NABANGA PROJECT NI 43-101 TECHNICAL REPORT – PRELIMINARY ECONOMIC ASSESSMENT
Prepared by:
Patrick Pérez, P.Eng. DRA/Met-Chem
Yves A. Buro, P.Eng. DRA/Met-Chem
Ewald Pengel, MSc., P.Eng. DRA/Met-Chem
François Thibert, MSc., P.Geo. SEMAFO Inc.
Richard Roy, P.Geo. SEMAFO Inc.
Effective Date: September 30, 2019 Report Date: November 14, 2019
DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page i
TABLE OF CONTENTS
1 SUMMARY ...... 1 1.1 Property Description and Location ...... 1 1.2 Accessibility, Climate, Local Resources, Infrastructure, and Physiography ...... 1 1.3 History ...... 2 1.4 Geological Setting and Mineralization ...... 3 1.5 Exploration Work and Drilling ...... 4 1.5.1 Exploration ...... 4 1.5.2 Drilling...... 5 1.6 Mineral Processing and Metallurgical Testing ...... 6 1.7 Mineral Resources Estimate ...... 6 1.8 Mining Method ...... 7 1.9 Recovery Methods ...... 7 1.10 Project Infrastructure ...... 7 1.11 Market Studies and Contracts ...... 8 1.12 Environmental Studies, Permitting and Social or Community Impact ...... 8 1.13 Capital and Operating Costs ...... 8 1.13.1 Capital Cost Estimate (Capex) ...... 8 1.13.2 Operating Cost Estimate (Opex) ...... 9 1.14 Economic Analysis ...... 10 1.15 Interpretation and Conclusions ...... 11 1.16 Recommendations ...... 11 1.16.1 Mining and Geology ...... 11 1.16.2 Process...... 11 1.16.3 Environment ...... 12 2 INTRODUCTION ...... 13 2.1 Terms of Reference Scope of Study ...... 13 2.2 Qualified Persons ...... 13 2.3 Effective Date and Declaration ...... 14 2.4 Site Visit ...... 14 2.5 Units and Currency ...... 15 3 RELIANCE ON OTHER EXPERTS ...... 16 4 PROPERTY DESCRIPTION AND LOCATION ...... 17 4.1 Property Location ...... 17 4.2 Property Description ...... 18 4.2.1 Nabanga Research Permit (Exploration Permit) ...... 18 4.2.2 Industrial Operating Permit (Mining Permit) ...... 20 4.2.3 Mining Agreement ...... 20 4.3 Royalties, Back-in Rights, Payments, Encumbrances ...... 20 4.4 Environmental Liabilities...... 21 4.5 Required Permits ...... 21 4.6 Significant Factors or Risks ...... 21 4.7 Conclusions ...... 21
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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 22 5.1 Accessibility ...... 22 5.2 Climate ...... 22 5.3 Local Resources and Infrastructure ...... 24 5.4 Physiography ...... 24 5.5 Vegetation ...... 24 5.6 Surface Rights ...... 24 6 HISTORY ...... 25 6.1 Prior Ownership and Ownership Changes ...... 25 6.2 Previous Mineral Exploration Work ...... 25 6.3 Previous Mineral Resources and Production ...... 26 6.3.1 Historical Resources – Snowden 2012 ...... 26 6.3.2 Historical Resources – Snowden 2015 ...... 27 6.3.3 Historical Resources – Conclusions ...... 28 6.4 Previous Mineral Production ...... 28 7 GEOLOGICAL SETTING AND MINERALIZATION ...... 29 7.1 Regional Geology ...... 29 7.2 Structural Evolution of the Eburnean Orogeny ...... 30 7.3 Local Geology ...... 31 7.4 Property Geology ...... 31 7.5 Mineralization ...... 33 7.6 Alteration ...... 34 7.7 Structure ...... 34 8 DEPOSIT TYPES ...... 35 9 EXPLORATION ...... 36 10 DRILLING ...... 38 10.1 Drilling by Orbis ...... 38 10.1.1 Methodology ...... 38 10.1.2 Drill Hole Collar Location ...... 39 10.1.3 Downhole Deviation Survey ...... 39 10.1.4 Logging Core and RC Chips ...... 40 10.1.5 RC Sampling by Orbis ...... 40 10.1.6 Core Sampling by Orbis...... 41 10.1.7 Drill Samples Recovery ...... 41 10.2 Drilling by SEMAFO ...... 42 10.2.1 2016 Drilling Program ...... 42 10.2.2 2017 Drilling Program ...... 43 10.2.3 2018 Drilling Program ...... 44 10.2.4 Drill Hole Collar Locations ...... 45 10.2.5 Downhole Deviation Survey ...... 45 10.2.6 Logging Core and RC Chips ...... 46 10.2.7 Magnetic Susceptibility Determinations ...... 46 10.2.8 RC Sampling ...... 46 10.2.9 Core Sampling ...... 46 10.2.10 Drill Sample Recovery ...... 47 10.2.11 Conclusions ...... 47
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11 SAMPLE PREPARATION, ANALYSIS AND SECURITY ...... 48 11.1 Samples from Orbis – 2010-2013 Programs ...... 48 11.1.1 Introduction ...... 48 11.1.2 Sample Preparation ...... 48 11.1.3 Sample Analysis ...... 48 11.1.4 Internal QA/QC Procedures ...... 48 11.1.5 Laboratories Certification and Accreditation ...... 49 11.1.6 QA/QC Protocol – Orbis ...... 49 11.1.7 Samples Security, Chain of Custody ...... 51 11.1.8 Specific Gravity ...... 52 11.1.9 Conclusions ...... 52 11.2 Samples from SEMAFO – 2016-2018 Programs ...... 52 11.2.1 Sampling...... 52 11.2.2 QA/QC Protocol – SEMAFO ...... 52 11.2.3 Introduction ...... 52 11.2.4 Certified Reference Materials (CRMs; "Standards") ...... 53 11.2.5 Coarse Duplicate Samples ...... 55 11.2.6 Coarse Blank Samples ...... 60 11.2.7 Sample Preparation and Analysis ...... 60 11.2.8 SEMAFO's Mana Mine Laboratory ...... 61 11.2.9 Density...... 62 11.2.10 SEMAFO's Secondary Laboratory Certification and Accreditation ...... 62 11.2.11 Conclusions, Recommendations ...... 62 12 DATA VERIFICATION ...... 64 12.1 Twinning of RC Holes by Orbis ...... 64 12.2 Orbis Database (Assay) by Snowden ...... 68 12.3 Site Visits ...... 68 12.4 Verifications by DRA/Met-Chem ...... 68 13 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 69 13.1 Metallurgical Test Work ...... 69 13.1.1 Comminution Test Work ...... 69 13.1.2 Cyanidation Test Work ...... 70 13.1.3 Flotation Test Work ...... 70 14 MINERAL RESOURCE ESTIMATES ...... 71 14.1 Introduction ...... 71 14.2 Drill Hole Database Construction and Validation ...... 71 14.3 Geological Modelling ...... 72 14.4 Basic Statistics Calculations ...... 73 14.5 Composite Data ...... 73 14.6 Capping ...... 74 14.7 Density ...... 74 14.8 Variogram Modelling ...... 75 14.9 Resource Block Modelling ...... 80 14.10 Grade Interpolation Methodology ...... 81 14.11 Resource Validation ...... 82 14.12 Resources Definitions and Classification ...... 83
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14.12.1 CIM Guidelines & Definition ...... 83 14.12.2 Nabanga Classification ...... 85 14.13 Mineral Resource Statement ...... 85 15 MINERAL RESERVE ESTIMATES ...... 86 16 MINING METHOD ...... 87 16.1 Geotechnical Assessment ...... 87 16.2 Hydrology and Hydrogeology ...... 88 16.3 Transition Between Open Pit and Underground Mining ...... 88 16.4 Open Pit Mining ...... 89 16.4.1 Pit optimization ...... 89 16.4.2 Cut-Off Grade (COG) ...... 90 16.4.3 Pit Design and Sequencing ...... 90 16.5 Underground Mining ...... 93 16.5.1 Mining Method and Cut-Off Grade (COG) ...... 93 16.5.2 Mine Design ...... 96 16.5.3 Mine Services ...... 96 16.5.4 Operation ...... 100 16.5.5 Mine Production Plan ...... 100 17 RECOVERY METHODS ...... 107 17.1 Nabanga Processing Plant ...... 107 17.2 Design Criteria ...... 107 17.3 Material Balance and Water Balance ...... 108 17.4 Flow Sheets and Process Description ...... 109 17.5 Crushing ...... 110 17.6 Grinding...... 112 17.7 Sulphide Flotation ...... 112 17.8 Intensive Cyanidation and Carbon-in-Leach (CIL) Cyanidation...... 113 17.9 Elution and Carbon Regeneration and Electro-winning and Refining ...... 115 17.10 Grinding Media, Concentrator Reagents and Refinery Chemicals ...... 115 17.10.1 Grinding Media ...... 115 17.10.2 Flotation reagents ...... 115 17.10.3 Leach Chemicals ...... 116 17.10.4 Carbon System Chemicals ...... 116 17.10.5 Refinery Fluxes ...... 116 17.10.6 Other Chemicals ...... 116 17.11 Water and Air Services...... 117 17.11.1 Water Supply: ...... 117 17.11.2 Air Supply: ...... 117 17.12 Overall Gold Recovery ...... 118 18 PROJECT INFRASTRUCTURE ...... 119 18.1 Roads ...... 122 18.1.1 Main Access Road ...... 122 18.1.2 Site Roads ...... 122 18.1.3 Mine Roads ...... 122 18.2 Power Supply ...... 123 18.2.1 Power Demand ...... 123
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18.2.2 Power Plant and distribution – reticulation network ...... 124 18.2.3 Reticulation network ...... 124 18.2.4 Main Electrical Equipment ...... 124 18.3 Control System ...... 125 18.3.1 Automation Process Network ...... 125 18.3.2 Process Control System ...... 126 18.3.3 Wiring and Junction Boxes ...... 126 18.3.4 SCADA ...... 126 18.3.5 SCADA and PLC Power Sources ...... 126 18.3.6 Redundancy ...... 127 18.3.7 Process Analog Instruments ...... 127 18.4 Communication System (Local and External) ...... 127 18.4.1 Telecommunication Local System ...... 127 18.4.2 Telecommunication and Mobile Radio Systems ...... 127 18.4.3 Telecommunication Services ...... 127 18.4.4 Telecommunications Distribution ...... 128 18.4.5 Corporate Network ...... 128 18.4.6 Camera System ...... 128 18.5 Tailings Storage Facility ...... 128 18.6 Site Buildings ...... 129 19 MARKET STUDIES AND CONTRACTS ...... 130 19.1 Contracts ...... 130 20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT ...... 131 20.1 Legal Framework and Permits to Obtain ...... 131 20.1.1 Policies and Strategies for Environmental Protection ...... 132 20.1.2 Legal Framework ...... 132 20.1.3 Mining Code ...... 133 20.1.4 Institutional Framework...... 134 20.1.5 Permits to Obtain ...... 135 20.2 Baseline Studies ...... 138 20.3 Project Impacts, Risk Analysis, Environmental and Social Management Plan ...... 138 20.3.1 Risk Analysis ...... 139 20.3.2 Environmental and Social Management Plan ...... 139 20.4 Acid Rock Drainage & Waste Disposal ...... 139 20.5 Closure, Decommissioning, and Reclamation ...... 139 20.6 Conclusion ...... 140 21 CAPITAL AND OPERATING COSTS ...... 141 21.1 Capital Cost Estimate (Capex) ...... 141 21.1.1 Scope of the Estimate ...... 141 21.1.2 Capital Costs Summary ...... 142 21.1.3 Basis of Estimate General ...... 143 21.1.4 Methodology ...... 145 21.1.5 Qualifications ...... 147 21.2 Operating Cost Estimate (Opex) ...... 149 21.2.1 Introduction ...... 149 21.2.2 Summary of SEMAFO Personnel Requirements ...... 149 21.2.3 General and Administrative Costs ...... 149
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21.2.4 Mining Operating Costs ...... 150 21.2.5 Process Operating Costs ...... 150 22 ECONOMIC ANALYSIS ...... 152 22.1 Assumptions ...... 152 22.1.1 Macro-Economic Assumptions ...... 152 22.1.2 Mineral Royalties ...... 153 22.1.3 Taxation Regime ...... 153 22.1.4 Project Financing ...... 153 22.1.5 Technical Assumptions ...... 153 22.2 Cash Flow Analysis and Financial Results ...... 153 22.3 Sensitivity Analysis ...... 158 23 ADJACENT PROPERTIES ...... 160 24 OTHER RELEVANT DATA AND INFORMATION ...... 161 25 INTERPRETATION AND CONCLUSIONS ...... 162 25.1 Conclusions ...... 162 25.2 Risk Evaluation ...... 162 26 RECOMMENDATIONS ...... 163 26.1 Mining and Geology ...... 163 26.2 Geotechnical ...... 163 26.3 Hydrogeology and Hydrology ...... 163 26.4 Process Recommendations ...... 164 26.4.1 Trade-Off Studies ...... 164 26.4.2 Other Test Works ...... 165 26.5 Environment ...... 165 26.6 Proposed Work Program ...... 165 27 REFERENCES ...... 166 28 ABBREVIATIONS ...... 167 29 CERTIFICATES OF QUALIFIED PERSONS ...... 178
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LIST OF TABLES
Table 1.1– Nabanga Deposit – Historical Mineral Resources (at 0.5 g/t Au Cut-Off) – August 2012 ...... 3 Table 1.2 – Nabanga Deposit – Historical Mineral Resources (at 5.0 g/t Au Cut-Off) – June 2015 ...... 3 Table 1.3– Nabanga Permit –Holes Drilled by Orbis in 2010-2013 ...... 5 Table 1.4 – Summary of the Drill Programs by SEMAFO in 2016-2018 ...... 6 Table 1.5 – Nabanga – Mineral Resources Estimate as of December 31, 2018, using a Cut-Off Grade of 3.0 g/t Au ...... 6 Table 1.6– Initial Capex Summary ...... 9 Table 1.7 – Summary of Operating Costs ...... 10 Table 1.8 - Financial Model Indicators...... 10 Table 2.1 – Qualified Persons and their Respective Sections of Responsibilities ...... 13 Table 4.1 – Nabanga Permit Boundaries (UTM, WGS84, Zone 31N) ...... 18 Table 5.1 – Nabanga Region – Summary of Monthly Weather Averages ...... 22 Table 6.1 – Nabanga Deposit – Historical Mineral Resources (at 0.5 g/t Au Cut-Off) – August 2012 ...... 27 Table 6.2 – Nabanga Deposit – Historical Mineral Resources (at 5.0 g/t Au Cut-Off) – June 2015 ...... 27 Table 9.1 – Summary of the Auger Holes drilled Entirely or Partly on the Nabanga Permit in 2018 ...... 36 Table 10.1 – Nabanga Permit – 2010-2013 Holes Drilled by Orbis ...... 38 Table 10.2 – Summary of the 2016-2018 Drill Programs by SEMAFO ...... 44 Table 11.1 – List and Details on the CRMs Used by SEMAFO ...... 53 Table 11.2 – QC Samples – Comparison of the Results from the Pairs of Duplicate Analyses from Diamond Drill Holes (SMF BF Laboratory) ...... 55 Table 11.3 – Results from the CRMs Analyzed at ALS Ouagadougou ...... 56 Table 12.1 – Comparison of Twinned Holes NARC033 and NADD023 (5 m separation) ...... 65 Table 12.2 – Comparison of Twinned Holes NARC184 and NADD002 (2 m separation) ...... 66 Table 12.3 – Comparison of Twinned Holes NARC134 and NADD013 (1 m separation) ...... 67 Table 12.4 – Comparison of Twinned Holes NARC154 and NADD012 (3 m separation) ...... 67 Table 13.1 – Bond Work Index Tests ...... 69 Table 13.2 – SMC Test Results ...... 69 Table 13.3 – Standard Leach Test 0.035% w/v NaCN Concentration Test Results ...... 70 Table 13.4 – Intensive Leach Test 0.5% w/v NaCN Concentration Test Results ...... 70 Table 14.1– Summary of Database Entries by Hole Type...... 72 Table 14.2 – List of Fields Contained in the Drill Hole Database ...... 72 Table 14.3 – General Statistics on the Gold Values (Au) within the HG Wireframe ...... 73 Table 14.4 – Nabanga Block Models Parameters ...... 80 Table 14.5 – Interpolation Parameters ...... 81 Table 14.6 – Basic Statistical Parameters from the Assays, Composites, and Blocks ...... 82 Table 14.7 – Nabanga – Mineral Resources Estimate as of December 31, 2018, using a COG of 3.0 g/t Au ...... 85 Table 16.1 – Value of Mineralized Material Every 20 m Slice for Underground and Open Pit Scenarios ...... 89 Table 16.2 – Pit Optimization Parameters ...... 90 Table 16.3 – Open Pits Statistics ...... 92 Table 16.4 – UG Cut-Off Grade Parameters ...... 94 Table 16.5 – Nabanga Life of Mine Plan Statistics ...... 102 Table 17.1 – Design Criteria ...... 108 Table 17.2 – Concentrator Summarised Process Mass Balance ...... 108 Table 17.3 – Crusher Parameters ...... 110 Table 17.4 – Grinding Parameters ...... 112 Table 17.5 – Flotation Parameters ...... 113 Table 17.6 – Leaching Parameters ...... 114
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Table 17.7 – Grinding Media and Reagent Consumption...... 117 Table 17.8 – Gold Recovery for the Nabanga Mineralized Material ...... 118 Table 18.1 – Estimated Total Power Demand Consumption ...... 123 Table 18.2 – Tailings Design Basis ...... 129 Table 20.1 – List of Required Permits ...... 136 Table 21.1 – Initial Capex Summary ...... 142 Table 21.2 – Currency Exchange Rates ...... 144 Table 21.3 – Summary of Operating Costs ...... 149 Table 21.4 – SEMAFO Personnel Requirement ...... 149 Table 21.5 – Summary of General and Administrative Costs ...... 150 Table 21.6 – Summary of Mining OPEX ...... 150 Table 21.7 – Summary of Process OPEX ...... 151 Table 22.1 – Financial Model Indicators ...... 152 Table 22.2 – Macro-Economic Assumptions ...... 152 Table 22.3 – Project Evaluation Summary – Base Case ...... 155 Table 22.4 - Cash Flow Statement – Base Case ...... 156 Table 26.1 – Estimated Budget for Next Phase ...... 165
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LIST OF FIGURES
Figure 4.1 – Nabanga Location Map ...... 17 Figure 4.2 – Boundaries of the Original Yactibo Group Permit ...... 19 Figure 4.3 – Nabanga Permit Location ...... 19 Figure 5.1 – Wind Rose for Fada N'Gourma ...... 23 Figure 7.1 – Simplified Geological Map of the Leo–Man Shield ...... 30 Figure 7.2 – Nabanga Property Geology ...... 32 Figure 9.1 – Summary of Exploration and Drilling on the Nabanga Permit (Black and Red: Holes by Orbis; Blue, Yellow and Green: Holes by SEMAFO)...... 37 Figure 10.1 – Histogram of Recovery of RC Drilling by Orbis ...... 42 Figure 10.2 – Drilling Program 2016-2018: Holes in Grey Auger and in Black RC/DD...... 45 Figure 11.1 – Results from the Main Standard used by Orbis ...... 50 Figure 11.2 – Scatterplot of Original and Field Duplicate Samples ...... 51 Figure 11.3 – QC Samples – CRMs – Time Series Standardized Variable (Z-score); (Extreme Values Removed) 54 Figure 11.4 – QC Samples – CRMs – Time Series Moving Average ...... 54 Figure 11.5 – QC Samples – Comparison of the Results from the Pairs of Coarse Duplicates (Laboratory Rejects) from Diamond Drill Holes Analyzed by the SMF BF Laboratory – Mean vs. Half Relative % Difference Plot...... 57 Figure 11.6 – QC Samples – Comparison of the Results from the Pairs of Coarse Duplicates (Laboratory Rejects) from Diamond Drill Holes Analyzed by ALS – Q-Q Plot ...... 58 Figure 11.7 – QC Samples – Comparison of the Results from the Pairs of Coarse Duplicates (Laboratory Rejects) from Diamond Drill Holes Analyzed by ALS – Precision Plot (Mean vs Half Absolute % Difference) ...... 59 Figure 14.1 – Variograms of 1 m Composites for South Zone (Short and Long) ...... 77 Figure 14.2 – Variograms of 1m Composites for Central Zone (Short and Long) ...... 78 Figure 14.3 – Variograms of 1m Composites for North Zone (Short and Long) ...... 79 Figure 14.4 – Map of the 3 Grids Used for Block Models ...... 80 Figure 16.1 – Double Lane Ramp Design ...... 91 Figure 16.2 – Nabanga Pit Designs...... 91 Figure 16.3 – Pit Sequencing ...... 92 Figure 16.4 – Open Pit Sequence ...... 93 Figure 16.5 – Longitudinal View of the MSO Results Per Zone ...... 95 Figure 16.6 – Longitudinal Retreat Mining ...... 95 Figure 16.7 – Isometric View – Nabanga Underground Mine ...... 98 Figure 16.8 – Longitudinal View – Nabanga Underground Mine ...... 98 Figure 16.9 – Ventilation Network ...... 99 Figure 16.10 – Underground Production Plan ...... 101 Figure 16.11 – Underground Development Plan ...... 101 Figure 16.12 – Underground Development at Year 2 ...... 103 Figure 16.13 – Underground Development at Year 3 ...... 103 Figure 16.14 – Underground Development at Year 4 ...... 104 Figure 16.15 – Underground Development at Year 5 ...... 104 Figure 16.16 – Underground Development at Year 6 ...... 105 Figure 16.17 – Underground Development at Year 7 ...... 105 Figure 16.18 – Underground Development at Year 8 ...... 106 Figure 16.19 – Underground Development at Year 9 ...... 106 Figure 17.1 – Concentrator Water Balance ...... 109 Figure 17.2 – Simplified Flow Sheet of Nabanga Mineral Processing Facility ...... 111 Figure 18.1 – General Site Layout ...... 120 Figure 18.2 – Process Plant Layout ...... 121 Figure 22.1 – After-Tax Cash Flow and Cumulative Cash Flow Profiles ...... 154
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Figure 22.2 – Pre-Tax NPV5 % & IRR Sensitivity to Capital Expenditure, Operating Cost and Price ...... 158 Figure 22.3 – After-Tax Royalty NPV5 % & IRR Sensitivity to Capital Expenditure, Operating Cost and Price .... 159
LIST OF APPENDICES
Appendix A – Process Design Criteria Appendix B – Process Plant Mass Balance
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1 SUMMARY
1.1 Property Description and Location
The Nabanga exploration permit ("Permis de recherche") is located in the south-east of Burkina Faso, approximately 250 km to the southeast of Ouagadougou. The permit is centered on latitude 11° 18' 03” North and longitude 0° 29' 58” East in the geographic coordinate system. The Nabanga Permit has an irregular shape covering 178.50 km².
The southeastern sector of the Project is approximately 5 km to the NW extremity of the Kompienga Lake formed by the hydro-electrical dam built some 25 km further to the SE.
SEMAFO formerly held five contiguous exploration permits collectively known as the Yactibo Permit Group. Two (2) permits were relinquished and the Kamsongo, Napadé and the Nabanga permits were retained.
The Nabanga Permit is registered under the name of Birimian Resources Sarl, which is 100% owned by SEMAFO. The latest renewal brings the expiry date of the Nabanga Permit to April 1, 2020.
The Nabanga Permit is mainly regulated by the 2003 Mining Code. The Mining Code gives the exploration permit holder the exclusive right to explore for the minerals requested, and the right to convert it into an Industrial Operating Permit (Permis d'exploitation industrielle). The exploration permit is accompanied by a Mining Agreement (Convention minière) that supplements the provisions of the Mining Code.
The permit is subject to a 10% carried ownership interest to the Government of Burkina Faso, and a graduated royalty based on the gold price. A Net Profit Royalty (NPR) of 1% is assigned to the original vendor.
1.2 Accessibility, Climate, Local Resources, Infrastructure, and Physiography
The Property is accessible from Ouagadougou by travelling 220 km to the east via all-weather, paved (Route Nationale) RN04 to Fada-N’Gourma. Proceeding approximately 80 km southward along paved RN18 will lead to about 15 km east of the Property. A gravel road that crosses the Kompienga River provides the main access to the Property and different sectors of the Property can be reached by various roads, drill tracks and trails.
The Property is located in the southern portion of the Sahelian-Sudanian zone, under semi-arid climatic conditions characterized by a rainy season from May to September and a dry season between October and April, when the harmattan, a dry wind blowing off the Sahara Desert, carries large amounts of dust over hundreds of kilometers. The highest average temperature of 32.1°C was recorded in April, and the lowest averaged 25.0°C in January.
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Fada-N'Gourma is the nearest town where basic services and supplies are available. It is serviced by a hospital, hotels, and a 991 m long dirt-surface landing strip. Specialized services and equipment have to be sourced in Ouagadougou. Mining manpower would likely be from local and expatriate personnel. Power will probably come from on-site generation, but may be from the Kompienga dam power station or from a hydro-power line10 km away from the project. The Kompienga River and Lake, as well as a tributary that crosses the center of the northern tier of the Permit may represent a source of water for a mining operation.
1.3 History
The Nabanga Permit was originally granted to Mr. Sayouba Sawadogo in 2008 for a surface area of 238 km2 that was reduced by 25% to its current 178.50 km2 upon the second renewal. The permit was subsequently acquired by Birimian Resources Sarl, a company 100% owned by SEMAFO under the acquisition of Orbis Gold Limited (Orbis) in March 2015.
Birimian Resources was a wholly-owned subsidiary of Mt Isa Metals Limited, an Australian mining company that changed its name for Orbis in 2012. Five (5) contiguous exploration permits forming the Yactibo Permit Group were acquired between 2007 and 2011.
No exploration is known to have occurred on the Yactibo permits prior to 2008. Orbis initiated exploration on the Nabanga property in 2010 with reconnaissance and field mapping, delineating the artisanal mining sites and interpreting satellite imagery data.
Several soil sampling programs were initiated across the Yactibo Permit Group, as well as airborne magnetic-radiometric survey that showed the Nabanga structure extending over approximately 9 km.
In 2011, Orbis had confirmed the presence of a gold deposit on the Nabanga 500 m long Central and 600 m long North Zones by RC holes.
Additional drilling in 2011-2012 provided the data for the initial resource estimate prepared by Snowden in 2012. Deeper holes were drilled in 2013 and soil sampling and Induced Polarization (IP) surveys identified possible strike extensions and new mineralized bodies. Several phases of historical metallurgical test work were completed for the Nabanga Project.
An initial Joint Ore Reserves Committee (JORC) compliant resources estimate was performed by Snowden in August 2012 (Table 1.1). The estimate is based on 239 RC and 14 diamond drill holes and a cut-off grade of 0.5 g/t Au, a minimum mining width of 1.5 m and a top cut at 70 g/t Au. Grade interpolation was performed using the Ordinary Kriging method.
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Table 1.1– Nabanga Deposit – Historical Mineral Resources (at 0.5 g/t Au Cut-Off) – August 2012
Resources Tonnes Grade Contained Gold Material Category (Mt) (g/t Au) (oz x 1,000) *) Oxide 0.2 5.14 30.9 Inferred Fresh 3.0 6.53 628.5 Total 3.2 6.45 659.4
*) Small discrepancies may occur due to rounding.
The 2012 resources were revised by Snowden in 2015 (Table 1.2) based on the same drill data, but applying modified parameters, notably a 5.0 g/t Au cut-off, and following the CIM Definition and the NI 43-101 Standards of Disclosure.
Table 1.2 – Nabanga Deposit – Historical Mineral Resources (at 5.0 g/t Au Cut-Off) – June 2015
Resources Tonnes Grade Contained Gold Material Category (Mt) (g/t Au) (oz x 1,000) *) Oxide 0.08 8.9 20.0 Inferred Fresh 1.76 10.1 570.0 Total 1.84 10.0 590.0
*) Small discrepancies may occur due to rounding.
The Reader is cautioned that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.
The Property has not been the subject of prior commercial gold production, but gold has been extracted for an unknown period of time from shallow artisanal workings.
1.4 Geological Setting and Mineralization
Burkina Faso is located in the Leo-Man Shield that forms the southern exposure of the West African Craton (WAC). The Leo-Man Shield comprises an Archean nucleus, the Kénéma–Man domain (3.60–2.70 Ga) centered on Liberia, and the Paleoproterozoic Baoulé-Mossi domain (2.30 – 2.00 Ga) predominant in Burkina Faso. The Baoulé-Mossi Domain consists of granite-gneiss terrains, “Birimian” greenstone belts and basins and fluvio-deltaic basins of the Tarkwaian Sequence. Sediments of Neoproterozoic age overlie the Paleoproterozoic basement in the SE of Burkina Faso (Volta sedimentary basin) and skirt the western edge of the country (Taoudeni basin).
The N-NE trending, elongated greenstone belts comprise mafic volcanic rocks and intermediate and acid effusive suites that are typical of volcanic arc environments, interbedded with volcano- sedimentary units. They are bound by older granite-gneiss terrains and have been intruded by syn- to late-tectonic granite bodies.
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The present-day structural trend of litho-tectonic units and the dominant fabrics in the WAC formed during the Eburnean orogeny. An early phase of contractional (collisional) deformation and medium- to high-grade metamorphism followed by two (2) phases of transcurrent tectonics and low-grade metamorphic overprint is generally recognized.
The Yactibo Permit Group straddles a major NE-trending shear separating the Youga Belt in the northwest from the Diapaga Belt in the southeast, the latter hosting the Nabanga deposit and being the easternmost Belt in Burkina Faso.
The Diapaga Belt is dominantly comprised of intermediate volcanic rocks, sediments, granite and gneiss, with subordinate common mafic volcanic rocks and intrusive complexes. Conglomeratic sediments are mapped as Tarkwaian equivalents. Early banded or and/or foliated granitoids are also very common.
The main feature on the Nabanga property is the granodiorite hosting the mineralization. The gold- bearing structure is associated with a sheared quartz vein containing a large proportion of the high gold values, bracketed by an alteration halo of lower gold grade accompanied by minor amounts of pyrite. The Nabanga mineralization is manifest on surface by a NE-trending zone of shallow artisanal gold mining site that extends over 3.5 km. The gold mineralization is hosted in a single structure extending over 3.5 km, but traceable over several kilometers to the southwest, based on magnetic surveys and exploration work. The deposit is of high-grade gold type, which is evidenced by a moderate reduction of the contained gold ounces as the cut-off grade is raised.
The overall strike of the Nabanga mineralized structure is NE-SW and the dip varies between 55° and 70° toward the northwest. The mineralization is arranged as a series of shoots plunging at 45º to the NE and has an average thickness of 3.2 m. The Nabanga deposit is divided into the North, Central and Southern Extension Zones by three (3) cross faults.
The weathering profile in the Nabanga deposit area is relatively shallow, generally reaching a depth of 10 m to 21 m.
1.5 Exploration Work and Drilling
1.5.1 EXPLORATION
Exploration by SEMAFO on the Nabanga and the adjacent permits included establishment of a line grid, field mapping, outlining the different artisanal mining sites, geochemical surveying and an IP/Resistivity and magnetometer survey.
In 2018, 38 quarter core samples were submitted for whole-rock analysis in order to characterize the geochemical signature of the hydrothermal alteration associated with the mineralization.
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In addition, 8,530 auger holes were drilled to a nominal depth of 5 m for a cumulative total of 41,923 m, which generated 16,994 samples. Generally, auger drilling returned anomalous values that confirmed the regional NE geophysical trends in the Yactibo permits.
1.5.2 DRILLING
A total of 441 holes were drilled by Orbis between 2010 and 2013, as listed in Table 1.3.
Table 1.3– Nabanga Permit –Holes Drilled by Orbis in 2010-2013
RC Basis of Minimum Maximum Total Date - From Date - To Type Count Pre- Snowden Length Length Meterage collar Resources Nov. 24, 2010 Dec. 11, 2010 RC 31 21.0 150.0 2,824.00 √ Jun. 4, 2011 Aug. 3, 2011 RC 60 17.0 114.0 4,392.00 √ Nov. 4, 2011 Jun. 17, 2012 RC 270 18.0 210.0 29,003.00 √ Feb. 20, 2012 May 10, 2012 DDH 28 17.5 443.0 6,821.25 3,481.0 √ Nov. 22, 2012 Feb. 13, 2013 DDH 47 19.0 432.2 12,641.33 8,324.5 Feb. 12, 2013 Mar. 20, 2013 RC 5 139.0 200.0 761.00 Total RC 366 17.0 210.0 36,980.00 Total DDH 75 17.5 443.0 19,462.58 11,723.5 Grand Total 441 58,442.58
A combination of 103 RC and core holes were drilled by SEMAFO between 2016 and 2018, as listed in Table 1.4.
In 2016, SEMAFO drilled into the SW Extension of the Nabanga mineralized structure, in an attempt to upgrade the potential of the property and was successful explaining the soil anomalies, although none was found to be of economic potential.
Drilling in 2017 was aimed primarily at six (6) proximal targets, parallel to and along, the southwest extension of the Nabanga deposit. Generally, the holes returned low gold values and suggested lack of continuity. SEMAFO re-logged the RC chips from previous RC holes and incorporated the new data into the database.
In 2018, SEMAFO updated the database used by Snowden in 2012 and 2015, added the holes drilled by Orbis in 2012-2013 and the holes drilled by SEMAFO in 2016 and 2017.
SEMAFO re-interpreted the geology and structure of the Nabanga deposit and updated the 3D geology model. The resulting interpretation suggested a shallower northerly plunge than previously interpreted of the high-grade shoots, strongly suggesting they could extend along strike and beyond the depth of 200 m. A program of 25 diamond drill holes successful supported the new interpretation of higher-grade gold mineralization arranged as shallowly-plunging shoots. This, in contrast to the
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previous interpretation by Orbis, strongly suggested that the deposit is still open at depth and to the north, along the extension of the plunging shoots.
Table 1.4 – Summary of the Drill Programs by SEMAFO in 2016-2018
DIAMOND RC HOLES DRILL HOLES YEAR Total Length Total Length Count Count (m) (m)
2016 15 1,959.0 - - 2017 63 6,861.0 - - 2018 - - 25 7,148.2 TOTAL 78 8,820.0 25 7,148.2 GRAND TOTAL 103 holes for a total of 15,968.20 m
1.6 Mineral Processing and Metallurgical Testing
The tests performed included ball mill work index test and SMC tests. ALS Ammtec performed two (2) Bond Work Index tests. The material can be considered very hard and medium abrasive.
ALS Ammtec performed cyanidation tests. Process development undertaken during the PEA aimed at establishing a conventional sulphide gold/silver processing facility considering the constructability, operability and maintainability of the processing facility. The results indicate that material has refractory characteristics as a high concentration of cyanide is required to obtain good leach results.
1.7 Mineral Resources Estimate
The Mineral Resources are stated using an average density and are constrained in a preliminary optimized pit. The resources are estimated exclusively for the mineralization within the fresh (sulphide) portion of the Nabanga deposit. Table 1.5 presents the results from the resource estimate.
Table 1.5 – Nabanga – Mineral Resources Estimate as of December 31, 2018, using a Cut-Off Grade of 3.0 g/t Au
Tonnage Grade Total Gold Resource Category (Mt) (g/t Au) (x1,000 oz) Inferred 3.4 7.7 840
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1.8 Mining Method
The current mine plan is to mine Nabanga mineral resources via open pit mining and underground mining methods. It is estimated that approximately 616 Kt of mineralized material is extractable by conventional surface mining operations and 2.364 Mt of mineralized material is extractable by underground methods.
The total life of the mine has been estimated at approximately nine (9) years, including the preproduction period.
A production schedule (mine plan) was developed for the Project to produce a total of 571K oz of gold over the life of mine, using the mill recovery of 92% and a processing rate of 360 Kt of mineralized material per year.
1.9 Recovery Methods
The Nabanga concentrator is designed to process a nominal 360,000 tonnes of Run-of-Mine (ROM) mineralized material per year and produce 571,000 troy ounces of gold over the life of mine.
The ROM mineralized material will crush by a jaw crusher and transported the stockpile by conveyor. The crushed material is reclaimed from the stockpile by apron feeder to a SAG mill. The SAG mill discharge is screened, and the screen oversize is returned back in the SAG mill. The SAG screen undersize is pumped to the ball mill circuit in closed circuit with the cyclones. The cyclone overflow flows to the flotation circuit. The flotation concentrate will ground to 25 micron and processed in an intensive leach reactor, while the flotation tailings will be processed in a more conventional CIL circuit, followed by the carbon elution, electro-winning and refining. The estimated combined recovery gold recovery is 92%.
1.10 Project Infrastructure
The existing (Route Nationale) RN18 connects the town of Fada-N’Gourma to the border between Burkina Faso and Togo. Starting approximately 80 km South of Fada-N’Gourma, a new access road connecting the RN18 to the Nabanga Process Site will be developed over a distance of about 11 km.
The power demand of SEMAFO site was determined to be 5.8 MW based on the estimated connected loads, running loads and running power for the process operation. The power plant consists of five (5) generators using Light fuel oil (LFO) fuel (four in operation, one in stand-by, maintenance or repair).
In addition to the concentrator building that will house the processing equipment, the site will include administration and mine offices, an accommodation camp and cafeteria, a warehouse, a metallurgical laboratory, and a security gate.
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1.11 Market Studies and Contracts
SEMAFO is an existing producer and seller of gold. Historical data for gold prices, therefore, exists and as such no formal market studies for this Technical Report have been undertaken.
A long-term gold price projection of $1,300 US per ounce for this Technical Report is considered reasonable on the basis of actual gold prices which reflect SEMAFO conservative outlook of the future market for gold.
1.12 Environmental Studies, Permitting and Social or Community Impact
The project site is located approximately 250 km south-east of Ouagadougou, the country’s capital. Access to Nabanga is by means of (Route Nationale) RN04, an all-weather paved road from Ouagadougou, the capital of Burkina Faso, through Fada N’Gourma. From there, travel is via RN18, an all-weather paved road to within approximately 15 km of the Nabanga Project. An unsealed dirt road, which crosses the Kompienga River, is then used to access the Nabanga property approximately 15 km to the west of RN18, although other similar dirt access roads can be used to access other parts of the property.
As part of the environmental approval process in Burkina Faso, SEMAFO will carry out an Environmental and Social Impact Assessment (“ESIA”) for the Nabanga Project. Currently, little has been done to collect environmental and social data in the Project area since the Project is only at an early stage of study.
This Project being still at an early stage, no formal process for the Environmental and Social Impact Assessment (“ESIA”) has been started at this stage.
The typical approach that will be developed by SEMAFO during the future ESIA will take into consideration the social and environmental concerns of all interested parties. These concerns will be integrated into the initial stages of the project design. This approach aims at maximising the project’s integration into the environment and minimising its negative impacts to increase the environmental and social acceptability of the Project.
1.13 Capital and Operating Costs
1.13.1 CAPITAL COST ESTIMATE (CAPEX)
The Project scope covered in this Report is based on the construction of a greenfield mining and processing facility with an average mill feed capacity of 360,000 tonnes per year of mineralized material and producing an average of 70,000 oz per year of gold.
The capital and operating cost estimates related to the mine, the concentrator, and all required facilities and infrastructure have been developed by DRA/Met-Chem or consolidated from external sources.
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All amounts are in United States dollars (USD), unless otherwise specified.
The Capex consists of direct and indirect capital costs as well as contingency. Provision for sustaining capital is also estimated and included in the financial evaluation as well as the closure and rehabilitation costs.
The Capex includes the material, equipment, labour and freight required for the mine pre- development, processing facilities, tailings storage and management, as well as all infrastructure and services necessary to support the operation.
The Capex prepared for this PEA is based on a Class 4 type estimate as per the Association for the Advancement of Cost Engineering (“AACE”) Recommended Practice 47R-11 with a target accuracy of ±35%. Although some individual elements of the Capex may not achieve the target level of accuracy, the overall estimate falls within the parameters of the intended accuracy.
Table 1.6 presents a summary of the pre-production initial capital and the sustaining capital costs for the Project.
Table 1.6– Initial Capex Summary
Total Costs Area Area Description (‘000 USD) Direct Costs 1000 Mining 8,744 2000 Crushing 5,451 3000 Concentrator 22,181 4000 Tailings Management System 2,440 5000 General Site Infrastructure 7,130 6000 Electric Power Plant 6,615 Sub-total – Direct Costs 52,561 Indirect Costs 9000 Indirect Costs 17,187 9000 Contingency 13,949 Sub-total –Indirect Costs 31,136 TOTAL: 83,697
Numbers may not add due to rounding.
1.13.2 OPERATING COST ESTIMATE (OPEX)
The total estimated operating costs of the Project (including mining, processing, site services and general administration) is presented in the following Table 1.7.
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The sources of information used to develop the operating costs include in-house databases and outside sources. All amounts are in United States dollars (USD), unless otherwise specified.
Table 1.7 – Summary of Operating Costs
Cost per ounce Total Costs Description of gold produced (‘000 USD) ($/oz) Mining 154,570 271 Processing 139,090 244 General & Administration 39,860 70 Government Royalty 44,420 78 TOTAL OPEX 377,940 662 Sustaining Capex 55,910 98 TOTAL AISC 433,850 760
Numbers may not add due to rounding
1.14 Economic Analysis
The project has been evaluated using Discounted Cash Flow (DCF) analysis. Cash inflows were estimated based on annual revenue projections. The economic/financial assessment of the project is based on Q3-2019 price projections and cost estimates in U.S. currency.
The Net Present Value (NPV) of the project was calculated by discounting back cash flow projections throughout the Life-of-Mine (LOM) to the Project’s valuation date using a base case discount rate of 5%. The Internal Rate of Return (IRR) and the payback period were also calculated.
The financial indicators under base case conditions are presented in Table 1.8.
Table 1.8 - Financial Model Indicators
Base Case Financial Results Unit Value
Pre-Tax NPV @ 5 % M USD 146.7 After-Tax NPV @ 5 % M USD 99.8 Pre-Tax IRR % 31.4 After-Tax IRR % 22.6 Pre-Tax Payback Period years 3.5 After-Tax Payback Period years 4.4
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1.15 Interpretation and Conclusions
The economic analysis of the Project has demonstrated the potential viability of the project over its 9 years life of mine expectancy with recommendations to proceed to next level of Feasibility Study. At an average gold price of $1,300/oz, the financial results indicate a pre-tax NPV of 146.7 M USD at discount rates of 5%. The pre-tax IRR is 31.4% with a payback period of 3.5 years. The after-tax NPV is 99.8 M USD at discount rates of 5%. The after-tax IRR is 22.6% and the payback period is 4.4 years.
1.16 Recommendations
1.16.1 MINING AND GEOLOGY
It is recommended to increase the level of knowledge of the mineral resources of the Nabanga Project with the implementation of a new drilling campaign.
More geotechnical test work and analysis will be needed. A geotechnical drilling program will have to be carried out, with the following objectives:
• Final pit slope angle assessment;
• Geotechnical and ground support recommendations for portals excavations;
• Recommendation for stope dimension;
• 2D/3D finite element analysis, and ground support recommendations for the development areas and the production stope.
A hydrogeological investigation program should take place, possibly at the same time than the next geotechnical drilling program (for assessment of the pit slope and stope stability). Pumping / dewatering hole could be designed at this point if required.
Similarly, a hydrology assessment should take place as part of the PFS or DFS of the Nabanga Project. This work is required to evaluate the rain water contribution on the overall drainage.
1.16.2 PROCESS
A number of Trade-Off Studies (TOS) will need to be conducted during the next stage of the Project. The TOS identified during the PEA stage of the Project are the following:
• Grinding using one large mill or two smaller ones;
• Intensive Cyanide Leaching of Flotation Concentrate;
• Carbon-in-Leach versus Carbon-in-Pulp (CIL vs. CIP);
• Post Leach Thickener versus No Thickening;
• Large Bulk Reagent Tanks versus Totes.
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Additional physical characterisation test work will need to be completed. It is understood that further testing on mineralized material from deeper in the deposit and variability testing will need to be completed.
Flotation concentrate characterisation including settling, filtering and physical characterisation for regrind mill requirements will be required.
Additional leach and adsorption testing will be required including variability test work to determine any differences throughout the orebody with respect to recovery and reagent consumption. Testing will also be used to determine the requirement of Oxygen vs. Air as a leach circuit oxidant and also to determine the optimum leach and adsorption circuit flowsheet, CIP vs. CIL.
1.16.3 ENVIRONMENT
It is recommended to perform the following work in connection with environmental activities:
• Extend soil and surface water surveys to select the best location for the tailing ponds, waste rock and overburden piles;
• Carry out a hydrogeological study to collect field data in order to estimate from groundwater flow modeling dewatering rates and impacts;
• Start the environmental studies required to support permitting requirement and to optimize the site layout;
• Identify environmental requirement for site closure and estimate the cost.
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2 INTRODUCTION
2.1 Terms of Reference Scope of Study
The following Technical Report (herein after “the Report”) is a review and compilation of the exploration and metallurgical work performed by SEMAFO on the Nabanga Property (“Property”).
Met-Chem, a division of DRA Americas Inc. (“DRA/Met-Chem”) has provided engineering and integration services for all aspects of the NI 43-101 Technical Report Preliminary Economic Assessment (“PEA”) on the Nabanga Project (“Project”) with the participation of SEMAFO in Sections 4 to 12, and 14 as well as 23 of the Report.
SEMAFO acquired the Property from Orbis Gold Limited (“Orbis”). SEMAFO initiated a takeover of Orbis in December 2014 and Orbis was delisted from the Australian Securities Exchange (“ASX”) on March 16, 2015.
2.2 Qualified Persons
At the request of SEMAFO, DRA/Met-Chem has been mandated to prepare a NI 43-101 Report for the Project with the participation of specialized consultants. Table 2.1 provides a detailed list of qualified persons as defined in Section 1.5 of NI 43-101 and their respective sections of responsibility.
Table 2.1 – Qualified Persons and their Respective Sections of Responsibilities
Section Title of Section Qualified Persons
1 Summary Patrick Pérez and related QPs 2 Introduction Patrick Pérez 3 Reliance on Other Experts Patrick Pérez 4 Property Description and Location Richard Roy Accessibility, Climate, Local Resources, 5 Richard Roy Infrastructure and Physiography 6 History Richard Roy 7 Geological Setting and Mineralization Richard Roy 8 Deposit Types Richard Roy 9 Exploration Richard Roy 10 Drilling Richard Roy 11 Sample Preparation, Analysis and Security 11.1 Samples from Orbis – 2010-2013 Programs Richard Roy 11.2 Samples from SEMAFO – 2016-2018 Programs François Thibert 12 Data Verification François Thibert
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Section Title of Section Qualified Persons
13 Mineral Processing and Metallurgical Testing Ewald Pengel 14 Mineral Resource Estimates François Thibert 15 Mineral Reserve Estimates N/A 16 Mining Methods Patrick Pérez 17 Recovery Methods Ewald Pengel 18 Project Infrastructure Patrick Pérez 19 Market Studies and Contracts Patrick Pérez Environmental Studies, Permitting and Social or 20 Patrick Pérez Community Impact 21 Capital and Operating Costs Patrick Pérez and related QPs 22 Economic Analysis Patrick Pérez 23 Adjacent Properties Richard Roy 24 Other Relevant Data and Information Patrick Pérez 25 Interpretation and Conclusions Patrick Pérez and related QPs 26 Recommendations Patrick Pérez and related QPs 27 References Patrick Pérez and related QPs
2.3 Effective Date and Declaration
This Report is considered effective as of September 30, 2019 and is in support of SEMAFO's press release, dated September 30, 2019, entitled “SEMAFO: Positive PEA Results for Nabanga After-tax Net Present Value of $100 Million.”
This Report provides an independent Technical Report for the estimate to complete for the Nebanga Project, in conformance with the standards required by NI 43-101 and Form 43-101F1.
It should be understood that the Mineral Reserves presented in this Report are estimates of the size and grade of the deposits based on a number of drill holes and samplings and on assumptions and parameters currently available. The level of confidence in the estimates depends upon a number of uncertainties which include, but are not limited to, future changes in product prices and/or production costs, differences in size and grade and recovery rates from those expected, and changes in Project parameters.
2.4 Site Visit
This Section provides details of the personal inspection on the Property.
Richard Roy, P. Geo., visited the Nabanga site several times, the last visit being between the 24th of May 2018 and the 27th of May 2018. During his visits, he has completed a personal inspection of
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several sectors of the Property, as well as a few outcrops and several old drill sites. Selected drill cores were examined, the database, core logging and sampling activities, QA/QC procedures and geological interpretation were reviewed.
2.5 Units and Currency
In this Report, all currency amounts are in US Dollars (“USD” or “$”) unless otherwise stated. Quantities are generally stated in Système international d’unités (“SI”) metric units, the standard Canadian and international practices, including metric tonne (“tonne”, “t”) for weight, and kilometre (“km”) or metre (“m”) for distances. Abbreviations used in this Report are listed in Section 28.
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3 RELIANCE ON OTHER EXPERTS
The Qualified Persons (QPs) prepared this Report using reports and documents as noted in Section 27. The Authors wish to make clear that they are qualified persons only in respect to the areas in this Report identified in their “Certificates of Qualified Persons”, submitted with this Report to the Canadian Securities Administrators.
A draft copy of the Report has been reviewed for factual errors by SEMAFO. Any changes made as a result of these reviews did not involve any alteration to the conclusions made. Hence, the statement and opinions expressed in this Report 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.
The QPs who prepared this Report relied on information provided by experts who are not QPs. The QPs who authored the sections in this Report believe that it is reasonable to rely on these experts, based on the assumption that the experts have the necessary education, professional designations, and / or relevant experience on matters relevant to the Technical Report.
The QPs used their experience to determine if the information from previous reports was suitable for inclusion in this Technical Report and adjusted information that required amending. This Report includes technical information, which required subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, the QPs do not consider them to be material.
The main documents used to describe the different transactions involving the changes of ownership of the Nabanga permit and the acquisition by SEMAFO of Orbis Gold under Sections 1.1 Property Description and Location, 1.3 History, 4.3 Royalties, Back-in Rights, Payments, Encumbrances, 6.1 Prior Ownership, 9 Exploration are listed below:
• Arrêté No. 2008-08-059/MCE/SG/DGMGC, April 1, 2008.
• "Demande de renouvellement exceptionnel de permis de recherche Dynikongolo et Nabanga, à Birimian Resources SARL, No. 2018-073/MMC/SG/DGCM, 20 février 2018".
• "Bulletin de liquidation à Birimian Resources SARL, No. 2018/DF-RE/PR-17/013, 20 février 2018".
• SEMAFO's Annual Information Forms for the years of 2015 to 2018.
DRA/Met-Chem has relied on reports and opinions provided by SEMAFO for information in Section 20 pertaining to Environment Studies, Permitting and Social or Community Impact. DRA/Met-Chem has reviewed the content of this Section and believes that it provides current and reliable information on environmental, permitting and social or community factors related to the Project.
DRA/Met-Chem is relying on the previous NI 43-101 reports and its referenced documents in relation to all pertinent aspects of the Property. The Reader is referred to these data sources, which are outlined in the “References”, Section 27 of this Report, for further details.
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4 PROPERTY DESCRIPTION AND LOCATION
4.1 Property Location
The Project lies within the Nabanga exploration permit ("Permis de recherche") located in the south- east of Burkina Faso, approximately 250 km to the southeast of Ouagadougou (Figure 4.1).
Figure 4.1 – Nabanga Location Map
The permit is centered on latitude 11° 18' 03” North and longitude 0° 29' 58” East in the geographic coordinate system, or on UTM (Universal Transverse Mercator) coordinates 1 250 000 mN and 225 000 mE, in Reference System WGS 84 Zone 31N.
The Project is located in the Soudougui Department of the Koulpélogo Province in the East Central Administrative Region of Burkina Faso, close to the limit with the East Region.
The southeastern sector of the Project is approximately 5 km to the NW extremity of the Kompienga Lake formed by the hydro-electrical dam built some 25 km further to the SE.
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4.2 Property Description
SEMAFO formerly held five (5) contiguous exploration permits collectively known as the Yactibo Permit Group: the Yacti, Ouargaye, Kamsongo and Napadé permits, the latter two being adjacent to the north and the west of the Nabanga Permit.
4.2.1 NABANGA RESEARCH PERMIT (EXPLORATION PERMIT)
The Nabanga Research Permit ("Permis de recherche") is mainly regulated by Law No. 031-2003/AN of May 8, 2003 of the 2003 Mining Code (Section 1, Art. 11-16).
In Burkina Faso, an exploration permit is granted by Decree of the Minister of Mines. An exploration permit may be granted for a maximum area of 250 km2. The exploration permit is valid for an initial period of three years and may be renewed for two consecutive periods of three years, provided the titleholder complies with the rights and obligations set by the Mining Legislation. On the second renewal, at least 25 per cent of the original area must be relinquished.
The Mining Code gives the exploration permit holder the exclusive right to explore for the minerals requested, on the surface and in the subsurface, within the boundaries of the permit.
The Nabanga Permit was originally granted to Mr. Sayouba Sawadogo on April 1, 2008 and is now registered under the name of Birimian Resources Sarl, which is 100% owned by SEMAFO. The latest renewal by SEMAFO brings the expiry date of the Nabanga Permit to April 1, 2020.
The Nabanga Permit has an irregular shape covering 178.50 km2, the outline of which is detailed in Table 4.1 and depicted in Figure 4.2.
Table 4.1 – Nabanga Permit Boundaries (UTM, WGS84, Zone 31N)
Apex Easting (mE) Northing (mN)
A 220 795.6 1 255 312.6 B 234 795.7 1 255 312.6 C 234 795.7 1 247 989.6 D 229 925.6 1 247 989.6 E 229 925.6 1 242 726.6 F 227 125.6 1 242 726.6 G 227 125.6 1 238 312.6 G 220 795.6 1 238 312.6
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Figure 4.2 – Boundaries of the Original Yactibo Group Permit
Source: SEMAFO
Figure 4.3 – Nabanga Permit Location
Source: SEMAFO
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The application for a permit must be filed along with a program of exploration works that the applicant contemplates to carry out during the first year of validity of the permit, along with the related budget of such program. The titleholder of a permit must start exploration works within the covered perimeter no later than 6 months from the date of validity of the permit and must continue work diligently.
4.2.2 INDUSTRIAL OPERATING PERMIT (MINING PERMIT)
The exploration permit also gives the holder the exclusive right, at any time, to convert the exploration permit into an Industrial Operating Permit (Permis d'exploitation industrielle).
The mining permit (Section 2, Art. 17-23 of the 2003 Mining Code) grants its holder the exclusive right to exploit mineral deposits within the covered perimeter. Mining permits are valid for an initial period of twenty years and are renewable for five-year periods, until the reserves have been depleted.
The application for a mining permit must be filed along with a feasibility study and a plan for the development and exploitation of the deposits which shall include, inter alia, an environmental and social impact study and a mitigation and rehabilitation plan.
4.2.3 MINING AGREEMENT
The Mining Code provides that exploration and exploitation permits shall be accompanied by a mining Agreement (Section 4, Art. 30-39, 2003 Code; Convention minière) that the State will conclude with the titleholder.
The Mining Agreement supplements the provisions of the Mining Code and specifies the rights and obligations of the parties. The Mining Agreement guarantees stabilization of financial and customs regulations and rates during the period of the exploitation to reflect the rates in place at the date of signing. The Agreement is valid for a maximum of 25 years and is renewable for successive periods of 10 years.
4.3 Royalties, Back-in Rights, Payments, Encumbrances
The mining permits are subject to a 10% carried ownership interest to the benefit of the Government of Burkina Faso. In addition, once a Mining Convention is signed and an exploitation permit is awarded by the government, a graduated royalty applies on the prevailing gold price. Below $1,000 /oz, the royalty is 3% on the revenue. At a gold price between $1,000 and $1,300 /oz, the royalty is 4%. Above $1,300 /oz, a 5% royalty needs to be paid. Moreover, a 1% development tax must be applied on the total revenue.
The Nabanga Permit is subject to a Net Profit Royalty (NPR) of 1% assigned to the original vendor (Mr. Sayouba Sawadogo).
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4.4 Environmental Liabilities
All phases of development on the Property will be subject to some environmental regulations related to various concerns, such as maintenance of air and water quality standards, land reclamation, waste disposal.
DRA/Met-Chem is not aware of any environmental liabilities to which the Property may be subjected. However, DRA/Met-Chem has not investigated any environmental liabilities that may arise from previous work.
4.5 Required Permits
SEMAFO completed exploration work and drilling between 2016 and 2018 on the Nabanga Permit. Future exploration or drilling can probably be completed under the existing permits or it can be reasonably be expected that new permits will be issued by the Authorities.
The Project will require a number of approvals, permits and authorizations throughout all stages of development and an Industrial Operating Permit will be required prior to initiation of mining.
4.6 Significant Factors or Risks
DRA/Met-Chem is not aware of any significant factors, risks or encumbrances that may affect access, title, or the right or ability to perform work on the Property. DRA/Met-Chem has not verified the validity of title or rights on the Property. DRA/Met-Chem relied on information provided by SEMAFO on these matters (Section 3 – Reliance on Other Experts).
4.7 Conclusions
Although a new Mining Code has been formalized into law in 2015, importantly, it respects clauses in existing Mining Agreement. Indeed, the new Mining Code (Article 212) specifies that the mining titles and Mining Agreements valid at the time the 2015 Code became effective remain valid for the duration and substances for which they were granted.
The information in this section is only a summary description of some of the rights and obligations pertaining to the mining exploration titles in Burkina Faso. The reader seeking full, up to date and official information on titles or rights and obligations of the permit holders should refer to the Mining Code of Burkina Faso available at the following link:
http://www.eisourcebook.org/cms/February%202016/Burkina%20Faso%20Mining%20Code%20(in %20French%20).pdf
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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY
5.1 Accessibility
The Property is accessible from Ouagadougou by travelling 220 km to the east via all-weather, paved Highway RN04 (Route Nationale) to Fada-N’Gourma. Proceeding approximately 80 km southward along paved RN18 will lead to about 15 km east of the Property. A gravel road that crosses the Kompienga River provides the main access to the Property and different sectors of the Property can be reached by various roads, drill tracks and trails.
Access during the rainy season may be problematic, particularly where the road crosses the river. However, alternate routes allow access from the west for the greater part of the year. Access to the different prospect areas is excellent during the dry season.
5.2 Climate
Three (3) principal climate zones can be distinguished in Burkina Faso:
• Sahelian zone in the north: semi-arid steppe, characterized by three to five months of rainfall;
• Sahelian-Sudanian zone, across the center of Burkina Faso;
• Sudanian zone in the south: increasingly tropical wet-dry type climate, with a greater variability of temperature and greater total rainfall than in the north.
The Property is located in the southern portion of the Sahelian-Sudanian zone. The region is under semi-arid climatic conditions, characterized by a rainy season from May to September and a dry season between October and April (Table 5.1). During the dry season, the harmattan, a dry wind blowing from the northeast or east off the Sahara desert, carries large amounts of dust over hundreds of kilometers.
A hot, dry period extends from February to April in the Project area. April is the warmest month of the year, with an average temperature of 32.1°C, and January has the lowest average temperature at 25.0°C (Table 5.1).
The prevailing winds in the Fada-N'Gourma region blow from the N-NE and S-SW (Figure 5.1).
Table 5.1 – Nabanga Region – Summary of Monthly Weather Averages
May
July
April
June
March
August
January October
ANNUAL
February
December
November September Temperature (°C) Average 1) 25.0 27.8 30.8 32.1 30.9 28.6 26.8 25.8 26.2 28.1 26.9 25.2 27.8 Minimum1) 16.6 19.5 23.1 25.4 25.0 23.2 22.1 21.4 21.2 21.2 18.1 16.7 Maximum1) 33.5 36.2 38.6 38.9 36.8 34.0 31.5 30.3 31.3 35.0 35.7 33.8
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May
July
April
June
March
August
January October
ANNUAL
February
December
November September Average Precipitation 8221) Rainfall (mm) 2) 1 2 11 41 94 122 179 238 166 52 5 3 914 Number of Days 3) 0.5 0.2 0.9 2.5 6.1 9.2 12.9 15.0 11.9 2.7 0.3 0.3 62.5 Average Wind Speed 9.4 9.0 8.3 9.4 9.4 7.9 7.6 7.2 5.8 6.5 6.8 7.9 7.9 (km/h) 3) Sources: 1) Fada-N'Gourma Weather Station; Approximately 90 km to the N; https://en.climate-data.org/africa/burkina-faso/east/fada-n-gourma-766890/#climate-graph 2) Youga Weather Station; Approximately 100 km to the W-NW; https://en.climate-data.org/africa/burkina-faso/central-east/youga-490707/#climate-graph; 3) Diapaga Weather Station; Approximately 180 km to the E-NE; Years on Record: 112; http://www.weatherbase.com/weather/weather.php3?s=600267&cityname=Diapaga%2C+Est%2C+Burkina+Faso&units= https://en.climate-data.org/africa/burkina-faso/central-east/youga-490707/ – climate-graph
Figure 5.1 – Wind Rose for Fada N'Gourma
LEGEND Wind Speed (km/h) 0 >1 >5 >12 >19 >28 >38 >50 >61
Hours per Year of Wind Blowing from 1 59 58 14 19 1 0 0 0 the Indicated Direction.
(Source: Meteoblue; https://www.meteoblue.com/en/weather/historyclimate/climatemodelled/fada-n%27gourma_burkina-faso_2360886)
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A year-round mining operation under the weather conditions at the Nabanga site is possible, as attested by the currently active mines in the region..
5.3 Local Resources and Infrastructure
Fada-N’Gourma is the nearest town where basic services and supplies are available. With a population of 61,000 (2017 census, Wikipedia or 33,910, World Population Review), it is serviced by a hospital, hotels, and a 991 m long dirt-surface landing strip.
Specialized services and equipment have to be sourced in Ouagadougou. Mining manpower would likely be from local and expatriate personnel.
Currently power is envisioned to be provided by on-site generation, but may be either from the Kompienga dam power station or from the regional electricity grid. The Project is 10 km away from a hydro-power line.
Kompienga Lake is in close proximity to the southwestern sector of the Property and the Kompienga River crosses the easternmost part of the Nabanga Permit. An east-flowing tributary crosses the center of the northern tier of the Permit. These bodies may represent a source of water for the mining operation.
Development of the Project will require upgrade or construction of infrastructure, such as: road access, water and power supply, waste dumps and tailings storage areas, leach pad areas, as well as plant and camp facilities.
5.4 Physiography The Property area is at an elevation of about 200 m above seal level and is relatively flat.
The mineralized structures form minor, elevated ridges which may rise some 10 m to 20 m above the surrounding plains.
5.5 Vegetation
The majority of the Burkina Faso landmass, including the region of the Project, lies on a savannah plateau 200 m to 300 m above sea level. Plant life in the Project area is mainly restricted to tall grass, shrubs and isolated trees, notably of the acacia species. A tendency towards deforestation by human causes is observed.
Scattered plots of cultivated land are present on the Project around small settlements of a few houses.
5.6 Surface Rights
Considering the size of the Property and of the adjacent permits owned by SEMAFO, it seems clear that the surface area is sufficient to accommodate the needs of a mining operation.
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6 HISTORY
6.1 Prior Ownership and Ownership Changes
The Nabanga Permit was originally granted to Mr. Sayouba Sawadogo on April 1, 2008 by decree No. 2008-08-059/MCE/SG/DGMGC. The permit covered a surface area of 238 km2 that was reduced by 25% to its current 178.50 km2 upon the second renewal. The permit was subsequently acquired under an option agreement by Birimian Resources Sarl, a company 100% owned by SEMAFO under the acquisition of Orbis Gold Limited (Orbis) in 2015. The permit is currently registered under the name of Birimian Resources Sarl.
Birimian Resources was a wholly-owned subsidiary of Mt Isa Metals Limited (Mt Isa), a mining company headquartered in Brisbane, Australia, that changed its name for Orbis on August 9, 2012. Mt Isa acquired a series of mining properties in Burkina Faso and started exploration in 2010.
Five contiguous exploration permits forming the Yactibo Permit Group were acquired in 2007 and 2008: the Ouargaye, Nabanga, Kamsongo, Yacti and Nabanga permits. The Napadé permit was added in 2011 and filled a gap between the Nabanga and Kamsongo permits.
6.2 Previous Mineral Exploration Work
No exploration is known to have occurred on the Yactibo permits prior to their acquisition by Mr. Sawadogo in 2008.
Orbis initiated exploration on the Property in 2010 with reconnaissance and field mapping, which identified the trends of several artisanal mining sites. Grab samples of quartz vein material collected from heaps or from shallow artisanal workings at Nabanga averaged 14.48 g/t Au. The presence of gold mineralization in the Nabanga structure was defined over a 3.5 km strike length and a sub- parallel gold-bearing structure was identified.
High-resolution satellite imagery (QuickBird imaging satellite) was also acquired by Orbis and was used primarily for regional reconnaissance mapping and delineating the artisanal mining sites. These data were utilized to identify targets for follow-up exploration.
Several regional soil sampling programs were initiated across the Yactibo Permit Group in 2010. The soil samples were collected at 100 m intervals along 800 m nominal line spacing. Infill sampling followed within Kamsongo and Natanga along distances reduced to 50 m between the samples along lines 200 m apart. The results from the soil samples correlated well with the rock chips samples (Snowden Mining Industry Consultants Pty. Ltd.; "Snowden" Report, 2015).
A high-resolution airborne magnetic-radiometric survey was flown by New Resolution Geophysics (NRG) of South Africa and Australia in 2011. The Nabanga structure was interpreted from the magnetic data to extend over approximately 9 km along the strike. Details and illustrations on the geochemical and geophysical surveys are presented in Snowden's 2015 technical report.
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On January 20, 2011, Orbis announced that they had confirmed the presence of a gold deposit on the Property by widely-spaced RC drill holes. In September 2011, Orbis issued a press release about the discovery of the 500 m long Central and 600 m long North Zones at Nabanga.
Additional drilling in 2011-2012 provided the data for the initial resource estimate prepared by Snowden in 2012. The resource focused on a 2.3 km length of the known Nabanga structure.
On September 25, 2013, Orbis announced that deeper drilling had shown that the Nabanga mineralization extends below 200 m from surface, although follow-up drilling suggested a significant weakening of the structure below 200 m.
Soil sampling and Induced Polarization (IP) surveys conducted in 2013 identified linear anomalies that were interpreted as possible strike extensions or potential new "Nabanga-style" mineralized bodies. The samples from trenches excavated in 2013 across many of the anomalies indicated a potential strike extension to the SW of the known main Nabanga mineralized structure.
In 2014, Pells Sullivan Meynink, Australia, was contracted by Orbis to provide conceptual geotechnical risk assessment study pertaining to a potential future underground or open pit mining operation at the Nabanga project (Orbis Gold Ltd; Nabanga Project, Scoping Study; Preliminary Geotechnical Assessment; Rep. PSM1814-003R; January 2014).
Several phases of historical metallurgical test work were completed for the Nabanga Project. These are documented in reports that include:
• Pathfinder Exploration Pty Ltd – Petrographic and Mineragraphic Descriptions – 2011;
• Pathfinder Exploration Pty Ltd – SEM Analyses of Samples NARC040; 66 m to 67 m and 69 m to 70 m, for Mt Isa Metals Ltd. – 6 June 2012;
• JK Tech – SMC Test Report; Mt Isa Metals Ltd. (Orbis Gold former name) – July 2012;
• Lycopodium – Nabanga Project Metallurgical Testwork Review – October 2012;
• Knight Piesold Consulting – Memorandum to Mt Isa Metals Ltd. (Orbis Gold former name); re: Preliminary Waste Rock Geochemical Characterisation – 22 October 2012;
• ALS Metallurgy – Metallurgical testwork conducted upon samples from Nabanga Gold Project for Orbis Gold – November 2013 (Chemical analyses, Gravity separation; Flotation; Leaching).
Additional information and illustrations on previous exploration are provided in Snowden's 2015 technical report.
6.3 Previous Mineral Resources and Production
6.3.1 HISTORICAL RESOURCES – SNOWDEN 2012
A previous JORC (2004) compliant resources estimate for the gold mineralization of Nabanga was performed by Snowden for Orbis in August 2012 and is documented in a technical report (Snowden;
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Orbis Gold Ltd; Nabanga Mineral Resource Estimate; Project No. 03771; September 2012) and an internal report by Orbis entitled: "Nabanga Gold Project, Nabanga Desktop Study – For Internal Use Only; 9/01/2015; Report No: INT – RPT 047, Rev 2g). The results from this historical estimate are presented in Table 6.1.
The holes drilled until August 2, 2012 were used in the resources estimate. Out of 390 RC and core holes drilled on the Property, 385 of which were RC holes, and a database of 17,895 assay results, Snowden used all 239 RC holes and 14 diamond drill holes directly associated with the zone of interest in their resources estimate.
The resource was estimated applying a cut-off grade of 0.5 g/t Au, a minimum mining width of 1.5 m and a top cut at 70 g/t Au. Grade interpolation was performed using the Ordinary Kriging method.
Table 6.1 – Nabanga Deposit – Historical Mineral Resources (at 0.5 g/t Au Cut-Off) – August 2012
Resources Tonnes Grade Contained Gold Material Category (Mt) (g/t Au) (oz x 1,000) *) Oxide 0.2 5.14 30.9 Inferred Fresh 3.0 6.53 628.5 Total 3.2 6.45 659.4
*) Small discrepancies may occur due to rounding.
6.3.2 HISTORICAL RESOURCES – SNOWDEN 2015
The resources estimated in 2012 were revised by Snowden in 2015 based on the same drill data they used in 2012 (as at August 2, 2012) but applying modified parameters. The CIM Definition Standards and the NI 43-101 Standards of Disclosure were used for the revised resources.
The 2015 historical resources are the subject of a published NI 43-101 Technical Report by Snowden: "SEMAFO Inc., Yactibo Permit, Nabanga Gold Deposit; Project No. AU4582, NI 43-101 Technical Report, June 2015; Final".
The revised resources of 2015 are presented in Table 6.2.
Table 6.2 – Nabanga Deposit – Historical Mineral Resources (at 5.0 g/t Au Cut-Off) – June 2015
Resources Tonnes Grade Contained Gold Material Category (Mt) (g/t Au) (oz x 1,000) *) Oxide 0.08 8.9 20.0 Inferred Fresh 1.76 10.1 570.0 Total 1.84 10.0 590.0
*) Small discrepancies may occur due to rounding.
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6.3.3 HISTORICAL RESOURCES – CONCLUSIONS
DRA/Met-Chem and the issuer are neither treating the 2012 nor the 2015 historical estimates as current mineral resources or mineral reserves. The historical resources are only mentioned in this Report for the sake of completeness and they are superseded by the more recent and current estimate presented in this Report. The present resource includes the results from the later holes by Orbis disregarded in the 2012 estimate and from additional drilling by SEMAFO in 2016-2018. Although the historical resources are based on the CIM definitions, the QP has not done sufficient work to classify the historical estimates as current estimates, and an updated estimate is used in this Report.
No reserves were estimated by Orbis for the Nabanga mineralization.
The Reader is cautioned that Mineral Resources, that are not Mineral Reserves, do not have demonstrated economic viability.
6.4 Previous Mineral Production
The Property has not been the subject of prior commercial production from gold mineralization. However, gold has been extracted for an unknown period of time by the local community from shallow artisanal workings ("Orpaillage"). Gold is still mined manually on these sites, with only minor use of semi-mechanized machinery such as small crushers. Free gold is recovered by gravity (panning) and the tails are often sold for further processing of the fine gold. Alternatively, the unprocessed extracted gold-bearing material may be sold for off-site processing.
Artisanal mining sites within the Nabanga permit occur along 3.6 km of the NE-SW trending Nabanga mineralization. The vertical extent of the workings is unknown; however, they are thought to reach a maximum depth of approximately 40 m below surface although the majority of the workings are probably less than 10 m deep.
It appears that the miners have predominantly targeted the higher-grade vein within the core of the Nabanga mineralization, possibly leaving out the lower-grade material in the hanging-wall and footwall selvages. Although the tonnage and grade of the mineralization extracted from Nabanga are unknown they are not material to the resources estimate by virtue of the relatively limited tonnage removed.
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7 GEOLOGICAL SETTING AND MINERALIZATION
7.1 Regional Geology
Burkina Faso is located in the Leo-Man Shield (Rise) that forms the southern exposure of the West African Craton (WAC). The Leo-Man Shield comprises an Archean nucleus, the Kénéma–Man domain (3.60–2.70 Ga) centered on Liberia, and the Paleoproterozoic Baoulé-Mossi domain (2.30 – 2.00 Ga) to the east (Figure 7.1).
The geology of Burkina Faso is dominated by the Baoulé-Mossi Domain consisting of granite-gneiss terrains, “Birimian” greenstone belts and basins (Junner, 1940; Bessoles, 1977) and the fluvio-deltaic basins of the Tarkwaian Sequence (e.g. Baratoux et al., 2011; Gasquet et al., 2003). Sediments of Neoproterozoic age overlie the Paleoproterozoic basement in the SE of Burkina Faso (Volta sedimentary basin) and skirt the western edge of the country (Taoudeni basin).
The Paleoproterozoic Birimian granite-greenstone province was established and structured during the Eburnean Orogeny (2,195 Ma to 2,067 Ma; Pohl, 1988).
The N-NE trending, elongated greenstone belts comprise mafic tholeiitic volcanic rocks together with intermediate and acid calc-alkaline effusive suites that are typical of volcanic arc environments, interbedded with volcano-sedimentary units. They are bound by older granite-gneiss terrains and have been intruded by syn- to late-tectonic granite bodies.
Multiple episodes of intrusion have been distinguished in the granitoid domain (2,153 Ma to 2,097±7 Ma), such as the plutons belonging to the tonalite, trondhjemite, granodiorite (TTG) suite and granitic intrusions (Gasquet et al., 2003). The syn-orogenic foliated granitoid batholiths were emplaced into the basin centres and the late orogenic, unfoliated intrusions occur within the Upper Series volcanic belts (Leube et al., 1990).
The Birimian sedimentary basins are elongated or broad units dominated by greywacke and shales, and they occasionally contain lavas and chemical sedimentary rocks.
The rocks of the Birimian Supergroup have been subdivided into two main series regarded as being time equivalents (Appiah et al., 1991):
• Lower Series; chiefly sediments: argillite, phyllite, schist and meta-greywacke with subordinate volcanic rocks, metamorphosed to greenschist facies (Appiah et al., 1991; Dzigbodi-Adjimah, 1993);
• Upper Series; predominantly volcanic rocks: andesitic tuff, tholeiitic basalt, with associated basic intrusive rocks and interbedded graphitic phyllite.
Younger granitoids and lavas (2.10 and 2.07 Ga) were mostly emplaced in the western portion of the Baoulé-Mossi domain and along the contact zone with the Archean craton (Egal et al., 2002; Hirdes and Davis, 2002; Liégeois et al., 1991).
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The Tarkwaian basins are generally bounded by regional-scale shear zones, tectonised sequences made up of conglomerate, quartzite and shale that are discordant on Birimian formations (Davis et al., 1994; Perrouty et al., 2012). The Tarkwa formation (2,081 ±25 Ma to 1,968 ±49 Ma; Hirdes et al.; 1987) was defined in Ghana where it covers vast areas and hosts large gold deposits. Although the composition of the conglomeratic facies varies according to the rocks forming the substratum, all were deposited after the first tectono-metamorphic phase (D1). No Tarkwaian rocks have been found within the main Birimian basins of the Lower Series.
Figure 7.1 – Simplified Geological Map of the Leo–Man Shield
Source: SEMAFO
7.2 Structural Evolution of the Eburnean Orogeny
The present-day structural trend of litho-tectonic units and the dominant fabrics in the WAC formed during the Eburnean orogeny (Bonhomme, 1962; Eisenlohr and Hirdes, 1992; Feybesseet et al., 2006; Lemoine et al., 1990; Milési et al., 1989; Milesi et al., 1992; Perrouty et al., 2012). The number of deformation phases characterizing the Eburnean orogeny is still under debate. However, the
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presence of an early phase of contractional deformation and medium- to high-grade metamorphism followed by two phases of transcurrent tectonics and low-grade metamorphic overprint is generally recognized (Block et al., 2015; Caby et al., 2000;Debat et al., 2003; Galipp et al., 2003; Ganne et al., 2012; John et al.,1999; Klemd et al., 2002; Liégeois et al., 1991; Opare-Addo et al.,1993; Pitra et al., 2010; Triboulet and Feybesse, 1998).
The first deformation stage (D1; around 2,100 – 2,090 Ma) is characterized by collision thrusting in a southerly direction of the Lower Proterozoic over the Archean craton (Milesi, 1989). The second transcurrent tectono-metamorphic phase (D2; 2,096 – 2,073 Ma in the north of the Ivory Coast; Milesi, 1989) caused regional folds and N-S to NE-SW strike-slip faults. The subsequent transcurrent D3 deformation phase (2,091 ± 33 Ma to 2,073 ± 7 Ma in southern Mali) is marked by the development of large-scale folds controlled by regional shear zones responsible for the general NE- SW structuring of most of the WAC. The D1-D3 deformations form a continuous and overlapping time sequence.
7.3 Local Geology
The Yactibo Permit Group straddles a major NE-trending shear separating the Youga Belt in the northwest from the Diapaga Belt in the southeast. The Nabanga deposit is located to the southeast of the shear, in the Diapaga Belt, which is the easternmost Belt in Burkina Faso (Figure 7.2).
The Diapaga Belt is dominantly comprised of metamorphosed intermediate volcanic rocks, sediments, foliated or migmatitic granite and gneiss, with less common mafic volcanic rocks and mafic-ultramafic intrusive complexes. Conglomeratic sediments are present and are mapped as Tarkwaian equivalents. Early banded or and/or foliated granitoids, which may be ascribed to the TTG Group, are also very common.
7.4 Property Geology
The main feature on the Property is the granodiorite hosting the mineralization. The intrusion is in contact on the north with an assemblage of volcanic rocks (Figure 7.2).
The weathering profile in the Nabanga deposit area is relatively shallow and the contact with the fresh rock is generally crossed at a vertical depth of 10 m to 21 m.
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Figure 7.2 – Nabanga Property Geology
NORTH
Source: SEMAFO
The main lithologies encountered on the Nabanga permit consist of granodiorite, diorite, amphibolite and the quartz vein.
7.4.1.1 Granodiorite
The granodiorite is medium- to coarse-grained and contains 10-20% anhedral quartz, 40-60% plagioclase, 15% K-feldspar and some amphibole, biotite and magnetite (Rugless, 2011). Weak chlorite-epidote-sericite alteration is observable. The granodiorite often exhibits weak to very weak foliation that increases as the quartz-gold mineralization is approached.
The intrusive includes rafts or xenoliths of chlorite-sericite-biotite schist preserving relict phyric plagioclase that are interpreted to represent a basalt or a basaltic andesite metamorphosed to low- grade greenschist facies. Trace pyrite occurs within the mafic schist that appears to be largely unmineralized.
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7.4.1.2 Amphibolite
In the immediate proximity of the Nabanga mineralization, the granodiorite is interpreted to intrude a sequence of amphibolite (porphyritic basalt to andesite precursor; Rugless, 2011) that is represented by multiple, disconnected lenses within the granodiorite. The geophysical data suggest the granodiorite is semi-concordant with the meta-volcanic and/or meta-sedimentary layering. The amphibolite is made up of chlorite and actinolite, with minor relict feldspar phenocrysts and rare sulphides, and exhibits a weak to moderate pervasive foliation.
7.4.1.3 Diorite
A fine- to medium-grained diorite identified by drilling is believed to represent a different phase from the granodiorite, owing to its very low magnetic susceptibility, but does not appear to have any impact on the mineralization.
7.4.1.4 Quartz Vein
A massive, white quartz vein hosts the majority of the gold mineralization.
7.5 Mineralization
The gold-bearing structure at Nabanga is predominantly hosted within a magnetite-rich granodiorite intrusive that is part of the Birimian Diapaga greenstone belt. The mineralization is associated with a sheared quartz vein bracketed by a distinctive alteration halo and development of minor (<1%) pyrite. The Nabanga mineralization is manifest on surface by a NE-trending zone of shallow artisanal gold mining site that extends over a 3.5 km strike length.
The main Nabanga gold mineralization is hosted in a single structure extending over 3.5 km, but traceable over several kilometers along the strike to the southwest, beyond the limits of the surface gold workings, based on data from high resolution magnetic surveys and exploration work.
The main deposit is of high-grade gold type, which is evidenced by a moderate reduction of the contained gold ounces as the cut-off grade is raised (Historical Resources, 6.3.1and 6.3.2).
The overall strike of the Nabanga mineralized structure is NE-SW and the dip varies between 55° and 70° toward the northwest. It has an average horizontal thickness of 3.2 m. A large proportion of the high gold values are contained within the quartz vein. The mineralization is arranged as a series of shoots plunging at 45º to the NE. The mineralization is often lower-grade and/or narrower where the main zone is hosted in the amphibolite.
The Nabanga deposit is divided into the North, Central and Southern Extension Zones. A sub- parallel, 800 m long structure referred to as the Nabanga North Lode is present to the NW.
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Scanning Electron Microscope (SEM) analyses of specimens collected from RC hole NARC040 show that the gold occurs as fine (>10 µm) gold ± silver telluride inclusions within pyrite grains
(Rugless, 2012). Cavalerite (gold telluride, AuTe2) inclusions were observed under the SEM within pyrite grains.
Additional information and illustrations such as photomicrographs of thin sections under the petrographic and electron microscopes are provided in Snowden's technical report (2015).
7.6 Alteration
The granodiorite host has been variously altered by the mineralizing event, as evidenced by the presence of a sericite-biotite-hematite-chlorite assemblage. Sulphide minerals, mostly pyrite and chalcopyrite, are relatively uncommon (Pathfinder Exploration Pty. Ltd.; Rugless, 2011).
7.7 Structure
The Nabanga mineralized structure is crosscut by three (3) faults highlighted in the geophysical surveys and confirmed by drilling. The faults are inferred to be sub-vertical and steeply dipping (80º) to the NW. The mineralized structure tends to rotate east and to become shallower as it approaches the faults, which would be indicative of an oblique fault movement.
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8 DEPOSIT TYPES
The Yactibo Permit Group is located within the world-class Birimian Gold Province that hosts most of the major gold deposits in West Africa, notably in Ghana, Mali, Burkina Faso, Ivory Coast, and Senegal. For example, the Obuasi Mine in Ghana started production in 1897 with over 20 million ounces of gold extracted until its suspension in 2014. It is currently under re-development. Other large gold deposits include Sadiola (8 million oz Au) and Morila (5.9 million oz Au) in Mali.
Most Birimian deposits correspond to orogenic quartz lodes with gold or gold associated with sulphides or tourmaline.
A widely used classification of the West African primary gold deposits has been developed by Milesi et al. (1989), based on the type of host rock, the host structure, the geometry of the deposit and its paragenesis:
• Type 1: Gold associated with sulphides in tourmalinised turbidites; pre-orogenic deposits (Loulo in Mali); • Type 2: Gold with disseminated sulphides and associated hydrothermal alteration in tholeitic volcanic rocks; syn-orogenic deposits (Yaouré in Ivory Coast; Syama in Mali); • Type 3: Gold-bearing paleo-placer quartz pebble conglomerates; syn-orogenic deposits (Tarkwa, Tereberie, Iduaprem in Ghana); • Type 4: Mesothermal, late-orogenic lode quartz with gold-bearing arsenopyrite associated with regional structures in metasediments or dykes and sills; tardi-orogenic deposits (Ashanti, Prestea, Konongo in Ghana; Afema in Ivory Coast); • Type 5: Mesothermal, native gold and polymetallic sulphides in quartz veins in shear zones usually exhibiting evidence of hydrothermal alteration; tardi-orogenic deposits (Sabodala in Senegal; Poura in Burkina Faso; Kalana in Mali).
The secondary gold mineralization has been classified as:
• Alluvial and eluvial placers; mined by “orpailleurs” throughout the shield; • Colluvial, transported material; nugget gold in laterite with no evident underlying primary mineralization.
The Nabanga mineralization is considered to belong to the intrusion-hosted, epigenetic, mesothermal class of gold deposits that includes world-class deposits such as Fort Knox and Pogo (Alaska), Vasilkovskoe (Russia), Ahafo and Chirano (Sefwi Belt of Ghana).
Alternatively, the Nabanga deposit belongs to a distinct group that has been generally recognized and accepted during the last decade as a type of gold-only deposits associated with moderately reduced granodiorite-granite stocks and batholiths (Thompson and Newberry, 2000. Lang and Baker, 2001.Hart, 2005).
SEMAFO has applied this model as a basis on which the exploration and development programs were planned.
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9 EXPLORATION
Exploration by SEMAFO on the Nabanga and the adjacent permits included establishment of a line grid, field mapping, outlining the different artisanal mining sites, geochemical and geophysical surveying and an important auger drilling program (Figure 9.1).
In May and June 2017, SEMAFO completed regional reconnaissance mapping.
Systematic auger drilling, logging and whole-rock analysis by X-Ray Fluorescence (XRF) of the recovered material was carried out on the Nabanga, Kamsongo and Napadé permits. Sahara Geoservices, an exploration services company with an office in Ouagadougou, was hired to carry out the auger drilling programs.
In 2018, 8,530 holes were drilled to a nominal depth of 5 m for a cumulative total of 41,923 m, which generated 16,994 samples (Table 9.1). Three gas-powered auger drill rigs were mobilized on April 14, 2018 and started investigating the Nabanga North sector. The holes were drilled at 25 m spacing along lines 800 m apart, oriented along azimuth N135⁰. Between June 26 and July 14, 2018, three rigs were active on the Nabanga Extension block. The holes were drilled along a grid at 400 m by 25 m intervals, with the lines oriented N135°. Auger drilling is a cost-effective method for geochemical sampling. Vertical holes are drilled to the in-situ saprolite horizon and samples taken from both the laterite/saprolite interface and within the saprolite. Sample size is approximately 2-3 kg.
Generally, auger drilling returned anomalous values that confirmed the regional NE geophysical trends in the Yactibo permits.
Table 9.1 – Summary of the Auger Holes drilled Entirely or Partly on the Nabanga Permit in 2018
Period Meterage Holes Samples Sector (2018) (m) April 1,232 6,297 2,461 Nabanga North May 3,187 14,669 6,322 Yactibo South Yactibo South; June 2,798 14,408 5,586 Infill at Nabanga Extension Grid straddling Nabanga, Napadé and July 1,313 6,549 2,625 Kamsongo permits; Infill at Nabanga Extension TOTAL 8,530 41,923 16,994
In July 2018, SEMAFO located and mapped the artisanal mining sites on the Nabanga permit.
An IP/Resistivity and magnetometer survey was completed on the Nabanga Extension grid between July 1 and 29, 2018.
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In August 2018, SEMAFO collected 38 quarter core samples from holes NADD18-0005 and NADD18-0021 drilled in 2018 in the Nabanga permit to be submitted for whole-rock analysis. The purpose of this lithogeochemical work was to characterize the geochemical signature of the hydrothermal alteration at Nabanga, more specifically, the hematitic alteration.
Exploration work outside of the Nabanga permit is not part of the present study, but the results from exploration and development work by SEMAFO on the adjacent permits add to the general understanding of the general geology, structure and the controlling factors on the gold mineralization.
Figure 9.1 – Summary of Exploration and Drilling on the Nabanga Permit (Black and Red: Holes by Orbis; Blue, Yellow and Green: Holes by SEMAFO)
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10 DRILLING
10.1 Drilling by Orbis
10.1.1 METHODOLOGY
A first RC drilling program completed by Orbis (Mt. Isa) in 2010 over the entire 3.5 km length of exposed artisanal workings of Nabanga confirmed the discovery of the Nabanga gold deposit. A press release from Mt. Isa dated January 20, 2011 announced that 20 out of 21 widely spaced holes intersected gold mineralization of greater than 0.5 g/t Au, with an average of 5.66 g/t Au over a core intersection 4.6 m. (Source: Minenportal https://www.minenportal.de/unternehmen_degelistet.php? mid=1373&sid=32288&lang=en).
This was followed by additional drilling in mid-2011 and in 2011-2012, which culminated with an initial mineral resource estimate in 2012.
Follow-up drilling in 2013 suggested a significant weakening of the Nabanga structure below the vertical depth of 200 m. A further five (5) stepout RC holes drilled in 2013 along strike to the northeast intersected insignificant mineralization. However, SEMAFO showed that the holes had been designed to test a misinterpreted plunge of the high-grade shoots, which led Orbis to conclude that the known mineralization was closed off to the northeast.
Drilling at Nabanga by Orbis was performed by a combination of RC and diamond drill holes. Except for a few cored from surface, the diamond drill holes were generally pre-collared using RC drilling to a maximum depth of 200 m below surface, down to approximately 10 m above the interpreted top of mineralization. The RC holes were extended by coring (NQ core tails, 47.6 mm diameter) through the mineralization. These holes are identified as Multi-purpose (MP) holes in the database.
A total of 441 holes were drilled by Orbis between 2010 and 2013, and 390 of them were used in the initial resource estimation of 2012 (Table 10.1). All the holes drilled at Nabanga are listed in Table 10.1, but the RC or RAB holes drilled outside of the Nabanga permit are omitted.
Table 10.1 – Nabanga Permit – 2010-2013 Holes Drilled by Orbis
RC Basis of Minimum Maximum Total Date - From Date - To Type Count Pre- Snowden Length Length Meterage Collar Resources Nov. 24, 2010 Dec. 11, 2010 RC 31 21.0 150.0 2,824.00 √ Jun. 4, 2011 Aug. 3, 2011 RC 60 17.0 114.0 4,392.00 √ Nov. 4, 2011 Jun. 17, 2012 RC 270 18.0 210.0 29,003.00 √ Feb. 20, 2012 May 10, 2012 DDH 28 17.5 443.0 6,821.25 3,481.0 √ Nov. 22, 2012 Feb. 13, 2013 DDH 47 19.0 432.2 12,641.33 8,324.5 Feb. 12, 2013 Mar. 20, 2013 RC 5 139.0 200.0 761.00
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RC Basis of Minimum Maximum Total Date - From Date - To Type Count Pre- Snowden Length Length Meterage Collar Resources Total RC 366 17.0 210.0 36,980.00 Total DDH 75 17.5 443.0 19,462.58 11,723.5 Grand Total 441 58,442.58
Drilling the Nabanga deposit was carried out on NE-SW sections, on a nominal spacing of 40 m along strike and 20 m across strike. Drilling in the sector between 1 249 850 mN and 1 250 300 mN was tightened to approximately 30 m. As all holes were drilled at plunges ranging from 50º to 70º to the SE, they intersected the mineralization at angles of no less than 45º.
The main Nabanga deposit is divided into the North, Central and Southern Extension Zones. Drilling aimed at a sub-parallel structure to the NW, referred to as the Nabanga North Lode, returned 2 m of core length at 8.32 g/t Au as the best value. The majority of the results suggest that the mineralization in the North Lode is narrow, discontinuous and generally of low-grade (>0.3 g/t Au).
10.1.2 DRILL HOLE COLLAR LOCATION
No topographic survey had been conducted on the deposit at the time of the resource estimation and Snowden recommended Orbis undertake a survey to enable a more accurate depiction of surface.
A Digital Terrain Model (DTM) produced as part of the airborne geophysical survey was conducted over the project area in 2011 but was only used for drill planning purposes.
Upon completion of drilling, the collar locations were surveyed using a Trimble GeoExplorer 6000 Differential Global Positioning System (DGPS). The stated accuracy of the instrument is 1.0 cm in the X-Y directions and 1.5 cm in the Z direction.
Snowden used the collar elevations (RL) to produce a topographic surface in the immediate Nabanga area by triangulation for use in the resource estimate. Whilst not optimal, given the relatively flat topography and the early project status at the time, using this surface to constrain the resource was considered as reasonable.
10.1.3 DOWNHOLE DEVIATION SURVEY
The deviation path of all drill holes was surveyed using digital, single and multi-shot cameras. A Campteq and a Reflex Ez-trac multi-shot camera, both with an accuracy of ±0.5° for azimuth measurements and ±0.2° for dip measurements, were used for the downhole surveys. A reading was taken below the casing, at about 6 m and every 30 m thereafter. The 30 m spacing was adjusted, or one reading was added in order to have a measurement at the bottom of the holes.
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Snowden noted that the instruments can be influenced by the magnetic interference from the surrounding rocks. However, Snowden thought that this interference was unlikely to be material to the resource estimate, given the low magnetic susceptibility around the deposit (in the order of 24,600n Tesla). Yet, Snowden recommended assessing the accuracy of the downhole surveys using a method that is not impacted by magnetic interference, such as the gyroscope or optical technique. Snowden checked the data for possible unusual deviations both in plan and section and no issues were identified except for three holes. DRA/Met-Chem did not detect any abnormal deviation in the azimuth or the plunge reported in the Downhole Survey Table of the drill database.
10.1.4 LOGGING CORE AND RC CHIPS
Upon completion of each core hole, the geologist at the drill site took a set of photographs and performed a quick log prior to moving the boxes to Orbis onsite facilities for detailed logging and sampling. The quick logs and the photographs were sent on a daily basis to the Orbis Brisbane office for review. The RC holes were logged at the rig site.
Detailed logging of the drill core and RC chips included recording of lithology, alteration, sulphides percentage, vein composition and structural elements. Prior to 2013 logging was hand-written on a template and manually entered into the database. All logging after 2013 was done with Micromine Geobank Mobile software. The geologists used a library of geology codes to log in order to preserve consistency.
10.1.5 RC SAMPLING BY ORBIS
The drill cuttings from the RC holes were collected into polyethylene bags at the discharge of the cyclone for every meter. The mass of the samples was reduced to 2 kg using riffle splitters. The riffle splitters were cleaned after each sample with a brush.
The 2-kg, one-meter samples of visually identified mineralization were placed into a polyethylene bag with a sample tag and shipped to the laboratory. Outside of the mineralized intervals, the samples for analysis were composited to 4-m lengths. The bags had been pre-labelled with the hole ID and the "From" and "To" meterages. The QC samples were inserted into the sample stream at this stage.
Assays of 0.25 g/t or more on any 4-m composite were flagged for re-analysis of the four (4) original, individual 1-m intervals of the composite and of adjacent composites to bracket the identified mineralized section.
For wet samples, holes were poked in the polypropylene bags containing the 1-m samples and the water was allowed to drain. The samples were then placed onto a black plastic sheet and, once dry they were homogenized by hand and split using the same process as the dry samples. Records of sampling show that RC samples were collected dry, moist and wet respectively 77%, 9% and 13% of the time.
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RC drilling is designed to drill under dry or damp conditions. Even though DRA/Met-Chem realizes that it is possible to obtain good quality RC samples under wet drilling conditions, this is not recommended. Possible loss of fines in the slurry water or fines drained away through the holes pierced in the bags may compromise the sample quality. Adequate processing of wet samples involves the use of micro-pore bags that allow the water to drain while retaining the fine sample material or flocculation/decanting process.
However, Snowden (2005) compared the analytical results from the wet RC samples and the corresponding core samples and did not see any significant difference.
After extracting the 2-kg sub-samples, the entire 1-m sample bags for the initial RC holes were saved for future reference. Subsequently, Orbis only retained 2- to 3-kg splits from the full 1-m samples for storage in an enclosed shed on site. DRA/Met-Chem did not check this information provided by Snowden, and assumes this is still valid.
The RC samples were shipped to the SGS Minerals Services (SGS) laboratory in Ouagadougou for assay.
10.1.6 CORE SAMPLING BY ORBIS
Core sampling was done on a nominal 1-m interval, adjusted to respect the lithological, mineralization or alteration contacts. The core was cut using a diamond blade saw and the right-hand half was systematically placed into a polyethylene bag and used for analytical work. The QC samples were inserted while preparing the core samples.
The core samples were shipped to the SGS laboratory in Ouagadougou for assay.
10.1.7 DRILL SAMPLES RECOVERY
The global recovery of RC chips averaged 80% to 90%, but the values are distributed about two peaks, at 60% and at 100% (Figure 10.1). Recovery was calculated on the basis of the volume of a 5¼ inch (127 mm hammer) diameter hole and an average bulk density of 2.8.
The presence of two populations and the recoveries exceeding 100% suggests that some difficulties were experienced obtaining accurate measurements. It is generally accepted in the industry that a recovery of no less than 80% for a sample interval is required for a sample to be representative.
In Snowden's opinion, the recovery was acceptable for representative samples and subsequent mineral resources estimation. The core recovery was reported by Snowden as excellent.
DRA/Met-Chem did not verify this information provided by Snowden.
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Figure 10.1 – Histogram of Recovery of RC Drilling by Orbis
10.2 Drilling by SEMAFO
10.2.1 2016 DRILLING PROGRAM
In 2015, SEMAFO re-evaluated the results from all previous exploration and drill holes completed by Orbis and concluded that the property warranted additional drilling. SEMAFO designed the first phase of drilling aimed at investigating the SW Extension of the Nabanga mineralized structure, in an attempt to upgrade the potential of the Property.
Major Drilling mobilized a DB-30 multi-purpose rig and started a program of reverse circulation (RC) holes. A total of 15 holes were drilled into the Nabanga SW Extension to a depth of 130 m to 135 m for a total of 1,959 m, between April 25 and May 8, 2016 (Table 10.2). The holes were drilled at an inclination of -50° toward azimuth 135°.
Drilling continued until the month of July on the Kamsongo permit, but these holes outside of the Nabanga permit are not included in this Report.
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The holes were designed to intersect at right angle the strike indicated by the major magnetic lineaments, the known Nabanga structure, the quartz veins observed in the artisanal mining sites ("orpaillage"), as well as the trend of the soil anomalies detected by a local geochemical survey.
The program was successful explaining the soil anomalies but, although widespread anomalous gold values were obtained, none of economic potential was encountered.
10.2.2 2017 DRILLING PROGRAM
10.2.2.1 Drilling
The 2017 drilling program was completed by Forages Technic-Eau Burkina Sarl (FTE), a division of FTE headquartered in Canada. Drilling started with a T3W Reverse Circulation rig on March 1, 2017 and 63 RC holes for a total of 6,861 m were completed on April 4, 2017 (Table 10.2). The length of the holes varied between 30 m and 150 m.
Drilling was aimed primarily at six (6) proximal targets, parallel to and along, the southwest extension of the Nabanga deposit: Nabanga North, North-West, Centre, South-West (parallel and NW of Nabanga main deposit), South Extension (alongstrike), South-East (parallel and SE of Nabanga main deposit).
As for the 2016 program, the holes in 2017 were planned to cross the known NE trends, with added information gained from the trenches excavated in 2013 over several magnetic anomalies, namely the N and S Extensions. The holes were drilled toward azimuth 145° (135° over Nabanga SW) at an inclination of -50°.
Generally, the drilled portions of the structures returned low gold values and suggested lack of continuity.
Auger drilling completed by SEMAFO in 2017 on the Nabanga, Kamsongo and Napadé permits is addressed under Section 9 – Exploration, as this technique was used as a tool to map the lithologies and explore for additional mineralization.
10.2.2.2 Re-logging of RC Chips
In November 2017, SEMAFO initiated a program of re-logging the chips from the RC holes drilled at Nabanga. In March 2018, 10,523 m from 105 RC holes were re-logged, with another 447 m in six (6) holes re-logged in April 2018.
The purpose of re-logging was to gather consistent rocks descriptions in order to assist with the re- interpretation of the geology and structure of the deposit. This effort was part of constructing a new model to help with subsequent development drilling.
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10.2.3 2018 DRILLING PROGRAM
In 2018, SEMAFO updated the database used by Snowden for the resources estimations of 2012 and 2015 that were based on the holes drilled until August 2, 2012. SEMAFO entered into the database the holes drilled by Orbis in 2012-2013 and the holes drilled by SEMAFO in 2016 and 2017. The new descriptions of geology from re-logging the RC chips in 2017-2018 were incorporated into the database.
Building on these data, SEMAFO re-interpreted the geology and structure of the Nabanga deposit and updated the 3D geology model. The resulting interpretation suggested a shallower northerly plunge than previously interpreted of the high-grade shoots, strongly suggesting they could extend along strike and beyond the depth of 200 m.
SEMAFO commissioned Geodrill Limited who mobilized a 900-9 diamond drill rig on March 25, 2018, a second core drill rig (900-10) on May 14, and a third drill (200/3) in June. A program of 25 diamond drill holes, for a total of 7,148.20 m, was completed between March 26 and June 28, 2018 to test the new interpretation by SEMAFO (Table 10.2). The holes were drilled toward azimuth 145° and started at a plunge of -60° and -70°. The length of the holes ranged from 126 m to 456 m. The rigs were equipped to retrieve NQ core size and all core was oriented.
The 2018 program was successful in supporting the new interpretation of the Nabanga gold mineralization being hosted in a series of shallowly-plunging shoots of higher-grade mineralization, in contrast to the previous interpretation by Orbis. SEMAFO's interpretation and drilling results strongly suggested that the deposit is still open at depth and to the north, along the extension of the plunging shoots.
Table 10.2 – Summary of the 2016-2018 Drill Programs by SEMAFO
DIAMOND RC HOLES DRILL HOLES YEAR Total Length Total Length Count Count (m) (m)
2016 15 1,959.0 - - 2017 63 6,861.0 - - 2018 - - 25 7,148.2 TOTAL 78 8,820.0 25 7,148.2 GRAND TOTAL 103 holes for a total of 15,968.20 m
Auger drilling completed by SEMAFO in 2018 on the Nabanga permit is addressed under Section 9- Exploration, as exploration was the prime purposes of drilling these holes.
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Figure 10.2 – Drilling Program 2016-2018: Holes in Grey Auger and in Black RC/DD
10.2.4 DRILL HOLE COLLAR LOCATIONS
All drill hole collars were surveyed using a Trimble Geoexplorer 6000 differential global position system.
10.2.5 DOWNHOLE DEVIATION SURVEY
Holes were all surveyed downhole using a REFLEX GYRO™ electronic surveying tool. Both the azimuth and dip were recorded at intervals of approximately 5 m intervals downhole. The REFELX GYRO™ system is not affected by magnetic interference from either the drill rods or magnetic minerals. The accuracy of the system is reported to be ±0.5° when measuring azimuth and ±0.2° when measuring dip angle.
The first reading was taken below the casing at about 6 m and every 5, 10, 15 or 20 m thereafter in the 2016 program, every 5 m in 2017 and every 25 m or 50 m in the core holes of 2018.
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DRA/Met-Chem checked the results in the Downhole Survey Table of the drill database and did not detect any abnormal deviation in the azimuth or the plunge reported.
10.2.6 LOGGING CORE AND RC CHIPS
RC chips logging is performed by the geologists at the drill using a small quantity of cuttings for each meter drilled. This material is saved in chip boxes for future reference or re-logging. Recovery, lithology, mineralization (quartz, sulphide), alteration, texture and weathering were recorded.
The core logging geologists records core recovery, RQD, lithology, quartz veins (type, percentage), sulphides, supergene alteration (weathering), oxidation, hydrothermal alteration, texture and structural elements (contacts, fractures, faults, shears, foliation schistosity). A table containing the specific gravity results was ignored as the data were considered as unreliable. A table containing the specific gravity results was ignored as the data were considered as unreliable. A total of 2,507 measurements of α and β angles (strike and dip) of planar elements from 2018 oriented core samples are available in the database.
All the data are recorded directly into dedicated software, Geobank Mobile and subsequently entered into an electronic database under the supervision of the project geologist and the exploration manager. The master database is managed by a GIS specialist.
10.2.7 MAGNETIC SUSCEPTIBILITY DETERMINATIONS
SEMAFO started a program of determination of the magnetic susceptibility of the lithologies intersected by core drilling on July 5, 2018. The magnetic susceptibility was measured systematically on 4,497 m of core in 14 holes in July and over 1,403 m from six (6) holes in August. Five (5) readings were systematically taken over 1-m intervals and the final result was calculated as the average of the five (5) measurements.
10.2.8 RC SAMPLING
RC chips sampling is performed by the geologists at the drill site at fixed intervals of 1-m as drilling progresses. The samples dispatched to the laboratory are reduced to 2 kg by multi-stage riffle splitter. The remainder of the sample is saved for future reference.
10.2.9 CORE SAMPLING
Core sampling was done on a nominal 1-m interval, adjusted to respect the lithological, mineralization or alteration contacts. The core was cut using a diamond blade saw and one half placed into a polyethylene bag and used for analytical work. The QC samples were inserted while preparing the core samples.
DRA/Met-Chem noticed that the samples occasionally straddle the main contacts.
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The core samples were shipped to the SGS Minerals Services laboratory (SGS) in Ouagadougou for assay.
10.2.10 DRILL SAMPLE RECOVERY
Core recovery was excellent with 95% of the core drilled at a rate of ≥95% recovery.
10.2.11 CONCLUSIONS
DRA/Met-Chem has not observed any drilling, sampling or recovery factors that could materially impact the accuracy and reliability of the results.
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11 SAMPLE PREPARATION, ANALYSIS AND SECURITY
11.1 Samples from Orbis – 2010-2013 Programs
11.1.1 INTRODUCTION
The drill core and RC samples were submitted to three (3) different laboratories, all at their facilities in Ouagadougou, for preparation and assay:
• BIGS: core samples;
• ALS: samples from holes NARC001 to NARC091;
• SGS: RC samples subsequent to NARC091.
11.1.2 SAMPLE PREPARATION
The samples at SGS were dried at 105° for a minimum of 6 hours, weighted and crushed to 80% passing 2 mm (10 mesh). A 1.5-kg coarse split was extracted with a rotary or riffle splitter and reduced to 85% passing 75µm (200 mesh) in a bowl and puck pulveriser. A 200-g sub-sample was collected by scooping from the 1.5-kg pulp and used for analysis.
At ALS Minerals Services (ALS), the samples were crushed to 70% passing 2 mm and quartered to obtain a 250-g split. The 250-g riffle split was pulverized to better than 85% passing 75μm and quartered to get a 50-g split.
BIGS Global Burkina (BIGS) crushed the samples to <2mm using a Boyd jaw crusher and prepared a sample of approximately 1 kg using a riffle splitter to be processed by an LM2 pulveriser to 90% passing 75µm.
11.1.3 SAMPLE ANALYSIS
All samples were analyzed by Fire Assay (FA) on 50 g aliquots with an Atomic Absorption Spectroscopy (AAS) finish. The lower detection limit was 0.01 ppm for SGS and ALS and 0.001 ppm at BIGS. The samples exceeding the upper detection limit were re-assayed using a 50-g fire assay procedure with a gravimetric finish.
11.1.4 INTERNAL QA/QC PROCEDURES
All preparation equipment at SGS was flushed with barren material prior to commencement of the job and the crushers and pulverisers were cleaned with compressed air between each sample. SGS screened one in 30 samples to ensure a grind of 85% passing 75µm and all the samples were screened if this sample failed the test.
Each batch of 84 samples at SGS included four Certified Reference Materials (CRMs; commonly referred to as "Standards"), two (2) coarse and two (2) pulp duplicates and one (1) coarse and one
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(1) pulp blank. A minimum of 5% check assays are performed on all batches, depending on the number of anomalies present within a given batch.
The ALS internal QA/QC system involved insertion of two (2) CRMs, one (1) coarse and two (2) pulp duplicates and one (1) blank pulp into each analysis batch of 84 samples.
BIGS' internal quality control at the sample preparation stage includes preparation repeats (duplicate samples), cleaning preparation equipment by crushing and pulverizing barren material. BIGS introduced one (1) blank, two (2) CRMs and four repeats (including two (2) preparation repeats) into the batches of 50 samples.
Assay reports from the three (3) laboratories were submitted as digital data files and as signed certificates in PDF format.
11.1.5 LABORATORIES CERTIFICATION AND ACCREDITATION
ALS is certified to ISO 9001:2015 standards and accredited to ISO 17025:2005 at all of their locations. Their internal QA/QC protocol includes verifications at the preparation stage, insertion of blank, standard and various duplicate samples, as well as periodical internal and external instrument calibrations. The laboratories participate in approved proficiency testing and round robin programs and they operate under a Laboratory Information Management Systems (LIMS).
SGS was accredited after Orbis had completed its drill programs. Indeed, on July 15, 2015 SGS announced that their analytical laboratories in Ouagadougou were officially recognized by the South African National Accreditation System (SANAS) for meeting the requirements of the ISO/IEC 17025 standard for specific registered tests for the minerals industry.
The BIGS laboratory was established in 2007 and achieved certification to ISO-17025 standards specific for analytical laboratories and ISO-9001 for quality management system by SANAS in 2009.
SGS, ALS and BIGS are all commercial laboratories well established in the mining industry and are independent of SEMAFO.
11.1.6 QA/QC PROTOCOL – ORBIS
Orbis used standard, blank and duplicate samples to monitor the laboratory performance.
Orbis used 22 different CRMs sourced from Ore Research and Exploration Pty Ltd, (OREAS) in Australia. The CRMs are sourced from the Magdala Lode at the Stawell Gold Mine in Victoria and from the Cracow Gold Mine in Queensland. The grade and matrix from the CRMs matched the style of mineralization at Nabanga.
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A standard was inserted into the primary sample stream at the end of the visually mineralized intersections and at the bottom of the holes. A total of 502 analyses (2.8%) of standards were performed out of a total of 17,968 samples submitted to the laboratories.
Of the 502 analyses, 25 lay outside the control limits of the mean and three standard deviations from the expected grade (Figure 11.1).
Figure 11.1 – Results from the Main Standard used by Orbis
For each drill hole, Orbis inserted a blank sample at the top of the sample sequence and one in the mineralized intersections. No blank samples were placed into the un-mineralized portions of the holes. The blank material was sourced from the Bobu quarry in Burkina Faso. Although this blank is not certified, it provides an indication of contamination when positioned directly after a mineralized intercept and may also pick up sample mis-sequencing.
In addition, DRA/Met-Chem believes a coarse blank is preferable to a pulp to monitor possible contamination by the crusher. Nonetheless, the use of certified blank material is recommended in a future program as part of the QA/QC procedures.
On average, the background grade of the uncertified blank material is estimated at 0.01 g/t Au. Six (6) assays above 0.20 ppm Au may be the result of inherent variability of the blank and one (1) value at 1.82 ppm Au may be due to a sample mix-up. The blank samples analyzed at SGS show the highest variability of the three laboratories. The few samples with values exceeding the expected 0.01 ppm tenor do not point to a significant impact on the reliability of the results that are considered as acceptable for a resources estimation in the inferred category.
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Orbis riffle split 418 field duplicate samples to be submitted to the laboratory with the sample batches. The duplicate samples were prepared from the intervals with the best expected mineralization and were followed by a blank sample. No duplicate samples were requested in the un-mineralized portions of the holes.
Snowden calculated a Pearson correlation coefficient of 0.98 for the duplicate sample pairs, which is considered as excellent for a gold deposit. Overall, the original and duplicate populations compared well, with some outliers at higher grades typical of gold deposits containing visible particles (Figure 11.2).
Figure 11.2 – Scatterplot of Original and Field Duplicate Samples
Additional tests on the reliability of the results from the duplicate samples are provided in Snowden's 2015 Report.
11.1.7 SAMPLES SECURITY, CHAIN OF CUSTODY
Transportation of the samples during the drill programs was overseen by security guards and the samples were securely stored on site at Nabanga. SGS picked up the samples from the Project site twice a week unless otherwise organized. Personnel releasing the samples for shipment to the laboratories assumed responsibility for the sample security.
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11.1.8 SPECIFIC GRAVITY
Bulk density determinations using the Archimedes method were completed on site by Orbis from fresh, non-porous drill core samples. The bulk density of a total of 139 samples was determined, including five (5) duplicated measurements, on the following lithologies:
▪ Granodiorite (84);
▪ Amphibolite (30);
▪ Quartz Vein (20).
The initial 46 measurements were taken using the water displacement method, but they had to be disregarded due to inaccuracies caused by the use of inappropriate apparatus. The density on the subsequent samples was determined by the weight-in-air-weight-in water technique.
In view of the low number of density determinations, Snowden used default bulk densities in the 3D resource model.
11.1.9 CONCLUSIONS
Generally, on the brief review of the Orbis data, DRA/Met-Chem believes that the independent QC samples and the analytical work, primarily completed by SGS, are of sufficient precision and accuracy for use in an estimation of resources at the Inferred level.
11.2 Samples from SEMAFO – 2016-2018 Programs
11.2.1 SAMPLING
The individual diamond drilling samples represent a length of approximately one meter of core that was cut lengthwise using a diamond blade rock saw. One half of the core was kept on site for reference, and its counterpart was sent for preparation and gold assaying to ALS in Ouagadougou.
For RC drilling, the individual samples corresponding to each 1-meter advance of drilling were homogenized and riffle-split to an approximately 2-kilogram subsample that was sent for preparation and assay to ALS in Ouagadougou, or to the Mana Mine laboratory.
Each sample was fire-assayed for gold content on a 50-gram subsample. In addition to the laboratory's own QA/QC program, a QA/QC system was enforced by SEMAFO throughout the sampling program, using duplicate and blank samples, and recognized industry CRMs.
11.2.2 QA/QC PROTOCOL – SEMAFO
11.2.3 INTRODUCTION
DRA/Met-Chem reviewed the evaluation of the results from the QC samples by SEMAFO and generally agrees with their conclusions. The information and conclusions in this section are extracted from the memo by SEMAFO, as well as the selected control charts reproduced in this section.
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SEMAFO shipped 112 batches of samples to ALS in Ouagadougou in 2018. A total of 8,153 analyses were added to the Yactibo drill database, including re-analyses from a few holes drilled by Orbis. In addition, another 248 CRMs, 439 quarter core duplicates and 220 blanks samples were analyzed. Finally, 405 pulp duplicate samples were analyzed at SEMAFO's Mana laboratory to serve as external check by a secondary laboratory.
11.2.4 CERTIFIED REFERENCE MATERIALS (CRMS; "STANDARDS")
During the course of 2018, SEMAFO used ten (10) different CRMs with gold values ranging from 0.609 g/t Au to 5.931 g/t Au (Table 11.1). The grades of the CRMs selected by SEMAFO, as well as their oxide or sulphide matrix, match the samples from the Nabanga deposit.
Table 11.1 – List and Details on the CRMs Used by SEMAFO
Certified Value Standard CRM Count (g/t Au) Deviation OxE113 23 0.609 0.014 OxF125 18 0.806 0.02 OxF142 23 0.805 0.019 SF85 56 0..848 0.018 SG56 20 1.027 0.033 SG66 24 1.086 0.032 SH82 9 1.333 0.027 Si81 8 1.79 0.03 SJ80 49 2.656 0.057 SL61 18 5.931 0.177
The CRM are used to monitor the laboratory accuracy and are adequate to detect bias or drift of the results when they are reviewed over a certain length of time. The CRMs were purchased from Rocklabs, New Zealand and were prepared from powered feldspar, basalt and fine gold. The CRMs denoted with an S- prefix contain pulverized pyrite (2.6% to 3.0% sulphur).
Out of the 248 CRMs analyzed at ALS, 212 (85%) yielded results within three standard deviations from the mean, which is considered as acceptable (Figures 11.3 to 11.4 and Table 11.3).
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Figure 11.3 – QC Samples – CRMs – Time Series Standardized Variable (Z-score); (Extreme Values Removed)
Figure 11.4 – QC Samples – CRMs – Time Series Moving Average
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The batches containing six (6) of the CRMs that failed the test were re-analyzed and the new results were within the three (3) standard deviations threshold. Four (4) batches were left in the database although they were outside of the threshold. The assays associated with the remaining 26 CRMs were not re-analyzed since the batches they were in did not contain any significant gold values.
The mobile average plot of the CRM results does not indicate an analytical bias but displays a significant increase in failures in early June. Notably, the Z-score plot that incorporates the results from all the CRMs in one graph shows that CRMs SG56 and Si81 returned an anomalous number of failures.
11.2.5 COARSE DUPLICATE SAMPLES
Quarter core from 439 samples were used to serve as duplicates to monitor the precision of the analyses. No duplicate samples were prepared from RC cuttings as most of the 2018 holes were diamond drill holes.
The mean of the duplicate analyses is 9% lower than that of the original (Table 11.2). Indeed, 68% of the original assays are higher than the duplicates in the individual pairs. It is noteworthy that the proportion of original samples that have a higher value than the duplicate samples remains essentially the same, regardless of the range of gold values, except for the pairs of samples for which the values exceed 1 g/t Au.
The duplicate assays for the samples that yielded values higher than 1 g/t Au are higher in 54.4% of the cases and lower in 45.6% of the cases. In this case, 37 samples (54.4%) exceeded the values of the duplicate analysis and an opposite relationship is true for 31 samples (45.6%). Only 32% of the pairs have a difference less than 10%. Different statistical tests show a negative bias in the duplicate assays (Figures 11.5 to 11.7).
Table 11.2 – QC Samples – Comparison of the Results from the Pairs of Duplicate Analyses from Diamond Drill Holes (SMF BF Laboratory)
Samples Within % Original ≥ Duplicate Criteria Count Relative Difference Duplicate > Original ±10% ±20% 300 139 141 234 All Samples 439 (68.3%) (31.7%) (32.1%) (53.3%) 278 120 123 216 Au≥0.025 g/t 398 (69.8%) (30.2%) (30.9%) (54.3%) 241 89 100 176 0.025 g/t ≤Au<1.0 g/t 330 (73.0%) (27.0%) (30.3%) (53.3%) 37 31 23 40 Au ≥ 1.0 g/t 68 (54.4%) (45.6%) (33.8%) (58.8%)
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Table 11.3 – Results from the CRMs Analyzed at ALS Ouagadougou
PERIOD (date) CERTIFIED OBSERVED ACCEPTED FAILURES Above CRM Count Rate Below 3 Rate From To Value S.D. Mean S.D. Min. Max. Count Count 3 (%) StdDev (%) StdDev _OxE113 23 04-16-2018 07-10-2018 0.609 0.014 0.61 0.02 0.59 0.65 23 100.0 0 0 0 0.0 _OxF125 18 05-22-2018 06-19-2018 0.806 0.02 0.85 0.13 0.80 1.38 16 88.9 2 0 2 11.1 _OxF142 23 07-04-2018 08-04-2018 0.805 0.019 0.82 0.05 0.61 0.87 20 87.0 3 1 2 13.0 _SF85 56 01-05-2018 07-21-2018 0.848 0.018 0.85 0.03 0.79 0.99 54 96.4 2 1 1 3.6 _SG56 20 05-30-2018 08-11-2018 1.027 0.033 1.16 0.35 0.62 1.98 11 55.0 9 3 6 45.0 _SG66 24 04-21-2018 06-19-2018 1.086 0.032 1.06 0.08 0.90 1.34 20 83.3 4 2 2 16.7 _SH82 9 01-05-2018 06-19-2018 1.333 0.027 1.33 0.10 1.07 1.43 7 77.8 2 1 1 22.2 _Si81 8 06-09-2018 06-14-2018 1.79 0.03 2.41 1.43 1.82 5.96 2 25.0 6 0 6 75.0 _SJ80 49 05-30-2018 08-11-2018 2.656 0.057 2.78 0.76 1.84 6.73 41 83.7 8 4 4 16.3 _SL61 18 04-16-2018 08-01-2018 5.931 0.177 5.86 0.11 5.64 6.00 18 100.0 0 0 0 0.0
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Figure 11.5 – QC Samples – Comparison of the Results from the Pairs of Coarse Duplicates (Laboratory Rejects) from Diamond Drill Holes Analyzed by the SMF BF Laboratory – Mean vs. Half Relative % Difference Plot
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Figure 11.6 – QC Samples – Comparison of the Results from the Pairs of Coarse Duplicates (Laboratory Rejects) from Diamond Drill Holes Analyzed by ALS – Q-Q Plot
Duplicate Relative Original Au Au Difference Mean 0.96 0.87 -9% Standard Error 0.17 0.13 Median 0.18 0.15 Mode 0.08 0.01 Standard Deviation 3.60 2.63 Variance 12.92 6.91 Kurtosis 172.89 46.29 Asymmetry 11.68 6.28 Range 59.80 25.5 Minimum 0.01 0.01 Maximum 59.80 25.50 Sum 421.37 381.56 Count 439 439
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Figure 11.7 – QC Samples – Comparison of the Results from the Pairs of Coarse Duplicates (Laboratory Rejects) from Diamond Drill Holes Analyzed by ALS – Precision Plot (Mean vs Half Absolute % Difference)
0-5% 141 32.1% 5-10% 98 21.2% 10-20% 83 18.9% >20% 122 27.8% Total Pairs 444
Considering that the duplicate samples consist of quarter core, the difficulty to replicate the original analyses can partly be explained by the lower representativity of the duplicate samples. Indeed, the effects on the analyses of the volume variance between the mass of the original half core and of the quarter core in gold deposits have been widely documented.
In addition, the high variability of the gold values present in the Nabanga deposit contributes to the variance between the original and duplicate assays.
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11.2.6 COARSE BLANK SAMPLES
Blank samples are used to detect possible sample-to-sample contamination resulting from crushers or pulverizes that are not properly cleaned after each sample. Mis-sequencing of the samples by the geologists or at the laboratory may also be detected by the blank QC samples.
A total of 220 blank samples were inserted into the sample stream sent to ALS. All the blank samples returned gold values below the fail-pass threshold set at 0.08 g/t Au.
Barren sandstone originating from a site close to the Mana mine served as blank samples. This material, albeit not certified, provides the advantage of being a coarse blank that will be crushed by the laboratory. Although not ideal, the use of this non-certified, coarse blank is considered as adequate for the need of an estimation of resources in the inferred category.
11.2.7 SAMPLE PREPARATION AND ANALYSIS
11.2.7.1 Primary Laboratory
Preparation and analysis of all the samples were performed at the ALS laboratory in Ouagadougou.
At the preparation stage, the samples were crushed to 70% passing 2 mm (10 mesh) and quartered to obtain a 250-g split of crushed material. The 250-g riffle split was pulverized to better than 85% passing 75μm (200 mesh) and quartered to get a 50-g split. The remainder of the coarse rejects and of the pulps was saved for future reference.
All samples were analyzed for gold by the Fire Assay method on 50-g aliquots with a lower detection limit of 0.01 ppm gold and an AAS finish. The samples grading over the upper detection limit of 100 ppm gold were re-assayed using a 50-g fire assay procedure with a gravimetric finish.
11.2.7.2 Secondary Laboratory (Check Laboratory); SEMAFO's Mana Laboratory
A series of selected pulps that had been analyzed at the ALS facilities between April 16 and August 11, 2018 was shipped to the SMF BF laboratory at the Mana mine at the end of every quarter. The samples were selected on the basis of analytical results from completed core and RC holes. The samples were bundled in separate batches corresponding to their position in the alteration profile (Oxide-Transitional-Sulphide). The 405 pulps were comprised of 19 oxide and 386 sulphide samples and were re-analyzed during the third quarter of 2018.
The main purpose of these check analyses was to ensure that coherent inter-laboratories results were yielded.
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11.2.8 SEMAFO'S MANA MINE LABORATORY
11.2.8.1 Preparation, Analyses, QA/QC
Although the samples were shipped to SMF BF under the form of pulps, the preparation stage is described for comparison with the sample preparation in commercial laboratories.
Upon receival, the samples are registered and weighed before being dried for between 8 and 10 hours depending on moisture. Every ±2 kg sample is crushed to 70% passing 2 mm (10 mesh) and quartered to obtain a 250-g split of crushed material. The 250-g riffle split is pulverized to 85% passing 75μm (200 mesh) and quartered to obtain a 50-g sub-sample. The remainder of the coarse rejects and of the pulps is saved for future reference.
All preparation equipment is flushed with barren material prior to the commencement of each run. The crushers and pulverizes are cleaned by compressed air between each sample.
All samples were analyzed using 50-g Fire Assay method with an AAS finish with a lower detection limit of 0.01 ppm gold. The samples grading over 15 ppm gold were re-assayed using a 50-g fire assay procedure with a gravimetric finish.
The SMF BF's internal QA/QC protocol involves insertion of one CRM, one field duplicate, two pulp duplicates and one blank pulp into each analysis batch of 20 client samples. A minimum of 5% additional pulp check assays are performed on all batches, depending on the number of anomalies present within a given batch.
The laboratory submits the assay reports as digital data files.
11.2.8.2 Results
The average grade of the original oxide samples is 0.265 g/t Au and varies between 0.08 g/t and 0.86 g/t Au, as compared to the average grade of 0.257 g/t Au for the re-assay of the pulps. On average, the check samples have a lower grade of 2.87%, with 56% of these check samples yielding a grade higher than in the original corresponding sample. 39% of the sample pairs exhibit a variance of 10% or less in grade. Statistical tests confirm the absence of bias, although a statistically valid conclusion cannot be reached on the basis of such a low population, that is, 19 samples.
The average grade of the original sulphide samples is 0.95 g/t Au and ranges from 0.08 g/t Au and 37.40 g/t Au. The grade of the duplicate samples is 2.32% lower, on average, and 56% of these check samples yielded a grade higher than in the original sample. 57% of the sample pairs exhibit a variance of 10% or less in the grade. The statistical tests seem to confirm the presence of a significant bias toward the grade of the duplicate pulp analyses by SMF BF that are systematically higher than the original assays. However, the impact of this bias is probably negligible, considering
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that close to 60% of the sample pairs yielded a difference that is lower than 10% and that the difference between the average of the original and check samples populations hardly reaches 2%.
11.2.9 DENSITY
SEMAFO started completing density determinations on July 5, 2018 and 2,178 samples from 21 holes were processed over a two-month period.
The weight-in-air-weight-in-water method was used to measure the bulk density. The core samples submitted to the measurements were selected based on lithology and intensity of supergene alteration.
The results obtained for these bedrock samples were inconsistent and deemed unreliable to include in the resource estimation. The data should be re-evaluated and correction be made to ensure reliability before including the density measurement into the database.
11.2.10 SEMAFO'S SECONDARY LABORATORY CERTIFICATION AND ACCREDITATION
Several batches of pulps from samples that had been analyzed at the ALS facilities were shipped to SEMAFO's internal laboratory (SMF BF) at their Mana mine in Burkina Faso to serve as a check on the original analyses by a secondary laboratory.
The SMF BF laboratory operates under conditions consistent with industry best practices and standards. The facility utilizes a QA/QC system that includes insertion of standard, blank and duplicate samples and participates in regular round robin programs to monitor for bias.
Although the SMF BF laboratory is not independent of SEMAFO, the use of this facility strictly as a secondary laboratory is acceptable for a check on resources at the PEA level.
11.2.11 CONCLUSIONS, RECOMMENDATIONS
Although the standards in several batches failed the acceptability tests, they did not indicate a systematic bias and corrective actions were taken by SEMAFO for the batches with significant values.
The blanks did not indicate any material contamination of the samples or sample mis-sequencing.
Some bias was observed in the quarter core duplicate samples that returned an average grade 9% lower than the average of the original. The volume variance (half core original vs quarter core duplicate) and the nugget effect of the deposit are partly responsible for the variance. Noteworthy is the fact that the number of pairs where the duplicate samples yielded a higher value than the original sample is close to equal to the number with the opposite relationship.
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The secondary laboratory returned values with an average high bias of 2.32% relative to the original pulps. Considering this relatively modest bias and the fact that 60% of the samples exhibit a variance of less than 10%, DRA/Met-Chem believes that the duplicate samples reproduce reasonably well the original data.
Although the results from some QC samples did not perform in an outstanding manner, they are reasonably reliable for the needs of an estimation of resources exclusively in the lower level of confidence, the inferred category.
DRA/Met-Chem believes the analytical data from the drilling programs of 2016-2018 are sufficiently reliable to be used in a preliminary study.
However, for a future drilling program, DRA/Met-Chem recommends some actions for improvement in the quality and heightened confidence of the results:
• Document, to the extent possible, the causes of the significant increase in the failures on the CRMs assays in June 2018 and on the corrective actions taken by ALS to remedy the situation;
• Use the second half core rather than quarter core stubs as duplicate samples in order to obtain a more representative duplicate sample; alternatively, two quarter core samples may be prepared as the original and duplicate samples. A higher variability can be expected in the results, but the volume variance would not be involved from two (2) samples of the same mass;
• Perform density measurements, using proper methodology, including use of standards, duplicate determinations and calibration of the scale, and ensure proper validation the results before entry into the database.
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12 DATA VERIFICATION
12.1 Twinning of RC Holes by Orbis
Four (4) RC holes were twinned by Orbis using HQ diameter core holes (Table 12.1). DRA/MC reviewed the data and calculated the average grade of the mineralized intersections as logged by the geologists and recorded in the Lithology table of the drill database (Tables 12.1 to 12.4).
The NARC033/NADD023 and NARC184/NADD002 pairs exhibited high repeatability as quartz and pyrite are present in both holes, and the average gold grades are equivalent.
Quartz was found in both in NARC134/NADD013 and quartz and pyrite in the pair of NARC154 and NADD012 holes. However, the gold grades returned by the twinned holes provide conflicting information when gauged against the selected cut-off grade (COG) of 3.0 g/t Au applied in the resource estimation (over/below COG).
On the basis of the average grade of the mineralized intervals, the agreement between the results from the original and the twin holes is moderate. Part of the variability between the original and twinned holes can be accounted for by the large difference in volume (volume variance) between the samples retrieved by the RC holes (53/4 in or 133.35 mm diameter) and the core drilling (half of 63.5 mm core) that may impact the analytical results by virtue of the volume variance.
The comparison of the results from twinned holes should not be limited to the average grade of a mineralized interval to assess the repeatability. DRA/Met-Chem believes that twinned holes should also be compared on positioning and thickness of the mineralisation and on significant geological features. This is particularly true where the studied variable is characterized by high short-range variability, like in a gold deposit.
The position and length of the mineralized intervals match in all holes in the respective pairs. The analytical results show that no downhole contamination in the RC holes occurred and they do not
suggest a systematic bias in the average grades.
A striking feature observed in all the original RC holes is the presence of one or two high-grade gold samples (7.400 to 17.500 g/t Au). Remarkably, every core hole picked up these high gold values intersected in the RC hole except for the NARC134/NADD013 pair, which can reasonably be attributed to the intrinsic nugget effect of the deposit.
Although the number of twinned holes required for conclusive statistical analysis is not reached, the data provided by the four pairs of holes suggest a moderate to fair level of repeatability of the original holes.
The cross sections showing the twinned holes are presented in Snowden's Technical Report (2015).
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Table 12.1 – Comparison of Twinned Holes NARC033 and NADD023 (5 m separation)
INTERVAL GRADE HOLE ID DESCRIPTION (m) (g/t Au) NARC033 66-67 0.580 67-80 m: Trace to 3% pyrite; alteration, foliation/schistosity; NARC033 67-68 16.850 NARC033 68-69 11.300 67-68 m: quartz vein. NARC033 69-70 1.070 NARC033 70-71 0.370 NARC033 71-72 0.150 NARC033 72-73 0.100 NARC033 73-74 0.100 NARC033 74-75 0.120 NARC033 75-76 0.960 NARC033 76-77 1.090 NARC033 77-78 1.170 TOTAL 12 m 0.282 NADD023 60-61 0.270 60.0-70.7 m: strong deformation, 1 to 4% pyrite; quartz veinlets; NADD023 61-62 0.460 NADD023 62-63 10.300 62.9-63.9 m: quartz vein. NADD023 63-64 16.100 NADD023 64-65 1.960 NADD023 65-66 0.120 NADD023 66-67 1.220 NADD023 67-68 0.610 NADD023 68-69 0.210 NADD023 69-70 0.640 NADD023 70-71 0.490 TOTAL 11 m 0.294
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Table 12.2 – Comparison of Twinned Holes NARC184 and NADD002 (2 m separation)
INTERVAL GRADE HOLE ID DESCRIPTION (m) (g/t Au) NARC184 86-87 1.240 86-101 m: Trace to 3% pyrite; NARC184 87-88 3.560 86-89 m: Quartz vein; NARC184 88-89 17.500 96-97 m; Quartz vein. NARC184 89-90 0.430 NARC184 90-91 0.120 NARC184 91-92 0.190 NARC184 92-93 0.060 NARC184 93-94 0.060 NARC184 94-95 0.170 NARC184 95-96 0.190 NARC184 96-97 0.780 NARC184 97-98 0.250 NARC184 98-99 0.080 NARC184 99-100 2.470 TOTAL 14 m 1.936 NADD002 75-76 1.810 75-85 m: Mineralized interval; NADD002 76-77 0.320 83-84 m: Quartz vein, trace pyrite; NADD002 77-78 1.420 75-83 m: core non recovered; 84-85 m: core non recovered. NADD002 78-79 4.400 NADD002 79-80 2.160 NADD002 80-81 0.420 NADD002 81-82 3.400 NADD002 82-83 1.000 NADD002 83-84 2.140 NADD002 84-85 10.310 TOTAL 10 m 2.738
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Table 12.3 – Comparison of Twinned Holes NARC134 and NADD013 (1 m separation)
INTERVAL GRADE HOLE ID DESCRIPTION (m) (g/t Au) NARC134 36-37 0.440 37-39 m: 2-3% quartz vein. NARC134 37-38 9.130 NARC134 38-39 11.000 NARC134 39-40 0.760 NARC134 40-41 0.270 TOTAL 5 m 4.320 NADD013 37-38 0.510 37.1-41.4 m: Mineralized interval; NADD013 38-39 3.200 38.2-39.9 m: Shear, quartz vein. NADD013 39-40 0.840 NADD013 40-41 0.130 NADD013 41-42 0.540 TOTAL 5 m 1.044
Table 12.4 – Comparison of Twinned Holes NARC154 and NADD012 (3 m separation)
INTERVAL GRADE HOLE ID DESCRIPTION (m) (g/t Au) NARC154 55-56 0.200 55-64 m: Quartz vein, 2-3% pyrite; NARC154 56-57 7.400 64-65 m: 2-3% pyrite. NARC154 57-58 0.820 NARC154 58-59 1.690 NARC154 59-60 1.660 NARC154 60-61 1.420 NARC154 61-62 1.020 NARC154 62-63 0.810 NARC154 63-64 0.260 NARC154 64-65 0.290 TOTAL 10 m 1.557 NADD012 53-54 1.010 53-59 m: Shear, 4-5% pyrite, trace arsenopyrite; NADD012 54-55 0.600 57-58 m: quartz vein; NADD012 55-56 0.240 59-60 m: core non recovered. NADD012 56-57 0.940
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INTERVAL GRADE HOLE ID DESCRIPTION (m) (g/t Au) NADD012 57-58 8.100 NADD012 58-59 15.430 NADD012 59-60 0.440 TOTAL 7 m 3.822
12.2 Orbis Database (Assay) by Snowden
Snowden checked a random selection of 10% of all assay certificates, against the data in the database and found minor but no material issues.
12.3 Site Visits
One Snowden's QP geologist visited the site between March 13 and 17, 2015. Snowden did not collect independent QP check samples, but considering the QA/QC results, they did not believe independent sampling was required at that stage.
12.4 Verifications by DRA/Met-Chem
DRA/Met-Chem reviewed the measurements of the path of the drill holes and found no unusual deviation.
DRA/Met-Chem reviewed the detailed analysis of the QC samples results by François Thibert, P. Geo.; a QP of SEMAFO and generally agrees with the conclusions.
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13 MINERAL PROCESSING AND METALLURGICAL TESTING
The information in this Section of this Report under E. Pengel’s responsibility is largely drawn or summarized from the Report available on SEDAR entitled: “Yactibo Permit Group, Nabanga Gold Deposit, prepared by Snowden Mining Industry Consultants for SEMAFO Inc., Project No. AU4582, NI 43-101 Technical Report, June 2015” as well as “Nabanga Desktop Study – For Internal Use Only, Report # INT-RPT 047, dated January 2015, prepared by Orbis Gold Limited.”
13.1 Metallurgical Test Work
DRA/Met-Chem has reviewed the metallurgical test work results attached to the reports above mentioned, which included:
• Lycopodium Minerals QLD Pty Ltd, “Nabanga Project Metallurgical Test Work Review”, dated October 2012, Lycopodium Report 3110-STY-002_A, prepared for Orbis Gold.
• JK Tech Pty Lts, “SMC Test Report” dated July 2012, Report 12001/P79 prepared by JKTech Pty Ltd for Mt Isa Metals.
13.1.1 COMMINUTION TEST WORK
The tests performed included ball mill work index test and SMC tests. ALS Ammtec performed two (2) Bond Work Index tests. The material can be considered very hard and medium abrasive. The test results are in Table 13.1.
Table 13.1 – Bond Work Index Tests
BWi Ai Bond Index Sample (kWh/t) (g) Granodiorite 23.5 0.365 Quartz 22.9 0.262
Further comminution tests were SMC tests. These tests will indicate the type of grinding circuit can be selected. In general, the material is very hard. The results are in Table 13.2.
Table 13.2 – SMC Test Results
DWi DWi Mia Mih Mic Sample A b (kWh/m³) (%) (kWh/t) (kWh/t) (KWh/t) Granodiorite 8.08 79 21.7 16.7 8.6 89.1 0.39 Quartz 7.41 72 20.7 15.6 8.1 67.4 0.55
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13.1.2 CYANIDATION TEST WORK
ALS Ammtec performed cyanidation tests. Process development undertaken during the PEA aimed at establishing a conventional sulphide gold/silver processing facility considering the constructability, operability and maintainability of the processing facility. The results indicate that material has refractory characteristics as a high concentration of cyanide is required to obtain good leach results. Tables 13.3 and 13.4 list the results for standard and intensive leach tests.
Table 13.3 – Standard Leach Test 0.035% w/v NaCN Concentration Test Results
P₈₀ Grind Size Oxide Material Sulphide Material (mm) Au recovery Au feed grade Au recovery Au feed grade (%) (g/t) (%) (g/t) 0.106 70 7.0 47 7.5 0.053 78 7.7 56 7.6 0.025 81 8.3 60 7.3
Table 13.4 – Intensive Leach Test 0.5% w/v NaCN Concentration Test Results
P₈₀ Grind Size Oxide Material Sulphide Material (mm) Au Recovery Au Feed Au recovery Au Feed (%) Grade (g/t) (%) Grade (g/t) 0.106 90 7.3 81 7.1 0.053 95 7.4 86 8.2 0.025 99 7.2 95 7.3
13.1.3 FLOTATION TEST WORK
As reported in the Desktop Study, the report was done using a feed grind size F₈₀ = 0.075 mm. The flotation concentrate yield was 5% of the feed containing 79.2 percent of the gold in the feed. The flotation condition details were not specified, except that 20 g/t CuSO₄ was used as activator and PAX and MIBC as collector and frother respectively. Further, flotation was performed at a natural pH of 8.3.
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14 MINERAL RESOURCE ESTIMATES
14.1 Introduction
The Mineral Resources for the Nabanga deposit were estimated by SEMAFO's resource geologist. The resource estimate has an effective date of December 31, 2018 and supersedes the previous estimates completed in 2012 and 2015 by Snowden.
The resources estimate includes exclusively the fresh (sulphide) portion of the Nabanga deposit. The resources in the oxide and transitional zones were excluded from this study as SEMAFO considers they only represent a marginal proportion of the entire deposit.
The results from this estimate served as the basis for this Preliminary Economic Assessment (PEA). The results from the PEA were presented in a press release dated September 30, 2019 and entitled: "SEMAFO: Positive PEA Results for Nabanga; After-tax Net Present Value of $100 Million".
The methodology and parameters used in the resources estimate are presented in the following sections.
14.2 Drill Hole Database Construction and Validation
SEMAFO updated the database used by Snowden for the resources estimations of 2012 and 2015 that were based on the holes drilled until August 2, 2012. SEMAFO entered into the database the latest holes drilled by Orbis in 2012-2013 and the holes drilled by SEMAFO between 2016 and 2018. The new lithological descriptions deriving from re-logging by SEMAFO of the RC chips from previous RC holes were incorporated into the database.
The data required for the resources estimation was imported into dedicated software MICROMINE's Geobank Database that provides an environment for capturing, validating, storing and managing data from diverse sources. Several steps of validation of the database by the field and project geologists and by the GIS geologists were performed prior to extracting and importing the required fields from the master database into the resources database. Discrepancies in parameters such as collar location and elevation, permitted minimum/maximum values, overlap and gaps in the lithologies or sample intervals, sample numbering were checked. The Geobank module added a layer of validation during data transfer.
The data used for the resource estimate includes the results from RC and core drilling by Orbis and SEMAFO. A total of 456, RC, 50 MP, and 37 core holes directly associated with the North, Central and South Zones of the Nabanga deposit were used in the estimate (Table 14.1). The core holes pre-collared by RC drilling are referred to as MP (Multi-Purpose) holes.
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Table 14.1– Summary of Database Entries by Hole Type
Sum of Cumulative Number of Hole Type Count Assayed Length (m) Assays Length (m) DDH 37 8,468 8,137 7,635 RC 456 47,678 26,356 34,815 MP 50 16,228 3,368 4,061 Trenches - - - - Total 543 72,374 37,861 46,511
The fields imported from the geologists' tables for the resources estimate are listed in Table 14.2.
Table 14.2 – List of Fields Contained in the Drill Hole Database
Collar Survey Regolith Lithology Density Assays Hole ID Hole ID Hole ID Hole ID Hole ID Hole ID Location X Zone From From From From Location Y Azimuth (°) To To To To Location Z Dip (°) Profile Code Litho Code Density Sample ID Length (m) Depth (m) Au (g/t) Azimuth (°) Dip (°)
14.3 Geological Modelling
Building on drilling data gathered by Orbis, SEMAFO started by constructing a 3D model using the historical data and re-interpreted the geology and structure of the Nabanga deposit. The new interpretation stemming from this work differed from Orbis' and was successfully tested by drilling in 2018.
Based on this information and an updated database, SEMAFO resource geologist interpreted the geology and mineralization in 2D, on cross-sections 40 m and 80 m apart. The outline of the polygons was digitized by snapping on the contacts of logged lithologies, quartz percentage, and gold assay data. A modelling COG of 0.2 g/t Au with a minimum width of 2 metres were used to delineate the mineralized envelope.
Two (2) domains of gold mineralization were created to separate the high-grade, central quartz- filled portion of the shear and the lower-grade sheared and altered granodiorite halo bracketing it. A cut-off grade of ≥ 0.2 g/t Au to ≤ 1 g/t Au was used to constrain the low-grade domains and a cut- off of >1 g/t Au was applied to delineate the high-grade domain, before the grade interpolation process.
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The mineralized volumes were created by connecting the polygons from section to section to create a 3D geological model (wireframe).
The grade estimation was constrained to within the two (2) mineralized envelopes (i.e. namely Low Grade (LG) and High Grade (HG)), in order to avoid smearing of values between the low- and high- grade mineralization and prevent the influence from the barren host rock or isolated values of adjacent mineralization on the estimation process. For the purpose of the Reporting of Resource Estimate for the Nabanga deposit, only the blocks constrained within the High Grade envelope have been considered.
The late North and South faults offsetting the deposit were also modeled.
The Micromine software, Version 16.1 was used by SEMAFO to construct the 3D model.
14.4 Basic Statistics Calculations
Basic statistics were calculated on the raw data to provide easy visualization of the interpolated quality element (Au) (Table 14.3).
The assays below the detection limit were reset to a value equivalent to half the detection limit.
Table 14.3 – General Statistics on the Gold Values (Au) within the HG Wireframe
Composites (1m) Raw Data Count 938 951 Min 0.01 0.005 Max 70 154 Range 69.99 153.995 Mean 7.15 7.49 Median 2.69 2.69 Mode 70 1.85 Stdev 11.897 14.142 Variance 141.546 199.988 COV 166% 189%
14.5 Composite Data
The assayed intervals of the HG wireframe represent a total length of 936.8 m. A statistical analysis of the sample length show that the majority (90%) of assay intervals (mode) have a length of 1.0 m whereas 8% have a length of less than 1.0 m and 2% have a length greater than 1.0 m. Minimum
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sample length is 0.45 m and maximum sample length is 4.0 m where only three (3) samples had a length over 2.0 m.
The drill hole data were composited downholeusing the compositing module in Datamine Software. By setting the mode to 1, it forces all samples to be included in one of the composites by adjusting the composite length, while keeping it as close as possible to interval. The maximum possible composite length will then be 1.5*interval. .
Prior to grade interpolation, capped assay intervals within mineralized intercepts were composited into 1-m intervals. The average gold grade of nearby sub-blocks was estimated from only nearby composites included in the same mineralized envelope. Compositing standardizes the length of assay intervals within the mineralized intercepts. The composite length is selected to reflect the average width (along the X axis) of blocks being interpolated. Thus, blocks can be interpolated from composites having a length that compares with the intercept length of an inclined hole within a block. This ensures that block grades will include some of the dilution of the drill hole intercepts within blocks and that the estimates above cut-off are realistic. Datamine adjusts the length of composites to the length of mineralized intercepts.
Within the HG wireframe, a total of 938 composites were downloaded into the datamine software. The composite statistics for Au were compared to the results from the raw data (Table 14.3).
14.6 Capping
The study based on log probability plot and drawing curves of % cumulated metal vs. % cumulated sample proportion indicated a top-cut grade of 70 g/t Au. Capping the extremely high values was done on the raw data, prior to compositing.
The impact of capping on the total gold content of the deposit at 70 g/t Au is not significant. This value caps the grade of 0.95% of the samples and reduces the total gold content by approximately 4.2%.
14.7 Density
The density used for the estimation of the resource was set at 2.70 t/m³. An average density was used for the entire portion of the deposit, since the estimate only includes the fresh (non-weathered) portion of the mineralization. The assigned bulk density of 2.70 is based on the average of the measurements for quartz vein material by Orbis, and was used by Snowden in 2012 and 2015. The density determinations by SEMAFO’s geologists were rejected as unreliable.
The information on oxidation contained in the drill hole database for all the holes (n=626) was used to generate a surface at the interface at the top of the Transitional Zone and the Fresh Zone to cut the portions of the blocks extending into the weathered portion of the laterite profile.
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Depth of those contacts are derived from description along drill holes. The corresponding control points are connected by a series of strings on individual EW sections. A 3-D surface model is created by triangulation for both zones. Modelled surfaces may change abruptly from one section to the next. Those artefacts indicate that the interpreted contact surfaces are given with a significant uncertainty, but generally those contacts are rather flat. The delineation of the different oxidation facies may have an impact on tonnage (and metal content) of the estimated resources, through density. It also may impact the reserves through mill recovery, maximum pit slope angles as well as mining and processing costs.
14.8 Variogram Modelling
The spatial continuity of top-cut composites within mineralized solids of the South, North and Central Zones was assessed with variograms or correlograms. A correlogram examines the decrease of the correlation (measured by a correlation coefficient from -1 to +1) of composite grades in any given direction as the distance between composites increases along that direction. Composite pairs in the appropriate direction +/- an “angular tolerance” are selected and classified into various distance bins based on a given lag along the direction being considered. A correlation coefficient is calculated for each bin from the pairs available in that bin. By graphing 1-correlogram, a curve is drawn showing increasing average grade differences between composites as the distance increases, which is like a traditional variogram. The sill of this “pseudo-variogram” is generally around 1.0 and it corresponds to a correlation coefficient of zero (i.e., no correlation between composite grades). Correlograms are generally preferred to traditional variograms since they are considered more robust with respect to outliers and non-stationary features such as trends and proportional effect (i.e., variability increasing with grade). Correlograms have been calculated but they are presented herein as variograms by graphing the function: 1-correlogram.
All variogram analysis and modelling were performed using the Geostat+, Version 2.0 module
3D variograms of 1-m composite data for the South, North and Central Zones have been computed separately along horizontal strike direction of N070, N060 and N050 respectively and dip directions (dipping 65° to the north-north-east). Also, variograms along the drill hole direction (short distance variogram) were computed separately from those computed along the average strike and dip directions (long distance variogram) to facilitate the interpretation. In all three (3) zones (see Figures 14.1 to 14.3), short distance variograms are characterized by a relative nugget effect of 30% with a short range of 3 m. For the South and North Zones, variograms along strike and dip display a 40-50 m range. The proposed model is made of one exponential component that is isotropic with a range of 30 m. For the Central Zone, the variogram along dip (-60 to N330) displays a 60 m range whereas the variogram along strike (horizontal N060) displays a 40 m range. The model is made of one anisotropic exponential component with ranges of 36 m and 30 m.
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Based on the variogram study, the search ellipsoid parameters determined for grade interpolation are as follows:
• 50 to 125 m along the major axis;
• 50 to 125 m along the semi-major axis;
• 10 to 25 m along the minor axis.
Due to the relatively simple geometry of the Nabanga deposit, no subdivision into different structural domains was necessary to allow the search ellipse to properly code all the blocks during grade interpolation, but three (3) sectors were separated by two (2) faults, with horizontal directions of the mineralization North 70°, North 60°, and North 50°.
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Figure 14.1 – Variograms of 1 m Composites for South Zone (Short and Long)
Nabanga - HG South Zone 1m composites (top-cut 70g/t Au) 1.400 ABS,Au70 1.260 ! 18
1.120
SILL 0.980 ! 155 ! 365 ! 236 ! ! 95 58 ! 0.840 16
! 0.700 565
0.560
0.420
0.280
! 33 0.140
Distance 0.000 0.0 1.0 2.0 3.0 4.0 5.0 6.0 7.0 8.0 9.0 10.0
Variable : Au70 Date : 05-11-2019 Variogram : Absolute File : varcmphgshort.gsd ! Direction : DH Azimuth 145.00: Dip -60.00: Tolerance :40.00 Lag Dist : 1.10
Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 70.0/0.0/-60.0)
Nabanga - HG South Zone 1m composites (top-cut 70g/t Au) 1.400 ABS,Au70 1.260
1.120 ! ! ! 8 45 14 ! 87
SILL ! 0.980 1
0.840
0.700 ! 17
0.560
0.420
0.280
0.140
Distance 0.000 0.0 20.0 40.0 60.0 80.0 100.0 120.0 140.0 160.0 180.0 200.0
Variable : Au70 Date : 05-11-2019 Variogram : Absolute File : varcmphgslong.gsd ! ! Direction Strike: Dip Azimuth :70.00 340.00 Dip : 0.00 -60.00 Tolerance :20.00 20.00 Lag Dist :45.00 20.00 Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 70.0/0.0/-60.0)
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Figure 14.2 – Variograms of 1m Composites for Central Zone (Short and Long)
Nabanga - HG Central Zone 1m composites (top-cut 70g/t Au) 1.400 ABS,Au70 1.260 ! 14
1.120
SILL 0.980 ! ! 14 ! 103 224 ! ! 155 ! 40 0.840 63
! 0.700 321
0.560
0.420
0.280
0.140 ! 25
Distance 0.000 0.0 1.0 2.0 3.0 4.0 5.0 6.0 7.0 8.0 9.0 10.0
Variable : Au70 Date : 05-11-2019 Variogram : Absolute File : varcmphgcshort.gsd ! Direction : DH Azimuth 145.00: Dip -60.00: Tolerance :40.00 Lag Dist : 1.10
Gamma = N(0.3000) + E(0.7000, 12.0/10.0/1.0, 330.0/-60.0/0.0)
Nabanga - HG Central Zone 1m composites (top-cut 70g/t Au) 1.400 ABS,Au70 1.260
! 124
1.120
! 457 ! ! 686 SILL 182 ! ! ! ! 0.980 175 247642 249 ! ! 355 318 ! 314 ! ! 90 0.840 111
! 61
0.700
0.560
0.420
0.280
0.140
Distance 0.000 0.0 20.0 40.0 60.0 80.0 100.0 120.0 140.0 160.0 180.0 200.0
Variable : Au70 Date : 05-11-2019 Variogram : Absolute File : varcmphgclong.gsd ! ! Direction Strike: Dip Azimuth :60.00 330.00 Dip : 0.00 -60.00 Tolerance :20.00 20.00 Lag Dist :45.00 20.00 Gamma = N(0.3000) + E(0.7000, 12.0/10.0/1.0, 330.0/-60.0/0.0)
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Figure 14.3 – Variograms of 1m Composites for North Zone (Short and Long)
Nabanga - HG North Zone 1m composites (top-cut 70g/t Au) 1.400 ! 5 ABS,Au70 ! 23 1.260 ! 13
1.120
! 39 ! 115 SILL ! 0.980 1 ! 63
0.840
! 0.700 195
0.560
0.420
0.280
0.140
Distance 0.000 0.0 1.0 2.0 3.0 4.0 5.0 6.0 7.0 8.0 9.0 10.0
Variable : Au70 Date : 05-11-2019 Variogram : Absolute File : varcmphgnshort.gsd ! Direction : DH Azimuth 145.00: Dip -60.00: Tolerance :40.00 Lag Dist : 1.10
Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 50.0/0.0/-65.0)
Nabanga - HG North Zone 1m composites (top-cut 70g/t Au) 1.400 ABS,Au70 1.260
! 1.120 ! 241 47 ! 138 ! 188 ! 66 ! 40 SILL ! 232 ! ! 224 38 0.980 ! 109 ! 316 ! 61 ! 0.840 99 ! 346
0.700
0.560
0.420
0.280
0.140
Distance 0.000 0.0 20.0 40.0 60.0 80.0 100.0 120.0 140.0 160.0 180.0 200.0
Variable : Au70 Date : 05-11-2019 Variogram : Absolute File : varcmphgnlong.gsd ! ! Direction Strike: Dip Azimuth :50.00 320.00 Dip : 0.00 -65.00 Tolerance :20.00 20.00 Lag Dist :45.00 20.00 Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 50.0/0.0/-65.0)
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14.9 Resource Block Modelling
A block model was created using Datamine® software package. A unique block model was created for the South, Central and North Zones with parent block size of 5 m by 5 m by 5 m. Splitting of the parent block into sub-cells was done by using the parameter MAXDIP in Datamine. The degree of splitting is made dependant on the dip of the wireframe. By setting MAXDIP to 0 and SPLITS to 3 allow all parent cells that are intersected by the wireframe to be divided into 8 sub-cells in Y and Z, with minimum sub-block size of 0.005 m along X, 0.625 m along Y and 0.008 m along Z. Parent cells were estimated using a discretization of 3 m by 3 m by 3 m in the X, Y, and Z dimensions, respectively.
The block model was rotated by 55 to the East, which coincide with the general orientation of the mineralized structure. The coordinates of the origin correspond to the rotated model. Parameters are presented in Table 14.4. The block model extent is illustrated in Figure 14.4.
Table 14.4 – Nabanga Block Models Parameters
Parent Block Size (m) Block Grid Origin Block Model Rotation X Y Z X Y Z
All Zone 55 5 5 5 224,400 1,249,100 -270
Figure 14.4 – Map of the 3 Grids Used for Block Models
North
Central
South
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14.10 Grade Interpolation Methodology
Average block or sub-block grades were interpolated from nearby composites within the HG mineralized solid. Block grade interpolation was estimated by Ordinary Kriging (OK). The different variogram models used for kriging have been presented in a previous section.
Each block and sub-block are first coded according to its position relative to the oxidation surfaces. Bulk density is assigned according to the ore type of the block or the sub-block.
Three (3) successive interpolation passes were used in the estimation.
In the first pass, the search ellipsoid size was set at 50 m x 50 m x 10 m in the long, intermediate and short axes. The minimum and maximum of composites to interpolate a block were set at 7 and 25 and the maximum of composite per hole was restricted to 3. These constraints ensure that at least 3 holes are required to allow a block to be interpolated during this pass.
In the second pass, the search ellipsoid size was set at 100 m x 100 m x 20 m in the long, intermediate and short axes. The minimum and maximum of composites were set at 7 and 25 and the maximum of composite per hole was restricted to 3. Under these constraints, 3 holes are required to allow a block to be interpolated during this pass.
In the third pass, the search ellipsoid size was relaxed to 125 m x 125 m x 25m in the long, intermediate and short axes, to ensure that all the blocks situated within the envelope that were not coded in the previous passes will be blocks will be captured and coded. The minimum and maximum of composites were set at 4 and 25 and the maximum of composite per hole was restricted to 3. Under these constraints, two (2) holes are required to allow a block to be interpolated during this pass.
A summary of the block model set-up and parameters is presented in Table 14.5.
Table 14.5 – Interpolation Parameters
Items Description
Grade Interpolation Method Composites equal length 1m composites High Values Capping 70 g/t Au Elipse Orientation Based on variogram
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Items Description
Interpolation Pass Pass 1 Pass 2 Pass 3 Min. Number of Composites/Block 7 7 4 Max. Number of Composites/Block 25 25 25 Max. Number of Composites/Hole 3 3 3 Ellipse Size on the Major Axis (Strike) 50 100 125 Ellipse Size on the Semi-Major Axis (Dip) 50 100 125 Ellipse Size on the Minor Axis (Downhole) 10 20 25
14.11 Resource Validation
The first step in the Mineral Resources validation consisted in a visual validation: the estimated blocks have been compared to the composite and raw assays grades, on sections, plans and in 3D view. The correlation is considered to be very good.
The second step involved generating basic statistics of blocks and comparing them to the composite and assays used as input. The results from the descriptive statistics are presented in Table 14.6.
The comparisons show a good correspondence between the interpolated blocks (output of the resource model) and the composites (in/out of the model). As it could be expected, the ordinary krigging has a smoothing effect on the grade and limit the range of extreme values.
Table 14.6 – Basic Statistical Parameters from the Assays, Composites, and Blocks
North Zone Central Zone South Zone Parameter Assay Comp. Block Assay Comp. Block Assay Comp. Block Count 359 351 1,789,562 493 491 1,851,122 99 96 657,321 Average 6.98 6.98 5.7377 8.38 7.75 6.8891 4.89 4.71 4.7843 Median 2.92 2.92 2.67 2.68 2.1 2.12 Standard Deviation 10.52 10.19 3.063 16.88 13.45 4.052 9.56 8.56 3.341 Variance 110.71 103.83 9.385 285.08 180.76 16.42 91.30 73.27 11.16 COV 150.65 146.00 53.39 201.51 173.47 58.82 195.23 181.55 69.83 Range 63.905 63.38 23.9673 153.99 69.99 27.6357 60.11 60.11 19.1821 Minimum 0.005 0.02 0.5514 0.01 0.01 0.7607 0.09 0.09 0.6522 Maximum 63.91 63.40 24.5187 154 70 28.3964 60.2 60.2 19.8343
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14.12 Resources Definitions and Classification
14.12.1 CIM GUIDELINES & DEFINITION
The guidelines from the latest (May 10, 2014) version of the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards – For Mineral Resources and Mineral Reserves were followed in the resource classification: Selected items from the definitions are reproduced below as the full and official text is freely available on the CIM Website.
In addition to conformity with the guidelines adopted by CIM, the criteria used by SEMAFO's resource geologist for classifying the resources are based on the quality of the dataset and the certainty of continuity of geology and grade.
Definitions are from the Canadian institute of Mining, Metallurgy and Petroleum (CIM):
Mineral Resource
Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured categories. An Inferred Mineral Resource has a lower level of confidence than that applied to an Indicated Mineral Resource. An Indicated Mineral Resource has a higher level of confidence than an Inferred Mineral Resource but has a lower level of confidence than a Measured Mineral Resource.
A Mineral Resource is a concentration or occurrence of diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals in or on the Earth’s crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.
The term Mineral Resource covers mineralization and natural material of intrinsic economic interest which has been identified and estimated through exploration and sampling and within which Mineral Reserves may subsequently been defined by the consideration and application of technical, economic, legal, environmental, socio-economic and governmental factors. The phrase ‘reasonable prospects for economic extraction’ implies a judgment by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction. A Mineral Resource is an inventory of mineralization that under realistically assumed and justifiable technical and economic conditions might become economically extractable. These assumptions must be presented explicitly in both public and technical reports.
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
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information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes.
Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration. Confidence in the estimate is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure. Inferred Mineral Resources must be excluded from estimates forming the basis of feasibility or other economic studies.
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 drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.
Mineralisation may be classified as an Indicated Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such as to allow confident interpretation of the geological framework and to reasonably assume the continuity of mineralization. The Qualified Person must recognize the importance of the Indicated Mineral Resource category to the advancement of the feasibility of the project. An Indicated Mineral Resource estimate is of sufficient quality to support a Preliminary Feasibility Study which can serve as the basis for major development decisions.
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 drill holes that are spaced closely enough to confirm both geological and grade continuity.
Mineralisation or other natural material of economic interest may be classified as a Measured Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such that the tonnage and grade of the mineralization can be estimated to within close limits and that variation from the estimate would not significantly affect potential economic viability. This category requires a high level of confidence in, and understanding of, the geology and controls of the mineral deposit.
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14.12.2 NABANGA CLASSIFICATION
The mineral resource classification is based on certainty of geology and grades, which is, in most cases, related to the drilling density. Although the mineralized shear exhibits strong geological continuity and although a portion of the Nabanga deposit has been more densely drilled, all the resources were graded as Inferred Resources.
SEMAFO elected to constrain the resources using a Cut-Off Grade (COG) calculated on the basis of an potential underground exploitation in order to only state the resources that have good prospects for eventual economic extraction, in compliance with the CIM requirements.
14.13 Mineral Resource Statement
The Mineral Resources are stated using an average density and are constrained in a preliminary optimized pit. The resources are estimated exclusively for the mineralization within the fresh (sulphide) portion of the Nabanga deposit. Table 14.7 presents the results from the resource estimate.
Table 14.7 – Nabanga – Mineral Resources Estimate as of December 31, 2018, using a COG of 3.0 g/t Au
Total Gold Resources Category Tonnage (Mt) Grade (g/t Au) (x1,000 oz) Inferred 3.4 7.7 840
The Reader is cautioned that mineral resources that are not mineral reserves do not have demonstrated economic viability.
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15 MINERAL RESERVE ESTIMATES
This Report is a PEA Report, and as such, no Mineral Reserves have been estimated for the Nabanga Project, as per NI 43-101 regulations.
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16 MINING METHOD
The PEA was based on Mineral Resources Estimate published by SEMAFO with an effective date of December 31, 2018. DRA/Met-Chem has verified that the results of the PEA continue to be relevant and are valid for the updated Mineral Resource Estimate.
The Nabanga deposit will be mined by the use of the Open Pit and Underground mining methods, as described in this section.
16.1 Geotechnical Assessment
Peel Sullivan Meynink had been contracted by Orbis to provide preliminary geotechnical advice on the Nabanga Project (PSM 2014). No further work has been accomplished as part of the Nabanga PEA Project since then.
Peel Sullivan Meynink stated that rock mass where the open pit and the underground excavation could be located at Nabanga has a very good rock mass quality (RQD between 90 and 100). A high strength (UCS>80 MPa) rock is the dominant rock class.
The overall slope angles for the open pit was determined using an empirical method based on the rock mass ratings for different rock unit classes versus slope height. This method was considered adequate given the preliminary nature of the study (PEA Level), and given the relatively shallow depth of the proposed open pit excavation. However, for a subsequent evaluation stage (at PFS of DFS level), further investigation will be required to increase the amount of confidence.
The preliminary geotechnical pit slope design parameters are the following:
• An overall slope angle of 55° to 60° for the pit walls developed in a very good rock mass quality;
• An overall slope angle of 40° to 42° for the upper part of the pit wall in the poor-quality rock (overburden).
All the stopes are likely to be excavated in very good rock mass quality. With this condition, the preliminary geotechnical stope design parameters are:
• The stope width varied with the width of the mineralized material in the range of 1 to 7 m;
• The recommended maximum stope length is 30 m;
• Minimum crown pillar thickness is equal to the typical stope width.
Since the access, ramp and material sills are in a very good rock mass quality, the preliminary geotechnical ground support parameters are for an opening of 5.0 x 5.0 m:
• 2.5 m resin rebar or swellex on a pattern of 2.0 x 2.0 m for the roof (minimum);
• 1.8 m split set on a pattern 2.0 x 2.0 m for the wall (added for the study);
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• Galvanized screen (5.6 mm, 100 mm aperture) (added for the study).
The wall support and screen were not mentioned in the study or not recommended for excavation support; however, they have been added for safety reasons on miscellaneous possibility of small rock-falls condition.
For the vertical development, unsupported shafts up to 4 to 5 m diameter are expected to have few stability issues in very good rock mass. If the rock mass quality decreases, additional ground support will be required. All the vertical development connecting to surface will be designed to avoid any fault or instability zones.
Two (2) major faults intersect the mineralization zone. These structures have been described as highly to moderately weathered, with a RQD below 10. There is no evidence of gouge or clay. The orientation of both fault with the mineralization should not be problematic with respect to the open pit walls stability or underground stope stability.
16.2 Hydrology and Hydrogeology
No field work has been done so far on the Project concerning the hydrogeology, or the surface hydrology.
A river is located at the north-east of the mineralized zone, approximately 300 m away. During the rainy season, this river may not be enough to drain all the water, and a large flooded zone (swamp area) could be created along the river.
16.3 Transition Between Open Pit and Underground Mining
In order to determine the transition level between the Open Pit operation and the Underground operation, a calculation of the value of the mineralized material for each slice of 20 m from surface to the bottom of the mineralization has been estimated for both the Open Pit and Underground mining methods. The mining costs used are the ones described in the cut-off grade calculation in Sections 16.4 and 16.5. This method results in the optimum value of the mineralized material zone without taking in consideration timing and mine scheduling.
The optimal open pit limits were established with the NPV Scheduler (from Datamine software package) which uses the Lerch-Grossman algorithm for pit optimization. The pits limits were optimized using the parameter described in Section 16.4.
The Stope Optimizer tool from Deswik was used to define the potential stoping area. The parameters used for the Stope Optimizer are described in Section 16.5.
A pit value was evaluated for every 20 m step as it was mined by its own until the ultimate pit.
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Table 16.1 – Value of Mineralized Material Every 20 m Slice for Underground and Open Pit Scenarios
Elevation Underground Open Pit UG Open Pit mineralized mineralized Level Tonnes Au (g/t) Tonnes Au (g/t) material material value value SURFACE_180EL 100,000 5.8 240,000 5.5 5,700,000 23,300,000 180EL-160EL 220,000 5.6 250,000 6.6 10,700,000 24,400,000 160EL-140EL 200,000 5.9 180,000 7.5 12,100,000 16,400,000 140EL-120EL 250,000 6.1 130,000 8.5 16,900,000 13,200,000 120EL-100EL 240,000 6.7 80,000 9.9 19,200,000 10,100,000 100EL-80EL 280,000 6.3 150,000 6.7 21,000,000 3,500,000 80EL-60EL 300,000 6.6 210,000 5.4 25,400,000 19,100,000 60EL-40EL 290,000 7.1 70,000 7.7 29,100,000 300,000
It is to be noted that a significant portion of Optimized Stopes has been generated below the 40 m elevation, whereas no Pit Shell has been generated below that horizon.
Above elevation 140 m, the value of the 20 m slice is higher with the open pit mining and below, the value is better with the underground. The transition of mining method between surface and underground will be at elevation 140 m.
16.4 Open Pit Mining
16.4.1 PIT OPTIMIZATION
Open-pit optimization was conducted on the deposit to determine the economic pit limits. The optimization was carried out during the initial stage of the Project using initial cost, sales price, and pit and plant operating parameters.
The optimizer operates on a net value calculation for all the blocks in the model (i.e., revenue from sales of gold less operating cost).
Since this study is at a PEA level, inferred material has been considered in the optimization and mine plan.
Table 16.2 presents the pit optimization parameters.
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Table 16.2 – Pit Optimization Parameters
Description Unit Value
Mining Cost – Mineralized material US$/t (mined) 2.47 Mining Cost – Waste US$/t (mined) 2.36 Processing Cost US$/t (milled) 41.5 G&A US$/t (milled) 14.4 Refining US$/oz 3 Sales Price US$/oz 1,300 Mill Recovery % 81 State Royalties % 5 Development Tax % 1 Open Pit Dilution % 12.5 Overall Pit Slope ° 50 to 60
16.4.2 CUT-OFF GRADE (COG)
Using the economic parameters presented above, a COG of 2g/t of gold was calculated for the Open Pit Project. The COG is used to determine whether the material being mined will generate a profit after paying for the processing, transportation and administrative costs. Material that is mined below the cut-off grade is sent to the waste dump. The COG has been calculated according to the following formula:
(푀푖푛푖푛푔 푐표푠푡 + 푃푟표푐푒푠푖푛푔 푐표푠푡 + 퐺&퐴) ∗ 퐷푖푙푢푡푖표푛 퐶푂퐺 = (푆푎푙푒푠 푃푟푖푐푒 − 푅표푦푎푙푡푦) ∗ 푀푖푙푙 푅푒푐표푣푒푟푦)
16.4.3 PIT DESIGN AND SEQUENCING
A conventional truck/shovel operation will be used for all open pits. Drill and blast will be required almost at the beginning of the excavation work because there is almost no overburden (saprolite). Given the expected short life of the Open Pit operation, it is anticipated that a contractor will be selected and be responsible for all aspect of surface mining (operation, maintenance, office, garage erection and management).
The ramps and haul roads were designed with an overall width of 15.5 m. For double lane traffic (Figure 16.1), industry practice indicates the running surface width to be a minimum of 2.5 times the width of the largest truck. The overall width of a 50-tonne haul truck is 3.9 m which results in a running surface of 9.75 m. The allowance for berms and ditches increases the overall haul road width to 15.5 m. A maximum ramp grade of ten (10) % was used. This grade is acceptable for a 50-tonne haul truck.
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Figure 16.1 – Double Lane Ramp Design
Several pits have been designed for the Project. Figure 16.2 shows a plan view of the Nabanga pit designs.
Figure 16.2 – Nabanga Pit Designs
NORTH
The sequence will be directly related to free the face as soon as possible to the start the portal and ventilation entrance. Figure 16.3 shows the pit sequence.
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Figure 16.3 – Pit Sequencing
Pit Sequencing
Month-7 Month-1 Month-2 Month-3 Month-4 Month-5 Month-6 Month-8 Month-9
Month-20 Month-10 Month-11 Month-12 Month-13 Month-14 Month-15 Month-16 Month-17 Month-18 Month-19 Month-21 Month-22 Month-23 Month-24 Month-25 Month-26 Month-27 Month-28 Month-29 Month-30 Month-31
Pit 4 Pit 7 Pit 6 Pit 5 Pit 3 Pit 2 Pit 1
Table 16.3 shows the pit tonnage and grade for each zone at the optimized depth calculated in the previous section. The mining dilution is included in the total which is estimated at 12%.
Table 16.3 – Open Pits Statistics
Mineralized Waste PIT material Grade (g/t) Ounces Strip tonnage tonnage Pit 1 23,000 3.04 2,240 258,000 11.3 Pit 2 81,000 6.65 17,235 2,315,000 28.7 Pit 3 121,000 7.42 28,862 3,423,000 28.3 Pit 4 17,000 4.63 2,500 350,000 20.8 Pit 5 187,000 6.33 38,087 3,352,000 17.9 Pit 6 25,000 4.20 3,417 634,000 25.1 Pit 7 162,000 6.77 35,310 3,725,000 23.0 Total 616,000 6.45 127,651 14,058,000 22.8
Numbers may not add due to rounding.
The first pits will be mined as part of the pre-production period (Year 0) during the construction of the process plant. The mineralized material will be stockpiled until the start of the commissioning of the process plant.
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The open pit mining will be completed during Year 2 of the operation, and the total duration of the open pit operation is expected to be approximately 31 months. Figure 16.4 shows the tonnage mined per year and grade of mineralized material mined.
Figure 16.4 – Open Pit Sequence
16.5 Underground Mining
16.5.1 MINING METHOD AND CUT-OFF GRADE (COG)
The mineralized zone is a sub vertical vein type of mineralization (65-70°) from surface to approximately 120 m deep, and then flattens a little bit to about 50° going down. The width of mineralization can vary between 1 and 7 m. As described in Section 16.1, the host rock and mineralized rock have good strong properties.
Long hole stoping method is suggested to be used for this mineralized zone. Due to the shallow vein, 40 m will be used between levels. An extra 0.5 m will be added on the walls to take in long hole dilution into consideration.
The following parameters presented in Table 16.4 have been considered by DRA/Met-Chem to estimate the cut-off grade for the UG operations.
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Table 16.4 – UG Cut-Off Grade Parameters
Description Units Value
Mining Cost – UG US$/t (mined) 52 Processing Cost US$/t (milled) 41.5 G&A US$/t (milled) 14.4 Refining US$/oz 3 Sales Price US$/oz 1,300 Mill Recovery % 81 State Royalties % 5 Development Tax % 1 UG Dilution % 10
Using the economic parameters presented above, a cut-off grade of 3.7 g/t of gold was calculated for the UG operation. This cut-off grade was then used in the Stope Optimizer software (Deswik SO), and mineable economical shapes have been generated. A minimum mining with of 1.5 m was used. Figure 16.5 shows a longitudinal view of the possible stoping below the pit outline.
Knowing that the ground has strong properties, a double decking is considered to minimize the amount of development required. Figure 16.6 shows the longitudinal retreat approach with up and down holes. Once the block is mined out, a wedge of cemented rock fill (4% cement) will be dumped in the stope to create an artificial pillar between the stopes, and the rest of the stope will be filled with waste rock.
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Figure 16.5 – Longitudinal View of the MSO Results Per Zone
2300 m
Figure 16.6 – Longitudinal Retreat Mining
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16.5.2 MINE DESIGN
The mining area is spread over 2 km, several ramps (3 portals) will be used to ease material access and increase mining flexibility. The accesses will be designed for each stope grouping. This will give the flexibility to mine material zone independently and to give several faces of attack in the mining sequence. Moreover, it will ease the ventilation network / alternative egress and shorten the hauling distance for truck.
Figure 16.7 shows an isometric view of the Nabanga underground mine and Figure 16.8 shows a longitudinal view.
All ramps and accesses will be developed with a dimension of 5 m by 5.5 m. This size will allow the use of large equipment for hauling to surface such as 40-t truck. The level (material zone and ventilation access) will have a size of 5 m Height by 4.8 m Width. Large equipment is also planned to be used for mucking the production stopes (14 t LHD).
There will be four (4) ventilation raises (alternative egress) for the whole underground mine, mainly one for each stope grouping (ramp). The size will be 4 m x 3 m.
When the stoping area cannot justify lowering the ramp to recover them, a bench will be taken to recover a part or the whole identified block. This will reduce the overall mining cost while increasing mineralized material recovery.
The dilution of 0.5 m on each wall was taken in consideration in the MSO calculation. This dilution gives an average external dilution of 30% dilution and an internal dilution of 15%. It was assumed the same dilution in the mineralized material development on average.
The mining recovery of the production stopes is 95%. The scatter stopes, which are far from the main access or lonely stopes at depth that could not justify a ramp or access, are left behind.
16.5.3 MINE SERVICES
16.5.3.1 Ventilation – Alternative Egress
The fresh air requirement will be established using the Canadian homologation ventilation standard (CANMET). At this point, no calculation was done on the quantity; however, the ventilation network for the underground mining operation was taken in consideration in the mine design. No mine ventilation simulation was done.
Figure 16.9 shows the suggested ventilation network. Each material zone or area will be ventilated independently, which will ease the distribution. Doors/walls will be installed at the level raise access to control the required flow on the level. The alternative egress will be parallel to the raises.
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Main fans will be installed on surface (extractor) and booster fans will be used to bring the air to the working face. The intake air will come from the main ramps and the exhaust will be in the raise system.
16.5.3.2 Water Supply
The water required for the drilling operation will be taken of the river at the north east of the pit to start of the project. This water will be pumped into pond on surface when required. The water will be pumped to a tank above the portal to supply the underground mine with the gravity. The pond will be also used to collect water from the underground mine dewatering network.
16.5.3.3 Dewatering
Sumps will be excavated at each level entrance to collect the water from the level and the ramp above. There will be drain hole to connect each sump from level to level as there will be excavated. The water will be pumped out of the mine to a settling pond on surface. This water will be recirculated as water supply. At this point, there is not dimensioning of the pumping facilities as no hydrogeology study has been done, but a provision has been taken in the cost evaluation for dewatering. It has been assumed that the mining contractor will provide all the pumping equipment.
16.5.3.4 Electrical Distribution
Diesel generators will generate the electrical power for the Nabanga Project. A cable will bring the electrical power in the pit down to a substation and after, inside the ramp. There will be electrical substations at each level.
16.5.3.5 Compressed Air
A compressor will be installed at the pit bottom, close to the portal. One (1) 160 kW electrical compressor will service the mining area. The compressed air network in the mine will be used mainly for drilling and explosive loading.
16.5.3.6 Mine Communication
A leaky feeder system will be installed underground. This will allow an effective communication between all underground personal and surface installation.
16.5.3.7 Mine Facilities
The main facilities for the underground operation will include office, warehouse and maintenance facilities such as wash bay, garage. There will be a fuel station and an area to store used oil which will be collected and disposed by the fuel company. A cement batch plant will be installed by the contractor for cement needs such as backfill, concrete floor or shotcrete.
No surface layout for the facilities disposition was done at this stage.
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Figure 16.7 – Isometric View – Nabanga Underground Mine
Figure 16.8 – Longitudinal View – Nabanga Underground Mine
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Figure 16.9 – Ventilation Network
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16.5.4 OPERATION
All of the underground mining will be contracted as a turnkey project. The contractor will oversee all mining aspects including mobile, fixed equipment and the maintenance, mine services and facilities. The contractor will put in place safety practices and working procedures jointly with SEMAFO.
Development drilling will be done by jumbo drill. These units will also be used to drill holes for the appropriate ground support. Long hole drilling will be done with an electro-hydraulic rig for large scale production drilling. An ITH will drill a slot raise with the V30 and ventilation raise.
There is no plan in the design to have powder magazine underground. The powder magazine will be located on surface.
It is expected to use 14 t LHD to clean the development face as well as the production stopes.
The mineralized material haulage will be done by low profile underground truck. The truck capacity will be 40-t nominal. The material will be hauled from underground to the Run off Mine Pad on surface. While on surface, this truck will also be used to transport waste materiel at the bottom of the pit for backfill purposes. The waste development rock will be hauled out of the underground mine and temporarily stored at the bottom of the pit.
Waste rock will be used for backfill. The rock will be carried down the stope as there is no plan to have a backfill raise. A batch plant will be supplied by the contactor for cement. This will be required for the cemented rock fill wedge that is needed to increase stability and avoid leaving material pillar between the stope.
16.5.5 MINE PRODUCTION PLAN
The underground mine has been designed for a 1,000 t/d production. The mineralized material will come from the development and from the long hole stope. At the beginning of the mine, a large proportion of mineralized material production will come from development. This proportion will decrease as more level will be developed and give the sequence flexibility required in a longitudinal retreat mining method.
Pits 4 and 7 will be mined first and the underground excavation will start roughly 6 months after to give enough time to liberate the face for the portal. The mining will start from the bottom of the mine and progress towards the top, ending with the surface pillar recovery.
Figures 16.10 and 16.11 show the production plan and development plan of the Nabanga underground mine.
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Figure 16.10 – Underground Production Plan
Due to the nature of the mining method, most of the development necessary to access the different stoping areas will be done in the early years of the underground operation. From Years 2 to 4, an average of 2,680 m of development will be accomplished, then approximately 1,350 m in Year 5 and 260 m in Year 6, whereas the underground operation will last an additional 3 years (until Year 9).
Figure 16.11 – Underground Development Plan
Table 16.5 shows the statistics of the life of mine plan. It includes the open pit production at the beginning and the underground operation following, as well as the details of the material planned to be processed at the plant.
Figure 16.12 to Figure 16.19 illustrate the progression of the development and the mining of the underground operation from Years 2 to 9.
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Table 16.5 – Nabanga Life of Mine Plan Statistics
Description Unit Year 0 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Total
OP Waste Rock t 2,894,907 5,401,199 5,762,218 14,058,323
OP Mineralized Material t 88,067 308,251 219,657 615,974
OP Mineralized Material Grade g/t 6.20 6.33 6.70 6.45
UG Mineralized Material t 88,411 359,981 358,593 360,361 360,726 391,598 360,883 83,907 2,364,459
UG Mineralized Material Grade g/t 6.96 5.75 6.87 7.14 6.81 5.51 6.67 6.96 6.48
UG Development m 2,696 2,677 2,677 1,349 260 9,659
OP + UG Mineralized Material t 88,067 308,251 308,068 359,981 358,593 360,361 360,726 391,598 360,883 83,907 2,980,434
OP + UG Mineralized Material Grade g/t 6.20 6.33 6.78 5.75 6.87 7.14 6.81 5.51 6.67 6.96 6.47
Material Processed – Tonnes t 330,000 360,000 360,000 360,000 360,000 360,000 360,000 360,000 130,434 6.47
Material Processed – Grade g/t 6.30 6.69 5.79 6.83 7.13 6.82 5.55 6.54 6.81 6.47
Material Processed – Gold oz 66,882 77,459 66,958 79,077 82,497 78,967 64,294 75,714 570,768 Contained
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Figure 16.12 – Underground Development at Year 2
Figure 16.13 – Underground Development at Year 3
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Figure 16.14 – Underground Development at Year 4
Figure 16.15 – Underground Development at Year 5
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Figure 16.16 – Underground Development at Year 6
Figure 16.17 – Underground Development at Year 7
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Figure 16.18 – Underground Development at Year 8
Figure 16.19 – Underground Development at Year 9
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 107
17 RECOVERY METHODS
17.1 Nabanga Processing Plant
At the Nabanga processing plant, gold will be recovered using conventional sulphide flotation to produce a sulphide concentrate for intensive leaching, followed by conventional CIL cyanidation of the flotation tailings for additional gold recovery.
The processing facility or concentrator consists of:
• One stage crushing;
• SAG mill and ball mill grinding and classification;
• Conventional sulphide froth flotation;
• Intensive cyanidation of the flotation concentrate;
• CIL leaching;
• Pressure Zadra elution;
• Electro-winning and refining for gold doré recovery;
• Reagent and utilities.
The concentrator is designed to produce gold doré bars containing gold and silver from mineralized material containing a Life-of-Mine (LOM) average grade of 6.5 g/t gold and 3.0 g/t silver.
17.2 Design Criteria
Process development undertaken during the PEA aimed at establishing a conventional sulphide gold/silver processing facility considering the constructability, operability and maintainability of the processing facility.
The plant design is based on processing 360,000 dry tonnes per year of mineralized material, with a gold recovery of 92% This estimate is based upon bench scale metallurgical testing as described in Section 13. There is some variability in the mineralized materials tested and therefore; the results obtained which vary depending on mineralized material composition.
The concentrator is designed to operate 365 days per year with an average utilisation of 91.3% following the crushing circuit.
The concentrator throughput capacity is based upon an average throughput rate of 1,080 dry tonnes per day or nominal 49.3 dry tonnes of material per hour.
Table 17.1 lists the main process design criteria. The complete process design criteria are detailed in Appendix A.
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Table 17.1 – Design Criteria
Plant Capacity Parameter Units Value Material processing rate t/y 360,000 Material processing rate t/d 1,080 Feed moisture content % 5.0 Feed gold grade g/t 6.5 Feed copper grade % 0.25 Crusher operational utilisation % 68.5 Concentrator operational utilisation % 91.3 Material processing rate – t/h 49.3 Nominal Flotation gold recovery % 76 Intensive leach gold recovery % 95 CIL gold recovery % 90 Total gold recovery % 92 Life of Mine – Gold production troy ounces 571,000
17.3 Material Balance and Water Balance
The PEA level mass and water balance for the processing facility has been calculated based upon the flowsheet developed and the process design criteria established.
Table 17.2 gives a summary of the mass and water balances at an average throughput rate of 1,080 t/d.
The throughput and flows average rates are expressed in t/d and m³/d. One m³/d of water is equivalent to one metric t/d.
Table 17.2 – Concentrator Summarised Process Mass Balance
Mass Entering Concentrator System Mass Exiting Concentrator System Dry Total Dry Total Water Water Streams Solids Mass Streams Solids Mass (m³/d) (m³/d) (t/d) (t/d) (t/d) (t/d) ROM material to To tailings 1,080.0 68.9 1,148.9 1,080.0 970.2 2,050.2 Concentrator storage facility Raw water — 222.1 222.1 Reclaim water — 679.2 679.2 Total Entering 1,080.0 970.2 2,050.2 Total Exiting 1,080.0 970.2 2,050.2
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The detailed process plant mass balance is presented in Appendix B.
Figure 17.1 shows the processing plant water balance. The tailings storage facility is not considered part of the processing facility water system and is added for illustrative purposes only.
Figure 17.1 – Concentrator Water Balance
17.4 Flow Sheets and Process Description
A simplified flow sheet is presented in Figure 17.2 which is indicative of the selected process.
The concentrator has nine (9) distinct sections with the flow sheet presented showing seven (7) of these sections.
Grinding media, reagents and utilities are not shown in the simplified schematic flow sheet below.
The following sections are:
• Crushing;
• Grinding;
• Sulphide flotation;
• Concentrate intensive cyanide leaching and CIL;
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• Carbon elution and regeneration and electro-winning and refining;
• Grinding media and reagents;
• Utilities and services.
The crushing facility will operate independently from the remaining concentrator sections.
The simplified flow sheet presented is simplistic, but representative of the process.
17.5 Crushing
The crusher is expected to receive mineralized material, containing 6.5 g/t gold and 3.0 g/t silver with a moisture content of 5%. ROM material is introduced into the primary crusher feed bin after passing through a static grizzly. Ore enters the jaw crusher via a vibrating feeder at the base of the hopper. Jaw crusher product is transported using a conveyor for to the mill feed stockpile.
The jaw crusher discharges material with a particle size distribution of 80% passing (P₈₀) of 93 mm.
The parameters are summarised in Table 17.3.
Table 17.3 – Crusher Parameters
Parameter Unit Value Crushing rate t/h 65.7 Crusher operating time h/day 16.4 SMC test mill work index (Mic) kWh/t 8.35 Primary Jaw Crusher kW 75
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Figure 17.2 – Simplified Flow Sheet of Nabanga Mineral Processing Facility
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 112
17.6 Grinding
One-stage crushing followed by SAG mill and ball milling was selected over conventional three-stage crushing and rod mill and ball mill grinding, which is usually more expensive and cumbersome. The low material competency of the material tested was conducive to SAG milling due to the low energy requirement necessary for material breakage.
Crushed material is drawn from the mill feed stockpile utilizing apron feeders located in the tunnel underneath the stockpile. Apron feeders discharge the material onto the SAG mill feed conveyor. The SAG mill operates in closed circuit with a SAG mill discharge screen. The ball mill operates in closed circuit classification with a hydro-cyclone cluster. Both SAG and ball mill are equipped with a trommel screens. The trommel screen protects the mill discharge pumps from steel chunks and large pebbles.
The SAG screen undersize and the ball mill discharge flow into the secondary cyclone feed pump box. The material is pumped to the cyclone cluster for classification. Cyclone underflow or coarse material reports back to ball mill whilst the cyclone overflow or fine material flows by gravity to flotation circuit feed.
Grinding circuit pulp densities, mill speed and ball mill load are controlled automatically to ensure that the grinding circuit produces a consistent product grind size P₈₀ of 53 microns.
The parameters are summarized in Table 17.4.
Table 17.4 – Grinding Parameters
Parameter Unit Value Processing rate t/h 49.3 SAG Mill dimensions m 5.0 D × 2.0 EGL SAG Mill power (VFD) kW 800 Ball Mill dimensions m 3.6 D × 6.3 EGL Ball Mill power (VFD) kW 1200 Cyclone diameter mm 250 Cyclones number 6 Grinding circuit P₈₀ mm 0.053
17.7 Sulphide Flotation
The flotation circuit consists of rougher flotation, concentrate regrind and classification followed by cleaning flotation stages.
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The rougher flotation section consists of 8 tank cells that float the sulphide particles containing gold to a rougher concentrate. The rougher flotation cells provide 30 minutes of retention time. Rougher flotation tailings are pumped to the CIL leach circuit.
Rougher sulphide concentrate is recovered using copper sulfate as an activator, potassium amyl xanthate (PAX) as a collector and MIBC as a frother.
Rougher concentrate is pumped to a vertical regrind mill operating in closed circuit with a cluster of cyclones. The target cyclone overflow product size is a P₈₀ of 20 microns.
There are three (3) sequential cleaning flotation stages. First cleaner concentrate is pumped to the feed of the second cleaners with second cleaner tailings returned to the feed of the first cleaner bank. Second cleaner concentrate reports to the feed of the third cleaner bank. The third cleaner concentrate is final concentrate with third cleaner tailings reporting back to the feed of the second cleaner bank. Total cleaner bank flotation retention time is 30 minutes. Final concentrate is pumped to the concentrate thickener for dewatering. First cleaner tailings are combined with the rougher flotation tailings stream and pumped to the carbon-in-leach circuit and for additional gold recovery. See a summary of flotation parameters in Table 17.5.
Table 17.5 – Flotation Parameters
Parameter Unit Value Roughers retention time min 30 Cleaners retention time min 30 Regrind mill kW 140 Regrind product size P₈₀ mm 0.020
The regrind mill size was based on experience with sulphide concentrate regrinding. Additional tests on rougher concentrate material targeting the specific regrind mill energy consumption will be required during the prefeasibility phase to confirm the size of the regrind mill.
17.8 Intensive Cyanidation and Carbon-in-Leach (CIL) Cyanidation
Sulphide concentrate is thickened to 63% solids in the concentrate thickener. Excess solution is returned to the process water tank.
The thickener underflow is pumped to the intensive cyanide leach reactor feed bin.
The intensive cyanide leach reactor operates as a batch process and uses high cyanide concentration solutions and mixing to dissolve the free gold particles recovered from the flotation circuit. The leached gold in solution recovered from the leach reactor is pumped to the electro- winning cells for gold recovery. The intensive leach tailings are rinsed and pumped to final tailings.
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The flotation tailings are pumped to the cyanidation leach circuit. The circuit feed is screened to remove trash material. Screen undersize flows by gravity to the pre-leach thickener with the pulp thickened to adjust the percent solids in preparation for cyanidation.
Pre-leach thickener underflow is pumped to the pre-oxidation leach tank. The pre oxidation discharge then flows by gravity to the first of 6 × CIL tanks for the dissolution of gold and silver. The cyanidation leach process is accomplished using oxygen and cyanide in an alkaline leach environment. Pre- oxidation is a common first step to elevate the pulp dissolved oxygen levels in the slurry prior to cyanidation which improves the rate of gold dissolution and reduces the consumption of cyanide and lime.
Activated carbon is added to the last leach tank and pumped being pumped forward from the last tank to first tank in counter current sequence such that all of leach tanks contain carbon. This is the Carbon-In-Leach (CIL) scenario. The activated carbon adsorbs precious gold and silver from the slurry. Inter-stage pumping screens pump the slurry forward to the next tank. The screens prevent carbon moving along with the slurry back to the next tank. Fresh carbon containing no gold or barren eluted carbon containing low gold grades are added to the last tank in series. The loaded carbon in the first tank contains the highest loaded carbon gold grades and is pumped batch wise to the loaded carbon screen.
Carbon-In-Leach (CIL) was selected as the recovery option over leaching followed by Carbon-In- Pulp (CIP) recovery. The leach and adsorption test results were not available when the circuit selection and engineering development took place. The CIL option prevents preg-robbing of gold from solution if the material contains naturally occurring active organic carbon or sulphide that will consume cyanide. In the CIL option activated carbon is added during the leaching stages which helps to minimise this preg-robbing phenomenon.
Flocculent is added to improve thickener settling rates. Final sulphide concentrate grades will be in the order of 87 g/t gold. Approximately 80% of the gold recovered from mill feed is from the flotation section.
The main leach parameters are summarised below in Table 17.6.
Table 17.6 – Leaching Parameters
Parameter Unit Value Leach retention time h 48 Leach tank dimensions mׇ ³ 500 Leach tank number 6 Carbon concentration g/L 15
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17.9 Elution and Carbon Regeneration and Electro-winning and Refining
Loaded carbon is cleaned of the pulp particles by wet screening on a vibrating screen and then flows by gravity to the elution area for gold desorption and recovery. Carbon is first acid washed using dilute hydrochloric acid to remove inorganic salts, water rinsed and then eluted or stripped of the contained gold in an elution column using a caustic cyanide solution at elevated temperature and pressure. The gold cyanide solution recovered is referred to as pregnant solution and reports to the gold room area for electro-winning.
At the completion of the elution cycle the stripped or barren carbon is then water rinsed and then fed to a rotary kiln for carbon regeneration or reactivation by heating to 700ºC in a moist environment. This regeneration step ensures that the carbon remains active with respect to gold recovery by removing organic material from the carbon particles prior to being reintroduced to the last CIL tank.
The intensive leach reactor leach solution is pumped directly into the pregnant solution tank. Pregnant solution reports to the electro-winning cells located in the refinery. Gold and silver from solution are deposited onto the electro-winning cell stainless steel cathodes. The deposited product is removed from the cathodes using a high pressure washer and then dried, mixed with fluxes (sodium nitrate, fluorspar, borax, silica sand and soda ash) and smelted in an electric induction furnace.
Flux is required to remove impurities in the gold sludge from the molten metal and is removed as slag. The doré (gold and silver) bars are poured into 1000 troy ounce moulds for shipment to the external gold refinery.
17.10 Grinding Media, Concentrator Reagents and Refinery Chemicals
The grinding balls will be stored in the drums that they arrived in. Fresh water will be used for reagent mixing. Reagent preparation is conducted in a separate area as there is a relatively high consumption of sodium cyanide. MIBC is supplied in liquid form and is ready for immediate use.
17.10.1 GRINDING MEDIA
• The SAG mill uses 125 mm forged steel grinding balls;
• The ball mill uses 75 mm forged steel grinding balls;
• The re-grind mill will use 25 mm steel grinding balls.
17.10.2 FLOTATION REAGENTS
• Copper sulfate (CuSO₄) is a sulphide activator to enhance the sulphide recovery in the rougher flotation circuit;
• Potassium amyl xanthate (PAX) is used as a collector to recover sulphides. The sulphide minerals bond the collector;
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• Methyl isobutyl carbinol (MIBC) is used a froth stabilizer.
17.10.3 LEACH CHEMICALS
• Lime (CaO) is used as pH modifier in the cyanide leach circuits. It is sometimes in the thickener to support settling;
• Sodium Cyanide (NaCN) issued for the dissolution of gold and silver from the pulp to form soluble cyanide complexes and in the stripping circuit;
• Activated Carbon (C) is used to adsorb dissolved gold and silver and transport them selectively to the stripping area for purification;
• Lead nitrate (PbNO₃) is used to enhance gold dissolution in cyanide circuits.
17.10.4 CARBON SYSTEM CHEMICALS
• Hydrochloric Acid (HCl) is used to remove inorganic compounds from the activated carbon prior to stripping;
• Caustic Soda (NaOH) is used during stripping to create a high pH environment and enhance the solution conductivity.
17.10.5 REFINERY FLUXES
• Sodium nitrate (NaNO₃) is a strong oxidiser used to remove sulfur compounds from the melt;
• Fluorspar (CaF₂) is used remove aluminium impurities and increases the fluidity of the melt;
;٠8H₂O) is used to dissolve (base) metals in the melt[Borax (Na₂[B₄O₅(OH)₄ •
• Silica (SiO₂) is used to adsorb dissolved (base) metals and form silicates thus transporting the unwanted impurities to the slag;
• Soda Ash (NaCO₃) is used to decrease slag viscosity.
17.10.6 OTHER CHEMICALS
Flocculant is used to aid with the settling of fine particles in the thickener. This will be in the form of a non-toxic, polyacrylamide polymer.
Antiscalant is used to prevent the pipes plugging up with scale.
Table 17.7 lists the estimated main consumables consumption rates.
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Table 17.7 – Grinding Media and Reagent Consumption
Parameter Unit Value
Primary mill grinding balls g/kWh 65 Regrind mill grinding media g/kWh 8 Quick lime g/tonne 2,391 Potassium amyl xanthate g/tonne 160 Methyl isobutyl carbinol g/tonne 120 Copper sulfate g/tonne 20 Sodium cyanide g/tonne 1,394 Activated carbon g/tonne 41 Lead nitrate g/tonne 99 Flocculant g/tonne 31 Antiscalant g/tonne 5
17.11 Water and Air Services
Water and air are part of the essential services required for concentrator operations.
17.11.1 WATER SUPPLY:
There are three (3) types of water used in the processing facility.
• Process Water: During normal operation about 80% of all process water returns as reclaimed water from the Tailings Management Facility and as return water from the Backfill Plant into the Process Water Tank.
• Fresh Water: The Fresh Water Well will be the main water source of all fresh water. The main use of fresh water will be for Gland Water and reagent mixing.
• Potable water: The concentrator will get potable water from external sources.
17.11.2 AIR SUPPLY:
• High Pressure Air: Two (2) air compressors will supply the concentrator plant with compressed air at 760 kPa.
• Low Pressure Air: Two (2) air blowers will supply low pressure air to the flotation circuit at 60 kPa.
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17.12 Overall Gold Recovery
The overall gold recovery has been derived from the metallurgical test work carried out under the supervision of Orbis. Based on these results, the overall gold recovery for the Nabanaga mineralized material has been estimated at 92%. Table 17.8 depicts the parameters used for the calculation.
Table 17.8 – Gold Recovery for the Nabanga Mineralized Material
Gold Recovery Material (%) Fresh Material Sulphide flotation to Intense Leach Reactor 95 Flotation tailings to standard CIL 80 Oxide Material 90 Overall Gold Recovery (weighted average) 92
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18 PROJECT INFRASTRUCTURE
This section summarizes the infrastructure, buildings and other facilities required for the operation of the Nabanga Project. It consists of the following infrastructure, which are all located on site:
• Roads;
• Power Supply;
• Control System;
• Communication System;
• Tailings storage facility (TSF);
• Site Buildings.
The overall site layout and access is shown in the Figure 18.1, and the following Figure 18.2 illustrates the details of the plant layout.
No detailed topographic survey has been provided to DRA/Met-Chem, and only a topographic surface made from exploration collar location has been used.
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Figure 18.1 – General Site Layout
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Figure 18.2 – Process Plant Layout
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18.1 Roads
18.1.1 MAIN ACCESS ROAD
The existing (Route Nationale) RN18 connects the town of Fada-N’Gourma to the border between Burkina Faso and Togo. Starting approximately 80 km South of Fada-N’Gourma, a new access road connecting the RN18 to the Nabanga Process Site will be developed over a distance of about 11 km.
18.1.2 SITE ROADS
Site and service roads will be 6 m wide, except for mine roads. They will provide access to:
• Process facility from the Camp;
• Administration, plant and mine offices;
• Warehouse;
• Tailings storage facilities; and
• Power house.
18.1.3 MINE ROADS
Provision for a network of 1 km of haulage roads has been made. Mine roads will be 15 m wide and will provide access to:
• ROM stockpiling area;
• Pit entrance of UG Decline; and
• Mine garage.
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18.2 Power Supply
18.2.1 POWER DEMAND
The power demand of the SEMAFO site was determined to be 5.8 MW based on the estimated connected loads, running loads and running power for the process operation. Table 18.1 presents the power demand breakdown by sectors.
Table 18.1 – Estimated Total Power Demand Consumption
Mechanical Power Demand Area Operating Power kWelec kWmec All process areas equipment excepts: 2,600 2162 SAG-2200 SAG mill 800 826 BAM-2400 Ball mill 1,200 1239 PROCESS ONLY 4,600 4,227
Administration Building 162 Laboratory 32 Mechanical Shop 46 Plant Office 50 Camp 100 people 73 Other loads (including process lighting) 127 SERVICES ONLY 490
UG Mine – Development Production 235 204 UG Mine – Vent Fans 645 673 UG Mine Workshop / Office / Store 5 UG Mine Main Pumps 90 83 UG Mine Auxiliary Pumps 65 60 UG Mine ONLY 1,035 1,055
Total 5,742 Note: The power demand was calculated using an average efficiency factor, load factor and diversity factor.
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18.2.2 POWER PLANT AND DISTRIBUTION – RETICULATION NETWORK
The site will be supplied at 6.6 kV, 3 phase, 50 Hz from a Power Plant installed adjacent to the concentrator. The power plant consists in 5 generators using LFO fuel, with operating mode N+1 (four in operation, one in stand-by, maintenance or reparation). Each engine provides 1,600 kWe @ 0.8 PF, 2000 kVA, 6.6 kV, 50 Hz, 1500 RPM. The choice of the LFO fuel is justified by the Mine life versus the capital investment for HFO fuel. Each DG unit shall be installed in a container. The net
output installed power for 5 engines is 8,000 kWe and the net output operation power for 4 engines
is 6,400 kWe (power at the busbars of the of the MV Switchgear, 6.6 kV, 3 phases, 50 Hz, part od Power Plant package).
The main components of the Power Plant are:
• 5 Main DG units;
• Control Unit;
• Batteries c/w Battery Charger;
• Control and Synchronisation Panel for 5 units; and
• 6.6 kV Switchgear (5 incomings from DG units and 3 feeders).
18.2.3 RETICULATION NETWORK
The reticulation network consists in a 6.6 kV and 0.4 kV buried cable network and 6.6 kV pole lines. The 6.6 kV buried cable network starts from the Main 6.6 kV MV Switchgear output feeders and is going to the power step-down transformers 6.6 – 0.4 kV installed for each Electrical Room of the process plant.
The 6.6 kV pole line supplies, via step-down transformers 6.6 – 0.4 kV and low voltage buried cables network, the remote loads of the site as: Gate House, Administration Building, Laboratory, Mechanical Shop, Plant Office, Camp, Mine Office.
18.2.4 MAIN ELECTRICAL EQUIPMENT
The electrical equipment, the motors, the cables and the other electrical materials shall be designed, constructed, tested and installed as per the International Electrotechnical Commission (IEC) standards. The main electrical equipment is related to the process areas and is installed in Electrical Rooms.
ER-100 is the dedicated prefabricated Electrical Room for Crusher. The electrical room comprises a 1 MVA, 6.6 – 0.4 kV dry transformer and the low voltage MCC: 100-LVMCC-01 equipped with starters and VFD-s to control the equipment in this area. In addition, the auxiliary transformers (lighting, services) and the Control Panels for the control and instrumentation are also included.
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ER-200 is the dedicated prefabricated Electrical Rooms for Concentrator; two transformers, liquid filled, 2.5 MVA each, 6.6 -0.4 kV are installed outside, in its vicinity. ER-200 comprises the equipment necessary for the concentrator. Besides the 400 V switchgears and MCC-s, there are also the two 6.6 kV VFD-s that are suppling the mills: 800 kW for SAG mill and 1200 kW for Ball mill.
ER-990 is the dedicated prefabricated Electrical Room for the UG Mine and comprises step-down dry transformers 6.6 – 1 kV for the mining equipment and dry transformers 6.6 – 0.4 kV to supply vent fans, main and auxiliary pumps, workshop office.
18.3 Control System
18.3.1 AUTOMATION PROCESS NETWORK
The SEMAFO Process Control System (“PCS”) will be based on a redundant Ethernet network in a ring type topology. The network links all the main automation equipment, such as Supervisory Control and Data Acquisition (“SCADA”) System, Historian, Human Machine Interface (“HMI”) and Process Control System processors.
The proposed network includes fibre optic linking of the following main areas of the SEMAFO plant:
• Central Control Room
• Power Plant
• Electrical Room ER-100 (Crushing Area);
• Electrical Room ER-200 (Concentrator Area);
• Electrical Room ER-990 (Mining Area);
• Gate House,
• Administration Building,
• Laboratory,
• Mechanical Shop,
• Plant Office,
• Camp,
• Mine Office.
Network automation communication services are:
• SCADA stations located in the Central Control Room;
• Process Control System processors inter-communication;
• PCS/Remote Input/Output (“I/O”) communication;
• PCS direct interface to the Motor Control Centers (“MCC“s);
• IEC61850 interface to the power distribution equipment;
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• Field device communication including communication with 3rd parties Programmable Logic Controller (“PLC”) supplied with mechanical equipment;
• Camera system installed in the plant for process control viewing purposes.
18.3.2 PROCESS CONTROL SYSTEM
The process control system will be of PLC type. A PCS system will be supplied to control each strategic areas of the plant with remote I/O racks located generally in the Electrical Rooms.
The main processors will be included to control the following sectors: crushing, concentrator UG Mine and the remote loads connected to the 6.6 kV pole line. The major equipment like the Mills could come with their own PLC and with a Local Control Panel. The 400 V MCC-s should be equipped with ″intelligent″ protection relay able to communicate. The protection relays shall be equipped with Ethernet ports. Near each motor shall be installed a Local Control Station. The central SCADA system has the capacity to control and supervise all the remote PCS equipment. In a communication outage situation, the critical equipment will be controlled locally.
18.3.3 WIRING AND JUNCTION BOXES
All the field instruments and switches will be wired to the PCS through junction boxes up to remote I/O racks situated in the various electrical rooms. The wiring system will include field junction boxes for instrument power supply, for digital signals and for analog signals. The motor thermistor signals will be wired directly to the related motor protection relays while equipment RTD signals will be connected directly to the PCS remote I/O-s. The junction boxes will be located and installed in all process areas of the plant. The junction boxes will be wired to the PCS I/O racks via multi-conductor cable.
18.3.4 SCADA
The SCADA system will be based on client/server technology and will include:
• Two (2) SCADA servers for redundancy;
• One (1) historian server;
• Two (2) HMI operator stations; and
• One (1) engineering station.
The system will be installed in the Central Control Room.
18.3.5 SCADA AND PLC POWER SOURCES
In case of plant power outage, the PCS, switches, main servers, phone system, and security systems will be fed by Uninterruptible Power Supply (“UPS”). UPS status will be monitored.
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18.3.6 REDUNDANCY
For the automation network, the redundant ring topology design insures a second route in case of a communication outage on one (1) segment.
18.3.7 PROCESS ANALOG INSTRUMENTS
Traditional 4-20 mA loop cabling with enabled HART protocol will be used as base solution. When available /requested, process analog instruments will support the Profibus protocol and they will be wired to Profibus Spur field boxes to the process controller by ProfibusPA fieldbus cables.
18.4 Communication System (Local and External)
18.4.1 TELECOMMUNICATION LOCAL SYSTEM
The telecommunication system will be based on Ethernet links throughout the plant buildings and administrative buildings following generally the electrical reticulation network (buried and/or installed on the pole lines).
Single-mode fiber optic backbone will be deployed through the plant to accommodate both automation and corporate services on the same fiber cable on different fiber.
18.4.2 TELECOMMUNICATION AND MOBILE RADIO SYSTEMS
The telecom service includes the tower located in a high elevation zone of the plant; it will be supplied by a third-party provider and will communicate with the Process plant communication interface.
The telecommunication systems will include:
• IP Phones;
• Process and Security Camera System;
• Fire Detection System;
• Access Control System (gate, door);
• Mobile Radio System.
The mobile radio system will be provided for the construction phase and the operation of the mine and plant site.
18.4.3 TELECOMMUNICATION SERVICES
The site will be connected to a local Internet Service Provider (“ISP”).
A backup system will use a cellular modem or satellite technology.
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The IP phone system will be connected to an Internet Telephone Service Provider (“ITSP”).
18.4.4 TELECOMMUNICATIONS DISTRIBUTION
During the construction phase, all communication services, such as Internet and phone, will be distributed via Wi-Fi, Wimax and Microwave point-to-point radios to reach all areas of the plant site.
All mine trucks and pick-ups will be equipped with a Wimax/Wi-Fi antenna that shall also act as a Wi-Fi local access point.
The telecommunication distribution will be through the plant fiber optic network covering the crushing and concentrator areas, UG Mine and Gate House, Administration Building, Laboratory, Mechanical Shop, Plant Office, Camp and Mine Office. If necessary, wireless communication will be provided for the other auxiliary outside of the plant.
18.4.5 CORPORATE NETWORK
The automation Ethernet backbone network, in a ring type topology described in the previous section will be used for the automation, the camera and security video, the IP phone system and the corporate network applications.
All the major network equipment will be located in dedicated server rooms located in the administrative office, the telecom shelter, the control room and electrical rooms.
Corporate services are:
• Wired/Wireless Phones and System Server;
• Process and Security Camera System;
• Access Control System (gate, door);
• Fire Detection.
18.4.6 CAMERA SYSTEM
A camera system, with recorder and a viewer, will be installed in the main gate office. Aside from the gate cameras, other cameras will be installed in the plant for process control purposes. One (1) viewing station will be installed in each control room for process control purposes.
18.5 Tailings Storage Facility
A PEA level assessment of tailings disposal requirements was performed. The assessment aimed to estimate the quantities of materials required for the construction of confinement dykes for the proposed tailings impoundment facility.
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The design considers the transfer of free water from the tailings storage facility to the plant to be used in processing. Table 18.2 summarizes the process information used for design of the Tailings Storage Facility (TSF).
The TSF will be located close to the processing plant to reduce the pumping costs of the tailings from the plant to the TSF. It is estimated that 3.0 Mt of tailings (approximately 2.0 Mm³ at 1,500 kg/m³) will be produced over the life of the mine.
Table 18.2 – Tailings Design Basis
Description Unit Value Life of Mine (LOM) Year 8 Tailings tonnage per Year kt/y of dry solids 360 Total TSF volume required Mm3 2.0
Additional tailings and site characterisation should be undertaken before a final selection is completed to confirm storage capacities and dam volume requirements.
18.6 Site Buildings
In addition to the concentrator building that will house the processing equipment, the site will include the following:
• Administration offices and mine offices;
• An accommodation camp, and cafeteria;
• A spare-parts warehouse;
• A metallurgical laboratory; and
• A security gatehouse.
All these buildings have been considered in the Capex, but no formal detailed design have been done at this stage of the Project.
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19 MARKET STUDIES AND CONTRACTS
SEMAFO is an existing producer and seller of gold. Historical data for gold prices therefore exists and as such no formal market studies for this Technical Report have been undertaken.
A long-term gold price projection of $1,300 per ounce for this Technical Report is considered reasonable on the basis of actual gold prices which reflect SEMAFO conservative outlook of the future market for gold.
19.1 Contracts
There are no material contracts or agreements in place as of the effective date of this Technical Report. SEMAFO has not hedged, nor committed any of its production pursuant to an off-take agreement.
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20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT
The Nabanga Project is administratively part of the Soudougui Department in the province of Koulpélogo in the Center-East Region of Burkina Faso, close to the limit with the East Region.
The project site is located approximately 250 km south-east of Ouagadougou, the country’s capital. Access to Nabanga is by means of (Route Nationale) RN04, an all-weather paved road from Ouagadougou, the capital of Burkina Faso, through Fada N’Gourma. From there, travel is via RN18, an all-weather paved road to within approximately 15 km of the Nabanga Project. An unsealed dirt road, which crosses the Kompienga River, is then used to access the Nabanga property approximately 15 km to the west of RN18, although other similar dirt access roads can be used to access other parts of the property.
As part of the environmental approval process in Burkina Faso, SEMAFO will carry out an Environmental and Social Impact Assessment (“ESIA”) for the Nabanga Project. Currently, little has been done to collect environmental and social data in the Project area since the Project is only at an early stage of study.
This Project being still at an early stage, no formal process for the Environmental and Social Impact Assessment (“ESIA”) has been started at this stage.
The typical approach that will be developed by SEMAFO during the future ESIA will take into consideration the social and environmental concerns of all interested parties. These concerns will be integrated into the initial stages of the project design. This approach aims at maximising the project’s integration into the environment and minimising its negative impacts to increase the environmental and social acceptability of the Project.
In addition, this approach will allow a better understanding of the social aspects arising from the resettlement of households due to the presence of the mine facilities.
DRA/Met-Chem has performed a review of existing information and desktop reports provided by SEMAFO. DRA/Met-Chem has prepared a summary of relevant environmental and social issues, which is presented in this Section.
20.1 Legal Framework and Permits to Obtain
The legal framework for the construction and operation of mining facilities in Burkina Faso includes national and international policies, regulations, and guidelines. The design and environmental management of the Project’s facilities and activities will be performed in accordance with this legal framework.
Burkina Faso has a regulatory framework for environmental and social management. The relevant Burkina Faso policies, laws and regulations pertaining to mine development were taken into account
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for the ESIA. Furthermore, SEMAFO will meet international standards taking into account relevant issues identified in the International Finance Corporation ("IFC") Performance Standards.
20.1.1 POLICIES AND STRATEGIES FOR ENVIRONMENTAL PROTECTION
Since the early 1990s, Burkina Faso has developed numerous policies and strategies for the management of their natural resources. A declaration of Mining Policy was formulated in 1995 that highlighted the importance of the private sector as an engine of economic development. Other policies on environmental protection include the following:
• Strategy for Accelerated Growth and Sustainable Development (SCADD: Stratégie de Croissance accélérée et de Développement durable);
• Government Program for Emerging and Sustainable Development (PAGEDD: Programme d’Action du Gouvernement pour l’Émergence et le Développement durable, 2011-2015);
• Rural Development Strategy (SDR: Stratégie de développement rural 2015);
• National Policy on Environmental Matters (PNE: Politique nationale en Matière d’Environnement);
• Environmental Plan for Sustainable Development Program (PEDD: Plan d’Environnement pour le développement durable);
• National Policy on Rural Land (PNSFMR: Politique Nationale de Sécurisation Foncière en Milieu Rural); and
• National Action Program for Adaptation to Climate Variability and Change (PANA: Programme d’action national d’adaptation à la variabilité et aux changements climatiques).
20.1.2 LEGAL FRAMEWORK
The Burkina Faso legal framework with respect to environmental and social aspects related to economic activities is supported by a number of laws and decrees:
• Environmental Code (Code de l’environnement).
• Mining Code (Code minier).
• Forest Code (Code forestier).
• Public Health Code (Code de la santé publique).
• General Local Authorities Code (Code général des collectivités territoriales).
• Act on Rural Land Tenure (Régime foncier rural).
• Act on Agrarian and Land Reorganization (Reorganisation agraire et fonciere).
• Law on Water Management (Loi d'orientation relative à la gestion de l'eau).
• Act on Pastoralism (Loi d’orientation relative au pastoralisme).
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Other relevant regulations include:
• Decree No. 2015-1205/PRES-TRANS/PM/MEF/MARHASA/MS/MRA/MICA/MME/MIDT/ MATD dated October 28, 2015 on setting standards for discharges of used waters.
• Decree No. 2015-1200/PRES-TRANS/PM/MERH/MME/MICA/MS/MIDT/MCT dated October 28, 2015 on the terms and conditions of environmental audit.
• Decree No. 2015-1187/PRES-TRANS/PM/MERH/MATD/MME/MS/MARHASA/MRA/MHU/ MIDT/MCT dated October 22, 2015 on conditions and procedures for carrying out and validating the strategic environmental assessment as well as the environmental and social impact study or notice.
• Decree No. 2007-853/PRES/PM/MCE/MECV/MATD dated 26 December 2007 on specific environmental regulations for the exercise of mining in Burkina Faso.
• Decree No. 2006-590/PRES/PM/MAHRH/MECV/MRA dated 6 December 2006 on the protection of aquatic ecosystems.
• Decree No. 2006-588/PRES/PM/MAHRH/MECV/MPAD/MFB/MS dated 6 December 2006 determining the perimeters of protection for water bodies and streams.
• Decree No. 2001-342/PRES/PM/MEE dated 17 July 2001 on the scope, content and procedure for Environmental Impact Assessment Study and Environmental Impact Instruction.
• Decree No. 2001-185/PRES/PM/MEE dated 7 May 2001 on setting standards for discharges of pollutants into the air, water, and soil.
20.1.3 MINING CODE
The Mining Code (Law N°036-2015/CNT pertaining to the Mining Code of Burkina Faso) is administered by the Ministère des Mines et des carrières (“MMC”) and provides the legal framework for the mining industry in the country. The state owns title to all mineral rights and these rights are acquired through a map-based system by direct application to the MMC.
The Mining Code guarantees a stable fiscal regime for the life of any mine developed. It also guarantees stabilization of the financial and customs regulations and rates during the operational period to reflect the rates in place at the date of signing. The Mining Code also states that no new taxes can be imposed with the exception of mining duties, taxes and royalties. However, the title holder can also benefit from any reductions of tax rates during the life of the operating license.
There are three (3) types of mining permits and three (3) types of authorizations according to the Mining Code:
Mining Permits
• Research Permit.
• Industrial Operating Permit.
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• Semi Mechanized Small-Scale Operating Permit.
Authorizations:
• Prospecting Authorization.
• Traditional Artisanal Mining Authorization.
• Quarrying Authorization.
Adopted in June 2015 by the National Transitional Council, the new mining code complies with Economic Community of West African States (ECOWAS) and the community mining policies of the West African Economic and Monetary Union (UEMOA from its French designation, Union économique et monétaire ouest-africaine) by integrating a number of provisions for a better contribution of mining projects to the country's economic development. In particular, the new Mining Code establishes the creation of a mining fund for local development to fund municipal and regional development plans, to which the mining license holders will contribute up to 1% of their monthly turnover or the value of extracted products during the month. Like the preceding Mining code, the new code requires mining projects to conduct an ESIA and to meet the Environmental Code requirements.
Other relevant changes in the new mining code include: the removal of the mining agreement for the research phase; application of a specific transaction tax on mining titles; the possibility of suspending or removing mining titles or authorizations without notice when required by public order. The new mining code provides for the adoption of decrees relating to the implementation of various provisions of this code. No such decree has been adopted up to and including the date of this Report.
20.1.4 INSTITUTIONAL FRAMEWORK
The main institutional stakeholders in the environment include:
• Ministry of Environment, the Green Economy and Climate Change (MEEVCC: Ministère de l’Environnement, de l’économie verte et du changement climatique).
• Ministry of Energy, Mines and quarries (MME: Ministère de l’Énergie, des Mines et des Carrières).
• Office of Mines and Geology (BUMIGEB: Bureau des Mines et de la Géologie du Burkina).
• Chamber of Mines of Burkina Faso (CMB: Chambre des mines du Burkina).
• National Commission of Mines (Commission nationale des Mines).
• National Council for the Environment and Sustainable Development (CONEDD: Conseil national pour l’Environnement et le Développement durable).
• National Bureau of Environmental Assessment (BUNEE: Bureau national des Évaluations Environnementales): this organization is part of MEEVCC and has the mandate to promote, regulate, and manage the environmental assessment process of the country. BUNEE holds
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sessions to review the terms of reference submitted by the project promoter. It formulates an opinion on the admissibility of studies and makes recommendations to MEEVCC on the environmental acceptability of projects.
• Technical Committee on Environmental Assessments (COTEVE: Comité technique sur les Évaluations Environnementales): this organization was created by Decree No. 2006- 025/MECV/CAB in 19 May 2006 establishing the powers, composition, and functioning of COTEVE. COTEVE is the technical and scientific framework to examine and analyse research reports and notices of environmental impacts presented by the project promoters to MEEVCC.
Other Ministries and Departments involved:
• Department of Infrastructures (Ministère des Infrastructures).
• Ministry of Territorial Administration, Decentralization and Homeland Security (Ministère de l'Administration territoriale, de la Décentralisation et de la Sécurité).
• Department of Health (Ministère de la Santé).
• Department of Agriculture and Food Security (Ministère de l’Agriculture et de la Sécurité Alimentaire).
• Department of, Hydraulic Resources and Sanitation (Ministère des Ressources Hydrauliques et de l’Assainissement).
• Department of Animal and Fishery Resources (Ministère des Ressources Animales et Halieutiques).
• Ministry of Women, National Solidarity and Family (MFSNF: Ministère de la Femme, de la Solidarité Nationale et de la Famille).
20.1.5 PERMITS TO OBTAIN
The Project will trigger a range of regulatory requirements and processes, which will require the application for, receipt of, and compliance with a variety of environmental permits and approvals from the relevant Burkinabé authorities.
The application for an Operating permit requires a Feasibility Study (FS) that must first be accepted by MEEVCC. The FS must include an ESIA, which must include a Resettlement Action Plan (RAP) that has been accepted by all stakeholders.
Once in production, a mining permit holder is required to open under his name, a Trust account named “Fonds de préservation et de réhabilitation de l’environnement minier” at the Banque Centrale des États de l’Afrique de l’ouest (“BCEAO”). This account must be funded annually on January 1 by an amount equal to the total rehabilitation budget presented in the ESIA, divided by the number of years of expected production to cover the costs of mine reclamation, closure and rehabilitation. The required permits and administrative procedures are presented in Table 20.1.
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Table 20.1 – List of Required Permits
Permit / Authorisation Main Requirements Timeframe Cost
Cost for review and validation of the ToR: Delivered by the MEEVCC on the basis of a compliant 500,000 FCFA. ESIA Report: Fees for the public inquiry and COTEVE Submission of the ESIA Terms of Reference (ToR) to the session: Approximately 3 million FCFA. BUNEE Before the expiry of Cost for file processing-Project value of 50 Environmental Validation of the ToR the Exploration billion FCFA and more: Compliance Certificate Submission of the draft version of the ESIA to the BUNEE License Flat Fee: 25 million FCFA. Public Inquiry Proportional rights: 0.02% of the total COTEVE Session investment cost = 26.3 million FCFA Final ESIA Report approved by the BUNEE (assuming a project cost of about 131.6 billion FCFA). Request to the MME specifying the type of industrial license being sought. References of the exploration license under which or from which the request is being made. Costs for the National Commission of Mines The mineral substances or for whom or which the license session: about 2 million FCFA. is sought. Definition of coordinates of the perimeter and requested area. Fixed duties on mineral titles: 5,000,000 FCFA. Location perimeter license requested on a topographic Proportional rights (area taxes): map at 1:200,000. 90 days before First five years: 7,500,000 FCFA/ km²/y. Permit for Industrial expiry of the Exploitation Detailed plan at an appropriate scale where the Year 6 to10: 10 million FCFA/ km²/y. coordinates of the perimeter requested are attached to Exploration Permit From the 11thyear: 15 million FCFA/ km²/y. outstanding points invariable ground and well defined. Cost of proportional royalty on gold production: Detailed statement showing the results of research 5% of turnover if gold is USD 1,300 and conducted and evidence of expenditures incurred during higher, 4% if USD 1,000-1,300 and 3% if less the last period of license validity. than USD 1,000. FS and Development Plan and operation of the deposit, including an ESIA and public inquiry results. Surface plan specifying land reserved for the operation and installation of industrial units.
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Permit / Authorisation Main Requirements Timeframe Cost
Commitment of the applicant, in the case of a large mining operations, the allocation of 10% of the share capital of the company operating free of all charges, to the benefit of the state. Mining Convention that the applicant intends to sign with the State; agreement consistent with the mining contract and attached to the operating license required.
Authorization for the Application to the Directorate of Legislation and 30 days after the Tax of raw water for mining and industrial Management of raw regulations of the Ministry in charge of water. quarter of the levy purposes: 200 FCFA /m³ a water No license or authorization required (besides the Mining Authorization for the 10 FCFA /m³ for every m³ backfill placed. Collection of raw water Licence) 20 FCFA /m³ for every m³ of concrete poured. for civil work However, tax must be paid. Authorization for Road Authorisation from the Ministry of Transport on the basis Depending on the infrastructure and technical
infrastructure of technical studies. studies that have been conducted. Authorization for Authorisation from the Ministry in charge of water on the Depending on the work and technical studies
Hydraulic work or dam basis of technical studies. that have been conducted.
Notes: a) Currently there is a Water Tax in Burkina Faso, however the Chamber of Mines, the Ministry of mines, the Ministry of Finances, and the Ministry of Water have not yet arrived at a consensus as to how to apply this to the Mining Sector. As a result mining companies have negotiated an exoneration of the water tax until the construction costs of the water reservoir are covered
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20.2 Baseline Studies
No environmental baseline studies have been undertaken by SEMAFO concerning the Nabanga Project at this stage.
20.3 Project Impacts, Risk Analysis, Environmental and Social Management Plan
The typical methodology used to identify and analyze the environmental impacts is based on an approach recognized by international funding agencies. This approach identifies the direct interactions between the project activities considered impact sources and the physical, biological, and human components. These interactions are customized according to project-specific phases (construction, operation and closure). All interactions identified are then analyzed on the basis of three criteria (intensity, extent and duration) to obtain a global indicator, the importance of the impact. Three (3) levels of the importance of the impact were determined: minor, medium, and major.
It is anticipated that the Nabanga Project will result in several negative impacts and some of the major impacts expected include the following:
• Surface water and groundwater.
• Noise levels.
• Social and economic aspects regarding resettlement.
• Cultural heritage.
The economic impact of the Project at local, regional, and national levels is positive. Beginning from the construction phase, direct and indirect jobs will be created, resulting in tangible economic benefits for both local and regional communities. The project will create hundreds of skilled and unskilled direct and indirect jobs, most of them awarded to Burkinabé workers. This job creation will increase household incomes and improve living conditions. In addition, the procurement of goods and services required for the construction, operation, and closure of the mine will bring significant economic benefits to local and regional businesses, the majority in terms of supplying food and/or various products.
The revenues generated by the mining operation will increase Burkina Faso’s internal revenue through taxes and royalties charged by the local authorities. These revenues should have a beneficial impact at both the local and regional levels through increased investments in social and health services, and educational facilities. In addition, SEMAFO typically supports a number of social programs for the displaced households as it was previously done for the Mana and Boungou operations, and in a broader context, the local and regional communities.
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20.3.1 RISK ANALYSIS
No formal Risk Analysis was conducted so far on the Nabanga Project as it is still at an early stage of development. Further analysis will be conducted in a later phase such as the Feasibility or Construction phase.
20.3.2 ENVIRONMENTAL AND SOCIAL MANAGEMENT PLAN
No formal Environmental and Social Management Plan (“ESMP”) has been developed so far on the Nabanga Project. The ESMP will cover all project phases and allow avoiding, minimising, enhancing, or compensating the various anticipated impacts by reducing them to an acceptable level for all stakeholders.
20.4 Acid Rock Drainage & Waste Disposal
At this stage, no Acid Rock Drainage (“ARD”) and Waste Disposal Management Plan have been developed. During the course of the Feasibility Study, further detailed test work will be undertaken and ultimately a formal ARD (if necessary) and Waste Disposal Management Plan will be developed.
20.5 Closure, Decommissioning, and Reclamation
Following operations, the site will undergo comprehensive decommissioning and reclamation. The decommissioning and reclamation plan will conform to Burkinabé requirements.
The Nabanga Project closure, decommissioning and reclamation costs are estimated to 5.0 MUSD. These costs were included in the financial analysis for these closure activities.
The items presented in the cost estimation are based on a Conceptual Closure and Reclamation Plan and has an accuracy level of up to ±50%. The Closure and Rehabilitation Plan and its costs estimates will have to be better detailed prior to provide the Reclamation Trust to BCEAO. It is of the QP’s opinion that this level of accuracy on the closure cost will have negligible effect on the global project economics.
This Mine Closure and Reclamation Plan include work to be conducted from the closure of the mine, at the end of operational activities, as well as progressive rehabilitation work.
Also, the closure, decommissioning and reclamation will be prepared to satisfy the concerns of all stakeholders. The closure and reclamation plan will be developed in accordance with the national guidelines for preparing a mining site rehabilitation plan. The mine closure plan will be approved before the start of operations. The main objectives of the Closure and Reclamation Plan include restoring ecosystems and take-over and recovery of land uses.
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20.6 Conclusion
As the Project is in its early stages of study, little has been done to date to promote the Project to the various stakeholders and investigate local environmental and social settings.
As the Project progresses, an environmental and social impact study will be initiated, as well as public hearings and stakeholder consultations. The mitigation measures and environmental management plans will be prepared and discussed during the public inquiry, as well as a mine closure and reclamation plan.
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21 CAPITAL AND OPERATING COSTS
The Project scope covered in this Report is based on the construction of a green field mining and processing facility with an average mill feed capacity of 360,000 tonnes per year of material and producing an average of 70,000 oz per year of gold.
The capital and operating cost estimates related to the mine, the concentrator, and all required facilities and infrastructure have been developed by DRA/Met-Chem or consolidated from external sources.
The capital and the operating costs are reported in United States Dollars (“$”).
21.1 Capital Cost Estimate (Capex)
The capital cost estimate (“Capex”) consists of direct and indirect capital costs as well as contingency. Provisions for sustaining capital are also included, mainly for tailings storage expansion. Amounts for the mine closure and rehabilitation of the site have been estimated as well. The working capital is discussed in Section 22 – Economic Analysis.
21.1.1 SCOPE OF THE ESTIMATE
The Capex includes the material, equipment, labour and freight required for the mine pre- development, processing facilities, tailings storage and management, as well as all infrastructure and services necessary to support the operation.
The Capex prepared for this PEA is based on a Class 4 type estimate as per the Association for the Advancement of Cost Engineering (“AACE”) Recommended Practice 47R-11 with a target accuracy of ±35%. Although some individual elements of the Capex may not achieve the target level of accuracy, the overall estimate falls within the parameters of the intended accuracy.
The reference period for the cost estimate is Q3 2019.
21.1.1.1 Major Assumptions
The Capex is based on the Project obtaining all relevant permits in a timely manner to meet the Project schedule.
The Capex reflects an Engineering, Procurement and Construction Management (“EPCM”) type execution wherein one EPCM contractor will provide the design, procurement and construction activities for the Project. All sub-contracts would be managed by the EPCM contractor.
All back-fill materials will be available from gravel pits or other sources located close to the site. Mine waste rock is not suitable for use in road construction. All excavated material will be disposed of within the site battery limits.
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21.1.1.2 Major Exclusion
The following items were not included in the Capex:
• Provision for inflation, escalation, currency fluctuations and interest incurred during construction;
• Project financing costs;
• All duties and taxes; however, they are considered in the economic analysis.
21.1.2 CAPITAL COSTS SUMMARY
Table 21.1 presents a summary of the pre-production initial capital and the sustaining capital costs for the Project.
a. Pre-production Initial Capital Cost
The pre-production initial capital cost for the scope of work is $ 83.7 M, of which $ 52.6 M is direct cost, $ 17.2 M is indirect cost, $ 13.9 M is contingency.
Provision for sustaining capital is $ 55.9 M, excluding the amounts for working capital.
Table 21.1 – Initial Capex Summary
Total Costs Area Area Description (‘000 USD) Direct Costs 1000 Mining 8,744 2000 Crushing 5,451 3000 Concentrator 22,181 4000 Tailings Management System 2,440 5000 General Site Infrastructure 7,130 6000 Electric Power Plant 6,615 Sub-total – Direct costs 52,561 Indirect Costs 9000 Indirect Costs 17,187 9000 Contingency 13,949 Sub-total –Indirect costs 31,136 TOTAL: 83,697
Numbers may not add due to rounding.
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b. Closure and Rehabilitation Costs
Based on site layouts, a provision of $ 5.0 M was estimated for the closure and rehabilitation of the mine site. Requirements were established and cost estimation was based on material take- off and unit rates from recent projects.
The expenses were accounted for in the economic analysis and incurred during the final years of the life of mine.
No provision is required for the dismantling and disposal of the industrial facilities as it is assumed that the costs will be compensated by the salvage value.
21.1.3 BASIS OF ESTIMATE GENERAL
The capital cost estimate covers the facilities included in the scope of the work described in previous sections.
The Capex is based on the following key assumptions:
• The proposed construction work week is 8 hours per day, 5 days per week. All construction workers will be either from the nearby area or housed in a camp provided by the Owner;
• Fluctuations to nominated currency exchange rates are excluded;
• Allowance for industrial dispute or lost time arising from industrial actions is excluded;
• All taxes and duties are excluded;
• Escalation is excluded;
• Project financing costs and interest during construction is excluded;
• No allowance is made for acceleration or deceleration of the Project schedule;
• Project insurance is included in the Owner’s cost.
The Capex is based on an advance period whereby the design concepts are frozen and basic engineering commences. The timing for this start is following approval of the EIS and obtaining initial permits. The work would then continue through its lifecycle until the end of the construction and commissioning.
a. Currency
The Capex base currency is United States Dollars (USD). The Capex consists of items quoted in various foreign currencies which have been converted into USD using exchange rates as of August 30, 2019. Table 21.2 shows the currency exchange rates used in this Report.
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Table 21.2 – Currency Exchange Rates
Currency Source Base Description Exchange Currency Currency Rate USD United States Dollar USD 1.000 CAD Canadian Dollar USD 0.752 EUR EURO USD 1.166
b. Deliverables
The Capex estimate was developed based on the following list of deliverables:
▪ Project description;
▪ Mine plan, complete with initial mining equipment and pre-production costs;
▪ Mechanical equipment list;
▪ MTO for major electrical equipment, including the power plant;
▪ MTO for tailings storage, including tailings’ roads as well as tailings and reclaim water pipelines;
▪ Overall general arrangement plan.
c. Estimating Software
The Capex was developed using MS Excel.
d. Labour Cost and Productivity
Labour manhours were developed internally, for each site activities. The productivity factors vary as a function of the expected qualifications as well as of the building height and the congestion; it varies from 1.16 to 1.64 with an overall weighted average of 1.33. It should be noted that a PF of 1.0 refers to projects being executed with better than average skill, base 40 hours workweek, within reasonable commuting distance, limited in-plant movement, favorable weather, etc.
Labour rates were developed based on salary information reflecting local labour. They are inclusive of salaries, contractors’ indirect costs, namely mob and demob, small tools, construction equipment, consumables, PPE, temporary site establishments, supervision, administration as well as overhead and profit.
It is assumed that the local communities can accommodate the direct and indirect workforce estimated for the Project, including occasional site visits and vendor representatives. The peak workforce is estimated to reach 300, with an average of 200. Local accommodation as well as rotational transportation costs are included as part of construction field indirect costs.
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e. Freight, Duties and Taxes
Based on recent surveys and studies and when not included in the cost, the freight was accounted for by adding a factor to the value of the goods; a factor of 10 % is applied.
All duties and taxes were excluded from the capital cost, but relevant factors were considered for the after-tax economic analysis.
21.1.4 METHODOLOGY a. Data Sources
A mechanical equipment list was developed by Engineering. Conceptual estimates, supplemented by general arrangements drawings, were used for civil works, including earthworks, concrete and structural steel. To ensure the entire scope coverage, some allowances were added, based on DRA/Met-Chem’s experience. Piping, electrical distribution downwards of electrical rooms as well as instrumentation and controls were factored from mechanical.
Budgetary quotations were obtained for major plant equipment with the balance of the plant equipment costs generally developed based on an internal database. Some equipment costs were estimated when no relevant data was available.
Rates for bulk material were estimated.
b. EPCM Services
While the Project may not ultimately be executed via the EPCM model, the cost estimate was structured on that basis. EPCM services consist of the following:
▪ EPCM team salaries, fringes, uplifts, recruitment, overhead, etc.;
▪ EPCM team expenses (i.e. business travelling, room & board, accommodation, etc.);
▪ Home office support and expenses (communications, IT services, IT equipment, courier, printing, office space, furniture, consumables, stationaries, etc.).
For the initial Capex, EPCM services costs are estimated at 14% of the direct costs.
c. Construction Field Indirect Cost
Site construction indirect costs are included as a percentage of direct costs:
▪ Site preparation for all temporary infrastructures and buildings, construction facilities, laydown areas, temporary services, etc.;
▪ Temporary roads, walkways, parking areas and fencing, c/w signage and including temporary lighting, complete with maintenance;
▪ Temporary buildings/construction facilities (offices – for EPCM and Owner’s staff, camp, cafeteria, laundry facilities, medical clinic, security gate/office, etc.), complete with
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mobilization, demobilization, rental, operations and maintenance. It should be noted that contractors will be responsible for the provision of their own temporary facilities.
▪ Temporary infrastructure for the supply of power, fuel, gas, water and communications. It should be noted that contractors will be responsible for their own temporary infrastructures.
▪ Temporary infrastructure for the management of sewerage and construction waste (dry and wet, hazardous and non-hazardous), including collection, treatment and disposal. It should be noted that contractors will be responsible for their own requirements.
▪ Pad preparation and fencing – only – of contractor’s pads are included in the construction field indirect costs;
▪ Field office supply (IT equipment, courier, printing, office space, furniture, consumables, etc.);
▪ Access control and monitoring;
▪ Temporary lay down and storage areas, as well as warehousing, complete with, but not limited to, materials management and materials handling equipment;
▪ Mobilization and demobilization of all above listed temporary site establishments and restoration back to original site conditions;
▪ Site surveying
▪ Site security;
▪ Light vehicles;
▪ First aid and medical services;
▪ General and final clean-up.
Construction field indirect costs are estimated at 10% of the direct costs. They are inclusive of vendor representatives for construction and commissioning.
d. Owner’s Costs
Owner’s costs were estimated at 4% of all direct costs.
e. Other Costs
Costs for spare parts, special tools and initial fills are estimated at 3.25% of equipment costs for the initial capex phase. It is assumed that none will be required for sustaining capex as they are included as part of normal operations.
f. Project Contingency
For initial Capex, the project contingency was assessed at 13.9 M, representing 20% of all costs.
g. Inflation
Inflation beyond this Capex estimate base date is explicitly excluded.
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h. Risks
Risks, complete with mitigation plans, are explicitly excluded from this Capex estimate.
21.1.5 QUALIFICATIONS
All estimates are developed within a frame of reference defined by assumptions and exclusions, grouped under the estimate qualifications. Assumptions and exclusions are listed in the following paragraphs.
21.1.5.1 Assumptions The following items are assumptions concerning the Capex: • Estimate is based on rotations schedule of 4 and 2, i.e. 4 weeks in and 2 weeks R&R, with travelling during the 2 weeks R&R;
• Estimate is based on 5 days at 8 hours per day workweek;
• Estimate assumes that labour skills will be medium;
• Estimate assumes aggregates used for fill, adequate both in terms of quality and quantity, will be available within a 5 km radius from site;
• Estimate assumes concrete, adequate both in terms of quality and quantity, will be available within a 15 km radius from site;
• Estimate assumes overburden disposal will be within a 5 km radius from the construction site;
• Estimate assumes fresh water, adequate both in terms of quality and quantity, is available locally at no costs and does not need any treatment to be used for concrete mix, leak/hydro testing, flushing, cleaning, etc.;
• Estimate assumes drinking water will be bottled;
• Estimate assumes EPCM and Owner’s teams will be in sufficient quantity so as not to delay contractors;
• Estimate assumed smooth coordination between contractors’ battery limits;
• Estimate assumes 40% of manual labour will be sourced within the Nabanga area while 60% will be a combination of remote workers and expats from neighboring countries;
• Estimate assumes no labour decree is in effect in Burkina Faso;
• Estimate assumes security fencing surrounding the mine, camp and processing facilities;
• Estimate assumes construction contracts types will be either lump sum, cost plus or unit rates;
• Estimate assumes no construction contracts will be time and materials;
• Estimate assumes no underground obstructions of any nature;
• Estimate assumes no hazardous materials in excavated materials.
• Estimate assumes no delay in Client’s decision making;
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• Estimate assumes no delay in obtaining permits and licenses of any kind;
• No interruption in job continuity;
• Estimate assumes normal peak workforce;
• Estimate assumes engineering progress prior to the execution will be sufficient so as to avoid rework.
21.1.5.2 Exclusions The following items are not included in the Capex: • Currency fluctuations;
• Any and all scope change;
• Inflation beyond the Capex estimate base date;
• Risk;
• Financing charge;
• Delays resulting from community relation, permitting, project financing, etc.;
• Any and all taxes, customs charges, excises, etc.;
• Closure costs.
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21.2 Operating Cost Estimate (Opex)
21.2.1 INTRODUCTION
This Section provides information on the estimated operating costs of the Project and covers mining, processing, site services and general administration. Table 21.3 presents the operating costs summary.
The sources of information used to develop the operating costs include in-house databases and outside sources.
Table 21.3 – Summary of Operating Costs
Cost per ounce Total Costs Description of gold produced (‘000 USD) ($/oz) Mining 154,570 271 Processing 139,090 244 General & Administration 39,860 70 Government Royalty 44,420 78 TOTAL OPEX 377,940 662 Sustaining Capex 55,910 98 TOTAL AISC 433,850 760
Numbers may not add due to rounding
21.2.2 SUMMARY OF SEMAFO PERSONNEL REQUIREMENTS
Table 21.4 presents the personnel requirements for the Project (excluding the mining contractors for the open pit and underground operations).
Table 21.4 – SEMAFO Personnel Requirement
Area Number Mine (excluding contractors) 17 Processing 69 Management & Administration 94 Total Manpower 180
21.2.3 GENERAL AND ADMINISTRATIVE COSTS
General and administration operating costs have been sub-divided in three (3) categories: Manpower, General Services, and Site Services. Manpower includes finance, purchasing, warehouse, health & safety, environmental, human resources, security and other support personnel.
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General services include various office-related costs, as well as the lodging and travel expenses for expatriate personnel. Site services comprise the costs for operation and upkeeping of site service- related facilities.
Table 21.5 – Summary of General and Administrative Costs
Cost/Year Cost/tonne Total Cost Area ($) ($/t) (%) Manpower 2,003,272 5.56 42 General Services 1,748,000 4.86 37 Site Services 1,015,100 2.82 21 Total G&A Costs 4,766,372 13.24 100
21.2.4 MINING OPERATING COSTS
The Open Pit operating costs have been estimated using similar contractors’ costs used in other SEMAFO mine sites in Burkina Faso. This cost does not include a cost increment to consider the deepening of the pit. It is an average cost over the pit life. The unit costs used are the following:
• Ore mining costs: $2.47/t;
• Waste mining costs: $2.36/t.
The underground operating costs have been estimated using similar contractors’ costs than the one of other SEMAFO UG operation. A detailed estimation exercise was done using these contractor rates, for manpower, equipment, and maintenance. The resulting unit costs for production and development are the following:
• UG mineralized material production unit cost: $53.68 per tonnes mined.
• UG lateral development unit cost: $5,789 per m of advance (using a section of 23.6 m² on average).
Table 21.6 – Summary of Mining OPEX
Total Cost Cost/tonne Total Cost Description ($) ($/t) (%) Open Pit mineralized material OPEX 27,649,594 44.89 18 Underground mineralized material OPEX 126,918,531 53.68 82 Total Mineralized Material Mining OPEX 154,568,125 51.86 100
21.2.5 PROCESS OPERATING COSTS
For a typical year at design processing rate, the estimated process operating costs are divided into six (6) main components: Manpower, electrical power, grinding media and reagent consumption, Maintenance parts, material handling and spare parts and miscellaneous. The breakdown of these
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costs is summarized in Table 21.7. These costs were derived from supplier information, DRA/Met- Chem’s database or factored from similar operations. The unit cost of on-site generated electricity was established at $ 0.30/kWh.
Table 21.7 – Summary of Process OPEX
Cost/t Cost/oz of gold Total Cost Operating Cost Area Cost/year 1) processed 2) produced 3) (%) Manpower 1,748,600 4.86 24.99 10.4 Electrical power 10,148,300 28.19 145.03 60.4 Grinding media and reagents 2,790,000 7.75 39.87 16.6 consumption Maintenance parts 4) 1,200,000 3.33 17.15 7.1 Material handling 513,800 1.43 7.34 3.1 Spare parts and 400,000 1.11 5.72 2.4 miscellaneous 5) Total Operating Cost 16,800,700 46.67 240.10 100.0
Note:
1) 1 USD is one United States of America Dollar (holds for all other tables)
2) Based on Plant throughput rate of 360 thousand tonnes per year (on a dry basis).
3) Based on gold production of 69,974 troy ounces per year.
4) Maintenance wear items estimated as 6% of total equipment capital cost.
5) Capital spare parts, estimated as 2% of total equipment capital cost.
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22 ECONOMIC ANALYSIS
The project has been evaluated using discounted cash flow analysis (DCF). Cash inflows were estimated based on annual revenue projections. The economic/financial assessment of the project is based on Q3-2019 price projections and cost estimates in U.S. currency. No provision was made for the effects of inflation. The evaluation was carried out on a 100 %-equity basis.
Current tax regulations in Burkina Faso were applied to assess the project’s tax rate and royalty regime. The Net Present Value (NPV) of the project was calculated by discounting back cash flow projections throughout the life-of-mine (LOM) to the Project’s valuation date using a base case discount rate of 5%. The internal rate of return (IRR) and the payback period were also calculated.
The financial indicators under base case conditions are shown in Table 22.1.
Table 22.1 – Financial Model Indicators
Base Case Financial Results Unit Value
Pre-Tax NPV @ 5 % M USD 146.7 After-Tax NPV @ 5 % M USD 99.8 Pre-Tax IRR % 31.4 After-Tax IRR % 22.6 Pre-Tax Payback Period years 3.5 After-Tax Payback Period years 4.4
22.1 Assumptions
22.1.1 MACRO-ECONOMIC ASSUMPTIONS
The main macro-economic assumptions used in the base case are given in Table 22.2. The gold price assumption is based on the latest corporate directives from SEMAFO. The sensitivity analysis examines a range of prices going from $1,000/oz to $1,600/oz of gold.
Table 22.2 – Macro-Economic Assumptions
Item Unit Base Case Value
Gold Price Forecast USD/oz. 1,300 Marketing and Refining Charges USD/oz. 3 Discount Rate % per year 5
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22.1.2 MINERAL ROYALTIES
Burkina Faso royalties on gold have been applied to the project. A 5% royalty has been considered for the Project; moreover, a 1% development tax has been applied on the total revenue.
22.1.3 TAXATION REGIME
Annual corporate tax liabilities were calculated under the Burkina Faso tax regime. These are based on the 2015 new mining tax laws applied to capital costs, operating costs, sales of gold and profits.
The following assumptions were applied when calculating corporate taxes:
• 27.5% of corporate tax rate (income tax)
• Taxes are payable periodically, assumed yearly
• The project’s assets are depreciated over their useful life, which is the total life of the project (8 years) which is in accordance with Burkina Faso Taxation legislation
22.1.4 PROJECT FINANCING
No assumptions have been made on the financing of the project for the purpose of determining the net revenue.
22.1.5 TECHNICAL ASSUMPTIONS
The first production year consists of a ramp-up period of 3 months followed by 9 months during which production meets full capacity.
The engineering and construction period have been estimated at approximately 24 months, which is relatively common for this type of project given its size.
22.2 Cash Flow Analysis and Financial Results
Figure 22.1 illustrates the after-tax cash flow, the income tax, and the cumulative cash flow profiles. The graph shows that the after-tax payback period has been estimated at approximately 4.4 years.
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Figure 22.1 – After-Tax Cash Flow and Cumulative Cash Flow Profiles
A summary of the evaluation results is provided in Table 22.3 and Table 22.4 denotes the cash flow statement for the base case scenario.
The summary and cash flow statement indicate that the total pre-production (initial) capital costs was evaluated at $ 83.7 M. The sustaining capital requirement was evaluated at $ 55.9 M. Mine closure costs have been estimated at $5.0M.
The cash flow statement shows the estimated capital spending schedule over the 18-month pre- production period of the Project (the Year -1 column covers a 6-month period).
The total revenue derived from the sale of the gold was estimated at $ 740.3 M, and the total operating costs were estimated at $ 333.5 M (excluding government royalties).
The financial results indicate a Pre-Tax Net Present Value (“NPV”) of $ 146.7 M at a discount rate of 5 %. The Pre-Tax Internal Rate of Return (“IRR”) is 31.4 % and the payback period is 3.5 years.
The After-Tax NPV is $ 99.8 M at a discount rate of 5 %. The After-Tax IRR is 22.6 % and the payback period is 4.4 years.
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Table 22.3 – Project Evaluation Summary – Base Case
Item Unit Value
Total Revenue M USD 740.3 Total Operating Costs (excludes government royalties) M USD 333.5 Total Government Royalty Payments M USD 44.48 Initial Capital Costs (excludes Working Capital) M USD 83.7 Sustaining Capital Costs M USD 55.9 Closure Costs M USD 5.0 Total Pre-Tax Cash Flow M USD 217.7 Pre-Tax NPV @ 5 % M USD 146.7 Pre-Tax NPV @ 8 % M USD 115.1 Pre-Tax NPV @ 10 % M USD 97.4 Pre-Tax IRR % 31.4 Pre-Tax Payback Period Years 3.5 Total After-Tax Cash Flow M USD 157.9 After-Tax NPV @ 5 % M USD 99.8 After-Tax NPV @ 8 % M USD 74.1 After-Tax NPV @ 10 % M USD 59.7 After-Tax IRR % 22.6 After-Tax Payback Period* Years 4.4
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Table 22.4 - Cash Flow Statement – Base Case
Description Total or Avg. Construction Year Year Year Year Year Year Year Year Year LOM Period 1 2 3 4 5 6 7 8 9 MINE SCHEDULE Open-pit (t) 615,974 88,067 308,251 219,657 Open-pit grade (g/t) 6.45 6.20 6.33 6.70 Open-pit waste (t) 2,894,907 5,401,199 5,762,218 UG mineralized material (t) 2,364,459 88,411 359,981 358,593 360,361 360,726 391,598 360,883 83,907 UG grade (g/t) 6,48 6.96 5.75 6.87 7.14 6.81 5.51 6.67 6.96 Total material mined (t) 2,980,434 88,067 308,251 308,068 359,981 358,593 360,361 360,726 391,598 360,883 83,907 Mineralized material grade (g/t) 6.47 6.20 6.33 6.78 5.75 6.87 7.14 6.81 5.51 6.67 6.96 PROCESSING SCHEDULE Material processed 2,980,434 330,000 360,000 360,000 360,000 360,000 360,000 360,000 360,000 130,434 Grade processed (g/t) 6.47 6.3 6.69 5.79 6.83 7.13 6.82 5.55 6.54 6.81 Recovery (%) 92 92 92 92 92 92 92 92 92 92 Production (oz) 570,768 61,531 71,263 61,602 72,750 75,897 72,649 59,151 69,656 26,268 REVENUES (in $) 740,286,092 79,806,232 92 427 652 79 897 446 94 357 380 98,438,174 94,226,215 76,718,696 90 344 453 34,069,843 OPERATING COSTS (in $) (377,935,899) (40,893,017) (46 773 854) (48 473 535) (44 778 921) (46,829,207) (49,732,430) (44,887,293) (42 833 341) (12,734,303) EBITDA (in $) 362,350,192 38,913,215 45 653 798 31 423 911 49 578 459 51,608,968 44,493,785 31,831,403 47,511,113 21,335,540
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Description Total or Avg. Construction Year Year Year Year Year Year Year Year Year LOM Period 1 2 3 4 5 6 7 8 9 Accounts receivable (in $) N/A 9,975,779 11,553,457 9,987,181 11,794,672 12,304,772 11,778,277 9,589,837 11,293,057 4,258,730 Accounts payable (in $) N/A (3,407,751) (3,897,821) (4,039,461) (3,731,577) (3,902,434) (4,144,369) (3,740,608) (3,569,445) (1,061,192) Working capital (in $) N/A 6,568,028 7,655,635 5,947,720 8,063,096 8,402,338 7,633,908 5,849,229 7,723,612 3,197,538 Change in working capital (in $) N/A (6,568,028) (1,087,608) 1,707,916 (2,115,376) (339,242) 768,430 1,784,678 (1,874,382) 4,526,073 3,197,538 Initial capex (83,695,551) (83,695,551) Sustaining capex (55,914,520) 556,818 19,500,314 13,722,562 13,730,456 7,357,276 1,047,095 Rehabilitation & closure costs (5,000,000) (1,250,000) (1,250,000) (2,500,000) CASH FLOW 217,740,122 (90,263,579) 37,268,790 27,861,401 15,585,973 35,508,761 45,020,122 45,231,369 28,707,020 50,787,186 22,033,079 Total cash cost /oz 662 665 656 787 616 617 685 759 615 485 All-in sustaining cost /oz 760 665 664 1,103 804 798 786 777 615 485
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22.3 Sensitivity Analysis
A sensitivity analysis has been carried out, with the base case described above as a starting point, to assess the impact of changes in total pre-production capital expenditure (“Capex”), operating costs (“Opex”) and gold price (“Price”) on the Project’s NPV @ 5 % and IRR. Each variable was examined one-at-a-time. An interval of 30 % with increments of 10 % was used for the three (3) variables.
The Pre-Tax results of the sensitivity analysis, as shown in Figure 22.2, indicate that, within the limits of accuracy of the cost estimates in this Report, the Project’s Pre-Tax viability does not seem significantly vulnerable to the under-estimation of capital and operating costs, taken one at-a-time. As seen in Figure 22.2, the NPV is more sensitive to variations in Opex than Capex, as shown by the steeper slope of the Opex curve. As expected, the NPV is most sensitive to variations in price. The NPV becomes negative at a price variation of more than -30%. This corresponds to a break-even gold price of about $950/oz.
Figure 22.2 – Pre-Tax NPV5 % & IRR Sensitivity to Capital Expenditure, Operating Cost and Price
The variations in internal rate of return, provides the same conclusions. Due to the different timing associated with Capex versus Opex, the IRR is more sensitive to variations in Capex than Opex for negative variations, but remains less sensitive to Capex for positive variations.
The same conclusions can be made from the After-Tax results of the sensitivity analysis as shown in Figure 22.3. It indicates that the Project’s after-tax viability is mostly vulnerable to a price forecast reduction, while being less affected by the under-estimation of capital and operating costs.
The NPV becomes negative at a price variation of more than 30% also, with a break-even gold price of about $970/oz.
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Figure 22.3 – After-Tax Royalty NPV5 % & IRR Sensitivity to Capital Expenditure, Operating Cost and Price
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23 ADJACENT PROPERTIES
Although some valid permits are present in the area of the Yactibo Project, they are all owned by independent prospectors and no information can be found with regards to exploration work.
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24 OTHER RELEVANT DATA AND INFORMATION
There is no additional information or explanation necessary to make this Technical Report understandable and not misleading.
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 162
25 INTERPRETATION AND CONCLUSIONS
This Report was prepared and compiled by DRA/Met-Chem under the supervision of the QPs at the request of SEMAFO. This Report has been prepared in accordance with the provisions of National Instrument 43-101 Standards of Disclosure for Mineral Projects.
25.1 Conclusions
The mineralized material planned to be mined as part of the Nabanga Project includes 616 Kt grading 6.45 g/t of gold from the Open Pits and 2.364 Mt grading 6.48 g/t of gold from the underground mine. The total mineralized material planned to be mined over the life of mine is 2.980 Mt at an average grade of 6.47 g/t of gold. The duration of the life of mine has been estimated at approximately 9 years. Given the relatively short duration of the operation, it has been assumed that a mine contractor would be hired to perform the Open Pit and Underground mining activities.
The processing plant is designed to process 360 Kt/y of run of mine to produce approximately 68,000 ounces per year of gold based on a recovery of 92%. A suitable process flowsheet including crushing, grinding, flotation, regrind, leaching, elution, and carbon regeneration.
Tailings storage facility, camp facilities, power generation facilities as well as infrastructure and services have been added to complete the investment cost of the Project.
The pre-production initial capital cost, at an accuracy level of ±35%, is evaluated at 83.7M USD while the sustaining capital requirement, at an accuracy level of ±50%, is 55.9M USD.
The life of mine all-in sustaining costs is evaluated at 760 USD/oz of gold produced.
Mine closure and rehabilitation cost have been estimated at 5M USD.
The economic analysis of the project has demonstrated the potential viability of the project over its 9 years life of mine expectancy with recommendations to proceed to next level of Feasibility studies. At an average gold price of $1,300/oz, the financial results indicate a Pre-Tax NPV of 146.7 M USD at discount rates of 5 %. The before-tax Internal Rate of Return is 31.4% with a payback period of 3.5 years. The After-Tax NPV is 99.8 M USD at discount rates of 5%. The After-Tax IIR is 22.6 % and the payback period is 4.4 years.
25.2 Risk Evaluation
The risks affecting the economic and technical viability of the Project will be reduced as geotechnical and hydrogeological studies, more metallurgical testing, and engineering are undertaken during the next phase.
As for all mining projects, external risks beyond the control of the project such as the political situation in the Project region, commodity prices, exchange rates and government legislations are much more difficult to anticipate and mitigate.
Negative variance to these risks from the assumptions used to build the block model may have an impact on Mineral Resource Estimates.
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 163
26 RECOMMENDATIONS
26.1 Mining and Geology
It is recommended to increase the level of knowledge of the mineral resources of the Nabanga Project with the implementation of a new drilling campaign.
This campaign should aim at bringing the resource category to the Measured and Indicated level as much as possible, and also to increase the resource base at depth and along strike.
Given the size of the project and its expected life of mine duration, a formal discussion with different potential mining contractors should be done.
The ventilation system should also be studied in more details in the next stages of the project, and a ventilation model should be developed.
26.2 Geotechnical
As the Nabanga Project will move ahead into the next phases of development, more geotechnical test work and analysis will be needed. A geotechnical drilling program will have to be carried out, with the following objectives:
• Final pit slope angle assessment;
• Geotechnical and ground support recommendations for portals excavations;
• Recommendation for stope dimension;
• 2D/3D finite element analysis, and ground support recommendations for the development areas and the production stope.
26.3 Hydrogeology and Hydrology
A hydrogeological investigation program should take place, possibly at the same time than the next geotechnical drilling program (for assessment of the pit slope and stope stability). Pumping / dewatering hole could be designed at this point, if required.
Similarly, a hydrology assessment should take place as part of the PFS or DFS of the Nabanga Project. This work is required to evaluate the rain water contribution on the overall drainage. The hydrological data will be used to design the drainage system of the site, sediment ponds / sumps to collect the water and pumping / piping systems to remove the water from the open pit. Although the annual average rainfall is low, management of surface stormwater will be important due to large inter- year variability, potential for heavy rains periods during the wet month and relatively flat ground.
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 164
26.4 Process Recommendations
26.4.1 TRADE-OFF STUDIES
There are a number of Trade-Off Studies (TOS) that will need to be conducted prior to or during the prefeasibility stage. These TOS were identified during the PEA stage of the Project, but were not conducted due to the timing and/or availability of metallurgical test work information during the PEA.
26.4.1.1 Grinding using one large mill or two smaller ones.
Since there will be a large variability in material hardness, a trade-off study should be done to investigate the most efficient method of absorbing these variations. Two (2) smaller mills will adapt to changes in material hardness easier than one large mill. Further, two (2) small mills are more energy efficient, but will have a larger Capex associated with them. The best selection needs to be investigated in the FS.
26.4.1.2 Intensive Cyanide Leaching of Flotation Concentrate
There are additional operating costs and issues associated with intensive cyanidation. Cyanide also dissolves copper sulphides and this has been an issue at operations that have of copper sulphides in the mill feed stream. It can lead to high cyanide consumption levels and produce doré with high copper levels. This will need to be investigated in the test work program at the next stage of project development.
26.4.1.3 Carbon–in-Leach versus Carbon-in-Pulp (CIL vs CIP)
CIL is the conservative approach of designing a gold operation. The CIL plant ensures that if preg- robbing (gold) particles are in the slurry the carbon adsorb the gold preferentially. However, this leads to a significant higher carbon inventory and lower gold loading on the carbon and thus less efficiency during stripping. CIP entails the addition of CIP tanks after leaching has been completed. The inventory and wear of carbon is thus less. A CIL vs. CIP trade-off study is recommended for the next stage of the project to validate the Capex impact.
26.4.1.4 Post Leach Thickener vs No Thickening
Some plants with high discharge cyanide levels have a post leach cyanide thickener after leaching. This reduces cyanide consumption.
26.4.1.5 Large Bulk Reagent Tanks versus Totes
The decision to use drums reagent in the PEA design is based upon prior experience. During the feasibility study, reagent costs for bulk versus totes should be evaluated in greater detail to determine the most cost-effective option. Generally, bulk reagent deliveries are often lower in cost and the issue dealing with empty totes is avoided. Evidently, there is an associated Capex cost associated with storage of bulk chemicals.
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 165
26.4.2 OTHER TEST WORKS
Additional physical characterisation test work will need to be completed. It is understood that further testing on mineralized material from deeper in the deposit and variability testing will need to be completed.
Flotation concentrate characterisation including settling, filtering and physical characterisation for regrind mill requirements will be required.
Additional leach and adsorption testing will be required including variability test work to determine any differences throughout the mineralized material with respect to recovery and reagent consumption. Testing will also be used to determine the requirement of Oxygen vs. Air as a leach circuit oxidant and also to determine the optimum leach and adsorption circuit flowsheet, CIP vs. CIL.
26.5 Environment
It is recommended to perform the following work in connection with environmental activities:
• Extend soil and surface water surveys to select the best location for the tailing ponds, waste rock and overburden piles;
• Carry out a hydrogeological study to collect field data in order to estimate from groundwater flow modeling dewatering rates and impacts;
• Start the environmental studies required to support permitting requirement and to optimize the site layout;
• Identify environmental requirement for site closure and estimate the cost.
26.6 Proposed Work Program
To ensure the potential viability of the mineral resources and confirm the potential financial evaluation of the Project, proposed activities to be undertaken in the next phase have been identified. These activities along with their estimated costs are shown in Table 26.1.
Table 26.1 – Estimated Budget for Next Phase
Estimated Budget Activities ($) Definition Drilling Campaign 1,000,000 Geotechnical and Hydrogeology Studies 500,000 Metallurgical Test Work Program 300,000 Environmental Studies 1,000,000 Feasibility Study 1,000,000 Sub-Total 3,800,000 Contingency (20 %) 760,000 Total 4,560,000
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 166
27 REFERENCES
Grenholm, Mikael; 2014; The Birimian event in the Baoulé Mossi domain (West African Craton); Regional and Global Context; Dissertations in Geology at Lund University, Master’s thesis, no 375 (45 hp/ECTS credits).
Milési, Jean-Pierre et al.; 1989; West African Gold Deposits in their Lower Proterozoic Lithostructural Setting; Chron, rech. Min. no 497, pp. 3-98, Édition du BRGM.
Ministère de l’Énergie, Burkina Faso; Projet d’appui au Secteur de l’Électricité (PASEL); Mars 2017; SFG3233; Notice d’impact environnemental et social de l’extension de la centrale thermique de Fada N’Gourma; Version actualisée.
Pells Sullivan Meynink; January 2014; Orbis Gold Ltd.; Nabanga Project Scoping Study: Preliminary Geotechnical Assessment; Report PSM1814-003R.
Snowden, 2012. Nabanga Mineral Resource Estimate, unpublished internal report prepared by Snowden Mining Industry Consultants for Orbis Gold Limited, project number 03771, September 2012.
Snowden, 2014. ITR (Valuation of Orbis Gold Ltd Mineral Assets, report prepared by Snowden Mining Industry Consultants for Orbis Gold Limited, project number AU4523, December 2014.
Snowden, 2015. Yactibo Permit Group, Nabanga Gold Deposit, prepared by Snowden Mining Industry Consultants for SEMAFO Inc., Project No. AU4582, NI 43-101 Technical Report, June 2015.
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 167
28 ABBREVIATIONS
The following abbreviations may be used in this Report.
Abbreviation Terms or Units
μg/m³ Microgram per Cubic Metre μm Microns, Micrometre ' Feet " Inch $ Dollar Sign $/m² Dollar per Square Metre $/m³ Dollar per Cubic Metre $/t Dollar per Metric Tonne % Percent Sign % w/w Percent Solid by Weight ¢/kWh Cent per Kilowatt hour ° Degree °C Degree Celsius 2D Two Dimensions 3D Three Dimensions
AAS Atomic Absorption Spectroscopy AI Abrasion Index Actlabs Activation Laboratories Ltd. AMSL Above Mean Sea Level ARD Acid Rock Drainage ASL Above Sea Level AWG American Wire Gauge az Azimuth
bank Bank Cubic Metre
BDF Bulk Density Factors BFA Bench Face Angle BIF Banded Iron Formation BIGS BIGS Global Burkina
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Abbreviation Terms or Units
BOF Basic Oxygen Furnace BQ Drill Core Size (3.65 cm diameter) BRGM Bureau de recherches géologiques et minières BSG Bulk Specify Gravity BTU British Thermal Unit BUMIFOM Bureau minier de la France Outre-Mer BWI Bond Ball Mill Work Index
CAD Canadian Dollar CAPEX Capital Expenditures CDE Canadian Development Expenses CDP Closure and Decommissioning Plan Ce Cesium cfm Cubic Feet per Minute CFR Cost and Freight Cg Graphitic CIF Cost Insurance and Freight CIL Carbon in Leach CIM Canadian Institute of Mining, Metallurgy and Petroleum CIP Carbon in Pulp Cl Clay CL Concentrate Leach cm Centimetre CNT Conseil National de transition Co Cobalt CO Carbon Monoxide
CO₂ Carbon Dioxide COG Cut Off Grade COV Coefficient of Variation CRM Certified Reference Materials CTAE Comité Technique d’Analyse Environnementale CTMP Centre de Technologie Minérale et de Plasturgie Inc. Cu Copper
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Abbreviation Terms or Units
CWI Crusher Work Index
d Day d/w Days per Week d/y Days per Year D2 Second Generation of Deformation D3 Third Generation of Deformation D4 Fourth Generation of Deformation dB Decibel dBA Decibel with an A Filter DDH Diamond Drill Hole deg Angular Degree DEM Digital Elevation Model DGPS Differential Global Positioning System DMS Dense Media Separator DT Davis Tube DTM Digital Terrain Model DWI Drop Weight Index DWT Drop Weight Test DXF Drawing Interchange Format
E East EA Environmental Assessment EAB Environmental Assessment Board EAF Electric Arc Furnace EBS Environmental Baseline Study EDS Energy-dispersive X-ray spectroscopy EHS Environment Health and Safety EIA Environmental Impact Assessment EIS Environmental Impact Statement EM Electromagnetic EMP Environmental Management Plant EOH End of Hole EP Environmental Permit
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Abbreviation Terms or Units
EPA Environmental Protection Agency EPCM Engineering, Procurement and Construction Management EQA Environmental Quality Act ER Electrical Room ESBS Environmental and Social Baseline Study ESIA Environmental and Social Impact Assessment
FA Fire Assay FOB Free on Board ft Feet
g Grams G&A General and Administration g/l Grams per Litre g/t Grams per Tonne gal Gallons GCP Ground Control Points GCW Gross Combined Weight GEMS Global Earth-System Monitoring Using Space GOH Gross Operating Hours GPS Global Positioning System Gr Granular
H Horizontal h Hour h/d Hours per Day h/y Hour per Year H₂ Hydrogen ha Hectare HDPE High Density PolyEthylene HF Hydrofluoric Acid HFO Heavy Fuel Oil HG High Grade
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Abbreviation Terms or Units
HL Heavy Liquid hp Horse Power HQ Drill Core Size (6.4 cm Diameter) HVAC Heating Ventilation and Air Conditioning Hz Hertz
I/O Input / Output ICP-AES Inductively Coupled Plasma – Atomic Emission Spectroscopy ICP-MS Inductively Coupled Plasma – Mass Spectroscopy ICP-OES Inductively Coupled Plasma – Optical Emission Spectroscopy ID Identification IDW Inverse Distance Method IDW2 Inverse Distance Squared Method IFC International Finance Corporation In Inches IR Infrared Radiation IRA Inter-Ramp Angle IRR Internal Rate of Return IT Information Technology
JORC Joint Ore Reserves Committee JV Joint Venture
KE Kriging Efficiency kg Kilogram kg/l Kilogram per Litre Kg/t Kilogram per Metric Tonne kl Kilolitre km Kilometre km/h Kilometre per Hour kPa Kilopascal KSR Kriging Slope Regression kt Kilotonne kV Kilovolt
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Abbreviation Terms or Units
kVA Kilovolt Ampere kW Kilowatt kWh Kilowatt-hour kWh/t Kilowatt-hour per Metric Tonne
L Line l Litre l/h Litre per hour lbs Pounds LFO Light Fuel Oil LG Low Grade LG-3D Lerchs-Grossman – 3D Algorithm Li Lithium LIMS Laboratory Information Management Systems LIMS Low Intensity Magnetic Separator LPA Lumière polarisée analysée LPNA Lumière polarisée non-analysée LOI Loss on Ignition LOM Life of Mine LV Low Voltage
m Metre m/h Metre per Hour m/s Metre per Second m² Square Metre m³ Cubic Metre m³/d Cubic Metre per Day m³/h Cubic Metre per Hour m³/y Cubic Metre per Year mA Milliampere MCC Motor Control Center MEB Microscopie électronique à balayage mg/l Milligram per Litre MIBK Methyl Isobutyl Ketone
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Abbreviation Terms or Units
min Minute min/h Minute per Hour Min/shift Minute per Shift ml Millilitre ML Metal Leaching MLA Mineral Liberation Analyzer mm Millimetre mm/d Millimetre per Day Mm³ Million Cubic Metres MMER Metal Mining Effluent Regulation MMU Mobile Manufacturing Units MOLP Multiple Objective Linear Programming MOU Memorandum of Understanding Mt Million Metric Tonnes Mt/y Million of Metric Tonnes per year MV Medium Voltage MVA Mega Volt-Ampere MW Megawatts MWh/d Megawatt Hour per Day My Million Years
N North NAG Non-Acid Generating NaCN Sodium cyanide Nb Number NE Northeast NFPA National Fire Protection Association NGR Neutral Grounding Resistor Ni Nickel NI National Instrument Nm³/h Normal Cubic Metre per Hour NPV Net Present Value NQ Drill Core Size (4.8 cm diameter) NSR Net Smelter Return
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Abbreviation Terms or Units
NTP Normal Temperature and Pressure NTS National Topographic System NW North West
O/F Overflow OB Overburden OK Ordinary Kriging OPEX Operating Expenditures OREAS Ore Research and Exploration Pty Ltd. oz Ounce (troy) oz/t Ounce per Short Ton
P&ID Piping and Instrumentation Diagram PAG Potential Acid Generating PAPs Project Affected Persons Pd Palladium PEA Preliminary Economic Assessment PF Power Factor PFS Pre-Feasibility Study PGE Platinum-Group Element PGGS Permit for Geological and Geophysical Survey ph Phase (electrical) pH Potential Hydrogen PIR Primary Impurity Removal PLC Programmable Logic Controllers PP Preproduction ppb Part per Billion ppm Part per Million PR Permis de recherche psi Pounds per Square Inch Pt Platinum P-T Pressure-Temperature PVC Polyvinyl Chloride
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project / Page 175
Abbreviation Terms or Units
QA/QC Quality Assurance/Quality Control QKNA Quantitative Kriging Neighbourhood Analysis QP Qualified Person
RAP Resettlement Action Plans RCMS Remote Control and Monitoring System RER Rare Earth Magnetic Separator RMR Rock Mass Rating ROM Run of Mine rpm Revolutions per Minute RQD Rock Quality Designation RWI Bond Rod Mill Work Index
S South S Sulfur S/R Stripping Ratio SAG Semi-Autogenous Grinding SANAS South African National Accreditation System Sc Scandium scfm Standard Cubic Feet per Minute SCIM Squirrel Cage Induction Motors SE South East sec Second SEM Scanning Electronic Microscope Set/y/unit Set per Year per Unit SG Specific Gravity SGS-Lakefield SGS Lakefield Research Limited of Canada SIR Secondary Impurity Removal SMC SAG Mill Comminution SNRC Système National de Référence Cartographique SPI SAG Power Index SPLP Synthetic Precipitation Leaching Procedure SPT Standard Penetration Tests SR Stripping Ratio
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Abbreviation Terms or Units
SW South West SW Switchgear
t Metric Tonne t/d Metric Tonne per Day t/h Metric Tonne per Hour t/h/m Metric Tonne per Hour per Metre t/h/m² Metric Tonne per Hour per Square Metre t/m Metric Tonne per Month t/m² Metric Tonne per Square Metre t/m³ Metric Tonne per Cubic Metre t/y Metric Tonne per Year Ta Tantalum TCLP Toxicity Characteristic Leaching Procedure TER Travail d’Études et de Recherches TIN Triangulated Irregular Network ton Short Ton tonne Metric Tonne ToR Terms of Reference TSS Total Suspended Solids
U Uranium U/F Under Flow ULC Underwriters Laboratories of Canada USA United Stated of America USD United States Dollar USGPM Us Gallons per Minute UTM Universal Transverse Mercator
V Vertical V Volt VAC Ventilation and Air Conditioning VFD Variable Frequency Drive VLF Very Low Frequency
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Abbreviation Terms or Units
W Watt W West WAC West African Archean Craton WHIMS Wet High Intensity Magnetic Separation WHO World Health Organization WRA Whole Rock Analysis Method WSD World Steel Dynamics w/v Weight/volume wt Wet Metric Tonne
X X Coordinate (E-W) XRD X-Ray Diffraction XRF X-Ray Fluorescence
y Year Y Y coordinate (N-S)
Z Z coordinate (depth or elevation) Zr Zirconium
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29 CERTIFICATES OF QUALIFIED PERSONS
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final // DRA Americas, Inc.
555 René-Lévesque Blvd West / 6th floor / Montréal / Quebec / Canada / H2Z 1B1 T +1 514 288-5211 / E [email protected] / met-chem.com
CERTIFICATE OF AUTHOR
To accompany the Report entitled “NI 43-101 Technical Report – Preliminary Economic Assessment for the Nabanga Project” which is effective as of September 30, 2019 and issue on November 14, 2019 (the “Technical Report”) prepared for SEMAFO Inc. (the “Company”).
I, Patrick Pérez, P. Eng., do hereby certify:
1. I am Project Manager and Senior Mining Engineer with Met-Chem, a division of DRA Americas Inc, with an office at 555 René-Lévesque Blvd. West, 6th Floor, Montreal, Canada.
2. I am a graduate from “Ecole Nationale Supérieure de Géologie de Nancy”, in France, with a M.Sc. in Geological Engineering obtained in 2003.
3. I am a registered member of APEGS (Association of Professional Engineers and Geoscientists of Saskatchewan), membership #16131.
4. I have worked as a Senior Mining Engineer or Project Manager continuously since my graduation from university. I have been employed since my graduation in 2003. I have gained relevant experience on deposits similar to the Nabanga Project, including:
a) Work on gold mining operations and projects hosted in Birimian greenstone belt formations in West Africa, in both Open Pit and Underground environment; b) I have also participated and supervised several mineral resource estimates or engineering studies for different projects at various stages of development. Hands-on experience for gold in Ivory Coast, Mali, Burkina Faso, and Canada; c) Design, supervision and implementation of mining programs; d) Review, audits, interpretation of geoscientific data; e) Experience in on several projects in weathered terranes under tropical conditions (West Africa and New Caledonia); and f) Participation in the preparation of parts of NI 43-101 compliant Technical Reports.
5. I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be an independent qualified person for the purposes of NI 43-101;
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6. I am independent of the issuer applying all of the tests in section 1.5 of NI 43-101.
7. I have participated in the preparation of this Technical Report and am responsible for Sections 2, 3, 15, 16, 18 to 20, 21 (except for 21.2.5), 22, and 24 and parts of Sections 1 and 25 to 27.
8. I have not visited the property site.
9. I have had no prior involvement with the property that is the subject of the Technical Report.
10. I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this Report;
11. Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from the Company, or any associated or affiliated companies;
12. I have read NI 43-101 and Form 43-101F1 and have prepared the Technical Report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with generally accepted Canadian mining industry practice, and as of the date of the 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.
Dated this 14th day of November 2019
“Original Signed and sealed” Patrick Pérez, P. Eng. Project Manager
/ Page 2 of 2 // DRA Americas, Inc.
555 René-Lévesque Blvd West / 6th floor / Montréal / Quebec / Canada / H2Z 1B1 T +1 514 288-5211 / E [email protected] / met-chem.com
CERTIFICATE OF QUALIFIED PERSON
To accompany the Report entitled “NI 43-101 Technical Report – Preliminary Economic Assessment for the Nabanga Project” which is effective as of September 30, 2019 and issued on November 14, 2019 (the “Technical Report”) prepared for SEMAFO Inc. (the “Company”).
I, Yves A. Buro, P. Eng., do hereby certify: 1. I am a Senior Engineer with Met-Chem, a division of DRA Americas Inc, with an office at 555 René-Lévesque Blvd. West, 6th Floor, Montreal, Canada. 2. I am a graduate of University of Geneva, Switzerland, with the equivalent of a B.Sc. ("Licence ès Sciences de la Terre") and an M.Sc. ("Diplôme d'Ingénieur-Géologue") in Geology obtained in 1975 and 1976. 3. I am a registered member of the Ordre des ingénieurs du Québec (OIQ), membership # 42279. 4. I have worked continuously in mineral exploration and production since my graduation from University. Experience relevant to the Nabanga Project includes hands-on involvment with more than 30 gold projects at all phases of development: x Surface exploration: o Mapping, exploration (Canada, Algeria, Mali, Guinea, Ghana, Ivory Coast, Central African Republic, Gabon, Colombia); o Drilling, data interpretation (Canada, Mali, Algeria, Ghana, Mauritania); x Underground mine exploration, development and prodution: o Mapping, interpretation, drilling (Canada) and sampling (Mali, Colombia); o Mine development (Canada, Burkina Faso); mine geologist (Canada – 3.5 years); x Project audits, evaluation: o QP site visit (exploration projects, active mines) and participation in the preparation of NI 43-101 reports (Canada, Burkina Faso, Mali, Mongolia); o Due diligence (Mauritania), opportunity studies (Canada, Cameroon); x Metallurgical sampling: drilling to generate gold mineralized material (Mali). 5. I have read the definition of "qualified person" set out in National Instrument 43Ǧ101 (NI 43Ǧ
/ Page 1 of 2
101) and certify that by reason of my education, affiliation with a professional association (as defined by NI 43Ǧ101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43Ǧ101. 6. I am independent of the issuer applying all of the tests in section 1.5 of NI 43Ǧ101. 7. I have contributed in Sections 4 to 12 inclusively, and parts of Sections 1 and 25 to 27 in the preparation of this Technical Report. 8. I have not visited the property site. 9. I have had no prior involvement with the property that is the subject of the Technical Report. 10. I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this Report; 11. Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from the Company, or any associated or affiliated companies; 12. I have read NI 43-101 and Form 43-101F1 and have prepared the Technical Report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with generally accepted Canadian mining industry practice, and as of the date of the 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.
Dated this 14th day of November 2019
“Original document signed and sealed”
“Yves A. Buro”
Yves A. Buro, P. Eng.
/ Page 2 of 2 // DRA Americas, Inc.
555 René-Lévesque Blvd West / 6th floor / Montréal / Quebec / Canada / H2Z 1B1 T +1 514 288-5211 / E [email protected] / met-chem.com
CERTIFICATE OF AUTHOR
To accompany the Report entitled “NI 43-101 Technical Report – Preliminary Economic Assessment for the Nabanga Project” which is effective as of September 30, 2019 and issue on November 14, 2019 (the “Technical Report”) prepared for SEMAFO Inc. (the “Company”).
I, Ewald Pengel, M.Sc, P.Eng., Quebec, do hereby certify that:
1. I am a Senior Process Engineer with Met-Chem, a division of DRA Americas Inc, with an office at 555 Rene-Levesque Blvd. West, 6th Floor, Montreal, Canada;
2. I am a graduate from Queen's University, Kingston, Ontario with a B.Sc. in Metallurgical Engineering in 1982 and the University of Pittsburgh, Pittsburgh, Pennsylvania (USA) with a M.Sc. in Mining Engineering in 1985;
3. I am a registered member of Professional Engineers Ontario (# 90520297) and I am a member of the Canadian Institute of Mining Metallurgy and Petroleum.
4. I have worked for more than 30 years in the mining industry since my graduation from University. I have gained relevant experience on deposits similar to the Nabanga Project, including hands-on experience in froth flotation, fine grinding and gold in Canada, the USA and South America.
5. I have read the definition of “qualified person” set out in the National Instrument 43-101 and certify that by reason of my education, affiliation with a professional association and past relevant work experience, I fulfil the requirements to be an independent qualified person for the purposes of NI 43-101;
6. I am responsible for Sections 13 and 17 and contributed part of Sections 1, 21, and 25 to 27 of the Technical Report;
7. I have not visited the property site.
8. I have had no prior involvement with the property that is the subject of the Technical Report.
9. I have no personal knowledge as of the date of this certificate of any material fact or change, which is not reflected in this Report;
/ Page 1 of 2 10. Neither I, nor any affiliated entity of mine, have earned the majority of our income during the preceding three (3) years from the Company, or any associated or affiliated companies; 11. I have read NI 43-101 and Form 43-101F1 and have prepared the Technical Report in compliance with NI 43-101 and Form 43-101F1; and have prepared the report in conformity with generally accepted Canadian mining industry practice, and as of the date of the 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.
Dated this 14th day of November 2019
“Original Signed and sealed” Ewald Pengel, M.Sc., P.Eng. Senior Process Engineer
/ Page 2 of 2
Certificate of Qualified Person for Richard Roy, P.Geo.
I, Richard Roy, P.Geo. of Verdun (Québec) Canada, do hereby certify that: a) I am Vice-President, Exploration at SEMAFO lnc. ("Semafo") with an office located at 100 Alexis-Nihon Blvd, 7th floor, Saint-Laurent (Quebec) Canada H4M 2P3; b) This certificate regards the technical report entitled “NI 43-101 Technical Report – Preliminary Economic Assessment for the Nabanga Project” which is effective as of September 30, 2019 and issued on November 14, 2019 (the “Technical Report”); c) I am a graduate from Concordia University (B.Sc. Geology, 1988); d) I am a member in good standing (#536) of the Ordre des Géologues du Québec (Order of Geologist of Quebec); e) I have worked as a geologist continuously since my graduation from university. I have over 30 years of experience in the base and precious metal mineral resource industry including nine years’ experience in underground mine geology. f) I am a "Qualified Person" for purposes of National Instrument 43-101 (the "Instrument”); g) I visited the Nabanga property between May 24 and 27 2018; h) I have participated in the preparation of this Technical Report and am responsible for Sections 4 to 10 and 23 and parts of Sections 1, 11, and 27 of the Technical Report; i) I am not independent of Semafo as described in section 1.5 of the Instrument; j) I work for Semafo and have prior and current involvement with the property that is the subject of the Technical Report; k) I have read the Instrument and the sections of the Technical Report that I am responsible for which have been prepared in compliance with the Instrument; and l) As of the effective date of the Technical Report, to the best of my knowledge, information, and belief, the parts of the Technical Report that I am responsible for, contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. ______Signed and dated this 14th day of November 2019 at Montréal, Québec.
(signed) Richard Roy Richard Roy, P.Geo. Vice-President, Exploration SEMAFO Inc.
Téléphones : 514 744-4408 SEMAFO inc. 1-888-744-4408 100, boul. Alexis-Nihon e Télécopieur : 514 744-2291 7 étage Saint-Laurent (Québec) H4M 2P3
Certificate of Qualified Person for François Thibert, MSc, P.Geo.
I, François Thibert, MSc, P. Geo. of Montréal (Québec) Canada, do hereby certify that: a) I am Manager, Estimate Group Resources and Reserves at SEMAFO lnc. ("Semafo") with an office located at 100 Alexis-Nihon Blvd, 7th floor, Saint-Laurent (Quebec) Canada H4M 2P3; I am the Corporation’s “qualified person”; b) This certificate regards the technical report entitled “NI 43-101 Technical Report – Preliminary Economic Assessment for the Nabanga Project” which is effective as of September 30, 2019 and issued on November 14, 2019 (the “Technical Report”); c) I am a graduate from HEC Montreal (Quebec) (Certificate in Project Management in 2010), University of Montreal (Quebec) (Master of Science (M.Sc.) in Igneous Petrology (1993) and Bachelor of Science (B.Sc.) in Geology (1987)); d) I am a member in good standing (#1444) of the Ordre des Géologues du Québec (Order of Geologist of Quebec); e) I have worked as a geologist continuously since my graduation from university. I have more than 25 years’ experience in mining exploration in gold, and base metal projects across Canada and worldwide. I have participated and supervised several mineral resource estimates for different exploration projects at various stages of exploration. f) I am a "Qualified Person" for purposes of National Instrument 43-101 (the "Instrument”); g) I have never visited the Nabanga property; h) I have participated in the preparation of this Technical Report and am responsible for Section 12 and 14 and parts of Sections 1, 11, and 27 of the Technical Report; i) I am not independent of Semafo as described in section 1.5 of the Instrument; j) I work for Semafo and have prior and current involvement with the property that is the subject of the Technical Report; k) I have read the Instrument and the sections of the Technical Report that I am responsible for which have been prepared in compliance with the Instrument; and l) As of the effective date of the Technical Report, to the best of my knowledge, information, and belief, the parts of the Technical Report that I am responsible for, contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. ______Signed and dated this 14th day of November 2019 at Montréal, Québec.
(signed) François Thibert François Thibert, MSc, P.Geo. Manager, Estimate Group Resources and Reserves SEMAFO Inc.
Téléphones : 514 744-4408 SEMAFO inc. 1-888-744-4408 100, boul. Alexis-Nihon e Télécopieur : 514 744-2291 7 étage Saint-Laurent (Québec) H4M 2P3
SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project
Appendix A – Process Design Criteria
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final Process Design Criteria
Project Code: C3852
Project Name: Preliminary Economic Assessment, Nabanga Project
Document No.: C3852-PROC-DC-002
Client : SEMAFO Inc.
Process Engineer : Ewald Pengel, P.Eng.
Issue Date : 20 September 2019
Revision : B
APPROVALS
DISCIPLINE LEAD: Ewald Pengel, Senior Metallurgist
APPROVED BY (DRA): Patrick Perez, Project Manager
APPROVED BY (CLIENT):
Revision Description Revision No. Date Status
Internal DRA review A 14 août 2019 Complete Issued for report B 20 septembre 2019 Complete C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
1.0 TABLE OF CONTENTS, SOURCE CODES, TERMINOLOGY AND UNITS
1.1 Table of Contents PAGE 1.0 TABLE OF CONTENTS, SOURCE CODES, TERMINOLOGY AND UNITS 2 2.0 GENERAL PROCESSING DESIGN CRITERIA 5 3.0 CRUSHING CIRCUIT DESIGN CRITERIA 8 4.0 GRINDING CIRCUIT DESIGN CRITERIA 10 5.0 SULFIDE FLOTATION CIRCUIT DESIGN CRITERIA 13 6.0 INTENSIVE CYANIDE LEACHING CIRCUIT DESIGN CRITERIA 16 7.0 CYANIDE DESTRUCTION CIRCUIT DESIGN CRITERIA 19 8.0 CONSUMABLES AND REAGENTS CIRCUIT DESIGN CRITERIA 20 9.0 PLANT UTILITIES 24 10.0 MASS BALANCE 26 11.0 WATER BALANCE 30
1.2 Source of Process Design Criteria A = Semafo Inc. B = DRA Recommendation C = DRA Calculation D = DRA Layout E = Laboratory test work and pilot plant results F = I = Standard Industry Practice J = Map L = Literature M = Meteorological data O = Other operation references R = Supplier recommendation S = Supplier documents T = Test work documents X = To be determined
1.3 Terminology 1.3.1 Nominal Used in this context, nominal is a synonym for "normal" or "expected". The values displayed in the mass balance are nominal values. These values are the expected tonnages when the plant is fully operational and normal operating conditions.
1.3.2 Design These values are used to size equipment. Normally these values are nominal values with a design factor. In some cases these values are based off client needs.
1.3.3 Operating Time Crushing plant operating time (measured in hours) is the period that the crusher discharge conveyor operates under normal load. Processing plant operating time (measured in hours) is the period that the SAG Mill feed conveyor operates under normal load.
1.3.4 Operating Percentage Operating percentage is operating time period divided by the calendar time period expressed in a percentage.
1.3.5 Weight recovery All weight recoveries are based on New Feed, unless stated otherwise.
1.3.6 New Feed New Feed is the ore placed on the SAG Mill feed conveyor material by the Apron Feeder underneath the stockpile.
1.4 Colour coding 1.4.1 Black Final information from either test work, supplier, DRA or client experience.
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1.4.2 Blue Verified information. This is usually based on preliminary information or basic calculations.
1.4.3 Red Unverified information, entry is best guess or based on database.
1.4.4 Red with yellow background Unverified information, which may be not to be used by other groups
1.5 Units 1.5.1 Units usage SI units are used as a base. More conventional units in process engineering are used as well as some metric units. Standard SI formatting for numbers and values has been used.
1.5.2 Unit symbols °C = Degree Celsius % = Percent % v/v = Percent material by volume % w/w = Percent solid by weight AMSL = Above Mean Sea Level d = Day d/w = Days per week d/y = Days per year deg = Angular degree E = East g = Gram g/d = Grams per day g/L = Grams per litre g/t = Grams per tonne h = Hour h/d = Hours per day hp = Horsepower Hz = Hertz in = Inch km/h = Kilometre per hour kPa = Kilopascal (103 kilogram per square metre) kW = Kilowatt kWh/t = Kilowatt hour per metric tonne m = Metre m² = Square metre m³ = Cubic metre m³/d = Cubic metre per day m³/h = Cubic metre per hour
m³/m²•h¯¹ = Cubic metre per square metre per hour micron = Micrometre min = Minute mm = Millimetre m/s = Metre per second MW = Molecular weight N = North Nm³/h = Normal cubic metre per hour ph = Phase (electrical) pH = Liquid acidity, or the negative of the logarithm to base 10 of the molar concentration of hydrogen ions. psi = Pounds per square inch
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RPM = Revolutions per minute SG = Specific gravity (or relative density) t = Metric tonne T = Tesla ( 1 Tesla = 10 000 Gauss) t/m² = Metric tonne per square metre
t/m²•h¯¹ = Metric tonne per square metre per hour t/m³ = Metric tonne per cubic metre t/y = Metric tonne per year t/d = Metric tonne per day t/h = Metric tonne per hour V = Volt y = Year B = Major item or design criteria updated for Revision B
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 2.0 GENERAL PROCESSING DESIGN CRITERIA 2.0.1 General Design Base Design production rate (dry) t/y 360 000 A Throughput = 1080 t/d Operating days per year d/y 333 C Operating days per week d/w 7 A Operating hours per day h/d 24 I
2.1 Project Geography and Weather 2.1.1 Property Location Plant location South-east Burkina Faso J Site Elevation m AMSL 200 J Average longitude N 11° 18' 03" J Average latitude E 0° 29' 58" J
2.1.2 Climate at Ouagadougou Minimum temperature °C 16.0 15.0 M Maximum temperature °C 40.0 40.0 M Maximum relative humidity % 67% 75.0% M
2.2 Ore Characteristics 2.2.1 Design Ore Grades Gold Au g/t 6.5 7.5 E Silver Ag g/t 3.0 4.1 A
2.2.2 Ore Specifications Design ore dry specific gravity 2.78 2.78 E ROM porosity factor % 0.0% 0.0% B Estimated Apparent density t/m³ 2.78 2.78 C ROM swell factor % 57% 60% B & L Estimated Dry bulk density t/m³ 1.77 1.74 C Moisture in ore (assumed) % w/w 5.0% 6.0% B Estimated Wet bulk density t/m³ 1.86 1.84 C Angle of repose deg 37.0 37.0 L
2.2.3 Ore Physical competency specifications and indices Abrasion index (Ai) g 0.308 E JK - breakage resistance of large particles (A) 77.5 E JK - breakage resistance of small particles (b) 0.46 E
JK - resistance abrasion grinding (ta) 0.33 E SMC test mill work index (Mic) kWh/t 8.35 E JK - drop-weight index (DWi) kWh/m³ 7.74 E SMC test mill work index (Mia) kWh/t 21.19 E SMC test mill work index (Mih) kWh/t 16.14 E Bond ball mill work index (BWi) kWh/t 23.20 E
2.2.4 Design Recoveries Nominal Design Orbis Gold Ltd., Report No: INT-RPT-047 - 2015-01-09; Rev 2g Sulfide flotation - gold recovery % 78% 79% E pg. 80
Sulfide concentrate leach - gold recovery % 94% 95% E pg. 78, at P80 < 10 micron
Oxide tailings leach - gold recovery % 88% 90% E pg. 78, at P80 < 53 micron Overall- gold recovery % 93% 94% C
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 2.3 Processing Criteria 2.3.1 General Processing Facility Design Criteria Daily processing facility rate (dry average) t/d 1 080 1 404 C 30% design factor Crusher circuit operating percentage h 6 000 A Concentrator operating percentage h 8 000 A Orbis Gold Ltd., Report No: INT-RPT-047 - 2015-01-09; Rev 2g, pg. 82 Crusher circuit operating percentage % 68.5% C Concentrator operating percentage % 91.3% C
2.3.2 Equipment Sizing Criteria Crushing plant equipment design factor % 30% B Concentrator plant equipment design factor % 25% B Concentrator plant slurry pump design factor % 10% B
2.3.3 Crusher Area Criteria Crusher circuit average hrs operating per day h/d 16.4 C Crushing circuit rate (wet) t/h 65.7 85.4 C 30% design factor
Crusher - Feed F100 mm 636 C Diagonal from grizzly opening
Crusher - Feed F80 mm 310 S From simulation
Crusher design - Product P80 mm 93 S From simulation
2.3.4 Concentrator Area Criteria Concentrator average hrs operating per day h/d 21.9 C Nominal processing circuit rate (dry) t/h 49.3 61.6 C Percent passing through grizzly feeder of New Feed % 37.5% C & S From simulation
2.3.5 Grinding Circuit Criteria
Primary grinding - Feed F80 mm 93 S From simulation
Primary grinding - Transfer T80 mm 1.80 B
Secondary grinding design - Product P80 mm 0.075 0.075 E Gravity circuit weight recovery (New Feed) % 1.05% 1.31% E
2.3.6 Sulfide Flotation Circuit Criteria
Flotation - Feed F80 mm 0.075 0.094 E Rougher concentrate weight recovery (based on fresh feed) % 15.0% 18.8% E
Regrind mill - Feed F80 mm 0.075 0.094 E
Regrind mill - Product P80 mm 0.020 0.015 E Cleaner concentrate weight recovery (New Feed) % 7.5% 9.4% B & E Cleaner concentrate weight recovery (New Feed) % 5.0% 6.3% B & E
2.3.7 Concentrate Intensive Leaching Circuit Criteria Intensive cyanide leach time h 20.0 25.0 E Leach recovery 86% Sulfide concentrate thickener underflow solids %w/w 70.0% 70.0% B
2.3.8 Cyanide Leaching Circuit Criteria Cyanide leach time h 50.0 62.5 E Leach recovery 95% Pre-leach thickener underflow solids %w/w 52.0% 50.0% B
2.3.9 Cyanide Destruction B Not required A E-Mail from S.Duchesne 2019-08-15 Cyanide destruction time Cyanide destruction percent solids
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 2.4 Utility Specifications 2.4.1 Water Requirements Raw water source Collection basins A E-Mail from S.Duchesne 2019-08-15 Average daily raw water required m³/d 222 244 B & C 10% design factor Raw water specific gravity 1.00 1.00 I Raw water solids density % w/w 0% 0% I Process water specific gravity 1.00 1.00 I Process water solids density % w/w 0% 0% I Process water recycling rate % 70% 70% B
2.4.2 Air Requirements High pressure air pressure kPa 800 B Plant air and flotation column High pressure air volume Nm³/h X Low pressure air pressure kPa 180 250 B Conventional cleaner scavenger cells Low pressure air volume Nm³/h 1 178 1 470 C & R
2.4.3 Electrical Requirements B Power source Locally generated A E-Mail from S.Duchesne 2019-08-15 High voltage V 11 000 A E-Mail from S.Duchesne 2019-08-15 Medium voltage V 6 600 A E-Mail from S.Duchesne 2019-08-15 Low voltage V 690 X E-Mail from S.Duchesne 2019-08-15 Low voltage V 380 X E-Mail from S.Duchesne 2019-08-15 Phase ph 3B Frequency Hz 50 B
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 3.0 CRUSHING CIRCUIT DESIGN CRITERIA 3.0.1 General Crushing circuit throughput rate (dry) t/h 65.7 85.4 B & C 30% design factor Crushing circuit throughput rate (wet) t/h 69.9 90.9 B & C 30% design factor ROM moisture % w/w 6.0% C Ore specific gravity 2.78 E
3.1 Primary Crushing Circuit 3.1.1 Rock Breaker Type Hydraulic B Feed top size to the crusher mm 1 000 1 300 B & C 30% design factor Power installed kW B
3.1.2 Grizzly Type Two inch bars, wedge cover wear plates B Feed top size to the crusher mm 636 636 C Diagonal from grizzly opening Grizzly opening Width mm 450 B Length mm 450 B Grizzly size Length m D By Met-Chem designer Width m D By Met-Chem designer
3.1.3 Crusher Feed Hopper Type Steel chute B Truck size t 45.0 58.5 B & C 30% design factor Crusher feed hopper capacity truck loads 2.0 2.0 B Crushed ore bulk density t/m³ 1.84 1.84 C Angle of repose deg 37.0 37.0 L Hopper capacity t 90.0 117.0 C Hopper volume m³ 48.8 63.4 D Dust suppression (Yes or No) Yes Yes B
3.1.4 Vibrating Grizzly Feeder Type of feeder Vibrating B Number 1B Drive (variable speed or fixed speed) Variable speed B Feeder throughput rate (wet) t/h 69.9 90.9 C 30% design factor Wet ore bulk density t/m³ 1.84 1.84 C Angle of repose deg 37.0 37.0 L Feed top size to the crusher mm 636 636 C Diagonal from grizzly opening Type of feeder Bars as part of feeder B Grizzly opening mm 64 64 C & S From simulation Percent Passing less than 64 mm of New Feed % 37.5% 37.5% C & S From simulation Power installed kW R Dust suppression (Yes or No) Yes Yes B
3.1.5 Primary Crusher Type of crusher Jaw A Type C116 equivalent Number 1B Feed throughput rate (wet) t/h 43.7 56.8 B & C 30% design factor Crusher closed setting mm 100 100 B
Crusher feed – F80 mm 361 361 C & S From simulation
Crusher product – P80 mm 105 105 C & S From simulation
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS Reduction ratio 3.44 3.44 C % of jaw crusher product less than 20 mm % 13.0% 13.0% C & S From simulation Power installed kW S Dust suppression (Yes or No) Yes Yes B
3.1.6 Sacrificial Conveyor Conveyor type Belt B Feed throughput rate (wet) t/h 69.9 90.9 B & C 30% design factor
Total conveyor feed F100 mm 93 S From simulation Conveyor size Width mm R Power installed kW R Dust suppression (Yes or No) Yes Yes B
3.1.8 Stockpile Conveyor Conveyor type Belt B Feed throughput rate (wet) t/h 69.9 90.9 B & C 30% design factor
Total crusher product P100 mm 93 C & S From simulation Conveyor size Width mm R Power installed kW R Dust suppression (Yes or No) Yes Yes B
3.2 Crushing Circuit Auxiliaries 3.2.1 Jaw Crusher Monorail Type of crane Monorail B Number of cranes 1B Required lift capacity t 25 B Required span m D By Met-Chem designer Required lifting height m D By Met-Chem designer Power installed kW R
3.2.2 Tramp Metal Magnet Monorail Type of crane Monorail B Number of cranes 1B Required lift capacity t 5.0 B Required span m D By Met-Chem designer Required lifting height m D By Met-Chem designer
3.2.3 Tramp Metal Magnet Type Oil-cooled manual cleaning magnet B Self cleaning (Yes or No) No B Location At crusher discharge tail pulley B Magnet field strength Gauss 815 O Power installed kW 6.5 B
3.2.4 Crusher Area Sump Pump Type Vertical cantilever pump B Number 1B Drive Fixed speed drive B Power installed kW R
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 4.0 GRINDING CIRCUIT DESIGN CRITERIA 4.0.1 General Nominal hourly processing circuit rate (dry) t/h 49.3 61.6 B & C 25% design factor Design feed moisture % w/w 6.0% B Design ore specific gravity 2.78 E
4.1 Primary Milling Circuit 4.1.1 Coarse Ore Stockpile Live capacity (design) h 24.0 24.0 A To be verified Live wet capacity (design) t 1 258 1 640 B & C 30% design factor Number of stockpile 1 1 B ROM Moisture % w/w 6.0% 6.0% A Wet bulk density t/m³ 1.84 1.84 C Angle of repose deg 37.0 37.0 B Stockpile volume m³ 682 889 C Stockpile bottom diameter m D By Met-Chem designer Stockpile height m D By Met-Chem designer Covered or Open Open Open B
4.1.2 SAG Mill Feeder Type of feeder Apron B Number of feeders 2B Drive (variable speed or fixed speed) Variable speed B Feeder nominal capacity (wet) t/h 52.4 65.5 B & C 25% design factor Wet ore bulk density t/m³ 1.84 1.84 C
Design feed size (F100) mm 183 183 C & S From simulation Feeder size Width mm D By Met-Chem designer Length mm D By Met-Chem designer Power installed per unit kW R Total installed power kW C Dust suppression (yes or no) Yes Yes B
4.1.3 SAG Mill Feed Conveyor Type of conveyor Belt B Number of conveyors 1 B Conveyor capacity (wet) t/h 74.6 93.2 B & C 25% design factor
Design feed size (F100) mm 183 C & S From simulation Conveyor size Width mm R Height differential m D By Met-Chem designer Conveyor horizontal length m D By Met-Chem designer Power installed kW R
4.1.4 Primary Grinding Mill Type of mill Semi-Autogenous Grinding B Number of mills 1A Drive (variable speed or fixed speed) Variable speed B Fresh feed throughput rate (dry) t/h 49.3 61.6 B & C 25% design factor Discharge pulp solids density % w/w 76.0% 76.0% B SAG mill feed water m³/h 14.0 14.0 C Discharge trommel wash water m³/h 3.0 3.0 B
Design fresh feed size (F80) mm 93 93 S
Design transfer size (T80) mm 1.80 1.80 B
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS Mill size Diameter m R Length m R Power installed kW R
4.1.5 SAG Mill Screen Type of screen Single deck horizontal, vibrating R Number of screens 1B Screen throughput rate (dry) t/h 56.7 70.8 B & C 25% design factor Percentage oversize of SAG mill feed (New Feed) % 15.0% 15.0% B Screen deck oversize specific gravity 2.757 2.757 B guestimated Screen oversize percent moisture % w/w 9.0% 9.0% B guestimated Screen wash water m³/h 5.0 5.0 B Number of decks 1 1 B Screen opening size mm 9 × 9 R squares
Final product size - passing (P80) mm 1.8 B Screen size Width m R Length m R Power installed kW R
4.2 Secondary Milling Circuit 4.2.1 Cyclone Feed Pump Pump type Horizontal centrifugal, metal B Number of pumps 2B Total discharge flow m³/h 218 240 B & C 10% design factor Average operating density % w/w 52.0% 52.0% C Power installed kW R
4.2.2 Secondary Cyclones Cyclone type High gravity B Number of cyclones 8AE-Mail from S.Duchesne 2019-08-15 New feed throughput rate (dry) t/h 49.3 61.6 B & C 25% design factor Total cyclone feed throughput rate (dry) t/h 197.1 246.4 B & C 25% design factor Cyclone feed flow rate m³/h 272.3 340.4 B & C 25% design factor Proportion circulating load (based on new feed) % 300% 300% B Size (Diameter) mm 380 380 A & C To be confirmed Cyclone feed solids density % w/w 57.2% 57.2% C Cyclone overflow solids density % w/w 35.0% 35.0% B Cyclone underflow solids density % w/w 72.5% 72.5% B Cyclone underflow specific gravity 2.813 2.813 B
4.2.3 Secondary Grinding Mill Type of mill Ball F Number of mills 1B Drive (variable speed or fixed speed) Fixed speed A E-Mail from S.Duchesne 2019-08-15 Feed nominal throughput rate (dry) t/h 49.3 61.6 B & C 25% design factor Total feed entering ball mill t/h 147.8 C
Design new feed size - Feed (F80) mm 1.80 1.80 B
Design product size - Product (P80) mm 0.075 0.075 E Feed pulp solids density % w/w 72.0% 72.0% B Ball mill feed water m³ 1.4 1.8 B & C 25% design factor Discharge trommel wash water m³ 3.0 3.8 B & C 25% design factor
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS Mill size Diameter m R Length m R Power installed kW R
4.3 Grinding Circuit Auxiliaries 4.3.1 Apron Feeder Monorail Type of crane Monorail B Number of cranes 3B Required lift capacity t 1.0 B Required span m D By Met-Chem designer Required lifting height m D By Met-Chem designer
4.3.2 Grinding Circuit Area Overhead Crane B Type of crane Overhead B Number of cranes 1B Required crane capacity t 15 A E-Mail from S.Duchesne 2019-08-15 Required span m D By Met-Chem designer Required lifting height m D By Met-Chem designer
4.3.3 Automated Ball Feeder B Type Screw feeder A Number 0BBall bucket only recommended by; e-Mail from S.Duchesne 2019-08-15 Drive Timer B Size C Power installed kW C
4.3.4 Stockpile and Grinding Area Sump Pump Type Vertical cantilever pump B Number 2B Drive Fixed speed drive B Power installed kW R
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REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 5.0 SULFIDE FLOTATION CIRCUIT DESIGN CRITERIA 5.0.1 General Number of conditioning stages 2 B For rougher and cleaners only Number of rougher stages 1 B Rougher and scavenger combined Number of cleaner stages 3 B First cleaner, second cleaner and third cleaners Number of re-grinding stages 1 B After roughers Flotation cell aeration factor % 15.0% I For all flotation cells Flotation scale -up factor from lab tests 2.5 I For all flotation cells
5.1 Sulfide Rougher Flotation Circuit 5.1.1 Sulfide Rougher Flotation Circuit Type of cells Conventional R Number 6R10 m³ each Rougher throughput rate t/h 49.3 61.6 B 25% design factor Rougher design flow rate m³/h 109.2 136.5 C
Design feed size (F80) mm 0.075 0.094 E 0 Rougher feed pulp density %w/w 34.0% B Rougher flotation time min 30.0 E & I 2.5 × lab test time. Rougher weight recovery (New Feed) % 15.0% 18.8% E Rougher concentrate solids specific gravity 3.202 B Guestimate Rougher concentrate pulp density %w/w 40.0% B Rougher concentrate launder water m³/h 3.5 C
5.1.2 Sulfide Rougher Flotation Concentrate Pump Pump type Vertical tank, SRL B froth pump ae single pumps Number of pumps 1B Total discharge flow m³/h 16.9 18.5 C 10% design factor Average operating density % w/w 33.7% C
5.1.3 Sulfide Rougher Flotation Tailings Pump Pump type Horizontal centrifugal, SRL B Number of pumps 2B Total discharge flow m³/h 95.8 105.4 C 10% design factor Average operating density % w/w 34.2% C
5.2 Sulfide Cleaner Flotation Circuit 5.2.1 Re-grind Cyclone Feed Pump Pump type Horizontal centrifugal, SRL B Number of pumps 2B Total discharge flow m³/h 30.4 33.4 C 10% design factor Average operating density % w/w 49.9% C
5.2.2 Sulfide Regrind Cyclones Cyclone type High gravity B Number of cyclones 8AE-Mail from S.Duchesne 2019-08-15 Total cyclone feed throughput rate (dry) t/h 23.4 29.3 B 25% design factor Cyclone feed flow rate m³/h 30.4 38.0 C 25% design factor Proportion circulating load (based on new feed) % 200% B Size (Diameter) mm 150 A & C To be confirmed Cyclone feed water m³ 2.66 B Cyclone feed solids density % w/w 49.9% C Cyclone overflow solids density % w/w 33.0% B Cyclone underflow solids density % w/w 71.0% B Cyclone underflow specific gravity 3.523 B
C3852 Process Design Criteria and Mass Balance - Rev B Page 13 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS
5.2.3 Sulfide Regrind Mill Type of mill Tower mill R Variable speed Number of mills 1B Mill throughput rate (dry) t/h 7.4 9.2 C 25% design factor Total feed entering ball mill t/h 14.8 18.5 C
Design new feed size - Feed (F80) mm 0.075 0.094 E
Design product size - Product (P80) mm 0.020 0.015 E Feed pulp solids density % w/w 70.0% B Mill feed water m³/h 0.30 C
5.2.4 Sulfide First Cleaner Flotation Circuit Type of cells Conventional R Number 4R First cleaner throughput rate t/h 8.6 10.8 C 25% design factor First cleaner flow rate m³/h 20.2 25.2 C 25% design factor
Design feed size (F80) mm 0.020 0.015 E First cleaner feed pulp density %w/w 35.0% B First cleaner flotation time min 15.0 E & I 2.5 × lab test time. First cleaner flotation cell aeration factor % 15.0% I For cleaner flotation cells First cleaner concentrate wt. distribution (New Feed) % 7.5% 9.4% B & E First cleaner concentrate solids specific gravity 3.628 B Estimate First cleaner concentrate pulp density %w/w 32.0% B First cleaner concentrate launder water m³/h 2.0 C
5.2.5 Sulfide First Cleaner Tailings Pump Pump type Horizontal centrifugal, SRL B Number of pumps 2B Total discharge flow m³/h 14.9 16.4 C 10% design factor Average operating density % w/w 27.1% C Power installed kW B
5.2.6 Sulfide First Cleaner Concentrate Pump Pump type Vertical tank, SRL B Number of pumps 1B Total discharge flow m³/h 9.9 10.9 C 10% design factor Average operating density % w/w 29.3% C Power installed kW B
5.2.7 Sulfide Second and Third Cleaner Flotation Circuit Type of cells Conventional R Number 5B3 → 2nd & 2 → 3rd cleaner cells Final cleaner design capacity t/h 3.7 4.6 C 25% design factor Final cleaner flow rate m³/h 9.9 12.4 C 25% design factor Final cleaner feed pulp density %w/w 28.0% B Final cleaner flotation time min 15.0 E & I 3 × lab test time. Cleaner flotation cell aeration factor % 15.0% I For cleaner flotation cells Final cleaner concentrate wt. distribution (New Feed) % 5.00% 6.25% B & E Final cleaner concentrate solids specific gravity 3.701 B Estimate Final cleaner concentrate pulp density %w/w 20.0% B Final cleaner concentrate launder water m³/h 1.3 C Final cleaner volume m³ 3.6 C Installed power per cell kW R Total power installed kW R
C3852 Process Design Criteria and Mass Balance - Rev B Page 14 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS
5.2.8 Sulfide Second Cleaner Tailings Pump Pump type Horizontal centrifugal, SRL B Number of pumps 2B Total discharge flow m³/h 2.9 3.2 C 10% design factor Average operating density % w/w 32.4% C Power installed kW B
5.2.9 Sulfide Second Cleaner Concentrate Pump Pump type Vertical tank, SRL B Number of pumps 1B Total discharge flow m³/h 10.8 11.9 C 10% design factor Average operating density % w/w 26.8% C Power installed kW B
5.2.10 Sulfide Third Cleaner Concentrate Pump Pump type Vertical tank, SRL B Number of pumps 1B Total discharge flow m³/h 8.3 9.2 C 10% design factor Average operating density % w/w 24.3% C Power installed kW B
5.3 Sulfide Flotation Circuit Auxiliaries 5.3.1 Sulfide Flotation Area Sump Pump Type Vertical cantilever pump B Number 1B Drive Fixed speed drive B Power installed kW R
C3852 Process Design Criteria and Mass Balance - Rev B Page 15 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 6.0 INTENSIVE CYANIDE LEACHING CIRCUIT DESIGN CRITERIA 6.0.1 General Intensive leach time h 20.0 25.0 B & E Leach time h 50.0 62.5 B & E Effective volume factor % 92% B Leach solution pH 10.5 11.0 E
6.1 Sulfide Concentrate Dewatering Circuit 6.1.1 Sulfide Concentrate Thickener Type of thickener High rate B Number of thickeners 1 B Thickener throughput rate (dry) t/h 2.46 3.08 C 25% design factor Thickener feed flow rate m³/h 8.33 10.42 C 25% design factor Thickener feed pulp density %w/w 30.0% 25.0% B Thickener dilution water flow rate m³/h 0.0 0.0 C Thickener overflow pulp density % 0.0% 0.0% I Thickener underflow pulp density % 70.0% 70.0% C Thickener underflow solids specific gravity 3.701 C Total power installed kW R
6.1.2 Sulfide Concentrate Thickener Overflow Pump Pump type Horizontal centrifugal B Pumps back to the process water tank Number of pumps 2B Total discharge flow m³/h 6.6 7.3 C 10% design factor Average operating density % w/w 0.0% 0.0% C Power installed kW R
6.1.3 Sulfide Concentrate Thickener Underflow Pump Pump type Horizontal centrifugal, SRL B Number of pumps 2B Total discharge flow m³/h 1.7 1.9 C 10% design factor Average operating density % w/w 70.0% 70.0% C Power installed kW R Variable speed
6.1.4 Sulfide Concentrate Holding Tank Type of vessel Tank with agitator B Number of tanks 1E Stock tank throughput rate (dry) t/h 2.46 3.08 C 25% design factor Stock tank flow rate m³/h 1.72 2.15 C 25% design factor Stock tank dilution water flow m³/h 0.0 B Average operating density % w/w 70.0% 65.0% C Stock tank retention time (each) h 20.0 25.0 C 25% design factor Stock tank volume m³ 34.4 43.0 C Total power installed kW B
6.1.5 Sulfide IRL Feed Pump Pump type Horizontal centrifugal, SRL B Number of pumps 2B Total discharge flow m³/h 1.7 B Average operating density % w/w 70.0% C Power installed kW R Variable speed
C3852 Process Design Criteria and Mass Balance - Rev B Page 16 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 6.2 Intensive Cyanide Leaching 6.2.1 Leach Reactor Type Conical tank B Number of bins 1B Throughput rate (wet) t/h 2.46 3.08 C 25% design factor Percent solids %w/w 70.0% 70.0% C Retention time h 20.0 25.0 C 25% design factor Tank volume m³ 5.0 R
6.3 Leach Circuit 6.3.1 Pre-leach Thickener Type of thickener High rate B Number of thickeners 1 B Thickener throughput rate (dry) t/h 46.8 58.5 C 25% design factor Thickener feed flow rate m³/h 126.3 157.8 C 25% design factor Thickener feed pulp density %w/w 30.0% 25.0% C Thickener dilution water flow rate m³/h 15.5 19.4 C 25% design factor Thickener overflow pulp density % 0.0% 0.0% B Thickener underflow pulp density % 52.0% 50.0% B Total power installed kW B
6.3.2 Pre-Leach Thickener Overflow Pump Pump type Horizontal centrifugal, Metal B Pumps back to the process water tank Number of pumps 2B Total discharge flow m³/h 66.0 72.6 C 10% design factor Average operating density % w/w 0.0% 0.0% C Power installed kW R
6.3.3 Pre-Leach Thickener Underflow Pump Pump type Horizontal centrifugal, SRL B Number of pumps 2B Total discharge flow m³/h 60.2 66.3 C 10% design factor Average operating density % w/w 52.0% 52.0% C Power installed kW C
6.3.4 Carbon-in-Leach Tanks Type of vessel Tank with pumping screen and agitator B Swept screen Number of tanks 6BStandard minimum number Leach throughput rate (dry) t/h 46.8 58.5 C 25% design factor Leach feed flow rate m³/h 60.2 75.3 C 25% design factor Leach feed pulp density %w/w 52.0% 50.0% C Leaching time h 50.0 50.0 E Leach oxygen flow rate Nm³/h TBC C Leach oxygen pressure required kPa 500 C Effective volume factor 92.0% B Total leaching design volume m³ 3 765 C Activated carbon specific gravity 2.3 I Leach tank carbon pulp density g/L 15 15 E Total carbon inventory t 56 C Installed power per tank kW B Total installed power kW B
C3852 Process Design Criteria and Mass Balance - Rev B Page 17 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS 6.3.5 Interstage Screen with Pump Type of screen Mineral Processing Separation B Number of screens 7B Number in operation 6 C Number of spares (standby) 1 B
Design feed size (F80) mm 0.075 0.094 S
Maximum leach circuit feed size (F100) mm 0.150 B MPS feed pulp density %w/w 52.0% 50.0% B & C MPS feed flow rate m³/h 60.2 66.3 C 10% design factor Carbon absorption time h TBC E Total power installed kW B
6.3.6 Carbon-in-Leach Tank Carbon Transfer Pump Pump type Screw, recessed impeller B Number of pumps 6 B Total discharge flow m³/h 25.0 27.5 C 10% design factor Average operating density % w/w 52.0% 50.0% B & C Power installed kW 7.5 B Variable speed
6.4 Intensive Cyanide Leach Area Circuit Auxiliaries 6.4.1 ICL Area Sump Pump Type Vertical cantilever pump B Number 1B Drive Fixed speed drive B Power installed kW R
6.4.2 Leach Area Sump Pump Type Vertical cantilever pump B Number 1B Drive Fixed speed drive B Power installed kW R
C3852 Process Design Criteria and Mass Balance - Rev B Page 18 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE 7.0 CYANIDE DESTRUCTION CIRCUIT DESIGN CRITERIA B CANCELLED 7.0.1 General Intensive leach time h 2.0 2.0 B & E Effective volume factor % 80% B
7.1 Sulfide Concentrate Dewatering Circuit 7.1.1 Cyanide Destruction Type of vessel Tank with oversized agitator and air spargers B Number of tanks 1B
Detox throughput rate t/h #REF! #REF! C 25% design factor Detox feed flow rate m³/h #REF! #REF! C 25% design factor Detox dilution water flow rate m³/h 0.0 0.0 C 25% design factor Detox tank pulp density %w/w 52.0% 52.0% B Based on experience Detox tank retention time h 2BTo be verified with test work Detox tank volume m³ #REF! C Detox air flow rate Nm³/h 4 000 B Estimated Detox air pressure required kPa 500 B Total power installed kW 150 B
7.1.2 Cyanide Destruction Discharge Pump Pump type Horizontal centrifugal B Pumps back to the process water tank Number of pumps 2B Design operating flow rate m³/h 62.0 68.2 C 10% design factor Average operating density % w/w 52.7% 52.7% C Power installed kW R
7.2 Cyanide Destruction Area Circuit Auxiliaries 7.2.1 Cyanide Destruction Area Sump Pump Type Vertical cantilever pump B Number 1B Drive Fixed speed drive B Power installed kW R
C3852 Process Design Criteria and Mass Balance - Rev B Page 19 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE 8.0 CONSUMABLES AND REAGENTS CIRCUIT DESIGN CRITERIA 8.0.1 General Concentrate Dewatering Average daily tonnes t/d 1 080 A & C Average throughput rate t/h 49.3 C Average daily operating hours h/d 21.9 C
8.1 Wear Items 8.1.1 Jaw Crusher Liners
Crusher specific energy = Wc kWh/t 0.166 C Crusher liner wear rate kg/kWh 0.022 C Average liner replacement point t 450 000 B Average liner replacement cycle d 417 C Main frame liner replacement point t 50 000 B To be verified with supplier Main frame liner replacement cycle d 46 C Counter weight liner replacement point t 50 000 B To be verified with supplier Counter weight liner replacement cycle d 46 C
8.1.2 SAG Mill Liners and Lifters Shell liner and lifter replacement point t 150 000 B To be verified with supplier Shell liner and lifter replacement cycle d 139 C Feed and discharge end with grates replacement point t 300 000 B To be verified with supplier Discharge end with grates replacement cycle d 278 C Pulp lifters replacement point t 600 000 B To be verified with supplier Pulp lifters replacement cycle d 556 C
8.1.3 Ball Mill Liners and Lifters Daily ball mill circuit throughput t/d 1 080 C Lifter replacement point t 900 000 B To be verified with supplier Lifter replacement cycle d 833 C
8.2 Grinding Media 8.2.1 SAG Mill Balls
SAG mill specific energy = WT kWh/t 8.41 C SAG mill ball wear rate g/kWh 105 C & L SAG mill ball wear rate g/t 887 B Seems very high SAG mill annual ball charge t 319.2 399 C SAG mill ball diameter mm 125 B
8.2.2 Ball Mill Balls Daily Ball mill circuit throughput t/d 1 080 C
Ball mill specific energy = WB kWh/t 23.2 C Ball mill ball wear rate g/kWh 105 C & L Ball mill ball wear rate g/t 2 446 C Seems extremely high Ball mill annual ball charge t 880.7 1 101 C Ball mill ball diameter mm 75 B
8.2.3 Copper Regrind Daily Stirred mill circuit throughput t/d 324 M Verti mill specific energy = WB kWh/t 15.0 C Verti mill ball wear rate g/kWh 81 C & L Verti mill ball wear rate g/t 1 222 C Seems extremely high Verti mill annual ball charge t 144.5 181 C Verti mill ball diameter mm 25 B
C3852 Process Design Criteria and Mass Balance - Rev B Page 20 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE 8.3 Concentrator Plant Reagents and Chemicals 8.3.1 Collector - PAX Orbis Gold Ltd., Report No: INT-RPT-047 - 2015-01-09; Rev 2g, pg. 659 Average collector consumption g/t 160 200 E 25% design factor Daily usage t/d 0.17 C Annual usage t/y 57.6 72.0 C 25% design factor
8.3.2 Frother - MIBC Average frother consumption g/t 120 150 E 25% design factor Daily usage t/d 0.13 C Annual usage t/y 43.2 54.0 C
8.3.3 Copper Sulfate Orbis Gold Ltd., Report No: INT-RPT-047 - 2015-01-09; Rev 2g, pg. 79 Average copper sulfate consumption g/t 20 25 E 25% design factor Daily usage t/d 0.02 C Annual usage t/y 7.2 9.0 C
8.3.4 Flocculant Average flocculant consumption g/t 50 63 B 25% design factor sulfide concentrate Average flocculant consumption g/t 30 38 E 25% design factor pre-leach (sulfide flotation tailings) Daily usage t/d 0.03 C Annual usage t/y 11.2 14.0 C 25% design factor
8.3.5 Limestone Average lime consumption to flotation g/t 200 B To flotation feed in g/t fresh feed Average lime consumption to intensive leach g/t 680.0 B To intensive leach reactor in g/t sulfide concentrate Average lime consumption to conventional leach g/t 2270 B To pre-leach thickener in g/t flotation tailings Average lime consumption to CN detox g/t 0.0 B To cyanide destruction in g/t final tailings Average lime consumption g/t 2391 2 988 C 25% design factor Daily usage t/d 2.6 3.2 C Annual usage t/y 861 1 076 C
8.4 Leaching Chemicals 8.4.1 Sodium Cyanide Orbis Gold Ltd., Report No: INT-RPT-047 - 2015-01-09; pg. 650 & Average cyanide consumption to intensive leach g/m³ 1.0 B To intensive leach reactor in g/t sulfide concentrate Average cyanide consumption to conventional leaching g/m³ 1 140 B To pre-leach thickener in g/t flotation tailings Average cyanide consumption g/t 1 394 1 742 B 25% design factor Daily usage t/d 1.5 1.9 C Annual usage t/y 501.7 627.2 C
8.4.2 Activated Carbon Average carbon wear rate % 0.08% 0.08% B Carbon inventory in CIL t 56 C Daily usage t/d 0.04 C Annual usage t/y 14.8 18.5 C
8.4.3 Lead Nitrate Average lead nitrate consumption to intensive leaching g/m³ 200 E To intensive leach reactor in g/t sulfide concentrate Average lead nitrate consumption to conventional leaching g/m³ 75 E To pre-leach thickener in g/t flotation tailings Average lead nitrate consumption g/t 99 123 E 25% design factor Daily usage t/d 0.1 0.1 C Annual usage t/y 35.5 44.4 C
8.5 Solutions Chemicals 8.5.1 Hydrochloric Acid
C3852 Process Design Criteria and Mass Balance - Rev B Page 21 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE Average acid consumption kg/t Carbon 60.0 B Carbon washing t/d 3.0 3.8 B 25% design factor Daily acid usage t/d 0.2 0.2 C Annual usage t/y 60 75 C
8.5.2 Caustic Soda Average base consumption kg/m³ Barren 2.0 B Expected barren solution volume m³/d 90 B Daily usage kg/d 180 225 C 25% design factor Annual usage t/y 60 75 C
8.6 Refinery Chemicals 8.6.0 Refinery Parameters Gold and silver in ore kg/d 9.2 11.5 C Assume 90% doré recovery Doré purity in electro-winning cells % 70% 70% I Doré sludge in electro-winning cells kg/d 13.2 16.5 C
8.6.1 Sodium Nitrate Average nitre consumption kg/t sludge 250 E Average daily usage kg/d 3.3 C Annual usage kg/y 1 099 1374 C 25% design factor Dry nitre bulk density t/m3 1.35 1.35 I
8.6.2 Fluorspar Average collector consumption kg/t sludge 150 E Average daily usage kg/d 2.0 C Annual usage kg/y 660 824 C 25% design factor Dry fluorspar bulk density t/m3 1.60 1.60 I
8.6.3 Borax Average borax consumption kg/t sludge 500 500 E Average daily usage kg/d 6.6 C Annual usage kg/y 2 199 2748 C 25% design factor Dry borax bulk density t/m3 0.96 0.96 I
8.6.4 Silica Sand Average sand consumption kg/t sludge 200 200 E Average daily usage kg/d 2.6 C Annual usage kg/y 879 1099 C 25% design factor Dry sand bulk density t/m3 1.56 1.56 I
8.6.5 Sodium Carbonate Alternative refinery chemical Average soda ash consumption kg/t sludge 50.0 50.0 E Average daily usage kg/d 0.7 C Annual usage kg/y 220 275 C 25% design factor Dry sand bulk density t 1.10 1.10 C
8.7 Cyanide Detoxification Chemicals B CANCELLED 8.7.0 Cyanide Destruction Parameters Residual cyanide in slurry g/t 250 C Normal operations Solution volume entering detox m³/d #REF! C Cyanide entering detox kg/d #REF! C Copper sulfate ratio per cyanide kg/kg 0.2 I SMBS ratio per cyanide kg/kg 4.9 I
C3852 Process Design Criteria and Mass Balance - Rev B Page 22 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE
8.7.1 Copper Sulfate Average copper sulfate consumption g/t #REF! E Average daily usage kg/d #REF! C Annual usage t/y #REF! #REF! C 25% design factor
8.7.2 Sodium Metabisulfite Average SMBS consumption g/t #REF! E Average daily usage kg/d #REF! C Annual usage t/y #REF! #REF! C 25% design factor
8.8 Reagent Area Circuit Auxiliaries 8.8.1 Reagent Area Sump Pump Type Vertical cantilever pump B Number 2B Drive Fixed speed drive B Power installed kW R
C3852 Process Design Criteria and Mass Balance - Rev B Page 23 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE 9.0 PLANT UTILITIES 9.0.1 General Selected daily ore processing rate (nominal) t/d 1 080 C Annual ore processing rate t/y 360 000 A Raw water specific gravity 1.00 I Raw water solids density % w/w 0.00 I Process water specific gravity 1.00 I Process water solids density % w/w 0.00 I
9.1 Water Services 9.1.1 Raw Water System Type of tank Closed top, flat bottom B Number of tanks 1B Capacity m³ 472 520 C 10% design factor Well pump 2BVertical turbine, direct drive Distribution pump 2BHorizontal, direct drive Design operating flow rate m³/h 10.1 11.2 C 10% design factor
9.1.2 Fire Water System Type of tank Closed top, flat bottom B Number of tanks 1B Capacity m³ 895 985 C 10% design factor Diesel pump 1 Electric pump 1 Jockey pump 1
9.1.3 Gland Water System Type of tank Closed top, flat bottom B Number of tanks 1B Capacity m³ 29.7 32.7 C 10% design factor Gland water pump 2BHorizontal, direct drive Design operating flow rate m³/h 20.0 22.0 C 10% design factor
9.1.4 Potable Water System Type of tank Closed top, flat bottom B Number of tanks 1B Capacity m³ 12.2 13.4 C 10% design factor Potable water pump 2 B Horizontal, direct drive Water treatment system 1 B Filter, UV light and chlorine Design operating flow rate m³/h 12.0 13.2 C 10% design factor
9.1.4 Process Water System Type of tank Open top, flat bottom B Number of tanks 1B Capacity m³ 761.8 838.0 C 10% design factor Process water pump 2 B Horizontal, direct drive Process water operating flow rate m³/h 113.7 125.1 C 10% design factor Reclaim water pump 2 B Horizontal, direct drive Reclaim water operating flow rate m³/h 31.0 34.1 C 10% design factor
9.2 Pressurised Air 9.2.1 Plant Air Compressors Compressor type Rotary screw B
C3852 Process Design Criteria and Mass Balance - Rev B Page 24 of 25 C3852 — PROCESS DESIGN CRITERIA
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019
REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE Number of compressors 2 B Normal discharge pressure kPa 700.0 770.0 B 10% design factor Air receiver 2B2 regular receivers and 1 instrument
9.2.2 Flotation Air Blowers Blower type Multistage, Centrifugal B Number of blowers 2B Normal discharge pressure kPa B Air Flow Nm³/h B
C3852 Process Design Criteria and Mass Balance - Rev B Page 25 of 25 SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project
Appendix B – Process Plant Mass Balance
November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final Mass Balance
Project Code: C3852
Project Name: Preliminary Economic Assessment, Nabanga Project
Document No.: C3852-PROC-DC-002
Client : SEMAFO Inc.
Process Engineer : Ewald Pengel, P.Eng.
Issue Date : 20 September 2019
Revision : B
APPROVALS
DISCIPLINE LEAD: Ewald Pengel, Senior Metallurgist
APPROVED BY (DRA): Patrick Perez, Project Manager
APPROVED BY (CLIENT):
Revision Description Revision No. Date Status
Internal DRA review A 14 août 2019 Complete Issued for report B 20 septembre 2019 Complete 10.0 MASS BALANCE
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019 SOLIDS WATER SLURRY TOTAL NAME h/d t/d t/h m³/h SG m³/d t/h m³/h SG t/h m³/h % w/w SG CRUSHING Vibrating Grizzly Feeder Vibrating feeder 16.44 1 080.0 65.7 23.6 2.785 68.9 4.2 4.2 1.000 69.9 27.8 94.0% 2.515 Jaw crusher feed 16.44 675.0 41.1 14.7 2.785 43.1 2.6 0.0 1.000 43.7 14.7 94.0% 2.515 Jaw crusher by-pass 16.44 405.0 24.6 8.8 2.785 25.9 1.6 0.0 1.000 26.2 8.9 94.0% 2.515
Jaw Crusher Jaw crusher feed 16.44 675.0 41.1 14.7 2.785 43.1 2.6 0.0 1.000 43.7 14.7 94.0% 2.515 Jaw crusher product 16.44 675.0 41.1 14.7 2.785 43.1 2.6 0.0 1.000 43.7 14.7 94.0% 2.515
Coarse Ore Stockpile Jaw crusher by-pass 16.44 405.0 24.6 8.8 2.785 25.9 1.6 0.0 1.000 26.2 8.9 94.0% 2.515 Jaw crusher product 16.44 675.0 41.1 14.7 2.785 43.1 2.6 0.0 1.000 43.7 14.7 94.0% 2.515 Coarse ore stockpile 16.44 1 080.0 65.7 23.6 2.785 68.9 4.2 0.0 1.000 69.9 23.6 94.0% 2.515
PRIMARY GRINDING SAG Mill SAG mill feed from stockpile 21.92 1 080.0 49.3 17.7 2.785 68.9 3.1 3.1 1.000 52.4 20.8 94.0% 2.515 SAG mill feed from screen crusher return 21.92 162.0 7.4 2.7 2.757 16.0 0.7 0.7 1.000 8.1 3.4 91.0% 2.381 SAG mill feed water 21.92 - - - - 307.3 14.0 14.0 1.000 14.0 14.0 0.0% 1.000 Total SAG mill feed 21.92 1 242.0 56.7 20.4 2.781 392.2 17.9 17.9 1.000 74.6 38.3 76.0% 1.948 SAG mill trommel water 21.92 - - - - 65.8 3.0 3.0 1.000 3.0 3.0 0.0% 1.000 SAG mill discharge 21.92 1 242.0 56.7 20.4 2.781 458.0 20.9 20.9 1.000 77.6 41.3 73.1% 1.879
SAG Mill Screen SAG mill screen pump discharge 21.92 1 242.0 56.7 20.4 2.781 458.0 20.9 20.9 1.000 77.6 41.3 73.1% 1.879 SAG mill screen wash water 21.92 - - - - 109.6 5.0 5.0 1.000 5.0 5.0 0.0% 1.000 SAG mill screen feed with wash water 21.92 1 242.0 56.7 20.4 2.781 567.6 25.9 25.9 1.000 82.6 46.3 68.6% 1.784 SAG mill screen oversize to SAG mill 21.92 162.0 7.4 2.7 2.757 16.0 0.7 0.7 1.000 8.1 3.4 91.0% 2.381 SAG mill screen undersize to cyclone pump 21.92 1 080.0 49.3 17.7 2.785 551.5 25.2 25.2 1.000 74.4 42.9 66.2% 1.737
SAG Mill Screen Oversize Screen oversize 21.92 162.0 7.4 2.7 2.757 16.0 0.7 0.7 1.000 8.1 3.4 91.0% 2.381 Return to SAG mill 21.92 162.0 7.4 2.7 2.757 16.0 0.7 0.7 1.000 8.1 3.4 91.0% 2.381
SECONDARY GRINDING Cyclone Feed Pumps SAG mill screen undersize 21.92 1 080.0 49.3 17.7 2.785 551.5 25.2 25.2 1.000 74.4 42.9 66.2% 1.737 Ball mill discharge 21.92 3 240.0 147.8 52.6 2.813 1 325.8 60.5 60.5 1.000 208.3 113.0 71.0% 1.866 Cyclone pump box dilution water 21.92 - - - - 1 357.4 61.9 61.9 1.000 61.9 61.9 0.0% 1.000 Cyclone feed 21.92 4 320.0 197.1 70.3 2.806 3 234.7 147.6 147.6 1.000 344.7 217.8 57.2% 1.582
Secondary Grinding Cyclones Cyclone feed 21.92 4 320.0 197.1 70.3 2.806 3 234.7 147.6 147.6 1.000 344.7 217.8 57.2% 1.582 Cyclone overflow to sulfide flotation 21.92 1 080.0 49.3 17.7 2.785 2 005.7 91.5 91.5 1.000 140.8 109.2 35.0% 1.289 Cyclone underflow 21.92 3 240.0 147.8 52.6 2.813 1 229.0 56.1 56.1 1.000 203.9 108.6 72.5% 1.877
C3852 Mass Balance - Rev B Page 2 of 5 10.0 MASS BALANCE
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019 SOLIDS WATER SLURRY TOTAL NAME h/d t/d t/h m³/h SG m³/d t/h m³/h SG t/h m³/h % w/w SG Secondary Grinding - Ball Mill Cyclone underflow 21.92 3 240.0 147.8 52.6 2.813 1 229.0 56.1 56.1 1.000 203.9 108.6 72.5% 1.877 Ball mill dilution water 21.92 - - - - 31.0 1.4 1.4 1.000 1.4 1.4 0.0% 1.000 Ball mill feed 21.92 3 240.0 147.8 52.6 2.813 1 260.0 57.5 57.5 1.000 205.3 110.0 72.0% 1.866 Ball mill trommel water 21.92 - - - - 65.8 3.0 3.0 1.000 3.0 3.0 0.0% 1.000 Ball mill discharge 21.92 3 240.0 147.8 52.6 2.813 1 325.8 60.5 60.5 1.000 208.3 113.0 71.0% 1.866
SULFIDE FLOTATION CIRCUIT Sulfide Rougher Flotation Secondary cyclone overflow 21.92 1 080.0 49.3 17.7 2.785 2 005.7 91.5 91.5 1.000 140.8 109.2 35.0% 1.289 Sulfide rougher tailings to pre-leach thickene 21.92 918.0 41.9 15.4 2.722 1 762.7 80.4 80.4 1.000 122.3 95.8 34.2% 1.289 Sulfide rougher flotation concentrate 21.92 162.0 7.4 2.3 3.202 243.0 11.1 11.1 1.000 18.5 13.4 40.0% 1.379 Sulfide rougher concentrate launder water 21.92 - - - - 75.9 3.5 3.5 1.000 3.5 3.5 0.0% 1.000 Total rougher flotation concentrate 21.92 162.0 7.4 2.3 3.202 318.9 14.5 14.5 1.000 21.9 16.9 33.7% 1.302
Sulfide Regrind Cyclone Sulfide rougher concentrate 21.92 162.0 7.4 2.3 3.202 318.9 14.5 14.5 1.000 21.9 16.9 33.7% 1.302 Regrind discharge 21.92 324.0 14.8 4.2 3.523 138.9 6.3 6.3 1.000 21.1 10.5 70.0% 2.005 Sulfide final cleaner tailings 21.92 27.0 1.2 0.4 3.492 56.2 2.6 2.6 1.000 3.8 2.9 32.4% 1.270 Sulfide regrind cyclone feed dilution water 21.92 - - - - 58.3 2.7 2.7 1.000 2.7 2.7 0.0% 1.000 Cyclone feed 21.92 513.0 23.4 6.9 3.413 516.1 23.5 23.5 1.000 47.0 30.4 49.9% 1.544 Regrind cyclone overflow to first cleaner 21.92 189.0 8.6 2.7 3.241 383.7 17.5 17.5 1.000 26.1 20.2 33.0% 1.296 Regrind cyclone underflow to regrind mill 21.92 324.0 14.8 4.2 3.523 132.3 6.0 6.0 1.000 20.8 10.2 71.0% 2.034
Sulfide Rougher Concentrate Regrind Sulfide regrind mill feed 21.92 324.0 14.8 4.2 3.523 132.3 6.0 6.0 1.000 20.8 10.2 71.0% 2.034 Sulfide regrind mill feed water 21.92 - - - - 6.5 0.3 0.3 1.000 0.3 0.3 0.0% 1.000 Sulfide regrind discharge to cyclone feed 21.92 324.0 14.8 4.2 3.523 138.9 6.3 6.3 1.000 21.1 10.5 70.0% 2.005
Sulfide First Cleaner Flotation Cyclone overflow 21.92 189.0 8.6 2.7 3.241 383.7 17.5 17.5 1.000 26.1 20.2 33.0% 1.296 Sulfide first cleaner dilution water 21.92 - - - - 1.1 0.0 0.0 1.000 0.0 0.0 0.0% 1.000 Total first cleaner feed 21.92 189.0 8.6 2.7 3.241 441.0 20.1 20.1 1.000 28.7 22.8 30.0% 1.262 Sulfide first cleaner tailings to leaching 21.92 108.0 4.9 1.6 3.000 290.6 13.3 13.3 1.000 18.2 14.9 27.1% 1.220 Sulfide first cleaner flotation concentrate 21.92 81.0 3.7 1.0 3.628 150.4 6.9 6.9 1.000 10.6 7.9 35.0% 1.296 Sulfide 1st cleaner concentrate launder wate 21.92 - - - - 44.6 2.0 2.0 1.000 2.0 2.0 0.0% 1.000 Total 1st cleaner flotation concentrate 21.92 81.0 3.7 1.0 3.628 195.1 8.9 8.9 1.000 12.6 9.9 29.3% 1.270
Sulfide Second and Third Cleaner Flotation Sulfide first cleaner flotation concentrate 21.92 81.0 3.7 1.0 3.628 195.1 8.9 8.9 1.000 12.6 9.9 29.3% 1.270 Sulfide final cleaner tailings 21.92 27.0 1.2 0.4 3.492 56.2 2.6 2.6 1.000 3.8 2.9 32.4% 1.270 Sulfide final flotation concentrate 21.92 54.0 2.5 0.7 3.701 138.9 6.3 6.3 1.000 8.8 7.0 28.0% 1.257 Sulfide final concentrate launder water 21.92 - - - - 29.2 1.3 1.3 1.000 1.3 1.3 0.0% 1.000 Total Sulfide final flotation concentrate 21.92 54.0 2.5 0.7 3.701 168.0 7.7 7.7 1.000 10.1 8.3 24.3% 1.216
C3852 Mass Balance - Rev B Page 3 of 5 10.0 MASS BALANCE
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019 SOLIDS WATER SLURRY TOTAL NAME h/d t/d t/h m³/h SG m³/d t/h m³/h SG t/h m³/h % w/w SG INTENSIVE CYANIDE LEACHING Sulfide Concentrate Thickener Sulfide rougher flotation concentrate 21.92 54.0 2.5 0.7 3.701 168.0 7.7 7.7 1.000 10.1 8.3 24.3% 1.216 Sulfide thickener dilution water 21.92 ------1.000 - - 0.0% 1.000 Sulfide thickener feed 21.92 54.0 2.5 0.7 3.701 168.0 7.7 7.7 1.000 10.1 8.3 24.3% 1.216 Sulfide thickener overflow to PW tank 21.92 - - - - 144.9 6.6 6.6 1.000 6.6 6.6 0.0% 1.000 Thickener underflow to ILR 21.92 54.0 2.5 0.7 3.701 23.1 1.1 1.1 1.000 3.5 1.7 70.0% 2.044
Sulfide Concentrate Stock Tank Concentrate thickener underflow 21.92 54.0 2.5 0.7 3.701 23.1 1.1 1.1 1.000 3.5 1.7 70.0% 2.044 Sulfide stock tank dilution water 21.92 ------1.000 - - 0.0% 1.000 Sulfide stock tank discharge 21.92 54.0 2.5 0.7 3.701 23.1 1.1 1.1 1.000 3.5 1.7 70.0% 2.044
Intensive Leach Reactor Sulfide stock tank discharge 21.92 54.0 2.5 0.7 3.701 23.1 1.1 1.1 1.000 3.5 1.7 70.0% 2.044 Intensive leach discharge to detox 21.92 54.0 2.5 0.7 3.701 23.1 1.1 1.1 1.000 3.5 1.7 70.0% 2.044
CYANIDE LEACHING Pre-Leach Thickener Sulfide rougher tailings 21.92 918.0 41.9 15.4 2.722 1 762.7 80.4 80.4 1.000 122.3 95.8 34.2% 1.277 Sulfide first cleaner tailings 21.92 108.0 4.9 1.6 3.000 290.6 13.3 13.3 1.000 18.2 14.9 27.1% 1.220 Pre-leach thickener dilution water 21.92 - - - - 340.7 15.5 15.5 1.000 15.5 15.5 0.0% 1.000 Pre-leach thickener feed 21.92 1 026.0 46.8 17.0 2.749 2 394.0 109.2 109.2 1.000 156.0 126.3 30.0% 1.236 Pre-leach thickener overflow to PW tank 21.92 - - - - 1 446.9 66.0 66.0 1.000 66.0 66.0 0.0% 1.000 Thickener underflow to filter holding tank 21.92 1 026.0 46.8 17.0 2.749 947.1 43.2 43.2 1.000 90.0 60.2 52.0% 1.494
Carbon-in-Leach Tanks Leach feed 21.92 1 026.0 46.8 17.0 2.749 947.1 43.2 43.2 1.000 90.0 60.2 52.0% 1.494 Leach tailings to cyanide destruction 21.92 1 026.0 46.8 17.0 2.749 947.1 43.2 43.2 1.000 90.0 60.2 52.0% 1.494
TAILINGS POND Tailings Pond Pump Box Intensive leach discharge 21.92 54.0 2.5 0.7 3.701 23.1 1.1 1.1 1.000 3.5 1.7 70.0% 2.044 Leach tailings 21.92 1 026.0 46.8 17.0 2.749 947.1 43.2 43.2 1.000 90.0 60.2 52.0% 1.494 Tailings pond pump box dilution water 21.92 ------1.000 0.0 0.0 0.0% 1.000 Tailings pond pump box to tailings pond 21.92 1 080.0 49.3 17.7 2.785 970.2 44.3 44.3 1.000 93.5 62.0 52.7% 1.510
WATER SERVICES Fresh Water Tank IN Raw water source pump 21.92 222.1 10.1 10.1 1.000 10.1 10.1 0.0% 1.000
OUT Gland water system 21.92 - - - 1.000 - - 0.0% 1.000 Potable water system 21.92 - - - 1.000 - - 0.0% 1.000 Process water pump 21.92 222.1 10.1 10.1 1.000 10.1 10.1 0.0% 1.000 Total OUT fresh water distribution to process 222.1 10.1 10.1 1.000 10.1 10.1 0.0% 1.000
C3852 Mass Balance - Rev B Page 4 of 5 10.0 MASS BALANCE
Preliminary Economic Assessment Revision: A Nabanga Project — SEMAFO Inc. 06 novembre 2019 SOLIDS WATER SLURRY TOTAL NAME h/d t/d t/h m³/h SG m³/d t/h m³/h SG t/h m³/h % w/w SG Process Water Tank IN Raw water source pump 21.92 222.1 10.1 10.1 1.000 10.1 10.1 0.0% 1.000 Sulfide thickener overflow to PW tank 21.92 144.9 6.6 6.6 1.000 6.6 6.6 0.0% 1.000 Pre-leach thickener overflow to PW tank 21.92 1 446.9 66.0 66.0 1.000 66.0 66.0 0.0% 1.000 Reclaim water pump 21.92 679.2 31.0 31.0 1.000 31.0 31.0 0.0% 1.000 Total IN process water from process and reclaim water pond 2 493.1 113.7 113.7 1.000 113.7 113.7 0.0% 1.000
OUT SAG mill feed water 21.92 - - - - 307.3 14.0 14.0 1.000 14.0 14.0 0.0% 1.000 SAG mill trommel water 21.92 - - - - 65.8 3.0 3.0 1.000 3.0 3.0 0.0% 1.000 SAG mill screen wash water 21.92 - - - - 109.6 5.0 5.0 1.000 5.0 5.0 0.0% 1.000 Cyclone pump box dilution water 21.92 - - - - 1 357.4 61.9 61.9 1.000 61.9 61.9 0.0% 1.000 Ball mill dilution water 21.92 - - - - 31.0 1.4 1.4 1.000 1.4 1.4 0.0% 1.000 Ball mill trommel water 21.92 - - - - 65.8 3.0 3.0 1.000 3.0 3.0 0.0% 1.000 Sulfide rougher concentrate launder water 21.92 - - - - 75.9 3.5 3.5 1.000 3.5 3.5 0.0% 1.000 Sulfide regrind cyclone feed dilution water 21.92 - - - - 58.3 2.7 2.7 1.000 2.7 2.7 0.0% 1.000 Sulfide regrind mill feed water 21.92 - - - - 6.5 0.3 0.3 1.000 0.3 0.3 0.0% 1.000 Sulfide first cleaner dilution water 21.92 - - - - 1.1 0.0 0.0 1.000 0.0 0.0 0.0% 1.000 Sulfide 1st cleaner concentrate launder wate 21.92 - - - - 44.6 2.0 2.0 1.000 2.0 2.0 0.0% 1.000 Sulfide final concentrate launder water 21.92 - - - - 29.2 1.3 1.3 1.000 1.3 1.3 0.0% 1.000 Sulfide thickener dilution water 21.92 ------1.000 - - 0.0% 1.000 Sulfide stock tank dilution water 21.92 ------1.000 - - 0.0% 1.000 Pre-leach thickener dilution water 21.92 - - - - 340.7 15.5 15.5 1.000 15.5 15.5 0.0% 1.000 Tailings pond pump box dilution water 21.92 ------1.000 - - 0.0% 1.000 Total OUT process water to process 2 493.1 113.7 113.7 1.000 113.7 113.7 0.0% 1.000
MASS BALANCE SUMMARY STREAMS - IN SAG mill feed from stockpile 1 080.0 49.3 17.7 2.785 68.9 3.1 3.1 1.000 52.4 20.8 94.0% 2.515 Raw water source pump - - - - 222.1 10.1 10.1 1.000 10.1 10.1 0.0% 1.000 Reclaim water pump - - - - 679.2 31.0 31.0 1.000 31.0 31.0 0.0% 1.000 TOTAL IN 1 080.0 49.3 17.7 2.785 970.2 44.3 44.3 1.000 93.5 62.0 52.7% 1.510
STREAMS - OUT Tailings pond pump box to tailings pond 1 080.0 49.3 17.7 2.785 970.2 44.3 44.3 1.000 93.5 62.0 52.7% 1.510 TOTAL OUT 1 080.0 49.3 17.7 2.785 970.2 44.3 44.3 1.000 93.5 62.0 52.7% 1.510
C3852 Mass Balance - Rev B Page 5 of 5