nabanga project...semafo ni 43-101 technical report preliminary economic assessment – nabanga...

DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report

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

SEMAFO NI 43-101 Technical Report Preliminary Economic Assessment – Nabanga Project

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November 2019 DRA/Met-Chem Ref.: C3852-Nabanga-PEA-Final Report 002-C3852-NI43-101_PEA_Nabanga_Final

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.


/ Page 3


Table 1.1– Nabanga Deposit – Historical Mineral Resources (at 0.5 g/t Au Cut-Off) – August 2012

Resources Category

Material Tonnes

(Mt) Grade

(g/t Au) Contained Gold (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 Category

Material Tonnes

(Mt) Grade


Oxide 0.08 8.9 20.0


Total 1.84 10.0 590.0


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.


/ Page 4


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.


/ Page 5


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

Date - From Date - To Type Count Minimum Length

Maximum Length

Total Meterage

RC Pre-

collar

Basis of Snowden

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


/ Page 6


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

YEAR

RC HOLES DIAMOND

DRILL HOLES

Count Total Length

(m) Count

Total Length (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

Resource Category Tonnage

(Mt) Grade

(g/t Au) Total Gold (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.


/ Page 8


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.


/ Page 9


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

Area Area Description Total Costs (‘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.


/ Page 10


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

Description Total Costs (‘000 USD)

Cost per ounce of gold produced

($/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.


/ Page 12


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.


/ Page 13


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

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

Richard Roy

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


/ Page 14


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

20 Environmental Studies, Permitting and Social or Community Impact

Patrick Pérez

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


/ Page 15


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.


/ Page 16


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.


/ Page 17


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.


/ Page 18


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


/ Page 19


Figure 4.2 – Boundaries of the Original Yactibo Group Permit

Source: SEMAFO

Figure 4.3 – Nabanga Permit Location

Source: SEMAFO


/ Page 20


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).


/ Page 21


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

http://www.eisourcebook.org/cms/February%202016/Burkina%20Faso%20Mining%20Code%20(in%20French%20).pdf

http://www.eisourcebook.org/cms/February%202016/Burkina%20Faso%20Mining%20Code%20(in%20French%20).pdf


/ Page 22


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

Jan

uar

y

Feb

ruar

y

Mar

ch

Ap

ril

May

Jun

e

July

Au

gu

st

Sep

tem

ber

Oct

ob

er

No

vem

ber

Dec

emb

er

AN

NU

AL

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

https://www.britannica.com/place/Sahara-desert-Africa


/ Page 23


Jan

uar

y

Feb

ruar

y

Mar

ch

Ap

ril

May

Jun

e

July

Au

gu

st

Sep

tem

ber

Oct

ob

er

No

vem

ber

Dec

emb

er

AN

NU

AL

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 (km/h) 3)

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

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 the Indicated Direction.

1 59 58 14 19 1 0 0 0

(Source: Meteoblue; https://www.meteoblue.com/en/weather/historyclimate/climatemodelled/fada-n%27gourma_burkina-faso_2360886)

https://en.climate-data.org/africa/burkina-faso/east/fada-n-gourma-766890/#climate-graph

https://en.climate-data.org/africa/burkina-faso/central-east/youga-490707/#climate-graph

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/%20-%20climate-graph

https://www.meteoblue.com/en/weather/historyclimate/climatemodelled/fada-n%27gourma_burkina-faso_2360886


/ Page 24


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.


/ Page 25


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.


/ Page 26


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;


/ Page 27


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 Category

Material Tonnes

(Mt) Grade


Oxide 0.2 5.14 30.9


Total 3.2 6.45 659.4


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 Category

Material Tonnes

(Mt) Grade


Oxide 0.08 8.9 20.0


Total 1.84 10.0 590.0



/ Page 28


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


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.


/ Page 29


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).


/ Page 30


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


/ Page 31


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.


/ Page 32


Figure 7.2 – Nabanga Property Geology

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.

NO

RT

H


/ Page 33


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.


/ Page 34


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 (2018)

Holes Meterage

(m) Samples Sector

April 1,232 6,297 2,461 Nabanga North

May 3,187 14,669 6,322 Yactibo South

June 2,798 14,408 5,586 Yactibo South;

Infill at Nabanga Extension

July 1,313 6,549 2,625

Grid straddling Nabanga, Napadé and 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


Maximum Length

Total Meterage

RC Pre-

Collar

Basis of Snowden

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

https://www.minenportal.de/unternehmen_degelistet.php?mid=1373&sid=32288&lang=en

https://www.minenportal.de/unternehmen_degelistet.php?mid=1373&sid=32288&lang=en


/ Page 39



Maximum Length

Total Meterage

RC Pre-

Collar

Basis of Snowden

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.


/ Page 40


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.


/ Page 41


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.


/ Page 42


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.


/ Page 43


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.


10.2.2.1 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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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

YEAR

RC HOLES DIAMOND

DRILL HOLES

Count Total Length

(m) Count

Total Length (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.


/ Page 46


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.


/ Page 47


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


/ Page 49


(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

CRM Count Certified Value

(g/t Au) Standard 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).


/ Page 54


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)

Criteria Count Original ≥ Duplicate

Duplicate > Original

Samples Within % Relative Difference

±10% ±20%

All Samples 439 300

(68.3%) 139

(31.7%) 141

(32.1%) 234

(53.3%)

Au≥0.025 g/t 398 278

(69.8%) 120

(30.2%) 123

(30.9%) 216

(54.3%)

0.025 g/t ≤Au<1.0 g/t 330 241

(73.0%) 89

(27.0%) 100

(30.3%) 176

(53.3%)

Au ≥ 1.0 g/t 68 37

(54.4%) 31

(45.6%) 23

(33.8%) 40

(58.8%)


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Table 11.3 – Results from the CRMs Analyzed at ALS Ouagadougou

CRM Count

PERIOD (date) CERTIFIED OBSERVED ACCEPTED FAILURES

From To Value S.D. Mean S.D. Min. Max. Count Rate (%)

Count Below 3 StdDev

Above 3

StdDev

Rate (%)

_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


/ Page 58


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

Original Au Duplicate

Au Relative

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)

HOLE ID INTERVAL

(m) GRADE (g/t Au)

DESCRIPTION

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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HOLE ID INTERVAL

(m) GRADE (g/t Au)

DESCRIPTION

NARC184 86-87 1.240 86-101 m: Trace to 3% pyrite;

NARC184 87-88 3.560 86-89 m:

96-97 m;

Quartz vein;

Quartz vein. NARC184 88-89 17.500

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:

75-83 m:

84-85 m:

Quartz vein, trace pyrite;

core non recovered;

core non recovered. NADD002 77-78 1.420

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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HOLE ID INTERVAL

(m) GRADE (g/t Au)

DESCRIPTION

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


HOLE ID INTERVAL

(m) GRADE (g/t Au)

DESCRIPTION

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


/ Page 68


HOLE ID INTERVAL

(m) GRADE (g/t Au)

DESCRIPTION

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.


/ Page 69


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

Bond Index Sample BWi

(kWh/t) Ai (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

Sample DWi

(kWh/m³) DWi (%)

Mia (kWh/t)

Mih (kWh/t)

Mic (KWh/t)

A b

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


/ Page 70


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 (g/t)

Au recovery (%)

Au feed grade (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 Grade (g/t)

Au recovery (%)

Au Feed 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.


/ Page 71


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.


/ Page 72


Table 14.1– Summary of Database Entries by Hole Type

Hole Type Count Cumulative Length (m)

Number of Assays

Sum of Assayed

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.


/ Page 73


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


/ Page 74


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.


/ Page 75


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.


/ Page 76


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°.


/ Page 77


Figure 14.1 – Variograms of 1 m Composites for South Zone (Short and Long)

SILL

0.0 1.0 2.0 3.0 4.0 5.0 6.0 7.0 8.0 9.0 10.0

0.000

0.140

0.280

0.420

0.560

0.700

0.840

0.980

1.120

1.260

1.400

ABS,Au70

Distance

Nabanga - HG South Zone 1m composites (top-cut 70g/t Au)

Variable : Au70 Date : 05-11-2019

Variogram : Absolute File : varcmphgshort.gsd

Direction :

Azimuth :

Dip :

Tolerance :

Lag Dist :

Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 70.0/0.0/-60.0)

! DH

145.00

-60.00

40.00

1.10

!

!!

!

! !

!

!

!

565

365

236

155

95

58

33

18

16

SILL

0.0 20.0 40.0 60.0 80.0 100.0 120.0 140.0 160.0 180.0 200.0

0.000

0.140

0.280

0.420

0.560

0.700

0.840

0.980

1.120

1.260

1.400

ABS,Au70

Distance

Nabanga - HG South Zone 1m composites (top-cut 70g/t Au)


Variogram : Absolute File : varcmphgslong.gsd

Direction :

Azimuth :

Dip :

Tolerance :

Lag Dist :

Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 70.0/0.0/-60.0)

!

Strike

70.00

0.00

20.00

45.00

!

Dip

340.00

-60.00

20.00

20.00

!

!!

17

887! !

!

45 1

4

1


/ Page 78


Figure 14.2 – Variograms of 1m Composites for Central Zone (Short and Long)

SILL

0.0 1.0 2.0 3.0 4.0 5.0 6.0 7.0 8.0 9.0 10.0

0.000

0.140

0.280

0.420

0.560

0.700

0.840

0.980

1.120

1.260

1.400

ABS,Au70

Distance

Nabanga - HG Central Zone 1m composites (top-cut 70g/t Au)


Variogram : Absolute File : varcmphgcshort.gsd

Direction :

Azimuth :

Dip :

Tolerance :

Lag Dist :

Gamma = N(0.3000) + E(0.7000, 12.0/10.0/1.0, 330.0/-60.0/0.0)

! DH

145.00

-60.00

40.00

1.10

!

!!

!

!!

!

!

!

321

224

155 1

03

63 4

0

25

14

14

SILL

0.0 20.0 40.0 60.0 80.0 100.0 120.0 140.0 160.0 180.0 200.0

0.000

0.140

0.280

0.420

0.560

0.700

0.840

0.980

1.120

1.260

1.400

ABS,Au70

Distance

Nabanga - HG Central Zone 1m composites (top-cut 70g/t Au)


Variogram : Absolute File : varcmphgclong.gsd

Direction :

Azimuth :

Dip :

Tolerance :

Lag Dist :

Gamma = N(0.3000) + E(0.7000, 12.0/10.0/1.0, 330.0/-60.0/0.0)

!

Strike

60.00

0.00

20.00

45.00

!

Dip

330.00

-60.00

20.00

20.00

! !

! !

175

642

457

686

!

! !

!

!

!

!

!

!

!

111

355

318

247

314

249

182

124

90

61


/ Page 79


Figure 14.3 – Variograms of 1m Composites for North Zone (Short and Long)

SILL

0.0 1.0 2.0 3.0 4.0 5.0 6.0 7.0 8.0 9.0 10.0

0.000

0.140

0.280

0.420

0.560

0.700

0.840

0.980

1.120

1.260

1.400

ABS,Au70

Distance

Nabanga - HG North Zone 1m composites (top-cut 70g/t Au)


Variogram : Absolute File : varcmphgnshort.gsd

Direction :

Azimuth :

Dip :

Tolerance :

Lag Dist :

Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 50.0/0.0/-65.0)

! DH

145.00

-60.00

40.00

1.10

!

!

!

!

!

!

!

!

195

115

63

39

23

13

5

1

SILL

0.0 20.0 40.0 60.0 80.0 100.0 120.0 140.0 160.0 180.0 200.0

0.000

0.140

0.280

0.420

0.560

0.700

0.840

0.980

1.120

1.260

1.400

ABS,Au70

Distance

Nabanga - HG North Zone 1m composites (top-cut 70g/t Au)


Variogram : Absolute File : varcmphgnlong.gsd

Direction :

Azimuth :

Dip :

Tolerance :

Lag Dist :

Gamma = N(0.3000) + E(0.7000, 10.0/10.0/1.0, 50.0/0.0/-65.0)

!

Strike

50.00

0.00

20.00

45.00

!

Dip

320.00

-65.00

20.00

20.00

!

!

!

!

99

241

316

346

!

!!

!

!

!

!!

!

!

61

232

224

188 1

38

109

66

40

47

38


/ Page 80


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

Block Model Rotation Parent Block Size (m) Block Grid Origin

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


/ Page 81


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


/ Page 82


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

Parameter North Zone Central Zone South Zone

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


/ Page 83


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


/ Page 84


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

Resources Category Tonnage (Mt) Grade (g/t Au) Total Gold (x1,000 oz)

Inferred 3.4 7.7 840

The Reader is cautioned that mineral resources that are not mineral reserves do not have



/ Page 86


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.


/ Page 89


Table 16.1 – Value of Mineralized Material Every 20 m Slice for Underground and Open Pit Scenarios

Elevation Underground Open Pit UG mineralized

material value

Open Pit mineralized

material value

Level Tonnes Au (g/t) Tonnes Au (g/t)

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:

𝐶𝑂𝐺 =(𝑀𝑖𝑛𝑖𝑛𝑔 𝑐𝑜𝑠𝑡 + 𝑃𝑟𝑜𝑐𝑒𝑠𝑖𝑛𝑔 𝑐𝑜𝑠𝑡 + 𝐺&𝐴) ∗ 𝐷𝑖𝑙𝑢𝑡𝑖𝑜𝑛

(𝑆𝑎𝑙𝑒𝑠 𝑃𝑟𝑖𝑐𝑒 − 𝑅𝑜𝑦𝑎𝑙𝑡𝑦) ∗ 𝑀𝑖𝑙𝑙 𝑅𝑒𝑐𝑜𝑣𝑒𝑟𝑦)

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

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.

NO

RT

H


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Figure 16.3 – Pit Sequencing

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

PIT Mineralized

material tonnage

Grade (g/t) Ounces Waste

tonnage Strip

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


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.

Mo

nth

-1

Mo

nth

-2

Mo

nth

-3

Mo

nth

-4

Mo

nth

-5

Mo

nth

-6

Month

-7

Mo

nth

-8

Mo

nth

-9

Mo

nth

-10

Mo

nth

-11

Mo

nth

-12

Mo

nth

-13

Mo

nth

-14

Mo

nth

-15

Mo

nth

-16

Mo

nth

-17

Mo

nth

-18

Mo

nth

-19

Month

-20

Mo

nth

-21

Mo

nth

-22

Mo

nth

-23

Mo

nth

-24

Mo

nth

-25

Mo

nth

-26

Mo

nth

-27

Mo

nth

-28

Mo

nth

-29

Mo

nth

-30

Mo

nth

-31

Pit Sequencing

Pit 4 Pit 7 Pit 6 Pit 5 Pit 3 Pit 2 Pit 1


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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

Figure 16.6 – Longitudinal Retreat Mining

2300 m


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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.


/ Page 97


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


/ Page 99


Figure 16.9 – Ventilation Network


/ Page 100


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.


/ Page 102


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 Contained

oz 66,882 77,459 66,958 79,077 82,497 78,967 64,294 75,714 570,768


/ Page 103


Figure 16.12 – Underground Development at Year 2



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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 – Nominal

t/h 49.3

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

Streams Dry

Solids (t/d)

Water (m³/d)

Total Mass (t/d)

Streams Dry

Solids (t/d)

Water (m³/d)

Total Mass (t/d)

ROM material to Concentrator

1,080.0 68.9 1,148.9 To tailings storage facility

1,080.0 970.2 2,050.2

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;


/ Page 110


• 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


/ 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


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


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


Leach retention time h 48

Leach tank dimensions m ׇ 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;


/ Page 116


• 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;

• Borax (Na₂[B₄O₅(OH)₄]٠8H₂O) is used to dissolve (base) metals in the melt;

• 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.


/ Page 117


Table 17.7 – Grinding Media and Reagent Consumption


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.


/ Page 118


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

Material Gold Recovery

(%)

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.


/ Page 120


Figure 18.1 – General Site Layout


/ Page 121


Figure 18.2 – Process Plant Layout


/ Page 122


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.


/ Page 123


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

Area Mechanical

Operating Power kWmec

Power Demand kWelec

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.


/ Page 124


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.


/ Page 125


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;


/ Page 126


• 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.


/ Page 128


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.


/ Page 129


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

Environmental Compliance Certificate

Delivered by the MEEVCC on the basis of a compliant ESIA Report:

Submission of the ESIA Terms of Reference (ToR) to the BUNEE

Validation of the ToR

Submission of the draft version of the ESIA to the BUNEE

Public Inquiry

COTEVE Session

Final ESIA Report approved by the BUNEE

Before the expiry of the Exploration License

Cost for review and validation of the ToR: 500,000 FCFA.

Fees for the public inquiry and COTEVE session: Approximately 3 million FCFA.

Cost for file processing-Project value of 50 billion FCFA and more:

Flat Fee: 25 million FCFA.

Proportional rights: 0.02% of the total investment cost = 26.3 million FCFA (assuming a project cost of about 131.6 billion FCFA).

Permit for Industrial Exploitation

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.

The mineral substances or for whom or which the license is sought. Definition of coordinates of the perimeter and requested area.

Location perimeter license requested on a topographic map at 1:200,000.

Detailed plan at an appropriate scale where the coordinates of the perimeter requested are attached to outstanding points invariable ground and well defined.

Detailed statement showing the results of research conducted and evidence of expenditures incurred during the last period of license validity.

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.

90 days before expiry of the Exploration Permit

Costs for the National Commission of Mines session: about 2 million FCFA.

Fixed duties on mineral titles: 5,000,000 FCFA.

Proportional rights (area taxes):

First five years: 7,500,000 FCFA/ km²/y.

Year 6 to10: 10 million FCFA/ km²/y.

From the 11thyear: 15 million FCFA/ km²/y.

Cost of proportional royalty on gold production: 5% of turnover if gold is USD 1,300 and higher, 4% if USD 1,000-1,300 and 3% if less than USD 1,000.


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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 Management of raw water

Application to the Directorate of Legislation and regulations of the Ministry in charge of water.

30 days after the quarter of the levy

Tax of raw water for mining and industrial purposes: 200 FCFA /m³ a

Authorization for the Collection of raw water for civil work

No license or authorization required (besides the Mining Licence)

However, tax must be paid.

10 FCFA /m³ for every m³ backfill placed.

20 FCFA /m³ for every m³ of concrete poured.

Authorization for Road infrastructure

Authorisation from the Ministry of Transport on the basis of technical studies.

Depending on the infrastructure and technical studies that have been conducted.

Authorization for Hydraulic work or dam

Authorisation from the Ministry in charge of water on the basis of technical studies.

Depending on the work and 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

Area Area Description Total Costs (‘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



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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

Source Currency

Description Base

Currency

Currency Exchange

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;


/ Page 148


• 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.


/ Page 149


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

Description Total Costs (‘000 USD)

Cost per ounce of gold produced

($/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.


/ Page 150


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

Area Cost/Year

($) Cost/tonne

($/t) Total Cost

(%)

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

Description Total Cost

($) Cost/tonne

($/t) Total Cost

(%)

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


/ Page 151


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

Operating Cost Area Cost/year 1) Cost/t

processed 2) Cost/oz of gold

produced 3) Total Cost

(%)

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 consumption

2,790,000 7.75 39.87 16.6

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 miscellaneous 5)

400,000 1.11 5.72 2.4

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.


/ Page 152


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 IRR % 31.4


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.


/ Page 154


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.


/ Page 155


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 IRR % 31.4

Pre-Tax Payback Period Years 3.5

Total After-Tax Cash Flow M USD 157.9





After-Tax Payback Period* Years 4.4


/ Page 156


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


/ Page 157


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.


/ Page 159


Figure 22.3 – After-Tax Royalty NPV5 % & IRR Sensitivity to Capital Expenditure, Operating Cost and Price


/ Page 160


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.


/ Page 161


24 OTHER RELEVANT DATA AND INFORMATION

There is no additional information or explanation necessary to make this Technical Report

understandable and not misleading.


/ 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.


/ 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.


/ 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.


/ 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

Activities Estimated Budget

($)

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


/ 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.


/ 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


/ Page 168



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


/ Page 169



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


/ Page 170



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


/ Page 171



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


/ Page 172



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


/ Page 173



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


/ Page 174



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


/ Page 175



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


/ Page 176



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


/ Page 177



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


/ Page 178


29 CERTIFICATES OF QUALIFIED PERSONS

// 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

/ Page 1 of 2

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;

/ Page 2 of 2

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


555 René-Lévesque Blvd West / 6th floor / Montréal / Quebec / Canada / H2Z 1B1 T +1 514 288-5211 / E [email protected] / met-chem.com

/ Page 1 of 2

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:

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);

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);

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);

Metallurgical sampling: drilling to generate gold mineralized material (Mali).

5. I have read the definition of "qualified person" set out in National Instrument 43 101 (NI 43

/ Page 2 of 2

101) and certify that by reason of my education, affiliation with a professional association (as defined by NI 43 101) and past relevant work experience, I fulfill the requirements to be a "qualified person" for the purposes of NI 43 101.

6. I am independent of the issuer applying all of the tests in section 1.5 of NI 43 101.

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.







“Original document signed and sealed”

“Yves A. Buro”

Yves A. Buro, P. Eng.


555 René-Lévesque Blvd West / 6th floor / Montréal / Quebec / Canada / H2Z 1B1T +1 514 288-5211 / E [email protected] / met-chem.com

/ Page 1 of 2

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;




/ Page 2 of 2




“Original Signed and sealed” Ewald Pengel, M.Sc., P.Eng. Senior Process Engineer

Téléphones : 514 744-4408 1-888-744-4408 Télécopieur : 514 744-2291

SEMAFO inc. 100, boul. Alexis-Nihon 7e étage Saint-Laurent (Québec) H4M 2P3

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 1-888-744-4408 Télécopieur : 514 744-2291

SEMAFO inc. 100, boul. Alexis-Nihon 7e 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.



Appendix A – 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 :

Revision : B

Internal DRA review A

Issued for report B

Status

Complete

Complete20 septembre 2019

Revision Description Revision No. Date

14 août 2019

APPROVED BY (DRA): Patrick Perez, Project Manager

APPROVED BY (CLIENT):

Process Design Criteria

DISCIPLINE LEAD: Ewald Pengel, Senior Metallurgist

APPROVALS

20 September 2019

A

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 CriteriaA = 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 Terminology1.3.1 Nominal

1.3.2 DesignThese 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 TimeCrushing 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 PercentageOperating percentage is operating time period divided by the calendar time period expressed in a percentage.

1.3.5 Weight recoveryAll weight recoveries are based on New Feed, unless stated otherwise.

1.3.6 New FeedNew Feed is the ore placed on the SAG Mill feed conveyor material by the Apron Feeder underneath the stockpile.

1.4 Colour coding1.4.1 Black

Final information from either test work, supplier, DRA or client experience.

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.

06 novembre 2019

Preliminary Economic AssessmentNabanga Project — SEMAFO Inc.

C3852 — PROCESS DESIGN CRITERIA

Revision:

C3852 Process Design Criteria and Mass Balance - Rev B Page 2 of 25

A06 novembre 2019



Revision:

1.4.2 BlueVerified information. This is usually based on preliminary information or basic calculations.

1.4.3 RedUnverified information, entry is best guess or based on database.

1.4.4 Red with yellow backgroundUnverified information, which may be not to be used by other groups

1.5 Units1.5.1 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


A06 novembre 2019



Revision:

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

= Major item or design criteria updated for Revision BB


A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS2.0 GENERAL PROCESSING DESIGN CRITERIA2.0.1 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 Weather2.1.1 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 Characteristics2.2.1 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 SpecificationsDesign 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 indicesAbrasion 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

06 novembre 2019Revision:




A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS




2.3 Processing Criteria2.3.1 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 CriteriaCrushing 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 CriteriaCrusher 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 CriteriaConcentrator 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 CriteriaPrimary 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 CriteriaFlotation - 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 CriteriaIntensive 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 CriteriaCyanide 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 Not required A E-Mail from S.Duchesne 2019-08-15

Cyanide destruction time

Cyanide destruction percent solids

B


A





2.4 Utility Specifications2.4.1 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

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 3 B

Frequency Hz 50 B

B


A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS3.0 CRUSHING CIRCUIT DESIGN CRITERIA3.0.1 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 Circuit3.1.1 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


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 1 B

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


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


3.1.5 Primary Crusher

Type of crusher Jaw A Type C116 equivalent

Number 1 B

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

06 novembre 2019



Revision:


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06 novembre 2019



Revision:

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


3.1.6 Sacrificial Conveyor

Conveyor type Belt B


Total conveyor feed F100 mm 93 S From simulation

Conveyor size

Width mm R



3.1.8 Stockpile Conveyor

Conveyor type Belt B


Total crusher product P100 mm 93 C & S From simulation

Conveyor size

Width mm R



3.2 Crushing Circuit Auxiliaries3.2.1 Jaw Crusher Monorail

Type of crane Monorail B

Number of cranes 1 B

Required lift capacity t 25 B

Required span m D By Met-Chem designer

Required lifting height m D By Met-Chem designer


3.2.2 Tramp Metal Magnet Monorail



Required lift capacity t 5.0 B



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 1 B

Drive Fixed speed drive B



A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS4.0 GRINDING CIRCUIT DESIGN CRITERIA4.0.1 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 Circuit4.1.1 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 2 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


4.1.4 Primary Grinding Mill

Type of mill Semi-Autogenous Grinding B

Number of mills 1 A


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


Preliminary Economic AssessmentNabanga Project — SEMAFO Inc. 06 novembre 2019

Revision:


A




Revision:

Mill size

Diameter m R

Length m R


4.1.5 SAG Mill Screen

Type of screen Single deck horizontal, vibrating R

Number of screens 1 B

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


4.2 Secondary Milling Circuit4.2.1 Cyclone Feed Pump

Pump type Horizontal centrifugal, metal B

Number of pumps 2 B

Total discharge flow m³/h 218 240 B & C 10% design factor

Average operating density % w/w 52.0% 52.0% C


4.2.2 Secondary Cyclones

Cyclone type High gravity B

Number of cyclones 8 A E-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 1 B

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


A




Revision:

Mill size

Diameter m R

Length m R


4.3 Grinding Circuit Auxiliaries4.3.1 Apron Feeder Monorail



Required lift capacity t 1.0 B



4.3.2 Grinding Circuit Area Overhead Crane

Type of crane Overhead B


Required crane capacity t 15 A E-Mail from S.Duchesne 2019-08-15



4.3.3 Automated Ball Feeder

Type Screw feeder A

Number 0 B Ball 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


Number 2 B



B

B


A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS5.0 SULFIDE FLOTATION CIRCUIT DESIGN CRITERIA5.0.1 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 Circuit5.1.1 Sulfide Rougher Flotation Circuit

Type of cells Conventional R

Number 6 R 10 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 1 B

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 2 B



5.2 Sulfide Cleaner Flotation Circuit5.2.1 Re-grind Cyclone Feed Pump


Number of pumps 2 B



5.2.2 Sulfide Regrind Cyclones

Cyclone type High gravity B

Number of cyclones 8 A E-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



Revision:


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Revision:

5.2.3 Sulfide Regrind Mill

Type of mill Tower mill R Variable speed

Number of mills 1 B

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


Number 4 R

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


Number of pumps 2 B




5.2.6 Sulfide First Cleaner Concentrate Pump

Pump type Vertical tank, SRL B

Number of pumps 1 B




5.2.7 Sulfide Second and Third Cleaner Flotation Circuit


Number 5 B 3 → 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


A




Revision:

5.2.8 Sulfide Second Cleaner Tailings Pump


Number of pumps 2 B




5.2.9 Sulfide Second Cleaner Concentrate Pump


Number of pumps 1 B




5.2.10 Sulfide Third Cleaner Concentrate Pump


Number of pumps 1 B




5.3 Sulfide Flotation Circuit Auxiliaries5.3.1 Sulfide Flotation Area Sump Pump


Number 1 B




A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS6.0 INTENSIVE CYANIDE LEACHING CIRCUIT DESIGN CRITERIA6.0.1 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 Circuit6.1.1 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 2 B




6.1.3 Sulfide Concentrate Thickener Underflow Pump


Number of pumps 2 B



Power installed kW R Variable speed

6.1.4 Sulfide Concentrate Holding Tank

Type of vessel Tank with agitator B

Number of tanks 1 E

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


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


Number of pumps 2 B

Total discharge flow m³/h 1.7 B


Power installed kW R Variable speed



Revision:


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Revision:

6.2 Intensive Cyanide Leaching6.2.1 Leach Reactor

Type Conical tank B

Number of bins 1 B

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 Circuit6.3.1 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


6.3.2 Pre-Leach Thickener Overflow Pump

Pump type Horizontal centrifugal, Metal B Pumps back to the process water tank

Number of pumps 2 B




6.3.3 Pre-Leach Thickener Underflow Pump


Number of pumps 2 B



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 6 B Standard 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


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Revision:

6.3.5 Interstage Screen with Pump

Type of screen Mineral Processing Separation B

Number of screens 7 B

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


6.3.6 Carbon-in-Leach Tank Carbon Transfer Pump

Pump type Screw, recessed impeller B

Number of pumps 6 B


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 Auxiliaries6.4.1 ICL Area Sump Pump


Number 1 B



6.4.2 Leach Area Sump Pump


Number 1 B




A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE7.0 CYANIDE DESTRUCTION CIRCUIT DESIGN CRITERIA 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 Circuit7.1.1 Cyanide Destruction

Type of vessel Tank with oversized agitator and air spargers B

Number of tanks 1 B

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 2 B To 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 2 B

Design operating flow rate m³/h 62.0 68.2 C 10% design factor



7.2 Cyanide Destruction Area Circuit Auxiliaries7.2.1 Cyanide Destruction Area Sump Pump


Number 1 B





Revision:

B


A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE8.0 CONSUMABLES AND REAGENTS CIRCUIT DESIGN CRITERIA8.0.1 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 Items8.1.1 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 Media8.2.1 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



Revision:06 novembre 2019


A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE




8.3 Concentrator Plant Reagents and Chemicals8.3.1 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


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



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)


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 Chemicals8.4.1 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



8.4.2 Activated Carbon

Average carbon wear rate % 0.08% 0.08% B

Carbon inventory in CIL t 56 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



8.5 Solutions Chemicals8.5.1 Hydrochloric Acid


A





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 Chemicals8.6.0 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


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


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



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



Dry sand bulk density t 1.10 1.10 C

8.7 Cyanide Detoxification Chemicals 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

B


A





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 Auxiliaries8.8.1 Reagent Area Sump Pump


Number 2 B




A

REF. ITEM UNITS NOMINAL DESIGN SOURCE COMMENTS / REFERENCE9.0 PLANT UTILITIES9.0.1 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 Services9.1.1 Raw Water System

Type of tank Closed top, flat bottom B

Number of tanks 1 B

Capacity m³ 472 520 C 10% design factor

Well pump 2 B Vertical turbine, direct drive

Distribution pump 2 B Horizontal, direct drive


9.1.2 Fire Water System


Number of tanks 1 B

Capacity m³ 895 985 C 10% design factor

Diesel pump 1

Electric pump 1

Jockey pump 1

9.1.3 Gland Water System


Number of tanks 1 B

Capacity m³ 29.7 32.7 C 10% design factor

Gland water pump 2 B Horizontal, direct drive


9.1.4 Potable Water System


Number of tanks 1 B


Potable water pump 2 B Horizontal, direct drive

Water treatment system 1 B Filter, UV light and chlorine


9.1.4 Process Water System

Type of tank Open top, flat bottom B

Number of tanks 1 B


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



Revision:


A




Revision:

Number of compressors 2 B

Normal discharge pressure kPa 700.0 770.0 B 10% design factor

Air receiver 2 B 2 regular receivers and 1 instrument

9.2.2 Flotation Air Blowers

Blower type Multistage, Centrifugal B

Number of blowers 2 B

Normal discharge pressure kPa B

Air Flow Nm³/h B




Appendix B – Process Plant 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 :

Revision : B

Internal DRA review A

Issued for report B

Status

Complete

Complete20 septembre 2019

Revision Description Revision No. Date

14 août 2019

APPROVED BY (DRA): Patrick Perez, Project Manager

APPROVED BY (CLIENT):

Mass Balance

DISCIPLINE LEAD: Ewald Pengel, Senior Metallurgist

APPROVALS

20 September 2019

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

SOLIDS WATER SLURRY TOTAL

10.0 MASS BALANCE


Revision: A06 novembre 2019

C3852 Mass Balance - Rev B Page 2 of 5

NAME h/d t/d t/h m³/h SG m³/d t/h m³/h SG t/h m³/h % w/w SG SOLIDS WATER SLURRY TOTAL

10.0 MASS BALANCE



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



10.0 MASS BALANCE



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



10.0 MASS BALANCE



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


nabanga project...semafo ni 43-101 technical report preliminary economic assessment – nabanga...

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