow, at 15% lime concentration, flows by gravity to the agitated milk of lime storage tank. The tank provides 8 hours of storage capacity. Milk of lime is pumped continuously from this tank to the lime loop for distribution to various users. Excess milk of lime returns via the loop to the storage tank.

Cyanide Solution Preparation

Cyanide is used as a cobaltite / arsenopyrite depressant in the bismuth flotation stage, and in the autoclave residue cyanidation steps.

Cyanide is delivered as a powder in 1 t bags. For each mixing batch, one bag is emptied into an agitated cyanide mixing tank. Fresh water and a small dose of sodium hydroxide are added to the tank to prepare a batch of 10 wt% cyanide solution. Sodium hydroxide maintains the alkaline conditions necessary to prevent the formation of toxic hydrogen cyanide gas. Cyanide solution is distributed from the tank to various users in the process plant. One batch of cyanide solution is sufficient to satisfy the plant‟s daily requirement.

The mixing tank is sealed. Dust and fumes in this area are removed with a reagent exhaust fan to a reagent scrubber. Sodium hydroxide solution is the scrubbing agent.

Sulphuric Acid

Sulphuric acid, at 98% concentration, is used in copper releach, zinc and nickel ion exchange, cobalt dissolution, bismuth leach and cobalt pressure leaching during start-up to initiate the reaction.

P&E Mining Consultants Inc., Report No. 247 Page 179 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Sodium Metabisulphite

If cobalt electrowinning encounters operational issues, cobaltic sludge might be formed and accumulate at the bottom of the EW cells. The sludge would need to be removed from the cell and releached into solution. Sodium metabisulphite is a contingency reagent that is used only if sludge releaching is required. There is no requirement in normal operations.

Lignosol

Lignin sulphonate (Lignosol) is a surfactant added regularly to the cobalt slurry in the cobalt autoclave surge tank before the slurry is fed to the autoclave. Lignosol consumption is based on an addition rate of 6 kg of lignosol per tonne of autoclave feed specified for laboratory testwork. It is expected that consumption will be reduced greatly during operation, as is found commonly in other industrial POX circuits.

Sodium Hydroxide

Sodium hydroxide is used for pH adjustment and hydroxide precipitation is used in the second stage of cobalt precipitation, nickel ion exchange, effluent treatment, cyanide preparation and the bismuth CLER circuit.

Sodium Carbonate

Sodium carbonate is used in copper precipitation, the first stage of cobalt precipitation, cobalt electrowinning and in zinc and nickel precipitation.

Sodium carbonate is delivered in bulk by truck or railcar and off-loaded into a storage bin. The bin is equipped with a bin vent and a rotary feeder, which discharges the solids into an agitated mixing tank. Fresh water is added to the tank to make a 15 wt% sodium carbonate solution. The solution is transferred to a day tank, from where it is pumped to various users as required.

Flotation Reagents

Reagents used in flotation are PAX and MIBC. PAX is the collector in the roughing and scavenging flotation stages, while MIBC is the frother. The flotation reagent preparation area has a monorail hoist for pump maintenance and lifting of PAX bags and MIBC drums, as required.

Flocculant

Flocculant is used in various locations to aid solid-liquid separation in thickeners and clarifiers.

Hydrogen Peroxide

Small amounts of hydrogen peroxide are added to the last cobalt Fe-As precipitation tank as an additional dose of oxidant in a polishing step to remove iron and arsenic from solution. Peroxide also is added to the cyanide destruction tanks. The consumption of hydrogen peroxide in cobalt Fe-As precipitation is based on pilot plant testwork usage. Depending on the effectiveness of oxidation using enriched air in actual operations, this usage will be reduced.

P&E Mining Consultants Inc., Report No. 247 Page 180 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Manganese Sulphate

Manganese sulphate (MnSO4) at 32% strength is used in the cobalt electrowinning circuit to prevent corrosion of the lead anodes. Fresh water is added to an agitated mixing tank in the cobalt electrowinning area for the slurrying and preparation of MnSO4.

Hydrochloric Acid

Hydrochloric acid is used in the RO unit of the water treatment plant. HCl at 36 wt% strength is delivered and stored on-site in drums. A drum pump transfers the acid to the RO unit.

Lead Nitrate

Lead nitrate is added in the Merrill Crowe circuit to create preferential reduction sites on the zinc particles. Powdered lead nitrate is delivered and stored on-site in drums. The drums are emptied into a bin and screw-fed into the first of two Merrill Crowe precipitation tanks.

Zinc Dust

Zinc dust is added in the Merrill Crowe circuit to precipitate gold via a cementation reaction on the metallic zinc particles. Zinc dust is delivered and stored on-site in drums. The drums are emptied into a bin and screw fed into the first of two Merrill Crowe precipitation tanks in series.

Diatomaceous Earth (Bodyfeed and Precoat)

Diatomaceous earth is used in various locations of the process plant to aid solid-liquid separations in the filters.

Diatomaceous earth, in 25 kg bags, is added to two agitated mixing tanks. Gold barren solution is used to prepare a 30 wt% solution, which is pumped, to the various users.

Ammonium Hydroxide

Ammonium hydroxide is used in the nickel IX circuit to strip / clean the nickel IX resin, should the copper column fail to remove all the copper from the solution and copper loads onto the nickel resin.

Ammonium hydroxide at 25 wt% strength is delivered and stored on-site in drums. A drum pump transfers the solution to a tank, before it is pumped to the nickel IX system.

Annual consumption of ammonium hydroxide is estimated to be 5 t/a.

Guar Gum

Guar gum is added to the electrolyte solution as a deposit smoothing and levelling agent. Provision could be made to allow for the addition of guar gum to the cobalt electrowinning tankhouse, if required.

Guar gum, in 25 kg bags, can be added to an agitated mixing tank. Process water is used to prepare a 10 wt% solution. P&E Mining Consultants Inc., Report No. 247 Page 181 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Sodium Lauryl Sulphate (SLS)

Provision is made to allow for the addition of Sodium lauryl sulphate (SLS) to the cobalt electrolyte circulating tank. SLS is used in the cobalt electrowinning to minimize mist generation. SLS in 25 kg bags is added to an agitated mixing tank where the strong electrolyte solution is used to prepare the solution.

Gold Fluxes

Fluxes used in gold refining are silica, borax, sodium nitrate and sodium carbonate. These fluxes are used in the induction crucible furnace to purify the gold, as flux bonds with contaminants and floats on top of the melted gold. Since the gold casting will be a mixture of gold sands from both the EW circuit and the Merrill Crowe circuit, it is difficult to predict actual flux requirements before operation.

Silica, borax and sodium nitrate are delivered in 25 kg bags, and sodium carbonate is drawn from the sodium carbonate bin. The total flux required is ratioed to the gold production in a 1:1 mass ratio for cost estimating purposes.

Sodium Chloride

Sodium chloride is used in the bismuth CLER circuit. Sodium chloride will come in tote bags as solid crystals with a purity of 97 to 99%.

17.5 EXECUTION PLAN

17.5.1 NICO Concentrator Overview

The NICO Concentrator Project Execution Plan aims to develop the NICO site from its present conditions to an operational state by implementing all facilities to produce concentrates for shipment to the metallurgical process plant.

The construction strategy is based on the condition that an all-weather road (“AWR”) will be available at the time of starting main site construction activities. This approach will reduce constrain related to site access and the logistics to support the construction work will be less complex by having all year around access.

The Project Execution Plan is outlined in six phases that encompasses all activities from construction of the AWR through start-up of the plant. This schedule presents a total of 24 months for implementing the NICO Project. It is expected that a 12-month period is required to complete the AWR and a period of 12 months for completing the main construction work and plant commissioning and commence start up. This schedule considers implementation of early activities before completion of the AWR by using a winter road and / or seasonal road during the winter of the previous year. The main camp should also be serviceable at the time that AWR is completed.

Construction of the NICO facility will require a full detailed implementation plan for incorporating engineering, construction and logistic activities that are required for a remote site location development. The present execution plan considers starting the engineering and P&E Mining Consultants Inc., Report No. 247 Page 182 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. procurement (E/P) phase at least 9 months before the completion of the AWR. In consideration that implementing this phase will help to firm up the construction plan, it will be desirable to initiate this phase as early as possible, and holding equipment purchase orders until a decision to proceed is obtained.

Optimization of the Project Execution Plan will be possible while completing engineering work considering the site location, early work implementation, equipment fabrication and modularization to minimize field construction and the appropriate logistics to support field activities.

17.5.2 NICO Construction Schedule Presentation

The NICO construction schedule is summarized in Figure 17.1. The overall schedule considers six phases for implementation of the NICO plant:

 Phase 1: All-Weather Road completion, estimated at 15 months  Phase 2: Detail Engineering / Procurement, estimated at 12 months. Manufacturing and delivering of the equipment could take up to 15 months. This phase could take between 18 to 21 months to complete.  Phase 3: Exploration / Early Work Program, requiring 9 months.  Phase 4: Pioneering Construction, requiring 5 to 7 months.  Phase 5: Main Construction Work, requiring 17 months for completion which includes 5 months for site preparation before the completion of the AWR and 12 months for the Process Plant work.  Phase 6: Plant Commissioning and Start Up, requiring 3 months.

P&E Mining Consultants Inc., Report No. 247 Page 183 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Figure 17.1 Project Schedule NICO Concentrator

Year -2 -1 1 2 3 Month -15 -14 -13 -12 -11 -10 -9 -8 -7 -6 -5 -4 -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 Q4 Q1 Q2 Q3 Q4 Q1 Q2 Q3 Q4 Q1 Q2 Q3 Q4 Q1 Q2 Q3 Duration No. Items (months) 10 11 12 1 2 3 4 5 6 7 8 9 10 11 12 1 2 3 4 5 6 7 8 9 10 11 12 1 2 3 4 5 6 7 8 9 10 11 12 1 2 3 4 5 6 7 8 9 No.

MILESTONES

M1 ♦ PERMIT FOR AWR M1

M2 ♦ START DETAIL ENGINEERING START DET ENGINEERING M2

M3 LAND USE PERMITS/WATER LICENCES LAND USE PERMITS/WATER LICENCES M3

M4 ♦ COMPLETION OF AWR AWR COMPLETION M4

M5 ♦ PLANT START UP PLANT START UP M5

PH1 FEED REPORT 1 TECHNICAL REPORT CONCENTRATOR 1 Report delivery 1 Dec-11 2 FEED 2 Updated Feed 3 Feb-12

I PHASE 1 AWR CONSTRUCTION I

3 Winter Road Winter Road 3

4 SOR SOR 4

5 Spur road to site TBC Spur road to site 5

6 AWR CONSTRUCTION TBC 1 AWR PHASE 6 CONSTRUCTION AWR CONSTRUCTION

II Ph 2 -Det Eng/Procurement

7 Detailed Engineering - Concentrator 12 Detailed Engineering - Concentrator 7

8 Design Pkg's Long Lead Equipment 4 Design Pkg's Long Lead Equipment 8

9 Procure Long Lead Equipment 8 Procure Long Lead Equipment 9

10 Manufacture & Deliver - Long Lead Equipment 12 Manufacture & Deliver - Long Lead Equipment 10

11 Procure Remaining Equip 5 Procure Remaining Equip 11

12 Manufacture & Deliver - Remaining Equipment Manufacture & Deliver - Remaining Equipment 12

13 Camp Sel/Proc 2 Camp Sel/Proc 13

14 Trade Contracts 6 2 -Det Ph Eng/Procurement Trade Contracts 14

III Ph 3 - Exploration/ Early Work Program III

15 Exploration Supplies 3 Exploration Supplies 15

16 Exploration Support Program 6 Exploration Support Program 16

17 Expand Pioneer Camp 3 Expand Pioneer Camp 17 Program

18 Site Preparation:Roads, Camp;Truck shop 6 Work Early Site Preparation:Roads, Camp;Truck shop 18 Ph 3 3 -Ph Exploration/ IV Ph 4- Pionering Construction

19 Pionering Camp Ready 0 Pionering Camp Ready 19

20 Mobilization to Staging Point 3 Mobilization to Staging Point 20

21 Main Camp Construction 3 Main Camp Construction 21

22 Batch plant - Foundation-Piling 3 Batch plant - Foundation-Piling 22

Construction Ph 4- Ph Pionering 23 Const. Diesel Gensets 3 Const. Diesel Gensets 23

V Ph 5 - Construction

24 Main Camp Available 0 Main Camp Available 24

25 Excavate Foundation: Stockpile -truck shop plant 3 Excavate Foundation: Stockpile -truck shop plant 25

26 Foundation: Truck shops and plant 5 Foundation: Truck shops and plant 26

27 Def. Power Gensets/Power plant 5 Def. Power Gensets/Power plant 27

28 Construction - Plant/Infrastructure 12 Construction - Plant/Infrastructure 28

29 Foundations 5 Foundations 29

30 Bldg Envelopes 6 Bldg Envelopes 30

31 Structural Steel, tanks, Platework 5 Structural Steel, tanks, Platework 31

32 Major Equipment Placement 5 Major Equipment Placement 32 Ph 5 5 -Ph Construction 33 Piping/Mechanical 8 Piping/Mechanical 33

34 Electrical 7 Electrical 34

35 Instrumentation 7 Instrumentation 35

36 Punchlist Completion 3 Punchlist Completion 36

37 Punchlist Completion 3 Pre-Commissioning 37

VI Commissioning and Start Up

38 Commissioning 2 Commissioning 38

39 Plant Start-up... 2 Plant Start-up... 39

40 Ship Product to SMPP... Ship Product to SMPP... 40

Ph 6 6 -Ph Commission

P&E Mining Consultants Inc., Report No. 247 Page 184 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

17.5.3 Saskatchewan Metallurgical Processing Plant (SMPP) Overview

The SMPP Project Execution Plan aims to develop the SMPP site near Langham, Saskatchewan by implementing all facilities to treat concentrates from the NICO Concentrator and produce valuable metals and metal compounds for sale.

The construction strategy considers the completion of an all-weather road to NICO mine and concentrator as the baseline to start main construction work.

The Project Execution Plan is outlined in four phases to complete all activities from start of the Detail Engineering through start-up of the plant. This schedule presents a total of 32 months for implementing the SMPP Project. It is expected that a 12-month period is required to complete the EP Contract and a period of 20 months for completing the main construction work and reach plant start-up. This schedule considers implementation of early activities before completion of AWR by implementing an early work phase for site development.

Construction of the SMPP facility will require a full detailed implementation plan for incorporating engineering, construction and logistic activities that are required for Northern site location development. The present execution plan considers starting the engineering and procurement (E/P) phase at least 12 months before the completion of the AWR. In consideration that implementing this phase will help to firm up the construction plan, it will be desirable to initiate this phase as early as possible, and holding equipment purchase orders until a decision to proceed is obtained.

17.5.4 SMPP Construction Schedule Presentation

The SMPP construction schedule is summarized in Figure 17.2. The overall schedule considers four phases for implementation of the NICO plant:

 Baseline: All-Weather Road completion, estimated at 15 months.  Phase 1: Detail Engineering / Procurement, estimated at 15 months. Manufacturing and delivering of the equipment could take up to 15 months. This phase could take between 15 to 18 months to complete.  Phase 2: Early Work Program, requiring 6 months.  Phase 3: Main Plant Construction Work, requiring 17 months for completion which includes 5 months for site preparation and 12 months for the process plant.  Phase 4: Plant Commissioning and start up, requiring 5 months which includes 2 months for start-up after receiving concentrate.

P&E Mining Consultants Inc., Report No. 247 Page 185 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Figure 17.2 Project Schedule SMPP

2011 2012 -2 -1 0 1 2 3 -21 -20 -19 -18 -17 -16 -15 -14 -13 -12 -11 -10 -9 -8 -7 -6 -5 -4 -3 -2 -1 0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 Q4 Q1 Q1 Q2 Q3 Q4 Q1 Q2 Q3 Q4 Q1 Q2 Q2 No. No. 10 11 12 1 2 3 1 2 3 4 5 6 7 8 9 10 11 12 1 2 3 4 5 6 7 8 9 10 11 12 1 2 3 4 5 6 7 8 9

MILESTONES

M1 ♦ START AWR CONSTRUCTION M1

M2 ♦ START DETAIL ENGINEERING START DET ENGINEERING M2

M3 LAND USE PERMITS/WATER LICENCES LAND USE PERMITS/WATER LICENCES M3

M4 ♦ COMPLETION OF AWR AWR COMPLETION M4

M% PLANT START UP PLANT START UP

FEED UPDATE

Technical Report

Technical Report Update Dec-11

FEED UPDATE

Update completion Feb-12 FEED UPDATE FEED

I AWR CONSTRUCTION

1 AWR CONSTRUCTION 1 AWR AWR

AWR CONST. COMPLETION CONSTRUCTIO Ph 1 -DET ENG./PROCUREMENT

2 EP Contract Award 0 EP Contract Award 2

3 Detailed Engineering - SMPP 12 Detailed Engineering - SMPP 3

4 Design Pkg's Long Lead Equipment 5 Design Pkg's Long Lead Equipment 4

5 Design Pkg's Other Equipment 6 Design Pkg's Other Equipment 5

6 Procure Long Lead Equipment & Materials 9 Procure Long Lead Equipment & Materials 6

7 Procure Other Equipment & Materials 7 Procure Other Equipment & Materials 7

8 Manufacture & Deliver - Long Lead Equip 12 Manufacture & Deliver - Long Lead Equip 8

9 Mfr. & Del. Additional Equip 6 Mfr. & Del. Additional Equip 9

10 Manufacture & Deliver - Bulk Materials 8 Manufacture & Deliver - Bulk Materials 10

11 CMPC Selection - SMPP 0 CMPC Selection - SMPP 11 Ph 1 -DET ENG./PROCUREMENT ENG./PROCUREMENT 1 -DET Ph 12 Trade Contracts 6 Trade Contracts 12

IV Ph 2 EARLY WORK

13 Land Use Permits/ Water Licences 0 Land Use Permits/ Water Licences 13

14 Mobilize to Site 4 Mobilize to Site 14

15 Site Prep/ Roads 4 Site Prep/ Roads 15

16 Off site work 9 16

Power Line Ph Ph 2WORK EARLY Ph 3 -Construction

17 Residue Storage Facility 5 Residue Storage Facility 17

18 Process Water Pond 3 Process Water Pond 18

19 Foundations 5 Foundations 19

20 Structural Steel, tanks, Platework 6 Structural Steel, tanks, Platework 20

21 Bldg Envelopes 6 Bldg Envelopes 21

22 Major Equipment Placement 5 Major Equipment Placement 22

23 Piping 10 Piping 23

24 Electrical 9 3 -Construction Ph Electrical 24

25 Instrumentation 10 Instrumentation 25

26 Trade Completion Testing 3 Trade Completion Testing 26

VI Commissioning and Start Up

27 Commissioning 3 Commissioning 27

28 Receive Product from NICO 0 Receive Product from NICO 28

29 Start-up 2 Start-up 29

Products to sale

Ph 4 4 -Ph Commissioning

P&E Mining Consultants Inc., Report No. 247 Page 186 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

18.0 PROJECT INFRASTRUCTURE

18.1 NICO MINE SITE AND CONCENTRATOR

The location of the various facilities and infrastructure was established with consideration given to the following site-specific objectives:

 Provide safe working facility  Minimize environmental impact  Minimize earthworks  Minimize mining haul distance  Minimize distance from accommodation facility to work areas  Utilize heat recovery from power plants to heat the process areas  Provide aesthetically pleasing facility for accommodation and dining  Design a compact plant for optimizing heating requirements.

The Project has been developed to meet safety, environmental and operations objectives for the facility using design criteria implemented by the engineering team.

18.1.1 Access Road

The planned completion of the all-weather road, serving the Whatì and Gamètì communities and the NICO spur will coincide with the start of construction. The road, referred to in Item 5, consists of the GNWT portion from Behchokö to the La Marte Crossing, and an additional segment to the junction with the NICO spur. The spur and segment from La Marte is 51 km across various terrains with a crossing of the Marian River.

Access to the NICO site is from the southwest off the proposed alignment of the Whatì and Gamati roads as shown in Figure 18.2. The route will encounter a river crossing at the Marian River, which is within 1.5 km of the mine site boundary. A steel bridge (Bailey-type construction) capable of supporting a 100-t load is proposed. The bridge width will accommodate 6 m and will have unobstructed height.

The road from the Marian River and through to the mine and process plant is planned for 6-m width. The road passes to the north of the Burke Lake and then turns due north to the site facilities.

The access road passes through swampy / low areas initially which will require drainage / culverts and rock fill. The area also transverses across rock outcrop where blasting will be required for gradient and drainage control.

The final road system around the site will be designed and constructed to minimize mining activities from the initial construction site and, ultimately, the operating facilities. Refer to Figure 18.2, to view the site road plan. Additionally, waste haulage to disposal and to areas of tails embankment construction will be isolated from other site road traffic flow to the maximum extent possible.

P&E Mining Consultants Inc., Report No. 247 Page 187 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

18.1.2 Access Road Geotechnical

The proposed access road to the NICO mine site begins approximately 24 km northeast of the community of Whati, NT. The route to get to the beginning of the currently proposed access road is along a former a winter road to Gameti (formerly Rae Lakes), NT. This route is part of the proposed all weather access road from Rae-Edzo to Gameti. From this starting point, the proposed access road route to NICO is approximately 32 km long, and crosses the Marian River about 3 km south of the NICO mine site.

EBA reviewed available information in 2004 and selected the initial road route. In July 2004, EBA conducted an investigation along the proposed route (EBA file: 1700127.001). This investigation involved walking the route to classify terrain types, and conducting manual test pitting to obtain soil samples for the purpose of subgrade classification and the identification of potential borrow sources for road embankment material. Five terrain classifications were made.

This included sections where ice-rich permafrost is likely present and subgrade conditions are poor, and others where bedrock is shallower and subgrade conditions are good. Based on the conditions encountered, conceptual road embankment structure designs were developed. Locations where major culverts would be required for drainage were also identified, as well as the location of the Marian River crossing, which will require a bridge.

Seven potential borrow pits were identified and delineated. The borrow sources contain granular materials that are expected to be suitable for use as general embankment fill. Further investigation would be required to estimate suitable material quantities within these sources, but it is expected that these sources will provide a sufficient quantity of general fill to construct the access road.

In 2007, EBA conducted preliminary foundation design of the bridge crossing (EBA file: 1700127.005). A site with outcropping bedrock at both abutments was selected and surveyed. The site hydrology assessment was conducted by Golder Associates Ltd. The design proposes a 25 m (80‟) pre-engineered steel girder bridge. The proposed clear width is 6 m, which is effectively a single-lane. The type of deck has not been selected. While the bridge could be supported on concrete abutments, timber crib and granular fill abutments were proposed to limit concrete volumes at the remote site.

EBA began work in 2009 to refine the conceptual design and improve material quantity estimates. This is documented in a report submitted in April 2011 (EBA file: Y14101101.005). A helicopter fly-over of the route was conducted by EBA. LiDAR data, and satellite imagery along the route was provided. The LiDAR data provided EBA with a detailed topography along the route.

The 2011 preliminary design was based on assumed traffic volumes of 3 fuel and supply loads, and 5 loads of concentrate per day. Similar to the 2005 conceptual design, five terrain classifications were made, with a road embankment structure design for each. The road route selected was almost the same as in 2005, but the LiDAR data allowed EBA to adjust the route to avoid low spots and areas with steep changes in elevation. In addition, the details of the access to the mine site have changed.

Additional borrow sources for granular material were identified in areas where roadway cuts may be possible. The borrow sources identified were located north of the Marian River crossing. P&E Mining Consultants Inc., Report No. 247 Page 188 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Material quantities available within any of the borrow sources identified cannot be accurately estimated without further investigation of the sources.

Drainage paths where culverts will be required were identified more accurately in the 2011 design. A 2007 report by Golder Associates Ltd. provided an assessment of Fish and Fish Habitat of watercourses along the proposed road route. The report indicated that there were five water crossings within the road route. Four of the five crossings will require large culverts and the Marian River crossing will require a bridge. The report indicated that the Marian River supports fish habitat. Approximately 65 culverts will be required in total.

The preliminary road design is conservative, and assumes the road will be built primarily by embankment construction with very little roadway excavation. The road is not designed to maintain the permafrost below. The embankment may need to be thickened in some areas as operation of the road begins and some settlement occurs due to permafrost thaw. Annual maintenance of the road surface will be required as well.

The 2011 report did not include a cost estimate for construction of the road. EBA has updated the cost estimate from 2005, using the material quantity requirements established in the 2011 report and adjusted 2005 unit rate prices for inflation. The updated cost estimate totals to about $15.6M. While much of the design has been done, this estimate includes allowances of 5% for engineering, 15% for construction services and a 15% contingency. The bridge is estimated to cost about $400k., the estimated road construction and engineering cost amounts to about $480k per kilometer, excluding the cost of the bridge.

18.1.3 Plant Site Geotechnical

The NICO plant site is proposed to be east of the open pit. EBA, A Tetra Tech Company conducted a geotechnical investigation of the plant site area (EBA 2011). Thirty one vertical boreholes were drilled at the proposed process plant, accommodation complex, truck shop, and associated site facilities, and three angle boreholes were drilled at the proposed primary crushing station site. Three ground temperature cables (GTCs) were installed.

The subsurface conditions at the proposed process plant, accommodation complex, truck shop, and associated facilities locations comprise a discontinuous veneer of weathered frost-shattered rock material or a thin mantle of organic material underlain by bedrock or a blanket of bouldery till up to 3.5 m thick in bedrock depressions. The underlying bedrock consists of metagreywacke with quartz and feldspar, amphibole porphyry, and rhyolite. The bedrock is predominantly high strength with rock quality ranging from poor, predominantly in the uppermost portion, to excellent. A possible southwest-northeast oriented fault was identified in three boreholes. Permafrost was not identified in the plant site area in the three holes that the ground temperature cables were installed. It is planned that the overburden and the bedrock will be removed to create level surfaces for the plant site facilities. The plant site facilities will be founded on bedrock.

The subsurface conditions of the proposed primary crushing station site, comprise exposed bedrock that is predominantly of rhyolite and feldspar porphyry, with rock quality ranging from very poor to good throughout the borehole depths.

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Figure 18.1 Plant Site Borehole Locations

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18.1.4 Power Plant

The power plant will be located close to the process plant with sufficient access surrounding the equipment with one standby unit installed next to the truck shop.

The estimated power demand for the concentrator will be 10 MW with a running load estimated to be 6.92 MWe and annual energy consumption of 60 591 MWh.

18.1.5 Fuel Depot

Eight envirotanks will be ganged together in a bermed area to serve as the fuel storage facility. A manifold will facilitate distribution of diesel from a supply tank feed to a filling station for the fuel truck.

18.1.6 Process Facility

The process structures are founded on bedrock. They will be constructed with insulated wall panels. The plant is show in Figure 18.3.

The complex will be heated using hydronic units with heated glycol / water supplied from waste heat produced from the site power generators. A backup diesel-fired boiler will also provide makeup heat as required.

18.1.7 Accommodation Complex

The facility will be of modular construction, capable of being installed with minimal site preparation on cribs or piles.

The units will be reconfigurable from double occupancy during construction to single occupancy for operation.

A kitchen and dining area will serve the operation workers in one shift and the construction crew in two.

18.1.8 Corridors / Utilidors

Connecting walkways and utility trays will be in covered utilidors providing services between the process plant, assay lab, truck shop, dining hall and camp facility

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Figure 18.2 NICO Site Layout

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Figure 18.3 Plant Layout

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18.1.9 Water Supply

Fresh water will be extracted from Lou Lake and pumped to the plant water storage tanks through a heat traced and insulated pipe.

A skid-mounted water treatment package will provide potable water.

Site run-off will be collected in contact ponds, and allowed to settle naturally. Overflow will be directed into holding ponds prior to discharge or reclaimed. Excess water flow will be directed into the pit for closure.

18.1.10 Co-Disposal Facility

18.1.10.1 Tailings and Waste Rock Co-Disposal Facility

The mining process will generate a total of 29.9 Mt of tailings and 96.9 Mt of mine rock. Both these waste streams will be disposed together in a facility referred to as CDF. The general arrangement plan of the CDF is shown on Figure 18.4.

Figure 18.4 General Arrangement Plan of NICO Co-Disposal Facility

The rationale for selecting a CDF over a more conventional arrangement (i.e., separate facilities for tailings and mine rock) is discussed in Section 20.

The CDF will be contained on all sides by a Perimeter Dyke comprising a prism of mine rock at least 25 m thick, which will be raised continually in 5 metre lifts using the upstream construction method (Figure 18.5). The Perimeter Dyke will be free draining. To retain tailings particles, the P&E Mining Consultants Inc., Report No. 247 Page 194 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Perimeter Dyke will incorporate a filter comprising either a zone of sand or a non-woven geotextile. The dyke is designed to safely store the Probable maximum precipitation (PMP) rainfall event during operational years and an emergency spillway will be provided for post- closure to safely convey the 1:10,000 year rainfall storm. Additionally, the facility is designed to be stable under a 1:2,500 return period earthquake, at minimum.

Figure 18.5 Typical Cross-Section of the Co-Disposal Facility Perimeter Dyke

Inside the Perimeter Dyke, the CDF will comprise a “layer cake” of alternating layers of mine rock and tailings about 5 m thick. The tailings layers will be created by constructing a series of tailings disposal cells using waste rock, as shown on Figure 18.6. The cell perimeter berms will be constructed by end dumping mine rock. The berm will have a nominal crest width of 6 m to allow vehicle access. Each tailings disposal cell will be filled with non-segregating thickened tailings. The thickened tailings will be discharged through a series of spigot discharge points from the berms of the tailings disposal cells. A 5 m thick layer of waste rock will be pushed over the tailings disposal cells as soon as they are filled to maximize mixing of the waste rock and the tailings. The tailings pipeline system will have a spill collection system.

The perimeter berms of the tailings disposal cells will be permeable. Therefore, the tailings “bleed water” and run-off will seep through them. A good portion of this water will be conveyed to the Reclaim Pond area within the CDF. Since the CDF and the Perimeter Dyke are generally free draining, some of the tailings bleed water and run-off water will seep through the facility and report to five topographically low areas downstream of the CDF. Seepage Collection Ponds (SCPs) will be constructed downstream of the CDF at each of these topographically low areas to intercept the seepage water (Figure 18.7).

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Figure 18.6 Typical Layered Co-disposal Scheme

The SCPs will be generally shallow ponds and they will be contained by low permeability dams. The dams will be constructed of rock fill with internal liner system as shown on Figure 18.7. Rip-rap will be provided on the upstream face of the dam. Transition and bedding layers will be provided above the rock fill to ensure filter compatibility. The liner will be embedded within the bedding layer. Each of the SCP will be protected by an emergency spillway to safely convey storm events.

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Figure 18.7 Typical Cross-Section of Seepage Collection Ponds Dams

There have been three rounds of geotechnical investigation campaigns within the footprint areas of the CDF and associated water management facilities.

18.1.11 Water Treatment

The site sewage system includes rotating biological contactors (RBC) provided as a containerized unit.

Treatment of run-off, pit water and tailings reclaim, all site waters will be contained in a run-off pond and treated prior to discharge to the environment. The seepage ponds collect only water from the Co-Disposal Facility (CDF). The properties of these flows are reported under separate cover by Golder.

18.1.12 Plant Mobile Equipment

The following list (Table 18.1) shows the fleet of mobile equipment to be in service at NICO.

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TABLE 18.1 PLANT MOBILE EQUIPMENT FOR NICO 1 Boat @ NICO 2 Snowmobile 1 Skid Steer Loader (Bobcat S70) 1 Track Dozer (Caterpillar D8T, Komatsu D155AX-6) 1 Potable Pipe Fusion Machine (McElroy Rolling 250, TracStar 250) 1 Surface Grader (Caterpillar 14M, Komatsu GD 675) 1 Telescopic Handler (Caterpillar TL943) 1 Self Propelled Telescopic Boom (Terex Genie) 1 Boom Truck (Western Star 4900 SA) 1 Medium Forklift (Clerk ECX 32) 1 Walkie Straddle Pallet Truck (Clerk CSM 15) 1 Fire Truck 4 Heater - Equipment Thawing 1 Ambulance 2 ATV - Enviromental 4 Light Duty Truck - Services/Maintenance/Security/Spare (Ford F-250) 1 Potable Air Compressor (Compair DLT1303) 2 Potable Generator (Honda 5kW)

18.2 SASKATCHEWAN METALLURGICAL PROCESSING PLANT (SMPP)

18.2.1 Plant Site and Facility

The SMPP site location is in proximity of

 Technically-competent labour pool  Access to transportation network, and  Connectivity to energy utilities.

The vicinity meets the needs of a metallurgical processing facility.

18.2.2 Process Facility

The processing plant is to be erected on a greenfield site with buffer zone, a surrounding berm and perimeter rows of trees to contribute to aesthetics. A residue storage facility is planned for the northern portion of the site. These features are shown on Figure 18.8.

The plant building will be of structural steel with metal roof and insulated wall panels. The facility will be erected with footings and a concrete floor. The current footprint depicts a linear building in which the mechanical equipment is placed by the order of appearance on the flow sheet, attributing to the functionality, material handling characteristics and purity concerns. Figure 18.9 depicts the process building.

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Figure 18.8 SMPP Site Layout

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Figure 18.9 SMPP Plant Layout

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18.2.3 Administrative, Laboratory and Warehouse Building

The two-storey facility will be of structural steel with a prefinished metal roof and insulated wall panels. There will be entrance doors for product pallets carried by propane, powered forklifts and mandoors for personnel. The design will be finished with windows for top floor offices. The administrative building will be the first structure on the approach by road, with the main parking lot being adjacent.

Additional outdoor space will provide warehousing needs in a fenced area with fabric matting for short-term storage and gravel surface for longer periods.

18.2.4 Railway Access

The CN mainline traverses the southwestern portion of the property. A siding will be constructed to receive railcars for concentrate and reagent.

A trackmobile will move the railcars to the process area according to production scheduling and process needs. The siding will be suitable for weekly receiving and storage of bulk concentrate in gondola cars.

18.2.5 Oxygen Plant

A facility will produce the oxygen required for the process. This unit is situated between a bermed area at the west end of the main plant. The oxygen plant is a large power consumer, and will be located close to the incoming electrical feed.

18.2.6 Utilities

Electricity and natural gas are provided by the locally-operating providers, SaskPower and TransGas, respectively.

Sewage will be collected from underground holding tanks to be hauled to the local wastewater treatment facility.

Fresh water will be sourced from three of four wells distributed across the grounds with sufficient open space to permit drill rigs to mobilize for rehabilitation dictated by results of annual inspection testing.

Brine effluent will be deep well injected into the saline aquifer at 800-m depth below surface per MDH report.

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18.2.7 Site Drainage

Ponds are situated to collect:

 Site run-off  Process water  Process residue storage facility (PRSF)  Future water reuse opportunities (area reserved).  Process Residue Storage Facility

Additional relevant information on the process residue storage facility and perimeter drainage is provided in Section 20.3 of this report.

18.2.8 Plant Mobile Equipment

Forklifts are to be used for moving packages of bulk concentrate and product at receiving and shipping, respectively. The units are operated on cyclic schedules, whereby they are run for periods of up to an hour and then will be idle for the next hour. On occasion, they will be called to transport a skid of reagent drums from storage to the production area.

A boom truck and boom will be required for maintenance work around boilers, cell houses and air compressor/blower areas.

Sitework involving haulage of residue and land spreading in the RSF will employ a dump truck and grader. The dump truck will load at the plant and empty at the RSF around the clock. The grader will operate for one 8-h shift per day.

A rubber-tired loader is utilized for general site work and snow removal.

The mobile railcar mover will move groups of railcars from the siding to unloading points, unlatch and return to the siding for other cars. When the car has been emptied, the mover is dispatched to relocate it to a layover siding.

Light vehicles will consist of a fleet of pickup trucks which will be divided by each work division as shown in Table 18.2.

TABLE 18.2 PLANT MOBILE EQUIPMENT FOR SMPP 1 Forklift, 5 tons 2 Small Forklift, 3 tons 1 Boom Truck (Freightliner M2-106) 1 Grader (Caterpillar 120) 1 Dump Truck (Volvo VHD 20t) 4 Ford Truck (Ford F-250) 1 Portable Welding Machine (Lincoln Vantage 500) 1 Portable Generators (Honda EB6500XA) 2 Portable Pump (Honda WT 40XK2A) 1 Manlift - Scissor Lift (Genie GS-3268) 1 Skid Steer Loader (Bobcat, Caterpillar 262C) 1 Trackmobile Railcar Mover (Viking)

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19.0 MARKET STUDIES AND CONTRACTS

19.1 COBALT

19.1.1 Uses

Cobalt is a high strength magnetic metal used to make steel alloys and chemicals. Cobalt has growing consumption in super alloys and cobalt sulphate needed to manufacture high performance lithium-ion and nickel-metal hydride rechargeable batteries used in portable electronic devices and hybrid-electric cars.

Metallic uses include super alloys for the aerospace industry to make power and jet engine turbines, cutting tools and cemented carbides used to machine steel, and electromechanical devices such as magnets, electric motors, generators, transformers and magnetic storage tape and hard disks. The most important factor contributing to the growth in demand for cobalt is its use in chemicals, and particularly for the manufacture of high performance lithium-ion and nickel- metal hydride rechargeable batteries used in portable electronic devices such as cellular telephones and computers and in plug-in and hybrid-electric cars. Cobalt sulphate heptahydrate is preferred by many battery manufacturers for this purpose and sells for an approximate 21 to 22% premium over 99.8% specification high grade cobalt metal cathodes. Cobalt chemicals are also used to make catalysts for petroleum refining and to manufacture plastics. They are also used as pigments and as the source of Vitamin B12.

19.1.2 Supply

The past three years have seen refined cobalt production rates outpacing demand, most noticeably in 2010 when production rose by almost 20,000 MT over 2009 levels. The majority of the production increase was the result of rampant expansion of refining activities in China, where processing of ores/concentrates and intermediate cobalt materials from the Democratic Republic of the Congo (“DRC”) took place on an unprecedented scale. A surplus of supply over demand (refined cobalt) is projected in 2012, although this can be expected to reverse in 2013 as demand levels catch up to and overtake production.

The rate of growth in production slowed significantly in 2011 (China actually decreased its imports of DRC ores/concentrates and its production of refined cobalt products), though still registered a surplus over demand. A contraction of refined cobalt production in China is anticipated for 2012 and carrying into 2013, as refiners there work off the inventory overhang that has appeared. Additional refined cutbacks can be expected at marginal producers and primary producers such as CTT and Kasese.

The majority (75%) of new cobalt production projected over the next 2 – 3 years will originate in the DRC; most of this new DRC cobalt in turn (88%) will be in the form of intermediate cobalt products (hydroxide and carbonate), with the balance being metal (cathode). It is interesting to note that going forward, there are no new projects slated to produce only sulphides (concentrates). The largest producer of refined cobalt in 2011 at 34,969 MT was again China. Chinese processors have been importing concentrates and intermediate forms of cobalt from the DRC and upgrading these to produce cobalt chemicals for lithium ion battery markets.

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Significant cobalt is also mined from nickel-cobalt laterite deposits, which have much higher capital and processing costs relative to sulphide deposits such as NICO, which is also planned to be reliable Canadian-based supplier.

The amount of new cobalt production reported to be coming on line from new projects over the next few years is significant, although it must be noted that most of this is not in refined form, ie: most new cobalt production will be in the form of intermediates (mainly cobalt hydroxides). In the case of cobalt hydroxides from lateritic nickel mines (ie: HPAL processes), the output is far from assured given the problems and delays experienced with recent such projects (New Caledonia, Ambatovy, Ramu).

19.1.3 Demand

The largest market for cobalt is the rechargeable battery segment (lithium ion, nickel metal hydride, and nickel cadmium batteries). In 2011, battery applications represented an estimated 30% (21,600 MT) of total cobalt demand in the form of cobalt chemicals. Cobalt in battery applications has experienced very strong growth from 2000 – 2011, with annual growth rates averaging approximately 20%. Almost half of the net increase in global demand for cobalt since 2000 stems from battery growth.

The other major market for cobalt is super-alloy applications, accounting for 22% (15,840 MT) and hardmetals accounting for 11% (7,920 MT) of cobalt demand. Cobalt demand for superalloys applications is in the form of metal. Super-alloys demand has seen strong growth in 2011, posting an increase of 9% over 2010 levels. The outlook for the near future remains equally buoyant, as commercial aircraft and engine OEMs have strong production schedules out to 2014 and order backlogs. Hardmetals applications of cobalt saw sharp declines in 2009, as automotive, construction, and mining industries all cut demand in the wake of the 2008 economic downturn. A strong recovery in 2010 carried over into 2011, as demand from automotive and industrial markets remained strong, offsetting softer demand from construction markets.

Demand for refined cobalt is expected to continue at healthy levels, the result of a very strong outlook for the battery and super-alloy markets.

19.1.4 Pricing

Cobalt prices are expected to remain soft during the summer 2012. During the late third quarter of 2012 some firming up of market prices is forecasted as the effects of growing demand combine with refined cobalt production corrections and improving market sentiment to shore up pricing. By the end of 2012, we are projecting Metal Bulletin High Grade pricing to hit US$16/lb. In the following years, pricing can be expected to hit US$18/lb - $22/lb. The emergence of stocks of unprocessed ores/concentrates in the DRC (and also at HPAL nickel producers) in the coming years can be expected to serve as a “limiter” to refined cobalt market prices (given the ability of Chinese refiners to ramp up refining when prices strengthen).

NICO cobalt products will consist of cobalt cathode, cobalt carbonate, and cobalt sulphate. Of the three cobalt products, cobalt sulphate offers the highest realized market price in terms of cobalt value. This is followed by cobalt cathode and subsequently cobalt carbonate.

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Figure 19.1 Application of Industrial Usage

4% Batteries (27%) 5% Superalloy (19%) 6% 27% Hard Materials (13%) 7% Colours (10%) Catalysts (9%) 9% Magnets (7%) Hardfacing & Other Alloys (6%) 10% 19% Tyre Adhesives, Soaps, Driers (5%)

13% Feedstuffs (4%)

Figure 19.2 Global Electric Vehicle Battery Sales

Figure 19.3 Proportion of World Cobalt Production (%)

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19.2 BISMUTH

19.2.1 Uses

In broad terms, there are two broad classifications into which bismuth applications can fall: Chemical and Metal. Chemical applications account for approximately 65% of all bismuth uses. The most common starting forms for most chemical forms of bismuth are bismuth nitrate and bismuth oxides.

Bismuth is a soft metal with very high density and low melting temperature, and is scientifically recognized as one of the safest elements for human consumption. This, together with antibacterial properties, is why bismuth is used in pharmaceuticals and medicines, including Pepto-bismol®, bandage dressings, cosmetics, and some medical devices. The physical properties of bismuth are otherwise similar to lead, but unlike lead, bismuth is not toxic and is therefore used to replace lead in paint pigments, free-machining steel, galvanizing alloys, ceramic glazes, radiation shielding, ammunition, greases, plumbing solders and brasses, and electronics solders. Many of these new applications result from legislation that has banned the use of lead, particularly in potable water plumbing sources in developed countries as well as electronics in the European Union. Bismuth is also one of the few elements that expand when cooled making it important in the manufacture of dimensionally stable alloys and compounds, including metal castings and coatings that could crack from shrinkage during cooling such as automotive anti-corrosion alloy electro-plated on premium automobiles and galvanizing. Bismuth is also used for frit coatings on automotive glass to protect windshield seals from degradation from exposure to ultraviolet radiation and changing temperatures. Some super conductors, fire sprinkler systems, fire retardants, compact discs, and heat transfer alloys used to generate electricity all use bismuth.

19.2.2 Supply

The majority of bismuth supply in 2011 originated from China (85%). China has been the world‟s dominant supplier for decades produced primarily as a by-product of copper, lead, and tungsten mining and refining. Recent de-stocking has resulted in bismuth supply exceeding true demand levels. This situation is expected to rectify itself in the near future, as demand for bismuth registers growth. Supply is expected to remain stable during 2012 and 2013; many of the smaller mines in China will be forced to discontinue operations as a result of environmental clampdowns by the government and/or the unfavourable economic operating conditions prevailing at this time. Other projects not yet operative are not expected to begin for the time being (Nui Phao in Vietnam and Bonfim in Brazil).

The bismuth market is between 15,000 and 20,000 tonnes per year. During 2011 supply of bismuth has outpaced demand, as some de-stocking took place. This situation is soon expected to correct itself.

The NICO deposit contains 15% of global bismuth reserves.

19.2.3 Demand

Bismuth demand is poised to grow significantly on account of its non-toxic nature, combined with physical properties that allow it to be substituted in place of more harmful metals (eg: lead in plumbing and soldering). Recent changes in laws governing plumbing for drinking water P&E Mining Consultants Inc., Report No. 247 Page 206 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. require the elimination of lead. Demand growth is expected to be in the 8% to 10% range annually over the coming 3 years. Bismuth demand is forecasted to enjoy healthy growth due to its role as a non-toxic substitute for lead, especially in such applications as free cutting (machining) steels and coppers/brasses.

19.2.4 Pricing

Since the beginning of 2012, prices have mainly moved sideways posting a small uptick in February 2012 to US$10.50/lb to $11.40/lb, but retreating again by mid-March and remaining at similar levels through to the time of this writing ($10.10/lb to $10.80/lb on June 8). Pricing is expected to remain sluggish during the summer months of 2012, though market prices are forecasted to firm up during the fourth quarter, reaching a level of US$12.50/lb by year-end. With continued healthy demand growth and supply being kept in check, prices are forecasted to increase by US$1.50 in 2013 reaching $$14.00/lb, and then hitting $15.50/lb by the end of 2014. Longer term pricing may see even higher pricing if growth momentum in Pb-substitution applications is speeded along by regulations/legislation.

Figure 19.4 Applications and Uses of Bismuth

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Figure 19.5 World Bismuth Reserves

19.3 PRECIOUS AND BASE METALS

The NICO deposit contains in-situ values of gold and copper. The gold contained in the deposit is about 1.1 million in-situ ounces and at current metal prices, is the dominant revenue producing metal in the deposit. The gold contained in the NICO deposit is a highly liquid product that mitigates project risk from cobalt and bismuth price volatility and can also be used as an attractive financing option to develop the project. Gold and copper are freely traded on the open market.

Figure 19.6 Historical and Forecast Gold Price

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19.4 CASH FLOW COMMODITY PRICES

The Base Case and Escalated Metal Price Case metal price assumptions used in the cash flow model were established from a marketing report prepared by Skybeco. The Current Metal Price Case and 3-Year Trailing Average Metal Price Case use metal prices that were obtained from Metals Bulletin and Kitco as at May 31, 2012. These cases reflect the currently low prices for cobalt in the economic analysis. The Skybeco report and Company marketing information were used to support the average 22% premium price for cobalt sulphate over high grade cobalt metal cathode. The Escalated Metal Price Case and Optimistic Metal Price Case use a par U.S. to Canadian dollar exchange rate.

Base Case Price assumptions are US$1,450/troy ounce (“oz”) for gold, US$20/pound (“lb”) for cobalt, US$11/lb for bismuth and US$3.50/lb for copper at an exchange rate of US$0.95 = C$1. The 3-year Trailing Average Prices Case are as at May 31, 2012 and are US$1,359.94/oz for gold, US$18.53/lb for cobalt, US$9.83/lb for bismuth and US$3.51/lb for copper and an exchange rate of US$0.98 = C$1. The Current Price Case uses prices as at May 31, 2012 and are US$1,558.00/oz for gold, US$15.23/lb for cobalt, US$10.55/lb for bismuth and US$3.40/lb for copper and an exchange rate of US$0.97 = C$1. The Escalated Price Case uses metal price assumptions of US$1,800.00/oz for gold, US$22.50/lb for cobalt, US$12.50/lb for bismuth and US$4.00/lb for copper and an exchange rate of US$1 = C$1. For the Optimistic Price Case uses US$2,000.00/oz for gold, US$25.00/lb for cobalt, US$15.00/lb for bismuth and US$4.50/lb for copper at an exchange rate of US$1 = C$1.

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20.0 ENVIRONMENTAL STUDIES PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1 NICO MINE SITE

20.1.1 Environmental Study Results

DAR was submitted to the MVEIRB in May 2011. This section provides a summary of the impact predictions to the biophysical environments from the NICO Project, which were presented in the DAR.

The active mine area will be small (approximately 485 hectares [ha]), with limited changes made to the natural flow of water. The NICO Project will have a minimal effect on water quantity, air, soils, vegetation, and wildlife and fish health. Closure, caribou and water quality have been identified as the most important concerns related to the environment by the communities. People should not be able to observe a change in the availability of wildlife due to effects of the NICO Project, relative to current natural changes in population size. Changes in water, soils, and plants caused by the NICO Project in the small area at and near the mine site will not affect the health of wildlife, or the health of people that eat wildlife.

20.1.1.1 Caribou

Caribou are important to the aboriginal people and other residents of the NWT; they are a critical component of the diet of many northerners, and are the most important resource harvested by aboriginal groups on traditional lands.

Caribou may be in the area near the NICO Project during winter. Because of noise during operations, caribou might not use the immediate area around the NICO Project during construction and operation. This will be a temporary effect and will disappear at closure.

The calculated cumulative levels of disturbance from the NICO Project and other past, present and reasonably foreseeable projects on caribou seasonal (winter) and annual ranges is well below the ecological threshold for identifying effects of habitat fragmentation. Approximately 85 ha will be permanently altered by the NICO Project. This is mainly the open pit, and the CDF that will fill the valley beside it which will remain as local features on the landscape. The habitat lost will be very small compared to the amount of existing caribou habitat. The surface of the CDF will be allowed to re-vegetate after closure, as will other areas of disturbance which will minimize permanent habitat loss.

Traffic associated with the proposed all-season Tłįchǫ road and the access road into the NICO Project could affect behaviour and movement of caribou. There will be more traffic during the construction period than later during operations. During mining, noise from vehicles along the access road and the Tłįchǫ road will occur year round; however, barren-ground caribou are only in the area in the winter. Although trucks might affect caribou movement and behaviour, the effects will be similar to what already occurs when the winter roads are in use. With the P&E Mining Consultants Inc., Report No. 247 Page 210 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. development of the all-season roads, hunters would be able to make more use of vehicles (including snowmobiles) to access areas in the region, and for longer periods of time compared to the winter roads. Harvesting will affect the caribou population only during winter when caribou are on their winter range. Should harvesting on the all-season roads become a concern, the Tłįchǫ Government or the Wek‟èezhìı Renewable Resources Board could enact restrictions to control the harvest. So increased harvesting pressures can be controlled and the development of the NICO Project should not change the amount of harvesting in the area.

Effects from the mine will be limited to a very small area and unlikely to be a major contributing factor to changes in the abundance and distribution of the caribou herds. Changes in water, soils and plants will also be minimal so that the NICO Project will not affect the health of caribou, or the health of people that eat caribou. People should not observe a change in the availability of caribou due to effects from the NICO Project, relative to current natural changes in population size.

20.1.1.2 Water Quality

The NICO Project includes an ETF that will be based on reverse osmosis (“RO”) with chemical and biological treatment of the brine (this represents a change from the IX-based system that was proposed in the DAR). The RO treatment system produces an effluent quality that is projected to meet all Site Specific Water Quality Objectives (SSWQOs), is robust to changes in influent quality, and is expected to produce an effluent quality that is consistent. The RO system will produce a secondary waste in the form of precipitated metal hydroxides that is stable and compatible for disposal at the site.

A receiving water quality model was developed to predict and evaluate changes in water quality due to NICO Project activities. The model included loadings from the RO treatment system interpolated annually during operations between the “early years” and “worst case” effluent quality projections. The model also included highly conservative loading of metals associated with deposition of fugitive dust emissions and conservative estimates of seepage concentrations from Wetland Treatment Systems to NICO Lake during the post-closure period. Fugitive dust deposition was noted as the largest source of projected changes in receiving water quality for the majority of metals during construction and operations.

The receiving water quality projections for NICO Lake, Peanut Lake, Burke Lake, and the Marian River are discussed below.

Total dissolved solids and associated minerals, such as chloride, potassium and sulphate in the nearby lakes are expected to be higher during operations and after closure than before mining due to discharges from the ETF during operations into Peanut Lake, and seepage into the wetland treatment systems that will flow into NICO Lake after closure. These dissolved minerals are not expected to impact aquatic life. Chloride concentrations are projected to remain well below the Canadian Council of Ministers of the Environment (“CCME”) water quality guideline for the protection of aquatic life (WQG) in NICO Lake and downstream waters; sulphate concentrations are projected to remain well below SSWQO concentrations in NICO Lake and downstream waters.

Projected changes in nutrient concentrations are not expected to result in residual effects to aquatic life or changes in trophic status in NICO Lake and in the downstream environment. Ammonia concentrations are projected to remain below SSWQO concentrations in NICO Lake P&E Mining Consultants Inc., Report No. 247 Page 211 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. and Peanut Lake and below CCME guideline concentrations in Burke Lake and at the confluence with the Marian River; nitrate concentrations are projected to remain below CCME guideline concentrations in NICO Lake and downstream waters. Phosphorus concentrations are not expected to increase significantly.

For the purposes of the environmental assessment, Fortune derived a “worst case scenario” for water quality predictions. Many of the assumptions used were very conservative in nature which allowed for variability in the mine discharges. Water quality under standard operating conditions will be considerably better than the worst case present in the environmental assessment. Under worst case conditions, receiving water concentrations of many metals are projected to change as a result of NICO Project activities:

 Some metals (including aluminum, arsenic, cadmium, cobalt, iron, and selenium) are projected to increase above background levels during operations then decline after closure.  Levels of other metals (including barium, chromium, and vanadium) are projected to increase during operations and remain above background levels after closure.  Levels of some other metals (including antimony, lead, manganese, mercury, molybdenum, thallium, uranium, and zinc) are projected to increase during operations and early stages of closure.  For most of these metals, the increased levels predicted are lower than the water quality guidelines or SSWQOs, which indicates that the water quality will be safe for the environment.

Notable metals concentrations in the assessment are discussed in more detail below:

 Total aluminum concentration projections are above the SSWQO in NICO Lake, Peanut Lake and Burke Lake during operations primarily due to conservative estimates of loading from dust deposition, whereas dissolved aluminum concentrations were projected to remain below the SSWQO in waters downstream of NICO Lake. Aluminum concentration projections are above the CCME WQG during operations in Burke Lake and occasionally in the Marian River, with concentrations declining during closure. Aluminum concentrations in the ETF effluent discharge are less than the SSWQO value, and the projected values in the receiving environment during operations are primarily due to conservative estimates of loading from dust deposition. Aluminum concentrations have also occasionally exceeded the CCME WQG under measured baseline conditions in NICO Lake, Peanut Lake, Burke Lake and the Marian River.  Arsenic concentrations are projected to remain below the SSWQO value in NICO Lake, Peanut Lake and Burke Lake, and below the CCME WQG in the Marian River. As noted in the DAR, arsenic concentrations in Burke Lake are projected to be above the CCME guideline during operations, primarily due to dust deposition, and to decline to levels below the CCME guideline during the closure period. Arsenic concentrations in the Burke Lake watershed are elevated under baseline conditions, frequently exceeding the CCME WQG in NICO Lake.  Total iron concentrations are projected to exceed the SSWQO in NICO Lake, Peanut Lake, and Burke Lake primarily as a result of the deposition of fugitive dust. However, dissolved iron concentrations may only exceed SSWQO values occasionally in NICO Lake, and are expected to remain below SSWQO values in Peanut Lake and Burke Lake. Iron concentrations have also exceeded the CCME P&E Mining Consultants Inc., Report No. 247 Page 212 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

WQG under measured baseline conditions in NICO Lake, Peanut Lake, Burke Lake, and the Marian River.  Although levels of chromium, mercury, silver, and thallium in NICO Lake may be higher than the water quality guidelines or SSWQOs after the mine has closed, and levels of chromium may also occasionally be higher than water quality guidelines in Peanut Lake following closure, no metals are expected to be at levels above water quality guidelines or SSWQOs in Burke Lake or the Marian River. These metals concentrations are not expected to impact aquatic life.

Modelled cadmium, mercury and silver concentrations are influenced by inputs derived from monitoring data that have a large proportion of results lower than method detection limits (“MDLs”). MDL values or half the MDL values were used to derive input concentrations and loading for these metals. This will result in an overestimation of modelled concentrations and is therefore another source of conservatism in the modelling.

The operations water quality projections described above include use of a robust treatment technology with an effluent quality meeting all SSWQOs. The contribution of the ETF effluent to changes in receiving environment metals concentrations are generally small relative to contributions from dust deposition during operations. Projected operations concentrations, including those that exceed CCME WQGs, present “negligible” to “low and likely negligible” risks to aquatic life under estimated worst case conditions.

Exceedances of CCME WQGs for those parameters that do not have derived SSWQOs in the DAR does not necessarily imply toxicity will occur, only that effects may potentially occur. CCME WQGs are conservative, and are intended to protect all forms of aquatic life during all life stages. As stated above, the water quality predictions used in the environmental assessment are a worst case scenario for the purposes of identifying potential impacts and risks to the environment. As discussed in the DAR, follow-up monitoring will be used when operations commence to assess whether effects are occurring and to track the uncertainties of projections. Fortune is committed to undertaking regular monitoring and follow-up testing of water quality and aquatic health during the NICO Project.

The weight of evidence from the analysis of the primary pathways predicts that the incremental impacts from the NICO Project will result in changes to water quality in NICO, Peanut, and Burke Lakes under worst case conditions, but that these changes, incorporating the evaluation of risk to aquatic life, will not have a significant adverse impact on the suitability of water in these lakes to support a viable and self-sustaining aquatic ecosystem.

As noted in the DAR, there were a number of uncertainties in the assessment of effects to water quality, and conservative assumptions were applied in the receiving water quality modelling so that potential effects would not be underestimated. The major uncertainties included metals loading due to dust deposition during construction and operations, and design requirements of wetland treatment systems during the post closure period.

While conservative assumptions were used in the assessment to provide confidence that changes to water quality will not be worse than projected, they also provide an upper bound in order to develop adequate mitigation. Projected water quality is based on several inputs (i.e., surface flows, groundwater flows and seepage, background water quality and geochemical characterization), all of which have inherent variability and uncertainty. As such, it is suggested that water quality projections should not be used to predict absolute concentrations, but rather as P&E Mining Consultants Inc., Report No. 247 Page 213 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. a planning tool and to develop monitoring plans. It is anticipated that ETF discharge and seepage will be monitored during operations to compare to projections contained in the DAR. If it is identified that the quality of discharge or seepage varies from the projections, adaptive management strategies will be triggered.

The air quality and deposition rate projections used the maximum emission rates from the NICO Project during construction and operations, and projected annual deposition rates based on maximum daily road dust emissions during summer and winter. Emissions of road dust from on- site haul roads, the primary sources of particulate matter and metal compounds, do not include potential mitigating effects of weather (such as precipitation or snow-covered ground), which results in an overestimation of annual fugitive dust emissions and deposition rates. Furthermore, the receiving water quality model assumed no retention on the landscape and that deposited dust particles, and associated metals, less than 10 microns in diameter would remain in suspension indefinitely. These highly conservative assumptions have likely resulted in an overestimation of metals concentrations in receiving waters during operations.

Conservatism was also applied to the quality of discharges from the Wetland Treatment Systems after closure, in that the projected outflow quality is generally expected to be better than the influent quality projections that were applied to the outflows. However, influent concentrations were only adjusted to cap projected exceedances of SSWQO concentrations at the respective objective concentrations. No further adjustments were applied due to uncertainty regarding the constituent-specific effectiveness of the planned passive treatment system. Therefore, projected water quality presented in the DAR for closure periods is expected to be overestimated for constituents without SSWQO values.

Work is ongoing towards defining the effectiveness of wetland treatment systems. Results of bench scale testing of a passive treatment system, constructed to demonstrate the feasibility of passive treatment as a post-closure water treatment strategy for the NICO Project, showed effective removal of aluminum, arsenic, cadmium, cobalt, lead, iron, selenium, and uranium from a concentrated brine created from process water generated by pilot plant operations. The bench test results appear to support the assertion that the receiving water quality model assumptions regarding Wetland Treatment System effluent concentrations are conservative.

20.1.2 Waste and Tailings Disposal, Site Monitoring and Water Management

20.1.2.1 Co-disposal Facility

The mining process will generate a total of 29.9 Mt of tailings and 96.9 Mt of mine rock. Both these waste streams will be disposed together in a facility referred to as the CDF. The general arrangement plan of the CDF is shown on Figure 18.4.

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The co-disposal of tailings and mine rock was selected as the preferred mine waste management system for the following reasons:

 Reduces the footprint area requirement (relative to separate facilities for tailings and mine rock);  Maximizes the rate of consolidation of the tailings;  Improves the stability of the disposal facility;  Reduces metal leaching and acid mine drainage;  Minimizes freeze dry and dusting;  Reduces the mine hauling distance and tailings pumping length; and  Allows progressive closure.

The CDF will be contained by a Perimeter Dyke comprising a prism of mine rock at least 25 m thick. The Perimeter Dyke will be raised continually in 5 m lifts using the upstream construction method. Inside the Perimeter Dyke, the CDF will comprise a “layer cake” of alternating layers of mine rock and tailings about 5 m thick. The tailings layers will be created by constructing a series of tailings disposal cells using waste rock. Typically, each tailings disposal cell will be a nominal square of 200 m by 200 m. The tailings disposal cell will be filled with non-segregating thickened tailings. A layer of waste rock will be pushed over the filled tailings disposal cells using bulldozers.

The Perimeter Dyke will be free draining but it will retain tailings particles. Five SCPs will be constructed downstream of the CDF at topographically low areas (Figure 18.7) to intercept any tailings water that may seep through the Perimeter Dyke. Water collected in the SCPs will be pumped to the Process Plant for re-use.

There have been three rounds of geotechnical investigation campaigns within the footprint areas of the CDF and associated water management facilities.

20.1.2.2 Water Management

The major components of the water management system of the Project will comprise: Lou Lake, the CDF Reclaim Ponds, five SCPs, a Surge Pond, Open Pit sumps, a Process Plant Runoff Pond, sewage treatment plant (STP) and an ETF.

During Operation

Lou Lake will be the source of fresh water for Process Plant, dust control, and potable water. The CDF will collect tailings water, precipitation and runoff. This water will be temporarily stored in Reclaim Ponds that will be created within the CDF. A movable pump barge will be used to pump the supernatant and run-off water from the Reclaim Ponds back to the Surge Pond. SCPs No. 1, 2 and 3 will be located in three topographic lows adjacent to the western end of the CDF and they are designed to intercept seepage from the CDF which would otherwise flow to NICO Lake. SCPs No. 4 and 5 will be located north and southeast of the Open Pit, to collect localized seepage from the CDF. The open pit will collect direct precipitation, run-off and groundwater seepage. This contact water will be collected in sumps and pumped to the Surge Pond. The Surge Pond will be located in a topographic low north of the Process Plant to temporarily store contact water pumped back from the SCPs and the Reclaim Pond. The Process Plant Runoff Pond will be a small lined pond to be constructed within the area of the Process Plant to collect local runoff. Water will be pumped from the Surge Pond either to the Process Plant for reuse or to the P&E Mining Consultants Inc., Report No. 247 Page 215 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

ETF for treatment. The ETF will use a RO system Treated water from the ETF will be pumped through a diffuser directly into Peanut Lake. Effluent from the STP will be discharged through the same diffuser in Peanut Lake. Water balance analysis indicates that the average flow discharged into Peanut Lake will be relatively small flow of about 290,000 m3/year.

After Closure

The water which accumulates in SCP Nos. 1, 2, 3, and 5, as well as the Surge Pond, will be passively treated in Wetland Treatment Systems and then released directly into NICO Lake. (This is subject to the demonstration of the technical performance of the Wetland Treatment Systems during operations. Continued treatment in the ETF is a contingency.) At the end of mining, pumping of water out of the Open Pit will cease and the Open Pit will slowly fill with water. Backflooding of the Open Pit is beneficial because higher water levels will reduce any localized areas of potentially acid generating rock that will be exposed to atmospheric conditions, thus reducing the total metal loading from the pit wall runoff over time. Modelling indicates that it will take roughly 120 years for the Open Pit water level to rise to Elev. 260 m, at which time the flooded Open Pit may begin to overflow. Just prior to pit overflow, the water quality at the top of the flooded Open Pit will be evaluated, and a decision will be made regarding post-overflow treatment.

Closure

For closure, a soil cover will be placed over the entire CDF facility. On the top of the CDF, the cover system will comprise 0.5 m of glacial till underlain by 0.25 m of sand. The sand layer will act as capillary break to minimize the potential for upward flux of tailings pore water, reducing the potential for arsenic uptake by vegetation. The top surface of the closed CDF will be sloped at about 2% to enhance the water shedding capacity of the facility. On the surface of the sloped perimeter dyke, the cover will comprise 1 m of glacial till. Progressive reclamation of the CDF will be undertaken during the operating life of the mine. By the time the mine operation is completed, about 85% area of the CDF will have been covered and revegetated. The CDF covers will be completed and revegetated within 2 years after operations cease.

As stated earlier, the seepage water that reports to SCP Nos. 1, 2, 3 and 5 will be routed through Wetland Treatment Systems into NICO Lake. The Wetland Treatment Systems will be constructed and tested prior to the end of mine operations. The ETF will be maintained on site for 10 years as a backup to the Wetland Treatment Systems.

At the end of mining, pumping of water out of the Open Pit will cease and the Open Pit will slowly fill with water. Backflooding of the Open Pit is beneficial because higher water levels will reduce any localized areas of potentially acid generating rock that will be exposed to atmospheric conditions, thus reducing the total metal loading from the pit wall runoff over time. Modelling indicates that it will take roughly 120 years for the Open Pit water level to rise to Elev. 260 m, at which time the flooded Open Pit may begin to overflow. Just prior to pit overflow, the water quality at the top of the flooded Open Pit will be evaluated, and a decision will be made regarding post-overflow treatment. The options include the following:

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There are several alternatives for the treatment of the overflow water, including:

 Relying on stratification of water in the flooded open pit to ensure that the surficial water is suitable for discharge to Peanut Lake;  Treating the water in the flooded open pit prior to overflow by chemical or biological means;  Treatment of the overflow water in Wetland Treatment System No. 4 to be constructed in the former Contingency Pond area prior to discharge into Peanut Lake; or  Building a new Effluent Treatment Facility (ETF) for active treatment prior to discharge into Peanut Lake.

20.1.3 Project Permitting

The NICO Project is regulated by the WLWB under the MVRMA. The MVRMA implements provisions of land claim agreements and establishes co-management boards as institutions of public government. The Tłįchǫ Government and the WLWB regulate the use of settlement and Crown land and water in their respective settlement areas. For developments that may have effects that extend beyond the Mackenzie Valley, the WLWB regulates the use of land and water within Tłįchǫ territory. If a proposed development has potential to cause a significant adverse effect on the environment, or if it is likely to cause public concern, the development can be referred to the Mackenzie Valley Review Board.

The MVRB is established under the authority of the MVRMA to review the potential environmental effects of developments proposed within the Mackenzie Valley area of the NWT. There are 3 stages in the environmental assessment process in the Mackenzie Valley. The MVRB provides the following description of the stages:

 Preliminary Screening: A Preliminary Screening is a quick review of a proposed development‟s application to decide if the development might have significant adverse impacts on the environment, or might cause public concern. If so, the application is referred to the second stage – Environmental Assessment.  Environmental Assessment: The MVRB conducts Environmental Assessments. This stage is a more thorough study of a proposed development‟s application to decide if the development is likely to have significant adverse impacts on the environment, or likely to cause public concern.  Regulatory Phase: Once Environmental Assessment process is complete, the NICO Project would go back to the Wek‟èezhìı Renewable Resources Board for the regulatory phase of permitting. In this phase, the details of the water license and land use permits are established. Once received, the project would proceed to construction when ready. There are a number of other permits that must also be obtained during the regulatory phase to allow for construction and operations to proceed.

20.1.3.1 NICO Project and the Impact Assessment Process

Fortune submitted applications for a Type A Land Use Permit and a Type A Water License to the WLWB on 5 November 2008. The WLWB then sent the applications to interested parties and regulators for review in January 2009. The WLWB initiated a Preliminary Screening and Indian and Northern Affairs Canada (INAC) (now Aboriginal Affairs and Northern Development P&E Mining Consultants Inc., Report No. 247 Page 217 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Canada (“AANDC”) referred the applications to the MVRB for an EA on 27 February 2009. The MVRB notified Fortune of the decision on 2 March 2009. Proof of community engagement for the Type A Water License was submitted to the WLWB on 21 January 2009.

Community scoping sessions were held in Yellowknife (20 April 2009), Whatì (27 April 2009), Behchokǭ (4 May 2009), Gamètì (7 May 2009), and Wekweetì (2 and 3 November 2009). The TOR for the NICO Project were issued in draft on 15 September 2009, and in final form on 30 November 2009. The MVRB developed the draft TOR based on the information gathered, and public interest demonstrated, at the community scoping sessions and from comments provided by regulators, community governments, and stakeholders. The TOR provided direction for Fortune to organize existing information into a stand-alone DAR, [the NWT analogue to an EIS, which was submitted on 20 May 2011. The DAR was used to inform interested parties about the development during the analytical phase of the environmental assessment.

In October 2011, regulators, community governments, and stakeholders submitted Information Requests (“IRs”) as follow-up to the DAR. The purpose of an IR is to give parties the information needed to help reach conclusions about potentially significant impacts. Fortune submitted responses to these IRs in December 2011. A Technical Session was held in Yellowknife, NWT from 7 to 9 February 2012, open to regulators, community governments, and stakeholders to further discuss key topics that parties require clarification on or wish to discuss in greater detail. A second round of IRs was received by Fortune issued on 20 April 2012. Fortune submitted its responses to the second round of IRs on May 11, 2012. No further technical meeting will be held to discuss this second round of IRs. Technical reports are to be submitted on 15 June 2012, and Public Hearings are tentatively scheduled for the week of August 27-31, 2012.

After the hearings, the MVRB will make a recommendation to the Federal Minister and, assuming the recommendation is affirmative, the file will be returned to the WLWB for the regulatory stage of permitting. The regulatory phase involves the development and issuance of a water license and land use permit that defines the specific conditions under which the mine must be constructed and operated. Once these authorizations are in place, other permits, licenses, and authorizations can be obtained.

20.1.3.2 Reclamation Bond Requirements

Fortune will be required to provide a security bond (or another acceptable form of financial assurance) to cover the projected costs of closure and reclamation of the NICO Project. The bond will be held by the government through AANDC.

20.1.4 Social and Community Requirements and Considerations

The NICO Project is anticipated to have significant positive impacts on the economics of the Tłįchǫ communities, and both positive and negative (but not significant) impacts on the social and cultural environments. The NICO Project is a small development compared to other mines in the NWT, but it will contribute to the overall labour, financial, physical, human, and social resources of both the Northwest Territories and more specifically the nearby communities. Benefits will result despite some employment challenges, which include previous experience, employability and availability in light of minimum education and skill requirements, advancement opportunities, employee retention, criminal records, and drug and alcohol use. Overall, the NICO Project is expected to have few negative effects on people, business capacity,

P&E Mining Consultants Inc., Report No. 247 Page 218 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. or public infrastructure and services. Most negative effects can be overcome with plans that Fortune has developed.

Because the NICO Project is a small development, it will have limited effects on society and culture. The NICO Project is not expected to either substantially increase or decrease education levels or health and wellness indicators.

In general, this project will increase the amount of money in the area through additional wages and business activities, with secondary benefits such as improved roads and spending. Money will be made through employment and in contracts to provide supplies and services. Increased spending of money earned by the NICO Project employees and contractors will spread the benefits. Much of the additional spending is expected to be in Yellowknife, unless the mine is used as an opportunity to open businesses with a greater variety of goods or services in the smaller communities. The presence of an all-season road would make it easier to supply businesses in these communities.

There may be a need to hire workers from outside the NWT due to a shortage of trained mine workers, but as some of the existing mines begin to wind down over the next few years, some experienced workers will be able to shift to the NICO Project.

The NICO Project is expected to contribute social benefits with increased labour force participation, especially for the potentially-affected communities closest to the mine site. The location of the NICO Project is a substantial benefit to the Tłįchǫ communities since flexibility with shift rotation and time for cultural traditions will be more easily accommodated. More women may be able to participate in the NICO Project due to its close proximity to their home communities.

Workers will have the ability to purchase vehicles and recreational vehicles for pursuing traditional cultural activities, such as trapping and hunting. They may make improvements to their houses and the quality of food and clothing. Additional income and time off will also allow some workers from the communities to be more fully engaged in traditional activities, including hunting, fishing and trapping.

In summary, impacts to economics in the area are expected to be positive and to last until the mine is closed. The positive and negative impacts to health and wellness, and public safety are expected to be small. Post-development conditions are expected to be similar to existing conditions. Impacts on infrastructure are expected to be positive and are expected to last through the life of the mine.

20.1.4.1 Corporate Commitments and Agreements

Fortune has also agreed to fund a TK Study that will focus on providing traditional knowledge and land use practice information for the environmental review of the NICO Project. This study, P&E Mining Consultants Inc., Report No. 247 Page 219 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. which will be carried out by the Tłįchǫ Government, will contribute to the environmental assessment process.

The NICO Project is expected to employ the equivalent of about 231 people during construction who will come into site at different times to work on specific jobs. The greatest number of work opportunities, about 233 jobs, will be during the first eight months of operation when the underground and open pit will both be operating. After the underground is finished, there will be about 127 jobs available during the last 16 years of operations. Employment during closure and reclamation is expected to be fewer than 100 jobs. During operations there may be jobs that can accommodate a different work rotation allowing people to work at NICO who cannot balance the two weeks in and two weeks out with their home life.

Fortune will provide employees with a comprehensive benefits package that is competitively balanced in the Northwest Territories with the following included:

 Health care;  Group registered retirement savings plan;  Group life insurance;  Medical travel assistance;  Education assistance program;  Site allowance;  Northern travel allowance; and  Employee assistance program.  Hiring preference will be given to local aboriginal and northern residents as part of Fortune‟s commitment to provide employment and business opportunities to northerners, which also includes support of apprentices at the mine.

20.1.5 Mine Closure Requirements and Costs

Mine closure and reclamation is discussed in Chapter 9 of the DAR, because it is an important part of the Project Description. The next step is to develop a formal Electro Recovery (“CLER”) dimensionally stable anodes (DSA)Xstrata Process Support Laboratories (XPS) Bench Face Angle („BFA‟) and Inter-Ramp Angle („IRA‟) and Overall Slope Angle („OSA‟)methyl isobutyl carbinol (“MIBC”) vacuum pressure swing adsorption (“VPSA”)Ion Exchange (IX) Site Specific Water Quality Objectives (“SSWQOs”), reverse osmosis (“RO”) CCME Aboriginal Affairs and Northern Development Canada (“AANDC”) Conceptual Closure and Reclamation Plan (“CCRP”) according to the requirements of the AANDC and the Wek‟èezhìı Renewable Resources Board. The CCRP will include an estimate of closure costs and a financial assurance will be posted based on this estimate. The amount and form of this financial assurance will be subject to negotiations between Fortune and AANDC. As suggested by MVLWB (2009), the CCRP will be updated to an Electro Recovery (“CLER”) dimensionally stable anodes (DSA)Xstrata Process Support Laboratories (XPS) Bench Face Angle („BFA‟) and Inter-Ramp Angle („IRA‟) and Overall Slope Angle („OSA‟)methyl isobutyl carbinol (“MIBC”) vacuum pressure swing adsorption (“VPSA”)Ion Exchange (IX) Site Specific Water Quality Objectives (“SSWQOs”), reverse osmosis (“RO”) CCME Aboriginal Affairs and Northern Development Canada (“AANDC”) Conceptual Closure and Reclamation Plan (“CCRP”) Interim Closure and Reclamation Plan (ICRP) and that in turn will be reviewed and updated every 3 years during operations.

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Fortune has met with AANDC to begin initial discussions for scenarios for bonding for closure in terms of the financial amounts that would need to be in place, particularly around 2 key issues. The first is the reclamation required immediately at the end of the mine life. The second is for dealing with potential water quality issues once the open pit overflows, if required, subsequent to the mine closure. The amount of the security required will be agreed upon with AANDC during the regulatory phase. The security will be deposited at agreed upon dates and milestones to ensure that the funds required for future reclamation will be available.

The security will be held in-trust and deposited in low-risk investments intended to meet or exceed the cost of inflation over time so that additional funds will not be required in the future. Fortune will not be able to access the funds held in-trust and only the designated beneficiary, currently the Receiver General for Canada, will be able to release any funds. Criteria will be established and will need to be met prior to release of any security held by the beneficiary. Funds may be released by the beneficiary back to Fortune, only if Fortune has satisfied its reclamation obligations. To the extent funds are held in-trust to cover unforeseen future reclamation costs or in the event that certain reclamation activities are not completed the regulatory authorities will have the right to use the security funds to fulfill any necessary obligations.

Fortune currently has posted a letter of credit in favour of the Receiver General for Canada for existing environmental liabilities related to advanced exploration activities. The letter of credit is secured by investment accounts held with a large Canadian financial institution. Under this arrangement, the security currently held is in excess of the amount required by regulatory authorities for future reclamation.

20.2 SMPP

MDH, a member of the SNC-Lavalin Group, completed an EIS of the Fortune Saskatchewan Metals Processing Plant (SMPP), which is proposed to be a hydrometallurgical facility to process high-value metal concentrate produced from the NICO mine in the NWT. The SMPP will use proven hydrometallurgical processes to produce cobalt, bismuth, and copper cathode products, as well as bismuth ingots, gold doré, and a small amount of nickel precipitate. Annually, the SMPP will process approximately 65,000 tonnes of metal concentrate produced from the NICO mine.

20.2.1 Waste and Tailings Disposal

A process residue storage facility (“PRSF”) will be used to permanently store process residues generated from the metals processing plant. It is expected that approximately 158,000 tonnes of residue will be produced each year. The PRSF will be an engineered containment facility, designed to minimize the potential impact to the surrounding environment. FML is planning to construct the PSRF directly north and northeast of the proposed plant site. This area of land is sufficiently large to allow for a sizable buffer zone around the PRSF.

The PRSF will be divided into cells to provide containment and storage of the process residue. This cellular design minimizes the active footprint, will allow for liner repairs (if required), and enable active decommissioning throughout the project life. Construction of the containment cells will be in pairs with each cell capable of accommodating between 2 and 2.5 years of process residue. Eighteen years of residue storage capacity is provided by PRSF Cells 1 through 8. An additional 7 years of storage capacity, for a total facility life of 25 years, can be obtained from the construction of PRSF Cells 9 and 10, at the discretion of FML. The 18 year facility covers 42.8 ha and accommodates 1.51 million m3 (Mm3) of residue. The 25 year facility covers 57.9 ha P&E Mining Consultants Inc., Report No. 247 Page 221 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. and will contain 2.09 Mm3 of residue. These additional cells will have an increased capacity to each store 3.5 years of process residue. Cell geometry will vary, however, each cell consists of a 4.65 m deep excavation and 2 m high containment dykes. The containment dykes are planned to be constructed from select material obtained from the cell excavation. The containment dykes will be constructed with a 10 m top width, 3H:1V interior side slopes, and 7H:1V exterior side slopes. The dyke top width and shallow exterior slopes were designed to accommodate use by haul trucks from the plant site. Additional details regarding the PRSF can be found in Appendix A (Volume II) of the EIS document.

A „dry tomb‟ approach was selected for containment and long-term storage of the SMPP residue, such that each cell is constructed above the groundwater table and capped with a „store and release‟ engineered cover system after being filled with residue. A 0.5 m freeboard allowance in the containment cells will be in place prior to construction of the cover system. This cover system will control dust, limit water and oxygen ingress, and support vegetation. Store and release covers may consist of single or multiple soil layers with a vegetated cover. The cover soil acts to store moisture that is released back to the environment through evaporation and/or transpiration (evapotranspiration) processes. Storage and evapotranspiration then limits the net percolation of moisture into the subsurface. Additionally, water stored in the cover soil will limit the amount of oxygen ingress through the saturated cover system which may react with the stored waste.

Each SMPP cell will have a dual containment liner and a leak detection system. The primary liner will be a composite liner consisting of a geomembrane placed directly over approximately 0.45 m of compacted soil. Leak detection is provided by a geocomposite material installed beneath the primary liner. Secondary containment is provided by approximately 0.2 m of a compacted soil liner under the geocomposite material. The primary environmental receptor for the downward migration of contaminants from the PRSF is the Dalmeny Aquifer (comprised of the Upper and Lower Floral Aquifers). There is approximately 9 m to 18 m of low conductivity till between the base of the PSRF and the Dalmeny Aquifer, providing a high level of secondary containment for the process residue.

A leachate collection system was also provided for each cell to collect any fluid (i.e. leachate, precipitation, snowmelt, etc.) that accumulates when the cell is open. This system sits directly over the primary liner and consists of a perforated drainage pipe in a herring bone pattern with a granular cover. The base of the cells will be graded towards a granular sump where access to the leachate collection system will be provided through a perforated pipe.

Perimeter ditches around the PRSF facility and a runoff collection pond dedicated to the PSRF will collect any runoff once the cells are capped, prior to the establishment of vegetation. This collected runoff may be directed to the process water storage pond, for use in the facility, or monitoring may indicate its suitability to be released to the environment. It should be noted that this surface runoff from the capped PRSF cell will not be exposed to the process residue beneath and may eventually be eliminated as the vegetation on the capped mounds will intercept runoff.

20.2.1.1 Reclamation

Reclamation will occur at the SMPP site on an ongoing basis as the portions of the PRSF will be capped and vegetated while the facility is in operation. This will provide a “working model” to identify potential concerns and problems while FML is at the site, to ensure all issues are

P&E Mining Consultants Inc., Report No. 247 Page 222 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. identified and any adjustment to the design can occur, if needed, before additional cells are constructed.

Once the SMPP facility ceases operations, reclamation will generally consist of re-contouring, replacing topsoil, and re-vegetating to restore the land surface to as near as possible to the original conditions. Surface drainage conditions will be restored similar to conditions before site construction. All ditches will be filled in and contoured to blend with the pre-development terrain and drainage patterns. Salvageable topsoil from site decommissioning and possibly facility construction will be used to re-contour the landscape, where applicable. Plant species selected to provide a vegetative cover on each cap will be compatible with the surrounding vegetation. This will ensure that the established vegetation will provide a self-sustaining vegetative cover, according to the cover system design criterion. The establishment of a permanent vegetative cover should include appropriate pre-field preparation, site preparation, revegetation, and post planting management.

20.2.2 Waste Water Injection Well

The proposed SMPP facility has been designed with one active injection well (and one backup injection well) installed in the sands of the Lower Cretaceous aged Mannville Group. The Mannville Group is the primary target for completion of the well, as it is the first available aquifer horizon that is saline, is not used as a potable water source, and has already been used as a waste water injection horizon in the Saskatoon area. The Mannville Group is located at a depth approximately 475 m below ground level and is approximately 50 m thick in the vicinity of the SMPP site.

The injection well(s) will be used to inject brine solution process water. The brine solution process water does not have significant heavy metal contents, but is high in the common ions 2+ 2- found in natural groundwater: sodium (Na+), magnesium (Mg ), sulphate (SO4 ) and chloride (Cl-). The anticipated approximate total dissolved solids (TDS) concentration of this wastewater is 47,500 mg/L, consisting primarily of sodium and sulphate ions with subsidiary magnesium and chloride ions (Table 20.1). This wastewater is planned to be injected into the Mannville Group, which has an anticipated TDS content of 30,000 to 70,000 mg/L in the vicinity of the SMPP.

TABLE 20.1 PROCESS/BRINE SOLUTION CHARACTERISTICS Stream Properties Elemental Composition (mg/L) Percent Solids 0.02 Na 12,930 Mn 0.00 As 0.00 Temperature (°C) 39 Mg 1,610 Fe 0.00 Cd 0.00 pH 10 - 11 Al 0 Co 0.01 Au 0.00 Specific Gravity 1.03 S 10,400 Ni 0.00 Pb 0.00 Cl 1,620 Cu 0.00 Bi 0.00 Ca 180 Zn 0.00 2- Note: Assuming sulphur (S) in solution exists as sulphate ion where 1 mg/L S is approx. 3 mg/L SO4 .

The location of the proposed injection wells (IW1 and IW2) are based on the required 75 m buffer from above ground utilities, 40 m buffer from below ground utilities, 200 m from the property line, and 100 m from any building, as outlined in The Oil and Gas Conservation Act (1985) and The Oil and Gas Conservation Regulations (1985).

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20.2.2.1 Site Monitoring

Fortune environmental monitoring programs will be conducted to comply with the requirements of Saskatchewan‟s Environmental Management and Protection Act and Regulations (2002) and The Clean Air Act and Regulations (2003). Results of environmental monitoring will be reported annually to the MOE in accordance with the approval to operate. The results of monitoring conducted throughout the year will be compiled in an annual environmental report, submitted to MOE. Data will be collected in a manner that complies with industrial, federal, and provincial standards to ensure data collection is consistent and information is representative of the conditions that exist at the site.

As part of the reporting, a summary and discussion of any potential impacts to the environment surrounding the facility and PRSF area from SMPP operations will be required. A summary of the action FML proposes to take as a result of any potential impacts will also be required for submission to MOE. The collection of baseline environmental data at the proposed site, ongoing site monitoring, and the ability to monitor regional environmental receptors (i.e. wells, surface water, vegetation, etc.) will provide an effective method to determine if the proposed SMPP is impacting the surrounding environment.

Site monitoring at the SMPP will include:

 Groundwater monitoring;  Surface water monitoring;  Air emissions;  Soils; and  Containment dykes.

FML will annually update emergency plans, if necessary, for general spill reporting, PRSF contingency plans, berm/dyke failure, and other emergency situations at the site. The procedures will be amended as required to reflect the addition of new process and activities at the proposed SMPP.

20.2.2.2 Post-Decommissioning Monitoring

Long-term monitoring of the site, including the cover systems and monitoring wells, will occur following the closure of the plant and the capping of the containment cells. This monitoring will likely consider surface and groundwater chemistry, dyke stabilities, and permanent cover of vegetation. A detailed plan to monitor the site after decommissioning will be developed in consultation with the provincial government. It may be necessary to install further monitoring wells over time to comply with MOE environmental regulations.

The PRSF at the SMPP will remain in-place, along with associated leak detection monitoring wells. This will require monitoring to occur for a period of time required by provincial regulators, after the site is decommissioned. Initially monitoring events are planned to occur biannually until minimal change is detected. Annual monitoring will occur thereafter. If significant environmental changes occur during annual monitoring, the monitoring will revert to a biannual frequency until any trend stabilizes. According to provincial guidelines, once it is determined that the decommissioning and reclamation at the SMPP is successful and closure objectives and requirements are met, FML may apply to have the site placed into provincial custody. P&E Mining Consultants Inc., Report No. 247 Page 224 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

A contingency plan, specific to post-decommissioning, will be developed to determine appropriate courses of action and contact people to inform in the event of an accidental release to the environment.

20.2.3 Storage Ponds

The operation of the SMPP requires the construction of five storage ponds for site runoff, process water, brine solution, cobalt solution, and runoff from the PRSF. An additional effluent water re-use pond was included in the design in the event that FML is able to implement additional water re-use programs in the future. Preliminary sizing of these storage ponds has been completed and final designs are pending.

Each of the five ponds will be enclosed by a perimeter dyke with a design dyke height of 1.0 m. The exterior dyke slopes are planned to be 3H:1V for construction and 1H:1V for post- construction. These dykes are designed to be 1 m high, only. The interior dyke slopes and excavation slopes vary depending on the storage pond. Containment dykes will provide the ponds with freeboard allowance, while limiting excavation depths and providing a barrier around the ponds, for safety purposes.

The process water, cobalt solution, and brine solution ponds were designed to accommodate the following storage volumes:

 Process water pond – 5,000 m3;  Cobalt solution pond – 300 m3; and  Brine solution pond – 125 m3.

These ponds were combined into one pond facility with three cells which will be located within the plant site.

The site runoff pond was designed to accommodate the runoff from a 1:50 year precipitation event over the plant site area including the buildings, parking lot, and rail areas. Additional storage was provided in the site runoff pond to accommodate supplementary runoff from snowmelt. The additional storage will also prevent overflow in the case that a 1:50 year precipitation event occurs and the pond is not completely empty. The site runoff pond will also be constructed with a 0.45 m thick compacted soil liner. This pond is expected to collect runoff from precipitation falling onto the plant footprint and building exteriors, only. As previously mentioned, each building will be self-contained with sumps depositing any spilled material back into the tank it came from. A combination of site grading, ditches, and/or berms will direct surface flows into the runoff pond. The site runoff pond will be located on the west side of the plant site. Water collected in this pond will be tested and suitable water will be pumped into the site water system, to reduce the volume required from the primary groundwater supply well. If the water is unsuitable for re-use, it may be disposed in the waste water injection well or transported to an approved facility for disposal. It is anticipated that all water collected in the site runoff pond will be suitable for use as process water.

An additional runoff pond will be required to collect any water which comes from the PRSF, as previously mentioned. This residue storage collection pond will collect all water from the area during construction, residue placement, and once the cells have been covered. A series of ditches will collect water from the area and divert flows to the pond which will be centrally located. The P&E Mining Consultants Inc., Report No. 247 Page 225 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. pond was also sized to accommodate a 1:50 year precipitation event. The additional storage required for snowmelt and required capacity when the pond is not empty during a 1:50 year event will be provided by the ditches. The residue storage collection pond will also be constructed with a 0.45 m thick compacted soil liner.

The provisional design for the water re-use pond, located in the land reserve for future water reuse opportunities to the west of the plant site, includes a 0.45 m thick compacted soil liner. This pond has a required storage volume of 110,000 m3. The water re-use pond would be constructed at a later date, once it is determined that it is needed.

20.2.4 Anticipated Decommissioning Costs

A detailed decommissioning and reclamation plan will be developed once the project designs are finalized. An estimate of the decommissioning cost will be included in the plan. Financial assurances will be determined in consultation with MOE during the permitting of the proposed project. It can be expected that some form of an assurance fund (i.e. dedicated trust fund, irrevocable letter of credit, or term deposit) will be determined by FML and the appropriate regulatory agency.

20.2.5 Potential Social or Community Related Requirements

Fortune strives to maintain good public relations in their community. There have been several forums and opportunities for feedback from the public on the proposed project. Direct notification was provided to the municipal government and local landowners in the immediate vicinity of the proposed SMPP site. FML published advertisements in community newspapers and handed out pamphlets (door to door) in Dalmeny and Langham regarding open houses on 7 and 8 February 2011, respectively. Neighbouring landowners and the general public were invited to review and provide feedback regarding the proposed project. Public notice of the EIS will be provided in local papers by MOE and on the internet.

Fortune personnel have worked closely with Enterprise Saskatchewan, Saskatchewan Regional Economic Development Association (SREDA), and representatives from the towns of Langham and Dalmeny to keep government agencies and the public informed about the proposed SMPP. Various meetings and interviews have taken place from April 2009 to the present day with landowners, residents of Langham and Dalmeny, RM representatives, and various media outlets to keep people aware of FML‟s plans to develop the proposed project. Meetings at the homes of seven surrounding landowners occurred in November 2009, to introduce the project and discuss their concerns. These concerns and feedback became part of FML‟s continuous improvement for the project. Representatives from the Rural Municipality of Corman Park met with FML personnel on 24 June 2009 and 10 June 2010 to discuss the SMPP project and any updates on the project. The meeting provided a chance for the RM to voice their comments, concerns, and questions to FML, if any. The project was received favourably by the RM.

In August 2010, FML published a “frequently asked questions” pamphlet which they distributed to Enterprise Saskatchewan, landowners, and posted on their company website. Several meetings also occurred in April 2010 with surrounding residents with water well(s) within 4.5 km of the P&E Mining Consultants Inc., Report No. 247 Page 226 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. proposed project. Personnel from FML informed people of the SMPP project and obtained well usage information and permission to collect samples to obtain baseline water chemistry. Appointments were offered to well owners to accommodate their personal schedules.

Once the SMPP is operational, FML plans to distribute a newsletter twice a year to update the local residents on any changes regarding the facility. They may also hold yearly town meetings to address any concerns or questions about the facility and to provide updates. FML will have an open door policy when it comes to dealing with any questions or concerns from the public. A public relations officer will be designated to answer any questions on a daily basis. Contact information for this person will be widely distributed. FML may also sponsor local events (i.e. charity events, ball or hockey tournaments, community fundraisers), although the exact level of sponsorship cannot be detailed at this time.

20.2.5.1 Geotechnical Foundation Investigation

A geotechnical investigation for the proposed Fortune Minerals Limited Saskatchewan Metals Processing Plant project was completed by MDH, a member of the SNC-Lavalin Group, in June 2010. The scope of the project was to complete a geotechnical evaluation of the site in support of foundation design for the plant building and related geotechnical engineering works.

The field component of the investigation consisted of drilling 8 boreholes, excavating 16 test pits, and installing 1 piezometer to characterize the subsurface. Ground resistivity testing was also completed to collect necessary information for building grounding and cathodic protection for concrete reinforcement and other buried metal structures vulnerable to chloride induced corrosion.

A general description of the soils encountered, the soil properties (including resistivity), anticipated behaviour of soils during construction, and measured groundwater levels are also provided in the geotechnical investigation report. Geotechnical recommendations for shallow foundations, grade supported slabs, pile foundations, and other general geotechnical engineering parameters related to the plant building foundation were provided. The foundation design parameters were derived from calculations based on the Canadian Foundation Engineering Manual (2006) and other relevant geotechnical references.

20.2.5.2 Ground Water Supply

FML plans to use groundwater from the Upper Floral Aquifer (a regionally extensive and well- studied glacial blanket aquifer known locally as the Dalmeny Aquifer) to supply water to the facility. A 24-hour constant rate pumping test and 3D saturated-unsaturated finite-element groundwater flow modelling were used to evaluate water production at the SMPP. Analysis of the pumping test data indicated that the theoretical long-term yield from the primary production well (M2112-38) is 792 m3/d (121 IGPM). Therefore, the 183 IGPM of water required for facility operations will need to be obtained from multiple wells. MDH recommended two wells (with a third backup well) spaced at least 250 m apart.

Effects on groundwater due to the proposed project have been assessed according to baseline data collection, analysis, and detailed modelling studies. Hydrogeologic drilling, instrumentation, and testing programs were combined with numerical modelling to determine impacts to the groundwater in the major aquifers beneath the proposed facility. Analytical models were also

P&E Mining Consultants Inc., Report No. 247 Page 227 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. used to evaluate the potential effects of the injection of the brine solution process water on the deep well injection horizon, the Mannville Group.

20.2.5.3 Waste Characterization

The process residue reporting to the process residue storage facility (PRSF) is planned to consist of two residue streams from the mill operations. The cyanide leach residue is expected to report to the PRSF at a rate of 9 tonnes/hr and the Fe/As precipitate at 5.7 tonnes/hr. Geochemical characterization of the PRSF residue, including solids trace metal content, mineralogical composition, and leach potential, is currently in progress under the direction of the MDH Advanced Testing Laboratory.

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21.0 CAPITAL AND OPERATING COSTS

21.1 CAPITAL COST SUMMARY

21.1.1 Process Plants Capital Cost Summary

The start-up capital cost estimate for the NICO and SMPP Process plants are CAD 167.5 million and CAD 210.1 million respectively. Cost estimates were originally based on second quarter 2010, however approximately 18% of the mechanical equipment budgetary quotes have been updated to first quarter 2012. This initial capital cost assumes that the all-weather road for the NICO site in the North will have been completed before the commencement of its construction. These costs cover the period from the decision to proceed with the Project until the first day of production. The overall accuracy of the estimate is -15%/+15%.

In 2006, as a capital cost offsetting measure, Fortune purchased the Golden Giant Mill in Hemlo, Ontario to utilize the equipment for the NICO project. This facility was dismantled during 2008 and 2009 and the usable equipment has been stored and will be refurbished as required and used at the NICO operation and the Saskatchewan Metals Processing Plant.

The NICO operation will utilize the former Hemlo crushing and screening equipment, conveying equipment, underground ventilation fans and heaters, tanks, pumps, compressors, boilers, mobile generators, overhead cranes, electrical equipment and other miscellaneous equipment. The SMPP will utilize compressors, boilers, flotation cells and tanks, electrical equipment, overhead cranes, mobile equipment, laboratory equipment and other miscellaneous equipment.

A Summary of the initial capital cost is summarized in Table 21.1 below:

TABLE 21.1 CAPITAL COST SUMMARY OF THE PROCESS PLANTS Discipline Description Total, CAD Initial Capital Cost for the NICO Concentrator D0 Mining Not Included D1 Sitework 10,203,634 D2 Concrete 6,867,319 D3 Structural Steel 9,676,847 D4 Architectural 11,637,076 D5 Mechanical 28,491,821 D6 Electrical 13,573,689 D7 Instrumentation 8,068,275 D8 Piping 8,318,778 D9 Mobile Equipment 2,079,198 Dh HVAC 3,296,879 Ia Project Field Indirects 7,204,239 Ib Construction Camp & Catering 2,592,350 Ic Initial Fills 3,031,087 Id Freight 9,132,257 Ie Vendor Representatives 700,600 If Spare Parts 1,336,666 Ig Taxes, Duties, Insurance & Permits Not Included Ih Winter Work 762,291 Ij EPCM 16,352,889 P&E Mining Consultants Inc., Report No. 247 Page 229 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

TABLE 21.1 CAPITAL COST SUMMARY OF THE PROCESS PLANTS Discipline Description Total, CAD Ik Commissioning & Start Up 1,357,246 Il Third Party Engineering 200,000 Im HSE Program And Training 318,240 Is Construction Fuel 3,327,605 K Contingency 18,944,014 O Owner's Cost Not Included

Total Cost for the NICO Concentrator 167,473,000

Initial Capital Cost for SMPP D1 Sitework 8,676,447 D2 Concrete 7,035,331 D3 Structural Steel 10,097,209 D4 Architectural 10,332,860 D5 Mechanical 74,551,380 D6 Electrical 19,843,676 D7 Instrumentation 7,320,830 D8 Piping 17,190,202 D9 Mobile Equipment 1,162,262 Ia Temporary Bldgs & Facilities 3,549,347 Ib Temporary Construction Utilities & Services Not Included Ic Initial Fills 2,426,770 Id Freight 2,668,726 Ie Vendor Representative 781,500 If Spares 2,468,052 Ij EPCM 17,338,981 Ik Commissioning 2,117,035 Im Taxes, Duties, Insurance & Permits Not Included Io Third Party Engineering 555,592 K Contingency 22,015,801 O Owner's Cost Not Included

Total Cost for SMPP 210,132,000

Total Project Cost for Both Process Plants 377,605,000

21.1.2 Capital Cost Summary – Cobalt Sulphate Option

The start-up capital cost estimate for the SMPP Cobalt Sulphate option at the scoping study level was CAD 219.7 million. Cost estimates were originally based on second quarter 2010, however approximately 18% of the mechanical equipment budgetary quotes have been updated to first quarter 2012. These costs cover the period from the decision to proceed with the Project until the first day of production.

The accuracy range of the capital cost estimate is -10% to +25%. The accuracy prediction for the Project takes into account the current state of engineering and procurement. It is also based on

P&E Mining Consultants Inc., Report No. 247 Page 230 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. the fact that budgetary quotes were obtained for all the commodities, labour, and 92% of the Process mechanical equipment of the base SMPP plant.

The sustaining capital costs have been excluded from this estimate. A breakdown of the initial capital costs for the cobalt sulphate option is compared with the base case of the SMPP facility in Table 21.2 below.

TABLE 21.2 CAPITAL COST COMPARISON Description Cobalt Sulphate Option CAD Base Case CAD Mechanical Equipment 78,146,997 77,070,405 Bulk Commodities 60,460,790 54,760,081 Infrastructure 15,453,531 15,453,531 General 8,912,852 8,926,178

Total Direct Cost 162,974,170 156,210,195 Total Indirect Costs 56,115,831 53,921,804 Total Project Cost 219,090,001 210,132,000

The organic reagent used in the solvent extraction is relatively expensive. A price of $40 USD/kg was estimated for budgetary purposes for the OPEX/CAPEX evaluation.

21.1.3 Open Pit Equipment Capital Costs

The projected open pit capital and sustaining capital costs are shown in Table 21.3.

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TABLE 21.3 OPEN PIT EQUIPMENT CAPITAL AND SUSTAINING CAPITAL EXPENDITURES Capital Expenditure Timing Projected Open Pit Capital Expenditure (CAD$) Month 19 13,730k Month 20 1,801k Month 21 3,965k Month 24 63k Month 30 2,654k Month 35 60k Year 6 219k Year 9 1,720k Year 10 2,736k Year 11 1,837k Year 12 6,255k Year 13 538k Year 14 20k Year 16 126k Year 18 24k Year 19 10k

Total 35,758k(1) (1) Projected LOM pit equipment cost if purchased. The economic analysis assumes that open pit equipment would be leased.

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21.1.4 Underground Capital Costs

21.1.4.1 Pre-production Underground Capital Costs

All underground pre-production costs will be capitalized. Pre-production capital costs include contractor mobilization, and the dewatering and rehabilitation of the existing underground facilities. It is estimated that the underground pre-production period will be during May 2016. A summary of 2016 underground pre-production capital costs is presented in Table 21.4.

TABLE 21.4 SUMMARY OF 2016 UNDERGROUND PRE-PRODUCTION CAPITAL COSTS Description Estimated Cost ($) Contractor Mobilization and Dewatering 1,356,248 Diesel Fuel 50,381 Fortunes Indirect Labour 22,362 Fortunes Support Equipment 574,426 Electric Power 171,205

Total Estimated Cost 2,174,623

21.1.4.2 Underground Sustaining Capital Costs

A summary of sustaining capital purchases in 2016 and 2017 is presented in Table 21.5.

TABLE 21.5 SUMMARY OF SUSTAINING CAPITAL COSTS

Description / Year 2016 ($) 2017 ($) Total ($) Contract Mining 31,135,493 1,932,360 33,067,853 Diesel Fuel 1,100,838 49,373 1,150,210 Fortunes Indirect Labour 557,545 91,760 649,306 Fortunes Support Equipment 140,270 27,579 167,849 Electric Power 1,737,018 305,214 2,042,232 Sample Preparation & Assaying 86,518 7,483 94,001

Total Estimated Cost 34,757,683 2,413,768 37,171,451

21.2 OPERATING COST SUMMARY

21.2.1 Concentrator and Southern Metallurgical Processing Plant Operating Cost

The overall operating cost for the NICO Concentrator Plant and Southern Metallurgical Processing Plant is shown in Table 21.6.

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The exchange rate used to convert other currencies to Canadian dollars is CAD 1.00 = 0.92 USD.

TABLE 21.6 ANNUAL CONCENTRATOR AND SOUTHERN METALLURGICAL PROCESSING PLANT OPERATING COST SUMMARY CAD$/ t ore milled Processing Plants 41.51 G&A 10.85

Total 52.36

21.2.1.1 NICO Operating Cost

The overall operating cost for the NICO Concentrator Plant is shown in Table 21.7. The overall annual operating cost is approximately CAD 54.5 million, which equates to CAD$ 32.13/t of ore milled.

TABLE 21.7 ANNUAL NICO PROCESS PLANT OPERATING COST SUMMARY CAD/ t ore milled Processing Plant 24.01 G&A 8.12

Total 32.13(1)(2) (1) Nominal throughput of 1,695,060 dry tonnes per year (2) Reagents, consumables and maintenance costs for tailings from Golder Associates technical memorandum

The overall operating cost for the SMPP Plant is shown in Table 21.8. The overall annual operating cost is approximately CAD$ 34.3 million, which equates to CAD$ 20.23 / t of ore milled.

TABLE 21.8 ANNUAL SMPP PROCESS PLANT OPERATING COST SUMMARY CAD$/ t ore milled Processing Plant 17.50 G & A 2.73

Total 20.23

The operating cost for cobalt sulphate is 10.85% lower than the base case as shown in Table 21.9.

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TABLE 21.9 OPERATING COST SUMMARY With Cobalt Sulphate Base Case Description CAD/t bulk concentrate CAD/t bulk concentrate Labour 139.20 153.71 Power 60.79 71.84 Reagents 172.61 199.93 Maintenance Supplies 67.25 74.59 Infrastructure 8.89 8.89 Other 45.80 45.80

Total 494.55 554.76

It is to be noted that the unit operating costs shown above in Table 21.8 and Table 21.9are based on nominal average 1,695,060 dry t / year of ore milled and nominal 67188 dry t/ year of bulk concentrate processed at the SMPP. The financial model of FML uses the mine schedule as basis for the operating cost calculations.

21.2.2 Open Pit Operating Costs

The open pit mining costs include drilling and blasting, loading and haulage, pit wall pre-shear drilling, grade control sample preparation and assaying, and mine indirect operating costs. The projected open pit operating costs are shown in Table 21.10.The average LOM mining cost is $2.33/t mined.

The production drilling and blasting operating costs were estimated from first principles with input from drill and explosive suppliers. The estimated drilling and blasting costs include blasthole drill operating and maintenance labour costs and fuel, lubricants, filters, parts, and drill string and hammer and bit costs; and explosive, blasting agents and accessories costs including secondary blasting cost allowances; and the explosive supplier‟s plant and equipment and personnel costs. The explosive supplier will load the blastholes. The costs to load and haul ore to the mill crusher, low grade ore to the designated low grade ore stockpile and waste to the co- disposal facility were estimated from first principles with supplier input. The mine indirect operating costs include mine indirect labour costs and mine dewatering, road dressing and stemming rock crushing, geotechnical pit slope inspections, pick-up truck and mine shop operating costs.

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TABLE 21.10 OPEN PIT OPERATING COSTS Months Operating Cost Item 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 Drilling and Blasting ($k) 501 735 769 784 780 777 799 777 785 782 791 790 861 863 859 863 Loading and Hauling ($k) 615 913 944 974 990 968 956 946 978 974 976 985 1,226 1,129 1,144 1,169 Grade control sample preparation and assaying - - 6 25 25 14 54 27 23 27 34 32 14 17 19 19 costs ($k)

Total cost ($k) 1,117 1,649 1,720 1,784 1,795 1,759 1,810 1,752 1,786 1,784 1,802 1,809 2,101 2,010 2,023 2,052 $/t mined 1.72 1.65 1.64 1.70 1.71 1.68 1.72 1.67 1.70 1.70 1.72 1.72 2.00 1.91 1.93 1.95 Running average ($/t rock $1.72 $1.68 $1.66 $1.67 $1.68 $1.68 $1.69 $1.68 $1.69 $1.69 $1.69 $1.69 $1.72 $1.73 $1.75 $1.76 mined)

Months Year 4 Year 5 Years Operating Cost Item 35 36 Q1 Q2 Q3 Q4 Q1 Q2 Q3 Q4 6 7 8 Drilling and Blasting ($k) 870 874 2,595 2,618 2,490 2,479 2,469 2,486 2,596 2,596 6,510 6,446 6,446 Loading and Hauling ($k) 1,163 1,171 3,709 3,506 3,977 3,708 4,012 4,117 3,564 3,320 10,116 9,888 9,846 Grade control sample preparation and assaying 37 36 87 118 98 92 87 101 96 90 402 373 373 costs ($k)

Total cost ($k) 2,071 2,082 6,393 6,242 6,567 6,281 6,570 6,705 6,258 6,007 17,029 16,709 16,666 $/t mined 1.97 1.98 2.03 1.98 2.08 1.99 2.09 2.13 1.99 1.91 2.37 2.32 2.31 Running average ($/t rock mined) 1.77 1.78 1.82 1.84 1.87 1.88 1.90 1.92 1.92 1.92 1.99 2.03 2.06

Operating Cost Years Item 9 10 11 12 13 14 15 16 17 18 19 20 21 22 LOM Drilling and 6,446 5,128 5,128 4,826 5,249 5,028 4,502 4,453 4,426 4,426 3,676 4,274 3,790 280 115,642 Blasting ($k) Loading and 11,096 8,234 7,837 6,751 7,279 8,585 6,710 7,606 7,667 8,046 7,256 8,230 7,374 545 181,222 Hauling ($k) Grade control sample preparation 357 361 373 380 381 387 379 381 373 360 366 402 350 20 7,217 and assaying costs ($k)

Total cost ($k) 17,900 13,724 13,340 11,957 12,910 14,001 11,591 12,441 12,467 12,833 11,300 12,906 11,515 846 304,082 $/t mined 2.49 2.50 2.43 2.17 2.35 2.55 2.58 2.76 2.77 2.85 3.16 2.72 2.97 2.94 2.33 Running average 2.10 2.13 2.15 2.15 2.16 2.18 2.20 2.22 2.24 2.27 2.29 2.31 2.33 2.33 ($/t rock mined)

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21.2.3 Underground Operating Cost

All underground contractor operating costs are capitalized.

21.2.3.1 Fortune’s Labour Costs

A summary of labour rates, by position, for Fortune personnel required for the underground operation is presented in Table 21.11. These labour rates include base salary, planned overtime, travel time allowance and bonus. Burdens are assumed to be 35%.

A summary of total Fortune labour costs is presented in Table 21.12.

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TABLE 21.11 LABOUR RATES FOR UNDERGROUND FORTUNE PERSONNEL Camp Workers in Days per Travel Burden @ Position 4 week Base Salary Planned O/T Bonus Total Total/Yr/Man Year Allowance 35% Period /man Admin Clerk 2.0 182.63 p. y $50,000 $3,750 $4,762 $58,512 $78,991 Staff Mine Mine Engineer 1.0 182.63 p.y. $150,000 $11,250 $14,286 $175,536 $236,973 Staff Drafting Technician 1.0 182.63 p.y. $85,000 $6,375 $8,095 $99,470 $134,285 Mine Geologist 1.0 182.63 p.y. $100,000 $7,500 $9,524 $117,024 $157,982 Geological Technician S1 1.0 182.63 p.y. $85,000 $6,375 $8,095 $99,470 $134,285 Geological Technician S2 1.0 182.63 p.y. $85,000 $6,375 $8,095 $99,470 $134,285

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TABLE 21.12 TOTAL ESTIMATED LABOUR COST

Position / Year 2016 ($) 2017 ($) Total ($)

Administration Clerk $92,156 $16,193 $108,349 Mine Staff Mine Engineer $152,508 $28,831 $181,339 Drafting Technician $86,421 $11,190 $97,612 Mine Geologist $92,156 $13,165 $105,321 Geological Technician S1 $78,333 $11,190 $89,523 Geological Technician S2 $78,333 $11,190 $89,523

Total $579,907 $91,760 $671,667

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22.0 ECONOMIC ANALYSIS

The economic viability of the proposed Project including the development, operation and eventual closure of the NICO open pit, underground mine and the ore processing plant in the NT, bulk concentrate transportation, and bulk concentrate processing at the SMPP has been assessed using an after-tax discounted cash flow based economic analysis.

22.1 COBALT METAL OPTION

The economic analysis was initially conducted for the cobalt metal option where it is assumed that the Project will produce saleable gold doré, 99.8% cobalt cathode (“cobalt metal‟), 99.99% bismuth ingot, and copper metal precipitate products. The results of the economic analysis of the cobalt metal option based on five metal prices and US$:C$ exchange rate scenarios are shown in Table 22.1. Currencies are reported in Canadian dollars unless indicated otherwise.

The results of the economic analysis of the cobalt metal option for the base case metal prices and exchange rate scenario (US$1,450/oz Au, US$20/lb Co, US$11/lb Bi, US$3.50/lb Cu, US$0.95=$C1.00) show that the Project offers:

 An after-tax IRR of 9.6%  An after tax NPV (7%) of $101 M, after tax NPV (5%) of $207 M  Payback in about 6.8 operating years.

TABLE 22.1 (1) RESULTS OF THE ECONOMIC ANALYSIS FOR THE COBALT METAL OPTION Metal Price Pre-Tax After-Tax & Exchange $M $M $M $M IRR % IRR % Rate Case NPV (7%) NPV (5%) NPV (7%) NPV (5%) Base Case 10.8 164.5 293.2 9.6 101.0 207.1 Prices 3-yr Trailing Average 7.4 17.1 114.6 6.6 (15.3) 69.0 Prices Current Prices 7.1 2.1 99.7 6.2 (30.6) 53.4 Escalated 13.9 315.2 477.8 12.3 214.9 344.7 Prices Optimistic 18.3 539.5 749.8 16.3 387.5 551.3 Prices (1) The estimated capital and operating costs in the cobalt metal option have an accuracy of ±15%. See Table 22.3 and Section 25 for additional information.

22.2 COBALT SULPHATE OPTION

The cobalt sulphate option is similar to the cobalt metal option except that the Project would produce cobalt sulphate heptahydrate containing 20.9% cobalt (“cobalt sulphate” product) instead of cobalt cathode. This option is of interest due to the projected premium prices for cobalt sulphate. The results of the economic analysis of the cobalt sulphate option for the base case metal prices and exchange rate scenario (US$1,450/oz Au, US$20/lb Co, US$11/lb Bi, US$3.50/lb Cu, US$0.95=$C1.00) indicate that the Project offers:

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 An estimated after-tax IRR of 12.4%  An estimated after tax NPV(7%) of $212M, after tax NPV(5%) of $338M  Payback in 5.9 operating years (see also paragraph below).

The capital and operating cost estimates for the overall process plant with the cobalt sulphate option have an accuracy of -10% to +25%. This is a lower accuracy than the cobalt metal case as the cobalt sulphate circuit is estimated at a scoping level. The results of the economic analysis of the cobalt sulphate option indicate that this option is of definite interest. As such, it is recommended that further engineering for the cobalt sulphate circuit be carried out along with supporting metallurgical testwork to improve the overall accuracy of the cobalt sulphate option cost estimates to bring the overall accuracy of the cobalt sulphate option cost estimate to ±15% and that the economic analysis model then be re-run before deciding to proceed with this option. See Sections 25 and 26 for additional information.

The results of the economic analysis of the cobalt sulphate option for the base case and four other metal price and exchange rate scenarios based on currently available cost estimates are shown in Table 22.2.

TABLE 22.2 (1) RESULTS OF THE ECONOMIC ANALYSIS FOR THE COBALT SULPHATE OPTION Metal Price Pre-Tax After-Tax & Exchange $M $M $M $M IRR % IRR % Rate Case NPV (7%) NPV (5%) NPV (7%) NPV (5%) Base Case 14.0 308.5 466.0 12.4 212.6 338.7 Prices 3-yr Trailing Average 10.5 146.8 270.0 9.3 86.7 188.4 Prices Current 9.6 109.5 228.2 8.5 57.6 156.8 Prices Escalated 17.1 467.1 660.1 15.2 332.4 483.7 Prices Optimistic 21.6 707.0 951.1 19.3 514.5 702.3 Prices (1) The projected capital and operating costs in the cobalt sulphate option have an overall accuracy of -10% to +25%. See Table 22.3 and Sections 25 and 26 for additional information.

The economic parameters and key assumptions for the cobalt metal option and the cobalt sulphate option are shown in Table 22.3.

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TABLE 22.3 ECONOMIC PARAMETERS AND KEY ASSUMPTIONS Mineral Reserves The economic analysis is based on the updated mineral reserve estimate presented in Section 15. Mine Production Schedule The economic analysis is based on the mine development and production schedules shown in Section 16. Open Pit Tonnages Diluted ore tonnes 32,611,516 t Tonnes of waste 97,809,571 t Underground Mine Tonnages Diluted ore tonnes 376,044 t Tonnes of waste 68,524 t Mill Feed Tonnage And Average Grades Mill throughout 4,650 tpd Operating life 19.8 years Tonnes milled 32.987 Mt LOM average gold 1.023 g Au / t grade LOM average 0.1131% Co cobalt grade LOM average 0.1403% Bi bismuth grade LOM average 0.0374% Cu copper grade Gold grade variable recovery ranges from 56% to 85% Gold recovery 73.7% weight average for 5m x 5m x 5 m blocks. Cobalt recovery 84.4% Bismuth recovery 72.2% Copper recovery 41.2% Gold doré, 99.8% cobalt cathode (“cobalt metal‟) and/or cobalt sulphate heptahydrate Saleable metal containing 29% cobalt (“cobalt sulphate”), 99.99% bismuth ingot, and a copper metal products precipitate. Projected Metal Prices and US$:C$ Exchange Rates Current Price 3-year Trailing Base Case Escalated Optimistic Price Parameter Average Prices (as of Case (as of May 31, Price Case Case May 31, 2012) case 2012) Projected LOM Metal Production Gold 800,091 troy oz Au Cobalt 69,471,792 lbs Co Bismuth 73,636,507 lbs Bi Copper 11,187,952 lbs Cu Projected Metal Prices Gold 1,450 1,359.94 1,558 1,800 2,000 (US$/troy oz Au) Cobalt 20* 18.53* 15.23* 22.50* 25.00* (US$/lb Co) Bismuth 11 9.83 10.55 12.5 15.00 (US$/lb Bi) Copper 3.50 3.51 3.40 4.00 4.50 (US$/lb Cu) US$=C$1 0.95 0.98 0.97 1.00 1.00 * The projected cobalt sulphate price is 22% higher than the projected cobalt metal price. Projected Operating Costs Operating cost per tonne Area LOM operating cost Operating cost per tonne mined milled Open pit mining $282,810k $2.17 $8.67 Underground $37,351k $84.02 $99.34 P&E Mining Consultants Inc., Report No. 247 Page 242 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

TABLE 22.3 ECONOMIC PARAMETERS AND KEY ASSUMPTIONS mining $320,161k $2.45 $9.71 Total mining Milling $531,193k $16.10 Concentrate $238,038k $7.22 transport $679,218k $20.59 SMPP $1,448,449k $43.91 Total Shared Services $275,852k $8.36 (G&A)

Total $2,044,462k $61.97 Projected Capital Costs Area Projected capital cost Open pit $7,003k Underground mine $2,175k NWT Plant CapEx $182,042k SASK CapEx $208,090k Contingency $41,211k

Total capital $440,521k Sustaining capital $113,588k Accuracy of Estimates and Cost Basis The estimated costs for the cobalt metal option were developed to an accuracy of ±15%. The estimated costs for the cobalt sulphate solvent extraction option were developed to an overall accuracy of -10% to +25% with cost estimates for the cobalt sulphate circuit having scoping level accuracy (e.g. ±40%). The estimated NICO Concentrator and SMPP capital and Accuracy of operating costs are based on Q2, 2010 costs, however approximately 18% of the mechanical estimates equipment budgetary quotes have been updated to Q1 2012. The estimated open pit mining, underground mining costs and shared services costs are based on Q2, 2012 costs. The cost of procuring three used cone crushers, 1500 kVA generator set, air compressors, two boilers, and selected conveyors and electrical components included in the mill that Fortune Minerals purchased from the Golden Giant Mine at Hemlo are treated as sunk costs. It is assumed that an all-weather access road to the NICO site will be built before the Other assumptions commencement of the main construction and available over the life of the project. It is also assumed that the NICO and SMPP project will be developed as scheduled.

22.3 ECONOMIC ANALYSIS

The base case cobalt metal cashflow model including projected royalties and taxes is shown in Table 22.4.

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TABLE 22.4 COBALT METAL, BASE CASE METAL PRICES-EXCHANGE RATE CASHFLOW MODEL LOM 1-Jan-14 1-Feb-14 1-Mar-14 1-Apr-14 1-May-14 1-Jun-14 1-Jul-14 1-Aug-14 1-Sep-14 1-Oct-14 1-Nov-14 1-Dec-14 Unit Total 31-Jan-14 28-Feb-14 31-Mar-14 30-Apr-14 31-May-14 30-Jun-14 31-Jul-14 31-Aug-14 30-Sep-14 31-Oct-14 30-Nov-14 31-Dec-14 Projected metal production:

Gold doré troy oz Au 800,091 ------Cobalt Cathode (99.997%) pound 69,471,792 ------Bismuth Cathode (99.5%) pound 73,636,507 ------Copper Cathode (xx%) pound 11,187,952 ------Metal sales: Gold C$ M 1,221.19 ------Cobalt C$ M 1,462.56 ------Bismuth C$ M 852.63 ------Copper C$ M 29.68 ------Total metal sales C$ M 3,566.07 ------

Selling costs: C$ M 7.33 ------

Net revenue C$ M 3,558.74 ------

Operating costs Open Pit Mining C$ M 282.81 ------Underground Mining C$ M 37.35 ------Milling C$ M 531.19 ------Shared services incl G&A itemsC$ M 275.85 ------Transportation of Concentrate C$ M 238.04 ------SMPP C$ M 679.22 ------Total operating costs C$ M 2,044.46 ------

Reclamation Costs (Accretion) C$ M 30.97 - - 0.04 0.04 0.04 0.04 0.04 0.04 0.04 0.04 0.04 0.04 Royalties C$ M 51.54 ------Corporate administration C$ M 25.25 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 EBITA C$ M 1,406.52 (0.09) (0.09) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13)

Working capital adjustments C$ M 1.88 1.88 - (0.02) 0.00 (0.00) 0.00 (0.00) - 0.00 (0.00) 0.00 (0.00) Earmings before taxes C$ M 1,404.64 (1.97) (0.09) (0.12) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13)

Depreciations and amortizationC$ M 643.47 ------Corporate taxes C$ M 193.02 ------Earnings after tax C$ M 1,211.62 (1.97) (0.09) (0.12) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13) (0.13)

Capital costs: Open pit mine C$ M 7.00 ------Underground mine C$ M 2.17 ------Nico mill C$ M 182.04 - 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 SMPP C$ M 208.09 - 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 Contingency C$ M 41.21 - 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 Initial capital expenditures C$ M 440.52 - 17.25 17.25 17.25 17.25 17.25 17.25 17.25 17.25 17.25 17.25 17.25 Projected sustaining capital expendituresC$ M 113.59 ------

Projected after-tax cashflow C$ M 657.51 (1.97) (17.35) (17.37) (17.39) (17.39) (17.39) (17.39) (17.39) (17.39) (17.39) (17.39) (17.39)

IRR = 9.6% 7% NPV = $ 101 Million

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LOM 1-Jan-15 1-Feb-15 1-Mar-15 1-Apr-15 1-May-15 1-Jun-15 1-Jul-15 1-Aug-15 1-Sep-15 1-Oct-15 1-Nov-15 1-Dec-15

Unit Total 31-Jan-15 28-Feb-15 31-Mar-15 30-Apr-15 31-May-15 30-Jun-15 31-Jul-15 31-Aug-15 30-Sep-15 31-Oct-15 30-Nov-15 31-Dec-15 Projected metal production: Gold doré troy oz Au 800,091 ------513 780 966 Cobalt Cathode (99.997%) pound 69,471,792 ------189,720 288,695 330,939 Bismuth Cathode (99.5%) pound 73,636,507 ------117,528 178,841 257,471 Copper Cathode (xx%) pound 11,187,952 ------6,227 9,475 12,859 Metal sales: Gold C$ M 1,221.19 ------Cobalt C$ M 1,462.56 ------Bismuth C$ M 852.63 ------Copper C$ M 29.68 ------Total metal sales C$ M 3,566.07 ------

Selling costs: C$ M 7.33 ------

Net revenue C$ M 3,558.74 ------

Operating costs Open Pit Mining C$ M 282.81 ------3.60 1.96 1.97 1.94 Underground Mining C$ M 37.35 ------Milling C$ M 531.19 ------2.23 2.21 2.23 Shared services incl G&A itemsC$ M 275.85 ------0.68 0.66 0.68 Transportation of Concentrate C$ M 238.04 ------0.52 0.80 1.02 SMPP C$ M 679.22 ------Total operating costs C$ M 2,044.46 ------3.60 5.38 5.64 5.87

Reclamation Costs (Accretion) C$ M 30.97 0.05 0.04 0.05 0.04 0.05 0.04 0.05 0.05 0.04 0.05 0.04 0.05 Royalties C$ M 51.54 ------Corporate administration C$ M 25.25 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 EBITA C$ M 1,406.52 (0.14) (0.13) (0.14) (0.14) (0.14) (0.14) (0.14) (0.14) (3.74) (5.52) (5.77) (6.01)

Working capital adjustments C$ M 1.88 (0.00) 0.00 (0.00) 0.00 (0.00) 0.00 (0.00) - (1.45) (1.03) (0.39) (0.14) Earmings before taxes C$ M 1,404.64 (0.14) (0.13) (0.14) (0.14) (0.14) (0.14) (0.14) (0.14) (2.29) (4.49) (5.38) (5.87)

Depreciations and amortizationC$ M 643.47 ------0.09 1.13 1.67 2.05 Corporate taxes C$ M 193.02 ------Earnings after tax C$ M 1,211.62 (0.14) (0.13) (0.14) (0.14) (0.14) (0.14) (0.14) (0.14) (2.29) (4.49) (5.38) (5.87)

Capital costs: Open pit mine C$ M 7.00 - - - - 3.12 0.91 1.62 1.36 - - - - Underground mine C$ M 2.17 ------Nico mill C$ M 182.04 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 7.28 SMPP C$ M 208.09 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 8.32 Contingency C$ M 41.21 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 1.65 Initial capital expenditures C$ M 440.52 17.25 17.25 17.25 17.25 20.37 18.16 18.87 18.61 17.25 17.25 17.25 17.25 Projected sustaining capital expendituresC$ M 113.59 ------0.58 0.58 0.68 0.65

Projected after-tax cashflow C$ M 657.51 (17.39) (17.39) (17.39) (17.39) (20.51) (18.30) (19.01) (18.75) (20.13) (22.33) (23.32) (23.77)

IRR = 9.6% 7% NPV = $ 101 Million

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LOM 1-Jan-16 1-Feb-16 1-Mar-16 1-Apr-16 1-May-16 1-Jun-16 1-Jul-16 1-Aug-16 1-Sep-16 1-Oct-16 1-Nov-16 1-Dec-16

Unit Total 31-Jan-16 29-Feb-16 31-Mar-16 30-Apr-16 31-May-16 30-Jun-16 31-Jul-16 31-Aug-16 30-Sep-16 31-Oct-16 30-Nov-16 31-Dec-16 Projected metal production: Gold doré troy oz Au 800,091 800 811 669 556 554 685 1,223 6,314 9,584 5,809 11,932 9,569 Cobalt Cathode (99.997%) pound 69,471,792 301,980 302,527 307,264 295,525 310,200 335,207 367,852 310,019 286,913 373,663 215,477 172,144 Bismuth Cathode (99.5%) pound 73,636,507 345,825 343,488 315,144 360,517 319,461 439,884 382,538 441,764 429,133 511,261 283,154 231,301 Copper Cathode (xx%) pound 11,187,952 11,821 11,840 17,603 10,060 16,784 11,670 26,208 41,213 10,549 17,800 11,237 16,339 Metal sales: Gold C$ M 1,221.19 - - 6.23 1.40 0.90 1.03 1.79 8.92 14.07 9.38 17.36 14.91 Cobalt C$ M 1,462.56 - - 12.08 17.98 11.36 8.66 8.23 7.21 6.42 7.38 5.60 4.49 Bismuth C$ M 852.63 - - - 10.92 7.50 6.19 5.42 5.27 5.03 5.56 4.35 3.59 Copper C$ M 29.68 - - - 0.05 0.08 0.06 0.07 0.08 0.06 0.06 0.04 0.05 Total metal sales C$ M 3,566.07 - - 18.31 30.35 19.84 15.94 15.50 21.48 25.57 22.38 27.34 23.03

Selling costs: C$ M 7.33 - - 0.04 0.01 0.01 0.01 0.01 0.05 0.08 0.05 0.10 0.09

Net revenue C$ M 3,558.74 - - 18.28 30.34 19.83 15.93 15.49 21.43 25.49 22.32 27.24 22.94

Operating costs Open Pit Mining C$ M 282.81 1.95 1.92 1.96 1.95 1.96 1.97 2.00 1.91 1.91 1.94 1.94 1.95 Underground Mining C$ M 37.35 - - - - - 4.09 4.57 5.77 6.29 6.34 5.55 2.42 Milling C$ M 531.19 2.23 2.19 2.23 2.21 2.23 1.95 1.97 1.97 1.95 1.97 1.95 1.97 Shared services incl G&A itemsC$ M 275.85 0.68 0.65 0.68 0.66 0.68 0.65 0.67 0.67 0.65 0.67 0.65 0.67 Transportation of Concentrate C$ M 238.04 1.02 1.02 1.02 1.02 1.02 1.03 1.02 1.02 1.02 1.04 1.02 0.82 SMPP C$ M 679.22 - - 2.92 2.83 2.92 2.83 2.92 2.92 2.83 2.92 2.83 2.92 Total operating costs C$ M 2,044.46 5.88 5.77 8.81 8.68 8.81 12.51 13.14 14.25 14.65 14.87 13.94 10.75

Reclamation Costs (Accretion) C$ M 30.97 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 0.05 Royalties C$ M 51.54 ------Corporate administration C$ M 25.25 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 0.09 EBITA C$ M 1,406.52 (6.02) (5.90) 9.32 21.52 10.88 3.28 2.21 7.03 10.71 7.31 13.16 12.06

Working capital adjustments C$ M 1.88 (0.14) (0.13) 7.44 14.63 (8.27) (4.68) (1.63) (0.92) (0.56) 0.37 (1.23) 0.25 Earmings before taxes C$ M 1,404.64 (5.88) (5.77) 1.89 6.89 19.14 7.96 3.84 7.95 11.27 6.94 14.39 11.80

Depreciations and amortizationC$ M 643.47 2.41 2.31 2.32 2.34 2.41 3.55 4.23 5.30 5.90 6.23 6.42 4.34 Corporate taxes C$ M 193.02 ------Earnings after tax C$ M 1,211.62 (5.88) (5.77) 1.89 6.89 19.14 7.96 3.84 7.95 11.27 6.94 14.39 11.80

Capital costs: Open pit mine C$ M 7.00 ------Underground mine C$ M 2.17 1.09 1.09 ------Nico mill C$ M 182.04 7.28 7.28 ------SMPP C$ M 208.09 8.32 8.32 ------Contingency C$ M 41.21 1.65 1.65 ------Initial capital expenditures C$ M 440.52 18.34 18.34 ------Projected sustaining capital expendituresC$ M 113.59 0.59 0.59 0.61 0.59 0.59 1.53 0.62 0.62 0.62 0.62 0.68 0.62

Projected after-tax cashflow C$ M 657.51 (24.81) (24.70) 1.27 6.30 18.56 6.42 3.22 7.33 10.65 6.32 13.71 11.18

IRR = 9.6% 7% NPV = $ 101 Million

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LOM 1-Jan-17 1-Apr-17 1-Jul-17 1-Oct-17

Unit Total 31-Mar-17 30-Jun-17 30-Sep-17 31-Dec-17 Projected metal production: Gold doré troy oz Au 800,091 7,401 1,718 1,729 2,058 Cobalt Cathode (99.997%) pound 69,471,792 939,518 881,155 1,071,818 1,017,269 Bismuth Cathode (99.5%) pound 73,636,507 936,325 879,111 888,923 750,360 Copper Cathode (xx%) pound 11,187,952 32,499 74,835 93,851 144,071 Metal sales: Gold C$ M 1,221.19 12.36 2.95 2.65 3.12 Cobalt C$ M 1,462.56 18.48 18.58 21.87 21.50 Bismuth C$ M 852.63 10.70 10.34 10.34 9.11 Copper C$ M 29.68 0.10 0.17 0.22 0.33 Total metal sales C$ M 3,566.07 41.63 32.04 35.07 34.06

Selling costs: C$ M 7.33 0.08 0.02 0.02 0.02

Net revenue C$ M 3,558.74 41.55 32.02 35.05 34.04

Operating costs Open Pit Mining C$ M 282.81 6.01 5.80 6.00 5.77 Underground Mining C$ M 37.35 2.01 0.33 - - Milling C$ M 531.19 5.54 5.79 5.82 5.82 Shared services incl G&A itemsC$ M 275.85 2.21 2.25 2.27 2.27 Transportation of Concentrate C$ M 238.04 3.06 3.06 3.07 3.06 SMPP C$ M 679.22 8.43 8.51 8.60 8.60 Total operating costs C$ M 2,044.46 27.26 25.74 25.75 25.52

Reclamation Costs (Accretion) C$ M 30.97 0.16 0.16 0.16 0.16 Royalties C$ M 51.54 - - - - Corporate administration C$ M 25.25 0.27 0.27 0.27 0.27 EBITA C$ M 1,406.52 13.87 5.84 8.87 8.09

Working capital adjustments C$ M 1.88 1.26 (0.02) 0.70 (0.37) Earmings before taxes C$ M 1,404.64 12.61 5.86 8.16 8.46

Depreciations and amortizationC$ M 643.47 8.89 7.40 7.32 7.23 Corporate taxes C$ M 193.02 - - - - Earnings after tax C$ M 1,211.62 12.61 5.86 8.16 8.46

Capital costs: Open pit mine C$ M 7.00 - - - - Underground mine C$ M 2.17 - - - - Nico mill C$ M 182.04 - - - - SMPP C$ M 208.09 - - - - Contingency C$ M 41.21 - - - - Initial capital expenditures C$ M 440.52 - - - - Projected sustaining capital expendituresC$ M 113.59 1.79 1.86 1.84 3.73

Projected after-tax cashflow C$ M 657.51 10.82 4.00 6.32 4.73

IRR = 9.6% 7% NPV = $ 101 Million

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LOM 1-Jan-18 1-Apr-18 1-Jul-18 1-Oct-18

Unit Total 31-Mar-18 30-Jun-18 30-Sep-18 31-Dec-18 Projected metal production: Gold doré troy oz Au 800,091 2,869 4,378 3,977 3,243 Cobalt Cathode (99.997%) pound 69,471,792 926,968 893,116 926,079 987,569 Bismuth Cathode (99.5%) pound 73,636,507 731,526 726,199 777,112 871,746 Copper Cathode (xx%) pound 11,187,952 120,048 113,824 124,489 144,294 Metal sales: Gold C$ M 1,221.19 4.33 6.61 6.09 4.99 Cobalt C$ M 1,462.56 19.80 19.03 19.45 20.54 Bismuth C$ M 852.63 8.58 8.48 8.89 9.79 Copper C$ M 29.68 0.32 0.31 0.32 0.36 Total metal sales C$ M 3,566.07 33.03 34.42 34.76 35.68

Selling costs: C$ M 7.33 0.03 0.04 0.04 0.03

Net revenue C$ M 3,558.74 33.00 34.38 34.72 35.65

Operating costs Open Pit Mining C$ M 282.81 5.88 5.98 6.02 5.74 Underground Mining C$ M 37.35 - - - - Milling C$ M 531.19 5.76 5.79 5.82 5.82 Shared services incl G&A itemsC$ M 275.85 2.22 2.25 2.27 2.27 Transportation of Concentrate C$ M 238.04 3.07 3.06 3.06 3.07 SMPP C$ M 679.22 8.43 8.51 8.60 8.60 Total operating costs C$ M 2,044.46 25.37 25.60 25.77 25.50

Reclamation Costs (Accretion) C$ M 30.97 0.17 0.17 0.17 0.17 Royalties C$ M 51.54 - - - - Corporate administration C$ M 25.25 0.27 0.27 0.27 0.27 EBITA C$ M 1,406.52 7.19 8.34 8.50 9.70

Working capital adjustments C$ M 1.88 (0.42) (0.22) 0.15 0.55 Earmings before taxes C$ M 1,404.64 7.61 8.56 8.35 9.16

Depreciations and amortizationC$ M 643.47 7.26 7.45 7.57 7.58 Corporate taxes C$ M 193.02 - - - - Earnings after tax C$ M 1,211.62 7.61 8.56 8.35 9.16

Projected after-tax cashflow C$ M 657.51 5.49 6.75 1.45 7.34

IRR = 9.6% 7% NPV = $ 101 Million P&E Mining Consultants Inc., Report No. 247 Page 248 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

LOM 1-Jan-19 1-Jan-20 1-Jan-21 1-Jan-22 1-Jan-23 1-Jan-24 1-Jan-25 1-Jan-26 1-Jan-27

Unit Total 31-Dec-19 31-Dec-20 31-Dec-21 31-Dec-22 31-Dec-23 31-Dec-24 31-Dec-25 31-Dec-26 31-Dec-27 Projected metal production: Gold doré troy oz Au 800,091 10,513 19,947 45,465 52,936 90,040 51,417 18,924 19,448 17,658 Cobalt Cathode (99.997%) pound 69,471,792 4,086,118 3,977,380 3,586,066 3,462,063 2,780,495 3,999,657 3,963,636 3,864,741 3,724,753 Bismuth Cathode (99.5%) pound 73,636,507 3,749,568 4,548,892 5,017,567 4,835,535 4,519,980 3,331,389 4,073,441 4,663,226 4,723,929 Copper Cathode (xx%) pound 11,187,952 616,712 790,065 654,368 503,609 335,076 281,853 358,563 511,347 786,228 Metal sales: Gold C$ M 1,221.19 16.08 30.32 69.06 80.70 136.95 78.98 29.31 29.68 26.98 Cobalt C$ M 1,462.56 85.79 83.86 75.94 73.05 59.32 82.87 83.40 81.47 78.58 Bismuth C$ M 852.63 43.06 51.92 57.60 56.12 52.64 39.69 46.57 53.41 54.60 Copper C$ M 29.68 1.61 2.04 1.77 1.39 0.95 0.77 0.93 1.31 2.00 Total metal sales C$ M 3,566.07 146.54 168.15 204.37 211.25 249.85 202.31 160.21 165.87 162.15

Selling costs: C$ M 7.33 0.11 0.19 0.41 0.48 0.79 0.47 0.19 0.19 0.18

Net revenue C$ M 3,558.74 146.43 167.95 203.95 210.77 249.06 201.85 160.02 165.68 161.97

Operating costs Open Pit Mining C$ M 282.81 15.21 15.11 15.06 16.12 12.83 11.95 12.08 12.44 12.56 Underground Mining C$ M 37.35 ------Milling C$ M 531.19 27.31 27.36 27.31 27.31 27.31 27.36 27.31 27.31 27.31 Shared services incl G&A itemsC$ M 275.85 14.94 14.98 14.94 14.94 14.94 14.98 14.94 14.94 14.94 Transportation of Concentrate C$ M 238.04 12.26 12.26 12.26 12.26 12.26 12.25 12.26 12.26 12.26 SMPP C$ M 679.22 34.22 34.31 34.22 34.22 34.22 34.31 34.22 34.22 34.22 Total operating costs C$ M 2,044.46 103.94 104.02 103.80 104.85 101.56 100.85 100.81 101.18 101.29

Reclamation Costs (Accretion) C$ M 30.97 0.75 0.82 0.90 0.98 1.06 1.16 1.26 1.38 1.50 Royalties C$ M 51.54 - - - - - 5.55 10.00 5.45 5.29 Corporate administration C$ M 25.25 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 EBITA C$ M 1,406.52 40.64 62.02 98.16 103.85 145.33 93.18 46.84 56.58 52.80

Working capital adjustments C$ M 1.88 0.46 0.77 0.33 (0.30) (0.53) (0.44) 0.04 0.63 (0.03) Earmings before taxes C$ M 1,404.64 40.18 61.25 97.83 104.14 145.86 93.63 46.79 55.94 52.83

Depreciations and amortizationC$ M 643.47 30.66 30.76 30.79 31.62 31.73 31.80 32.63 32.74 33.67 Corporate taxes C$ M 193.02 - - - - - 19.13 8.91 11.86 11.05 Earnings after tax C$ M 1,211.62 40.18 61.25 97.83 104.14 145.86 74.50 37.88 44.08 41.78

Capital costs: Open pit mine C$ M 7.00 ------Underground mine C$ M 2.17 ------Nico mill C$ M 182.04 ------SMPP C$ M 208.09 ------Contingency C$ M 41.21 ------Initial capital expenditures C$ M 440.52 ------Projected sustaining capital expendituresC$ M 113.59 7.17 5.82 0.72 9.46 3.38 2.17 8.44 0.92 7.70

Projected after-tax cashflow C$ M 657.51 33.01 55.43 97.10 94.68 142.48 72.32 29.44 43.16 34.08

IRR = 9.6% 7% NPV = $ 101 Million

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LOM 1-Jan-28 1-Jan-29 1-Jan-30 1-Jan-31 1-Jan-32 1-Jan-33 1-Jan-34 1-Jan-35 1-Jan-36

Unit Total 31-Dec-28 31-Dec-29 31-Dec-30 31-Dec-31 31-Dec-32 31-Dec-33 31-Dec-34 31-Dec-35 31-Dec-36 Projected metal production: Gold doré troy oz Au 800,091 19,334 26,370 36,841 68,030 100,934 41,707 94,787 7,600 - Cobalt Cathode (99.997%) pound 69,471,792 3,679,012 3,481,136 3,308,041 2,688,857 3,462,421 2,528,888 3,693,742 1,153,170 - Bismuth Cathode (99.5%) pound 73,636,507 4,637,590 4,003,832 3,748,694 3,088,453 2,581,323 2,761,020 1,345,878 487,578 - Copper Cathode (xx%) pound 11,187,952 978,182 1,181,553 999,440 524,566 288,775 800,302 412,384 85,332 - Metal sales: Gold C$ M 1,221.19 29.49 40.16 56.09 103.43 153.63 64.42 143.99 12.83 - Cobalt C$ M 1,462.56 77.53 73.50 69.85 57.32 72.06 54.25 76.49 28.65 - Bismuth C$ M 852.63 53.78 46.93 43.68 36.39 30.41 31.84 16.87 7.09 - Copper C$ M 29.68 2.53 3.06 2.70 1.54 0.85 1.98 1.19 0.38 - Total metal sales C$ M 3,566.07 163.32 163.66 172.33 198.68 256.96 152.48 238.55 48.95 -

Selling costs: C$ M 7.33 0.19 0.25 0.34 0.60 0.89 0.39 0.83 0.10 -

Net revenue C$ M 3,558.74 163.13 163.41 171.99 198.08 256.07 152.10 237.71 48.86 -

Operating costs Open Pit Mining C$ M 282.81 11.05 11.04 11.17 11.52 10.08 11.65 10.45 2.45 - Underground Mining C$ M 37.35 ------Milling C$ M 531.19 27.36 27.31 27.31 27.31 27.36 27.31 27.31 16.44 - Shared services incl G&A itemsC$ M 275.85 14.98 14.94 14.94 14.94 14.98 14.94 14.94 8.67 - Transportation of Concentrate C$ M 238.04 12.26 12.26 12.25 12.26 12.26 12.26 12.26 2.96 - SMPP C$ M 679.22 34.31 34.22 34.22 34.22 34.31 34.22 34.22 34.14 - Total operating costs C$ M 2,044.46 99.96 99.78 99.90 100.26 98.98 100.39 99.18 64.66 -

Reclamation Costs (Accretion) C$ M 30.97 1.64 1.79 1.95 2.12 2.31 2.52 2.75 3.22 - Royalties C$ M 51.54 5.89 6.09 6.07 7.21 - - - - - Corporate administration C$ M 25.25 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 1.10 EBITA C$ M 1,406.52 54.55 54.66 62.98 87.39 153.68 48.09 134.69 (20.12) (1.10)

Working capital adjustments C$ M 1.88 (0.03) (0.62) (0.39) (1.10) 0.91 (1.51) 0.58 (2.67) 0.15 Earmings before taxes C$ M 1,404.64 54.57 55.28 63.37 88.49 152.78 49.60 134.10 (17.45) (1.24)

Depreciations and amortizationC$ M 643.47 33.52 33.57 33.70 33.75 33.59 33.96 33.80 7.77 - Corporate taxes C$ M 193.02 11.82 12.26 14.78 21.51 40.36 12.28 35.43 (5.56) (0.81) Earnings after tax C$ M 1,211.62 42.75 43.02 48.60 66.98 112.42 37.31 98.67 (11.89) (0.43)

Projected after-tax cashflow C$ M 657.51 42.08 42.56 47.74 66.62 112.07 33.57 98.33 (15.61) (25.09)

IRR = 9.6% 7% NPV = $ 101 Million

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22.4 PROJECTED OPERATING CASH COSTS

The projected cash operating costs for the cobalt metal option and cobalt sulphate option are shown in Table 22.5 and Table 22.6.

TABLE 22.5 PROJECTED OPERATING CASH COSTS FOR THE COBALT METAL OPTION Cash Cost Metal Price – Cash Cost Net of By-Product Credits Equivalent Oz Au Exchange Rate Gold Gold Bismuth Cases Cobalt $US/lb $US/equivalent oz US$/oz $US/lb Base Case 831.30A (356.70) (0.81) (8.63) 3-Year Trailing 859.94 (77.23) 1.98 (5.79) Average Current 990.44 142.52 (1.07) (4.83) Escalated 943.87 (551.70) (4.58) (13.05)

TABLE 22.6 PROJECTED OPERATING CASH COSTS FOR THE COBALT SULPHATE OPTION Cash Cost Metal Price – Cash Cost Net of By-Product Credits Equivalent Gold Oz Exchange Rate Gold Gold Bismuth Cases Cobalt $US/lb $US/equivalent oz US$/oz $US/lb Base Case 762.50 (738.75) (0.81) (12.78) 3-Year Trailing 788.54 (431.20) 1.98 (9.63) Average Current 921.45 (148.42) (1.07) (7.99) Escalated 868.38 (981.51) (4.58) (17.72)

22.5 SENSITIVITY ANALYSIS

A sensitivity analysis of the cobalt metal option base case metal prices and exchange rate (US$1,450/oz Au, US$20/lb Co, US$11/lb Bi, US$3.50/lb Cu, US$0.95=$C1.00) was conducted for the following cashflow model variables: gold price, cobalt price, bismuth price, copper price, and the four metal (i.e. Au, Co, Bi, Cu) prices; project operating costs; and project capital costs. Each variable was independently increased / decreased by ±10% and ± 15%. The results of the sensitivity analysis for the cobalt metal option base case price-exchange rate scenario are shown in Table 22.7 and Figure 22.1. The results of the sensitivity analysis indicate that the Project is most sensitive to concurrent changes in all four metal prices. The Project is relatively sensitive to operating cost changes and less sensitive to changes in capital costs.

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TABLE 22.7 SENSITIVITY ANALYSIS (COBALT METAL OPTION, BASE CASE METAL PRICES-EXCHANGE RATE) Projected after tax NPV (7%) (C$ M) Variable Variation in parameter -15% -10% 0% +10% +15% Gold price only 44.9 64.1 101 137.6 155.9 Cobalt price only 24.3 50 101 151.3 176.5 Bismuth price only 56.4 71.6 101 130.4 145.1 Copper price only 99.6 100 101 101.9 102.3 All 4 metals change (85.3) (20.3) 101 218.9 276.8 Operating costs 249.6 200 101 1.9 (47.6) Capital costs 162.4 141.9 101 60 39.5

Figure 22.1 Sensitivity Analysis (Cobalt Metal Option, Base Case Metal Prices-Exchange Rate)

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23.0 ADJACENT PROPERTIES

There are no properties of significance adjacent to the NICO Property.

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24.0 OTHER RELEVANT DATA AND INFORMATION

24.1 OPTION TO PRODUCE COBALT SULPHATE

Fortune Minerals Limited engaged Aker Solutions to prepare and evaluate the option to produce cobalt sulphate heptahydrate (CoSO4.7H2O) as an alternative to producing cobalt cathodes by electrowinning (Base Case in the FEED report). Capital and operating costs (“CAPEX”) and (“OPEX”) for this option were prepared to an overall accuracy of -10% / + 25%, whereas , the Base Case was prepared to an accuracy of +/- 15%.

SGS Lakefield conducted a test program to produce cobalt sulphate at a pilot plant scale. Jacobs‟s personnel witnessed this testwork and used the provisional results of the SGS Lakefield modelling to develop the flowsheet. At the time of compiling this study, pilot testwork was in progress. The results will be updated at the next stage of engineering for the cobalt sulphate circuit.

SGS Lakefield proposed a process block diagram, provided initial sizing of equipment, and developed inputs to of the design criteria.

24.2 SCOPE OF WORK

The process work on the cobalt sulphate circuit (Area 4710 to 4720) was carried out at a scoping level. Apart from the areas that were removed (see below), the rest of the metallurgical process circuits are unchanged and remain at the FEED level.

Work involved a review of the SMPP process block diagram, to develop the equipment list (based on the Base Case equipment list, Revision I.)

The following process areas were removed:

Area 4400:  Cobalt carbonate precipitation dissolution and filtration,  Cobalt precipitation stage #2,  Cobalt dissolution and filtration Area 4500  Zinc IX and nickel IX  Zinc IX and elution  Zinc and nickel precipitation Area 4600  Cobalt electrowinning  Cobalt stripping and degassing

The new equipment list includes:

Area 4700  4710 Cobalt solvent extraction and,  4720 Cobalt crystallization

Based on the equipment list, the plant layout was revised for the new equipment.

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SGS proposed two options designated as A (DWG-4700-B-002) and B (DWG-4700-B-001). Option B is similar to Option A but includes the cobalt precipitation with Na2CO3 and releaching with sulphuric acid of the solvent extraction purified strip liquor. The CAPEX and OPEX have been developed for option B.

24.3 PROCESS DESCRIPTION (OPTION B)

Cobalt concentrate received as slurry from the flotation circuit associated with the cobalt hydrometallurgical plant contains about 3.4 % cobalt. The repulped bismuth leached residue and cobalt concentrate slurry at 60 wt% solids is subjected to leaching in a five-compartment pressure oxidation autoclave, the same as with the cobalt cathode option. A nominal 60 minutes of retention time, sufficient to leach up to 95% cobalt and 76% of copper. The autoclave operates at 180 °C and 2,100 KPa pressure. The leaching process is highly exothermic (refer to DWG- 4700-B-001).

The discharge slurry from autoclave circuit is fed to a cobalt thickener. The thickener underflow is filtered and the residue which contains significant amount of gold will be leached with cyanide and precipitated with zinc powder in a Merrill Crowe circuit to produce gold doré, the same as with the cobalt cathode option. The thickener overflow solution is the cobalt PLS and is sent to the downstream iron-arsenic precipitation, also the same as with the cobalt cathode option. Lime is added to neutralize acid and gradually bring the solution pH up to a level 4.6.

The slurry is sent to solid liquid separation. The solids of the underflow are leached to recover copper, which is cemented with iron dust, the same as with the cobalt cathode option. The overflow is pumped to a polishing precipitation tank for further removal of copper with Na2CO3. The thickener underflow is recycled back to the Fe-As precipitation tanks. The partially treated PLS is purified and concentrated in a solvent extraction circuit, which is different from the cobalt cathode option.

The solvent extraction circuit consists of four extractions, two scrubbing, three primary cobalt strips and two secondary stages for impurity strip.

The partially purified cobalt pregnant solution is fed to a Solvent Extraction (“SX”) circuit utilizing 12% v/v Cyanex 272 organic extractant dissolved in a suitable hydrocarbon diluent (Exxol D80), some zinc, manganese and magnesium are co-extracted while calcium and nickel are rejected.

The extraction circuit consists of four counter current stages. Each stage is comprised of a single mixer and a settler for phase disengagement.

Raffinate solution treatment is still to be confirmed. One option is to neutralize with Na2CO3 to precipitate nickel. After solid liquid separation, the effluent will be sent to the water treatment plant and the solid are put to storage.

Scrubbing is required to remove as much cobalt as possible, loaded with impurities into the organic extractant. As in the extraction, this operation is done with counter-current organic and aqueous flow.

For scrubbing, the loaded organic composition from the extractant circuit is fed to a two stage scrubbing circuit, where the organic is scrubbed with the cobalt strip feed to remove magnesium P&E Mining Consultants Inc., Report No. 247 Page 255 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. and calcium. The scrubbed loaded organic composition is then fed to the three stage stripping section.

The stripping of the cobalt takes place as sulphate by contacting the scrubbed solvent extractant with the mother liquor from crystallization adjusting the pH. The organic extractant is regenerated; followed by the separation of the cobalt – loaded stripping solution from the solvent – extractant.

Then, the cobalt stripping liquor (CoSO4) is purified by removing manganese, which is oxidized and precipitated using O2 with 2.5% SO2 at a pH of 3.5.

The manganese removal will be done in an agitated reactor. Acid is generated which may require additional lime, and is hazardous due to the use of SO2.

The cobalt stripped organic (barren organic) goes to a two-stage zinc stripping. The reacted slurry from the zinc precipitation tank is pumped to a zinc filter. The Zinc carbonate will be filtered and stored.

The manganese purified cobalt strip solution is neutralized with sodium carbonate (Na2CO3) and cobalt carbonate is precipitated. After a solid liquid separation, the cobalt carbonate cake is releached with sulphuric acid.

Cobalt sulphate heptahydrate crystals are produced from the leached solution by evaporating the solution to concentrate the cobalt sulphate and by crystallizing it in three vessels of Continuous Stirred Glass Lined Reactors (CSTR).

Crystals are separated from the mother liquor in a decanter centrifuge. The mother liquor is recycled back to the scrubbing stage and the cobalt stripping stage.

The cobalt sulphate is packaged into 50 kg bulk bags in a bulk bag filling system, supplied with a platform scale.

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Figure 24.1 Crystalization Block Flow Diagram (Option A)

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Figure 24.2 Crystalization Block Flow Diagram (Option B)

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24.4 COBALT SULPHATE HEPTAHYDRATE GRADE

The solvent extraction route provides flexibility with respect to impurity removal, and could tolerate high levels of zinc and nickel. The following specifications for the CoSO4.7H2O product are to be achieved:

TABLE 24.1 TARGET PRODUCT SPECIFICATION Unit Spec Molecular CoSO4 7H2O Formula Co wt% >20.76 Ni ppm 30 Fe ppm 10 Ca ppm 10 Mg ppm 20 Cu ppm 10 Mn ppm 10 Zn ppm 10 Pb ppm 10 Na ppm 30 Si ppm * As ppm 10 - 15 Cl % 0.01 – 0.02 C ppm * S ppm * pH - tba Water Insoluble - tba *Trace (tbc)

24.5 CAPITAL COST SUMMARY

The start-up capital cost estimate for the SMPP Cobalt Sulphate option is CAD$ 219.7 million. Cost estimates are based on second quarter 2010. These costs cover the period from the decision to proceed with the Project until the first day of production.

The accuracy range of the capital cost estimate is -5% to +20%. The accuracy prediction for the Project takes into account the current state of engineering and procurement. It is also based on the fact that budgetary quotes were obtained for all the commodities, labour, and 92% of the Process mechanical equipment of the base SMPP plant.

The sustaining capital costs have been excluded from this estimate. Summary of the initial capital cost is summarized in Table 24.2 below. P&E Mining Consultants Inc., Report No. 247 Page 259 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

TABLE 24.2 CAPITAL COST COMPARISON Description With Cobalt Sulphate RevI. Base Case RevJ. CAD CAD Direct Costs General 8,912,852 8,926,178 Mechanical Equipment 78,146,997 77,070,405 Bulk Commodities 60,460,790 54,760,081 Infrastructure 15,453,531 15,453,531

Total Direct Cost 162,974,170 156,210,195 Total Indirect Costs 56,115,831 53,921,804 Total Project Cost 219,090,001 210,132,000

24.6 OPERATING COST SUMMARY

The operating cost for cobalt sulphate is 11.15% lower than the base case as shown in Table 24.3.

TABLE 24.3 OPERATING COST SUMMARY With Cobalt Sulphate Base Case Description CAD/t bulk concentrate CAD/t bulk concentrate Labour 139.20 153.71 Power 60.79 71.84 Reagents 174.52 201.80 Maintenance Supplies 67.25 74.59 Infrastructure 7.22 7.22 Other 30.85 30.85

Total 479.83 540.02

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25.0 INTERPRETATION AND CONCLUSIONS

25.1 MINING

The capital costs of mine-related infrastructure such as the mine dry, mine offices, mine rescue station are included in the processing plant capital cost. Similarly, the costs of operating these infrastructure facilities is in the processing plant operating cost and the G&A cost.

The cost of mine closure cost has not been estimated as part of the open pit and underground costs. Underground electrical power costs are projected to be $0.254 / kW hr.

25.2 GEOLOGY

The NICO deposit currently being developed by Fortune is a new cobalt-gold-bismuth-bearing IOCG deposit located 160 km northwest of Yellowknife in the NWT. A mineral Resource Estimate, was prepared by P&E utilizing the latest GEMS access database for the project. P&E concluded that the deposit shows good continuity of mineralization and consequently the NICO deposit resources were entirely in the Measured and Indicated category. The resource estimate is suitable for use in an economic evaluation of a mining operation including a feasibility study.

The 2011 P&E resource model with an effective date of 30 November 2011 was used in the determination of a mineral reserve which is summarized in the following Tables. The mineral reserve for NICO, was prepared using commodity price assumptions of $US17.00/lb Co, $US1,200/oz Au, $US9.00/lb Bi and a $CDN/$US0.95 exchange rate. The model was reported at an NSR cut-off grades of $CDN104.64/t for underground mineable mineralization and $CAD32.21/t for open pittable mineralization.

TABLE 25.1 (1)(2) UNDERGROUND RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 282 4.93 0.14 0.27 Probable 94 5.60 0.11 0.19

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TABLE 25.2 (1)(2) OPEN PIT RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 20,513 0.94 0.11 0.15 Probable 12,099 1.05 0.11 0.13

TABLE 25.3 (1)(2) TOTAL RESERVES Classification Tonnes (k) Au (g/t) Co (%) Bi (%) Proven 20,795 0.99 0.11 0.15 Probable 12,193 1.09 0.11 0.13

25.3 MINING METHODS

The pit slope design investigations indicate an engineering geology model comprised of strong competent rock masses. Consequently, pit slope designs were are on kinematic assessments of the structural fabric in the wall rocks. Results indicate that bench geometries with an inter-ramp slope angles of 50˚ and slightly steeper can be achievable with good blasting practices and scaling.

Underground geometries were based on the same engineering geology model. Analyses included i) semi-empirical stope design and ii) three-dimensional numerical modeling for evaluating the interaction of the open pit and underground openings. Recommended maximum hydraulic radius (HR) for unsupported walls ranges from 4.5m (Hanging wall stability) to 6.6m (sidewall stability). The stope designs conform to the HR recommendations and systematic ground support such as cable bolts are not required.

Pit/underground workings interactions were assessed with respect to induced stresses on the pillars as the pit deepens. This guided the selection of underground working geometries as follows: use stope dimensions that would minimize the potential for dilution; don‟t carry out any backfilling during underground operations; and minimize as much as practically possible the interaction of the stopes with the open pit.

To mitigate the hazard of open pit mining above unfilled workings, void filling of stopes beneath the pit floor working areas, walls or ramps will be accomplished from the open pit, as part of safe P&E Mining Consultants Inc., Report No. 247 Page 262 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. mining practices. Some stopes require drop raises to be drilled from the surface for placement of muck, and the stopes will be tightly filled prior to advancing the ramp over them. Other areas will be infilled as the pit deepens by blasting of the ribs and pillars above and between them.

25.4 NICO SITE - TAILINGS AND WASTE ROCK CO-DISPOSAL FACILITY

The mining process will generate a total of 29.9 Mt of tailings and 96.9 Mt of mine rock at the NICO site. Both these waste streams will be disposed together in a facility referred to as the Co- disposal Facility (CDF). The CDF will have a relatively small footprint area and will be entirely located within the valley of the Grid Ponds. Co-disposal has the advantage of significantly reducing the access of oxygen and water into the potentially acid generating mine rock, thus reducing both acid generation and metal leaching.

The CDF will be contained on all sides by a Perimeter Dyke, which will be a prism of rock at least 25 m thick. Inside the Perimeter Dyke, the CDF will comprise a “layer cake” of alternating layers of mine rock and non-segregating thickened tailings, each about 5 m thick. Runoff and bleed water from the tailings deposition will be “reclaimed” back to the process plant for reuse. Since the CDF and the Perimeter Dyke will be generally free draining, some of the tailings bleed water and run-off water will seep through the facility and report to five topographically low areas downstream of the CDF. Seepage Collection Ponds (SCPs) will be constructed at each of these topographically low areas to intercept the seepage water. The SCPs will be relatively shallow ponds which will be contained by low permeability dams.

25.5 NICO SITE – ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

25.5.1 Environmental Study Results

A DAR was submitted to the MVEIRB in May 2011. The conclusions of the DAR are discussed in detail under Item 20. Changes to the biophysical environment from the NICO Project are not predicted to result in significant adverse residual impacts to valued ecosystem components. Consequently, the NICO Project is not predicted to have significant adverse impacts on traditional and non-traditional land use practices.

The active mine area will be small (approximately 485 hectares [ha]), with limited changes made to the natural flow of water. The NICO Project will have a minimal effect on water quantity, air, soils, vegetation, and wildlife and fish health. Closure, caribou and water quality have been identified by the communities as the most important concerns related to the environment. People should not be able to observe a change in the availability of wildlife due to effects of the NICO Project, relative to current natural changes in population sizes. Changes in water, soils, and plants caused by the NICO Project in the small area at and near the mine site will not affect the health of wildlife, or the health of people that eat wildlife.

25.5.2 Water Management

As described in Item 20, the major components of the water management system of the Project will comprise: Lou Lake, the CDF Reclaim Pond, five SCPs, a Surge Pond, Open Pit sumps, a Process Plant Runoff Pond, sewage treatment plant (STP) and an ETF.

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During Operation: Lou Lake will be the source of fresh water for Process Plant, dust control, and potable water. SCPs No. 1, 2 and 3, which will be located in three topographic lows adjacent to the western end of the CDF, are designed to intercept seepage from the CDF which would otherwise flow to NICO Lake. The Surge Pond will temporarily store contact water pumped back from the SCPs and the Reclaim Pond. Water will be pumped from the Surge Pond either to the Process Plant for reuse or to the ETF for treatment. The ETF will use a RO system. Treated water from the ETF will be pumped through a diffuser directly into Peanut Lake. Treated effluent from the STP will also be discharged through the same diffuser into Peanut Lake. Water balance analysis indicates that the average flow discharged into Peanut Lake will be a relatively small flow of about 290,000 m3/year.

After Closure: The water which accumulates in SCP Nos. 1, 2, 3, and 5, as well as the Surge Pond, will be passively treated in Wetland Treatment Systems and then released directly into NICO Lake. At the end of mining, pumping of water out of the Open Pit will cease and the Open Pit will slowly fill with water. Modelling indicates that it will take roughly 120 years for the Open Pit water level to rise to Elev. 260 m, at which time the flooded Open Pit may begin to overflow. Just prior to pit overflow, the water quality at the top of the flooded Open Pit will be evaluated, and a decision will be made regarding post-overflow treatment.

25.5.3 Closure

Closure measures are described in Item 20. For closure, a soil cover will be placed over the entire CDF facility. The seepage water that reports to SCP Nos. 1, 2, 3 and 5 will be routed through Wetland Treatment Systems into NICO Lake. The Wetland Treatment Systems will be constructed and tested prior to the end of mine operations. The ETF will be maintained on site for 10 years as a backup to the Wetland Treatment Systems. Post-closure water management is described above.

25.5.4 Project Permitting

Item 20 outlines the permitting regime affecting the NICO project. It also provides details of the consultation and permitting activities undertaken to date. Significant milestones to date have included: submission of the DAR, [the NWT analogue to an EIS], on 20 May 2011, and a Technical Session held in Yellowknife, NWT from 7 to 9 February 2012. Public Hearings are tentatively scheduled for the week of August 27-31, 2012.

25.5.5 Social and Community Requirements and Considerations

As described under Item 20, the NICO Project is anticipated to have significant positive impacts on the economics of the Tłįchǫ communities, and both positive and negative (but not significant) impacts on the social and cultural environments. The NICO Project is a small development compared to other mines in the Northwest Territories, but it will contribute to the overall labour, financial, physical, human, and social resources of both the Northwest Territories and more specifically the nearby communities.

25.5.6 Mine Closure Requirements and Costs

Mine closure and reclamation is discussed under Item 20. The next step is to develop a formal Conceptual Closure and Reclamation Plan (“CCRP”) according to the requirements of the AANDC and the Wek‟èezhìı Renewable Resources Board. The CCRP will include an estimate P&E Mining Consultants Inc., Report No. 247 Page 264 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd. of closure costs and a financial assurance will be posted based on this estimate. The amount and form of this financial assurance will be subject to negotiations between Fortune and AANDC. Fortune has met with AANDC to begin initial discussions regarding scenarios for bonding for closure.

25.5.7 Metallurgy and Processing

The FEED study completed for the NICO Project is comprehensive covering all the main Project elements of mineral processing at the concentrator and hydromet processing at SMPP. The following is concluded:

 Available data is sufficient to meet the requirements of a FEED study.  The project is technically viable, subject to further works to be carried out in the detailed engineering. The technical issues are recorded in the FEED study. Recommendations to address the main issues are in Item 26 of this report.  The process plant is designed for approximately a 20 year Life of Mine at a production rate of approximately 1.7 Mt/a.

25.5.8 Project Execution

The most crucial element for project execution will be the construction of the all-weather road. This will provide year round access, allowing for standard Northern construction methodology.

25.5.9 Project Economic Analysis

The projected economic outcome of the project has been demonstrated using a discounted after- tax cashflow model that is based on the project development and operations schedule and incorporates projected revenues, capital and operating costs, working capital and interest cost allowances and taxes.

Potential significant risks and uncertainties that could reasonably be expected to offset the projected economic outcome are reviewed below.

Cobalt product Option

This Technical Report has assessed and demonstrated the technical and economic viability of the cobalt metal option. However based on the results of the economic analysis the cobalt sulphate option offers a higher IRR and NPV in comparison to the cobalt metal option and appears promising. As such, it is recommended that additional cobalt sulphate option metallurgical testwork and assessment be completed before Fortune Minerals commits to the cobalt sulphate option. See the cobalt sulphate option recommendation in Section 26.

Project Schedule

It is assumed that an all-weather access road to the NICO site will be built before the commencement of the main construction and available over the life of the project. The project schedule and operating cost estimates include allowances for envisaged delays / lost time due winter weather conditions. The cost estimates do not however include costs for a winter road or additional indirect operating costs due to schedule extending delays.

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There is a possibility that additional higher NSR value ore, in relation to typical NSR values for ROM pit ore, could be extracted from the underground mine during the early phase of the project.

25.6 MARKETING

25.6.1 Cobalt

The long-term outlook for cobalt is generally positive. It is anticipated that 2012 will see a small surplus of refined Co supply over demand, but expect this situation to switch to a small deficit in 2013. Demand will again exceed supply in 2014 and 2015, though supply will increase as Chinese refiners ramp up production to keep up with battery demand in Asia.

Towards the end of 2012 it is forecast pricing to be in the $16/lb range. While fluctuations will always be a part of market prices, longer term pricing beyond 2012 is expected to move within the range of US$18/lb to US$22/lb.

25.6.2 Bismuth

Bismuth supply is expected to remain stable during 2012 and 2013. Demand growth for bismuth is expected to be in the 8% to 10% range annually over the coming 3 years. Bismuth demand is forecasted to enjoy healthy growth due to its role as a non-toxic substitute for lead, from an environmental regulations point of view.

Bismuth prices are forecasted to increase by US$1.50 in 2013 reaching $$14.00/lb, and then hitting $15.50/lb by the end of 2014.

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26.0 RECOMMENDATIONS

26.1 GEOTECHNICAL

It is recommended that a geotechnical assessment of the exposed development ramp and drifts be carried out as part of any remediation work following dewatering and updated as additional drifts are advanced/excavated in order to confirm the rock mass fabric and parameters used in these geotechnical studies.

The open pit and underground geotechnical analyses have been carried out based solely on results of previous investigation drilling and summaries, completed prior to 2005. It is recommended that these numerical analyses be re-visited as soon as the presently flooded underground workings have been pumped out and a geotechnical assessment of the exposed development and test stope completed and rock mass parameters confirmed.

26.2 METALLURGY AND PROCESSING

26.2.1 NICO Concentrator – Detailed Engineering Execution

The engineering priorities for the next phase will be as follows:

 Initiate and complete remaining process testwork and investigations. Future Works (listed in item 26.2.2), particularly the trade-off study required for finalising the crusher selection, the testwork and trade-off study for the gravity gold circuit, review and confirmation of the flotation circuit, and size confirmation of the bulk concentrate filter.  Confirm concentrate feed rate and specification. This will be critical for the sizing of the autoclave.  Complete a HAZOP review to assess the FEED design from the operations viewpoint. Plant operability / maintainability from a worker health and safety aspect are to be reviewed in detail, in conjunction with Client specialists who are ultimately responsible for the operation of the „as-constructed‟ processing plant.

26.2.2 NICO Process Opportunities – Future Works

A complete list of opportunities and recommendations to be further reviewed and discussed in detail engineering are included below.

Crusher Selection

An opportunity exists for the replacement of the existing primary Jaw crusher with either a Sandvik hybrid crusher, or by engineered blasting.

A high level cost saving and operability comparison was carried out between a Jaw crusher and a Gyratory crusher. The conclusion was that a single C 160 Jaw crusher would be the least cost alternative and as such was adopted as the Project‟s primary crusher option. It is believed that the hybrid crusher option, as offered by Sandvik, could further reduce capital costs. However, since the hybrid crushing is relatively new to hard rock mining, sufficient due diligence and review must be performed. This trade-off is to be investigated during the next phase of engineering.

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There is also the opportunity to reduce the cost of the primary crusher by employing engineered blasting technique at the mine to decrease the size of the blasted ROM ore. More investigation is required at the next phase of the Project to determine if this technique is suitable.

Gravity Recoverable Gold

The gravity recovery testwork is based on the P3 and P4 composite, which respectively contain high bismuth and low cobalt in one sample, and high cobalt and low bismuth in the other, but have similar gold grades. Knelson modelled the potential gravity gold recovery of a conceptual stream that took on the characteristic of a blend of 50/50 P3 and P4 composite.

As pointed out by the Vendor, the results obtained from the model should be considered very preliminary at this stage, since the database of measured results is very limited. Further bench- scale work with representative drill core could be considered, in order to provide a more definitive prediction of the overall gold recovery benefit.

A trade-off exists for gold recovery versus equipment. A higher gold recovery can be resulted by using a larger gravity circuit, but this increase in revenue is offset by the increase in capital cost.

Primary Grind Size

The 2007 pilot plant conducted at SGS showed that the feed size to flotation can have an impact on bulk flotation performance. With an increasing size, there was no change to the recovery of cobalt, a slight decrease in the recovery of bismuth and a higher decrease in the recovery of gold. Further testwork is recommended to determine the optimal feed size to bulk flotation.

On-Stream Analyzer

Future considerations will be made to install an OSA system in the flotation area to provide assay information on nine streams in the flotation circuit.

Bulk Concentrate Pressure Filter

It is recommended that a sample of the bulk concentrate intermediate product is provided to the Vendor for confirmatory testing for filter selection and sizing. Opportunities exist for the material to filter more efficiently and effectively than originally estimated, hence reducing the size of the equipment and the cost of transportation.

Flotation FLEET Testwork

The recoveries in the pilot plant were lower than the results from the FLEET model. Additional tests are recommended to verify earlier pilot recovery results and regrind mill sizes.

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26.2.3 SMPP - Detailed Engineering Execution

The engineering priorities for the next phase will be as follows:

 Initiate and complete remaining process testwork particularly continuous IX, cyanide destruction and the production of cobalt pucks.  The Metsim model needs to be updated and remodelled based on the latest cobalt, gold and copper stage recoveries obtained from the 2010 – 2011 pilot tests  Confirm concentrate feed rate and specification. This will be critical for the sizing of the autoclave.  Complete a HAZOP review to assess the FEED design from the operations viewpoint. Plant operability / maintainability from a worker health and safety aspect are to be reviewed in detail, in conjunction with Client specialists who are ultimately responsible for the operation of the „as-constructed‟ processing plant.  Identification of long-lead items and finalisation of Vendor bid list.  Early engineering of the process building will be required to secure delivery of steel on-site at the appropriate time.  Update of 3D layout with recent process changes and certified drawing from the chosen suppliers of equipment.

26.2.4 SMPP Process Opportunities - Future Works

A complete list of opportunities and recommendations to be further reviewed and discussed in detail engineering are included below.

Regrind and Flotation

Two 355-kW SMDs were proposed for the regrind application. A further review in DE should be conducted to optimize the size, quantity and configuration.

Ion Exchange – Continuous Versus Fixed Bed

Further testwork is recommended to assess continuous IX columns as previous testwork was on fixed bed IX columns. CIX columns could be more cost effective.

Design of Cobalt Electrowinning Circuit

Further testwork is recommended to confirm adequate deposition of Cobalt pucks at higher current density and higher cobalt electrolyte concentration.

Cyanide Destruction

Cyanco Canada has been retained by Fortune Minerals to complete a testwork program to evaluate cyanide destruction options from the SMPP plant. Cyanco Canada will obtain the autoclave residue and perform cyanidation testwork on the material.

Since the most suitable cyanide destruction method was not determined by the release of this report, hydrogen peroxide was selected as the method of treatment for the SMPP FEED study. This selection should be reviewed at the completion of Cyanco‟s cyanide destruction testwork.

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Other Opportunities and Recommendations

Cobalt Sulphate Option

26.2.5 Bismuth Processing Plant

 Leach Temperature Optimization: According to prior SGS analyses, Leach Stages 1 and 2 temperature of 95 – 100°C were successfully used, however preliminary tests showed that the temperature can be reduced to 80°C without affecting the recoveries. There remains potential for improvement with respect to reduced leach temperature and related materials selection. The cost of PVDF (Kynar)- lined steel pipe, required for 95°C leaching circuit was estimated at $1MM while the cost of lower temperature materials is approximately 1/3. Further testing should be conducted to optimize leach temperatures as part of overall optimization testing.  Electrowinning: A test program is suggested to optimize the cathode/anode current density. Test work was conducted with a current density of 200 A/m2, which was determined adequate to produce Bi in powder form. There was insufficient concentrate remaining to test further, however, it has been proposed that a current density between 300-330 A/m2 is achievable. This would reduce the cathode requirement from 480 m2 to potentially as low as 300 m2, a 37.5% reduction in the size of the tankhouse. The potential savings in the EW cells would be $745,000 before the building; steelwork and other savings are taken into account. If a supply of Bi Con becomes available, it is recommended that a test plan for current density optimization be implemented. Concurrently the Anolyte and Catholyte recirculation rates should be optimized for best recoveries. Should an EW test be conducted, it would be prudent to perform the testing in new EW lab test cells that have some of the design features that were developed to the concept stage. For instance, the angle of repose of the Bi powder should be confirmed; and the potential for gravity collection and peristaltic pumping confirmed.  Iron Removal: Further work should be conducted to determine if the iron levels in anolyte are acceptable for direct injection to the saline aquifer. This would entirely eliminate the Fe/As circuit with a potential cost savings of $528,800.  Electrical: In the detailed engineering phase, the electrical design philosophy should be reviewed for some potential cost savings with alternate suppliers,

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components and reducing some redundant processors and network. The total potential savings in Electrical work and components is over $400,000.  Cost of Further Bismuth Processing Plant Testing: Any further testing is dependent on the production and delivery of enough Bismuth Concentrate. The cost of, design and fabrication of new test equipment, management of the test program; and testing is estimated at approximately $150k with a potential savings to the Bismuth Plant CAPEX of up to $3.8M.

In the detailed engineering a control estimate should be produced that will update the CAPEX (and OPEX) estimates to current costing

26.2.6 Cobalt Sulphate Option Recommendation

It is recommended that Fortune Minerals continue with metallurgical testwork to confirm the quality of the end product and to carry out further engineering to complete FEED for the cobalt sulphate circuit. The cost estimate in FEED for the cobalt sulphate option will have an overall accuracy of ±15%. The economic analysis could then be re-run before proceeding with this option. The pilot testwork, interpretation of results, process design, engineering, bid requests, bid evaluations and estimating can be completed over a 4 to 5 month time line at a cost of between $250k and $325k.

26.3 ENVIRONMENTAL-- PERMITTING, TAILINGS AND CLOSURE

It is recommended that pilot scale testing be carried out to support for detailed design of the tailings thickener, including flow loop studies.

Large scale test cells for co-disposal should be set up on site in the early stages of operations, once representative mine rock is available. These should incorporate lysimeters to allow sampling of the leachate over time.

The project schedule call for the first three Wetland Treatment Systems to be constructed, tested and optimized during the mine operating period. This work is to be completed prior to closure so that the Wetland Treatment Systems are verified and fully operational by the time of closure.

26.4 MARKETING

It is recommended that Fortune Minerals institute a Research and Development Group to increase demand by developing new uses for both cobalt and bismuth.

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27.0 REFERENCES

Aboriginal and Territorial Relations (2007): The Mackenzie Valley Resource Management Act. Minister of Public Works and Government Services Canada.

Aker Solutions (2009): Bulk Concentrate Load Out Containers, NICO Cobalt-Gold-Bismuth Project, July 2009

ALS Laboratory Group (2009): Certificate of Analysis, Analytical Report, September 18, 2009

Barton et al (1974): Rock Quality Index Q in “Engineering Classification of Rock Masses for the Design of Tunnel Support”. Rock Mechanics, Vol. 6, No. 4, pp. 183-236.

Brown, J., O.J. Ferrians, Jr., J.A. Heginbottom, and E.S. Melnikov. (1998): revised February 2001. Circum-arctic map of permafrost and ground ice conditions. Boulder, CO: National Snow and Ice Data Center/World Data Center for Glaciology. Digital media.

Bryan, D. (1981): Geological Report on the GAR 1-3 Claims, Noranda Exploration Company Limited. Department of Indian Affairs and Northern Development, Assessment Report No. 081145.

Bryan, D. (1982): Geological Report, GAR 1-5, 7 and 9 Claims (Jan-Dec 1981), Lou Lake Area, District of Mackenzie (85N/10). Department of Indian Affairs and Northern Development, Assessment Report No. 081422.

Campbell Johnston, (2008): Metso, Minerals Bruno Crushing Process Simulations, December 1, 2008

Clarke, S., (2008): SJC Engineering, Autoclave Corrosion Testing of Candidate Alloys for HPOX, October 17, 2008

Corriveau, L. (2007): Iron Oxide Copper-Gold Deposits: A Canadian Perspective. In: Goodfellow, W.D. (ed) Mineral Deposits of Canada: A Synthesis of Major Deposit-Types, District Metallogeny, the Evolution of Geological Provinces, and Exploration Methods. Geological Association of Canada, Mineral Deposits Division, Special Publication no. 5, p. 307-328.

EHA Engineering (2005): Review of Hydrometallurgical Testwork, August 2005.

EBA Engineering Consultants Ltd., (2005a): NICO Mine Access Route Evaluation, Final Report, March 2005. Report submitted to Fortune Minerals Ltd., London, Ontario.

EBA Engineering Consultants Ltd., (2005b): NICO Tailings Dam and Process Plant Facilities 2004 Geotechnical Site Investigation, Final Report, April 2005. Report submitted to Fortune Minerals Ltd., London, Ontario

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Fraser, J.A., Hoffman, P.F., Irvine, T.N. and Mursky, G. (1972): The Bear Province. In: Price, R.A. and Douglas, R.J.W. (eds) Variations in Tectonic Styles. Geological Association of Canada, Special Paper no. 11, p. 453-503.

Gandhi, S.S. (comp.) (2004): World Distribution of Iron Oxide ± Cu ± Au ± U (IOCG) Deposits. Geological Survey of Canada, Geoscience Data Repository, Mineral Resources Metadata. http://gdr.nrcan.gc.ca/minres/metadata_e.php?id=8.

Gandhi, S.S., Prasad, N. and Charbonneau, B.W. (1996): Geological and Geophysical Signatures of a Large Polymetallic Exploration Target at Lou Lake, southern Great Bear magmatic zone, Northwest Territories. Current Research 1996-E, Geological Survey of Canada, p. 147-158.

Goad, et al, (1998): The NICO and Sue-Dianne Proterozoic, Iron Oxide-hosted, Polymetallic Deposits, Northwest Territories: Application of the Olympic Dam Model in Exploration.

Goad, R.E., Mumin, A.H., Duke, N.A., Neale, K.L., Mulligan, D.L. and Camier, W.J. (2000a): The NICO and Sue-Dianne Proterozoic, Iron Oxide Hosted, Polymetallic Deposits, Northwest Territories: Applications of the Olympic Dam Model in Exploration. Exploration and Mining Geology, vol. 9, no. 2, p. 123-140.

Goad, R.E., Mumin, A.H., Duke, N.A., Neale, K.L. and Mulligan, D.L. (2000b): Geology of the Proterozoic Iron Oxide-Hosted, NICO Cobalt-Gold-Bismuth and Sue- Dianne Copper-Silver Deposits, Southern Great Bear Magmatic Zone, Northwest Territories, Canada. In: Porter, T.M. (ed.) Hydrothermal Iron Oxide Copper-Gold & Related Deposits: A Global Perspective. Australian Mineral Foundation, Adelaide, p. 249-267.

Golder Associates Ltd. (1998): Environmental Scoping for Fortune Minerals‟ NICO and Sue- Dianne Properties. Report No. 982-6849.

Golder Associates Ltd., (1999): A Review of Geotechnical Data and Recommendations for Pit Slope Design Configurations, NICO Deposit. July 19, 1999. Project 982- 1426.

Golder Associates Ltd., (2003): Factual Report on Geotechnical and Hydrogeological Investigations for the Proposed Open Pit and Underground Mine Workings. NICO Deposit – Northwest Territories. Project # 03-1117029.

Golder Associates Ltd. (2003a): Factual Report on Geotechnical and Hydro-Geological Investigations for the Proposed Open Pit and Underground Mine Workings, NICO Deposit, Northwest Territories. Draft report submitted to Fortune Minerals Limited, London, Ontario.

Golder Associates Ltd. (2003b): Heritage Resources Impact Assessment, Fortune Minerals NICO Gold Prospect, Northwest Territories, Archaeologists Permit #2003-942. Final report submitted to Fortune Minerals Limited, London, Ontario, 41p.

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Golder Associates Ltd., 2004: Technical Memorandum – Open Pit Slope Design Recommendations for NICO Project. Memorandum submitted to Fortune Minerals Ltd. on November 3, 2004.

Golder Associates Ltd (2004a): 2003 Environmental Surveys at Fortune Minerals‟ NICO Property. Final report submitted to Fortune Minerals Limited, London, Ontario, 59p.

Golder Associates Ltd (2004b): Open Pit Slope Design Recommendations, Fortune Minerals Limited, NICO Gold-Bismuth-Cobalt Project. Report 03-1117-029.

Golder Associates Ltd., (2005a): Factual Report on Geotechnical and Hydrogeological Investigations (2003) for the Proposed Open Pit and Underground Workings, NICO Deposit, Northwest Territories.

Golder Associates Ltd., (2005b): Technical Memorandum – Stoping Dimensions Based on Mathews/Potvin Stability Graph Analysis. Memorandum submitted to Robin Goad and Gene Puritch on January 25, 2005. Proejct # 05-1117032.

Golder Associates Ltd., (2007): Report on NICO Tailings Dams and Process Plant Facilities 2006 Geotechnical Site Investigation. Submitted to Fortune Minerals Limited. April 2007.

Golder Associates Ltd., (2010a): Technical memorandum on Geotechnical Underground Design for the NICO Project. Document DOC093 submitted to P & E Mining Consultants Inc. on February 18, 2010. Project # 08-11180043.

Golder Associates Ltd., (2010b): Technical memorandum on Geotechnical Underground Design for the NICO Project, including stopes OR080 and OR085. Document DOC0101 submitted to P & E Mining Consultants Inc. on April 23, 2010. Project # 08-11180043.

Golder Associates Ltd., (2010c): NICO Fortune Minerals Groundwater Modeling - Technical Supporting Document for the NICO Project EIA. February 2011. Project # 10-11180046.

Golder Associates Ltd. (2011): Fortune Minerals Limited Developer‟s Assessment Report. Project # 09-1373-1004.

Golder Associates Ltd. (2012): Technical Memorandum on Revision C Cost Estimate Summary for the NICO Tailings and Mine Rock Co-disposal Facility. Document DOC010 submitted to Fortune Minerals Limited on May 17, 2012. Project # 11-1118-0066.

Hennessey, B.T. and Puritch, E. (2004): An Updated Mineral Resource Estimate for the NICO Cobalt-Gold-Bismuth Deposit, Mazenod Lake District, Northwest Territories, Canada. NI 43-101 technical report prepared for Fortune Minerals Limited by Micon International Limited, November 2004.

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Hennessey, B.T., Puritch, E., Ward, I.R., Konigsmann, K.V., Hayden, A.S., Bocking, K.A. and Rougier, M. (2007): Technical Report on the Bankable Feasibility Study for the NICO Cobalt-Gold-Bismuth Deposit, Mazenod Lake Area, Northwest Territories, Canada. NI 43-101 technical report prepared for Fortune Minerals Limited by Micon International Limited, February 2007.

Herget, G., (1990): Rock stresses and rock stress monitoring in Canada. MRL 90-011(TR) December 1990.

Hildebrand, R.S. and Bowring, S.A. (1984): Continental Intra-Arc Depressions: A Non- Extensional Model for Their Origin, With a Proterozoic Example From Wopmay Orogen. Geology, vol. 12, p. 73-77.

Hildebrand, R.S., Hoffman, P.F. and Bowring, S.A. (1987): Tectono-Magmatic Evolution of the 1.9 Ga Great Bear Magmatic Zone, Wopmay Orogen, Northwestern Canada. Journal of Volcanology and Geothermal Research, vol. 32, p. 99- 118.

Hoek, E, Kaiser, P.K, and Bawden, W.F. (1995): Support of Underground Excavations in Hard Rock, Chapter 14 – The Stability Graph Method. A.A. Balkema, 1995.

Hoffman, P.F. (1973): Evolution of an Early Proterozoic Continental Margin: The Coronation Geosyncline and Associated Aulocogens of the Northwestern Canadian Shield. Royal Society of London Philosophical Transactions, Ser. A, 273, p. 547-581.

Hoffman, P.F. (1980): Wopmay Orogen: A Wilson Cycle of Early Proterozoic Age in the Northwest of the Canadian Shield. In: Strangway, D.W. (ed) The Continental Crust and its Mineral Deposits. Geological Association of Canada, Special Paper no. 20, p. 523-549.

Kilborn SNC-Lavalin (1998): Scoping Study Report, NICO Project, Northwest Territories. Report prepared for Fortune Minerals Limited, November 1998.

Knelson Research & Technology Centre (2011): Gravity Modeling Report for Fortune Minerals NICO Gold-Copper-Bismuth Project, Project KRTS 20621-1, January 28, 2011

Jackman, R., and Konigsmann, K., (2004): Report on Summary of Test Results and Processing Options, July 29, 2004

Jenike & Johanson (2009): Flow Property Test Results for Untreated NICO Ore, Report 9513-1, January 7, 2009

Lakefield Research Limited (1997a): An Investigation of the Recovery of Copper, Cobalt, Gold and Bismuth from NICO Project Samples. Progress report no. 1, project no. LR5070, April 30, 1997.

P&E Mining Consultants Inc., Report No. 247 Page 275 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Lakefield Research Limited (1997b): An Investigation into the Recovery of Metal Values from NICO Project Samples. Progress report no. 2, project no. LR5070, July 8, 1997.

Lakefield Research Limited (1998): An Investigation of the Recovery of Cobalt, Bismuth and Gold from NICO Project Samples. Progress report no. 3, project no. LR5070, June 12, 1998.

Lakefield Research Limited (2001): An Investigation into the Recovery of Cobalt and Bismuth from NICO Project Samples. Progress report no. 6, project no. LR10226- 001, September 25, 2001.

MacPherson, A.R., Proposed Grinding System for NICO Comminution Circuit, Report LR10044-103, December 10, 2004

Marinelli, F., and W. L. Niccoli. 2000. Simple analytical equations for estimating ground water inflow to a mine pit.

Mathews, K. E., Hoek, E., Wyllie, D.C. and Stewart, S.B.V. (1981): Prediction of Stable Excavations for Mining at Depth below 1000 metres in Hard Rock. CANMET Report DSS Serial No.OSQ80-0081, DSS File No. 17SQ.23440-0-9020, Ottawa: Dept. Energy, Mines and Resources.

Micon International Limited, (2007): NICO Cobalt-Gold-Bismuth Deposit, Bankable Feasibility Study, Volume I, February 2007

Molnar, R., (2012): SGS Minerals Services, FML CoSO4 Flowsheet and Balances, February 24, 2012

Mumin, A.H. (1997) A Qualifying Report on the Geology and Mineralization of the NICO 1 (F28905), NICO 2 (F28906), NICO 3 (F50933), NICO 4 (F18965), NICO 5 (F18966), NICO 6 (F50155), NICO 7 (F50156), NICO 8 (F50157), NICO 9 (F50158), NICO 10 (F50159), NICO 11 (F51389) & NICO 12 (F51390) Claims, Marian River Area, Mackenzie (South) District, Northwest Territories, Canada. Report prepared for Fortune Minerals Limited, April 1997.

Mumin, A.H. (1998a): Gold, Cobalt, Bismuth and Copper Resources of the Bowl Zone Deposit, NICO Property, Mackenzie (South) District, Northwest Territories. Report prepared for Fortune Minerals Limited, May 1998.

Mumin, A.H., (1998b): High-Grade Resources, Deep Exploration and Mining Potential of the NICO Property, Northwest Territories. Report prepared for Fortune Minerals Limited, June 1998.

MVRB (Mackenzie Valley Review Board). (2009): Terms of Reference for the Environmental Assessment of Fortune Minerals Ltd. NICO Cobalt-Gold-Bismuth-Copper Project EA 0809-004. Yellowknife, NWT.

P&E Mining Consultants Inc., Report No. 247 Page 276 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

MVLWB (Mackenzie Valley Land and Water Board). (2009): Closure and Reclamation Plans – Preparation Guidelines for Mines within the Mackenzie Valley. Draft 15 July 2009.

P&E Mining Consultants Inc., (2009): Table on 20091214_NICO Production Schedule Open UG (updated), December 15, 2009

Pocock, (2008): Hydrometallurgical Solid-Liquid Separation Report for SGS-Fortune, May 2008.

Pocock, (2008a): Concentrator Solid-Liquid Separation Report for SGS-Fortune, December 2008

Potvin, Y. (1988): Empirical Open Stope Design in Canada, Ph.D. thesis, Dept. Mining and Mineral Processing, Univ. of B. Columbia.

Samuels, M, (2007): Fortune Minerals Ltd, Bulk Sample Strategy for the NICO 200 t Pilot Plant Study, August 14, 2007

SGS, An Investigation into The Recovery of Gold from NICO Test Samples, Report 11758-1- No.5, November 12, 2010

SGS Lakefield Research, (2007): Flocculant Screening, Gravity Sedimentation and Pulp Rheology, December 2007

SGS Mineral Services, (2008): Flocculant Screening, Gravity Sedimentation, Pulp Rheology, Pressure Filtration and Vacuum Filtration Studies, May 2008

SGS Lakefield, (2009): A Pilot Plant Investigation into Cobalt, Bismuth, and Gold Recovery from the NICO Deposit (DRAFT), Report 11747-001, January 16, 2009

SGS, (2009a): NICO Process Pilot Plant Report 2, Cobalt Carbonate Leaching and Cobalt Metal Electrowinning, Report 11758-001, March 10, 2009

SGS, (2009b): Recovery of Cobalt, Gold and Bismuth from the NICO Deposit Polymetallic Concentrates (DRAFT), Report 11758-001, March 31, 2009

SGS, (2009c): An Investigation into FLEET Modelling and Optimisation of the NICO Bulk Flotation Circuit, Report 11747-003, November 18, 2009

SGS, (2011): Phase 3 Flotation Pilot Plant Testing of Samples From the NICO Deposit, Report, 11747-004, April 12, 2011

SGS, (2011a): Pilot Plant Investigation into Recovery of Cobalt, Gold and Bismuth from the NICO Deposit Polymetallic Concentrates, Report 11758-005-No.1, December 8, 2011

SNC-Lavalin Engineers & Constructors (1999): Geological Audit and Resource Estimate Report, NICO Project, Northwest Territories. Report no. 334164.

Starkey, J., (2008): Report of Ore Hardness Variability Testing, November 30, 2008 P&E Mining Consultants Inc., Report No. 247 Page 277 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

Strathcona Mineral Services Limited (2000): Scoping Review NICO Cobalt-Gold Project, Mazenod Lake District, Northwest Territories, Canada. Report prepared for Fortune Minerals Limited, April 2000.

Thalenhort, H. and Farquharson, G. (2002): Updated Review, NICO Cobalt-Gold-Bismuth Project, Mazenod Lake District, Northwest Territories, Canada. NI 43-101 technical report prepared for Fortune Minerals Limited by Strathcona Mineral Services Limited, April 2002.

Thomas, M. and Olson, R.A. (1978): Eldorado Nuclear Limited Exploration - 1977 and 1978, LOO, BW and C Mineral Claims, Lou Lake, Mackenzie Mining District, NWT. Department of Indian and Northern Affairs, Assessment Report No. 080960.

Walker, E.C. (1999): Mineralogy Report. Petrologic Inc. report prepared for SNC-Lavalin Engineers & Constructors Inc., Report P1990304

Xstrata Process Support, (2010): Erosion/Crevice-Corrosion Testwork for Titanium Autoclave with 30% solids in the Slurry, June 22, 2010

Xstrata Process Support, (2010): Erosion/Crevice-Corrosion Testwork for a Titanium Autoclave, Report Rev. 1, January 10, 2010

P&E Mining Consultants Inc., Report No. 247 Page 278 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

28.0 CERTIFICATES

CERTIFICATE OF QUALIFIED PERSON

EUGENE J. PURITCH, P.ENG.

I, Eugene J. Puritch, P.Eng., residing at 44 Turtlecreek Blvd., Brampton, Ontario, L6W 3X7, do hereby certify that:

1. I am an independent mining consultant and President of P & E Mining Consultants Inc. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada” (the “Technical Report”) with an effective date of July 2, 2012. 3. I am a graduate of The Haileybury School of Mines, with a Technologist Diploma in Mining, as well as obtaining an additional year of undergraduate education in Mine Engineering at Queen‟s University. In addition I have also met the Professional Engineers of Ontario Academic Requirement Committee‟s Examination requirement for Bachelor‟s Degree in Engineering Equivalency. I am a mining consultant currently licensed by the Professional Engineers of Ontario (License No. 100014010), Association of Professional Engineers and Geoscientists Saskatchewan (License No. 16216), Professional Engineers and Geoscientists Newfoundland and Labrador (License No. 05998), Profession Engineers and Geoscientists New Brunswick (License No. L4778) and registered with the Ontario Association of Certified Engineering Technicians and Technologists (License No. 45252), as a Senior Engineering Technologist. I am also a member of the National Canadian Institute of Mining and Metallurgy.

I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. I have practiced my profession continuously since 1978. My summarized career experience is as follows:

 Mining Technologist - H.B.M.& S. and Inco Ltd., ...... 1978-1980  Open Pit Mine Engineer – Cassiar Asbestos/Brinco Ltd., ...... 1981-1983  Pit Engineer/Drill & Blast Supervisor – Detour Lake Mine, ...... 1984-1986  Self-Employed Mining Consultant – Timmins Area, ...... 1987-1988  Mine Designer/Resource Estimator – Dynatec/CMD/Bharti, ...... 1989-1995  Self-Employed Mining Consultant/Resource-Reserve Estimator, ...... 1995-2004  President – P & E Mining Consultants Inc, ...... 2004-Present

4. I visited the NICO Property on July 10-11, 2004 and on April 24, 2012. 5. I am responsible for authoring Section 19 co-authoring Sections 1, 12, 15, 16, 25, and 26 of the Technical Report. 6. I am independent of Fortune Minerals Ltd. applying the test in Section 1.5 of NI 43-101. 7. I have had prior involvement with the project that is the subject of this Technical Report. The nature of my involvement is as a co-author of several technical reports, the most recent one titled: “Fortune Minerals Limited, Technical Report on the Bankable Feasibility Study for the Nico Cobalt-Gold-Bismuth Deposit, Mazenod Lake District, Northwest Territories, Canada” dated February 28, 2007. 8. I have read NI 43-101 and Form 43-101F1. This Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Eugene J. Puritch] ______Eugene J. Puritch, P. Eng.

P&E Mining Consultants Inc., Report No. 247 Page 279 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

WAYNE D. EWERT, P.GEO.

I, Wayne D. Ewert, P. Geo., residing at 10 Langford Court, Brampton, Ontario, L6W 4K4, do hereby certify that:

1. I am a principal of P & E Mining Consultants Inc. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada” (the “Technical Report”) with an effective date of July 2, 2012. 3. I graduated with an Honours Bachelor of Science degree in Geology from the University of Waterloo in 1970 and with a PhD degree in Geology from Carleton University in 1977. I have worked as a geologist for a total of 42 years since obtaining my B.Sc. degree. I am a P. Geo., registered in the Province of Saskatchewan (APEGS No. 16217), the Province of British Columbia (APEGBC No. 18965), the Province of Ontario (APGO No. 0866) and the Province of Newfoundland and Labrador (PEG No. 06005 ). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Principal, P&E Mining Consultants Inc...... 2004 – Present  Vice-President, A.C.A. Howe International Limited ...... 1992 – 2004  Canadian Manager, New Projects, Gold Fields Canadian Mining Limited ...... 1987 – 1992  Regional Manager, Gold Fields Canadian Mining Limited ...... 1986 – 1987  Supervising Project Geologist, Getty Mines Ltd...... 1982 – 1986  Supervising Project Geologist III, Cominco Ltd...... 1976 – 1982 4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible for authoring Sections 2, 4, 5 through 11, 23 and 27 as well as coauthoring Sections 1, 3 and 25 of this Technical Report 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: June 2, 2012 Signed Date: August 16, 2010

{SIGNED AND SEALED} [Wayne Ewert] ______Dr. Wayne D. Ewert P. Geo.

P&E Mining Consultants Inc., Report No. 247 Page 280 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE of AUTHOR

TRACY J. ARMSTRONG, P.GEO.

I, Tracy J. Armstrong, residing at 2007 Chemin Georgeville, res. 22, Magog, QC J1X 0M8, do hereby certify that:

1. I am an independent geological consultant contracted by P&E Mining Consultants Inc. and have worked as a geologist continuously since my graduation from university in 1982. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada” (the “Technical Report”) with an effective date of July 2, 2012. 3. I am a graduate of Queen‟s University at Kingston, Ontario with a B.Sc. (HONS) in Geological Sciences (1982). I am a geological consultant currently licensed by the Order of Geologists of Québec (License 566), the Association of Professional Geoscientists of Ontario (License 1204) and the Association of Professional Engineers and Geoscientists of British Columbia, (Licence No. 34720).

I have read the definition of "qualified person" set out in National Instrument 43-101 (NI 43-101) and certify that by reason of my education and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43-101. This report is based on my personal review of information provided by the Issuer and on discussions with the Issuer‟s representatives. My relevant experience for the purpose of the Technical Report is:

 Underground production geologist, Agnico-Eagle Laronde Mine 1988-1993  Exploration geologist, Laronde Mine 1993-1995  Exploration coordinator, Placer Dome 1995-1997  Senior Exploration Geologist, Barrick Exploration 1997-1998  Exploration Manager, McWatters Mining 1998-2003  Chief Geologist Sigma Mine 2003  Consulting Geologist 2003-to present

4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible for coauthoring of Section 12 of this Technical Report. 6. I am independent of issuer applying the test in Section 1.5 of NI 43-101. 7. I have not had prior involvement with the Property that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signing Date: August 16, 2010

{SIGNED AND SEALED} [Tracy J. Armstrong] ______Tracy J. Armstrong, P. Geo.

P&E Mining Consultants Inc., Report No. 247 Page 281 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

FRED H BROWN CPG

I, Fred H Brown, CPG, do hereby certify that:

1. I am an independent geological consultant at Suite B10, 1610 Grover St., Lynden, WA, USA. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective of July 2, 2012. 3. I am a graduate of New Mexico State University (B.Sc., 1987), and the University of the Witwatersrand (M.Sc., 2005). I have worked as an economic geologist continuously since my graduation from university in 1987. I am registered with the South African Council for Natural Scientific Professions as a Professional Geological Scientist (registration number 400008/04), the American Institute of Professional Geologists as a Certified Professional Geologist (certificate number 11015) and the Society for Mining, Metallurgy and Engineering as a Registered Member (#4152172). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience is as follows:

 Underground Mine Geologist, Freegold Mine, AAC ...... 1987-1995  Mineral Resource Manager, Vaal Reefs Mine, Anglogold...... 1995-1997  Resident Geologist, Venetia Mine, De Beers ...... 1997-2000  Chief Geologist, De Beers Consolidated Mines ...... 2000-2004  Consulting Geologist, P&E Mining Consultants Inc...... 2004-2012

4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible for authoring Section 14 of this Technical Report. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Fred Brown] ______Fred H. Brown, CPG

P&E Mining Consultants Inc., Report No. 247 Page 282 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

DAVID A. ORAVA, P. ENG.

I, David A. Orava, M. Eng., P. Eng., residing at 19 Boulding Drive, Aurora, Ontario, L4G 2V9, do hereby certify that:

1. I am an Associate Mining Engineer at P&E Mining Consultants Inc. and President of Orava Mine Projects Ltd. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada” (the “Technical Report”) with an effective date of July 2, 2012. 3. I am a graduate of McGill University located in Montreal, Quebec, Canada at which I earned my Bachelor Degree in Mining Engineering (B.Eng. 1979) and Masters in Engineering (Mining - Mineral Economics Option B) in 1981. I have practiced my profession continuously since graduation. I am licensed by the Professional Engineers of Ontario (License No. 34834119). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My summarized career experience is as follows:  Mining Engineer – Iron Ore Company of Canada...... 1979-1980  Mining Engineer – J.S Redpath Limited / J.S. Redpath Engineering...... 1981-1986  Mining Engineer & Manager Contract Development – Dynatec Mining Ltd...... 1986-1990  Vice President – Eagle Mine Contractors...... 1990  Senior Mining Engineer – UMA Engineering Ltd...... 1991  General Manager - Dennis Netherton Engineering ...... 1992-1993  Senior Mining Engineer – SENES Consultants Ltd...... 1993-2003  President – Orava Mine Projects Ltd...... 2003 to present  Associate Mining Engineer – P&E Mining Consultants Inc...... 2006 to present

4. I have not visited the property that is the subject of this Technical Report. 5. I am responsible for authoring Sections 22.3 through 22.5 as well as coauthoring 1, 3, 16, 21, 25 and 26 of the Technical Report. 6. I am an independent of the issuer applying all of the tests in Section 1.5 of NI 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [David Orava] ______David Orava, M. Eng., P. Eng.

P&E Mining Consultants Inc., Report No. 247 Page 283 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF AUTHOR

JAMES L. PEARSON, P.ENG.

I, James L. Pearson, P.Eng., residing at 5 Clubhouse Court, Bolton, Ontario, Canada, L7E 0B3, do hereby certify that::

1. I am an independent Mining Engineering Consultant, contracted by P& E Mining Consultants Inc. 2. This certificate applies to the technical report entitled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada” (the “Technical Report”) with an effective date of July 2, 2012. 3. I am a graduate of Queen‟s University, Kingston, Ontario, Canada, in 1973 with a Bachelor of Science degree in Mining Engineering. I am registered as a Professional Engineer in the Province of Ontario (Reg. No. 36043016). I have worked as a mining engineer for a total of 37 years since my graduation.

I have read the definition of "qualified person" set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:

 Review and report as a consultant on numerous exploration and mining projects around the world for due diligence and regulatory requirements;  Project Manager and Superintendent of Engineering and Projects at several underground operations in South America;  Senior Mining Engineer with a large Canadian mining company responsible for development of engineering concepts, mine design and maintenance;  Mining analyst at several Canadian brokerage firms

4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible co-authoring Sections 1, 15, 16, 21 and 25 of the Technical Report; 6. I am independent of the issuer applying all of the tests in Section 1.5 of NI 43-101. 7. I have had no prior involvement with the property that is the subject of the Technical Report. 8. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that Instrument and Form. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective date: July 2, 2012 Signing Date: August 16, 2012

{SIGNED AND SEALED} [James L. Pearson]

James L. Pearson, P. Eng.

P&E Mining Consultants Inc., Report No. 247 Page 284 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

TIMOTHY M. HAYES, P.ENG.

I, Timothy M. Hayes, P. Eng., residing in Mississauga, do hereby certify that:

1. I am employed as a project engineer with Jacobs Minerals Canada Inc. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective date of July 2, 2012. 3. I graduated with a Bachelor of Applied Science degree in Chemical Engineering from the University of Toronto in 1983. I have worked as a project engineer for a total of 22 years since obtaining my B.A.Sc. degree. I am a P. Eng., registered in the Province of Ontario (PEO No. 90220047) and hold the credential of Project Management Professional from the Project Management Institute (PMP Lic No. #1258047). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Sr. Project Engineer, Jacobs Minerals Canada Inc...... 2011 – Present  Project Engineer, Aker Metals, a division of Aker Solutions Inc...... 2008 – 2011  Applications Engineer, Komline-Sanderson Ltd ...... 1989 – 2008

4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible for authoring the portion of Item 18 and coauthored Items 24, 25 and 26 of this Technical Report, all related to the process plants. I oversaw the writing of and assume responsibility for item 17.5. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Timothy Hayes] ______Timothy M. Hayes P. Eng.

P&E Mining Consultants Inc., Report No. 247 Page 285 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

ALEXANDER DUGGAN, B.SC(HONS), M.SC, P.ENG.

I, Alexander O. Duggan, P. Eng., residing in Brampton, do hereby certify that:

1. I am employed as a Contractor in the capacity as Manager, Estimating and Planning with Jacobs Minerals Canada Inc. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective date of July 2, 2012. 3. I graduated with a Bachelor of Science, Honours degree in Civil Engineering from the University of Aston in Birmingham, England in 1982 and Master‟s degree Planning and Transportation from the University of Salford in Manchester, England in 1984. I have worked as an estimator for a total of 28 years since obtaining my M.Sc. degree. I am a P. Eng., registered in the Province of Ontario (PEO No. 100103898). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Manager, Estimating and Planning, Jacobs Minerals Canada Inc...... 2011-present  Manager, Estimating and Planning, Aker Metals, (a division of Aker Solutions Inc.) 2005-2007/2009-2011  Senior Estimator, AMEC Americas ...... 2007-2008  Senior Estimator, SNC Lavalin ...... 2004-2005

4. I have visited the Properties that are the subject of this Technical Report. 5. I am responsible for authoring the portion of Item 21.1.1 and 21.1.3 of this Technical Report, all related to the capital cost estimate of the process plants. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Alexander O. Duggan]

______Alexander O. Duggan, B.Sc(Hons), M.Sc, P. Eng.

P&E Mining Consultants Inc., Report No. 247 Page 286 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

Graham Holmes, P. ENG.

I, Graham P Holmes, P. Eng., residing in Mississauga, do hereby certify that:

1. I am employed as a process specialist with Jacobs Minerals Canada Inc. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective date of July 2, 2012. 3. I graduated with a Bachelor of Science degree in Mineral Process Engineering from the Royal School of Mines, London University in 1966. I have worked as a process engineer for a total of 46 years since obtaining my degree. I am a P. Eng., registered in the Province of Ontario (PEO No. 20196507). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

4. I visited the Property that is the subject of this Technical Report on April 24th 2012. 5. I am responsible for authoring the following sections of the report: item 1.14; 17.1; 17.2; 26.2.1 and 26.2.2 all related to the process plant. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. I have reviewed the parts of Section 13 of this Technical Report which pertain to mineral processing based on testwork performed by SGS and which were used to develop the design criteria for the concentrator. 10. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Graham Holmes] ______Graham P Holmes P. Eng.

P&E Mining Consultants Inc., Report No. 247 Page 287 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

DIOGENES A. UCEDA, P. ENG.

I, Diogenes A. Uceda, P. Eng., residing at 606 Moonrock Avenue, Sudbury, ON P3E 5Z6 do hereby certify that:

1. I am a Senior Process Engineer at Jacobs Minerals Canada Inc. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective of July 2, 2012. 3. I graduated with a Bachelor of Science degree in Metallurgical Engineer from the National University of Engineering in Lima-Peru in 1967, with a MSc degree in Colorado School of Mines in 1976 and with a PhD degree in Metallurgical Engineering from Rolla University in 1988. I have worked as a metallurgist for a total of 46 years since obtaining my B.Sc. degree. I am a P. Eng., registered in the Province of Ontario (Licensed P. Eng. No. 90349069). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Senior Process Engineer at Jacobs Minerals Canada Inc...... 2011 – Present  Principal Metallurgical Engineer, at Xstrata Mount Isa, Queensland Australia ...... …….2008 – 2010  Director of EHS, Xstrata Nickel ...... 2002 – 2008  Director of Metallurgy and Environment at Xstrata Falcondo, Dominican Republic ...... 1992 – 2002  Senior Electrometallurgist, Xstrata (Falconbridge) at Kidd Creek, Timmins ...... 1989 – 1992  Production Manager at Ilo Copper Refinery - Peru ...... 1976 – 1978  Senior Metallurgist at Cerro Verde Mines, Peru...... 1970 – 1974 4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible for authoring Item 24, Option to Produce Cobalt Sulphate and coauthored Item 17.3 SMPP Process Description, Item 26.6 Recommendations, Cobalt Sulphate Option, and Item 17.4 SMPP Plant Utilities of this Technical Report 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Diogenes Uceda] ______Dr. Diogenes A. Uceda P. Eng.

P&E Mining Consultants Inc., Report No. 247 Page 288 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

WADE H. SUMNERS, P.Ag., P. Biol.

I, Wade H. Sumners, P. Ag., P.Biol., residing at Dundurn, SK, do hereby certify that:

1. I am a Manager of Biological Services at MDH Engineered Solutions Corp., a Member of the SNC-Lavalin Group. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective of July 2, 2012 3. I graduated with a Bachelor of Science degree in Land Use and Environmental Studies from the University of Saskatchewan in 1996 and with an MSc degree in Biogeography from the University of Saskatchewan in 2005. I have worked as a biologist for a total of 16 years since obtaining my B.Sc. degree. I am a P.Ag., registered in the Province of Saskatchewan (SIA No. 20657) and a P.Biol. registered in the Province of Alberta (ASPB No.1964). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Manager of Biological Services, MDH Engineered Solutions Corp ...... 2011 – Present  Senior Biologist, MDH Engineered Solutions Corp ...... 2006 – 2011  Ecological Consultant, Saskatchewan Research Council ...... 2005  Resource Conservation Assistant, Meewasin Valley Authority ...... 2000 – 2005  Principle Investigator, Prince Albert Model Forest Association...... 1999 – 2000  Greenhouse Technician, Shand Greenhouse (SaskPower) ...... 1998 – 1999  Research Technician, Canadian Wildlife Service...... 1997 – 1998  Forestry Technician, Prairie Land Forest Service...... 1996 – 1997

4. I have visited the Property in Saskatchewan that is the subject of this Technical Report. 5. I am responsible for authoring Section 20.2 and coauthoring Sections 1 and 3 of this Technical Report 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have had prior involvement with the project that is the subject of this Technical Report, through the collection of baseline biological data and authoring an Environmental Impact Statement for the proposed Saskatchewan Metals Processing Plant. 8. I have read NI 43-101 and Form 43-101F1 and Section 20 of the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Wade Sumners] ______Mr. Wade Sumners P. Ag., P.Biol.

P&E Mining Consultants Inc., Report No. 247 Page 289 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

DAN A. MACKIE, P.ENG.

I, Dan A. Mackie, P.Eng., residing at 1109 Lakeshore Road, Burlington, Ontario, L7R 1A7, do hereby certify that:

1. I am President of Dan Mackie and Associates Inc. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada” (the “Technical Report”) with an effective date of July 2, 2012. 3. I graduated with a Bachelor of Engineering degree from McGill University in 1961. I am a registered engineer in the Province of Ontario, License No. 28242014. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  President, Dan Mackie & Associates...... 1982 – Present  Director, HG Engineering ...... 1982 – 1990  Chief Engineer, Hudson Bay Mining and Smelting ...... 1981 – 1982

4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible for authoring Sections 4.2.3, 18.2 and 26.2.5 of this Technical Report 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9 As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2010

{SIGNED AND SEALED} [Dan A. Mackie] ______Dan A. Mackie P.Eng.

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CERTIFICATE OF QUALIFIED PERSON

MARC ROUGIER, P.ENG.

I, Marc Rougier, P.Eng. residing at 1160 Forest Trail Place, Oakville, Ontario, do hereby certify that:

1. I am a Principal of Golder Associates Ltd. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective of July 2, 2012. 3. I graduated with a Bachelor of Engineering (Civil) degree from Queen‟s University, Kingston, Ontario in 1991with a B.Sc. degree in Geological Engineering, Geotechnical Option. I have worked as an engineer for a total of 21 years since obtaining my degree. I am a P.Eng., registered in the Province of Saskatchewan (APEGS No. 15666), the Province of Ontario (PEO No. 90423880), and the Northwest Territories (NAPEG Licence No. L2461). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Principal, Golder Associates Ltd...... 2012 – Present  Associate, Golder Associates Ltd...... 2006 – 2012  Geotechnical Engineer, Golder Associates Ltd...... 1997 – 2006  Geotechnical Engineer, Piteau Associates Engineering Ltd...... 1994 – 1997

4. I have visited the Property that is the subject of this Technical Report in September 2003 for three weeks and in September 2008 and for three days as part of investigations to address the mining geotechnical and physical hydrogeology aspects of the project. 5. I am responsible for authoring Item 16.3 relating to geotechnical and hydrogeological parameters relevant to mine or pit design as well as Section 1 of this Technical Report. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. (if had prior, state nature) 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Marc Rougier] ______Marc Rougier, P.Eng.

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CERTIFICATE OF QUALIFIED PERSON

KENNETH A. BOCKING, P.ENG.

I, Kenneth A. Bocking, P.Eng., residing at 254 Tracina Drive, Oakville, Ontario), do hereby certify that:

1. I am a Principal of Golder Associates Ltd. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective of July 2, 2012. 3. I graduated with a Bachelor of Engineering (Civil) degree from the University of Saskatchewan in 1974 and with a M.Sc. degree in Geotechnical Engineering from the University of Saskatchewan in 1978. I have worked as an engineer for a total of 38 years since obtaining my B.Eng. degree. I am a P.Eng., registered in the Province of Saskatchewan (APEGS No. 4131), the Province of Ontario (PEO No. 4253654), and the Northwest Territories (NAPEG Licence No. L400). I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Principal, Golder Associates Ltd...... 1998 – Present  Associate, Golder Associates Ltd...... 1986 – 1998  Geotechnical Engineer, Golder Associates Ltd...... 1984 – 1986  Geotechnical Engineer, Thurber Consultants Ltd...... 1978 – 1984

4. I have visited the Property that is the subject of this Technical Report on July 6 and 7, 2010. 5. I am responsible for authoring Item 18.1.8 and 18.1.8.1 with respect to the Tailings and Waste Rock Co-disposal Facility at the NICO Site and Item 20.1 (a) with respect to Environmental at the NICO site and (b) with respect to the Waste and Tailings Disposal, Site Monitoring and Water Management at the NICO site of this Technical Report. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. (if had prior, state nature) 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2012

{SIGNED AND SEALED} [Ken Bocking] ______Kenneth A. Bocking, P.Eng.

P&E Mining Consultants Inc., Report No. 247 Page 292 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.

CERTIFICATE OF QUALIFIED PERSON

ALEX MEZEI, P.Eng.

I, Alex Mezei, P. Eng., residing at 1009 Nornabell Ave., Peterborough, Ontario, Canada, K9K 2S8, do hereby certify that:

1 I am Director, Engineering Technology Services, Metallurgical Operations with SGS Mineral Services, Canada. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada”, with an effective of July 2, 2012. 3. I graduated with a Chemical Engineering Technology degree from the University of Timisoara, Romania in 1981, date since I worked as a metallurgist. I am a Professional Engineer registered in Ontario, # 90402769. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Director, Engineering Technology Services, SGS Mineral Services...... 2007 – Present  Senior Metallurgist / Project manager, SGS Mineral Services...... 2002 – 2004  Senior Metallurgist / Project manager, Lakefield Research Limited...... 1998 – 2002  Metallurgist / Project leader, Lakefield Research Limited...... 1994-1998  Application specialist, Asea Brown Boveri (ABB) Canada...... 1992 – 1994  Principal Research Metallurgist, Nonferrous Metallurgical Institute Bucharest, Romania ..... 1984 – 1992  Metallurgical Engineer in Training, Phoenix Nonferrous Metallurgical Complex, Romania . 1981 – 1984

4. I have not visited the Property that is the subject of this Technical Report. 5. I am responsible for reviewing items pertaining to Section 13 and 26.2.6 of this Technical Report pertaining to the hydrometallurgical flowsheet development for the recovery of cobalt, bismuth and gold. The review is based on the reports submitted by SGS under my signature, consisting of SGS proposed process descriptions, bench and pilot scale validation and subsequent generation of design criteria to be used by the client for engineering and feasibility studies. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have been involved with the project since 1998 as a Metallurgist focusing on the hydrometallurgical process testing and flowsheet development. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed date: August 16, 2012

{SIGNED AND SEALED} [Alex Mezei]

______Alex Mezei, P.Eng.

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CERTIFICATE OF QUALIFIED PERSON

WILLIAM T. HORNE, P. ENG.

I, William T. Horne, P. Eng., of EBA Engineering Consultants Ltd., Edmonton Alberta T5V 1B4, do hereby certify that:

1. I am a Principal Consultant of EBA Engineering Consultants Ltd. 2. This certificate applies to the technical report titled “Technical Report and Updated Mineral Reserve Estimate and Front-End Engineering & Design (Feed) Study on the NICO Gold-Cobalt-Bismuth-Copper Deposit Mazenod Lake Area, Northwest Territories, Canada” (the “Technical Report”) with an effective date of July 2, 2012. 3. I graduated with an Master of Science degree in Geotechnical Engineering from the University of Alberta in 1987 and with a B.Sc. degree in Civil Engineering from the University of Calgary in 1983. I have worked as a engineer for a total of 27 years since obtaining my B.Sc. degree. I am a P. Eng., registered in the Province of Alberta, the territory of Northwest Territories. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:  Principal Consultant, EBA Engineering Consultants Ltd...... 2006 – Present  Senior Project Manager, EBA Engineering Consultants Ltd...... 1999 – 2006  Geotechnical Engineer, EBA Engineering Consultants Ltd...... 1987 – 1999  Geotechnical Engineer, Acres International Ltd...... 1985 – 1987  4. I have visited the Property that is the subject of this Technical Report June 26, 2008 and July 14, 2009. 5. I am responsible for authoring Sections 18.1.2 and 18.1.3 of this Technical Report. 6. I am independent of the Issuer applying all of the tests in section 1.5 of National Instrument 43-101. 7. I have not had prior involvement with the project that is the subject of this Technical Report. 8. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith. 9. As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: July 2, 2012 Signed Date: August 16, 2010

SIGNED AND SEALED} [William T. Horne]

______Mr. William T. Horne, P.Eng.

P&E Mining Consultants Inc., Report No. 247 Page 294 of 294 NICO Gold-Cobalt-Bismuth-Copper Deposit – Fortune Minerals Ltd.