NI 43-101 Technical Report 31 December 2019 Bisie Tin Mine

NI 43-101 Technical Report – 31 December 2019 Bisie Tin Mine,

North Kivu Province, Democratic Republic of Congo

Prepared for Alphamin Resources Corp.

Authors: Qualified Person Mineral Resources: Jeremy Witley, B.Sc. Hons., M.Sc. (Eng.), Pr.Sci.Nat, FGSSA Qualified Person Ore Reserves: Vaughn Duke, Pr.Eng., PMP, B.Sc. Mining Engineering (Hons.), MBA, FSAIMM Document Reference Number: PR/SMS/0844/19 Effective Date (date of determination of technical and financial parameters): 31 December 2019 Report Date (date of finalisation): 11 February 2020

Bisie Tin Mine NI43-101

Bisie Tin Mine NI 43-101 Technical Report – 31 December 2019

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Forward Looking Statements

Sound Mining International (Pty) Ltd (SMI) has prepared this Technical Report which contains forward-looking statements. In preparing this

Technical Report, SMI has also utilised information provided by the MSA Group (Pty) Ltd (MSA), Alphamin Bisie Mining SA (ABM), and

its specialist consultants. These forward-looking statements are based on the opinions and estimates of SMI, MSA, ABM and its specialist consultants at the date the statements were made. The statements are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those anticipated in the SMI, ABM and the specialists’ forward-looking statements. Factors that could cause such differences include changes in world markets, equity markets, costs and supply of materials relevant to the ABM business, and changes to regulations affecting them. Although SMI believes the expectations reflected in its forward-looking statements to be reasonable, SMI does not guarantee future results, levels of activity, performance or achievements. Many of the issues that have been reported upon are of a complex nature and SMI has where possible, verified this information making due enquiry of all material issues that could positively or negatively influence the level of risk associated with Mineral Resources or Mineral Reserves. SMI has assumed that all of the information and technical documents received and reviewed in this Technical Report are accurate and complete in all material aspects. While SMI carefully reviewed this information, SMI has not conducted an extensive independent investigation to verify its accuracy and completeness. The information and conclusions contained herein are based on the information available to SMI at the time of preparation of this Technical Report. ABM agrees that neither it nor its associates will make any claim against SMI to recover any loss or damage suffered as a result of SMI’s reliance on the information provided by ABM for use in the preparation of this Technical Report. Furthermore, ABM has also indemnified SMI against any claim arising out of the assignment to prepare this Technical Report, except where the claim arises as a result of any proven wilful misconduct or negligence on the part of SMI. This indemnity is also applied to any consequential extension of work through queries, questions, public hearings or additional work required arising from SMI’s performance of the engagement. SMI reserves the right to, but will not be obligated to, revise this Technical Report and conclusions thereto if additional information becomes known to SMI subsequent to the date of this Technical Report. The authors of this Technical Report are not qualified to provide extensive commentary on legal issues associated with ABM’s right to the assets. SMI has undertaken a review of the legal aspects of the mining operation but no warranty or guarantee, be it express or implied, is made by the authors with respect to the completeness or accuracy of the legal aspects of this document. This document has been prepared as at the date stated on the cover page. Given the nature of this document and the opinions expressed within, developments after the date of this document are likely. This document takes no account of such potential future developments; therefore, SMI recommends that readers seek advice from SMI in the future to ascertain whether any such events have occurred or updated information has become available and should be considered. Operational Risks

The businesses of mining and mineral exploration, development and production by their natures contain significant operational risks. The businesses depend upon, amongst other things, successful prospecting programmes and competent management. Profitability and asset values can be affected by unforeseen changes in operating circumstances and technical issues. Political and Economic Risks

Factors such as political and industrial disruption, currency fluctuation and interest rates could have an impact on ABM’s future operations, and potential revenue streams can also be affected by these factors. The majority of these factors are, and will be, beyond the control of ABM or any other operating entity.



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Contents List

1. EXECUTIVE SUMMARY ............................................................................................................................................. 9 1.1. Property Description and Ownership .................................................................................................................. 9 1.2. Status of Activities .............................................................................................................................................. 9 1.3. Geology and Mineralisation .............................................................................................................................. 10 1.4. Status of Exploration ........................................................................................................................................ 11 1.5. Mineral Resource Estimate .............................................................................................................................. 11 1.6. Mineral Reserve Estimate ................................................................................................................................ 12 1.7. Mineral Processing ........................................................................................................................................... 14 1.8. Environmental Considerations, Permitting and Social or Community Impact ................................................... 15 1.9. Infrastructure and Logistics .............................................................................................................................. 15 1.10. Capital and Operating Costs ............................................................................................................................ 15 1.11. Economic Assessment ..................................................................................................................................... 16 1.12. Conclusions ...................................................................................................................................................... 17 1.13. Recommendations ........................................................................................................................................... 17

2. INTRODUCTION ....................................................................................................................................................... 18 2.1. Issuer and Terms of Reference ........................................................................................................................ 18 2.2. Principal Sources of Information ...................................................................................................................... 18 2.3. Qualified Persons Site Visit .............................................................................................................................. 18 2.4. Qualifications, Experience and Independence of Qualified Persons ................................................................ 19 2.5. Mining Legislation in the DRC .......................................................................................................................... 19

Mineral Property and Title .......................................................................................................................... 19 Sale of Mining Products ............................................................................................................................. 20 Surface Rights Title .................................................................................................................................... 20

3. RELIANCE ON OTHER EXPERTS ........................................................................................................................... 21 4. PROPERTY DESCRIPTION AND LOCATION .......................................................................................................... 22

4.1. Location............................................................................................................................................................ 22 4.2. Ownership ........................................................................................................................................................ 22 4.3. Mineral Tenure, Permitting, Rights and Agreements ........................................................................................ 22 4.4. Force Majeure .................................................................................................................................................. 23

5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ........................ 24 5.1. Accessibility ...................................................................................................................................................... 24 5.2. Climate and Physiography ............................................................................................................................... 24 5.3. Local Resources and Infrastructure ................................................................................................................. 25

6. HISTORY ................................................................................................................................................................... 26 6.1. Geophysics ...................................................................................................................................................... 26 6.2. Soil Geochemistry ............................................................................................................................................ 27 6.3. Rock Sample Geochemistry ............................................................................................................................. 27 6.4. Stream Sediment Sampling .............................................................................................................................. 27

7. GEOLOGICAL SETTING AND MINERALISATION ................................................................................................... 28 7.1. Geological Setting and Mineralisation .............................................................................................................. 28

Regional Structure ..................................................................................................................................... 28 7.2. Property Geology ............................................................................................................................................. 29

Stratigraphy ............................................................................................................................................... 29 Local Structure ........................................................................................................................................... 30 Mineralisation ............................................................................................................................................. 31

7.2.3.1. Mpama South Prospect ........................................................................................................................... 32 7.2.3.2. Mpama North ........................................................................................................................................... 34

8. DEPOSIT TYPES ...................................................................................................................................................... 37 9. EXPLORATION ......................................................................................................................................................... 38

9.1. Reconnaissance work by Mining and Processing Congo Sprl (2006-2008) ..................................................... 38 10. DRILLING .................................................................................................................................................................. 39

10.1. Drill Sample Recovery ...................................................................................................................................... 39 11. SAMPLE PREPARATION, ANALYSES AND SECURITY ......................................................................................... 40 12. DATA VERIFICATION ............................................................................................................................................... 41

12.1. Analytical Quality Control and Assurance conducted by Alphamin .................................................................. 41



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Results of the Blank Sample Analysis ...................................................................................................... 41 Results of the CRM Sample Analysis ....................................................................................................... 43 Results of the Core Duplicate Sample Analysis ........................................................................................ 45 Results of the Pulp Duplicate Sample Analysis ........................................................................................ 46 Results of the Second Laboratory Check Analysis ................................................................................... 47 Re-assaying of Pre-2015 Samples by ALS ............................................................................................... 54

12.1.6.1. Summary of the external Assay QA/QC Checks.................................................................................... 56 12.2. Data Verification ............................................................................................................................................... 57

Independent Check Sampling ................................................................................................................... 57 Visual Verification ..................................................................................................................................... 58 Verification of Drillhole Collars .................................................................................................................. 58 Summary of the Data Verification ............................................................................................................. 58

13. MINERAL PROCESSING AND METALLURGICAL TESTING METHODS ............................................................... 59 13.1. Original Test Work ........................................................................................................................................... 59 13.2. Front -end Engineering and Design (Feed) Test Work ..................................................................................... 59

14. MINERAL RESOURCE ESTIMATE ........................................................................................................................... 61 14.1. Mineral Resource Estimation Database ........................................................................................................... 61 14.2. Exploratory Analysis of the Raw Data .............................................................................................................. 61

Validation of the data ................................................................................................................................ 62 Statistics of the Sample Data .................................................................................................................... 63 Statistics of the Assay Data ...................................................................................................................... 63

14.2.3.1. Univariate Analysis ................................................................................................................................ 63 14.2.3.2. Bivariate Analysis .................................................................................................................................. 63 14.2.3.3. Relationship between tin grade and specific gravity .............................................................................. 64 14.2.3.4. Comparison between Laboratory Gas Pycnometer and On-Site Wet-Dry Determinations .................... 64

Summary of the Exploratory Analysis of the Raw Dataset ........................................................................ 65 Topography............................................................................................................................................... 65 Mineralised Zones .................................................................................................................................... 66 Oxidation/Weathering Surfaces ................................................................................................................ 68

14.3. Statistical Analysis of the Composite Data ....................................................................................................... 68 Cutting and Capping ................................................................................................................................. 71

14.4. Geostatistical Analysis ..................................................................................................................................... 72 Semi-variograms ....................................................................................................................................... 72 Indicator Semi-variograms ........................................................................................................................ 75 Above and Below Threshold Semi-variograms ......................................................................................... 75

14.5. Block Modelling ................................................................................................................................................ 77 Validation of the Block Model Volumes with the Wireframe Volumes ....................................................... 78

14.6. Estimation ........................................................................................................................................................ 78 Estimation of tin accumulation .................................................................................................................. 78

14.7. Validation of the Estimates ............................................................................................................................... 81 14.8. Mineral Resource Estimate .............................................................................................................................. 83

15. MINERAL RESERVE ESTIMATE .............................................................................................................................. 84 15.1. Cut-off grade .................................................................................................................................................... 84 15.2. Mine Design Criteria and Modifying Factors ..................................................................................................... 84 15.3. Disclosures ....................................................................................................................................................... 85

16. MINING METHODS ................................................................................................................................................... 86 16.1. Geological Considerations ............................................................................................................................... 86 16.2. Geotechnical Considerations ........................................................................................................................... 86

Geotechnical Data Acquisition .................................................................................................................. 86 Rock Mass Quality .................................................................................................................................... 87 Joint Orientations ...................................................................................................................................... 88 Intact Rock Strength Testing..................................................................................................................... 88 Derivation of Mechanical Properties for Intact Rock and Joints ................................................................ 88 Regional Stresses and Rock Strengths .................................................................................................... 89 Access Excavations and Support Design ................................................................................................. 90

16.3. Ground Water Considerations .......................................................................................................................... 90 16.4. Ventilation Considerations ................................................................................................................................ 91



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Design Criteria and Constraints ................................................................................................................ 91 Determination of Air Requirements ........................................................................................................... 91 Development Ventilation ........................................................................................................................... 92 Stoping Ventilation .................................................................................................................................... 92 Ventilation Controls ................................................................................................................................... 92 Geotechnical Data Acquisition .................................................................................................................. 92

16.5. Mine Design ..................................................................................................................................................... 93 16.5.1. Mining Method Description ....................................................................................................................... 93 16.5.2. Dilution ...................................................................................................................................................... 95 16.5.3. Development End Dimensions.................................................................................................................. 95 16.5.4. Development Advance Rate ..................................................................................................................... 96 16.5.5. Scheduling Delays .................................................................................................................................... 96 16.5.6. Primary Access and Development ............................................................................................................ 96 16.5.7. Stoping ..................................................................................................................................................... 99

Backfill .................................................................................................................................................... 100 Load, Haul and Logistics ........................................................................................................................ 101

Equipment selection and productivities ................................................................................................. 101 16.6. Mining Layout and Scheduling ....................................................................................................................... 105

Mineral Resources to Mineral Reserve Reconciliation ............................................................................ 105 Mining Schedule ..................................................................................................................................... 107

16.7. Underground Mine Infrastructure ................................................................................................................... 111 Service Water Handling .......................................................................................................................... 111 Dirty Water Handling ............................................................................................................................... 111 Compressed Air ...................................................................................................................................... 111 Electrical ................................................................................................................................................. 111 Communication ....................................................................................................................................... 111 Trackless Equipment Workshop ............................................................................................................. 111

17. RECOVERY METHODS .......................................................................................................................................... 112 17.1. Description ..................................................................................................................................................... 112 17.2. Test Work ....................................................................................................................................................... 113 17.3. Process Description ....................................................................................................................................... 113 17.4. Reagents ........................................................................................................................................................ 115

18. PROJECT INFRASTRUCTURE .............................................................................................................................. 116 18.1. Introduction .................................................................................................................................................... 116 18.2. Mining Infrastructure ...................................................................................................................................... 117 18.3. Services ......................................................................................................................................................... 118 18.4. Process Plant and Administration Infrastructure ............................................................................................ 119 18.5. Power Supply and Distribution ....................................................................................................................... 121

Power Supply.......................................................................................................................................... 121 18.6. Mine Residue Storage .................................................................................................................................... 121 18.7. Waste Controls ............................................................................................................................................... 121 18.8. Health and Safety ........................................................................................................................................... 121

19. MARKET STUDIES AND CONTRACTS ................................................................................................................. 122 19.1. Tin Market Analysis ........................................................................................................................................ 122

Tin Use by Sector ................................................................................................................................... 122 Tin Supply ............................................................................................................................................... 123 Tin Demand ............................................................................................................................................ 123 Tin Stockpiles ......................................................................................................................................... 123

19.2. Supply and Demand Balance ......................................................................................................................... 124 Forecast Tin Price ................................................................................................................................... 124

19.3. Marketing Contract ......................................................................................................................................... 124 20. ENVIRONMENTAL, PERMITTING AND SOCIAL IMPACT ..................................................................................... 125

20.1. Permits and Legal Aspects ............................................................................................................................. 125 20.2. Environmental Aspects................................................................................................................................... 125 20.3. Social Aspects ................................................................................................................................................ 126

21. CAPITAL AND OPERATING COSTS ...................................................................................................................... 128 21.1. Capital Expenditure ........................................................................................................................................ 128



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21.2. Operating Cost Estimate ................................................................................................................................ 128 22. ECONOMIC ANALYSIS .......................................................................................................................................... 130 23. ADJACENT PROPERTIES ...................................................................................................................................... 131 24. OTHER RELEVANT INFORMATION ...................................................................................................................... 132

24.1. Exploration ..................................................................................................................................................... 132 Bisie Ridge Works Programme ............................................................................................................... 132 Mpama South Works Programme ........................................................................................................... 132

24.2. Security .......................................................................................................................................................... 133 24.3. Risks .............................................................................................................................................................. 134

25. INTERPRETATION AND CONCLUSIONS .............................................................................................................. 135 26. RECOMMENDATIONS ........................................................................................................................................... 136 27. REFERENCES ........................................................................................................................................................ 137 28. DATE AND SIGNATURE PAGE .............................................................................................................................. 138 29. APPENDIX 1: THREATS AND OPPORTUNITIES .................................................................................................. 141 30. APPENDIX 2: CHECKLIST OF ASSESSMENT AND REPORTING CRITERIA ...................................................... 142 31. APPENDIX 3: ABBREVIATIONS, SI UNITS AND GLOSSARY OF TERMS ........................................................... 146 32. APPENDIX 4: MINERAL RESOURCE SPLIT OF RoM SCHEDULED FROM THE BLOCK MODEL...................... 149 33. APPENDIX 5: INVENTORY OF DRILLHOLE INTERSECTIONS WITHIN THE MINERAL RESOURCE ................ 150 34. APPENDIX 6: HISTOGRAMS AND LOG PROBABILITY PLOTS OF THE COMPOSITE DATA ............................ 155 35. APPENDIX 7: SEMI VARIOGRAMS ........................................................................................................................ 169

Table List

Table 1: Bisie Mpama North Mineral Resource at 0.50% Sn Cut-off grade as at 30 June 2019 ................................................................ 12 Table 2: Bisie Mpama North Zone Mineral Reserve at USD17,000/t Sn as at 31 December 2019 ............................................................ 13 Table 3: LoM Capital Provision Summary (in real terms) .......................................................................................................................... 16 Table 4: Direct On-mine Operating Costs ................................................................................................................................................. 16 Table 5: Indirect Steady State Off-mine Selling Costs ............................................................................................................................... 16 Table 6: Average Annual Steady State Operating Margin ......................................................................................................................... 17 Table 7: Sources of Information ................................................................................................................................................................ 18 Table 8: Site Inspections .......................................................................................................................................................................... 19 Table 9: Reporting Responsibility ............................................................................................................................................................. 19 Table 10: Other Experts ............................................................................................................................................................................ 21 Table 11: List of PE13155 Beacon Coordinates ........................................................................................................................................ 23 Table 12: List of Exploitation and Exploration Permits held by ABM .......................................................................................................... 23 Table 13: Survey Parameters for the Airborne Geophysical Survey .......................................................................................................... 26 Table 14: Frequency of QA/QC Samples Used ......................................................................................................................................... 41 Table 15: Tin Grade for CRMs used at Bisie and Number Analysed ......................................................................................................... 43 Table 16: Certified Values of CRMs used at Bisie ..................................................................................................................................... 43 Table 17: Summary of CRM Analyses ...................................................................................................................................................... 45 Table 18: Mean Values and Standard Deviation of Original Versus Field Duplicates at Bisie .................................................................... 45 Table 19: Mean Values and Standard Deviation of Original versus Pulp Duplicates at Bisie ..................................................................... 46 Table 20: Mean Values and Standard Deviation of ALS versus 2013 SGS Pulp Duplicates at Bisie ......................................................... 47 Table 21: Mean Values and Standard Deviation in ppm of ALS versus 2015 SGS Pulp Duplicate Assays at Bisie ................................... 49 Table 22: Mean Values and Standard Deviation in ppm of ALS versus 2016 SGS Pulp Duplicate Assays at Bisie ................................... 52 Table 23: Results of the BCS-CRM Number: 355 Assays for the 2016 Confirmation Assay Programme................................................... 53 Table 24: Check Sample Assay Results ................................................................................................................................................... 57 Table 25: Check Sample Assay Results – Re-assays ............................................................................................................................... 58 Table 26: Drillhole Cores for which Mineralisation was Visually Verified ................................................................................................... 58 Table 27: Summary of the raw validated sample data at Mpama North. .................................................................................................... 63 Table 28: Default Density Value Applied for each Rock Type in the Waste Zones at Bisie ........................................................................ 64 Table 29: Summary Statistics (de-clustered) of the Estimation 1m Composite Data ................................................................................. 68 Table 30: Summary Statistics (de-clustered) of the Estimation 1m Composite Data for Tin Accumulation................................................. 69



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Table 31: Summary Statistics (de-clustered) of the Estimation 1m Composite Data for Tin Accumulations Separated into High- and Low-

grade Populations ..................................................................................................................................................................................... 71 Table 32: Impact of Capping the Estimation Data ..................................................................................................................................... 72 Table 33: Semi-variogram Parameters – Main Vein System, Mpama North .............................................................................................. 74 Table 34: Indicator Semi-variogram Parameters – Main Vein System, Mpama North................................................................................ 75 Table 35: Above and Below Threshold Semi-variogram Parameters – Main Vein System, Mpama North ................................................. 75 Table 36: Block Model Prototype Parameters for Bisie, Mpama North (m) ................................................................................................ 77 Table 37: Dynamic Anisotropy Search Parameters for Bisie, Mpama North .............................................................................................. 78 Table 38: Validation of Block Model Filling ................................................................................................................................................ 78 Table 39: Search Parameters for Mpama North – Ordinary Kriging .......................................................................................................... 80 Table 40:Search parameters for Mpama North – indicator approach ........................................................................................................ 80 Table 41: Comparison between Drillhole and Model Data Values ............................................................................................................. 83 Table 42: Bisie Mpama North Zone Mineral Resource at 0.50% Sn Cut-Off Grade, 30 June 2019 ............................................................ 83 Table 43: Cut-off Grade Assumptions ....................................................................................................................................................... 84 Table 44: Mining Losses and Dilution ....................................................................................................................................................... 85 Table 45: Bisie Mpama North Zone Mineral Reserve at USD17,000/t Sn, at 1.6% Sn cut-off 31 December 2019 ..................................... 85 Table 46: Summary of Rock Mass Quality Data ........................................................................................................................................ 87 Table 47: Intact Rock Strength Testing ..................................................................................................................................................... 88 Table 48: Summary of Rock Strength Testing Results .............................................................................................................................. 88 Table 49: Summary of Hoek-Brown and Mohr-Coulomb Parameters ........................................................................................................ 89 Table 50: Summary of Intact Rock Strength.............................................................................................................................................. 89 Table 51: Rock Block, Rock Mass and Design Rock Mass Strength ......................................................................................................... 90 Table 52: General Systematic Support ..................................................................................................................................................... 90 Table 53: Ventilation Design Criteria and Occupational Exposure Levels ................................................................................................. 91 Table 54: Air Requirements based on Active Diesel-Powered Fleet .......................................................................................................... 92 Table 55: Open Stope Overbreak ............................................................................................................................................................. 95 Table 56: Development Dimensions and Overbreak ................................................................................................................................. 95 Table 57: Average Development and Stoping Advance Rates .................................................................................................................. 96 Table 58: Allocation of Excavation Type to Ground Support Class ............................................................................................................ 99 Table 59: Summary of Selected Mining Equipment ................................................................................................................................. 102 Table 60: Equipment Productivity/Efficiency ........................................................................................................................................... 104 Table 61: Layout Losses ......................................................................................................................................................................... 105 Table 62: Net Losses, Dilution and Pillar Recoveries .............................................................................................................................. 105 Table 63: Applying 1.6% Cut-Off Grade .................................................................................................................................................. 105 Table 64: RoM Feed to Plant .................................................................................................................................................................. 110 Table 65: Percent Mineral Resource Categories applied to Production Schedule ................................................................................... 110 Table 66: General Infrastructure – Completed and Planned .................................................................................................................... 116 Table 67: ABM’s Permitting Status ......................................................................................................................................................... 125 Table 68: LoM Capital Provision Summary (in real terms)....................................................................................................................... 128 Table 69: Direct On-mine Operating Costs ............................................................................................................................................. 128 Table 70: Mining Costs (2019) ................................................................................................................................................................ 128 Table 71: Indirect Steady State Off-mine Selling Costs ........................................................................................................................... 129 Table 72: Average Annual Steady State Operating Margin ..................................................................................................................... 130 Table 73: Proposed Activities, Timing and Costs of Mpama Ridge Exploration ....................................................................................... 132 Table 74: Proposed Activities, Timing and Costs of Mpama South Exploration ....................................................................................... 133



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Figure List

Figure 1: Corporate Ownership Structure* .................................................................................................................................................. 9 Figure 2: Location and Licences, DRC ........................................................................................................................................................ 9 Figure 3: Interpreted Geology Map of Permits .......................................................................................................................................... 10 Figure 4: Processing Plant Block Flow Diagram ....................................................................................................................................... 14 Figure 5: Location and Licences ............................................................................................................................................................... 22 Figure 6: Corporate Ownership Structure* ................................................................................................................................................ 22 Figure 7: Location and Licences ............................................................................................................................................................... 23 Figure 8: Busy Bee Charter Flight at Alphamin Airstrip ............................................................................................................................. 24 Figure 9: Average Annual Rainfall ............................................................................................................................................................ 24 Figure 10: Topography on the Bisie Ridge during Construction in 2018 .................................................................................................... 25 Figure 11: Lithological Units and Exploration Targets Delineated by Interpretation of Geophysical Data ................................................... 27 Figure 12: Regional Geological Map ......................................................................................................................................................... 28 Figure 13: Landsat Image showing the Limits of the Granite Pluton Adjacent to Bisie Ridge ..................................................................... 29 Figure 14: Bisie Stratigraphic Column ....................................................................................................................................................... 30 Figure 15: Structure and Geology Cross Section Interpretation of the Wedge Area at Section 9885615N................................................. 31 Figure 16: Mpama South Drill Collars, with Significant Intersects .............................................................................................................. 33 Figure 17: Schematic Section 9884650N showing Drillholes BGH001, BGC006, and BGH007 and 007A at Mpama South Prospect ....... 34 Figure 18: Schematic Section 9886000N showing Drillholes at Mpama North Prospect ............................................................................ 35 Figure 19: Mpama North Schematic Drillhole Locality Map showing Significant Intercepts and Drillhole Collars and Traces for Phase 1 and

Phase 2 Drilling ........................................................................................................................................................................................ 36 Figure 20: Mpama North Schematic Drillhole Locality Map showing Significant Intercepts for Phase 3 Drilling ......................................... 37 Figure 21: MINTEK Pilot Scale Shaking Table .......................................................................................................................................... 60 Figure 22: MINTEK Bench Scale Flotation cell ......................................................................................................................................... 60 Figure 23: Log Cumulative Frequency Plot of the Sample Sn Grade Data ................................................................................................ 63 Figure 24: Regression Analysis for SG versus Sn Grade .......................................................................................................................... 64 Figure 25: QQ Plot of Gas Pycnometer SG versus Wet-dry Method ......................................................................................................... 65 Figure 26: Isometric View of the DTM Created from the LIDAR Survey Data - View is Approximately to the North ................................... 66 Figure 27: Isometric Views of the Geological Model, View to East (top) and West (bottom) ...................................................................... 67 Figure 28: Histograms and Log Probability plots of the Geological Data ................................................................................................... 70 Figure 29: Indicator Semi-variograms for 80 Sn%t/m (Main Vein) ............................................................................................................. 75 Figure 30: Illustration of the Search Distance Derived from the Linear Portion of the Indicator Semi-variogram Model ............................. 79 Figure 31: Sections through Block Model and Drillhole Data illustrating Correlation between Model and Data .......................................... 81 Figure 32: Sectional Validation Plots for Sn% (model grades back calculated from the accumulations) .................................................... 82 Figure 33: Isometric View of the Mining Method ....................................................................................................................................... 93 Figure 34: Isometric Presentation of the Mining Method for Orebody Width up to 8m ............................................................................... 94 Figure 35: Isometric Presentation of the Mining Method for Orebody Width >8m but <15m ...................................................................... 94 Figure 36: Isometric Presentation of the Mining Method for Orebody Width >15m .................................................................................... 94 Figure 37: Schematic Access to Underground .......................................................................................................................................... 96 Figure 38: Current Development ............................................................................................................................................................... 97 Figure 39: Initial Development Design ...................................................................................................................................................... 97 Figure 40: Development Design Continued............................................................................................................................................... 98 Figure 41: Development Design End ........................................................................................................................................................ 98 Figure 42: Typical Stoping Layout ............................................................................................................................................................. 99 Figure 43: Stope Layout - Sliping ............................................................................................................................................................ 100 Figure 44: Mine Design with Stopes ....................................................................................................................................................... 100 Figure 45: Backfill Volumes .................................................................................................................................................................... 101 Figure 46: Loader (Epiroc ST-14) ........................................................................................................................................................... 102 Figure 47: Dump Truck (Epiroc MT42) .................................................................................................................................................... 102 Figure 48: Development Jumbo (Epiroc S1 D-DH) .................................................................................................................................. 103



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Figure 49: Support Jumbo (Epiroc BOLTEC 235) ................................................................................................................................... 103 Figure 50: Production Drill Rig (Epiroc SIMBA S7) .................................................................................................................................. 103 Figure 51: Applying 1.6% Cut-off Grade ................................................................................................................................................. 106 Figure 52: Mineral Resource Classification ............................................................................................................................................. 106 Figure 53: Mineral Resources with Mine Design ..................................................................................................................................... 107 Figure 54: Concentrator Block Flow Diagram .......................................................................................................................................... 112 Figure 55: General Site Layout ............................................................................................................................................................... 117 Figure 56: Refined Tin Consumption by End-use (2018) ......................................................................................................................... 122 Figure 57: Primary Tin Producers (2017) ................................................................................................................................................ 123 Figure 58: Forecast Production/Consumption Estimates to 2020 ............................................................................................................ 124 Figure 59: Adjacent Properties (DRC Mining Cadastre Portal) ................................................................................................................ 131 Figure 60: Plan View showing Tin Grade Distribution from Historical Drilling .......................................................................................... 133

Graph List

Graph 1: Blank Sample Results for Tin (Sn), Silver (Ag), Copper (Cu), Zinc (Zn) and Lead (Pb) ............................................................... 42 Graph 2: CRM Sample Results for Tin ...................................................................................................................................................... 44 Graph 3: Pulp Duplicate Results ............................................................................................................................................................... 46 Graph 4: Scatterplot for ALS versus SGS 2013 Tin Pulp Duplicate Assays ............................................................................................... 47 Graph 5: Control Chart for CRMs Assayed by SGS in 2013 ...................................................................................................................... 48 Graph 6: Scatterplot for ALS versus SGS 2015 Assays of 2014 Samples Tin Pulp Duplicate Assays ....................................................... 49 Graph 7: Control Chart for CRMs Assayed by SGS in May 2015 .............................................................................................................. 50 Graph 8: Scatterplot for ALS versus Set Point 2015 Assays of 2014 Samples Tin Pulp Duplicate Assays ................................................ 51 Graph 9: Control Chart for CRMs Assayed by Set Point in June 2015 ...................................................................................................... 51 Graph 10: Scatterplot of ALS versus SGS 2016 Assays of 2015 Samples Tin Pulp Duplicate Assays. by XRF......................................... 53 Graph 11: Scatterplot of ALS versus SGS 2016 Assays of 2016 Samples by Titration .............................................................................. 54 Graph 12: Control Charts for CRMs Assayed by ALS in 2016 for the Re-Assay Programme .................................................................... 55 Graph 13: Scatterplot for ALS 2016 Re-assays versus ALS Pre-2015 Assays .......................................................................................... 56 Graph 14: Scattergrams of Tin and Copper Check Sample Assay Results ............................................................................................... 57 Graph 15: Semi-variograms for Tin Grade (Main Vein) ............................................................................................................................. 73 Graph 16: Threshold Semi-variograms for Main Vein Zone <80 Sn%t/m .................................................................................................. 76 Graph 17:Threshold Semi-variograms for Main Vein Zone >80 Sn%t/m ................................................................................................... 77 Graph 18: Rock Mass Rating (RMR 89) for All Rock Types ...................................................................................................................... 87 Graph 19: Q-Index for All Rock Types ...................................................................................................................................................... 87 Graph 20: Africa Region K-Ratios ............................................................................................................................................................. 89 Graph 21: Backfill Volume ...................................................................................................................................................................... 101 Graph 22: Monthly Development Metres ................................................................................................................................................. 108 Graph 23: Annual Development Metres .................................................................................................................................................. 108 Graph 24: Production Profile ................................................................................................................................................................... 109 Graph 25: Average Monthly Product (Sn) ............................................................................................................................................... 109 Graph 26: Product (Sn) and Arsenic (As) ................................................................................................................................................ 110 Graph 27: Rom Feed with Grade ............................................................................................................................................................ 110



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1. EXECUTIVE SUMMARY

1.1. Property Description and Ownership Alphamin Resources Corp. (Alphamin), through its jointly owned subsidiary (Figure 1), Alphamin Bisie Mining SA (ABM), has legal title over three exploration permits (numbers: PR 10346; PR 5266; and PR 5267) and one mining permit (PE 13155) on which the Bisie Tin Mine (the Mine) is located.

Figure 1: Corporate Ownership Structure*

Source: Alphamin *Note: Through British Virgin Island (BVI) Companies

The licences are located in the North Kivu Province of the Democratic Republic of Congo (DRC). The Mine is located approximately 180km northwest of Goma, the capital of North Kivu Province and approximately 60km northwest of Walikale, a regional centre. The Mine can be accessed by road or by air from Goma (Figure 2).

Figure 2: Location and Licences, DRC

Source: Sound Mining (2019)

1.2. Status of Activities Considerable progress has been made on mining licence PE13155 since the previous NI43-101 Technical Report of March 2017 which presented the Feasibility Study results. The underground mine and processing plant at Mpama North were commissioned in 2019 and production has started. The December monthly year to date (YTD) forecast reports that 172kt of ore has already been processed at a grade of 5.30% Sn. The Mineral Resource for Mpama North was also updated in 2019 to account for these depletions, and to cater for a lowering of the depth from where it was believed the artisanal miners had impacted the Mineral Resource. These adjustments reduced the Mineral Resources as stated in the March 2017 Technical Report by approximately 340kt containing 9.8kt Sn. No further exploration drilling on the downdip extension to Mpama North occurred during the mine development phase.



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While the Mpama North mine is expected to reach the stated steady state production during 2020, planned process plant debottlenecking and capital initiatives are expected to facilitate additional throughput capacity at the plant. On this basis, an extensional drilling programme to Mpama North, and a drilling programme targeting a maiden Mineral Resource at Mpama South, have been planned and costed for implementation. Mpama South was a large centre of artisanal mining in the past, with extensive surface and rudimentary artisanal shafts located on the prospect, much like Mpama North originally. The mineralisation was delineated previously by seventeen (17) diamond drillholes which intersected similar tin mineralisation to Mpama North, albeit with higher grades of base metals. The proposed exploration and development program for 2020 will entail a boxcut, underground development and 6,000m of diamond drilling on Mpama South to facilitate a maiden resource declaration to support mining studies and a quick timeline to supplementing the current process plant feed so that the anticipated future spare capacity at the plant is fully utilised.

1.3. Geology and Mineralisation The Bisie Mine area is underlain by Kibaran Orogenic Belt lithologies, interpreted as being an inter-cratonic collision zone with different periods of extension and compression. The two units present at the mine are the lower Paleoproterozoic basement comprising Rusizian and Ante-Rusizian units composed mainly of dolomites, quartzites, amphibolites, mica schists and migmatite gneisses, as well as the upper Mesoproterozoic unit composed of dominant micaceous schists to red arenaceous phyllites with minor interbedded quartzites and amphibolites. Shales and conglomerates are also found in the upper parts of this sequence. Both units have been intruded by different generations of granites, starting in the Mesoproterozoic (± 1,375Ma) and continuing until 986Ma (Neoproterozoic) which are believed to be the last of the ‘tin granites’ and source of the numerous tin occurrences. ABM has conducted a detailed mapping exercise across the permits that included geological and structural interpretations from airborne geophysical campaign (Figure 3).

Figure 3: Interpreted Geology Map of Permits

Source: Data NRG; Interpretation G. J. Elliott, (2014)

The stratigraphic rock package hosting the deposit has been divided into five separate units, from hangingwall to footwall:

• Carbonaceous shale – dark grey to black, thinly bedded, fine grained, carbonaceous siltstone-shale

greater than 150m true thickness. Contains abundant quartz-tourmaline-carbonate veins and minor

pyrite;



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• Meta-sediments – pale grey, thinly bedded), fine-grained siltstone-shale 30m to 40m true thickness.

The host rock has been moderately silica-magnetite altered. Magnetite occurs as either discrete bands

(1-2cm), pervasive, disseminated alteration or stock-work veins;

• Quartz-sericite schist – in drill core appears more like a feldspar-rich, polymictic tuff. Pale to dark grey-

green, thinly bedded to massive, medium grained, feldspar-sericite rich tuff 80m to 90m true thickness;

• Mica schist – pale to dark grey, laminated to moderately banded, fine grained mica-rich schist 100m to

150m true thickness. Intensity of dark and light-coloured bands varies according to biotite-muscovite

content. Interpreted as a weak to moderately chlorite-talc-garnet alteration zone; and

• Amphibolite schist – dark green to black, moderately banded to massive, fine grained to porphyroblastic

(garnet), chlorite-amphobolite schist 20m to 30m true thickness. Moderate to intense chlorite-talc-

garnet alteration. Hosts the tin and base metal mineralisation at Bisie.

The mineralisation is associated with a steeply dipping (~65° east) north-south striking zone of intense chloritisation and shearing contained within micaceous schists. The main tin bearing chloritised zone is on average approximately 9m thick. Narrower subordinate zones occur several metres above and below the main zone in certain areas. The mineralisation occurs in the form of irregular high-grade veins of botryoidal cassiterite several tens of centimetres thick and lesser amounts of cassiterite blebs and vein fragments irregularly disseminated in the chlorite schist. The strength differential between lithologies often results in boudinage cassiterite lenses. The mineralised zone plunges approximately 35° to the north, although local steeper plunging high grade trends are evident. Copper, lead and zinc occur as chalcopyrite, galena and sphalerite in locally significant concentrations, together with silver. Two zones of mineralisation have been discovered at Bisie; these are known as Mpama North, which is the zone for which the Mineral Resource estimate described here-in applies, and Mpama South, which occurs about 750m to the south of Mpama North.

1.4. Status of Exploration The exploration permits PR10346; PR5266; and PR5267 are at much earlier stages of development than PE13155, though from the work conducted so far also show anomalous results of tin, precious and base metals and or in the past have supported artisanal miners to varying degrees. Work conducted on the exploration permits is limited to mapping, trenching, stream sediment sampling, geochemical soil sampling, ground and airborne geophysical surveys.

1.5. Mineral Resource Estimate Mr J C Witley of the MSA Group (MSA) (the independent Qualified Person (QP) for this Mineral Resource estimate), last visited the site in August 2015. During the site visits he conducted independent check sampling, inspection of the drillhole cores and sites, checked sample assays and confirmed the original sample assays were within reasonable limits for this style of mineralisation. The results of drilling at Mpama North are consistent with mineralisation observed by the QP. The assay results received from the primary laboratory have been confirmed by a quality assurance and quality control programme including duplicate assays completed by a second and third laboratory. The QP considers that the exploration work conducted by ABM was carried out using appropriate techniques for the style of mineralisation, and that the resulting database is suitable for Mineral Resource estimation. As at the effective date of this report, 34,963.55m in 171 diamond drillholes have been drilled and logged on the Mpama North deposit, as well as 27 PQ metallurgical holes. The Mineral Resource estimate was based on tin, copper, lead, zinc and silver assays and density measurements obtained from the cores of 122 NQ size diamond drillholes, which were completed by ABM between July 2012 and November 2015. In addition to the exploration drillholes, the split cores from 21 PQ size holes were used in the estimate. Mineral Resource estimation was carried out using Datamine Studio 3 software. A 0.35% Sn threshold was used to define the mineralised envelopes. Wireframes were constructed for the mineralised zones and a block model was constructed by filling the wireframe solids with parent cells of 20m in the approximate strike direction, 10m in the dip direction and 2m across the zone. Rotated block models were constructed in order to best fit the mineralised envelopes that dip steeply to the east. Semi-variograms were created for each of the estimated attributes and each attribute was estimated into the block models using ordinary kriging. Two statistical populations of tin grade were defined, the high-grade population being estimated separately from the lower grade and the estimates then combined. Search distances and orientations were aligned with the respective variogram range for each attribute estimated.



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For tin, accumulations of density and grade were used to appropriately reflect the relationship between tin grade and density and the tin grades were then back-calculated from the accumulation estimate. Outlier control was performed on the 1m composite data and included a restricted search distance for the high-grade tin population. The Mineral Resource estimate has been cut to a maximum elevation of 725RL allowing for the shallow area of Mpama North which has been partially depleted by artisanal mining and the quantity of remaining Mineral Resource in the affected area cannot be stated within reasonable limits. The maximum depth of the Mineral Resource is dictated by the location of the diamond drilling data. The Mineral Resource extends for approximately 700m in the northerly plunge direction and the deepest Mineral Resource reported is approximately 550m below surface, the mineralisation being open down-plunge. Estimates were extrapolated for a maximum distance of 20m up- or down-plunge from the nearest drillhole intersection. Extrapolation is minimal over most of the Mineral Resource as the up-and down dip limits have been well defined by the drilling. The Mineral Resource estimate has been depleted for underground development and stoping since the previous Technical report published in March 2017. The Mineral Resource estimate was completed by Mr J C Witley (B.Sc. Hons., M.Sc. (Eng.)) who is a geologist with twenty-seven years’ experience in base and precious metals exploration and mining as well as Mineral Resource evaluation and reporting. He is a Principal Resource Consultant for The MSA Group (an independent consulting company), is a member in good standing with the South African Council for Natural Scientific Professions (SACNASP) and is a Fellow of the Geological Society of South Africa (GSSA). Mr Witley has the appropriate relevant qualifications and experience to be considered a ‘Qualified Person’ for the style and type of mineralisation and activity being undertaken as defined in National Instrument 43-101 Standards of Disclosure of Mineral Projects. The Mineral Resource was estimated using The Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Best Practice Guidelines (2003) and is reported in accordance with the 2014 CIM Definition Standards, which were incorporated by reference into National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101). The Mineral Resource is classified into the Measured, Indicated and Inferred categories as shown in Table 1. The Mineral Resource (Table 1) is reported at a base case tin grade of 0.50%, which the QP considers will satisfy reasonable prospects for economic extraction given the high in-situ value of the mineralisation. All Mineral Resources are reported inclusive of Mineral Reserves.

Table 1: Bisie Mpama North Mineral Resource at 0.50% Sn Cut-off grade as at 30 June 2019

Category Quantity

(Mt) Quality (% Sn)

Content (kt Sn)

Cu (%)

Zn (%)

Pb (ppm)

Ag (g/t)

Measured Mineral Resource 0.33 4.75 15.6 0.22 0.12 0.006 1.4

Indicated Mineral Resource 3.99 4.59 183.4 0.32 0.16 0.010 2.8

Measured and Indicated 4.32 4.61 199.0 0.31 0.15 0.010 2.7

Inferred Mineral Resource 0.48 4.57 21.8 0.16 0.09 0.013 1.4

Source: MSA Notes:

• All tabulated data has been rounded and as a result minor computational errors may occur.

• Mineral Resources which are not Mineral Reserves have no demonstrated economic viability.

• The Mineral Resource is reported inclusive of Mineral Reserves.

• Alphamin has an 80.75% interest in ABM. The Government of the Democratic Republic of Congo (GDRC) has a non-dilutive, 5% share in ABM. The Gross Mineral Resource for the Mine is reported.

• Depleted by mining from mine surveys as at 30 June 2019 and an estimate of the extent of artisanal mining to 725mamsl.

1.6. Mineral Reserve Estimate Mr V G Duke of Sound Mining International Ltd (SMI) (the independent QP for this Mineral Reserve estimate), last visited the site in May 2019. During the site visits he conducted independent checks on the available supporting infrastructure, underground mining operation, processing plant and the Mineral Resource Management (MRM) function and responsibilities. The on-site observations were consistent with expectations formulated from reading the Definitive Feasibility Study (DFS) that was completed prior to construction of the mine. Appendix 4 contains the mine production schedule used to underpin the Mineral Reserve Statement as at 31 December 2019 (Table 2). It follows an assessment of the economic viability (Section 22) of the Mineral Resources that were scheduled for depletion before confirming them as Mineral Reserves.



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Table 2: Bisie Mpama North Zone Mineral Reserve at USD17,000/t Sn as at 31 December 2019

Classification

Quantity (Mt)

Quality (%)

Content (kt)

2019 2017 2019 2017 2019 2017

Proven Mineral Reserve 0.05 0.38 3.77 4.17 1.89 15.9

Probable Mineral Reserve 3.28 4.29 4.01 3.53 131.49 151.4

Total Mineral Reserves 3.33 4.67 4.01 3.58 133.38 167.3

Source: SMI Notes:

• The Mineral Reserve has been reported in accordance with the requirements and guidelines of NI-43101 and are 100% attributable to ABM.

• Apparent computational errors due to rounding are not considered significant.

• The Mineral Reserves are reported with appropriate modifying factors of dilution and recovery.

• The Mineral Reserves are reported at the head grade and at delivery to Plant.

• The Mineral Reserves are stated at a price of USD17,000/t Sn as at 31 December 2019.

• Although stated separately, the Mineral Resources are inclusive of the Mineral Reserves.

• No Inferred Mineral Resources have been included in the Mineral Reserve estimate.

• Quantities are reported in metric tonnes.

• The input studies are to the prescribed level of accuracy.

• The scheduled production includes 13.7% Inferred Mineral Resource, with most of these towards the tail end of the production forecast.

• The Mineral Reserve estimates contained herein may be subject to legal, political, environmental or other risks that could materially affect the potential exploitation of such Mineral Reserves.

The Mineral Reserve was estimated from a Life-of-Mine (LoM) design that supported appropriate sequencing and scheduling. Cognisance was taken of hydrogeology, geotechnical and ventilation criteria and various other modifying factors to ensure an acceptable level of accuracy. The planned mining method is comprised of open stoping with hydraulic backfill. The layout depends on individual stope characteristics relating to width, grade, ground conditions, ore type and other modifying factors. The mine design included pillars left in-situ to ensure stable and safe mining conditions underground. Differences between the new and past Mineral Reserve Statements is mostly accounted for by the change in mining method. The previous Sub-Level Caving design (SLC) included much higher levels of dilution and thus had a higher tonnage but lower average grade reserve than the new Mineral Reserve declared using the open stoping with hydraulic backfill method. In addition, Mineral Reserves decreased due to:

• some 172kt of ore at a grade of 5.30% Sn were depleted through mining in the interim;

• the Mineral Resource was trimmed at the top due to the updated estimated depth of artisanal working,

effectively removing one potential mining level from the mine design and reserves;

• the new Mineral reserves are declared at a 1.6% Sn cut-off versus the previous 1.4% Sn cut-off; and

• the SLC design had a high extraction ratio of 85%, whereas open stoping leaves behind pillars even

after reclamation, thus lowering the relative extraction ratio.

The production schedule is based on reasonable assumptions and calculations using appropriate software and best efforts. The production target (approximately 32ktpm) in terms of quantity mined could be exceeded or actual quantities mined could fall below the target. A variation of 10% either way will affect revenue, and therefore profitability, materially. It includes a planned grade according to the mine design and rate of mining and is directly related to the area being mined. Mining areas according to the life of mine plan as per the production schedule is essential to achieve revenue forecasts. Again, any deviation from the production schedule may affect revenue and profitability materially. Excessive dilution (in excess of planned dilution assumptions) will negatively affect the grade of ore sent for processing. Conversely, minimising dilution with good controls and mining practice, will improve the grade of ore and enhance revenue and profitability materially. It is therefore essential to implement the mine plan (based on the mine design and production schedule) by mining in the correct areas, at the planned grade, and in the correct quantities over time. Key assumptions, parameters and methods used to convert Mineral Resources to Mineral Reserves: Cut-off grade: A cut-off grade was calculated based on an assumed metal price (USD18,000/t Sn) and operating costs (USD172/t milled) with process recovery assumption (72%) and other factors, taking account of mining losses and geology. The calculation determined a cut-off grade of 1.6% Sn, which was used to limit the mine design to areas of the orebody where the in-situ grade exceeded 1.6% Sn. The mine scheduling further accounted for planned (5%) and unplanned dilution (15.7%), recovering some sill pillars, and previously mined out areas.



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Metal Price: The prevailing metal price over time will fluctuate according to supply and demand fundamentals. Any movement in price will affect the revenue stream directly and may result in a material improvement or deterioration, as the case may be. The operation will have no control over the metal price but may be able to implement contingency plans by mining in different areas with different grades or different quantities. Metallurgy: Metallurgical recoveries influence the revenue and profitability directly and therefore any improvement or deterioration in metallurgical recovery greater than 10% will have a material effect on profitability. The economic viability assessment assumed a metal recovery of 72%. The QP is satisfied that the Mineral Reserve estimate is consistent with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Best Practice Guidelines (2014) and is reported in accordance with the 2010 CIM Definition Standards. The Mineral Reserve was classified as Proven and Probable in accordance with the current Mineral Resource classification of Measured and Indicated. The terms ‘Mineral Resource’ and ‘Mineral Reserves’ have the meanings as ascribed by the Canadian Institute of Mining, Metallurgy and Petroleum, as the CIM Definition Standards on Mineral Resources and Mineral Reserves adopted by CIM council, as amended.

1.7. Mineral Processing The process plant (Figure 4) was designed to treat up to 390ktpa of ore to produce 15.5ktpa of tin concentrate at 62% tin, or 9.6kt tin contained.

Figure 4: Processing Plant Block Flow Diagram

Source: DRA

The processing plant (the Plant) entails the following processes:

• Crushing of Run-of-Mine (RoM) ore to -8mm;

• Screening of the crushed ore into -8mm +1mm and -1mm fractions;

• The -8mm +1mm is processed by jigging;

• Jig concentrate is milled to 80% -425µm and processed using gravity spiral concentrators and

shaking tables;

• Spiral concentrate is milled and sulphides removed by flotation to provide the bulk of the final

concentrate;



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• Spiral tailings are reground to 80% -106µm, sulphides are removed by flotation, and a final tin

concentrate produced by a second flotation stage;

• Jig tailings are discarded to a tailings stockpile, part of the tailing’s storage facility;

• The -1mm screened from the crushed ore is processed using gravity spiral concentrators. There

are two spiral concentration sections. The primary concentrate is combined with the jig spiral

concentrate for grinding and sulphide flotation. The secondary concentrate is combined with the

sulphide concentrates and tin flotation tailings to form a ± 10% tin low-grade concentrate;

• The -1mm spiral tails are thickened and discarded. Combined final concentrates are treated through

a magnetic separator to remove iron, then filtered and bagged for sale; and

• Combined final gravity concentrates and sulphide float product is treated through a magnetic separator to remove iron, then filtered and bagged for sale.

Lower than normal availabilities have been used in the design, owing to the remote location of the Plant. The crusher plant has been designed to run at 50% utilisation and the remainder of the Plant at 85%.

1.8. Environmental Considerations, Permitting and Social or Community Impact Mr H J Allison of Bahori Consulting (Pty) Ltd (Bahori) contributed to the Environmental and Social aspects at ABM. He was intricately involved in the development of the mine and has visited site several times from July 2017 to December 2019. ABM has the necessary Environmental permits to operate and has developed an environmental and social management plan. ABM will shortly submit its amended Environmental Impact Study (EIS), Mitigation and Rehabilitation Plan (MRP), and Environmental Management Plan for the Mine (PGEP or EMPP) to the Department responsible for the Protection of the Mining Environment (DPEM) in Kinshasa; for approval. The Mine is also committed to implementing its accredited Isometrix Integrated Risk System. The mine closure costs are reviewed and revised annually in line with best international industry practice. The QP considers the extent of all environmental, permitting and social or community impact liabilities, to have been appropriately provided for. The Mine is a significant employer in the Walikale Region. The mining operations contribute to the employment of local Congolese and the growth of the local and the DRC economies. ABM prioritises local employment and in the period since the last NI 43-101 submission (2017) achieved 100%. Stakeholder engagement is ongoing; through the Alliance Lowa ASBL (i.e. ABM Foundation) and this involves regular meetings with the community. Alliance Lowa ASBL is a partnership with all the local communities of Walikale in order to create social cohesion between the different tribes living in its area of influence. This has allowed key stakeholders from community to national government level and the private sector to collaborate this past year. ABM contributes to community development initiatives through Alliance Lowa ASBL and its partners, focussing on potable water supplies, alternative livelihood opportunities and business diversification, education, health care and local economic development projects.

1.9. Infrastructure and Logistics The mine is accessible by road from Goma, the capital city of North Kivu Province or via Uganda, as well as by airplane from regional locations to the local ABM airstrip which has been constructed and licenced for use on site. The main ports available to the Mine are Mombasa in Kenya and Dar es Salaam in Tanzania. The original site was established as an exploration camp in 2011. The facilities have expanded significantly over time and the Mine is now equipped with a clinic, emergency response room, security control room, offices, ICT infrastructure, accommodation, power generation, and messing facilities, Plant, tailings storage facilities, underground mine and an airstrip. This infrastructure supports the maintenance, mining, processing, tailings storage and mining operations.

1.10. Capital and Operating Costs Operations at Bisie Mine have reached commercial production but will ramp-up further during 2020 to a production target of 32ktpm as forecast by the 2019 LoM plan. Capital expenditure of approximately USD4 M has been allocated for on-mine capital in 2020. The capital requirements beyond 2020 have been catered for as a collective in the sustaining capital provision (Table 3).



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Table 3: LoM Capital Provision Summary (in real terms)

Source: Alphamin 2019

It is noted that the Mine is in operation and the current costs have been used as a baseline to project future costs. At a steady state of 32ktpm for the period January 2020 to December 2028 the on-mine operating costs are tabled below (Table 4).

Table 4: Direct On-mine Operating Costs

Description Cost

(USD/t ore) Cost

(USD/t Sn sold)

Mining 22.95 734.75

MRM Function 5.41 175.64

Processing 8.00 257.60

TSF 6.80 220.47

G&A 29.77 965.67

Diesel 20.05 650.40

Site Infrastructure 14.78 479.53

Transport and Duties 2.79 89.68

Wages and Salaries 56.83 1,843.70

Total LoM Operating Cost 167.38 5,417.44


The anticipated indirect unit costs for the period January 2020 to December 2028 are shown in Table 5.

Table 5: Indirect Steady State Off-mine Selling Costs

Description Cost

(USD/t Sn Sold)

Logistics - Mine to Delivery Point 753.84

Logistics - Delivery Point to Smelter 482.51

Road Maintenance 128.61

Treatment Charges 1,325.59

Clearing Agent Fees 74.67

Export Taxes and Fees 377.54

Marketing Commission 342.41

Government Royalty 548.60

Total 4,033.67


The average steady state (from January 2020 to December 2028) operating unit costs are expected to be USD9,451/t Sn sold (i.e. USD5,417/t Sn sold and USD4,034/t Sn sold). Capex over the corresponding period is USD366/t Sn sold, giving and all in sustaining cost USD9,817/t Sn sold.

1.11. Economic Assessment SMI has tested the economic viability of the Mineral Reserves as stated. The economic assessment (Table 6) takes account of all of the relevant technical, operating cost, capital expenditure and economic parameters but excludes any debt or financing structures. The assessment demonstrates that the Mineral Reserves are economically viable, with robust margins that remain positive over the full LoM plan.

Description Cost

(USD’000)

F2020 On-mine Capital 4,030

Mpama South Exploration Drilling/Studies 2,710

Exploration 530

Fine Tin Recovery Project 1,000

Sustaining Capital 48,830

Total Capital Provision 57,100



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Table 6: Average Annual Steady State Operating Margin

Rock Type Unit Total

Ore Mined kt 384

RoM Grade % 3.98

Development (ore and waste) m 2,500

Processing Recovery % 72

Payable Tin Sold t 11,004

Tin Price USD/t Sn Sold 17,000

Direct Steady State Unit Costs USD/t ore 167.38

Indirect Steady State Unit Costs USD/t Sn Sold 4,033.67

Gross Revenue USD M 187

On Mine Costs USD M 63

On Mine Margin USD M 124

Off Mine Costs USD M 44

Other Costs (including Mine closure provision) USD M 4

Pre-tax and Pre-funding Operating Margin (EBITDA) USD M 76

Source: SMI

The QP has also confirmed the economic viability of the Mineral Reserves as stated herein by inspecting the operations latest Business Plan and the associated post tax and pre-financing discounted cashflow at a real 31 December 2019 steady state price of USD17,000/t Sn sold. It was based on an underground mining operation with a processing throughput of 360ktpa to 390ktpa using Proven and Probable Mineral Reserves. Plant recoveries of 72% were applied for a final concentrate grading 62% Sn. The model considered corporate taxes of 30% and a minimum tax of 1% of turnover during the capital amortisation period. A 3.5% royalty on gross revenues less transport costs was also applied to the sale of tin in concentrate. The DCF returned a positive net present value over the full LoM at a real discount rate of 8%. The main uncertainties that may impact the cashflow forecasts are those related to tin price variations, the availability of the road to the mine site, forecast cost of fuel, logistics and salaries and wages and grade mined and processing recoveries. A sensitivity analysis shows that the value is most impacted by revenue followed by operating costs and capital costs.

1.12. Conclusions Mineral Resources of 4.80Mt at 4.60% (at a cut-off of 0.5% Sn) and Mineral Reserves of 3.33Mt and 4.01% (at a price of USD17,000/t Sn and cut-off of 1.6% Sn) are stated as at 30 June 2019 and 31 December 2019 respectively. These estimations are underpinned by a Mineral Resource model and a new LoM plan from 1 January 2020. This Plan takes account of some 126kt of ore that will have been mined since end June 2019 to end December 2019, at a grade of 5.83% Sn. The 2019 LoM and business plan have been developed to acceptable standards. The mine designs and layouts, dilution and economic parameters applied to the Mineral Resource estimates have been prepared to industry standard practices and economic viability of the Mineral Reserves has been clearly demonstrated to the satisfaction of the QP. Significant infrastructure has been installed on site and production has commenced. The mine utilises mining and processing practices that are proven in the mining and tin industry. ABM has demonstrated committed progress towards sustained operations at an increased steady state production target of 32ktpm run of mine for processing. This increase is due to a revised mining method that underpins the latest LoM plan. This new production target is anticipated to be reached in January 2020 and to continue until December 2029. This schedule does not include the upside potential of the proposed work programmes for 2020 including but not limited to: plant debottlenecking, installation of a fine tin recovery circuit, improved recoveries, anticipated additional production from Mpama North extension and Mpama South.

1.13. Recommendations The 2019 LoM plan indicates a robust production forecast in the context of appropriately calculated dilution, mining loss, and recovery factors. However, insufficient flexibility, geotechnical uncertainty and logistical challenges may impact operating margins, and these need to be proactively mitigated. ABM should continue to prioritise its exploration plans and this will result in increased flexibility for the planning and mining operations. Given the change in the mining method, management need to also closely monitor the ground conditions underground and the backfilling operation to limit delays in the mining sequence.



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2. INTRODUCTION

2.1. Issuer and Terms of Reference SMI has prepared this Technical Report for Alphamin. The purpose of this report is to provide an update following the change from a project in development to an operating mine, since the previous report entitled ‘NI 43-101 Technical Report – 23 March 2017 Updated Feasibility Study and Control Budget Estimate Report’. This update covers a revised LoM plan following construction, commissioning and a change in the mining method, and presents updated Mineral Resource and Mineral Reserve statements. Alphamin is a publicly traded company with a primary listing on the TSX Venture Exchange (TSX-V) under the symbol AFM and secondary listing on the Johannesburg Stock Exchange Alternative Exchange (JSE-AltX) under the symbol APH. The technical report has been prepared to comply with disclosure and reporting requirements set forth in the NI 43-101, Companion Policy 43-101CP, Form 43-101F1, and the CIM Definition Standards for Mineral Resources and Mineral Reserves adopted by the CIM Council 10 May 2014.

2.2. Principal Sources of Information This technical report has been compiled by SMI and it has been updated to reflect the change in development status and the change in mining method since the previous NI 43-101 Technical Report of 23 March 2017. Some sections remain largely unchanged with much of the text reproduced verbatim. It is based on technical information supplied by Alphamin and its subsidiary companies to the QPs and SMI for their independent review and comment. Table 7 summaries the key sources of information for the Technical Report.

Table 7: Sources of Information

Source: Alphamin

Mpama North on PE13155 is now considered by the QP to be a producing asset and Mpama South to represent an exploration project in a relatively advanced stage of exploration but pre-declaration of a maiden Mineral Resource. The remaining exploration licences are considered to represent early stage exploration projects which are inherently speculative in nature. However, the QP’s consider that the properties have been acquired on the basis of sound technical merit. The properties are also considered to be sufficiently prospective, subject to varying degrees of exploration risk, to warrant further exploration and assessment of their economic potential, consistent with the proposed programs. The QPs have provided consent for the inclusion of the Technical Report for Alphamin’s disclosure requirements and have not withdrawn that consent prior to lodgement.

2.3. Qualified Persons Site Visit Table 8 summarises the site visits to Mpama North carried out by the QPs.

Area/Discipline Source of Data

General, Ownership, Permitting etc. Alphamin

Geology Alphamin

Mineral Resources MSA Group

Geotechnical Data Latona Consulting, DRA

Mining and Life of Mine Plan Sound Mining and Alphamin

Process, Engineering and Infrastructure DRA and Alphamin

Economic Assessment SMI and Alphamin

Environment, Community Bahori Consulting and Alphamin



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Table 8: Site Inspections

Specialists Site Visit

V G Duke (SMI) Visited site as the QP responsible for the Mineral Reserves between 13 and 15 May 2019

J C Witley (MSA) Visited the site in July 2013, May 2014 and August 2015 as the QP responsible for the Mineral Resources

T Leach (Latona) Multiple site visits (August 2016 to July 2019) and contributed to the geohydrology and geotechnical review

H Allison Visited site multiple times from July 2017 to December 2019 and contributed to the environmental review

Source: SMI

2.4. Qualifications, Experience and Independence of Qualified Persons The QPs below, as defined by NI 43-101, were responsible for the preparation of this technical report. Table 9 lists the qualifications for each QP, as well as the section(s) of the report for which they are responsible.

Table 9: Reporting Responsibility

Technical Specialists Reporting Responsibility

Jeremy Witley, B.Sc. Hons., M.Sc. (Eng.), Pr.Sci.Nat, FGSSA Items 1.3, 1.4, 1.5, 6 to12, 14 and co-responsible

for items 1.12, 1.13, 2, 24.1 25 and 27

Vaughn Duke, Pr.Eng., PMP, B.Sc. Mining Engineering (Hons.), MBA, FSAIMM Items 1.1, 1.2, 1.6 to 1.13, 2 to 5, 13, and 15 to 27

inclusive

Source: SMI

2.5. Mining Legislation in the DRC The following review of mineral legislation in the DRC is summarised from Hubert André-Dumont (2013) and the Mining Code. The main legislation governing the mining industry in the DRC includes:

• Law number 007/2002 of 11 July 2002 as amended by Law 18/001 of 9 March 2018 (Mining Code);

• Mining regulation as enacted by decree 038/2003 of 26 March 2003, as amended and completed by

decree 18/024 dated 8 June 2018 (Mining Regulation); and

• Other decrees and ministerial orders which relate to subjects such as tax legislation, Valued Added

Tax (VAT), custom codes, labour, immigration, environmental protection and contractual or corporate

matters.

Mineral Property and Title All deposits of mineral substances in the DRC are held by the State, and the holder of mining rights gains ownership of the mineral products for sale. Under the 2002 Mining Code, mining rights are regulated by Exploration Permits (Permis de Recherches Minières or PRs), Exploitation Permits (Permis d’Exploitation or PEs) small-scale Exploitation Permits and tailings Exploitation Permits (Certificats d’Exploitation des Rejets or PERs). The DRC is divided into mining cadastral grids using a WGS84 geographic coordinate system. The grid defines uniform quadrangles or cadastral squares (carrés), each 84.955ha in area. The perimeter of a mining right is in the form of a polygon consisting of entire contiguous quadrangles subject to the limits relating to the borders of the DRC and those relating to reserved prohibited and protected areas as set forth on the 2003 Mining Regulations. Perimeters are exclusive and may not overlap subject to specific exceptions listed in the Mining Code and Mining Regulations. PEs are valid for 25 years and renewable for 15-year periods until the end of the mine’s life, provided the conditions laid out in the Mining Code have been met. The PEs were previously valid for 30 years when ABM converted its licence and therefore the new value does not apply to ABM. Granting of a permit is dependent on a number of factors that are defined in the Mining Code, including:

• Proof of the existence of a deposit which can be economically exploited, by presenting a feasibility study, accompanied by a technical framework plan for the development, construction, and exploitation of the mine;

• Proof of the existence of the financial resources required for execution of the project, according to a financing plan for the development, construction and exploitation of the mine, as well as the rehabilitation plan for the site when the mine is closed. This plan specifies each type of financing, the sources of planned financing and justification of their possible availability;

• Pre-approval of the project's environmental and social impact study (ESIS or EIS) and environment and social management plan (ESMP or EMPP);



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• Transfer to the DRC State 10% of the shares in the registered capital of the company applying for the licence. These shares are free of all charges and cannot be diluted. The requirements were previously only 5% when ABM converted its licence and thus the 10% value does not apply to ABM;

• Creation, upon each transformation, in the framework of a distinct mine or a distinct mining exploitation project, an affiliated company in which the applicant company holds at least 51% of the shares; and

• Filing of an undertaking deed whereby the holder undertakes to comply with the cahier des charges defining the social responsibility in relation to the local communities affected by the Mine’s activities.

The obligations to maintain the validity of the permit are set out in Articles 196 to 199 of the Mining Code and include presenting: Exploitation Permit evidence that the certificate at the beginning of works was duly delivered by the Cadastre Minier; evidence of payment of the annual superficiary rights payable per squares (carrés) and of the tax on the surface area of mining concessions; and by providing evidence of the capacity to treat (traiter) and transform the mineral substances in the DRC and filing an undertaking deed to treat and transform these substances within the Congolese territory. The PE, as defined in the Mining Code, grants the holder the exclusive right to carry out, within the perimeter over which it has been granted, and during its term of validity, exploration, development, construction and exploitation works in connection with the mineral substances for which the licence has been granted and associated substances (if the holder has applied for these). In addition, it entitles the holder, without restriction, to:

• Enter the exploitation perimeter to conduct mining operations;

• Build the installations and infrastructure required for mining exploitation;

• Use the water and wood within the mining perimeter for the requirements of the mining exploitation,

while complying with the requirements set forth in the EIS and the EMPP;

• Use, transport and freely sell the products originating from within the exploitation perimeter;

• Proceed with concentration, metallurgical or technical treatment operations, as well as the

transformation of the mineral substances extracted from the deposit within the exploitation perimeter;

and

• Proceed to carry out works to extend the mine.

A PE expires at the end of the appropriate term of validity if no renewal is applied for in accordance with the provisions of the 2002 Mining Code, or when the deposit that is being mined is exhausted

Sale of Mining Products Under the 2002 Mining Code and its amendments, the sale of mining products which originate from the exploitation permit is ‘free’, meaning that the holder of a PE may sell any licensed products to a customer of choice, at ‘prices freely negotiated’. The process plant at Mpama North produces a greater than 60% tin concentrate and as such can be freely exported.

Surface Rights Title The DRC State has exclusive rights to all land, but can grant surface rights to private or public parties. Surface rights are distinguished from mining rights, since surface rights do not entail the right to exploit minerals or precious stones. Conversely, a mining right does not entail any surface occupation right over the surface, other than that required for the operation. The 2002 Mining Code and its amendments, states that subject to any rights of third parties over the surface concerned, the holder of an exploitation mining right has, with the authorisation of the governor of the province concerned, and on the advice of the Administration of Mines, the right to occupy within a granted mining perimeter the land necessary for mining and associated industrial activities, including the construction of industrial plants and dwellings, water use, dig canals and channels, and establish means of communication and transport of any type. Any occupation of land that deprives surface right holders from using the surface, or any modification rendering the land unfit for cultivation, entails an obligation on the part of the mining rights holder to pay fair compensation to the surface right holders. The mining rights holder is also liable for damage caused to the occupants of the land in connection with any mining activity, even if such an activity has been properly permitted and authorised.



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3. RELIANCE ON OTHER EXPERTS Alphamin has warranted that it has openly provided all material information to SMI and the other relevant QPs who are responsible for sections of this technical report. To the best of Alphamin’s knowledge and understanding, the information provided is complete, accurate and true. The list of experts and their areas of expertise or opinion relied on in this technical report is presented in Table 10:

Table 10: Other Experts

Company Responsibility

Alphamin General items, introductions, security, administration, logistics; market studies and off-take contracts

Bahori Consulting Environmental, social and community aspects

Latona Consulting Geotechnical engineering consultancy

MSA Geology and mineralisation, exploration; drilling, sampling, QA/QC, and Mineral Resource estimation

SMI Mine design and production scheduling; Mineral Reserve estimation; and economic assessment

Source: SMI Note: Vaughn Duke as a QP has relied on information provided by Alphamin and other specialists (including T Leach (Latona) and H Allison (Bahori)) to complete various sections of this technical report. The Alphamin information has been from in-house mining and metallurgical specialists in the employ of Alphamin. The QP is satisfied himself of the competence of these specialists.



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4. PROPERTY DESCRIPTION AND LOCATION

4.1. Location Alphamin’s mineral assets are located in the Walikale District of the North Kivu Province of the DRC. The mine is approximately 180km northwest of Goma, the capital of North Kivu and approximately 60km northwest of Walikale, the regional centre. The mine can be accessed by road or by airplane from Goma (Figure 5).

Figure 5: Location and Licences

Source: Alphamin (2019)

4.2. Ownership ABM is owned 80.75% by Alphamin, a TSX-V listed company, 14.25% by the Industrial Development Corp., and 5% by the Government of the DRC. Figure 6 presents a simplified corporate structure.

Figure 6: Corporate Ownership Structure*

Source: Alphamin (2019) *Note: Through British Virgin Island (BVI) Companies

4.3. Mineral Tenure, Permitting, Rights and Agreements Alphamin, through its wholly owned subsidiary, ABM, has legal title over four exploration permits (Permis de Recherches PR 10346; PR 5266; and PR 5267) and one mining permit (Permis d’Exploitation PE 13155) on which the Bisie Tin Mine is located. This Technical Report primarily covers the PE13155 exploitation permit bounded by beacons with the geographical coordinates in Table 11.



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Table 11: List of PE13155 Beacon Coordinates

Beacon Longitude Latitude

1 27°39’00”E 01°03’00”S

2 27°39’00”E 01°00’00”S

3 27°46’30”E 01°00’00”S

4 27°46’30”E 01°07’00”S

5 27°44’30”E 01°07’00”S

6 27°44’30”E 01°05’30”S

7 27°42’30”E 01°05’30”S

8 27°42’30”E 01°04’30”S

9 27°41’00”E 01°04’30”S

10 27°41’00”E 01°03’00”S

Source: WAI Report

Figure 7 and Table 12 offer more detail on all of the permits.

Figure 7: Location and Licences

Table 12: List of Exploitation and Exploration Permits held by ABM

Licence Area Commodities Certificate Granted Expires

PE13155 151 Blocks

(±128.96km2) Tin and gold CAMI/6923/15 3 February 2015 3 February 2045

PR5266 42 Blocks

(±36.30km2) Tin, gold, copper, zinc, lead and

silver CAMI/CR/2935/2006 24 November 2014 23 November 2022

PR5267 380 Blocks

(±322.83km2) Tin and gold CAMI/CR/2936/2006 29 September 2006 23 May 2023

PR10346 77 Blocks

(±65.76km2) Coltan and gold CAMI/CR/6635/12 2 July 2009 1 July 2022


Although the exploitation permit for PE 13155 does not currently include silver, copper, lead and zinc, ABM’s applications for these additional elements on both PR 5266 and PR 10346 were accepted under Article 59 of the Mining Code and Articles 111-114 of the Mining Regulations, and the file is properly recorded with CAMI (DRC Cadastre Minier). Licences cover an aggregate area of approximately 610km2.

4.4. Force Majeure There is a history of Force Majeure concerns in the DRC, and ABM experienced complications in this respect with PR 5270, PR5267 and PR4246, as reported in the previous technical report (March 2017). These specific issues have since been resolved with no outstanding Force Majeure concerns. It is noted that ABM has relinquished PR4246 and in the processing of relinquishing PR5270. This reduces ABM’s exposure in this regard.



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5. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1. Accessibility The mining operation can be accessed by road with Uganda being the preferred route, aeroplane or helicopter from Goma, the capital city of North Kivu Province. There is 32km access road to site from the N3 national road. An airstrip (Figure 8) offers accelerated access to the mine.

Figure 8: Busy Bee Charter Flight at Alphamin Airstrip

Source: Alphamin

5.2. Climate and Physiography The climate is a tropical humid one with two wet seasons; April to June and September to January. The average annual rainfall is 1,300mm, with an average temperature range of 12°C, from a minimum of 18°C to a maximum of 30°C. The annual average temperature is 25°C. Operations can carry on all year round. Climatic data has been obtained for the town of Goma, some 180km South East of the mine site and the source of the closest reliable data, measured at the town’s airport as shown in Figure 9.

Figure 9: Average Annual Rainfall

Source: https://en.climate-data.org/



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The topography in the area (Figure 10) comprises an undulating hilly terrain with moderate to steep slopes and perennial streams and rivers in the valleys. The altitude ranges from 500mamsl to 870mamsl. The area is covered by dense equatorial forest vegetation and bedrock outcrops are poor. The slopes are covered by a mixture of clay soils and eluvial rubble, and clay (kaolin) soils with alluvium occur in the valleys.

Figure 10: Topography on the Bisie Ridge during Construction in 2018

Source: Alphamin

5.3. Local Resources and Infrastructure Nearby local resources are scarce and so basic supplies, fuel and consumables are imported via Uganda by road, which also serves staff living not too far from the mine and who return home on rotational cycles through Goma. The airstrip is used for special or emergency items as well as staff with more distal home rotational legs. Mining personnel are sourced from surrounding areas wherever possible as part of the agreement with local representatives and the Alliance Lowa ASBL initiative. Water is available from springs and rivers for processing and domestic use. Power is not available from the national power grid and is instead sourced from a bank of diesel generators. Telecommunications are available through cellular service providers with erected reception on the ridge. There are no significant population centres in the or on the mine. The area is remote in relation to the local communities, which reduces the potential impacts on the community. Disseminated small villages are present along the N3 national road between the regional centre Walikale and the Mine, scattered Artisanal Small-scale Miners (ASM) and prospectors are located throughout the broader area. Walikale is 75km away by road. There is sufficient space for use as storage areas and / or for waste disposal areas. The DRC State has exclusive rights to the land over the whole area, though, the Mining Code states that subject to any rights of third parties, the holder of an exploitation permit has the right to occupy the land necessary for mining and associated industrial activities, including the construction of industrial plants and dwellings, water use, dig canals and channels, and establish means of communication and transport of any type.



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6. HISTORY Tin bearing gossan was discovered on the Bisie ridge in 2002 and soon became the focus of large scale illegal artisanal small-scale mining operations. There are no known records of prior ownership nor the type, amount, quantity, or general results of exploration, development work or production carried out on the concession areas. There are no known historical Mineral Resource or Mineral Reserve estimates before ABM’s involvement either. Primary cassiterite was mined by artisanal means from two main areas, Mpama South and Mpama North, located along 1.5km of a ridge which extends over more than 14km over two of ABM’s mining and exploration licences (PE13155 and PR5266). The deepest artisanal mining workings were reported from Mpama North where the main tunnel was focused on a high-grade mineralised chute, locally known as the ‘Salon’ which reached an estimated depth of 70m. In 2006, Mining and Processing Congo SPRL (MPC), a former subsidiary of Kivu Resources, was granted four exploration permits (PR’s) within the Bisie Project area (PR4246, PR5266, PR5267 and PR5270). MPC was the first company to be granted legal title over the area and began work at Bisie in October 2006. At the end of November 2006. MPC carried out five reconnaissance exploration campaigns over an eighteen-month period from October 2006:

• The first campaign established an exploration camp and commenced geological mapping, focusing on

cassiterite and columbite-tantalite (coltan) occurrences and rock chip sampling;

• The second campaign continued with the geological mapping on a regional basis as well as mapping

artisanal workings;

• The third campaign included more detailed geological mapping and further mapping of the surface and

underground artisanal workings;

• The fourth campaign included the review of mapped data and artisanal workings and rock samples

were collected for mineralogical and petrological examinations; and

• The fifth campaign continued geological mapping including several traverses across the ridge as well

as continued mapping of hard rock and alluvial artisanal workings.

In August 2011, Alphamin acquired a 70% interest in the Bisie Tin Project, and by late 2011, Alphamin had acquired 100% of Bisie from Kivu Resources. In November 2015, Alphamin Resources Corp. announced that it had entered into an agreement with the Industrial Development Corporation of South Africa Limited (IDC) pursuant to which the IDC invested USD10 M directly in ABM, in three tranches. The final tranche was received in May 2016, and the IDC now holds 15% of the Class A shares of ABM (effective 14.25% economic interest).

6.1. Geophysics An airborne geophysical campaign was carried out over four of the licences: PE 13155 (at the time a part of PR 5266), PR 5266, PR 5267 and PR 10346 (Figure 11). The survey was flown in six separate blocks at various line spacing, collecting magnetic and radiometric data. The survey was flown by New Resolution Geophysics (NRG), using the Xplorer system. Survey parameters are shown in Table 13.

Table 13: Survey Parameters for the Airborne Geophysical Survey

Survey Name Country Line (km)

Line Spacing (m x m)

Line Direction

UTM Zone Average Sensor

Height (m)

Terrain

PR5266_PR10346 Block DRC 2,461 250m x 2,500m 90 355 28,2 Rugged

Bisie Block DRC 611 50m x 500m 90 355 31,4 Rugged

Wedge Block DRC 594 150m x 1,500m 90 355 27,6 Rugged

Access Block DRC 1,752 50m x 500m 90 355 28,7 Rugged

Tin Granite DRC 607 100m x 1,000m 90 355 29,2 Rugged

TG2 DRC 109 100m x 1,000m 90 355 29,8 Rugged

Block 5267 DRC 2,368 150m x 1,500m 90 355 30,6 Rugged

Umate Block DRC 457 75m x 750m 90 355 32,7 Rugged

Total 8,959

Source: MSA



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The data were processed by NRG and the resulting raster images were used to produce an updated geological and structural map of the area. The results were also used to delineate new potential targets by finding structures with the same magnetic and radiometric signature as the Mpama North and South prospects.

Figure 11: Lithological Units and Exploration Targets Delineated by Interpretation of Geophysical Data

Source: Geophysical data procured by NRG (2014) and interpreted by G. J. Elliott, Geophysical Consultant

6.2. Soil Geochemistry Soil samples of approximately 2kg each were collected from the B soil horizon approximately 60cm below surface. Samples were taken at 50m intervals along east-west lines spaced 200m apart. All samples were air dried and sieved to -2mm and then -180µm before being analysed on site using a Niton hand held XRF analyser. Approximately 100g of the sample of sieved soil passing 180µm was placed in a plastic zip lock bag. The sample was placed on the Niton stand and analysed using an analysis time of 30 seconds on the soil setting.

6.3. Rock Sample Geochemistry Forty-four rock samples were collected from the area to the west of the Bisie ridge on PR 5266 and PE 13155, over the interpreted massive intrusive pluton. The samples’ lithologies were mostly gneissic and granitic, with one mafic and one quartzite sample. These samples did not show significant tin content, but provided an insight into the geochemistry of the massive intrusion.

6.4. Stream Sediment Sampling Twenty-six stream sediment samples were collected to the northwest of the Mpama North prospect. Five-kilogram samples were collected and then concentrated through panning before being sent for analysis. Six of the samples reported results of over 5,000ppm Sn in concentrate. The samples were part of a sterilisation programme over potential tailings storage facility sites. Areas where positive results were achieved have been dismissed for a more favourable location for tailings disposal.



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7. GEOLOGICAL SETTING AND MINERALISATION

7.1. Geological Setting and Mineralisation The Mine is underlain by the Kibaran Orogenic Belt lithologies, interpreted as being an inter-cratonic collision zone with different periods of extension and compression (Pearl, 2011). Units present in the area include a Paleoproterozoic Basement comprising poorly exposed dolomites, quartzites, amphibolites, mica schists and migmatite gneisses and Upper Mesoproterozoic rocks comprising micaceous schists and arenaceous phyllites with minor interbedded quartzites and amphibolites. Shales and conglomerates are also found in the upper parts of this sequence (Figure 12).

Figure 12: Regional Geological Map

Sources: Geological map: shape file extracted from the «Carte Géologique et Minière de la République Démocratique du Congo » (Musée Royal de l’Afrique Centrale, Tervuren), Licences coordinates: CAMI.

Units have been intruded by different generations of granites, which started in the Mesoproterozoic (± 1,375Ma) and continued until the last so-called ‘tin granitei’ intrusion at about 986Ma (Neoproterozoic). These intrusions are commonly believed to be the source of the numerous tin occurrences in the region, with the granites themselves containing elevated levels of tin.

Regional Structure On a regional scale, the metamorphic rock units generally strike northwest-southeast. The ridge hosting the Bisie mineralisation strikes north-south as far as the Oso River in the north, after which the strike of the ridge changes to the northwest-southeast. Regional scale folding is evident in the satellite imagery (Figure 13). The source of the mineralising fluids is thought to be the massive plutonic intrusion to the west of the ridge with recent drilling and geophysics interpretation demonstrating there are multiple intrusive granitic phases, some of which are probably the source of the tin-bearing mineralising fluids at Bisie.



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Figure 13: Landsat Image showing the Limits of the Granite Pluton Adjacent to Bisie Ridge

Source: Landsat data (NASA), interpretation by John Barrett. (August 2012) Note the Oso River running from east to west in the upper part of the Figure.

7.2. Property Geology Detailed mapping carried out between August and September 2015 included surface mapping of rock types and three dimensional (3D) geological modelling using drilling data along the Mpama Ridge. The surface mapping data were entered into an Excel spreadsheet then plotted using GIS software Map info (Version 12.5). The 3D geology modelling of drilling data was conducted section by section between 9885500N and 9886140N using Micromine 2014 (Version 15). Queries were identified with the respective boreholes being re-logged, validated and corrections incorporated into the drilling database.

Stratigraphy The stratigraphic rock package hosting the Bisie deposit has been divided into five separate units (Figure 14). From hangingwall to footwall, a general description of the major units is as follows:

• Carbonaceous shale (CBSH) – dark grey to black, thinly bedded (0.5cm to 2cm), fine grained,

carbonaceous siltstone-shale greater than 150m true thickness. Contains abundant quartz-tourmaline-

carbonate veins and minor pyrite;

• Meta-sediment (METS) – pale grey, thinly bedded (0.5cm to 2cm), fine-grained siltstone-shale 30m to

40m true thickness. The host rock has been moderately silica-magnetite altered. Magnetite occurs as

either discrete bands (1cm to 2cm), pervasive, disseminated alteration or stock-work veins;

• Quartz-sericite schist (QSSH) – in drill core appears more like a feldspar-rich, polymictic tuff. Pale to

dark grey-green, thinly bedded (~1cm) to massive, medium grained (1mm to 5mm), feldspar-sericite

rich tuff 80m to 90m true thickness; and

• Mica schist (MSCH) – pale to dark grey, laminated to moderately banded (0.5cm to 5cm), fine grained

mica-rich schist 100m to 150m true thickness. Intensity of dark and light-coloured bands varies

according to biotite-muscovite content.



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• Amphibolite schist – dark green to black, moderately banded to massive, fine grained to porphyroblastic

(garnet), chlorite-amphobolite schist 20m to 30m true thickness. Moderate to intense chlorite-talc-

garnet alteration. Hosts the tin and base metal mineralisation at Bisie.

Figure 14: Bisie Stratigraphic Column


The stratigraphic package dips east between 60° and 75° and appears to steepen down-dip. The Mpama North ridge crest is more or less defined by the AMPH which probably resists erosion due to the massive, coherent nature of the rock and high chlorite content. The QSSH appears more susceptible to weathering and erosion due to its’ high feldspar-sericite content. The base of complete oxidation (BOCO) is approximately 10m along the ridge crest and approximately 50m in the OSSH. Overall the BOCO averages 30m.

Local Structure Surface mapping was combined with drillhole lithological logging and interpreted in three dimensions to assist in Mineral Resource estimation and mine planning. Structural mapping was combined with and draped over the results of a topographical Lidar survey. The structural mapping suggests the geometry of the local structures reflects the regional scale structure, with an overall north-south trending shear zone cross-cut by east-northeast trending faults. 3D geology modelling using drillhole data identified a dominant structure comprising a brittle-ductile shear zone, with a number of hangingwall and footwall fault splays, running parallel to mineralisation along the Mpama ridge. The interpretation suggests a number of the splay structures fan out towards the north while others appear to be anastomosing. South of 9885870N, the main structure occurs in the hangingwall of the deposit while north of 9885870N, the structure occurs in the footwall. Foliation is the most common feature observed in outcrop and is a reliable indicator of folding and faulting with a change in orientation suggesting either of the two. The regional foliation strikes north-south to north-northwest, dips steeply east-northeast between 60° and 65° and is more or less parallel to bedding. Structural interpretation using stereonets also suggests that the cassiterite veins hosted within the amphibolite unit occur parallel to the regional foliation. Folding is commonly observed on an outcrop scale. 3D modelling in cross section suggests the Wedge area, at the south end of Mpama North, has been folded into a broad, south plunging (20°-25°) anticline (Figure 15). The anticline structure is well represented on section 9885615N with the host lithologies in the footwall and hangingwall folded into a broad, open fold and the fold axis dipping steeply to the west. Further north and south, the fold structure weakens with the deposit becoming more tabular and dipping to the east.



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Figure 15: Structure and Geology Cross Section Interpretation of the Wedge Area at Section 9885615N


Parasitic folding is commonly seen in outcrop and in cross section interpretation from the 3D geology modelling. The folding appears to post-date mineralisation with the host rock and cassiterite veins folded about the westerly dipping fold axis. In contrast to the regional foliation which dips east, the fold axis dips to the west suggesting the folding is related to a different deformation event. Similarly, to foliation, quartz veining is a good indicator of faulting and orientation of stratigraphy with foliation parallel veins provides reliable orientation data in areas of limited outcrop. A number of quartz veins were mapped in the Wedge area and at Mpama North. In the Wedge area several north-northwest trending quartz veins cross cutting the foliation were projected below surface and correlated in drill holes west and east of the mineralised zone. The western quartz vein appears to truncate mineralisation up-dip and to the west and similarly the eastern quartz vein coincides with the termination of mineralisation down-dip towards the east. Several quartz veins were mapped in the main area of the Mpama North deposit however these are not considered to have any effect on the geometry of the mineralisation.

Mineralisation Tin mineralisation at Bisie is hosted within a north-south striking, eastward dipping, amphibolite unit along the Mpama Ridge. Structural and mineralogical evidence from drill core indicates cassiterite was emplaced first, followed by copper mineralisation in the form of chalcopyrite and bornite, then by lead and zinc mineralisation occurring as sphalerite and galena. There is also evidence of late-stage quartz-chalcopyrite veining which cross-cuts the mineralisation with veins trending north-northwest. Chlorite alteration is widespread and appears to be the result of late stage fluids entering the system. The host rocks are predominantly highly chlorite-altered amphibolites and fine to medium grained, mica-chlorite-garnet schists. The tin and copper mineralisation is predominantly found in zones dominated by intense chlorite alteration, however, cassiterite mineralisation with no chlorite alteration has been intersected in the hangingwall and footwall vein zones hosted in MSCH.



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The dominant structural control on mineralisation is a north-south trending, brittle-ductile shear zone that runs parallel to the cassiterite mineralised zones. It occurs predominantly as a single structure, with minor hangingwall and footwall splays particularly in the upper, more brittle parts of the structure. Tin mineralisation seems to be concentrated in two high-grade chutes, referred to by Alphamin as the upper high-grade chute and lower high-grade chute. Mineralisation between these two chutes is lower grade as these areas contain narrower, more widely spaced cassiterite veins. Both chutes run parallel to each other and plunge to the north at approximately 35°. The structural controls on these chutes is not well understood but is considered to be related to the shear zone. The highest-grade tin mineralisation has so far been found at Mpama North with cassiterite veins becoming wider and higher grade with depth. At Mpama South, mineralisation appears lower grade comprising at least four vein sets, however Mpama South is not well explored, particularly in the deeper areas.

7.2.3.1. Mpama South Prospect Drilling at Mpama South (19 drill holes for 3,364m) has been exploratory in nature and aimed to determine the extent and nature of the mineralisation. Two distinct mineralised zones have been intercepted, with an upper zone showing well-developed lead, zinc and silver mineralisation, and a lower zone rich in tin and copper. Figure 16 shows holes BGH001, 006 and 007A in cross-section intercepting the two zones (Figure 17). Drilling to date at Mpama South has delineated 260m of strike returning tin intercepts of 10m or more grading in excess of 0.7% Sn, for example:

• 32.2m at 0.76% Sn from 106.9m including 22.05m at 1.02% Sn

• 11m at 1.48% Sn from 71m, including 2.5m at 5.76% Sn.

• 32.8m at 2.46% Sn from 192.2m; and

• 1.65m at 6.57% Sn from 320m and 23.65m at 1.15% Sn from 325m. Copper and rare earth element (Cerium and Lanthanum) mineralisation is commonly associated with the tin rich zone, along with elevated lead and zinc. Significant results include:

• 11m at 0.88% Cu from 72m including 4.5m at 1.74% Cu;

• 35m at 0.77% Cu from 53m including 10m at 1.67% Cu; and

• 10.1m at 1,042g/t Ce from 162m. Drilling further identified a zone rich in silver, zinc and lead mineralisation in the Mpama South target area. Best results include:

• 19m at 197g/t Ag from 61m;

• 17.7m at 14.11% Zn from 61m including 13m at 18.09% Zn; and

• 14.75m at 10.82% Pb from 61m.

A drilling program has been designed with the aim of delineating a Mineral Resource at Mpama South in the 2020 financial year. The intention is to then immediately initiate the mining studies necessary for the production of additional ore to fill the anticipated future spare processing capacity in the plant as it is further de-bottlenecked and/or expanded.



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Figure 16: Mpama South Drill Collars, with Significant Intersects




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Figure 17: Schematic Section 9884650N showing Drillholes BGH001, BGC006, and BGH007 and 007A at Mpama South Prospect


7.2.3.2. Mpama North As at the effective date of this report, 34,963.55m in 171 diamond drillholes have been drilled and logged, as well as 27 PQ metallurgical holes. The latest phase of exploration was completed in November 2015. In addition to the first nine exploratory holes drilled during the first campaign, 28 holes (and one re-drill) were completed during the second program and 134 holes (and ten re-drills) in the latest phase of drilling. The drillholes were drilled from east to west along section lines spaced 50m apart with closer spaced drilling in the shallower areas. Along the section lines the drillholes intersected the mineralisation between 25m and 50m apart in most of the Mineral Resource area (Figure 20), with drilling being sparser, up to 100m apart, in the shallower portions of the down-plunge area. The ‘Wedge’ area is situated to the south of the Mpama North main zone and it was drilled at 25 m spacing along dip and strike. The mineralisation appears to be offset to the east from the main zone of mineralisation. Further drilling will be required to declare a Mineral Resources in the Wedge area despite positive results from the earlier exploration. This is because more definition is needed on the structural complexity, which precluded the declaration of mineral resources. An underground drilling campaign from production development planned on 3 Level, 4 Level, 5 Level, will contribute to unlocking the potential of this southern extension of Mpama North. In addition, the ability to extend the Mpama North resource at depth and along strike to the north also present a promising opportunity. The mineral resources at Mpama North were limited to the most northern drill fence which was also the deepest due to the size of the quasi-man-portable exploration drill rigs used in the previous drilling campaigns. Ongoing underground development deeper and closer to this northern extension provides an equally promising opportunity to use the underground diamond drill for exploration activities to extend Mpama North mineral resources. It is planned to develop a comprehensive drilling program and expansion programme when development has reached 7 Level which will be used as a drilling base (in the footwall) to delineate further extension north along strike and down dip for depth extension.



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An example of a drilled section is shown in Figure 18, Figure 19 and Figure 20, which show the collar locations and significant intercepts for Phases 1, 2 and 3 of drilling at Mpama North.

Figure 18: Schematic Section 9886000N showing Drillholes at Mpama North Prospect


All of the available drilling results as at the effective date of this report have been included in the Mineral Resource assessment, there being no outstanding data of significance. 27 PQ sized holes totalling 2,783.5m were drilled to obtain a metallurgical sample. Twenty-four of the holes were assayed and were drilled in three clusters approximately 25m apart. Within the clusters, the PQ holes were drilled approximately 5m apart. The drilling at Mpama North included nine holes to investigate metallurgical variability in different locations across the deposit (Figure 20). The exploration drillholes intersected variable mineralisation mostly within a persistent chlorite schist unit with massive cassiterite veins hosting the majority of the tin. The structural data suggests a shallow northerly plunge, which has been confirmed by infill drilling and drilling in the down-plunge area towards the north. Copper mineralisation at Mpama North is consistently associated with tin mineralisation, generally overlapping it. While copper grades are generally low, some holes returned relatively high-grades. BGC035 reported 14.8m at 1.03% Cu within an interval of 29m at 3.3% Sn.



36

Figure 19: Mpama North Schematic Drillhole Locality Map showing Significant Intercepts and Drillhole Collars and Traces for Phase 1 and Phase 2 Drilling




37

Figure 20: Mpama North Schematic Drillhole Locality Map showing Significant Intercepts for Phase 3 Drilling

Source: Alphamin, May 2016

8. DEPOSIT TYPES The tin deposit is a cassiterite-bearing stock-work or vein system adjacent and possibly distal to underlying source granite. From the composition of the mineralisation, it was concluded that the mineralisation has a low temperature origin, with a probable granitic source. Fluorine and lithium are absent from the ore forming fluids and base metal sulphides scarce in the cassiterite. The deposit has up to 0.5% rare earth elements (REE) and very high-grade tin (with some sample assays reaching 60% Sn). This may indicate the source granite to be at depth below the surface. Three-dimensional modelling of the geology presents the deposit as a number of steeply dipping tabular sheets of variable grade mineralisation consisting of irregular veins and disseminations of cassiterite that is complex on a small scale but on a larger scale the tabular sheets could be combined to reflect a relatively consistent Main Vein for modelling purposes.



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9. EXPLORATION

9.1. Reconnaissance work by Mining and Processing Congo Sprl (2006-2008)

Commencing in October 2006, MPC carried out five reconnaissance exploration campaigns over an eighteen-month period. On the first reconnaissance trip (October 2006), a camp was established and reconnaissance mapping commenced. Work was often interrupted by the 85th FARDC Brigade who were controlling between 300 and 400 artisanal miners. The trip was intended to locate and identify cassiterite and coltan (columbite-tantalite) occurrences and to undertake preliminary geological mapping and reconnaissance rock chip sampling. On the second trip, a two-day jungle walk was used as an opportunity to carry out field checking of the regional geology of the areas surrounding the Bisie Ridge and the artisanal workings. Further geological mapping and mapping of the artisanal workings was carried out. Trip three entailed detailed surface geological mapping and mapping of the underground artisanal workings. The work was conducted over a twelve-day period, using Global Positioning System (GPS) to record topographic information and note lithological contacts. The GPS was also used to measure and map the extent of the artisanal workings at surface. Three external geological consultants, Sam Mawson, Jerry Fiala and one consultant from Steffen, Robertson and Kirsten Consulting (SRK), accompanied employees of the MPC on the fourth trip. The consultants reviewed the existing geological data, examined the extent of the artisanal workings at Bisie and collected rock samples required for mineralogical and petrological examinations. In April 2008, MPC undertook its’ fifth reconnaissance campaign over 20 days. This included continuing reconnaissance geological mapping of the Bisie vein mineralisation, checking of outcrop localities and recording new artisanal workings on the Bisie ridge and in the alluvial ground in the surrounding valleys. Several traverses were undertaken from Bisie in order to map the host rocks for some distance both to the east and west of the deposit. Despite poor outcrop, the limited exposures indicated that the dominant unit is pale blue-green micaceous schist with thin beds of grey shale, lenses of clean quartzite and elongated irregular-shaped bodies of coarse-grained amphibolite, either weathered or altered to chlorite. A small weathered exposure of quartz-feldspar-hornblende granite was noted, located approximately 4km west of the Bisie workings. The granite was described as pink, medium-grained porphyritic granite with a slight fabric manifested by parallel elongation of the feldspar porphyroblasts. MPC also carried out a literature search and compiled a detailed review of information lodged at the Royal Africa Museum in Tervuren, Belgium. An additional search for information was carried out at the archives in Kalima. In both cases, information pertaining to regional lithological and mineral potential was found. There was little information on detailed geology and mineralisation, with only regional geological maps found. All known mineralisation in the Alphamin exploration permits was discovered post-colonialism (MPC Report, 2008). MPC purchased Landsat 7 Imagery and undertook detailed imagery interpretation. The interpretation was used to refine regional relationships between the Paleo- to Mesoproterozoic metasediments and the Neoproterozoic granites. On the existing regional geological maps, the metasediments show a strong northwest-southeast fabric, however from the Landsat Imagery, it was noted that the regional fabric of metasediments in the middle of the permit area is north-south, complicated by the intrusion of a granite pluton. This was confirmed by reconnaissance mapping (MPC Report, 2008). The granite contact is visible in the Landsat Imagery, approximately 3km to the west of the Bisie workings. It is thought to be the source of the quartz veining and the associated cassiterite mineralisation. The pure quartz veining and the dominant tin mineralisation are considered to represent part of the last phases of post-granite metallogenesis (MPC Report, 2008). MPC collected 38 rock samples in June 2007 as part of their reconnaissance work at Bisie, to be used for mineralogical and petrological examination. Most of the samples were collected from primary bedrock mineralisation in outcrop, artisanal workings and cuttings. Additional samples were collected from weathered mineralisation, fresh metasedimentary host rocks and alluvial gravels and talus slopes below the artisanal workings. These samples were sent to GET Company Ltd, Prague, Czech Republic, for analysis. Twenty-one samples were analysed for 31 elements including tin, tantalum, niobium and tungsten and REE’s using inductively coupled plasma mass spectrometry (ICP-MS) finish on a lithium borate fused bead. Fourteen other base metals were determined using ICP-MS on a four-acid digest. Thirty-nine polished thin sections were prepared from 24 of the samples for microscope mineralogical study. Eighteen thin sections were studied using the microprobe CAMECA SX100 at Masaryk University, Brno, Czech Republic (Breiter, 2007).



39

According to the studied samples, the Bisie mineralisation is unusual and very different to other classic tin deposits, specifically with reference to the presence of up to 0.5 % REE and the high-grade tin (50-90 % SnO₂ in some samples). From the mineral composition of the mineralisation, a low temperature origin is suggested. The mineralising fluids do not contain fluorine or lithium and base metal sulphides are scarce, suggesting a deep-seated granitic source. It was thought at the time that the shape of the mineralised bodies would be irregular, making exploration using modern exploration methods difficult.

10. DRILLING A total of 195 HQ size (63.5 mm diameter core) and NQ size (47.6 mm diameter core) diamond drillholes have been drilled on the Bisie Project to date, angled between 60° and 75° to intersect the mineralisation at ~right angles; 171 drillholes at Mpama North, 19 drillholes at Mpama South and five drillholes at Marouge. Six of the holes drilled at Mpama North were sterilisation holes relating to planned infrastructure, 13 were geotechnical holes and one was a water monitoring hole. A further 27 PQ size (85.0 mm diameter core) holes were drilled in order to obtain a sample for metallurgical test-work. Drillhole collars were positioned using a hand held GPS unit. Final collar surveys were completed with a digital GPS (DGPS), using reference base stations, by a certified surveyor. The final collar positions of five recent holes had not been surveyed as at the time of the Mineral Resource estimate. The collar positions of the sterilisation, geotechnical, water monitoring and abandoned holes were located by hand-held GPS. Down-hole surveys were conducted using a Reflex EZ-Trac digital survey instrument and multi shot receiver. All drillholes were surveyed at 30m intervals down the hole with the exception of the sterilisation, geotechnical, water monitoring and abandoned holes that were not surveyed. In the Phase 3 drilling, BGC065 and BGC075 were drilled 320m and 170m further south respectively from BGC033. BGC065 (drilled at -60°) intersected two narrow cassiterite veins of 1.03% Sn over 1m and 2.38% Sn over 0.70m between 61.0m and 66.75m. BGC075 (drilled at -60°) intersected a single cassiterite vein of 3.57% Sn over 0.8m at 75m down the hole. This confirms the potential to intercept significant mineralisation over the 650m zone separating the two prospects. This could be tested further in a future drilling program. Five holes were drilled at the Marouge Prospect, 2km to 3km south of Mpama South. No significant mineralisation was intersected.

10.1. Drill Sample Recovery Drill core was recovered from the core barrels and washed and placed into core trays at the drill site. Core recovery was measured and noted along with length of run and depth on the core blocks, which were inserted in the core trays at the end of each run. Each core tray was numbered and marked with the relevant drillhole number at the drill site. On completion of the drillhole, the core was airlifted back to the camp by helicopter where core recovery was checked during the marking process. Core recovery was generally very good within the mineralised zone and country rock. Most core loss was noted in the upper 65m and total core recovery averaged 89%, however at depths greater than 65m core recovery was over 95%.



40

11. SAMPLE PREPARATION, ANALYSES AND SECURITY The zones to be sampled were selected on the basis of visible mineralisation and Niton XRF analysis. A five to ten metre zone, where observations and Niton analysis did not identify any mineralisation, was sampled either side of any significant mineralisation identified. A nominal sample length of one metre was used, which was varied in order to honour individual zones of mineralisation intensity and lithological boundaries. The minimum sample length used was dependant on the nature of the mineralisation with the smallest sample being 0.15m in length, although the majority of the samples (all except 19) were at least 0.30m long. Maximum sample lengths did not exceed 2.0m. After sample mark up and completion of lithological logging the core was photographed. The core was then split longitudinally in half using a water-cooled diamond blade core saw along a cut line designed to separate the two halves of the core equally throughout the sample length. Core was held in closable Almonte core cutting boxes in order to reduce core loss and increase cutting accuracy. On completion of cutting of the sample, the cut core was replaced into the core tray with the half to be sampled facing upward. The half of the core sampled was the left-hand side relative to the low point of the core foliation. On completion of the cutting process, the core trays were moved to the sampling shed. The geologist or field assistant used the pre-printed sample number list to place the core samples (starting from the end of the sample and moving backward to the beginning of the sample) into pre-marked plastic sample bags. Cross checks were completed against the marked bags and sample numbers against the sample list to mitigate against sample swaps. A sample ticket corresponding to the number on the sample bag and sample sheet was placed inside the plastic bag which was then placed in the numbered sample bag along with the core sample and sealed with a cable tie. After final cross checking, the sample bags were tied closed using a plastic cable tie and then placed into poly-weave sacks which were in turn sealed with plastic cable ties. Each poly-weave sack was marked with a number and the sample numbers contained within, as well as the address of the laboratory. The poly-weave sacks were then transported to ABMs’ office in Bukavu or Goma via the contracted helicopter and packed into cardboard boxes which were labelled and shipped to the laboratory via air freight. At the laboratory, samples were first checked off against the list of samples supplied and then weighed and oven dried. The dried samples were crushed to 70% passing 2mm, from which a 250g split was taken and this was pulverised to 85% passing -75µm from which a sample for analysis was taken. Samples were submitted to the South African National Accreditation System (SANAS) accredited ALS Chemex (ALS) laboratory in Johannesburg where samples were analysed for tin using method code ME-XRF05 conducted on a pressed pellet with 10% precision and an upper limit of 10,000ppm. The upper limit was reduced to 5,000ppm from the second campaign onwards. Over limit samples were sent to ALS in Vancouver for ME-XRF10 which uses a Lithium Borate 50:50 flux to create a fused disk that is analysed by XRF with an upper detection limit of 60% and precision of 5%. Method code ME-ICP61 (HF, HNO3, HClO4 and HCl leach with ICP-AES finish) was used for 33 elements including base metals. ME-OG62, a four-acid digestion, was used on ore grade samples for lead, zinc, copper and silver. From January 2014 onward, high-grade samples were flagged and the laboratory was instructed to clean the crushers with coarse blank after such samples. From 2015, the pulveriser bowls were also cleaned with blank material following each flagged high-grade sample. In addition to elemental analysis, ALS conducted Specific Gravity (SG) measurements by gas displacement using a multi-pycnometer with a precision of +-10%. SG was also carried out routinely on-site on core sub-samples using Archimedes principle of weight in air versus weight in water. In the QP’s opinion, the sample preparation, security and analytical procedures used for the Bisie samples are adequate for the style of mineralisation at Bisie.



41

12. DATA VERIFICATION

12.1. Analytical Quality Control and Assurance conducted by Alphamin One certified reference material (CRM) sample was inserted on average for every 30 samples, with quarter core duplicate samples (field duplicate), or an instruction for a pulp duplicate analysis (laboratory duplicate) to be performed, and blank samples being inserted at the same frequency. From BGC038 onwards (i.e. the third phase of drilling) field duplicates were no longer taken in favour of the pulp duplicates. At least one of the blank samples was inserted immediately after a high-grade sample. The CRMs were sourced from Ore Research and Exploration of Australia (CRM numbers OREAS, 36, 140, 141, 142 and 163) and from Bureau of Analysed Samples Ltd of Great Britain (CRM number BCS-CRM 355). QA/QC samples were submitted on an approximately 11% basis. For the first two phases of drilling, quarter core samples were sent for assay along with the primary half-core samples, the duplicate assay immediately following the primary assay. At the end of the second phase of the exploration program, the pulp rejects were collected from ALS and 150 of these, together with CRMs, from 13 drillholes spread over the Mpama North area were sent to Société Générale de Surveillance Lakefield in Johannesburg (SGS) for check assay for tin, copper, zinc and lead. SGS is a SANAS accredited laboratory. In the third phase of drilling (i.e. BGC038 to BGC171), the quarter core field duplicates were discontinued and only pulp duplicates were used by the laboratory on instruction from Alphamin. Although these cannot be considered truly blind sample duplicates, they are useful for understanding the precision of the assaying as there is clear separation between the preparation and analytical processes at the laboratory. At least two CRMs, two duplicates and two blank samples were included with each mineralised intersection. At least one of the blank samples was inserted immediately after a high-grade sample. In 2015, pulp rejects were again collected from ALS for verification assay. 173 samples from 16 drillholes were sent for verification assay of tin, copper, zinc, lead and silver at SGS. In addition, 99 different samples were sent to Set Point (Johannesburg), which is a SANAS accredited laboratory, for verification assay of tin, copper, zinc, lead and silver. CRMs were included with the pulp rejects. In 2016, thirty-three pulp rejects were collected from ALS and sent for verification assay of tin at SGS. In 2016, a total of 98 pulp duplicates were re-labelled and re-submitted to ALS for re-assay. These were from samples originally assayed by ALS in 2013 and 2014 and covered the 1.5% to 60% Sn grade range. Included with the pulp duplicates were thirteen CRMs. Approximately 18% of the samples submitted were used to monitor the quality of the assaying. A summary of the quantity and proportion of Quality Assurance/Quality Control (QA/QC) samples used by Alphamin, outside of the laboratory’s own internal QA/QC, is shown in Table 14.

Table 14: Frequency of QA/QC Samples Used

Drillhole Number

Core Samples

Blanks CRMs Pulp Duplicates Core Duplicates Second

Laboratory Assays

(Number) (Number) (%) (Number) (%) (Number) (%) (Number) (%) (Number) (%)

BGH 001-009 1,005 20 2 26 3 16 2 9 1 0 0

BGC 001-009 546 0 0 92 4 57 2 46 2

150 6

BGC 010-037 2,046 87 4

BGC 038-146 4,174 184 4 186 4 140 3 0 0

173 at SGS in 2015 with 27 CRMs.

99 at Set Point with 15 CRMs

METBGC07-27 1,002 47 5 41 4 35 4 0 0

BGH 010-016 526 21 4 26 5 16 3 0 0

BGC147-171 540 25 5 29 5 17 3 0 0 33 6

Source: MSA

Results of the Blank Sample Analysis The blank samples used at Bisie were originally sourced locally from quartz veins. In July 2013, the blanks were changed to gabbro from Bukavu. The assays indicate that the gabbro contains fairly consistent trace quantities of copper (50ppm) and zinc (100ppm) and so the blank was changed back to quartzite in June 2015 (Graph 1). Some potential contamination is evidenced by tin (less than 0.20%). Elevated silver grades in the blank samples were more prevalent from the gabbro samples. Elevated copper, lead and zinc grades were not significant.



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The degree of potential contamination will not significantly affect the outcome of the tin grade estimate given the high-grade nature of the deposit and that all of the contamination is considerably below the tin cut-off grades considered for this deposit. Silver and lead grades are low and local inaccuracies may arise as a result of the levels seen in the blank samples.

Graph 1: Blank Sample Results for Tin (Sn), Silver (Ag), Copper (Cu), Zinc (Zn) and Lead (Pb)

Source: MSA

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14

BG

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16

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C1

23

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27

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31

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35

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R00

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136

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36

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46

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50

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61

BG

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65

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GC

167

Zn

pp

m

Blank Bisie Zn



43

Results of the CRM Sample Analysis Four different CRM’s were used with tin grades spanning the range of the Bisie mineralisation; however, several of the CRMs inserted in the earlier drilling phase were not of sufficient mass to be assayed for tin after analyses for other elements were complete and so tin assays were not completed for all of the CRM samples submitted Table 15. In the later phase of drilling, two additional CRM’s were introduced that were not certified for tin but did between them have silver, copper, zinc and lead grades that better covered the grade ranges of these elements at Bisie.

Table 15: Tin Grade for CRMs used at Bisie and Number Analysed

CRM Name Accepted Mean Sn

Grade (ppm) Standard Deviation Number Used

Number of Sn Assays Reported

BCS-CRM No355 314,200 2,200 93 86

OREAS36 Not certified for Sn

OREAS140 1,755 122 131 110

OREAS141 6,061 339 58 47

OREAS142 10,400 500 74 64

OREAS163 Not certified for Sn

Source: MSA

Table 16 shows the certified values for the elements of interest at Bisie for the CRMs used.

Table 16: Certified Values of CRMs used at Bisie

CRM Name

Sn (ppm)

Ag (g/t)

Cu (ppm)

Zn (ppm)

Pb (ppm)

Mean SD Mean SD Mean SD Mean SD Mean SD

BCS-CRM No355 314,200 2,200 - - 850 80 590 60 120 20

OREAS36 - - 10.17 0.63 151 5 42,300 600 5,790 130

OREAS140 1,755 122 1.03 0.11 1,529 82 1,706 123 26.7 0.8

OREAS141 6,061 339 1.58 0.11 2,453 98 3,637 178 59 3.8

OREAS142 10,400 500 1.22 0.12 1,466 65 2,436 82 54.3 3.8

OREAS163 - - 4.303 0.6 17,600 700 108 11 495 28

Source: MSA

Results from the CRM assays indicate that the tin assays for the lower grade CRMs (0.18%, 0.61% and 1.04% Sn) are accurate and precise. One assay for OREAS 142 reported a grade consistent with that expected from OREAS 141 and is presumed to have been mislabelled. One assay of OREAS 140 reported 27ppm Sn, which is suspected to be an incorrectly labelled blank sample. One assay of OREAS 141 reported 23ppm Sn, which is suspected to be an incorrectly labelled blank sample and another reported 1,885ppm Sn, which is expected to be a mislabelled OREAS 140CRM. The mean assay of the high-grade standard (BCS-CRM number: 355) by ALS prior to 2015 was 33.98% Sn, which is 8.1% higher than the accepted mean and all tin assays were well outside of the three standard deviation acceptance limit Graph 2. The mean assay of BCS-CRM number: 355 for the first two months of 2015 was 32.43% Sn, which is only 3% higher than the accepted mean. Most assays from March 2015 were close to the accepted mean and largely within tolerance, although with a slight high bias relative to the accepted mean.



44

Graph 2: CRM Sample Results for Tin

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Sn Sn Acc mean +2SD -2SD +3SD -3SD

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45

Source: MSA

A summary of the results of the CRM analyses is shown in Table 17. A failure was deemed an assay that fell outside of three standard deviations (SD) of the accepted mean value.

Table 17: Summary of CRM Analyses

CRM Name Sn Ag Cu Zn Pb

BCS-CRM Number: 355

All pre-March 2015 Fail Bias 8.1% high January to February 2015 Fail Bias 3% High Post February 2015 10/38 Fail, slightly high

N/A No failures No failures No failures

OREAS 036 N/A No failures 1/22 slight fail,

no overall bias 7/19 failed Biased low

2/22 failed Slight low bias

OREAS 140

One incorrectly labelled sample 28/107 failed. Slight low bias

3/131 failed 3/131 failed 70/131 failed Low bias Low-grade CRM – poor precision

OREAS 141 Two incorrectly labelled samples 15/56 failed.

Slight low bias 3/56 failed 2/56 failed

Slight low bias 10/56 failed Slight Low bias

OREAS 142 One incorrectly labelled sample 5/60 failed 3/60 failed 5/60 failed –

Slight low bias 13/60 failed Slight low bias

OREAS 163 N/A No failures No failures No failures 5/24 failed

Slight Low bias

Source: MSA

Aside from the high-grade tin CRM and the high-grade zinc CRM (OREAS 036, large proportions of failures were noted where the grade of the CRM is low and therefore the impact on the Mineral Resource estimate is negligible.

Results of the Core Duplicate Sample Analysis The 55 quarter core field duplicates showed differences in individual assays outside of 20% for numerous samples. The percentage difference between the mean of the two sample sets was also high Table 18.

Table 18: Mean Values and Standard Deviation of Original Versus Field Duplicates at Bisie

Sn (%)

Ag (g/t)

Cu (ppm)

Zn (ppm)

Pb (ppm)


Original 2.37 8.05 1.87 3.86 2,124 4,341 1,759 3,552 74 115

Field Duplicate 2.64 8.54 1.94 4.13 2,213 4,704 2,213 4,704 71 97

% Mean Difference 11 6 4 7 4 8 14 9 4 17

Source: MSA

It should be noted that poor precision can be expected with small samples (quarter core) taken from the nuggety irregular vein style tin mineralisation at Bisie, as also confirmed by the poor precision found with the quarter core check sample assay results. The use of coarse duplicates as a quality check was discontinued in Phase 3 of the drilling.

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46

Results of the Pulp Duplicate Sample Analysis Reasonable precision was noted between the pulp duplicates for high-grade tin. Above 2.5% Sn, only two out of 20 pairs had a mean difference of more than 10% (Graph 3). High percentage differences were noted for tin at grades of less than approximately 0.03% Sn. Poor repeatability was noted for silver at grades of less than 1g/t, copper at less than 100ppm, zinc at less than 300ppm and lead at less than 100ppm. The poor repeatability at such low grades will not significantly impact the Mineral Resource estimate. Overall the average grades of the original and pulp duplicate assays were similar there being no significant bias (Table 19).

Graph 3: Pulp Duplicate Results

Source: MSA

Table 19: Mean Values and Standard Deviation of Original versus Pulp Duplicates at Bisie

Sn (%)

Ag (g/t)

Cu (ppm)

Zn (ppm)

Pb (ppm)


Original 1.14 4.42 1.54 6.57 1,037 2,203 873 2,270 120 840

Pulp Duplicate 1.12 4.36 1.57 7.11 1,046 2,228 870 2,250 123 893

% Mean Difference 2 1 2 8 1 1 0 1 2 6

Source: MSA

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Bisie Pulp Duplicates Cu

Cu +10% +20% -10% -20%



47

Results of the Second Laboratory Check Analysis For the first and second phases of drilling, the SGS tin assays compared well with the ALS assays although they tended to be more variable (Table 20 and Graph 4), as evidenced by the slightly higher standard deviation. It is noted that the bias noted with the high-grade CRM assays was not repeated between SGS and ALS, there being no significant bias between SGS and ALS for the high-grade assays. The copper assays compared well between the two laboratories. For both zinc and lead, SGS reported lower grades than ALS.

Table 20: Mean Values and Standard Deviation of ALS versus 2013 SGS Pulp Duplicates at Bisie

Sn (%)

Ag (g/t)

Cu (ppm)

Zn (ppm)

Pb (ppm)


ALS 4.28 9.94 - - 2,874 3,937 1,381 3,248 66 99

SGS 4.30 10.49 - - 2,796 3,930 1,269 3,221 52 61

% Mean Difference 0 5 - - 3 0 8 1 25 47

Source: MSA

Graph 4: Scatterplot for ALS versus SGS 2013 Tin Pulp Duplicate Assays

Source: MSA

Five samples of BCS-CRM number: 355, OREAS 141 and OREAS 142 were included with the SGS assays. These indicate that the lower grade assays by SGS were accurate, although there was a tendency for SGS to under report the high-grade CRM assays, which is contrary to ALS which over assayed the grade of this CRM in 2013 (Graph 5).

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Oct 2013 - SGS vs ALS Sn assays



48

Graph 5: Control Chart for CRMs Assayed by SGS in 2013

Source: MSA

For the third phase of drilling (up to BGC075) 200 pulp samples prepared by ALS were sent for verification assay at SGS in Johannesburg. Together with the pulp duplicates of the core samples, a number of certified reference material samples were included. The SGS assays were lower than the ALS assays, particularly for the higher-grade samples (above 250,000ppm) (Graph 6).

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SGS Oct 2013 Standard BCS-CRM No355 - Sn

CRM Sn Assay CRM value -2SD +2SD -3SD +3SD

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SGS Oct 2013 Standard OREAS142 - Sn




49

Graph 6: Scatterplot for ALS versus SGS 2015 Assays of 2014 Samples Tin Pulp Duplicate Assays

Source: MSA

However, this cannot be taken as firm evidence that the ALS assays were necessarily too high as SGS had a strong tendency to under-assay the CRMs that were included with the duplicate pulp samples Table 21. There is a 6% mean difference between the ALS and SGS tin assays. SGS zinc and lead assays are significantly higher than those of SGS and copper assays are almost the same (Graph 7).

Table 21: Mean Values and Standard Deviation in ppm of ALS versus 2015 SGS Pulp Duplicate Assays at Bisie

Sn (%)

Ag (ppm)

Cu (ppm)

Zn (ppm)

Pb (ppm)


ALS 5.95 12.2 2.1 1.5 2,984 4,680 1,613 2,429 54 95

SGS 5.59 11.79 3.7 3.0 2,946 4,579 2,058 2,783 67 80

% Mean Difference 6 4 31 22 1 2 24 14 22 16

Source: MSA

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SGS Sn ppm

May 2015 SGS assays of 2014 samples - SGS vs ALS Sn assays



50

Graph 7: Control Chart for CRMs Assayed by SGS in May 2015

Source: MSA

As the results of the second laboratory verification were inconclusive, due to the low assay bias for SGS as shown by the CRMs, a second verification laboratory was chosen; Set Point (Johannesburg), which is a SANAS accredited laboratory. The Set Point assays were significantly lower than those by ALS for grades above 10,000ppm (Graph 8). It should be noted that Set Point has a lower detection limit of 100ppm for Sn and therefore all pairs were excluded from the scatterplot for which Set Point assays returned below the detection limit. For many of the assays that ALS returned values of well over 100ppm (in many cases over 1,000ppm), Set Point returned values below detection limit.

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SGS May 2015 Standard BCS-CRM No355 - Sn


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SGS May 2015 Standard OREAS 142 - Sn




51

The two assays by Set Point of CRM OREAS 140 (1,755ppm Sn) both returned values below detection limit (<100ppm Sn). The assays of OREAS 142 by Set Point were below the accepted mean, although within tolerance. Two of the three assays of OREAS 141 were well below tolerance limits and the other was significantly higher (Graph 9). The assays of the high-grade (CRM BCS-CRM number: 355) returned values higher than the accepted mean with the exception of one assay that returned a low value of 212,000ppm Sn, which may be an incorrectly labelled field sample. On the basis of the CRM assays by Set Point, the comparison with ALS is inconclusive.

Graph 8: Scatterplot for ALS versus Set Point 2015 Assays of 2014 Samples Tin Pulp Duplicate Assays

Source: MSA

Graph 9: Control Chart for CRMs Assayed by Set Point in June 2015

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June 2015 - Setpoint vs ALS Sn assays

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Setpoint June 2015 Standard BCS-CRM No355 - Sn


212000



52

Source: MSA

Thirty-three pulp rejects from the 2015 drilling were sent to SGS in 2016 for verification assaying of tin. Nine of the samples were above the upper detection limit of 5% for XRF at SGS and were assayed using a titration method. The samples assayed by XRF compared well with the ALS assays, however the samples assayed by titration returned lower values than ALS (Table 22 and Graph 10).

Table 22: Mean Values and Standard Deviation in ppm of ALS versus 2016 SGS Pulp Duplicate Assays at Bisie

Percentage Tin SGS by XRF

Percentage Tin SGS by Titration

Mean Standard Deviation Mean Standard Deviation

ALS 1.01 1.36 23.72 14.69

SGS 1.02 1.34 20.75 12.46

% Mean Difference 0.6 1.4 13.3 16.4

Source: MSA Note that the value that was reported at the upper detection limit by ALS was excluded from these statistics

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Setpoint June 2015 Standard OREAS142 - Sn

CRM Sn Assay CRM value -2SD -3SD -3SD +3SD

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Setpoint June 2015 Standard OREAS141 - Sn

CRM Sn Assay CRM value -2SD -3SD -3SD +3SD

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53

Graph 10: Scatterplot of ALS versus SGS 2016 Assays of 2015 Samples Tin Pulp Duplicate Assays. by XRF

Source: MSA

This is with the exception of one assay that was above the ALS upper detection limit of 60%, for which SGS returned a value of 68.2% (Graph 11). Two of the re-assays were of the high-grade CRM (BCS-CRM number: 355). In both cases SGS returned considerably lower values than the certified mean of the CRM, whereas the ALS assay was consistent with the certified mean value (Table 23).

Table 23: Results of the BCS-CRM Number: 355 Assays for the 2016 Confirmation Assay Programme

D19779 (% Sn)

D19681 (% Sn)

ALS 31.70 31.50

SGS 25.30 29.60

Certified Mean 31.42 31.42

Source: MSA

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2016 SGS assays of 2015 samples - SGS vs ALS Sn assays. SGS by XRF



54

Graph 11: Scatterplot of ALS versus SGS 2016 Assays of 2016 Samples by Titration

Source: MSA

For the 2016 assays, it can be concluded that the ALS assays have been confirmed by SGS for those conducted using XRF by SGS. For those assays performed using titration by SGS (>5% Sn), the comparison is inconclusive, as the SGS assays of the high-grade CRM were considerably lower than the certified mean value.

Re-assaying of Pre-2015 Samples by ALS In order to understand any error in the high-grade tin assays prior to 2015, a selection of pulp rejects from pre-2015 were re-labelled and sent to ALS for re-assay in 2016. Ninety-eight samples across a range of original assay values from 1.5% to 60% Sn were selected, this being the grade range in which the over-assaying could occur; the lower grade CRM assays being deemed to be accurate. The intention was to isolate the grade threshold at which a bias occurs and then re-assay all the pulps above that threshold. Included with the pulp rejects were ten CRM samples of the high-grade BCS-CRM number: 355 (31.4% Sn) and three of the lower grade OREAS 142 (1.04% Sn). ALS returned tin grades for BCS-CRM number: 355 mostly within tolerance with a slight high bias consistent with the post February assays of this CRM by ALS. The assays of OREAS 142 were also found to be accurate (Graph 12).

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2016 SGS assays of 2015 samples - SGS vs ALS Sn assays. SGS by titration



55

Graph 12: Control Charts for CRMs Assayed by ALS in 2016 for the Re-Assay Programme

Source: MSA

The 2016 assays compared reasonably well with the pre 2015 original assays, however the 2016 re-assays returned values on average 5% higher than the original (Graph 13).

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CRM Sn Value -2SD +2SD -3SD +3SD CRM Sn Assay

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Setpoint 2016 Standard OREAS 142 Sn for re-assays

CRM Sn Value -2SD +2SD -3SD +3SD CRM Sn Assay



56

Graph 13: Scatterplot for ALS 2016 Re-assays versus ALS Pre-2015 Assays

Source: MSA

12.1.6.1. Summary of the external Assay QA/QC Checks The field duplicate data indicate that the tin mineralisation is nuggety, which is expected in the high-grade vein style of mineralisation at Bisie. There is evidence that ALS may have over-estimated the high tin grade samples prior to 2015 as shown by the high-grade CRM and comparisons with one of the other check laboratories (Set Point). The re-assay program completed in 2016 demonstrated that the pre-2015 assays are of reasonable accuracy, the reason for the higher than expected high-grade CRM assays prior to 2015 remaining uncertain. The silver assays completed on the Bisie samples are of poor accuracy at grades of less than 1.5g/t, being biased lower than the CRMs accepted mean values, however given the low grade of silver in the Bisie deposit, the risk to the Mine is low. The copper assays are considered to be of good quality, as shown by the low failure rate for the CRMs and the good repeatability of the ALS assays by SGS. The zinc assays are accurate with a slight low bias for zinc indicated by the CRM analysis. The SGS zinc assays were lower than those of ALS. Overall the lead analyses were outside of acceptable limits both for the CRM assays and the second laboratory pulp duplicate assays; however, given the low grade of lead in the Bisie deposit the risk to the Mine is low. In summary, the quality of the assays is reasonable with the exception of the lead and low-grade silver assays which should be considered to be of low confidence. No batches of assays were failed on the basis of the QA/QC samples. In the QP’s opinion, the assay data can be used for Mineral Resource estimation. The pre-2015 assays can be used without modification as they have been demonstrated to be of reasonable accuracy by the re-assay program.

y = 1.0418x + 2423.2R² = 0.9803

0

100000

200000

300000

400000

500000

600000

700000

0 100000 200000 300000 400000 500000 600000 700000

Rea

ssay

Sn

(p

pm

)

Original Assay Sn (ppm)

ALS High-Grade Reassay



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12.2. Data Verification The Bisie site was visited by Mr J C Witley in July 2013, May 2014 and August 2015. The length of each site visit was three days. During the site visit a selection of cores were inspected, the positions of the drill sites verified and the exploration processes were reviewed. During the first site visit independent check sampling was carried out by the QP.

Independent Check Sampling As part of the data verification exercise, ten quarter core samples were taken from two drillholes (BGC018 and BGC024) across a variety of mineralisation intensities within the well mineralised zone. The samples were photographed on site and sealed and prepared for dispatch to ALS in Johannesburg. On arrival in Johannesburg, the samples were unpacked, photographed, verified against the photographs taken on site and then sent for assay. A comparison of the original assays with the independent check sampling assays is shown in Table 24 and the tin and copper results are shown as scattergrams in Graph 14.

Table 24: Check Sample Assay Results

Sample ID

Sn (%)

Cu (%)

Ag (g/t)

Pb (ppm)

Zn (%)

Original Check Original Check Original Check Original Check Original Check

BGC024_39 2.88 1.79 0.04 0.04 0.5 0.0 8 6 0.06 0.06

BGC024_40 2.01 6.78 0.07 0.08 0.0 0.0 5 15 0.06 0.06

BGC024_41 3.20 3.12 0.16 0.14 1.2 0.0 12 16 0.07 0.07

BGC024_42 1.80 1.95 0.14 0.14 1.3 0.0 15 13 0.08 0.08

BGC024_43 11.00 15.45 0.14 0.17 1.4 0.0 22 26 0.07 0.07

BGC018_45 1.04 1.11 0.59 0.58 5.0 5.1 35 38 0.63 0.73

BGC018_46 38.80 29.00 0.49 0.51 4.2 4.4 32 47 0.07 0.10

BGC018_47 0.35 0.35 0.59 0.61 4.2 4.1 22 31 0.10 0.10

BGC018_48 12.50 1.39 0.45 0.61 3.7 4.8 43 45 0.22 0.13

BGC018_49 1.68 4.49 0.79 0.71 6.4 5.7 45 43 0.09 0.08

Mean 7.53 6.54 0.35 0.36 2.8 2.4 24 28 0.14 0.15

Source: MSA

Graph 14: Scattergrams of Tin and Copper Check Sample Assay Results

Source: MSA

The check assays confirmed the presence of high-grade tin, although the individual sample assays did not compare well for the high-grade tin samples. Five of the samples were re-submitted with a new sample identity (ID) and re-assayed for tin Table 25.



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Table 25: Check Sample Assay Results – Re-assays

Sample ID

Sn (%)

Original Re-assay

BGC024_39 1.79 1.79

BGC024_40 6.78 6.14

BGC024_43 15.45 13.85

BGC018_46 29.00 29.50

BGC018_48 1.39 1.44

BGC018_49 4.49 4.37

Mean 9.82 9.52

Source: MSA

The results compared reasonably well with the original check sample results. The difference between the check sample and original sample is likely to be a result of the nuggety nature of the high-grade vein mineralisation, amplified by the small check sample size (quarter NQ core). No particular bias was noted for the tin assays with the high-grade check sample assays being either lower or higher than the original. No verification assays have been taken since July 2013, the presence of high-grade tin in cassiterite having been confirmed in 2013.

Visual Verification Tin mineralisation at Bisie is clearly visible in the cores, occurring as coarse grained cassiterite veins and finer disseminations. Most of the core observations confirmed the magnitude of the assayed grades within reasonable limits. A number of tin grades reported were significantly higher or lower than the observations in the remaining half core. This was considered a result of the irregular nuggety nature of the high-grade tin mineralisation. Cores were observed from a selection of drillholes, which represent a range of mineralisation intensities at Bisie (Table 26). The second site visit focussed on the short-range variability in the closely spaced metallurgical sample holes and the third site visit focussed on the extension drilling to the north at depth and the shallow area to the south. Comparisons were made between the observations and the assays received in order to verify the mineralisation.

Table 26: Drillhole Cores for which Mineralisation was Visually Verified

July 2013 May 2014 August 2015

BGC006 METBGC010 BGC077






BGH001 BGC044 BGC131

BGH006 BGC045 BGH014

BHH007A BGC046 BGH015

BGC047 BGH016

Source: MSA

Verification of Drillhole Collars Bisie was first visited in July 2013, which was aimed at initial Mineral Resource definition. Eleven of the drillhole collars were photographed and the locations verified against the surveyed coordinates by using a hand-held GPS. In May 2014, a number of the drillhole sites were visited including those from which the metallurgical samples were drilled. In August 2015, eleven of the drillhole collar positions were verified using hand-held GPS. All of the hand-held GPS measurements compared within reasonable limits to the final survey.

Summary of the Data Verification Independent check sampling by the QP confirmed the high tin grade samples characteristic of the Bisie mineralisation. The mineralisation shown by the tin assays was visually confirmed by the QP for a representative selection of drillhole cores. The locations of the drillholes were verified within reasonable limits by the checks performed. In the opinion of the QP, the data verification processes demonstrate that the database is adequate for the purpose of Mineral Resource estimation.



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13. MINERAL PROCESSING AND METALLURGICAL TESTING METHODS

13.1. Original Test Work A 7,643kg bulk metallurgical sample was constituted from half and three-quarter PQ size (85mm) core sourced from twenty-seven purpose drilled holes at Mpama North, and sent to Maelgwyn Mineral Services (MMS) in Johannesburg for metallurgical testing and piloting. Residual samples from the original phase of test work (completed by MMS) was made available for additional test work. No new samples were made available. The samples had been retained by MMS at their laboratory in Johannesburg. Sample Selection: Several samples made available to the previous phase of study, which included the following: Approximately 400kg naturally arising -1mm material Approximately 17kg -8mm +1mm DMS feed material Approximately 19kg -8mm +1mm DMS sinks material Approximately 13kg -8mm RoM material Review of Feasibility Study Metallurgical Test Work: Mineralogy and metallurgical test work were carried out in the last phase of work and was successful in identifying a potential processing route which could produce smeltable grade concentrate with acceptable levels of contaminants. Mineralogy: The test work identified Cassiterite as the only tin-hosting phase and that it was coarse grained. The majority of associated contaminants were identified as silicates (54%) of which 37% were chlorites. Additionally, iron-oxide/hydroxides, ilmenite and rutile were present with further minor associations with sulphides, carbonates and phosphates identified. It was concluded that gravity and flotation processes would be the likely processing route to be followed due to the liberation effects and the grain size observed. Metallurgical: Gravity pre-concentration of the -8mm +1mm stream was undertaken followed by grinding, spirals, grinding, sulphide flotation and oxide flotation to produce a high-grade concentrate. Additionally, liberation grind test work was undertaken to identify the preferred grind sizes for each stage of the process. The naturally arising -1mm material was concentrated through spirals, however further flotation and gravity test work on this stream was unsuccessful so only a low-grade spiral concentrate could be produced.

13.2. Front -end Engineering and Design (Feed) Test Work Economical assessments of the potential to sell the low-grade concentrate produced from the naturally arising -1mm stream were not favourable, so further test work was undertaken with Mintek with the aim of producing a single high-grade concentrate from the process, along with further potential optimisations. Mineralogy: Mineralogy test work confirmed previous results surrounding the main contaminants present in the process. It further expanded on the liberation characteristics of several key streams within the process and suggested potential routes for further optimisation. In summary, a stage wise size reduction followed by a gravity concentration step will produce high grade concentrates at each level of grind, with the majority of the liberated tin losses being incurred at each stage to the ultrafine fraction (-38µm). This suggests that that further work on recoveries in this fine size range would benefit the project in future optimisations. Metallurgical: The limited sample masses available for test work required that a lab scale, rather than a pilot or commercial scale test work be undertaken. Simplistically, two arms of test work were undertaken with the DMS sinks material (HG (high-grade)) and the naturally arising -1mm material (LG (low-grade)). The test work indicated that tin recoveries in excess of 80% at a grade of above 60% tin was possible through almost exclusively a gravity concentration process with flotation used for contamination removal only. The final processing flow sheet included: The split at 1mm was maintained, with the +1mm material considered the HG circuit and the -1mm material considered the LG circuit; The HG sample available was already a DMS sinks concentrate. This was milled prior to the shaking table test work, designed to simulate a spirals/shaking table arrangement, followed by further milling and sulphide rejection flotation to produce a >60% tin concentrate. All of the test work was carried out at Mintek and utilised a pilot scale shaking table and lab scale flotation cells;



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The LG sample was subjected to the shaking table test work, designed to simulate a spirals/ shaking table arrangement, followed by milling, a further stage of shaking tables and sulphide rejection flotation to produce a final concentrate of >60% Sn. All of the test work was carried out at Mintek and utilised a pilot scale shaking table and lab scale flotation cells; Sulphides removal by flotation was in excess of 90% effective, with the final quantities of contaminants in the concentrate within range to ensure no excessive smelter penalties. The flotation cells used were forced induction, conventional, lab scale units which are commonly scaled up and used within industry. Examples of the additional equipment used for the latest test work campaign are shown in Figure 21 and Figure 22 below.

Figure 21: MINTEK Pilot Scale Shaking Table

Source: DRA

Figure 22: MINTEK Bench Scale Flotation cell

Source: DRA



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14. MINERAL RESOURCE ESTIMATE MSA has completed a Mineral Resource estimate for the Mpama North prospect of the Bisie Project. To the best of the QP’s knowledge there are currently no title, legal, taxation, marketing, permitting, socio-economic or other relevant issues that may materially affect the Mineral Resource described in this Technical Report, aside from those already mentioned. The Mineral Resources presented herein, with an effective date of 30 June 2019, represent an update to the previous Mineral Resource estimate with an effective date of 09 May 2016. The Mineral Resource estimate incorporates drilling data from holes completed by Alphamin from July 2012 until November 2015 inclusive, which in the QP’s opinion were collected in accordance with CIM ‘Exploration Best Practices Guidelines’, 2000. Mining of the Mineral Resource has commenced at Bisie and this Mineral Resource estimate incorporates depletion by recent mining by ABM and a re-assessment of the extent of artisanal mining. The Mineral Resource was estimated using the 2003 CIM ‘Best Practice Guidelines for Estimation of Mineral Resources and Mineral Reserves’ and classified in accordance with the ‘2014 CIM Definition Standards’. It should be noted that Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. The Mineral Resource estimate was conducted using Datamine Studio 3 software, together with Microsoft Excel, JMP and Snowden Supervisor for data analysis. The Mineral Resource estimation was completed by Mr J Witley, the QP for the Mineral Resource.

14.1. Mineral Resource Estimation Database The database used for the Mineral Resource estimate consists of:

• information from diamond drillholes;

• collar surveys;

• down-the-hole-surveys;

• sampling and assay data;

• SG measurements;

• geology logs; and

• a Digital Terrain Model (DTM) based on a high-resolution LIDAR survey.

The principal sources of information used for the estimate include raw data generated during the course of the exploration drilling program conducted by Alphamin between July 2012 and November 2015 inclusive. The Mineral Resource estimate was based on tin, copper, lead, zinc and silver assays and density measurements obtained from the cores of 122 NQ size diamond drillholes. In addition to the exploration drillholes the split cores from 21 closely spaced PQ size holes were used in the estimate. These holes were drilled in three clusters for the purpose of obtaining a metallurgical test sample. The drillholes were angled downwards at between 60° and 75° to the west and planned to intersect the mineralised zones at a spacing of between approximately 25m and 100m on the plane of mineralisation. The cut-off date for inclusion of data into this estimate is 06 April 2016 at which time there were no outstanding drilling data of significance.

14.2. Exploratory Analysis of the Raw Data

The dataset examined consisted of sampling and logging data from Diamond Drillholes (DD). The following attributes are of direct relevance to the estimate:

• Tin (Sn), silver (Ag), copper (Cu), zinc (Zn), lead (Pb), sulphur (S) and arsenic (As) assays in parts per million (the tin, copper, zinc, lead and sulphur data were converted to per cent for the Mineral Resource estimation), SG measurements and Lithological Codes.

• The high-grade mineralisation occurs within a persistent zone of intense chloritisation termed amphibolite by the Alphamin geologists. Less continuous and narrower zones of mineralisation occur in places above and below the Main Vein zone, thus three zones were defined; the Main Vein, the Footwall Vein and the Hangingwall Vein zones.

• Visual inspection of the data showed that the well mineralised intersections occur within drillholes drilled between 25m and 50m apart along east-west fence lines spaced approximately 50m apart over a strike



62

length of approximately 600m, with some infill drilling at 25m line spacing. Two holes, BGC008 and BGC009, were drilled 150m and 280m to the north respectively, with both intersecting the prospective chlorite schist zone, however they were not significantly mineralised. The best intersection in BGC008 was 0.76% Sn over 0.4m and in BGC009 0.09% Sn over 1m. A number of intersections have been drilled to the south of the March 2015 Mineral Resource and significant mineralisation has been intersected. Examination of the drillhole data in section revealed that in the down dip areas a number of below threshold grade intersections occur that have constrained the down dip extent of most of the mineralised zone. However, the mineralisation clearly plunges to the north and is open down-plunge to the north.

Several of the drillholes, BGC013, 014, 015, 016 and 032 are affected by artisanal mining activity as evidenced by cavities and cored wooden mine supports. Each one of the affected holes is the uppermost hole in the respective fence line along which it was drilled. The data from these holes were used to inform the mineralisation model extents, but were not used for grade estimation as it was assumed that much of the high-grade tin mineralisation would have been removed by the artisanal miners and that the remaining mineralisation is not representative of the in-situ mineralisation. A summary of the drillhole data used for the Mineral Resource estimate is provided in Appendix 5.

Validation of the data The validation process consisted of:

• examining the sample assay, collar survey, down-hole survey and geology data to ensure that the data

were complete for all of the drillholes;

• examining the de-surveyed data in three dimensions to check for spatial errors;

• examination of the assay and density data in order to ascertain whether they were within expected

ranges; and

• checks for ‘From-To’ errors, to ensure that the sample data did not overlap one another or that there

were no unexplained gaps between samples.

The data validation exercise revealed the following:

• As at the effective date of this report data are available for BGC001 to BGC171;

• A number of the holes drilled were for geotechnical or sterilisation drilling to support other aspects of

the ongoing mining study (Section 16), many of them being outside of the Mineral Resource area;

• SGs were measured on pieces of the sample rather than for the entire sample and, given the variable

mineralisation at Bisie, will not always represent the density of the entire sample. As a result, individual

SG values measured in this way are considered unreliable due to the in-heterogeneous nature of the

tin mineralisation within the samples. SGs were later measured at the laboratory using a gas

pycnometer, however SGs were not determined for all samples in all holes;

• There are no unresolved errors relating to missing intervals and overlaps in the drillhole logging data;

• No default values, except for detection limit data, were found. Fifteen tin grades of 60%, 64 arsenic

grades of 10,000ppm and 141 sulphur values of 10% occur in the assay database, which is the upper

detection limit used by ALS for these elements;

• Examination of the drillhole data in three dimensions shows that the collars of the drillholes surveyed

by DGPS plot in their expected positions. However, five of the holes (BGC162 and BGC168 to

BGC171) had not yet been surveyed by DGPS and their elevations were derived from the LIDAR

survey topographic model;

• 2,698 laboratory SG measurements exist in the database for Mpama North. None of the values fell

outside of expected ranges for the rock types and mineralisation at Bisie; and

• Extreme assays were checked. The highest silver assay in the Bisie dataset is 775g/t followed by

113g/t and the next highest value is 109g/t. The 775g/t value was removed from the database for

estimation.



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Statistics of the Sample Data A total of 8,293 assayed samples occur in the Mpama North database of which 8,257 were assayed for tin.

Statistics of the Assay Data

14.2.3.1. Univariate Analysis Table 27 is a summary of the sample assay and SG data statistics in the raw data at Mpama North.

Table 27: Summary of the raw validated sample data at Mpama North.

Variable Number of Assays Mean Value Minimum Value Maximum Value

Sn (ppm) 8,257 15,337 <5 600,000*

Ag (g/t) 8,293 1.55 <0.5 113

Cu (ppm) 8,293 1,516 <1 103,500

Zn (ppm) 8,293 1,042 <2 121,000

Pb (ppm) 8,293 143 <2 51,500

S (%) 8,293 1.11 <0.01 10*

As (ppm) 8,293 265 <5 10,000*

Source: MSA *Note that the upper limit of detection for Sn value assayed was 60%, As 10,000 ppm and S 10%

The maximum assay value reported for tin is 600,000ppm, arsenic 10,000ppm and sulphur 10%. These are the maximum reported values for the assay methods used. Over-limit assays were not performed for arsenic and sulphur. There are sixty-four arsenic values, fifteen tin values and 141 sulphur values reported on the upper assay limit. The impact on the tin estimate will be negligible although a slight underestimation of arsenic and sulphur will occur. The tin sample data was examined in order to understand the general grade distribution and to determine thresholds that may be used to define mineralised envelopes. The data distribution is mixed containing several grade populations. A break in the log cumulative frequency plot was noted at approximately 0.35% Sn. Observation of the drillhole data in section revealed that approximately 0.35% Sn is also a practical grade threshold in which to constrain the mineralisation in a three-dimensional model. A distinct high-grade population was noted that should be considered in the grade estimation (Figure 23).

Figure 23: Log Cumulative Frequency Plot of the Sample Sn Grade Data

Source: MSA

14.2.3.2. Bivariate Analysis Scatterplots were made that compare the grades of each variable with one another in order to understand any relationships that may exist in the data that should be preserved in the Mineral Resource estimate. A linear relationship between tin grade and SG is observed with the grade of tin increasing with density. Very weak relationships are observed between copper and zinc and copper and sulphur.



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14.2.3.3. Relationship between tin grade and specific gravity It is expected that there will be a strong relationship between tin grade and SG as the tin bearing mineral at the Mine is cassiterite, which has a high SG (approximately 7.0). This relationship can be used to estimate the SG based on the tin grade. Not all samples were measured for SG and so the SGs were assigned to the mineralised samples without SG measurements using a polynomial regression on the average SG for a number of tin grade bins as shown in Figure 24. It is important to assign SG to samples as SG is used to composite the sample data to equal lengths using length and density weighting during the estimation process. For the un-mineralised zones, a constant value for each lithology, calculated using the mean value of that un-mineralised zone, was applied as shown in Table 28.

Table 28: Default Density Value Applied for each Rock Type in the Waste Zones at Bisie

Rock Type SG

Chlorite Schist (AMPH) 3.16

Chlorite Mica Schist (ASCH) 3.06

Core Loss (CL) 3.06

Mica Schist (MSCH) 2.98

Quartz Vein 2.78

Shear Zone 2.95

Fault Zone 3.06

Source: MSA

Figure 24: Regression Analysis for SG versus Sn Grade

Source: MSA

14.2.3.4. Comparison between Laboratory Gas Pycnometer and On-Site Wet-Dry Determinations The SG measurements were obtained by laboratory gas pycnometer. This differs from the generally accepted method of weight in air versus weight in water (the Archimedes principle) in that it does not take the porosity of the samples into account and therefore may over-estimate the density. SG was routinely determined on site using weight in air versus weight in water. These were made on 15cm lengths of core, rather that the entire sample, and, due to the heterogeneity of the cassiterite mineralisation, were not considered representative of the SG of the entire sample and so were not used in estimation. A comparison was made between the samples of drillholes that had both gas pycnometer and wet-dry SG measurements. A quantile-quantile (QQ) plot is shown in Figure 25, which demonstrates that the gas pycnometer method tends to over-estimate SG compared to the Archimedes method for values less than the mean value and under-estimates values for the denser samples. Overall the two methods compare well, with the average pycnometer SG being 3.38 and the average wet–dry SG being 3.39. Therefore, it is considered that use of the gas pycnometer SGs for density determination is a reasonable approach.



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Figure 25: QQ Plot of Gas Pycnometer SG versus Wet-dry Method

Source: MSA

Summary of the Exploratory Analysis of the Raw Dataset

• The database is robust;

• Most sample lengths are 1m or less;

• The host rock to the mineralisation is mainly chlorite schist (locally termed amphibolite);

• Parts of the uppermost portion of the deposit have been removed by artisanal miners and several of

the shallow drillholes are unrepresentative, having drilled through the workings;

• A threshold of 0.35% Sn is considered suitable to model zones of significant tin mineralisation;

• A high tin grade population occurs;

• There is a strong relationship between tin grade and SG, which was used to assign density to the

samples that were not measured for SG;

• Laboratory gas pycnometer SGs compare well with wet-dry determined SGs and use of the gas

pycnometer SGs for density determination is considered a reasonable approach; and

• The Mpama North mineralised zone is constrained by drilling down dip, although potential to extend

the mineralised zone exists down-plunge northwards and along strike to the south.

Topography Alphamin conducted a LIDAR survey in the second quarter of 2015 in order to provide for an accurate model of the topography. The processed data was provided as points on a 2m by 2m grid and a DTM was created from the point data (Figure 26). A LIDAR survey is considered one of the most accurate remote methods available to survey topography.



66

Figure 26: Isometric View of the DTM Created from the LIDAR Survey Data - View is Approximately to the North

Source: MSA

Mineralised Zones The drillhole data were examined for the occurrence of cassiterite mineralisation relative to the geological logging. Consistent with the October 2015 estimate, three zones of mineralisation were modelled (Figure 27):

• The Main Vein mineralisation consisting of a number of uncorrelated cassiterite veins within pervasively

chloritised schist (logged as amphibolite). This zone generally occurs over thicknesses of between 2m

and 22m with an average thickness of approximately 9m. This zone is generally the highest grade and

most consistent.

• Hangingwall Vein (HW Vein) mineralisation occurring within partly chloritised schist (logged as

amphibolite schist) and micaceous schist between 1m and 20m above the Main Vein. This zone of

mineralisation is generally between 0.5m and 4m wide and occurs in the northern area of the deposit

although it appears to taper out northwards. The middling between the Hanging Wall Vein and the Main

Vein decreases in areas and it is possible that this vein merges into the Main Vein in some parts of the

deposit.

• Footwall Vein (FW Vein) mineralisation, occurring within the micaceous schist and amphibolite schist

between 1m and 10m below the Main Vein. This zone is restricted to the southern areas, is very narrow

(<50cm) and high-grade in its most northern occurrences. Towards the south it thickens to

approximately 6m. It is possible that this vein merges into the Main Vein in some parts of the deposit.

All three zones have been modelled in a small area of the deposit, which coincides with an area of bonanza-style mineralisation where the metallurgical sample holes were drilled. It should be noted that the deposit is open down-plunge and opportunities to increase the size of the deposit exist. A northwest to southeast striking sub-vertical fault has been modelled in the south of the deposit. The interpreted fault has a down-throw of approximately 15m to the south. Several mineralised intersections have been drilled to the south of this fault; however, the structure of the deposit is uncertain in this area. A 20m strike by 85m dip slab of mineralisation (the S1 block) has been modelled on the basis of three drillhole intersections, however several mineralised intersections in the area remain unresolved. A sub-vertical fault was previously modelled in the north of the deposit striking from southwest to northeast. This was modelled on the basis of only two drillhole intersections and therefore the structure was uncertain. Two drillholes (BGC044 and BGC127) within the main area of mineralisation were not used when creating the October 2015 model. The positions of the mineralisation were not consistent with the surrounding drillholes and they were discarded from the estimate until the cause for the discrepancy was resolved. Inclusion of these holes and the new drilling data has led to a revised structural interpretation comprising two faults that strike northwards sub-parallel with the plunge of the Main Vein mineralisation. Although the exact nature of these faults is uncertain there is significant displacement (over 10m) associated with them. Low angle faulting sub-parallel to the mineralised unit has been described previously and it is possible that the displacements modelled relate to this type of structure.

Alphamin Resources Ltd.Bisie Tin Project, Mpama North DepositDTM created from LIDAR survey dataIsometric View looking approximately to the north

November 2015

J. Witley



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It should be noted that although the mineralisation has been modelled as steeply dipping slabs, the mineralisation within the individual zones occurs in the form of irregular high-grade veins of botryoidal cassiterite several tens of centimetres thick and lesser amounts in blebs and vein fragments irregularly disseminated in the schist. The brecciated nature of the tin veins may be an indication that the slabs are broken in places and additional faults that are not resolvable on the scale of the drilling may occur. Grade shells for the mineralised zones were created using a 0.35% Sn threshold. Points were created for the hangingwall and footwall positions of each grade zone and these were linked together in a series of strings that were used to create the grade shell wireframes (Figure 27). A ‘mined out zone’ was created that encloses the drillholes that have been identified as having intersected artisanal workings. This initially extended to 50m below surface and has since been extended to 75m below surface based on information gained during project development. It should be noted that although artisanal mining has taken place, the tin mineralisation was not completely removed and potential for remaining tin mineralisation exists within the ‘mined out zone’. Mining of the Mineral Resource has commenced at Bisie and this Mineral Resource estimate incorporates depletion by recent mining by ABM.

Figure 27: Isometric Views of the Geological Model, View to East (top) and West (bottom)

Source: MSA Note: the artisanal mined out zone is the area above the horizontal yellow dashed line. Main Vein Zone – light blue, S1 Block – dark blue, Hangingwall Vein Zone - red, Footwall Vein Zone - green, drillhole traces - white



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Oxidation/Weathering Surfaces No weathering or oxidation surfaces were modelled. Cassiterite is an oxide and is not affected by oxidation. Artisanal mining is assumed to have mined within and below the weathering zone and the majority of the remaining mineralised rock is assumed to be fresh.

14.3. Statistical Analysis of the Composite Data The data were composited to 1m lengths, de-clustered to a cell size of 8mX, 40mY and 40mZ and summary statistics were compiled for each mineralised zone (Table 29).

Table 29: Summary Statistics (de-clustered) of the Estimation 1m Composite Data

Variable Number of

Composites Minimum Maximum Mean CV Skewness

Main Vein (Main Block)

Sn (%) 1,481 0.01 60 4.35 1.94 3.4

Cu (%) 1,481 0.00 4.23 0.34 1.45 3.0

Zn (%) 1,481 0.01 3.31 0.16 1.36 5.5

Pb (%) 1,481 0.00 0.65 0.011 3.23 9.7

Ag (g/t) 1,481 0.00 45.67 2.83 1.58 3.5

As (ppm) 1,481 0 10,000* 490 2.73 4.5

S (%) 1,481 0.00 10.00* 1.28 1.28 2.6

SG 1,481 2.79 6.10 3.31 0.10 3.5

Hangingwall Vein

Sn (%) 133 0.00 23.40 1.91 1.80 3.7

Cu (%) 133 0.00 0.18 0.03 1.25 2.8

Zn (%) 133 0.00 3.36 0.12 3.14 6.7

Pb (%) 133 0.00 0.10 0.006 1.76 6.6

Ag (g/t) 133 0.00 14.39 0.41 3.62 8.4

As (ppm) 133 0 1316 47 2.50 7.5

S (%) 133 0 3.70 0.31 1.68 3.7

SG 133 2.80 3.85 3.21 0.04 1.6

Footwall Vein

Sn (%) 81 0.02 52.90 4.50 1.88 3.7

Cu (%) 81 0.00 2.51 0.24 2.14 3.7

Zn (%) 81 0.00 1.11 0.10 1.17 4.8

Pb (%) 81 0.00 0.40 0.016 3.11 5.7

Ag (g/t) 81 0.00 36.0 3.00 1.91 3.0

As (ppm) 81 0 8184 146 4.10 12.2

S (%) 81 0.05 6.65 1.23 1.06 2.8

SG 81 2.81 5.22 3.24 0.11 3.0

S1 Block (Main Vein)

Sn (%) 32 0.09 40.00 7.01 1.39 1.9

Cu (%) 32 0.00 1.47 0.33 1.19 1.7

Zn (%) 32 0.00 0.11 0.04 0.96 0.8

Pb (%) 32 0.00 0.01 0.002 0.79 2.6

Ag (g/t) 32 0.00 5.14 1.22 1.19 1.4

As (ppm) 32 0 8,992 1195 1.95 2.4

S (%) 32 0.23 10.00* 3.29 0.91 1.0

SG 32 2.98 4.50 3.34 0.10 1.7

Source: MSA



69

The statistical analysis revealed:

• Most of the data are in the Main Vein;

• There are few data for the S1 Block. It should be noted that this area is an extension of the Main Vein,

separated by interpreted faults;

• The S1 Block is high-grade; however, it is of small volume and is positioned in line with the main high-

grade trend of the deposit;

• High tin grades are found in the Footwall Vein where the mineralisation typically consists of a narrow

cassiterite vein;

• High-grades are common within the Main Vein Zone;

• The Hangingwall Vein is characterised by moderate tin and low base metal, arsenic and silver grades;

• Zinc and lead grades are low on average, although sporadic significant grades occur. Copper grades

are more significant;

• The histograms are strongly positively skewed, lessor so for SG;

• The coefficient of variation (CV) is moderate to high for tin grade being between approximately 1.4 and

1.9; but not extreme. High-grade cassiterite veins within a low-grade host gives rise to a bimodal

distribution of very high-grades within a larger low-grade population;

• The CVs for density are low;

• The CVs for copper, lead and zinc are moderate to high, with high CVs where a few particularly high

values occur together with many generally low values; and

• The generally high CVs and bimodal tin distribution indicates that linear estimation should be used with

caution.

A number of intersections are less than the chosen composite length of 1m, particularly in the Footwall Vein system where narrow (as little as 0.2m) high-grade intersections occur. These narrow intersections were retained for estimation. Estimation of tin grade directly will assign weights to the narrow intersections that are too high for the volume which they represent. Accumulations of tin grade times density times length were calculated so that the estimation takes into account both the weight and volume of the composite. Density*length was used to back-calculate the tin grade from the accumulations after interpolation. The summary statistics for the tin grade times density times length accumulations are shown in Table 30. The histograms and log probability plots for tin accumulation are shown in Figure 28 and the rest of the modelled attributes in Appendix 5. The distributions for the Hangingwall and Footwall Vein systems are less well defined that that of the Main Vein due to fewer data. The accumulation CVs are slightly higher than those for the grades, with the tin accumulation CV being approximately two or greater. The distributions for the S1 block of the Main Vein is based on few data so are not meaningful.

Table 30: Summary Statistics (de-clustered) of the Estimation 1m Composite Data for Tin Accumulation

Variable Number of


Main Vein (Main Block)

SNPCTDL 1,481 0.03 342.50 16.75 2.30 4.6

LENDEN 1,481 2.50 5.94 3.30 0.10 2.8

Hangingwall Vein

SNPCTDL 133 0 90.04 5.95 2.06 4.4

LENDEN 133 1.28 3.91 3.05 0.18 -2.0

Footwall Vein

SNPCTDL 81 0.06 146.54 13.31 1.96 3.7

LENDEN 81 0.98 4.62 2.92 0.24 -0.6


SNPCTDL 32 0.27 179.46 25.49 1.54 2.4

LENDEN 32 2.97 4.49 3.28 0.10 2.1

Source MSA Note: SNPCTDL is the product (accumulation) of Sn %, length (m) and SG. LENDEN is the product (accumulation) of length (m) and SG.



70

Figure 28: Histograms and Log Probability plots of the Geological Data

Source: MSA



71

The high-grade tin samples represent massive and semi-massive cassiterite veins. The Main Vein histogram exhibits bimodality with a distinct high-grade population. The high-grade population was separated from the rest of the data within the modelled zones using a threshold of 80 Sn%t/m as noted as a break in the log probability plot for the Main Vein. Approximately 7% of the Main Vein, 0.7% of the Hangingwall Vein and 6% of the Footwall Vein composites are greater than 80 Sn%t/m. Summary statistics of the separate grade populations for the high and lower grade tin accumulation data within the vein zones are shown in Table 31. Once separated into the two grade populations the CVs are considerably lower than the combined data.

Table 31: Summary Statistics (de-clustered) of the Estimation 1m Composite Data for Tin Accumulations Separated into High- and Low-grade Populations

Variable Number of


High-grade (>80 %t/m)

Main Vein 99 80.34 342.50 133.85 0.47 1.6

Hangingwall Vein 1 90.04 90.04 90.04 ND* ND*

Footwall Vein 5 117.38 146.54 139.98 0.08 -1.3

S1 Block 3 82.45 179.46 125.27 0.32 0.7

Lower Grade (<80 %t/m)

Main Vein 1,382 0.03 79.44 8.97 1.60 2.7

Hangingwall Vein 132 0.00 69.98 5.34 1.85 3.8

Footwall Vein 76 0.06 74.51 9.57 1.53 2.7

S1 Block 29 0.27 74.34 16.27 1.33 1.7

Source: MSA *ND = Not determined

Cutting and Capping The log probability plots and histograms of the composite data were examined for outlier values that have a low probability of re-occurrence. These values were capped to a threshold as shown in Table 32. Decisions on the capping threshold were guided by breaks in the cumulative log probability plots and the location of the high-grade samples with respect to other high-grade samples. Top cuts for tin accumulation were applied for a comparative estimate using ordinary kriging of the total tin data, but not when separately estimated using two data sets (>80%t/m and <80%t/m). The high-grade Tin values occur as a distinct population with a low coefficient of variation and no top-cuts were considered necessary as the high-grade data were estimated separately using a restricted search to limit the impact of the high-grade samples in areas away from them. The capping reduced the extreme CV’s but lead and arsenic CVs in the Main Vein Zone remained high (>2).



72

Table 32: Impact of Capping the Estimation Data

Attribute

Before Capping Cap Value After Capping

Number of Composites

Mean CV Number of

Composites Capped Mean CV

Main Vein

Sn % 1,481 4.35 1.94 37.1 31 4.17 1.78

Cu % 1,481 0.34 1.45 2.50 13 0.33 1.42

Zn % 1,481 0.16 1.36 1.01 16 0.15 1.07

Pb % 1,481 0.011 3.23 0.28 4 0.011 2.97

Ag g/t 1,481 2.83 1.58 23.3 10 2.77 1.49

As ppm 1,481 490 2.73 - 0 490 2.73

S % 1,481 1.28 1.28 - 0 1.28 1.28

Sn %t/m 1,481 16.75 2.30 150.00 33 15.26 1.83

Hangingwall Vein

Sn (%) 133 1.91 1.80 13.45 4 1.80 1.56

Cu (%) 133 0.03 1.25 - 0 0.03 1.25

Zn (%) 133 0.12 3.14 0.97 3 0.09 1.76

Pb (%) 133 0.006 1.76 0.027 3 0.006 0.89

Ag (g/t) 133 0.41 3.61 1.02 5 0.22 1.88

As (ppm) 133 47 2.50 580 2 44 1.64

S (%) 133 0.31 1.68 - 0 0.31 1.68

Sn (%t/m) 133 5.95 2.06 47.40 4 5.48 1.53

Footwall Vein

Sn (%) 81 4.50 1.88 18.55 6 3.75 1.41

Cu (%) 81 0.24 2.14 0.69 3 0.15 1.20

Zn (%) 81 0.10 1.17 0.35 1 0.10 0.89

Pb (%) 81 0.016 3.11 0.068 5 0.009 1.53

Ag (g/t) 81 3.00 1.91 19.85 2 2.86 1.78

As (ppm) 81 146 4.10 1130 3 107 1.76

S (%) 81 1.23 1.06 3.87 2 1.13 0.83

Sn (%t/m) 81 13.31 1.96 74.50 6 11.48 1.59


Sn (%) 32 7.01 1.39 - 0 7.03 1.39

Cu (%) 32 0.33 1.19 0.93 2 0.30 1.19

Zn (%) 32 0.04 0.96 - 0 0.04 0.96

Pb (%) 32 0.002 0.79 0.004 3 0.002 0.59

Ag (g/t) 32 1.22 1.19 - 0 1.22 1.19

As (ppm) 32 1,195 1.95 - 0 1,195 1.95

S (%) 32 3.29 0.91 - 0 3.29 0.91

Sn (%t/m) 32 25.49 1.54 82.40 3 21.69 1.54

Source: MSA

14.4. Geostatistical Analysis

Semi-variograms The 1m composite data were examined using semi-variograms that were calculated and modelled using Snowden Supervisor software. All attributes were transformed to normal scores distributions and then the spherical semi-variogram models were back-transformed to normal statistical space for use in the grade interpolation process. Semi-variograms were calculated on the 1m composite data and modelled within the plane of mineralisation with a plunge to the north, this being the major direction of continuity, and the minor direction being across strike. Rotations were aligned for all the attributes estimated. Normalised semi-variograms were calculated so that the sum of the variance (total sill value) is equal to one. Semi-variograms were modelled with either one or two spherical structures. The nugget effect was estimated by extrapolation of the first two experimental semi-variogram points (calculated at the same lag as the composite length) to the Y axis.



73

There were insufficient data to calculate robust semi-variograms for the Hangingwall and Footwall Vein zones and the South block, and so the semi-variograms for the Main Vein zone were applied to these zones. The orientation of the South Block is slightly different to the Main Vein and the rotation angles were adjusted slightly to cater for the orientation. Most variables show strong continuity in the down-plunge direction in excess of the drillhole spacing. The tin grade semi-variogram model exhibits a range of 72m along plunge and 63m perpendicular to the plunge direction. The across strike variogram exhibited a range of 5m. It should be noted that the total data set tin accumulation semi-variograms and tin grade semi-variograms were only used in the check estimates to validate the indicator-based estimate. The semi-variogram model parameters are shown in Table 33, the tin grade semi-variograms in Graph 15 and variograms for all attributes estimated are presented in Appendix 7. The reliability of the semi-variograms both in the plane of the mineralisation and across strike for each of the variables is generally moderate to good.

Graph 15: Semi-variograms for Tin Grade (Main Vein)

Source: MSA



74

Table 33: Semi-variogram Parameters – Main Vein System, Mpama North

Attribute Transform Rotation Angle Rotation Axis Nugget

Effect (C0)

Range of First Structure (R1) Sill 1 (C1) Sill 1

(C1)

Range of Second Structure (R2) Sill 2

(C2) 1 2 3 1 2 3 1 2 3 1 2 3

Grades and Density

Sn % NS 80 62 -135 Z X Z 0.42 72 63 5 0.58

Ag g/t NS 80 62 -135 Z X Z 0.09 50 50 4 0.60 125 85 12 0.31

Cu % NS 80 62 -135 Z X Z 0.03 200 85 11 0.97

Zn % NS 80 62 -135 Z X Z 0.07 50 65 4 0.63 110 65 11 0.30

Pb % NS 80 62 -135 Z X Z 0.34 120 70 9 0.66

S % NS 80 62 -135 Z X Z 0.16 45 65 3 0.37 120 80 12 0.47

As ppm NS 80 62 -135 Z X Z 0.17 30 15 6.5 0.54 170 50 6.5 0.29

Density NS 80 62 -135 Z X Z 0.26 20 20 3 0.35 140 55 17 0.39

Accumulations

Sn %.t/m NS 80 62 -135 Z X Z 0.44 75 64 5 0.56

Length*Density NS 80 62 -135 Z X Z 0.19 100 50 3 0.24 100 50 25 0.57

Source: MSA Note: rotation angle for S1 Block adjusted to 90, 55, 0135 to cater for the different orientation to the Main Vein.



75

Indicator Semi-variograms At Mpama North, extreme high-grades associated with vein cassiterite occur together with lower grade disseminated cassiterite. A threshold of approximately 80 Sn%t/m separates the two distributions. The continuity of the higher-grade mineralisation in the Main Vein zone was modelled using indicator semi-variograms, whereby the data was transformed to a value of zero if less than 80 Sn%t/m or one if greater than 80 Sn%t/m, and semi-variograms were calculated using the indicator values. The indicator model parameters are shown in Table 34 and the semi-variograms in Figure 29. The semi-variogram was modelled with a range of 60m in the plunge direction, 42m perpendicular to plunge and 5m across strike.

Figure 29: Indicator Semi-variograms for 80 Sn%t/m (Main Vein)

Source: MSA

Table 34: Indicator Semi-variogram Parameters – Main Vein System, Mpama North

Attribute Rotation Angle Rotation Axis

Nugget Effect (C0)

Range of First Structure

(R1)

Sill 1 (C1)

Range of Second Structure

(R2) Sill 2 (C2)

1 2 3 1 2 3 1 2 3 1 2 3

Indicator 80 Sn (%t/m)

80 62 -135 Z X Z 0.34 60 42 3.5 0.66 - - - -


Above and Below Threshold Semi-variograms The continuity of tin mineralisation was modelled separately for the high-grade subset (>80 Sn%t/m) and the low-grade subset (<80 Sn%t/m). Strong down-plunge continuity was noted for the below threshold population, however omni-directional semi-variograms in the plane of mineralisation were modelled for the above threshold population, as robust semi-variograms could not be modelled in all three directions. The semi-variograms are shown in Graph 16 and Graph 17 and the semi-variogram parameters in Table 35.

Table 35: Above and Below Threshold Semi-variogram Parameters – Main Vein System, Mpama North

Attribute Rotation Angle Rotation Axis

Nugget Effect (C0)

Range of First Structure

(R1)

Sill 1 (C1)

Range of Second Structure

(R2) Sill 2 (C2)

1 2 3 1 2 3 1 2 3 1 2 3

>80 Sn (%t/m)

80 62 -135 Z X Z 0.60 73 73 3 0.40 - - - -

<80 Sn (%t/m)

80 62 -135 Z X Z 0.40 70 35 3 0.22 70 63 5.5 0.38




76

Graph 16: Threshold Semi-variograms for Main Vein Zone <80 Sn%t/m

Source: MSA



77

Graph 17:Threshold Semi-variograms for Main Vein Zone >80 Sn%t/m

Source: MSA

14.5. Block Modelling Block models were rotated in the dip direction in order to best fit the orientation of the mineralised zones. The block model prototype parameters are shown in Table 36. The cells were split to a minimum sub-cell of 0.20mX by 2mY by 2mZ in order to fill the wireframe model boundaries accurately and create blocks where the mineralisation is narrow.

Table 36: Block Model Prototype Parameters for Bisie, Mpama North (m)

Block Size Model Origin Rotation Angle Rotation Axis Number of Cells

X Y Z X Y Z 1 2 3 1 2 3 X Y Z

2 20 10 582,500 9,885,000 -150 0 28 0 Z Y Z 642 108 112

Source: MSA

Block models were created by filling below the topographic surface and above and below the ‘mined-out’ area surface using the same model prototypes as shown in Table 36. The topographic model and ‘mined-out’ models were added to the mineralisation models, so that the block model cells were coded as either mined or un-mined, and the model cells above the topographic surface were removed. A waste model was made at least 20m either side of the vein models to assist with mine planning.



78

A dynamic anisotropy process in suitable software was used to control the estimate so that the search parameter directions were modified to follow the local shape of the mineralised zones. This was achieved by estimating a dip and direction into each block model cell based on the dip and direction of the wireframes. The search parameters used to estimate the dynamic angles into the block model are shown in Table 37.

Table 37: Dynamic Anisotropy Search Parameters for Bisie, Mpama North

Search Distance (m)

Search Angle Search Axis Number of Samples

Search Multiplier

1

Search Multiplier

2

1 2 3 1 2 3 1 2 3 Minimum Maximum

50 25 7 80 62 0 Z X Z 3 5 6 -

Source: MSA

Validation of the Block Model Volumes with the Wireframe Volumes The filling of the wireframes by the block model cells was validated by comparing the volume of the rotated block model with the volume of the wireframe (Table 38). The volumes compare well.

Table 38: Validation of Block Model Filling

Zone Wireframe Volume Block Model Volume Difference (%)

Main Vein 1,597,157 1,582,542 -0.9

Hangingwall Vein 30,497 30,477 -0.1

Footwall Vein 46,725 46,558 -0.4

South 1 Block 21,052 21,476 2.0

Source: MSA

14.6. Estimation Attributes were estimated into the individual mineralised zones using the capped 1m composite drillhole sample data for each zone. Ordinary kriging was used to estimate the attributes into the block model cells using parent cell estimation. The search distance and the rotation angles that defined the search ellipses were based on the semi-variogram model for each attribute. The minimum number of composites required for a high confidence estimate was eight and the maximum number was twenty-four. The minimum number of samples required for an estimate was reduced to four for the narrower Hangingwall and Footwall Vein zones and six for the S1 block. If an estimate was not achieved within the search ellipse volume, the search ellipse was expanded by 50% and again by a factor of ten. Should an estimate still not be achieved, a larger search ellipse was used sourcing up to thirty-six composites to estimate grades close to the local average when away from the drillhole data. This was only necessary for an insignificant number of cells. Discretisation was set at five for strike by four on dip by two in the across strike direction of the blocks. Dynamic anisotropy was used to guide the search in the local direction of the mineralised zones. The ordinary kriging search parameters are shown in Table 39. Ordinary kriging was also used for the waste model using the same estimation parameters as for the Main Vein zone.

Estimation of tin accumulation A different approach was used for the tin accumulation whereby the probability of a block having an accumulation value above the 80 Sn%t/m threshold was estimated for each block in the model using indicator kriging. In order to restrict the influence of the high-grade values away from the high-grade intersections the indicator estimate was restricted to a single short search distance. This distance was defined by projection of the linear part of the indication semi-variogram to the sill value rather than the range of the sill from the spherical model, as illustrated in Figure 30. Any cell that was not estimated with an indicator value in the first search was assigned a zero-indicator value (probability of above threshold being nil) and therefore samples from the high-grade population were restricted from estimating beyond a relatively short distance away from their locations.



79

Figure 30: Illustration of the Search Distance Derived from the Linear Portion of the Indicator Semi-variogram Model

Source: MSA

Each cell was then estimated with the tin accumulation values above 80 Sn%t/m and again with tin accumulation values below 80 Sn%t/m. The estimated accumulation value using the high-grade data was multiplied by the probability of the cell being above the threshold and the estimated accumulation value using the lower grade data was multiplied by the probability of the cell being below the threshold. The two values added together result in an estimated accumulation value for each cell in the block model. The estimation parameters for length times density were aligned with the tin accumulation so that both estimates use the same samples to ensure that the transformation back to the grade values is correct. The search parameters for the indicator estimation are shown in Table 40.

45 m 33 m



80

Table 39: Search Parameters for Mpama North – Ordinary Kriging

Attribute

Search Distance Search Angle Rotation Axis First Search Volume

Factor

Search Multiplier 2

Factor

Search Multiplier 3

1 2 3 1 2 3 1 2 3 Minimum Number

Maximum Number

Minimum Number*

Maximum Number

Minimum Number*

Maximum Number

Grades and Density

Sn (%) 72 63 5 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Ag (g/t) 125 85 12 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Cu (%) 200 85 11 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Zn (%) 110 65 11 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Pb (%) 120 70 9 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

S (%) 120 80 12 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

As (ppm) 170 50 6.5 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Density 140 55 17 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Accumulations

Sn (%t/m) 75 64 5 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Length Times Density 75 64 5 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Source: MSA Note: The minimum number of composites used for the Hangingwall and Footwall vein zones is 4 and for the S1 Block is 6. Rotation angle for S1 Block adjusted to 90, 55, 0135 to cater for the different orientation to the Main Vein.

Table 40:Search parameters for Mpama North – indicator approach

Attribute

Search Distance Search Angle Rotation Axis First Search Volume

Factor

Search Multiplier 2

Factor

Search Multiplier 3

1 2 3 1 2 3 1 2 3 Minimum Number

Maximum Number

Minimum Number*

Maximum Number

Minimum Number*

Maximum Number

Indicator 80 (Sn %t/m)

45 33 3.5 80 62 -135 Z X Z 4 10 - - - - - -

Sn %.t/m > 80 73 73 3 80 62 -135 Z X Z 4 10 1.5 4 10 10 4 10

Sn %.t/m < 80 70 35 5 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

L times SG (Sn%.t/m > 80)

73 73 3 80 62 -135 Z X Z 4 10 1.5 4 10 10 4 10

L times SG (Sn%.t/m < 80)

70 35 5 80 62 -135 Z X Z 8 24 1.5 8 24 10 8 24

Source: MSA Note: The minimum number of composites used for the Hangingwall and Footwall vein zones is 4 and for the S1 Block is 6 for the <80 Sn%t/m population. Rotation angle for S1 Block adjusted to 90, 55, 0135 to cater for the different orientation to the Main Vein.

81

The estimates of tin grade times density times length were divided by the estimate of density times length for each block in order to obtain the estimated grade for tin. As well as the indicator approach, tin accumulation was estimated using ordinary kriging and the two results were compared in order to ensure the indicator approach is robust.

14.7. Validation of the Estimates The models were validated by:

• visual examination of the input data against the block model estimates;

• sectional validation;

• comparison of the input data statistics against the model statistics; and

• comparison of the indicator model with the ordinary kriged model.

The block model was examined visually in sections to ensure that the drillhole grades were locally well represented by the model. The model validated reasonably well against the data, although there is a minor degree of smoothing, the high and low-grade areas being well represented by the model. Examples of sections showing the block model and drillholes shaded by Sn% are shown in Figure 31. Note that the section on 9,885,705mY illustrates both the Main Vein and the Footwall Vein zone and that the waste model surrounding the vein systems is included in both illustrations.

Figure 31: Sections through Block Model and Drillhole Data illustrating Correlation between Model and Data

Source: MSA Note: The sections are 255m apart There have been no material changes since the May 2016 geological Model

Sectional validation plots were constructed for tin grade in order to compare the average grades of the block model against the input data along a number of corridors in various directions through the deposit. Samples of the sectional validation plots for tin grade (model grades back calculated from the accumulations) are shown in Figure 32. These show that the estimates are smoother than the data, yet retain the broad grade trends across the deposit.

Alphamin Resources Corp.Bisie Tin Project, Mpama North DepositDip Section Sn block model and drillholes 9,885,960 mY Looking North

May 2016

J. Witley

125 m

Alphamin Resources Corp.Bisie Tin Project, Mpama North DepositDip Section Sn block model and drillholes 9,885,705 mY Looking North

May 2016

J. Witley

125 m

82

Figure 32: Sectional Validation Plots for Sn% (model grades back calculated from the accumulations)

Source: MSA Note: Rotated coordinates shown, not true coordinates.

As a further check, the length-density weighted grades of the drillholes were compared with the model grade (Table 41) it being expected that the model grade will be slightly lower than the data grade due to the skewed data distribution. The model and the data averages compare well for most areas and attributes, the comparison being poorer in the zones with less data that are sensitive to the arrangement of high and low grades. There are few samples in the South 1 Block, which covers a small area, and the estimate is sensitive to their location, hence the large difference between the input data and model grade, the indicator approach applied having restricted the impact of the particularly high tin grades. The higher arsenic value in the S1 Block model compared to the data is due to the highest arsenic values being in the central location of the three drillholes that has the most influence in the model. The large difference between the model arsenic and silver grades and the data for the Footwall Vein is due to the particularly high arsenic and silver grades being clustered towards the north edge of the model area that is well drilled. The de-clustering applied has not adequately catered for this and larger areas of the model are estimated with sparser lower grade data.

0

20

40

60

80

100

120

140

160

180

200

0

2

4

6

8

10

12

14

Nu

mb

er

of

Sam

ple

Co

mp

osi

tes

Gra

de

(%

)

Rotated Elevation (amsl)

Sectional Validation of Main Vein Sn Grade - by Elevation

Number of Samples Sample Grade Model Grade

0

20

40

60

80

100

120

140

160

180

200

0

2

4

6

8

10

12

14

Nu

mb

er o

f Sa

mp

le C

om

po

site

s

Gra

de

(%)

Rotated Y coord (m)

Sectional Validation of Main Vein Sn Grade - by Northing

Number of Samples Sample Grade Model Grade

83

Table 41: Comparison between Drillhole and Model Data Values

Variable Mean Model Mean Full Intersection

Composite (length density weighted)

Mean 1m Composite

Data

Mean 1m Composite Data

(Capped)

Main Vein

Sn % 4.32 (4.26) 5.11 4.35 4.17

Ag g/t 2.54 2.67 2.83 2.77

Cu % 0.31 0.32 0.34 0.33

Zn % 0.15 0.15 0.16 0.15

Pb % 0.010 0.011 0.011 0.011

S % 1.24 1.34 1.28 1.28

As ppm 443 399 490 490

Density 3.30 3.31 3.31 3.31

Hangingwall Vein

Sn % 1.96 (2.18) 1.72 1.91 1.80

Ag g/t 0.20 0.27 0.41 0.22

Cu % 0.03 0.04 0.03 0.03

Zn % 0.08 0.16 0.12 0.09

Pb % 0.005 0.006 0.006 0.006

S % 0.28 0.43 0.31 0.31

As ppm 46 50 47 44

Density 3.21 3.21 3.21 3.21

Footwall Vein

Sn % 4.28 (3.73) 3.93 4.50 3.75

Ag g/t 1.72 3.38 2.97 2.86

Cu % 0.12 0.24 0.24 0.15

Zn % 0.08 0.10 0.10 0.10

Pb % 0.008 0.020 0.016 0.009

S % 1.10 1.19 1.23 1.13

As ppm 82 159 146 107

Density 3.18 3.18 3.24 3.24

S1 Block

Sn % 5.48 (7.03) 8.18 7.01 6.73

Ag g/t 1.12 1.32 1.22 1.22

Cu % 0.27 0.34 0.33 0.30

Zn % 0.03 0.04 0.04 0.04

Pb % 0.002 0.002 0.002 0.002

S % 3.27 3.09 3.29 3.29

As ppm 1,405 1,158 1,195 1,195

Density 3.35 3.35 3.34 3.34

Source: MSA

14.8. Mineral Resource Estimate The Mineral Resource is classified into the Measured, Indicated and Inferred categories as shown in Table 42. The Mineral Resource is reported at a base case tin grade of 0.50%, which MSA considers will satisfy reasonable prospects for economic extraction given the high in-situ mineralised value.

Table 42: Bisie Mpama North Zone Mineral Resource at 0.50% Sn Cut-Off Grade, 30 June 2019

Classification Quantity

(Mt) Quality

(%) Content

(kt) Cu (%)

Pb (%)

Zn (%)

Ag (%)

Measured Mineral Resource 0.33 4.75 15.6 0.22 0.12 0.006 1.4

Indicated Mineral Resource 3.99 4.59 183.4 0.32 0.16 0.010 2.8

Measured and Indicated Resources 4.32 4.61 199.0 0.31 0.15 0.010 2.7

Inferred Mineral Resource 0.48 4.57 21.8 0.16 0.09 0.013 1.4

Source: MSA Notes:

• All tabulated data has been rounded and as a result minor computational errors may occur.

• Mineral Resources which are not Mineral Reserves have no demonstrated economic viability.

• Mineral Resources are reported inclusive of Mineral Reserves.

• Alphamin has an 80.75 percent interest in ABM. The Government of the Democratic Republic of Congo (GDRC) has a non-dilutive, 5% share in ABM. The Gross Mineral Resource for the Mine is reported.

• Depleted by mining from mine surveys as at 30 June 2019 and an estimate of the extent of artisanal mining to 725 mamsl.

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15. MINERAL RESERVE ESTIMATE Key assumptions, parameters and methods used to convert Mineral Resources to Mineral Reserves.

15.1. Cut-off grade A cut-off grade was calculated based on an assumed metal price (USD18,000/t Sn) and operating costs (USD172/t milled) with process recovery assumption (72%) and other factors, taking account of mining losses and geology. The calculation determined a cut-off grade of some 1.6% Sn, which was used to limit the mining schedule to areas of the orebody where the in-situ grade exceeded 1.6% Sn. Table 43 presents the cut-off grade calculation. The cut-off grade calculation was completed at the start of the study to identify and eliminate low grade (probably non payable) areas.

Table 43: Cut-off Grade Assumptions

Item Unit Value

Tin Price USD/tonne 18,000

Operating Costs Processing USD/t RoM 8

Mining USD/t RoM 25

Labour USD/t RoM 61

MRM USD/t RoM 6

Overheads (G&A) USD/t RoM 30

TSF USD/t RoM 6

Diesel (Power) USD/t RoM 20

Site Infrastructure USD/t RoM 14

Transport and Duties USD/t RoM 3

Total Operating Cost USD/t RoM 172

Geology Factor % 87

Mining Loss % 5

Plant Recovery % 72

Mill Head Grade % Sn 1.38

Breakeven In-situ Grade % Sn 1.6

Source: Sound Mining Note: The breakeven grade calculation was performed early in the development of the revised LoM plan and prior to completion of the latest business plan and associated budget forecasts. Assumptions used in the calculation may have changed in the interim.

15.2. Mine Design Criteria and Modifying Factors Mining losses and dilution: These are two components of modifying factors which are presented in Table 44. The mine design included pillars left in-situ to ensure stable and safe mining conditions underground. The mining schedule production results were compared to the Mineral Resource statement to check quantities mined, grade and content. The mine scheduling further accounted for planned (5%) and unplanned dilution (15.7%), recovering some sill pillars, and previously mined out areas. Metal Price: The prevailing metal price over time will fluctuate according to supply and demand fundamentals. Any movement in price will affect the revenue stream directly and may result in a material improvement or deterioration, as the case may be. The operation will have no control over the metal price but may be able to implement contingency plans by mining in different areas with different grades or different quantities. Metallurgy: Metallurgical recoveries influence the revenue and profitability directly and therefore any improvement or deterioration in metallurgical recovery greater than 10% will have a material effect on profitability. The Mine’s business plan assumes a metal recovery of 72%.

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Table 44: Mining Losses and Dilution

Activity Mining Loss

(%)

Planned Dilution

(%)

Unplanned Dilution

(%)

On Ore Development 5% 6% 0%

Slot Cuts 5% 5% 15.7%

Sliping 5% 6% 0%

Stoping 5% 5% 15.7%

Sill Pillar Recovery 50% 0% 0%

Source: Sound Mining

This table also indicates some recovery of sill pillars later in the life of mine plan and is added to the mineable inventory. Dilution is included in the mineable inventory at zero grade for sill pillar recovery.

15.3. Disclosures The terms ‘Mineral Resources’ and ‘Mineral Reserves’ have the meanings as ascribed by the Canadian Institute of Mining, Metallurgy and Petroleum, as the CIM Definition Standards on Mineral Resources and Mineral Reserves adopted by CIM council, as amended. The Mineral Reserve was estimated for a LoM design to support appropriate sequencing and scheduling. Cognisance was taken of hydrogeology, geotechnical criteria and various other modifying factors to ensure an acceptable level of accuracy. Appendix 5 contains the mine production schedule that supports this Mineral Reserve Statement. Only Measured and Indicated Mineral Resources are converted with 13.7% of the scheduled production being from Inferred Mineral Resources. The Mineral Reserve estimate is stated in Table 45 as at 31 December 2019. It follows an assessment of the economic viability (Section 22) of the Mineral Resources that were scheduled for depletion before confirming them as Mineral Reserves. The forecasts in the associated economic model include provision for remedial or mitigation measures to address risks discussed herein.

Table 45: Bisie Mpama North Zone Mineral Reserve at USD17,000/t Sn, at 1.6% Sn cut-off 31 December 2019

Classification

Quantity (Mt)

Quality (%)

Content (kt)

2019 2017 2019 2017 2019 2017

Proven Mineral Reserve 0.05 0.38 3.77 4.17 1.89 15.9

Probable Mineral Resource 3.28 4.29 4.01 3.53 131.49 151.4

Total Mineral Reserves 3.33 4.67 4.01 3.58 133.38 167.3

Source: Sound Mining Notes:

• The Mineral Reserve has been reported in accordance with the requirements and guidelines of NI43-101 and are 100% attributable to ABM.

• Apparent computational errors due to rounding and are not considered significant.

• The Mineral Reserves are reported with appropriate modifying factors of dilution and recovery.

• The Mineral Reserves are reported at the head grade and at delivery to Plant.

• The Mineral Reserves are stated at a price of USD17,000/t Sn as at 31 December 2019.

• Although stated separately, the Mineral Resources are inclusive of the Mineral Reserves.

• No Inferred Mineral Resources have been included in the Mineral Reserve estimate.

• Quantities are reported in metric tonnes.

• The input studies are to the prescribed level of accuracy.

• The scheduled production includes 13.7% Inferred Mineral Resource, with most of these towards the tail end of the production forecast.

• The Mineral Reserve estimates contained herein may be subject to legal, political, environmental or other risks that could materially affect the potential exploitation of such Mineral Reserves.

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16. MINING METHODS

16.1. Geological Considerations The ore body dips at approximately 60° to 65° to the east and strikes close to north-south. The mineralised zone plunges approximately 25° towards the north. The Mineral Resource of Mpama North was estimated for the portion of the deposit that is deeper than approximately 75mbs to exclude shallow areas that have been depleted by artisanal mining. It extends approximately 700m in the down plunge direction and to a depth of approximately 550mbs as dictated by the availability of the exploration drilling data. The mineralisation is open in a northerly plunging direction and for a limited extent to the south. The strike of the payable zone ranges from 490m in the shallow areas to 150m in the deeper areas of the mine. The deposit remains open at depth. The Main Vein zone of the deposit, which accounts for approximately 97.5% of the Mineral Resource (by tin content), is on average approximately 9m thick. It narrows (~2m) at the margins and can be up to 20m thick in the central, and generally higher grade, area. The zones that occur several metres above and below the main zone are generally considerably narrower than the Main Vein zone and cover areas of between 100m and 200m in the dip and strike directions. Borehole logging showed that there are three main rock types that need to be considered for mine planning, namely, chloritised schist, amphibolite schist and mica schist. Witley and Heins (2014) in their competent persons report (CPR) described the host rock and mineralised zones as follows: ‘The tin mineralisation found so far at the Bisie Project is confined to the north-south striking, easterly dipping metasediments that occur approximately 3 km east of a granite contact. Structural and mineralogical evidence suggests that the cassiterite was emplaced first, followed by copper in the form of chalcopyrite and bornite, then by lead and zinc mineralisation. Chlorite alteration is extensive in parts and is thought to be the result of late stage fluids entering the system. The tin and copper mineralisation are predominantly found in zones dominated by intense chloritic alteration, although mineralised zones with no chlorite have also been intersected by drill holes. The host rocks are predominantly highly chlorite-altered amphibolites, fine- to medium-grained chlorite schists and to a lesser extent, the adjacent biotite schists and quartz schists.’ Jackson (2015) in his structural analysis described the deposit as follows: ‘The Mpama North orebody comprises a sheeted set of cassiterite veins, 2mm to 1.8m thick, over a true width of 5m to 15m which form a tabular body striking north-south for 570m and dipping 50° to 65° to the east. The veins are hosted within an intensely chlorite altered amphibolite schist, approximately the same width as the orebody. The amphibolite is in turn hosted within a mica schist unit. A 2m to 3m wide chlorite schist halo commonly occurs along the contact between the amphibolite and mica schist.’ The conclusions drawn from the geological interpretation is that the deposit is largely altered and contains high quantities of chlorite.

16.2. Geotechnical Considerations The QP has reviewed the following geotechnical aspects:

• Geotechnical data acquisition which involves the logging of core, the selection of suitable samples, test

work and the interpretation of the raw data into geotechnical design criteria; and

• Geotechnical design with respect to the maximum spans for all excavations, the location of footwall

infrastructure, and the support requirements.

Geotechnical Data Acquisition The geotechnical data acquisition program commenced at a concept level of accuracy in May 2014, with the logging of un-orientated core from five boreholes. The database was later supplemented with information collected from seven bore holes drilled in November/December 2014. Six more holes were logged in April/May 2015 for the portal design. This information facilitated a comprehensive assessment of the geotechnical character of the orebody and surrounding rock mass.

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Rock Mass Quality The overall rock mass quality across the Main area is classified as “good” based on the RMR89 and Q Index systems. Note: the classification as good represents rock mass quality before adjustments. Graph 18, Graph 19 and Table 46 describe the rock mass quality.

Graph 18: Rock Mass Rating (RMR 89) for All Rock Types

Source: Bara Consulting

Graph 19: Q-Index for All Rock Types


Table 46: Summary of Rock Mass Quality Data

Rock Type RMR89 Q

Average Interpretation Average Interpretation

Chloritised Schist 65.88 Good 11.39 Good

Amphibolite Schist 73.79 Good 13.59 Good

Mica Schist 70.40 Good 10.72 Good


Further detail on spatial distribution across the area and variances with depth can be referred to in the context of the report.

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Joint Orientations The stereonet analysis using the DIPS software package indicated the presence of three dominant joint sets each per rock type and numerous randomly orientated joints that could not be classified as a set. The orientations of the dominant joint sets are summarised in Table 47.

Table 47: Intact Rock Strength Testing

Rock Type Dip

(°)

Azimuth

(°) Joint Set

Chloritised Schist

63 106 JS 1

63 84 JS 2

60 54 JS 3

Amphibolite Schist

51 39 JS 1

61 90 JS 2

33 193 JS 3

Mica Schist

60 52 JS 1

64 79 JS 2

62 104 JS 3

Source: Middindi Consulting

Intact Rock Strength Testing Uniaxial compressive strength (UCS) tests were carried out on rock samples selected from the boreholes logged on site. Table 48 shows the failure strength properties, Young’s Modulus and Poisson’s ratio for each rock types.

Table 48: Summary of Rock Strength Testing Results

Domain

Strength

Tangent

Elastic

Modulus

Secant

Elastic

Modulus

Poisson's

Ratio

Tangent

Poisson's

Ratio

Secant

UCS at 50%

UCS

at 50%

UCS Gpa

at 50%

UCS

at 50%

UCS

Unit Mpa Gpa Gpa

Chloritised Schist

Average 35.05 15.72 11.87 0.36 0.19

Minimum 22.60 10.90 8.67 0.10 0.07

Maximum 41.92 18.50 14.40 0.52 0.30

Standard Deviation 8.19 3.41 2.21 0.17 0.09

Amphibolite Schist

Average 73.34 39.73 47.76 0.35 0.27

Minimum 65.96 21.40 27.40 0.22 0.21

Maximum 81.84 61.70 69.30 0.47 0.32


Mica Schist

Average 53.05 28.05 23.07 0.34 0.20

Minimum 29.87 21.50 14.20 0.21 0.12

Maximum 83.33 37.00 29.60 0.59 0.32



Chloritised schist and Mica schist exhibit strength values that were lower than anticipated. This can be attributable to the presence of the foliations as well as the altered properties (infill and alteration) of the rock. Results obtained from triaxial compressive strength (TCS) tests further corroborates that the strength data is correct as the projected UCS using triaxial test data negates the effect of foliation.

Derivation of Mechanical Properties for Intact Rock and Joints The results from triaxial tests, UCS tests and data gathered during the geotechnical logging phase were analysed using the RocData program from the RocScience suite. The Hoek-Brown and Mohr-Coulomb parameters for intact rock and joints are summarised in Table 49.

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Table 49: Summary of Hoek-Brown and Mohr-Coulomb Parameters

Rock Type Hoek – Brown Parameters Mohr-Coulomb Parameters

mb s a c (MPa) φ (°)

Chloritised Schist 0.315 0.0057 0.502 0.77 23.54

Amphibolite Schist 0.969 0.0129 0.501 1.83 37.38

Mica Schist 0.389 0.0066 0.502 1.04 29.33


Regional Stresses and Rock Strengths In the absence of stress measurements, a benchmarking exercise was performed to derive the likely k-ratio (horizontal to vertical pre-mining stress ratio) that may exist for the depth of mining at the Bisie Project. The average (k= 1.6) of five reported k-ratios (Brown, 1980, Sheorey, 1983, Stacey, 1994) ranging between 80mbs to 500 mbs for the African region, was selected for Bisie (Graph 20). The intense chlorite alteration defines the general rock strength class as weak to medium strong.

Graph 20: Africa Region K-Ratios


From an intact strength perspective, the rock can be classified as medium strong to strong as described previously and shown in the Table 50. The rock mass (RBS) and design rock mass strengths (DRMS) classifies the rock mass as weak (Table 51).

Table 50: Summary of Intact Rock Strength

Statistics

Chloritised Schist

(MPa)

Amphibolite Schist

(MPa)

Mica Schist

(MPa)

UCS Tests Triaxial

Tests UCS Tests

Triaxial

Tests UCS Tests Triaxial Tests

Mean 35.05 26.00 73.34 66.00 53.05 39.54

Standard Deviation 8.19

NA

5.17

NA

21.28

NA Minimum 22.60 65.96 29.87

Maximum 41.92 81.84 83.33


Within the context of underground mining between depths of 100mbs to 500mbs, strengths in the range 35MPa to 50MPa are considered low and easily susceptible to failure.

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Table 51: Rock Block, Rock Mass and Design Rock Mass Strength

Rock Type IRMR

(MPa)

UCS

(MPa)

Rock Block Strength

(RBS) Rating

Rock Mass Strength

(RMS)

Chlorite Schist 43.83 35.00 28.00 4.00 11.15


Mica Schist 48.96 53.00 42.40 6.00 18.22

Mean 48.90 53.67 42.93 5.33 19.11

Rock Type Weathering

Adjustment

Orientation

Adjustment Blasting Adjustment DRMS (DRMS/RBS) %

Chlorite Schist 0.80 0.90 0.94 7.55 26.96


Mica Schist 0.80 0.90 0.94 12.33 29.08

Rock Type 0.80 0.90 0.94 12.94 29.48


Access Excavations and Support Design The orebody will be accessed via a central decline and a series of levels and Ore drives. A number of other ancillary excavations will also be utilised either for access to or extraction and transportation of the ore. The dimensions of each of the excavations was driven by the selected mining method, access design, equipment selection and ventilation requirements. Systematic support systems were derived using the guidelines for primary support design (based on rock mass classification) provided by Barton et al, (1974) as well as industry common practice. This approach aims to ensure that design is not conservative and that safety is not compromised. Parameters such as ESR, Q Index and joint roughness ratings (Jr) were taken into consideration. Several support classes were then defined based on the intensity of support that will be required and the quality of the rock mass encountered. These support classes are presented in Table 52, ranging from lowest intensity to the highest intensity. For example, Class I support calls for face nets and systematic 1.3m long split sets whereas Class IV support requires face nets, 2m long resin bolts and 4.5m long cable anchors coupled with 50mm of reinforced shotcrete.

Table 52: General Systematic Support


For example, class I support calls for face nets and systematic 1.3m long split sets whereas class IV support requires face nets, 2.2m long resin bolts as well as 4.5m long cable anchors coupled with 50mm of reinforced shotcrete. The support strategy for the excavations was determined based on their inherent ground conditions.

16.3. Ground Water Considerations A groundwater evaluation of the aquifer dynamics and characteristics was undertaken. It relied on observations from site, such as natural groundwater base-flow (including springs) conditions, inspection of exploration cores (‘extrapolated’ to known/similar hard rock aquifers), conditions experienced by artisanal miners and borehole testing of five boreholes. The five boreholes were converted from exploration boreholes. On-going monitoring of boreholes continues under current mine operations. Initially, very small volumes were pumped from the mine workings at the highest level of mining (710mamsl). Groundwater volumes of between 1.2Mℓ/d and 1.7Mℓ/d were predicted for mining level 620, which is the elevation of the Western Valley bottom. Mining has not yet reached these depths as at the reporting date. The deepest level modelled was 530mamsl (210m deep, 90m below the western valley bottom, 290m above the final mining depth) where groundwater volumes of between 3.3Mℓ/d and 5Mℓ/d were predicted.

Support

classes

Systematic

1.3m Long

Split Sets

2.2m Long

Resin Bolts

4.5m Long

Fully Grouted

Cable Anchors

50mm Thick

Reinforced

Shotcrete

100mm Thick

Reinforced

Shotcrete

Face Nets

I x x

II x x

III x x x

IV x x x x

V x x x x

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Some of the springs and seepage zones (termed ‘base-flow’) within the dewatering cone will stop flowing, or significantly reduce, when groundwater levels drop and flow directions change toward the mine. However, perched conditions will prevail at some localities due to the high rainfall. The wet and dry rainfall season flows inside the dewatering zone, were estimated at >10ℓ/s and ±5ℓ/s respectively for the most recent wet and dry rainfall seasons. The total impact on the pre-mining total base-flow from the mountain ridge can be described as ‘small’. Due to the small neutralising capacity and high acid potential in footwall material, the ‘residence time’ (i.e. time that mine water is in contact with broken rock/ore) will be very important. Pumping designs cater for mine water to be pumped out very quickly, no salt build-up will occur and it is likely that sulphate concentrations will be <500mg/ℓ; i.e. lower than the proposed guideline concentration. Although, as concentrations may exceed drinking water guidelines, it is expected to be below guidelines of the DRC Mining regulations and Equator Principles. Cd concentrations may marginally exceed drinking water guidelines and Equator Principles guidelines; however, no guidelines exist in the DRC Mining regulations. Cu, Ni, Pb and Zn concentrations are expected to exceed drinking water guidelines and Equator Principles guidelines, and may potentially exceed DRC Mining regulation guidelines. It is not known to what extent the mentioned concentrations may improve over time after continued flow through the upper mined-out levels. The groundwater quality of surrounding aquifers is unlikely to be impacted given that the groundwater will be flowing towards the mining excavations.

16.4. Ventilation Considerations Ventilation is the primary means of diluting and removing pollutants such as dust, blasting fumes, gases, diesel exhaust emissions and heat, while providing fresh air for employees. The ventilation design gives the intended ventilation method, layout and air quantities and is reported in detail in the previous NI43-101 report. Refer to Definitive Feasibility Study – Mining Study, dated June 2016, Report No. 2014-150-04, completed by Bara Consulting.

Design Criteria and Constraints The ventilation design criteria must conform to established international best practices to provide a safe and healthy underground working environment. Table 53 lists key criteria used in the previous study (Definitive Feasibility Study – Mining Study, dated June 2016, Report No. 2014-150-04, completed by Bara Consulting).

Table 53: Ventilation Design Criteria and Occupational Exposure Levels

Description Quantum

Design Intake Air Temperature (wet bulb/dry bulb) 22.0°C to 28.0°C

Design Reject Air Temperature (wet bulb/dry bulb) 30.0°C to 35.0°C

‘Withdraw from Working Place’ Wet Bulb Temperature 32.0°C

Air to Engine Rated Diesel Power Ratio at Point of Use 0.06m³/s/kW

Overall Air Leakage Factor for the Mine 20%

Declines and Intake Air Tunnels Air Velocity 5 m/s to 8m/s

Return Airways Air Velocity Up to 14m/s

Unequipped Air Raises and Raise Bored Holes Air Velocity Up to 22m/s

Return Air Raises with Emergency Ladders, Pipes and Cables Air Velocity Up to 20m/s

Minimum Design Specific Cooling Power (SCP) 300W/m²

Friction Factor – Declines and Haulages and Crosscuts (average blast) 0.012Ns²/m4

Friction Factor – Unequipped Raises (rough blast) 0.02Ns²/m4

Friction Factor – Ladder Way, Pipes and Cables Equipped Raises 0.03Ns²/m4

Flammable Gas (methane) Limit (within 150 mm (6”) from source) Max. 1.0%


Determination of Air Requirements The air requirements are based upon the active mining fleet and standard mine ventilation criteria. An air to diesel power ratio of 0.06m³/s/kW at point of use was used to determine ventilation requirements. This ratio is

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internationally accepted and assumes modern machinery, a good maintenance regime, pollution control measures such as catalytic converters and diesel filters are used combined with low sulphur diesel fuel. The air requirements are given in Table 54.

Table 54: Air Requirements based on Active Diesel-Powered Fleet

Equipment Type kW Number Units Total kW

Truck 40t MT42 (utilisation 100%) 388 4 1,552

LHD 14t ST14 (utilisation 100%) 250 3 750

Development Rigs and Bolters (utilisation 50%) 80 4 160

Long Hole Rig (utilisation 50%) 80 1 40

Service Vehicles (utilisation 50%) 80 6 240

Total kW Diesel Power in use 2,742

Air Requirements

Diesel Power in use x Dilution Rate in m³/s/kW

2,742 x 0.06 165

Leakage Allowance 20 % 30

Total Airflow Required m³/s 195


The table above gives the base airflow requirements for the mine based on the diesel fleet required. Note that drill rigs and utility vehicles are assumed to work at a relatively low utilisation. Air requirements were checked by ventilation simulation (Ventsim) in the previous study. The simulation showed that the ventilation requirement increases from 170m3/s at the start of operations to 210m3/s later in the life of operations when more levels are developed. This is approximately in line with the calculated air requirements. Good management of ventilation seals to minimise leakage will maintain air requirements as the mining operation is extended with depth. The return ventilation system will consist of a series of 3.5m x 3.5m drop raises going from level to level. These drop raises will be equipped with an emergency escape ladder as well as being used as a service raise. The air from the upcast drop raises and the old workings will be exhausted via a RAW to surface where air will be extracted at the fan station.

Development Ventilation Development operations are well underway and are being ventilated adequately, with ventilation columns delivering fresh air to the various ends underground.

Stoping Ventilation The stopes will be ventilated by retaining the development ventilation columns in the footwall and Ore drives where necessary, with the return air being extracted through these columns and up the exhaust ventilation drop raises to the top of the mine and out via the main RAW and the main fans.

Ventilation Controls The primary ventilation control consists of a series of seals installed in the level access drive immediately after a level has been mined out to prevent short circuiting of fresh air into mined out areas. Backfilled stopes will also minimise ventilation leakage. The main fans consist of 2 x 250kW axial fan units installed in parallel, situated at a single fan station at the upper terrace on surface at the return air decline. The fans have been installed and are operational.

Geotechnical Data Acquisition The geotechnical data acquisition program commenced at a concept level of accuracy in May 2014, with the logging of un-orientated core from five boreholes. The database was later supplemented with information collected from seven bore holes drilled in November/December 2014. Six more holes were logged in April/May 2015 for the portal design. This information facilitated a comprehensive assessment of the geotechnical character of the orebody and surrounding rock mass.

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16.5. Mine Design A mining method selection study (Refer to Definitive Feasibility Study – Mining Study, dated June 2016, Report No. 2014-150-04) considered a number of possibilities before trading off sub level caving (SLC) and cut and fill (C&F). SLC was initially selected but changed to C&F based on recommendations from SRK and concerns regarding surface subsidence, water and clay entry into stopes from hangingwall and poor ore recovery due to the steep orebody dip. However, the C&F mining method is expensive, limits operating flexibility and is unlikely to generate the required Plant throughput at steady state on its own. ABM has therefore implemented a mining method that uses open stope drilling and blasting methods to excavate the ore before backfilling the open stopes with waste rock. It is noted that open stoping with cemented hydraulic fill or paste fill was considered as one of the mining methods during the initial mining method selection process. The latest mining plan is to use waste rock as the fill rather than cemented hydraulic or paste fill.

16.5.1. Mining Method Description The stoping blocks have been designed to include the full width and depth of the ore body. Inter level waste drives have been positioned along strike in the footwall of the orebody. These service 9° ramps that are developed up or down to access the orebody. Ore drives will then be developed from the ramps, along strike (and parallel to the waste drives). A typical mining section would normally consist of three inter levels but the latest design discards an intermediate level as shown in Figure 33. This reduces the total cost of development over the LoM. The dip pillars are designed to be 3m wide along strike and will be spaced at distances ranging from 20m to 50m depending on the depth below surface and the width of the orebody. The current mine design uses a spacing of 35m. The stope heights are planned at 10m vertical intervals. Mining direction will be from the middle of the stope lay-out towards the outer limits of the ore body. Stoping direction between sill pillars will start on the bottom drive and progress upwards from the bottom to allow for fill to be placed in the mined-out stopes. Ore mining will be maintained on the outside of the levels, whilst waste packing will progress from the centre outwards, following the mining operations. On average, a sill pillar thickness of 8m every 58m vertically, will generate a 13.8% pillar loss on dip. The proposed dip pillars (3m every 35m on strike) will contribute some 8% extra pillar loss on strike. A total of 21.7% of the ore body will be lost due to strike and dip pillars. A lower mining rate coupled with a total extraction of 50% will be applied to the high-grade sill pillars during the scheduling phase.

Figure 33: Isometric View of the Mining Method

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Extra waste development may be placed on each inter level drive with a spacing of approximately 60m on strike. These drives will connect the waste drives with the Ore drives on a regular interval and will introduce some flexibility into the access development. The open stoping mining block dimensions for different orebody widths are illustrated in Figure 34, Figure 35 and Figure 36.

Figure 34: Isometric Presentation of the Mining Method for Orebody Width up to 8m

Figure 35: Isometric Presentation of the Mining Method for Orebody Width >8m but <15m

Figure 36: Isometric Presentation of the Mining Method for Orebody Width >15m

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16.5.2. Dilution The planned dilution that was calculated for the stoping, is presented in Table 55. These calculations assumed an average ore body width of 10m and overbreak into the sidewalls of 25cm. Unplanned dilution was assumed to be 10% resulting in a total of 25% dilution for the stopes.

Table 55: Open Stope Overbreak

Description Area Value

Hanging Wall

Planned Area (m2) 43

Broken Area (m2) 70.7

Overbreak Area (m2) 27.7

Overbreak (m) 0.25

Hanging Wall Total (m3) 6.925

Side Wall

Planned Area (m2) 43

Overbreak (m) 0.25

Side Wall Sub Total (m3) 10.75

Number of Sides 2

Side Wall Total (m3) 21.5

Total

Total Volume Overbreak (m3) 28.425

Planned Volume (m3) 180.6

Overbreak Total (%) 15.7


The unplanned dilution is included to take account of the extra waste that occurs as a consequence of the lens like attributes of the orebody in the vertical and horizontal planes. The density for all ore and waste segments will be applied as per the block model during the evaluation phase. Waste development dilution will be set at 10% and ore development and sliping will be set at 6%. Mining recovery and plant efficiency are at 95% and 72% respectively.

16.5.3. Development End Dimensions Access development (Table 56) was optimally positioned to facilitate entry into the centre of the ore body for each mining block. Waste and ore development will be advanced to the outer limits of the ore body and holed before any stoping can commence.

Table 56: Development Dimensions and Overbreak

Description Width

(m)

Height

(m)

Planned Overbreak

(%)

Advance Rate

(t/m)

Decline 4.2 5.1 10 70

Access Drive 4.0 4.2 10

55

Station Drive 4.2 5.1 70

Level Access Cross Cut 4.2 5.1 10 70

Ore Drive 4.0 4.0 10 52

Ventilation Drive 4.0 4.2 10 55

Drill Cubby 4.0 4.2 10 55

Service Raise (Ø) 3.0 - 29

Vent Pass (Ø) 3.0 - 30

Workshop 4.2 4.2 10 57

Workshop Cubby 4.2 4.2 10 57

Ore Medium Drive 6.0 4.0 10 78

Ore Narrow Drive 4.0 4.0 10 52

Open stope - Bottom Access Ore 4.0 4.0 10 52

Open stope - Bottom Access Waste 4.0 4.0 10 52

Open stope - Top Access Ore 4.0 4.0 10 52

Open stope - Top Access Waste 4.0 4.0 10 52


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16.5.4. Development Advance Rate Average advance rates were applied to development and stoping activities as presented in Table 57. The rates quoted are the maximum allowable rate of advance that Earthworks Production Scheduler (EPS) applied to any single mining activity.

Table 57: Average Development and Stoping Advance Rates

Mining Activity Ore/Waste Advance Rate

Decline (m/month) Waste 50

Drill Cubby (m/month) Waste 40

Ore Drive (m/month) Ore 50

Open stope - Top-Access (m/month) Waste 50

Open stope - Middle-Access (m/month) Waste 50

Service Raise Access (m/month) Waste 50

Development Sliping (tonnes/month) Ore 2,688

Stope Slot (tonnes/month) Ore 2,300

Open stope (tonnes/month) Ore 2,560

Ventilation Pass (m/month) Waste 30

Ventilation Raise Access (m/month) Waste 50

Workshop (m/month) Waste 50

Workshop Cubby (m/month) Waste 50


16.5.5. Scheduling Delays A period of seven days delay will be allowed between the completion of Ore drive development and sliping operations. A further seven-day delay will be included between the completion of development sliping and slot cutting activities. A three-day delay will be included between slot cutting and the start of stoping operations to simulate the drilling of parallel holes within the ore body. Waste development volumes available for waste packing will be reported as part of the mining schedule.

16.5.6. Primary Access and Development The current access to the underground workings consists of two adits as illustrated in the schematic diagram in Figure 37. The one adit is developed for ventilation purposes (return air) while the other adit provides access for men, material, equipment and fresh intake air. The only other underground workings consist of various abandoned artisanal workings which are expected to extend to maximum depth of 75m below surface.

Figure 37: Schematic Access to Underground


The existing development extends down to 4 Level as shown in Figure 38.

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Figure 38: Current Development


Future planned development carries on from the current development layouts. The underground development will consist of the decline ramp located in the footwall which will advance to the central position on each level where the footwall drives will be started. The connection from the decline ramp to the footwall drives will be extended to the orebody (one up and one down at 9° inclination) from where the Ore drives will be developed parallel to the footwall drives. The Ore drives will be located along the footwall contact of the orebody. These footwall and Ore drives will extend to the limits of the orebody. Connections between the footwall and Ore drives will be developed as and when required. Ventilation passes for return air will be established on either side of the ramp decline and located midway between the ramp and the orebody limits. These ventilation passes will be established from the waste footwall drive, and connected to surface. The development design is presented in Figure 39 and Figure 40, together with existing development.

Figure 39: Initial Development Design


Development activities are currently underway and the design indicates where the activities will be focussed, in relation to existing excavations.

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Figure 40: Development Design Continued


Figure 41 demonstrates the total amount of development included in the design and schedule.

Figure 41: Development Design End


All development will be completed using mechanised drill, blast, support, load and haul equipment. Waste rock from development will be hauled to a designated stope for backfilling or a designated unused excavation underground.

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All development drilling will be done by a single boom electrohydraulic drill rig (Atlas Copco S1-D-DH). The drill steel length is planned at 3.4m resulting in an expected advance per blast of 3m. The geotechnical assessment has resulted in the ground support requirements in the mine being divided into five classes (I to V). The Table in the geotechnical section of this document summarises the support requirements for each of these classes. Table 58 shows the percentage of each excavation type which is expected to be in a particular ground support class.

Table 58: Allocation of Excavation Type to Ground Support Class

Excavation Rock Type and Description

Support Class (%)

I II III IV V

Decline Predominantly mica schist. May intersect chloritised areas. Entrance will be highly weathered/transitional material.

60 20 10 10

Level Access Cross Cut Mica schist. 70 20 10

Passing Bay Predominantly mica schist. Could be amphibolite schist towards the orebody.

90 10

Footwall Drive Mica schist. 90 10

Ore Drive Predominantly amphibolite schist. May intersect chloritised areas.

90 10

Return Air Drive Predominantly mica schist. 90 10

Service Raise Mica schist. 90 10

Cross Cut Predominantly mica schist. May intersect chloritised areas. 70 20 10

Source: Alphamin

Cover drilling to identify potential inflows of groundwater will be required when developing the decline ramp and footwall drives. Every 35m of advance, a fan of two by 40m long holes will be drilled to provide full cover. These holes will be sealed if an accumulation of water is encountered. Where required cementation methods will be employed to stem inflows of water encountered during cover drilling.

16.5.7. Stoping Once the footwall and Ore drive development on a level is completed and the ventilation passes are established, thereby creating return air ventilation capabilities, stoping can commence. Stope preparation starts with the development of a horizontal drive from the footwall drive to the hangingwall in the orebody. A slot drive is developed along the hangingwall contact. Thereafter, a slot raise is developed to the level above and opened (retreated). Refer to Figure 42 which demonstrates the commencement of stoping diagrammatically. The slot raise will be mined and will hole between sublevels. Slot raises will be mined using longhole raising techniques. The drilling of the holes will be completed with the long-hole production rig. Blast designs will be detailed with an explosives expert to ensure correct explosives, accessories, spacing and burdens and timing of blast holes is optimised. Blast designs will be completed in consultation with a suitable explosives company to assist with ring designs, burdens, spacings, hole diameter, type of explosives and accessories.

Figure 42: Typical Stoping Layout


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The mining block is then sliped to the height of the horizontal drive and supported Figure 43. The long hole rig can then drill up holes for blasting the Ore to the next level above (10m).

Figure 43: Stope Layout - Sliping


Figure 44 presents the total mine design with stoping included.

Figure 44: Mine Design with Stopes


Backfill The mining method requires backfill to be placed in mined out stopes. The backfill material will be recovered from waste development. The voids created by mining includes the slot and stope excavations, but excludes voids created from development and sliping activities.

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The backfill volumes were calculated from the mining schedule and included all waste rock produced from development activities. This volume is indicated in Table 49 and is converted to a bulk volume using a factor of 1.6. It is clear that the development waste will be able to provide sufficient waste material from backfill requirements over the life of mine plan. The additional cost to transport waste back into the mined-out areas must be confirmed, but is not expected to be material as the trucks can transport this waste on the way back from dumping a load of ore.

Figure 45: Backfill Volumes

Volume Volume (m3) Comments

Development Waste Volume 399,624 From mining schedule

Waste Volume (bulk) 639,399 Bulking factor of 1.6

Voids from Slots and Stopes 498,937 From mining schedule


The schedule of waste production from development is compared to the creation of voids from stoping and slots in Graph 21. This shows that waste from development will be stockpiled in old unused underground excavations or on surface. Waste from development activities will be transported by means of LHDs or dump trucks. This waste will be stockpiled and then re-loaded with LHDs and transported back to mined out stopes for backfilling.

Graph 21: Backfill Volume


Load, Haul and Logistics Blasted ore and waste will be loaded by 14t capacity load haul dump units (LHD). Ore will be transported to the level access drive and tipped directly into a truck (40t capacity) for hauling to surface. If no truck is available the LHD will tip into a stockpile in the vicinity of the level access drive and re-handle the material into a truck when one is available. The trucks will transport ore to the RoM pad on surface. All material, consumables and small equipment will be transported into the mine by light vehicles (LDV) or by utility vehicles (UV). Underground personnel will travel from surface to their working place by means of light vehicles.

Equipment selection and productivities Previous studies completed a trade-off study to determine the optimal size/capacity of mining equipment. The trade-off study included a comparison with 6t LHDs and 20t dump trucks compared to 14t LHDs and 40t dump trucks. The study showed that the larger fleet was more cost effective as less units were required. The expected benefit of smaller excavations when using the smaller fleet is not applicable as the size of the decline is determined by the intake ventilation requirement, which is in fact greater in the case of the smaller fleet due to increased diesel-powered equipment in operation underground. Table 59 shows a summary of the selected equipment.

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Table 59: Summary of Selected Mining Equipment

Equipment Type Epiroc Model

Capacity Productivity/Efficiency Development Stoping

Development Drill Rig S1D-DH 1 drifter 170m/month (multiple heading) 41m/month (single end)

2 0

Stoping Drill Rig Simba S7 1 drifter 102m/shift 0 1

Loader ST-14 14t 93t/h 2 1

Truck MT42 40t 49t/h 2 2

Bolter Boltec 1 drifter 7 bolts/h 2 0

UV Material, equipment 2 1

LDV Toyota LC 4x4 Material, equipment, personnel 2 1


Note: Epiroc previously known as Atlas Copco

Figure 46 to Figure 50 show illustrations of the primary mining equipment.

Figure 46: Loader (Epiroc ST-14)


Figure 47: Dump Truck (Epiroc MT42)


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Figure 48: Development Jumbo (Epiroc S1 D-DH)


Figure 49: Support Jumbo (Epiroc BOLTEC 235)


Figure 50: Production Drill Rig (Epiroc SIMBA S7)


The productivity of the selected equipment fleet was calculated based on the following parameters:

• A shift cycle of two ten-hour shifts per day, five and a half days per week. This amounts to an average

of twenty-three workings days or forty-six shifts per month.

• Unproductive time is planned at 1.7 hours per shift which accounts for:

o Travelling time in and out of the mine;

o Pre-shift meeting;

o Rest and meal break; and

o Post production handover.

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Equipment availability and utilisation were applied to the remaining 8.3 hours per shift to estimate the total working hours per month for each unit of mining equipment. Based on the operating conditions, which include haul distances and estimated travelling speeds, the productivity of the loaders and trucks was estimated. Table 60 shows the productivity estimate for mechanised equipment, which are based on reasonable assumptions.

Table 60: Equipment Productivity/Efficiency

Equipment Type Drill Rig LHD Truck Bolter Long Hole Rig

Availability (%) 75 80 80 75 75

Utilisation (%) 551 81 81 542 36

Percussion (h/month) 249 n/a n/a 233 156

Engine (h/month) n/a 403 403 n/a n/a

Productivity (t/h) n/a 93 49 n/a n/a

Bolts (number/h) n/a n/a n/a 7.2 n/a

Ends/shift 1.6 n/a n/a n/a n/a

Quantity (t/month) n/a 37,262 19,791 n/a 28,096


1= percussion utilisation

2= power pack utilisation

The productivities for LHDs and dump trucks is largely dependent on average travel speed and haul distances. The productivity results calculated in the table above are based on the following assumptions:

• LHD travel speed = 4km/h; haul distance = 150m (one way);

• Dump Truck travel speed = 10km/h; haul distance = 800m (one way); dump truck has to wait for three

LHD loads before it can begin its journey (some 24 minutes).

Production requirement of 30ktpm needs the following equipment based on the table above:

• LHD 1;

• Long hole rig 1 (requires slight efficiency improvement to achieve 30ktpm);

• Dump truck 2.

The development process requires two suites of mechanised equipment as follows:

• Drill rig 2;

• Bolter 2;

• LHD 2;

• Dump truck 2;

• UV 2;

• LDV 2.

A model was developed to estimate the time for all development activities over a period of 23 days (46 shifts). This model assumed the following cycle times for the main activities:

• Drilling 6.4h (includes drilling, setting up rig, rig down, move to new end);

• Charge up 2h (includes hole cleaning, charging, timing, connection, clear up);

• Cleaning 2.8h (includes cleaning blasted rock with LHD);

• Support 4.4h (includes drilling, installing bolts, setting up, rig down, move to new end);

• Services 2h (includes extending pipes, ventilation, power, pumping, general);

• Preparation 2h (includes washing, checking blast, grades, lines, marking).

• Total time 19.6h

The model demonstrated that, for multi ends (at least three ends), a total average advance of 170m/month per suite of equipment can be achieved over the life of mine. However, this development rate could range between 150m/month to 200m/month depending on cycle time assumptions as described above.

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16.6. Mining Layout and Scheduling The mine has already commenced with operations and some Mineral Resources have been extracted. In addition, pillars within the orebody were designed according to geotechnical recommendations to maintain stability and provide a safe working environment underground. These pillars will reduce the mineable inventory and therefore form part of the modifying factors. These pillar losses are referred to as layout (or mine design) losses and are presented in Table 61.

Table 61: Layout Losses

Description Quantity

(kt) Grade

(%) Content

(kt) Comments

Mined-out 170.2 4.80 8.2 Forecast to 31 December 2019

Pillars / Layout 1,306.5 5.67 74.1 All pillars *

Total Layout Losses 1,476.7 5.57 82.3


*Note: This represents total pillar and layout losses. The reader is advised that 50% of sill pillar losses will in fact be mined as part of the

pillar recovery plan towards the end of the life of mine

The net losses and dilution from mining operations, combined with Mineral Resources remaining above 4 Level and sill pillar recovery was calculated and presented in Table 62.

Table 62: Net Losses, Dilution and Pillar Recoveries


(%) Comments

Resources above 4 Level 3.7

Mining Losses 5.0

Estimates Planned Dilution 5.0

Unplanned Dilution 15.7

Sill Pillar Recovery 50.0


Mineral Resources to Mineral Reserve Reconciliation The mining schedule was further adjusted by applying the 1.6% Sn cut-off grade and removing all areas where the in-situ grade was less than 1.6% Sn. Stopes that mined areas below this cut-off grade were removed from the mining schedule as indicated in Table 63 and Figure 51.

Table 63: Applying 1.6% Cut-Off Grade


(kt) Grade

(%) Content

(kt) Comments

Cut-off 541.6 1.07 5.8 From schedule


Figure 51 shows the grey areas mainly identified along the outer edges of the orebody which were removed from the final production schedule results.

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Figure 51: Applying 1.6% Cut-off Grade


It was then necessary to identify areas of the orebody which were categorised as Measured, Indicated and Inferred Mineral Resources. The geological block model was examined and the Mineral Resource categories were identified as shown in Figure 52.

Figure 52: Mineral Resource Classification


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The approximate areas with grades lower than 1.6% Sn are also referenced for clarity. It is clear that a large portion of Inferred Resources is removed from the production schedule, however, small amounts remain included in the production schedule. The mine design was then included (‘overlaid’) in Figure 53, with the cut-off areas, to demonstrate where the mine design is positioned in relation to the Mineral Resource categories.

Figure 53: Mineral Resources with Mine Design


Mining Schedule The production schedule is based on reasonable assumptions and calculations using appropriate software and best efforts. The production target (approximately 32ktpm) in terms of quantity mined could be exceeded or actual quantities mined could fall below the target. A variation of 10% either way will affect revenue, and therefore profitability, materially. The production schedule includes a planned grade according to the mine design and rate of mining and is directly related to the area being mined. Mining areas according to the life of mine plan as per the production schedule is essential to achieve revenue forecasts. Again, any deviation from the production schedule may affect revenue and profitability materially. Excessive dilution (in excess of planned dilution assumptions) will negatively affect the grade of ore sent for processing. Conversely, minimising dilution with good controls and mining practice, will improve the grade of ore and enhance revenue and profitability materially. It is therefore essential to implement the mine plan (based on the mine design and production schedule) by mining in the correct areas, at the planned grade, and in the correct quantities over time. Mining costs have a direct influence on profitability and therefore cut-off grades. Lower costs may result in lower cut-off grades which could increase the amount of economically viable ore for mining. This scenario will result in an increase in Mineral Reserves and assumes that all other modifying factors are constant. The implementation and construction schedule may have a material effect on the economics if the planned ramp up to steady state production is not achieved as per schedule. Late implementation will normally result in a slower ramp up with a corresponding negative impact on the operation’s cash flow. Mining schedules for the development and stoping were based on the mine design and developed as a basis for the latest business plan.

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Development: Development rates were estimated as per the design criteria presented in Table 57 and applied to the design. This process resulted in the development schedule presented in Graph 22 and Graph 23. The sliping metres were converted from cubic metres of sliping in the orebody to linear metres by using a factor of 16m2 per cubic metre.

Graph 22: Monthly Development Metres


Graph 23: Annual Development Metres


Stoping: Stoping production rate was based on the mine design and a process plant throughput rate of some 32ktpm, which resulted in the production profile presented in Graph 24. The production profile includes all sources of ore from stoping, slots, sliping, pillars and development and will be transported to the process plant for treatment.

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Graph 24: Production Profile


Product (Sn): The quantity of tin mined and sent to the process plant for treatment is indicated in Graph 25, together with an estimate of product recovered in the plant. It is noted that additional ore are likely to be available for mining following the proposed exploration campaigns and this is anticipated to both enhance throughput and extend the LoM beyond that reflected in Graph 24. Additional production is expected from Mpama North north extension, Wedge, depth extension and Mpama South.

Graph 25: Average Monthly Product (Sn)


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Total Transverse Slot Tonnes Total Transverse Stope Tonnes Sill Pillars Tonnes Mined

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The grade of Sn and Arsenic (As) feed to the process plant is compared in Graph 26.

Graph 26: Product (Sn) and Arsenic (As)


The RoM feed to the plant is presented in Table 64 and Graph 27 with content and annual grade.

Table 64: RoM Feed to Plant

Description 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 2032 Totals

RoM Ore (kt) 315 373 382 388 383 372 360 363 279 273 221 114 33 3,855

Content (kt) 13 16 15 16 16 17 14 16 9 7 8 7 2 154

Grade (%) 4.01 4.16 3.94 4.07 4.16 4.66 3.76 4.35 3.14 2.44 3.66 6.21 6.03 4.00


Graph 27: Rom Feed with Grade


The final production schedule was separated into the Measured, Indicated and Inferred Mineral Resources categories using the block model Table 65.

Table 65: Percent Mineral Resource Categories applied to Production Schedule

Mineral Resource Category

2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 2032

Ore Development (%)

Measured 2.8 0.1 0.0 0.0 0.1 0.0 0.0 0.0 0.5 0.0 0.0 0.0 0.0

Indicated 44.5 23.3 28.4 22.6 12.4 18.6 12.5 5.5 8.8 7.2 0.1 0.0 0.0

Inferred 4.7 3.5 4.4 6.9 3.5 9.7 8.7 7.1 7.4 11.9 0.1 0.0 0.0

Stoping (%)

Measured 2.6 3.8 0.0 0.0 0.3 0.0 0.0 0.0 5.5 0.0 0.0 0.0 0.0

Indicated 44.4 66.9 64.9 68.0 80.7 65.7 75.1 74.3 68.1 63.2 74.6 73.7 64.5

Inferred 1.1 2.4 2.3 2.5 2.9 6.1 3.7 13.1 9.6 17.7 25.2 26.3 35.5

Totals (%)

Measured 5.3 3.9 0.0 0.0 0.5 0.0 0.0 0.0 6.1 0.0 0.0 0.0 0.0

Indicated 88.9 90.2 93.3 90.6 93.2 84.3 87.6 79.8 76.9 70.4 74.7 73.7 64.5

Inferred 5.8 5.8 6.7 9.4 6.4 15.7 12.4 20.2 17.0 29.6 25.3 26.3 35.5


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16.7. Underground Mine Infrastructure

Service Water Handling The mine service water system is in operation and continues to be installed as the development advances. Service water is supplied to the underground mine by a service water header tank located on surface. The mine water header tank is the primary collection and distribution point for all mine water. Other than being responsible for the distribution of water to the underground workings, the tank is used for the distribution of excess mine water to the plant.

Dirty Water Handling The dirty water handling system collects and transfers mine production and development water from the mining areas. The system is designed to accommodate the service water load from the mining activities in addition to a constant ground water inflow rate of 5.0 Mℓ/day. Dirty water from the production areas will be transferred to dirty water pump stations by means of submersible pumps. A submersible pump will transfer this water to the dirty water pump station. Dirty water from the development ends will also be transferred to the dirty water pump stations by means of submersible pumps. The dirty water pump stations will be situated on every second level. The pump stations will transfer the dirty water to the settling facility, situated on surface.

Compressed Air The underground compressed air system will be used solely for the ventilation of refuge bays. The system will be designed on an air usage factor of 85ℓ/minute per man. The factor allows for respiratory requirements in emergency situations and dilution of CO2 build up in confined areas The refuge bays will be capable of accommodating 40 people and will be equipped with chemical toilets, first aid kits, telephone communication, CO/CO2 scrubbers and oxygen tanks.

Electrical The electrical reticulation and equipment required to supply the surface 400V and underground electrical loads is installed and will be extended as and when required. The major underground loads are centred around the mining development and stoping activities, dedicated development and production MSU’s will supply the mining loads. As the mining production deepens, the mined-out levels will require limited power for pumping and other services.

Communication Communication systems are installed and operational. Leaky feeder and telephone systems will be installed and extended as required.

Trackless Equipment Workshop A trackless mobile machinery (TMM) workshop has been established on surface near the portal entrance. The surface workshop will be used for major repairs and maintenance. Underground satellite workshops will be established as the mine deepens and more levels are established. The satellite workshops will be established on every second level and will be used for minor repairs and maintenance.

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17. RECOVERY METHODS

17.1. Description The process plant Figure 55 has been designed to treat between 360ktpa and 430ktpa of RoM ore from the mining operation. This will generate in the order of 15.5ktpa of tin concentrate.

Figure 54: Concentrator Block Flow Diagram

Source: DRA

The Plant comprises the following processes:

• Crushing of Run of Mine (RoM) ore to -8mm;

• Screening of the crushed ore into -8mm +1mm and -1mm fractions;

• The -8mm +1mm is processed by jigging;

• Jig concentrate is milled to 80 % -425µm and processed using gravity spiral concentrators;

• Spiral concentrate is milled and sulphides removed by flotation to provide the bulk of the final

concentrate;

• Spiral tailings are reground to 80% -106µm, sulphides are removed by flotation, and a final tin

concentrate produced by a second flotation stage;

• Jig tailings are discarded to a tailings stockpile, part of the tailing’s storage facility;

• The -1mm screened from the crushed ore is processed using gravity spiral concentrators. There are

two spiral concentration sections. The primary concentrate is combined with the jig spiral concentrate

for grinding and sulphide flotation. The secondary concentrate is combined with the sulphide

concentrates and tin flotation tailings to form a ± 10% tin low-grade concentrate; and

• The -1mm spiral tails are thickened and discarded. Combined final concentrates are treated through a

magnetic separator to remove iron, then filtered and bagged for sale.

• Combined final gravity concentrates and sulphide float product is treated through a magnetic separator

to remove iron, then filtered and bagged for sale.

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Lower than normal availabilities have been used in the design, owing to the remote location of the Plant. The crusher plant has been designed to run at 50% utilisation and the remainder of the Plant at 85%. Ore Handling and Preparation: The Plant is fed with ore by front-end-loader (FEL) from a mine stockpile. A primary jaw crusher will reduce the ore from approximately 450mm to less than 150mm. The ore is then screened and crushed in a secondary and a tertiary cone crusher. The tertiary crusher is in closed circuit with a screen. The final crusher product will be -8mm. Crushed ore is stored on a stockpile. A FEL reclaims the ore and discharges it into a hopper and belt feeder arrangement. This feeds a jig. Coarse Treatment (High Grade): The crushed fraction is jigged in two stages, roughing and cleaning. Jig floats are screened before the +1mm fraction is discarded. Jig sinks are milled to 80% -0.425mm to further liberate tin. A ball mill in closed circuit with a screen is used. Milled material is subjected to gravity concentration on a series of spiral concentrators with the removal of excess water and ultra-fines being done with hydrocyclones. These cyclones and spirals are configured as a rougher and scavengers in series. High-grade concentrate is reground in a closed-circuit ball mill to 80% -106µm to liberate fine tin. Sulphides are removed by flotation. Spiral concentrator middlings and tailings are reground in a closed-circuit ball mill to 80% -106µm to liberate fine tin. Sulphides are removed by flotation. The tin flotation circuit consists of six flotation cells configured as a rougher bank. The concentrate is sent for disposal in a storage system made of geotextile fabrics. This drains water and retains the solids in a storage pond. The water is recycled. The concentrates are treated by a low intensity magnetic separator to remove any magnetic minerals. Fines Treatment (Low Grade): The -1mm crushed fines is subjected to gravity concentration on a series of spiral concentrators. These are configured as a rougher and two cleaners in series. Concentrate is added to the jig spiral concentrate stream for treatment. Spiral middlings and tailings are further treated in a second series of spiral concentrators. The concentrate is classified as low-grade tin concentrate and added to the tin flotation tailings and sulphide concentrates. Filtration: The vacuum belt filter will produce a cake for bagging and dispatch with a moisture content of approximately 10% to 15%. Tailings: Tailings from the fines spiral circuits and recovered water from the jigs will be dewatered in a thickener. Recovered water will be re-used in the Plant. Thickened tailings will be pumped to a tailing’s storage facility. Services: Services will include reagent storage, make-up and distribution facilities, water storage and distribution and air services. Operation: Changes to the LoM plan have resulted in the planned ore feed increasing to about 384ktpa. The Plant will run at a reduced throughput during the 2020 year of the forecast LoM before reaching the new LoM steady state forecast of 384ktpa.

17.2. Test Work The process as described above is the result of laboratory test work at Mintek. The test work indicated that a tin concentrate with a grade greater than 60% tin can be achieved with an overall recovery of at least 80%. DRA worked together with the ABM’s representative, to oversee and monitor the test work. Owing to potential plant performance uncertainties, a process recovery of approximately 72% with a grade of approximately 62% tin concentrate has been specified for the LoM business planning.

17.3. Process Description The complete process design criteria can be found in DRA-C0216-PROC-DC-028. From a macro view point, the process consists of the following, with appropriate dewatering stages in between:

• Front end crushing and screening;

• Coarse pre-concentration with jigs;

• Several stages of milling and gravity concentration using spirals and shaking tables;

• Sulphide flotation and magnetic separation to remove impurities;

• Concentrate dewatering;

• Tailings handling;

• Reagents and services.

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Services include water and compressed air facilities. Front End Crushing and Screening: Ore from the mine will be trucked to the RoM pad and either tipped directly into the RoM feed bin or onto a RoM stockpile. The RoM material will then be withdrawn from the bin by apron feeder and through a jaw crusher, which will reduce the material from the mine (nominally -450mm to -150mm). The jaw crusher product will be conveyed to the secondary crusher sizing screen, whose oversize will be conveyed to the secondary crusher feed bin, while the screen undersize will feed the tertiary crusher screen. The tertiary crusher screen will also be fed by the products on the secondary and tertiary crushers in a closed-circuit arrangement. The tertiary crusher screen oversize will be conveyed to the tertiary crusher bin while the screen undersize will be conveyed to the plant feed stockpile. Under nominal conditions, all of the screening will be conducted dry, with the provision for intermittent wet screening capabilities on the tertiary crusher screen. Material will be withdrawn from the bottom of the secondary and tertiary crusher bins by vibrating feeder, and will be directed through the secondary and tertiary cone crushers respectively. The secondary crusher will produce a product of nominally -32mm while the tertiary crusher will produce a product of nominally -8mm, both of which will be collected on a common conveyor and conveyed back to the tertiary crusher screen. The crushing and screening circuit will operate on 5,200 hours annually and has a design throughput of 100 t/hr. The 1,900t plant feed stockpile will serve as the buffer between the crushing circuit and the remainder of the plant, which will operate on 7,000 hours annually. Jigs: The -8mm material is reclaimed from the plant feed stockpile via FEL, a small plant feed bin and a belt feeder at nominally 52t/hr and is conveyed to the rougher jig. The jig effects a density separation, with the less dense reject material gravitating to the tailings dewatering screen and the more-dense concentrate material gravitating to the cleaner jig. The cleaner jig also effects a gravity separation and the tailings is pumped back to the rougher jig while the concentrate is gravitated to the concentrate dewatering screen. Both the jig concentrate and tailings is dewatered on a screen with a cut size of 1mm, with the combined -1mm slurry being pumped to the low grade gravity concentration circuit. The -8mm+1mm jig tailings, at nominally 37 t/hr, are conveyed to a tailings stockpile for reload by FEL onto trucks for disposal at the TSF and the -8mm+1mm jig concentrate is conveyed to the HG mill feed bin. HG Gravity Circuit: Jig concentrate is withdrawn from the bottom of the HG mill feed bin at a controlled rate, nominally 4t/hr, by weigh feeder and fed into the primary peripheral discharge ball mill. The mill circuit includes a sizing screen and produces a product with a p80 of 425µm which is pumped to the HG gravity concentration circuit. The high grade (HG) circuit is arranged such that a rougher spiral produces a concentrate and tailings, with the concentrate feeding a shaking table and the tailings feeding a scavenger spiral. The rougher shaking table concentrate (‘HG gravity concentrate’) is pumped directly to the product magsep feed tank due to its low contaminants and high tin values, and the shaking table tails are combined with the rougher spiral tailings and pumped to the scavenger spiral. The scavenger spiral produces a concentrate which is gravitated to another set of shaking tables, and a tail (‘HG gravity tails’) which is directed to the LG regrind circuit. The shaking table concentrate (‘HG gravity middlings’) is gravitated to the HG regrind mill and the tailings is recycled to the feed of the scavenger spiral. LG Gravity Circuit: In this section the -1mm stream from the jig product and discard dewatering screens, nominally 13t/hr, is pumped to a rougher spiral which produces a concentrate that is gravitated to a bank of shaking tables and a tailings which is pumped to a scavenger spiral. The scavenger spiral produces a tailings stream (‘LG gravity tails’) which is pumped to the tailings thickener and a concentrate stream which gravitates to the shaking table bank mentioned previously. The low grade (LG) shaking tables produce a concentrate (‘LG gravity concentrate’) which reports to the HG regrind circuit, and a tails (‘LG gravity middlings’) which reports to the LG regrind mill. LG Regrind Gravity Circuit: The HG gravity tails along with the LG gravity middlings report to a regrind ball mill and screen, which produces a product with a p80 of 106µm. The stream reports to a bank of shaking tables which produce a concentrate (‘LG regrind concentrate’) that reports to the sulphide float feed, and tailings (‘LG regrind tailings’) which is pumped to the tailings thickener.

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HG Regrind Circuit: The feed to the HG regrind circuit consists of HG gravity middlings and LG gravity concentrate. The mill circuit consists of a ball mill in conjunction with a screen and produces a product with a p80 of 106µm. The ground stream is then combined with the LG regrind concentrate and pumped to sulphide flotation. Sulphide Flotation, Magnetic Separation and Filtration: Product from the HG regrind mill along with the LG regrind concentrate reports, via a two-hour buffer tank, to a bank of conventional sulphide rougher flotation cells. The sulphide minerals, which are a contaminant in the final concentrate, are floated off with the froth while the material which did not float is combined with the HG gravity concentrate and pumped to the product magnetic separator. The magnetic separator removes any free iron minerals which may still be present along with any iron added to the stream during the various grinding stages. The magnetic separator tails are combined with the flotation froth and pumped to the tailings thickener while the magnetic separator concentrate reports to the product filter via a two-hour buffer tank. The single concentrate stream is dewatered with a vacuum belt filter before being bagged and set aside for storage. Tailings, Thickening and Pumping: The various tails streams produced throughout the process is thickened and the water recovered. The various streams will be pumped and gravitated to a thickener feed box. After flocculant addition, the slurry will be gravitated to the thickener. Thickener underflow will be controlled by pumps which will pump tailings to the tailings disposal section. Thickener overflow will gravitate to the process water tank. Thickener underflow is collected in a surge tank from where it is pumped to the TSF using either of two sets of parallel pumps. Two tailings lines to the TSF are provided. Water Services: In the early stages, raw water will be pumped from a river to the raw water tank and later on, the excess water produced by the mine will be gravitated to the raw water tank instead. The water will then be filtered and pumped to the plant area to serve as level make up water in the process water tank, gland service water for the tailings pumps and make up water for the reagents. The water quality will be assessed, and if necessary, treated to provide potable water for general use. Thickener overflow water will gravitate to the process water tank, from where it will be distributed to the required process areas. Compressed Air Services: An operational and standby air compressor will be used with associated filters, dryers and receiver to provide instrument and plant air.

17.4. Reagents Flocculant: The required quantity of flocculant is emptied into a hopper. From the hopper the flocculant will be fed by screw feeder to a wetting head and then to a mixing/activation tank. After a suitable hydration time, typically two hours, the flocculant is transferred to a dosing tank. The flocculant is then pumped to the thickener. Frother: Frother from isotainers is pumped directly from the isotainer to the required addition points. Sulphide Collector: Xanthate pellets are weighed in the store and added by hand to the storage tank and diluted. The xanthate is then pumped to the sulphide flotation circuit.

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18. PROJECT INFRASTRUCTURE

18.1. Introduction

The mine site is accessible from Kisangani in the north by land and from Goma in the east by air. The main ports available to the site are Mombasa in Kenya and Dar es Salaam in Tanzania, with transport being by road. Air transportation to site for some items and equipment is available but the transportation of some products like explosives in the DRC is not permitted. A new access road was established from the national road to the mine site. The road from Goma to Walikale is approximately 220km and in poor condition, but may be upgraded by provincial authorities. The road from Walikale to Logu is approximately 60km. The road via Beni and Uganda is in reasonable condition and is used for the operational supplies and exporting concentrates. Alphamin has constructed an access road from Logu to Bisie (32km). Use of this road in wet conditions is difficult and the road will be upgraded. The mine site comprises of an established mining and mineral processing operation with all associated infrastructure for these activities and the staff that perform them. The site has offices, staff accommodation, first aid clinic, messing facilities, core storage facilities and is currently serviced with private company phone and internet services via satellite link. The mine infrastructure constructed and planned to be constructed Table 66.

Table 66: General Infrastructure – Completed and Planned

Infrastructure Description Completed Planned

Box Cut and Portal -

Spiral Decline

Ventilation Adits (temporary and permanent)

Underground Void (mining development and stoping)

Underground Workshops and Services

Mine De-watering and Discharge Reticulation (clean and dirty water)

Ventilation Fans (and associated infrastructure) -

Compressor House, Lamp Room and Crush -

Conveyors -

Rock Handling Facilities (ore and waste)

Mineralised Waste (waste rock dump/s)

Mineral Processing Complex and Laboratory -

Tailings Storage Facility

Surface and Sub-surface Parking -

Change House (i.e. mine and plant) -

Mine Offices, Training Centre and Clinic -

Stores and Electrical Cable Yard -

Workshop, Wash, Tyre Bay and Storage Area (new and defunct) -

Control and Spares Room -

Water Supply, Management and Treatment Infrastructure

Electrical Power Generation, Electrical Reticulation

Overland and Buried Pipelines

Laydown Areas

Roads (i.e. internal and haul)

32km Access Road (i.e. construction and maintenance)

IT and Communications

Accommodation (junior and seniors) -

Security Fencing

Domestic and Hazardous Waste Management and Treatment

Explosives Magazine -

Waste Oil Storage, Treatment and Disposal -

Fixed Wing Private Airstrip (Kokoli) -

Bridges (i.e. over Bisie River and rivers along access road)

Fuel Tanks at Logu and Bisie (i.e. storage and distribution)

Surface Substation -

Incinerator -

Topsoil Stockpiles

Borrow Pits

Prospecting at Mpama North and South

Bulk Sampling (i.e. PR5266, PR13278, PE75) -

Nuclear Storage Facility -

Source: ESIA 2014

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The general site layout is shown in Figure 55.

Figure 55: General Site Layout

Source: Alphamin

18.2. Mining Infrastructure The infrastructure at the portal is located over three terraces, to suit particular requirements and layouts. These terraced areas are:

• The portal terrace, where the primary mining surface infrastructure and access decline portal are

located;

• The return airway (RAW) terrace where the permanent ventilation fans and compressors are located;

and

• The uppermost terrace where the settlers are located.

The mining infrastructure has been designed to provide adequate support for the duration of the life of mine and includes the following: security and access control at portal, mining offices at portal, office complex at portal, potable water at portal, daily stores at portal, trackless mining machinery workshop, fire suppression at portal, old and new oil storage at portal, diesel storage and dispensing at portal, sewage handling at portal, lamp room at portal, service water tank at portal, compressors at portal and proto room at the Plant. Access and Security: A security fence was constructed around the perimeter of the terraced mining areas to secure the facilities. Access to the mining terraces is via the haul road from the Plant site and controlled by the security office at the main gates. Vehicles are permitted access to the terraced areas through the gates, with personnel access being limited to a turnstile system. Only mine owned, maintenance, stores and supervisory type vehicles are permitted access to the portal terrace area. The portal entrance is secured to control vehicle and personnel access to the underground workings, via a gate and turnstile system, which is monitored from the Decline Shaft Control Room. The access system operates in conjunction with the movement of personnel through the lamp room and is electronically controlled and monitored from the control room. Parking for light vehicles is available outside of the portal terrace fence, near the security office. Mine Offices: Offices for the personnel associated with mining, production activities, technical services, general management and administration personnel are located in the main offices adjacent to the Plant in converted containerised offices.

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Lamp Room and Crush: The containerised lamp room, control room and first aid building at the portal entrance comprises three essential facilities:

• The lamp room, holding racks of cap lamps and self-contained self-rescuers (SCSR), gas detection

devices and a lamp repair bay;

• The portal control room, for the monitoring of underground vehicles and emergency situations; and,

• The first aid station, where basic medical care can be administered to personnel located at or near the

terraces and underground.

The entry into and exit from the building is controlled and monitored specifically to account for personnel presence underground. Daily Stores: A containerised daily stores facility is provided at the portal terrace area, for holding and issuing consumable and routine spares items which are generally required on a daily basis, to support the underground mining and equipment maintenance activities. This store is supplied via the main stores which are located at the Plant. Mining Equipment Workshop: The trackless mobile machinery (TMM) workshop is established on surface near the portal entrance, on the portal terrace. The workshop has been designed to accommodate the planned mobile underground fleet and includes provision for carrying out the following activities:

• Vehicle and equipment inspection, cleaning and servicing;

• Mechanical repairs;

• Boiler making and fitting;

• Electrical repairs;

• Hydraulic component repairs and replacement hydraulic hose assemblies; and

• Tyre changing and repairs.

The workshop also includes a canteen, ablution facilities for personnel on the upper and lower terrace areas and offices to accommodate the workshop supervisory staff. Fuel Dispensing and Storage: The fuel storage and dispensing facility is located on surface at the portal terrace area, in close proximity to the portal entrance. There are two surface storage tanks and dispensing points for the refuelling of all underground mining and portal-based vehicles. Other surface support vehicles will refuel at the bulk supply facility located at the Plant. Fuel will be supplied to underground based equipment by means of purpose-built cassettes, which will be refilled at the portal diesel dispensing facility. Diesel supply to the portal storage facility is from the bulk fuel storage facility at the Plant, with diesel being dispatched as required. Oil Storage and Handling: Small quantities of new and used oil and old oil drums are stored in the vicinity of the TMM workshop and the diesel storage facility on the portal terrace, with new oil being supplied from the bulk storage facility located at the Plant. Old oil from the workshop and underground service bays, together with recovered oil from the separator and skimmer unit on surface, are returned to the bulk storage facility for further handling and disposal. Oil Removal and Separation: The oil separation and rope skimmer recovery system is installed near the TMM workshop to remove hydrocarbons from the run-off and wash water collected from the workshop, fuel dispensing and oil storage areas. Grey water is recovered and pumped to the settlers for re-use, and the extracted hydrocarbons are gathered into old-oil storage drums for disposal. Due to the high levels of rainfall, surface facilities are sheltered where possible to reduce the possible hydrocarbon contamination of run-off water.

18.3. Services Potable Water: Potable water for the surface infrastructure and underground mining is stored and supplied from the Plant. Potable water is distributed manually to all buildings, offices and the workshops on the portal terrace. Underground personnel will fill their personal water containers en-route to their working places, from the point of supply located outside the lamp room. Compressed Air: A set of compressors is housed under cover, on the RAW terrace and reticulated to provide compressed air to the underground refuge bays and the TMM workshop.

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Service Water: Service water is provided to the underground mining activities from a lined pond with a capacity of 210m3 located on the uppermost terrace, alongside the settlers. The pond is partitioned to provide two separate water levels to accommodate the required volumes of both the fire water and service water. The two supplies are independent of each other and reticulated as separate systems. The pond is supplied with clarified water from the settlers and any overflow is transferred to the Plant. Fire Protection and suppression: The larger compartment of the Braithwaite storage tank located on the upper terrace will be utilised for the provision of water for fire protection. Water will be drawn via the reticulation system to feed the various installations on the lower terrace area. The fire protection system also includes hydrants connected to the mains reticulation, fire hose reels and portable extinguishers located in all surface buildings. Dirty water handling: Dirty water from the surface and underground activities is transferred to the settler installation located on the uppermost terrace to remove the suspended solids. The installation includes a flocculent plant, settlers and mud press. As mentioned above, the clarified water is transferred to the service water tank for redistribution to the Plant and underground mining activities. A mud press system is planned for the collection and handling of solids, which, once processed, will be transported to the Plant for disposal.

18.4. Process Plant and Administration Infrastructure The Plant and administration facilities include the following:

• Access roads within the Plant site area;

• Plant administration buildings, including, but not limited to, security office, change house, workshop,

stores, main administration offices, medical facility and warehouses;

• Sewerage treatment and disposal;

• Water services including of raw water abstraction, potable water and fire water;

• Accommodation;

• Security; and

• Communications.

Access Roads within the Plant: The roads within the Plant area were stripped of organic material and compacted to facilitate access to the requisite areas within the plant. Drainage ditches and culverts were created in accordance with the requirements for site drainage and channelled to the storm water dam. Plant Administration Buildings: Buildings located in the Plant area consist of a security office, change house, Plant workshop, control room, general administration offices, medical facility, reagent warehouses and spares warehouse. Security Office and Change House: The Security office and gatehouse is located at the main site entrance and is constructed from 6m containers. The gatehouse controls all vehicles in and out of the Plant. There are ablutions for males and females. The Plant operates in shifts. There are not more than 60 people per shift and the ablutions only need to accommodate one shift at a time. The main access gate to the Plant has a security office for the control of vehicle access to the Plant. Plant Control Room: A dedicated plant control room is located in a stacked container arrangement. The control room houses the SCADA system and provides operators with an elevated view of the Plant. Plant Workshop: A workshop with an area of 480m2 was established adjacent to the Plant to enable repair of plant equipment. The workshop is a steel frame building equipped with a three-tonne overhead crane and has bays for servicing light vehicles. The workshop has separate areas for mechanical and electrical repairs. Provision was made for oil separation of any water leaving the facility. Offices for supervisory, workshop store, maintenance and planning personnel are provided in the form of a modular building situated adjacent to the workshop. Administration Buildings adjacent to the Plant: The administration building is constructed from converted containers. The building includes general areas for engineering, administration personnel and offices for the general manager, mining manager, plant superintendent, administration superintendent, chief geologist, plant maintenance superintendent and chief security officer.

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Assay Laboratory: A third party service provider ALS Chemex manages the assay laboratory which is also in the form of containerised construction. The laboratory was designed and supplied by a different independent third-party laboratory service provider. This laboratory conducts all of the onsite test work for the Plant and grade control. Medical Facilities: A containerised medical facility is available for the treatment of any injuries during the construction and operational phases, as well as treatment of sick personnel. Sewerage Treatment and Disposal: Sewerage from the plant is be gravity fed to a sewerage conservancy tank. The sewage is then pumped to a sewerage treatment plant. The sewerage plant treats the water in accordance with South African DWEA General Limits for the release of treated sewerage into the environment. Water Services: The operation is located in a net water surplus climate. To minimise the volume of noncontact surface rainfall run-off reporting to the TSF or the open pits, water diversion channels and ditches were constructed. The preliminary water balance for the site indicates that there is a surplus of contact water that will require discharge to the environment. Water is initially being sourced from the Bisie River by pumping from a position upstream of the road crossing. At a later stage it is expected that de-watering of the mine will provide water in excess of the plant requirements. This water will be filtered for use as spray and gland seal water and will be treated for use as potable water. Water is recovered from the jigging circuit, the tailings thickener and the tailings dam for re-use in the Plant. Storm Water Dam: A necessary facility to accommodate a 1:20 year storm event. Potable Water Distribution: Potable water is provided to the mine accommodation camp and plant area via a potable water treatment plant. Raw water is supplied to the potable water treatment plant from a nearby spring. The potable water plant designed to supply sufficient water for up to 1,000 people per day. Potable water is reticulated to all areas of the plant including a safety shower header tank. Fire System: The fire water system consists of a fire water dam built on top of the ridge. Fire water is reticulated throughout the mine, process plant and accommodation camps by gravity. In addition, two portable trailers with chemical firefighting equipment have been procured. Numerous portable fire extinguishers are located at strategic locations within the Plant facilities and the mine accommodation camp. Accommodation: The camp can provide accommodation suitable for housing 500 persons. This camp includes the following infrastructure:

• Kitchen and camp dining room;

• Entertainment area including a gym;

• Laundry;

• Sewerage treatment plant; and

• Security building.

The mine accommodation camp was designed to house junior staff in 20-man dormitories and twin shared senior staffing accommodation. Materials of construction included corrugated roofing and locally sourced timber as well as containers. Flooring was a combination of a locally mixed cement and soil combination. Security: The Plant site is enclosed within a security fence (additional to the greater project site). Access to the Plant area is be via gates located on access roads to the site. Additional fencing is provided for further safety and security within process Plant areas, such as fuel storage, transformers and substations, and as required. Communications: Communications is via a dedicated satellite link internet service. Limited local cell phone access is also available and a third-party telecommunications service provider mast (Vodacom) has been erected on site. Telephone facilities, mobile radios and satellite infrastructure are provided as necessary to staff. Access Road to Site: The site is accessible from Goma and Kisangani via National road network with the last leg of the road being a 32km access road linking the mine to the National road. Fixed Wing Private Airstrip (Kokoli Airstrip): The construction and licencing of a fixed wing private airstrip adjacent to the access road was completed as part of the construction works Alphamin uses the airstrip to move personnel and critical equipment to and from site.

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18.5. Power Supply and Distribution

Power Supply Power is provided from a diesel power station at 400V three- phase 50Hz. Cables link the Diesel Power Station to the Plant substation. The power reticulated to the mining operation is transformed into 11kV to avoid excessive voltage drop over the 1.5km powerline linking the mine with the diesel power generators. The diesel power station is modular, each of the three units can produce 1.2MW, and can be expanded in size as the pumping requirements expand with the depth of mining. Fuel storage for the generator plant is contained in self bunded tanks with capacity of 120kL. The Logu base camp is used as a delivery point for local suppliers where current bowser capacity is 440kL, increasing to a 1,000kL facility in F2020 at the fuel supplier’s expense. The fuel storage capacity at Bisie and Logu amounts to 2-months estimated consumption. The distribution voltage is 400V for small/medium sized motors, welding feeders, cranes etc. The MCC control voltage is 110V AC. The specification and selection of electrical equipment has been in accordance with South African Standards (SANS Standards).

18.6. Mine Residue Storage Three streams of mine residue (or waste) reporting from the Plant and the mine, require storage within the permit area. The residues requiring storage are:

• Waste rock from underground mining;

• Slurry Tailings (Spiral Tailings) (<1mm) hydraulically pumped from the Plant; and

• Coarse, Dry Tailings (Jig Tailings) (1-10mm) trucked from the Plant.

18.7. Waste Controls ABM has identified its key hazardous and general waste streams. In order to meet the objectives of its waste management standard, ABM will operate the following waste management hierarchy:

• Source, reduction and reuse;

• Reduction in potential waste coming to site;

• Avoiding waste generation;

• Recycling and composting;

• Where financially practical, collecting, sorting, and processing of waste materials that can be used as

resources elsewhere (i.e. wood off-cuts to community);

• Diversion of organic material from the management and accommodation camps (i.e. vegetation

trimmings and food scraps) from landfills to compost; and

• Disposal and incineration will be the final option in the waste hierarchy (i.e. to incineration and or

landfill).

ABM executes this waste management hierarchy through collection, storage, sorting and transporting the wastes to their respective end points.

18.8. Health and Safety A clinic is available on site for the treatment of minor injuries and illness. An airstrip is available to provide medical evacuation facilities if required. Goma can be reached within 30 minutes and Kenya medical facilities can be reached within five hours if required.

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19. MARKET STUDIES AND CONTRACTS

19.1. Tin Market Analysis According to the International Tin Association (ITA), for a decade, tin has experienced relatively stable production from a handful of significant miners and smelters with steady demand growth driven largely by increased use of tin solder in electronic applications, especially in China. Recent fears of global recession, and the USA-China trade disputes had a short-term negative effect on the tin market affecting demand, stocks and prices in 2019. These short-term negative factors temporarily overshadowed the longer-term trend and future expectations of continued tin demand growth and ongoing supply stagnation (Figure 56).

Tin Use by Sector

Figure 56: Refined Tin Consumption by End-use (2018)

Source: ITA 2019

Solder: Tin is the primary component of both leaded and lead-free varieties of solder used in electronics and continues to be the top use for the metal, representing approximately half of global consumption. The total tin usage in the solder sector in 2018 was 175kt, having grown at an estimated 2% from 2017. The long-term outlook for solder usage remains positive as the threat of miniaturisation is countered by the expanding demand from electronic applications, especially 5G communication electronics. Recent cyclical downturns in semiconductor sales coupled with the impacts of USA-China trade tensions have dampened the sector’s growth in the short-term. Chemicals: Tin use in chemicals overtook tinplate as the second largest tin application in 2014 and continues to grow in this respect. Important chemical applications of tin include PVC stabilisers, polyurethane foam manufacture and glass coatings. The total tin usage in the chemical sector in 2018 was 67kt tonnes, having grown at an estimated 3% from 2017. The long-term outlook is positive for chemicals, driven materially by the PVC stabiliser and polymer catalyst markets. Tin Plating: Tin use in tinplating remained the third largest application for tin in 2018, despite the long-term trend of slow decline in the sector. The total tin usage in the tin plating sector in 2018 was 52kt tonnes, having declined by an estimated 2% from 2017.The long-term outlook for tinplate is negative due to expected continuation of lower tin coating weights and competition from alternative packaging. Other: Other uses of tin, accounting for the remaining 20% or 78kt of tin in 2018, are dominated by lead acid batteries and usage in copper alloys, with relatively minor applications also in tin powders, wine capsules, tinned wire, and pewter.

Other ; 8%

Copper Alloys; 5%

Solder; 47%

Tin Plating; 14%

Chemicals; 18%

Lead acid batteries; 8%

Other Copper Alloys Solder Tin Plating Chemicals Lead acid batteries

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Tin Supply Primary mined tin continues to be dominated by China, Indonesia and Myanmar accounting for ~70% of global supply while 75% of refined production is derived from China, Indonesia and Malaysia. Primary mined tin represents ~68% of annual refined tin production with secondary sources recovered from tin alloys, notably solders, brass, bronzes and lead alloys account for 32% in 2018. The rest of the world (ROW) contributes around 10% of global supply. Annual tin supply has been relatively stable for at least a decade hovering between 330ktpa to 380ktpa (Figure 57).

Figure 57: Primary Tin Producers (2017)

Source: ITA 2019

In recent years artisanal and small-scale mining has accounted for as much as 60% of world production, although this share has now dropped to below 40%. The main centres of artisanal mining have been Indonesia, China, Bolivia and Central Africa. Depletion of ore deposits and stricter sourcing disciplines from smelters is expected to result in a continuing decline in Indonesian and other artisanal production centres production. Recent demand weakness and subsequent price contraction in 2019 has prompted several producing nations or companies to announce a cut in production for 2020. The collective proposed cuts by fourteen Chinese producers of 20.2kt as well as the Indonesian announced cut of 10kt could have a material market effect if implemented successfully. Additionally, Indonesia has curtailed exports of tin to many private smelters, and Myanmar output in 2018 is believed to have fallen 21% to only 35kt, both of which further constrict global concentrate supplies.

Tin Demand Tin demand has grown steadily at an average annual rate of ~1.9% since 1980 and 2.0% since 2013. This growth has been largely driven by the Chinese and emerging markets transition to high tin content lead-free solders. Solder in electronics is the largest end use for tin and accounts for 47% of tin consumption, followed by chemicals (18%) and tin plating (13%). China is a major manufacturer of electronic products and the largest consumer of tin globally (45% in 2018). The Chinese economic slowdown, semiconductor downturn, and USA-China trade wars have impacted tin demand in 2019 despite the long-term fundamentals of demand remaining positive.

Tin Stockpiles The historically range-bound tin supply coupled with the steady historical demand growth, had resulted in the tin market being in deficit for ten out of the last eleven years. Unsurprisingly, global tin stocks dropped from 66kt in 2009 to less than 20kt in 2018. However, with the recent drop in demand and a slip in prices, low stock levels have recovered to a degree from their low base in 2018 to around 28kt in October 2019.

China35%

Indonesia17%

Peru6%

Bolivia6%

Brazil9%

Myanmar17%

ROW10%

China Indonesia Peru Bolivia Brazil Myanmar ROW

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19.2. Supply and Demand Balance Industry commentators and the International Tin Association (ITA) maintain their forecasts of a long-term gap between supply and demand (Figure 58).

Figure 58: Forecast Production/Consumption Estimates to 2020

Source: ITA 2015 Tin Use Snap Survey

Forecast Tin Price Many industry organisations and commentators are expecting a continued price turnaround from the low price in mid-2019 of ~USD16,000 driven by increased solder demand, easing USA-China trade tensions and supply shortages in future. In October 2019, the London Metals Exchange (LME) forecast an average price of USD22,000/t for 2020. This 2020 forecast was in line with 1H 2018 prices where tin was the best performing base metal on the LME. Given the high-grade and low unit cost per tonne tin produced nature of the Bisie Mine, Alphamin is more resilient to low price conditions than most existing producers. Although most market commentators forecast long-term tin prices exceeding USD20,000/t, Alphamin has conservatively used a forecast of USD17,000/t for planning, which was below spot at the date of writing.

19.3. Marketing Contract Following an adjudication process, in December 2017, Alphamin concluded an offtake agreement with Gerald metals SA (Gerald) for its tin concentrates. The agreement is for a five-year period commencing from declaration of commercial production and for 100% of production at a minimum tin content of 60%. There is however a contract termination provision in the agreement in the event of a change in control. The price agreed will be the LME price less direct costs such as treatment charges, impurity charges and other expenses including transport, insurance, handling, sampling costs, RMI and or OECD costs to ensure conflict free sourcing and storage costs. Impurity charges apply to sulphur, fluorine, chlorine, arsenic, lead, zinc, silver, nickel, copper, cobalt and other metallics. Net payable tin will be calculated from the certified smelter assayer less applicable standard unit deductions and impurity unit deductions for iron, manganese and wolfram (tungsten). The terms, rates and charges are considered to be within industry norms.

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20. ENVIRONMENTAL, PERMITTING AND SOCIAL IMPACT

20.1. Permits and Legal Aspects ABM has received all of the approvals necessary for mining under the appropriate policies and legal framework within the DRC. The operation is also aligned the with the International Equator Principles (Version 3, 2013) and certain requirements of the International Finance Corporation (IFC). Current environmental assessment process: The approved Environmental and Social Impact study and the Environmental and Social Management Plan are in accordance with Article 463 and the associated Decrees. There is a five-yearly update requirement. Reclamation bond status: ABM have provided a financial guarantee to the with the DRC authorities for rehabilitation. This bank guarantee has been topped up three times (2016 – 2018) and the funds cannot be accessed by governmental departments. Exploration environmental status: The exploration permits have a Mitigation and Rehabilitation Plan (PAR). The PAR sets out the type of exploration activity in the area and describes what measures will be carried out to ensure impacts are minimised and any significant damage is repaired. Exploitation environmental status: ABM met its environmental obligations during the conversion of its exploration permit to an exploitation permit under the Mining Code which required the preparation of an ESIS. ABM is required to carry out an environmental audit every two-years from the date of approval of the initial ESIS and continues to meet this requirement. Permitting status: Table 67 contains the permitting issues that could potentially, materially impact ABM’s operations (i.e. issues noted in the December 2018 audit report).

Table 67: ABM’s Permitting Status

Description Details Granted/Renewed Renewal Date

Exploration PR 5266

Permis de Recherches Environmental Plan Environmental obligations under Environmental

Plan 24 November 2014 23 November 2019

Exploration PR 5267

Permis de Recherches - Currently under Force Majeur Environmental obligation under

Environmental Plan 29 September 2006 23 May 2023

Exploration PE 13155

Permis d'Exploitation - Mining Permit 151 Blocks (±128.96 km2)

3 February 2015 3 February 2045

ESIA - DRC PE 13155

Existing ESIA - Environmental Adjustment Plan None required under new legislation

ESIA - DRC PE 13155

Existing ESIA - Amendment 20 September 2019 19 September 2024

Commercial 001/ 2019

PV Chargement Mine - Mineral Export Document. Information requested from ABM

Commercial 006/ 354/ F. Kay, M.B.B/ 1.0/ 2017

Authorisation to import explosives Valid for three months,

certificates only available until August 2018

Commercial Authorisation for water purification. Information requested from ABM

Commercial Permit for radioactive bunker. Information requested from ABM

Commercial (1896)

Permit to store hydrocarbons 15 February 2018 Being renewed for

2019-2020

Construction Authorisation for the construction of specific

infrastructure. Information requested from ABM

Source: ABM

20.2. Environmental Aspects Through the conversion process of the Permis de Reserches (exploration license) into Permis de Exploitation (mining license), ABM was required to submit an Environmental and Social Impact Assessments (ESIA) and Management plans to the Cadastre Minier for approval. The plans were reviewed and approved by the relevant authority and as such, the mine has full environmental authorisation to operate. Compliance audits occur every two years to monitor progress against the Environmental and Social Management plans. Mine dewatering: ABM will need to pump underground water to surface at some stage. This water will be used in the Plant, and any excess will be discharged into the surrounding environment in a controlled manner.

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Acid mine drainage (AMD): Geochemical studies and groundwater modelling have identified the mining excavations, TSF and waste rock dump as the main potential sources of AMD and ground water impacts. The level of impact is yet to be determined through modelling, and accordingly a specialist is to be appointed to both assess and provide suitable management and monitoring guidance to ABM. Tailings storage facility (TSF): Work by Epoch Resources identified that contaminates (i.e. AMD or dust) from the TSF could impact on people’s wellbeing. Management are cognisant of this and have implemented measures to manage this risk and the risk of harm or damage due to a partial or total failure of the TSF. Sewage and washing water: Grey water will be disposed of as packaged sewage. When a sewerage tank is full, the tank is emptied and transported to the package treatment plant where it is treated to an effluent that meets the requisite DRC discharge limits (ABM Water Management Procedure, 2017) This treated effluent can then be discharged directly to the environment or used as process water. Sludge from the treatment package plant is occasionally removed and disposed of in accordance with the EHS guidelines for Water and Sanitation (IFC, 2007). Water management and monitoring: Surface or storm water contaminated by oils from the mining operation is captured and treated prior to release. A Water Quality data review was carried out in 2015. It serves as the baseline for comparing the quality all water on the mine. No subsequent data has been collected. A water management procedure was commissioned by ABM in 2017, to guide water management at the mine. An update to this is planned which will cover a review of all hydrological issues including:

• Flow monitoring;

• Probabilistic/seasonal water balance;

• Flood-line mapping; and

• Water quality associated with the Jig and Spiral Tailings Deposition Facilities (TDF), waste rock dump sewage and wash water, landfill leachate, run-off and stormwater.

Climate change: An analysis of the volume of Greenhouse Gas (GHG) emissions was undertaken as part of an Air Quality Impact Assessment study to understand the mine’s impact on climate change. ABM is required to pay a fee for each tree destroyed, and since having commissioned the operation, is not expecting any further material impacts in this regard. Air quality and dust: In line with Article 466 of the Regulations ABM is required to monitor air quality to demonstrate the efficacy of the mitigation measures. Terrestrial fauna and flora: The Mine operate in a dense moist forest where the quality of the terrestrial habitats ranges from highly sensitive and relatively undisturbed primary forests to highly impacted crop cultivated areas. The area enjoys high floral biodiversity with multiple species and two were identified as having conservation significance. The hinge-back tortoise (Kinixys erosa) is a protected species and the only globally threatened species observed was the African Grey Parrot (Psittacus erithacus) which is listed as Vulnerable by the IUCN. Soil and land use: Humus and topsoil are important to the rehabilitation of the area. The rehabilitation of borrow pits has started on the mine. Hazardous and general waste: ABM has implemented a waste management procedure which also provides guidelines for the operation of a landfill and incinerator. Rehabilitation and mine closure: ABM will monitor the ongoing rehabilitation measures at the TSF to see that they are sustainable. Success in this regard will reduce the closure liability including post mining aftercare and maintenance costs. ABM is providing for mine closure, which with monies already held in trust (~USD2 M) is sufficient to meet ‘Equator Principle’ equivalent standards, as at 31 December 2019. This also includes a provision of USD0.2 M for post closure activities which will include groundwater monitoring, sealing of the access portal, flooding of the mine and landscaping where appropriate to manage surface run-off. It is understood that the final mine water level will be deeper than the pre-mining groundwater table, which means that the mine is not likely to decant contaminated water. Mitigation measures have accordingly not been proposed for mine decant but this assumption should be verified prior to mine closure.

20.3. Social Aspects Previous socio-economic environment work identified the Manoire residents as the only direct Project-Affected Community (PAC) to be impacted by the mining operations. There are numerous towns and villages along major road networks that may be affected indirectly. More information regarding the following is contained in the previous NI43-101 report:

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• Demographics.

• Age, Gender and Ethnicity.

• Income and Employment.

• Cultivation and Livestock.

• Small Businesses and Other Income Opportunities.

• Income and Expenditure.

• Education and Skills.

• Health, Water and Sanitation. Social or community status quo: The local people are relatively poor and rely on subsistence farming for their livelihood. They generally suffer from malnutrition and disease (e.g. malaria, TB and HIV/Aids). Community and Government Stakeholder Relations: The Alliance Lowa ASBL is an ABM foundation created in partnership with all the local communities of Walikale to stimulate sustainable economic development in the region. Community Development: Progress in this regard is being tracked through the level of community employment and a range of community initiatives focusing on health, education and lifestyle improvement.

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21. CAPITAL AND OPERATING COSTS

21.1. Capital Expenditure Operations at Bisie Mine have reached commercial production but will ramp-up further during 2020 to a production target of 32ktpa as forecast by the 2019 LoM plan. Capital expenditure of approximately USD4 M has been allocated for on-mine capital in 2020. The capital requirements beyond 2020 have been catered for as a collective in the sustaining capital provision (Table 68).

Table 68: LoM Capital Provision Summary (in real terms)

Description Cost

(USD’000)

F2020 On mine Capital 4,030

Mpama South Exploration Drilling/Studies 2,710

Exploration 530

Fine Tin Recovery Project 1,000

Sustaining Capital 48,830

Total Capital Provision 57,100


21.2. Operating Cost Estimate It is noted that the Mine is in operation and the current costs have been used as a baseline to project future costs. At a steady state of 32ktpm for the period January 2020 to December 2028 the on-mine operating costs are tabled below (Table 69).

Table 69: Direct On-mine Operating Costs

Description Cost

(USD/t ore) Cost

(USD/t tin sold)

Mining 22.95 734.75

MRM Function 5.41 175.64

Processing 8.00 257.60

TSF 6.80 220.47

G&A 29.77 965.67

Diesel 20.05 650.40

Site Infrastructure 14.78 479.53

Transport and Duties 2.79 89.68

Wages and Salaries 56.83 1,843.70

Total LoM Operating Cost 167.38 5,417.44


Mining: The mining costs (Table 70) are broken down further since they contribute materially (i.e. ~14%) to the forecast steady state cost of USD167.38/t of ore reporting to the Plant for the period January 2020 to December 2028.

Table 70: Mining Costs (2019)

Description Cost

(USD/t ore)

Fleet Maintenance 11.71

Explosives 6.45

Cement 2.1

Backfill 1.41

Support 1.28

Total Mining Cost 22.95


MRM Function: The steady state annual cost of staffing the mineral resource management function amounts to about 3% of the direct steady state operating costs. Process: The steady state annual processing cost will require an average of 5% of the direct steady state operating budget.

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TSF: This expense consumes 4% of the budget forecast at full production. General and Administration (G&A): This expense collectively draws on 18% of the operating budget for the period January 2020 to December 2028 and includes numerous categories of which the largest are personnel transportation, Alliance Lowa ASBL, Bank Charges, Security, Medical and Clinics. Diesel: The annual power consumption in kilowatt hour was determined by multiplying the absorbed power requirements for each area by the anticipated annual operating hours. The annual cost of the power was then calculated by applying a unit power rate in USD/kWh to the price of fuel at USD1.18/ℓ, assuming on-site power generation using diesel gensets. This cost constitutes about 12% of the annual steady state requirement. Site Infrastructure: This amounts to USD14.78/t ore (~9%). Transport and Duties: A provision of 2% of the total annual steady state budget will cover these costs. Wages and Salaries: The steady state annual cost of staffing the mining and processing operation will amount to about 33% of the direct steady state budget. The anticipated indirect costs at steady state are shown in Table 71. Table 71: Indirect Steady State Off-mine Selling Costs

Description Total

(USD/t Sn Sold)

Logistics - Mine to Delivery Point 753.84

Logistics - Delivery Point to Smelter 482.51

Road Maintenance 128.61

Treatment Charges 1,325.59

Clearing Agent Fees 74.67

Export Taxes and Fees 377.54

Marketing Commission 342.41

Government Royalty 548.60

Total 4,033.67


The average steady state (from January 2020 to December 2028) operating unit costs are expected to be USD9,451/t Sn sold (i.e. USD5,417/t Sn sold and USD4,034/t Sn sold). Capex over the corresponding period is USD366/t Sn sold, giving and all in sustaining cost USD9,817/t Sn sold.

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22. ECONOMIC ANALYSIS Table 72 is used together with the capital and unit operating cost forecasts to examine the economics of the operation in terms of the pre-tax and prefunding steady state operating margins. Table 72: Average Annual Steady State Operating Margin

Rock Type Unit Total

Ore Mined kt 384

RoM Grade % 3.98

Development (ore and waste) m 2,500

Processing Recovery % 72

Payable Tin Sold t 11,004

Tin Price USD/t Sn Sold 17,000

Direct Steady State Unit Costs USD/t ore 167.38

Indirect Steady State Unit Costs USD/t Sn Sold 4,033.67

Gross Revenue USD M 187

On-mine Costs USD M 63

On-mine Margin USD M 124

Off-mine Costs USD M 44

Other Costs (including Mine closure provision) USD M 4

Pre-tax and Pre-funding Operating Margin USD M 76

Source: SMI

Ore Mined: The QP has examined the LoM production schedule which was generated using from the latest geological block model. Production is planned to reach an annual steady-state throughput of approximately 32ktpm or 384ktpa. The average grades over the LoM are 3.98%, which translates into about 11kt of tin sales every year after applying a processing recovery of 72%. RoM Grade: This is derived from the LoM plan in the context of the geological block model and applicable modifying factors. It represents the grade of the Mineral Reserves as stated herein. Development: The ore and waste development are forecast to still be relatively high during 2020 (i.e. 6,324m) but will average closer to 2,500m during steady state thereafter. Process Recovery: The Plant has been through periods of ‘teething problems but the recoveries have generally been increasing during the ramp up to steady state production with recent performance being in line with the forecasts from test work (i.e.72%). Payable Tin Sold: This is essentially the tin recovered after processing (Section 17). A concentrate grade of 62% is anticipated. Tin Price: This is the price that has been used for the business plan (Section 19). Direct Steady State Costs: These are as described in Table 69. Indirect Steady State Costs: These are as presented in Table 71. Other Costs: These costs are included to cover working capital and a provision closure and sustaining capital. The LoM environmental provision amounts to USD3.3 M. The QP has also confirmed the economic viability of the Mineral Reserves as stated herein by inspecting the operations latest Business Plan and the associated post tax and pre-financing discounted cashflow at a real 31 December 2019 steady state price of USD17,000/t Sn sold. The DCF returns a positive net present value over the full LoM at a real discount rate of 8%. The main uncertainties that may impact the cashflow forecasts are those related to price variations, the availability of the road to the mine site, forecast cost of fuel, logistics and salaries and wages. A sensitivity analysis show that the value is most impacted by revenue followed by operating costs and capital costs.

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23. ADJACENT PROPERTIES Alphamin still holds three of the original five permits which are adjacent or proximal to other prospecting and mining rights (Figure 59). Figure 59: Adjacent Properties (DRC Mining Cadastre Portal)

Source: http://drclicences.cami.cd/en/

The DRC’s online Mining Cadastre Portal reveals that the Mining rights to the south and west of the Alphamin PE 13155, are held by Societe Aurifere du Kivu et Maniema Sarl (SAKIMA), but it is unclear as to what stage of exploration or development these projects are at. There are numerous prospecting rights in the area that cover various minerals (e.g. Ag, Au, Sn, Nb, Ta, Cu, Co, Zn, Pb, diamonds, Fe, wolfram and monazite). These belong to other companies or individuals but Alphamin is unaware of the stage of exploration or development these properties.

http://drclicences.cami.cd/en/

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24. OTHER RELEVANT INFORMATION

24.1. Exploration

Bisie Ridge Works Programme The Bisie Ridge runs approximately north-south for over 16km, of which 13km falls within PE13155. The area is linked to high levels of localised residual magnetic intensity in association with geochemistry anomalies. An exploration programme is planned that entails establishing a 10km access track along the ridge from Mpama South, employing and mobilising a dedicated exploration team under ABM and conducting reconnaissance exploration activities such as mapping, pitting, trenching and further soil geochemistry. This reconnaissance work is anticipated to lead to further exploratory work including target drilling. Table 73 encompasses the high-level activities, costs and timing of the work programme.

Table 73: Proposed Activities, Timing and Costs of Mpama Ridge Exploration

Activity F2020

USD’000 F2021

USD’000 F2022

USD’000

Establish Phase 2 and Phase 3 Jeep Track (10.4km) 208

Exploration Team Employment Costs 240 240 240

Excavate Test Pits along Jeep Track 120 120 120

Soil Sample Analysis 25 25 25

Exploration Drilling at Identified Sites

540 700

Develop Conceptual Resource Model

25 50

Total Provision 593 950 1,135


Mpama South Works Programme Although ABM’s previous exploration successfully intersected the deposits from surface drilling at Mpama North and South, it was not without great effort with numerous lessons learnt. As a consequence of this, and given the easier access to equipment, power, water roads and skills, the proposed exploration programme for Mpama South has been planned very differently. Mpama South infill drilling will be conducted from an underground drilling drive excavated in the footwall of the interpreted Mpama South deposit. This approach leverages off the available equipment, expertise and existing infrastructure on site. It also avoids the additional costs and delays associated with environmental challenges and the logistics of having to clear forest areas for drill pads and roads or having to use a helicopter and specialised small rigs to move between areas. The programme entails the construction of an access road to the western side of Mpama South over an already environmentally disturbed area, before establishing a small box-cut and a 50m long mine portal in the ‘centre of gravity’ of the anticipated Mpama South deposit. A 25m drilling drive will then be developed from which underground fan drilling can occur. About 6km of drilling using thirty drillholes is anticipated to be sufficient for a good understanding of the size, shape and variability of the deposit and for the declaration of Mineral Resources. Based on the results of this drilling, subsequent studies will be undertaken to assess the potential planning, scheduling and constructing further operations at Mpama South. These activities include the immediate access and mining of early ore from Mpama South by March 2021 and will be achieved by an anticipated capital spend of some USD10 M from cashflows. Figure 60 and Table 74 encompass the high-level activities, costs and timing of this work programme.

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Table 74: Proposed Activities, Timing and Costs of Mpama South Exploration

Activity F2020

USD’000 F2021

USD’000

Establish Box-cut (25,000m3) 255

Develop Mine Portal (50m) 175

Develop Drill Drive (25m) 50

Drill 30 Holes (6,000m) 1,650

Develop Resource Model 25

Prepare Conceptual Mine Plan and Schedule 50

Establish Initial Operational Levels for further Drilling/Works 250 8,000*

Total Provision 2,455 8,000

Source: ABM *Note: This provision of USD8 M is supported by a budget which is to be confirmed.

Figure 60: Plan View showing Tin Grade Distribution from Historical Drilling

Source: Minex Consulting 2019 Note: red > 0.5% Sn, purple > 1.0% Sn

24.2. Security ABM’s security department work closely with various DRC State security structures, the United Nations, and with security personnel from other mining companies or NGOs in the area, to manage the security threat related to a number of opportunistic armed groupings that operate in logistically challenging areas, where the National Security Forces find it difficult to respond.

• The strategy at ABM is to continue to manage this risk in a collective and proportionate manner. Suitable internal and internationally accepted standards and guidelines exist, namely;

• the ‘ABM Corporate Security Framework’;

• the ‘ABM Mine Site Security Procedures’;

• the ‘Voluntary Principles on Security and Human Rights’;

• the ‘ASIS International Standards and Guidelines’;

• the United Nations (UN) standards and guidelines on the use of force, firearms; and

• the UN standards and guidelines for law enforcement officers and security practitioners. ABM has not reported any serious ABM related security threats since the previous NI43-101 report of March 2017.

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24.3. Risks Appendix 1 contains a summary of the more material risks that could impact the Mineral Resource and Mineral Reserves as stated herein. The following points reflect the view of the independent QP:

• A new mining method is being implemented which is planned to enhance the safety, flexibility and productivity of the operation. Strict mining control and discipline will be required to ensure planned production targets are achieved;

• Bisie Mine has significant upside given the ongoing exploration programmes and the remaining exploration potential. The planned 2020 exploration programme should be progressed timeously so that the anticipated upside potential can be included in future mine planning to facilitate additional flexibility and the overall robustness of the business;

• Ongoing infill exploration drilling. This has the potential to materially enhance the flexibility in the mining plan and the LoM plan;

• The quality of the RoM ore sent for processing could vary materially. Introduce stringent proactive quality and grade control protocols into the mining operations early to ensure the quality of the RoM sent for processing;

• Logistical challenges associated with operating in a remote area. AMB should proactively monitor and manage the condition of the access roads to site;

• Underground trucking of rock or transporting of material could impact efficiencies and result in reduced Plant throughput. Operational personnel must pay attention to underground road maintenance with particular emphasis on keeping them free of rocks and water;

• Logistical arrangements for the placing of backfill need to be carefully planned to prevent delays which would also negatively impact the Plant’s throughput;

• Maintaining good relations with the communities is always a priority;

• Significant infrastructure has been installed on site and production has commenced. Some 172kt of ore was mined by end December 2019, at a grade of 5.30% Sn;

• The fluctuating tin price can have either a positive or a negative impact on profitability;

• Sufficient dilution has been included in the mine plan and strict control in identifying ore and waste could enhance head feed grades; and

• Continued care in managing and maintaining social license to operate.

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25. INTERPRETATION AND CONCLUSIONS Mineral Resources of 4.80Mt at 4.60% (at a cut-off of 0.5% Sn) and Mineral Reserves of 3.33Mt and 4.67% (at a price of USD17,000/t Sn and cut-off of 1.6% Sn) are stated as at 30 June 2019 and 31 December 2019 respectively. These estimations are underpinned by a mineral resource model and 2019 LoM and business plan that have been developed to acceptable standards. The mine designs and layouts, dilution and economic parameters applied to the mineral resource estimates have been prepared to industry standard practices and economic viability of the Mineral Reserves has been clearly demonstrated to the satisfaction of the QP. The Mineral Resources and Mineral Reserves have been estimated using The Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Best Practice Guidelines (2003) and are reported in accordance with the 2014 CIM Definition Standards, which have been incorporated by reference into National Instrument 43-101 - Standards of Disclosure for Mineral Projects (NI 43-101). In addition to onsite inspections of the exploration and mining operations, the QPs have reviewed all of the available data and information related to the Mineral Resource and Mineral Reserve estimates and are of the opinion that:

• The mine has been constructed, ramped up and is operating successfully since the previous March

2017 NI43-101 Technical report. The revised mining method, mine design and LoM plan is logical and

pragmatic. The Mine has transitioned to owner operated mining and has progressed steadily beyond

its first-year of operation. This performance is expected to continue during 2020 as it continues to ramp

up to an enhanced steady state of approximately 32ktpm in January 2021, as forecast by the latest

business plans. Full production is then planned to continue to December 2029 after which Plant

throughput is forecast to reduce, notwithstanding additional production from the Mpama North

extensions and/or Mpama South.

• The Bisie Ridge upon which the Mine is located is under explored and ABM has aggressive exploration

plans in place to bring additional exploration targets to account, including the Mpama South deposit.

The QP notes that this is likely to enhance the steady state life of the operation well beyond 2029.

• Mine infrastructure, which includes access to site, an airstrip, a concentrator, underground mining

infrastructure, a diesel power plant with fuel storage capacity, accommodation, sewage and treatment

facilities and on-site security, has been established and is performing according to industry norms. The

mining equipment fleet will need to be expanded to service the increased annual production forecast.

• The concentrator performance and ramp up has proven successful in delivering saleable product at

designed levels. It makes use of proven processing methodologies and the forecast recoveries of 72%

in the latest LoM and business planning is supported by metallurgical test work.

• The economic analysis is not a valuation of the mining operation in terms of any of the international

valuation codes. Its purpose was to assess the robustness of the margins to confirm economic viability

of the Mineral Reserve estimate. Healthy and positive steady state operating margins are forecast and

this demonstrates that the Mineral Reserves are economically viable based on the new mining method,

LoM schedule, forecast operating costs, environmental and other provisions, and at the assumed price

of USD17,000/t Sn;

• The necessary security of tenure exists with all permits required for ongoing exploration and

exploitation in place;

• The ongoing operational activities are environmentally sound, utilising simple and proven management

plans. All environmental permits are in place and ABM is reportedly in the process of generating an

amended EIS, MRP, and EMPP for submission to the DPEM in Kinshasa. The mine closure plans and

costs were considered and adequate provision has been made by ABM in its business planning to

meet these obligations. These are reviewed and revised annually in line with international best practice.

• The mine is a significant employer of local Congolese and contributes to the growth of both the local

and regional economies. Stakeholder engagement is ongoing; through the Alliance Lowa ASBL which

involves regular meetings with the community.

• Cognisance needs to be taken of the various risks identified. Apart from the concerns related to

security, all of the risks to the Mineral Reserve are not dissimilar to any other mine operating in a remote

location with seasonal weather patterns and long logistical supply lines. The mitigation actions to

minimise the risks have been identified and are likely to be implemented. The QP's overall assessment

from a risk point of view is that the level of residual uncertainty constitutes a medium level of risk, which

suggests that they are manageable.

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26. RECOMMENDATIONS The following recommendations are made:

• The ongoing infill exploration drilling and planned 2020 exploration programme should continue to be

prioritised so that flexibility is afforded to future mine planning to ensure a consistent supply of RoM

feed to the Plant. This will also enable ABM to take timely advantage of any increase in plant throughput

capacity, as anticipated. SMI notes that ABM has catered sufficiently for ongoing infill drilling, MRM

skills enhancement, road and bridge monitoring, mining control, and community engagements in the

LoM operating budget. These are all part of the ongoing nature of the producing asset. USD2.46million

has been included in the LoM budget for the 2020 exploration program. The subsequent phase in 2021,

which is to “Establish Initial Operational Levels for further Drilling/Works” is contingent on the success

of the first phase (Table 74);

• ABM should enhance the skills in the MRM function to improve the quality of the tactical planning and

for better grade control to address the variability in RoM fed to the Plant;

• AMB should continue to proactively monitor and manage the condition of the access road and bridges

to site to limit the impact on supply lines and final product delivery;

• Strict mining control and discipline will be required to ensure that the planned production targets from

the new mining method are achieved. Ground conditions need to be monitored closely prior to, and

after, the placing of backfill. The logistical arrangements for the placing of backfill will also need to be

carefully planned to prevent operational delays. The operational personnel must also maintain and

keep the underground roads free of rocks and water;

• The possibility of not needing to fill the open stopes with waste in the narrower (<4m) portions of the

orebody should be evaluated over the coming year (2020); and,

• ABM needs to continue to nurture good working relationships with the local communities and authorities

to eliminate or at least minimise threats of production interruptions.

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27. REFERENCES

Author/Company Date Title

AlphaMin Bisie Mine 2018 ABM Environmental and Social Monitoring Procedure. January 2018.

AlphaMin Bisie Mine 2017 ABM Water Management Plan. December 2017

AlphaMin Bisie Mine 2018 ABM Water Monitoring Procedure. February 2018

AlphaMin Report 2013 Bisie Tin Project Summary, February 2013

Alphamin Resources Corp. 2019 Bisie DCF Model

Alphamin Resources Corp. 2019 Bisie Owner Mining Cost for Cut & Fill

Alphamin Resources Corp. 2016 Bisie Tin Project: NI 43-101 Technical Report – 19 February 2016 Feasibility Study Report

Alphamin Resources Corp. 2017 Bisie Tin Project: NI 43-101 Technical Report – 23 March 2017 Updated Feasibility Study and Control Budget Estimate Report

Alphamin Resources Corp. 2019 Bisie Tin Project: Rock Engineering Constraints

Alphamin Resources Corp. 2019 Mineral Resource Statement for the Bisie Tin Project

Alphamin Resources Corp. 2019 Mpama South 2020 Budget Model

Bahori Consulting 2019 NI 43 101_Report_Env_draft final

Bara Consulting (Pty) Ltd 2016 Bisie Tin Project Definitive Feasibility Study: Mining Study

Bara Consulting (Pty) Ltd 2016 Bisie Tin Project: Mine Design Criteria

Bara Consulting (Pty) Ltd 2016 Bisie Tin Project: Mining Method Trade-Off Scoping Study

Bara Consulting (Pty) Ltd 2016 Geotechnical Design Parameters for The Bisie Tin Project Alphamin Resources Corporation, North Kivu Province Democratic Republic of Congo

Barton et al. 1974 Guidelines for Primary Support Design (Based on Rock Mass Classification)

Brieter, K 2007 Mineralogical Composition of Tin-Bearing Ore Samples from Bisie and Kalimbi, D.R.C.

Epoch Resources (Pty) Ltd 2018 Bisie Tin Detail Design Tailings Storage Facility. Stormwater Diversion Layout, Detail and Setting Out Data. Reference: 127-059-004. July 2018

Epoch Resources (Pty) Ltd 2018 Tin Detail Design Tailings Storage Facility - Phase 1 Access Road and River Diversion Layout, Details and Setting Out Data. Reference: 127-059-023. July 2018

Epoch Resources (Pty) Ltd 2018 Tin Project Residue Disposal Facilities. Detailed Design Report. December 2018.

Garret, N 2008 Artisanal Cassiterite Mining and Trade in North Kivu – Implications for Poverty Reduction and Security

Hubert André-Dumont 2013 Mining in 31 Jurisdictions Worldwide

Industrial Technology Research Institute

2016 2016 Report on Global Tin Resources and Reserves

InRoads Consulting 2016 Bisie Tin Project. Definitive Feasibility Study Report on a Geotechnical Investigation for The Tailings Disposal Facility and Process Plant. Detailed Design Report. January2016.

Jackson, M 2015 Mpama North Structure and Hanging wall Beam Modelling, Bisie, DRC

Latona Consulting (Pty) Ltd 2019 Further Rock Engineering Recommendations for Cut and Fill Mining

Latona Consulting (Pty) Ltd 2018 Rock Engineering Recommendations for Cut and Fill Mining

Latona Consulting (Pty) Ltd 2018 Updated Geotechnical Assessment and Recommendations for Bisie Tin SLC

MEG Mine Search 2013 Project Profile Report

MPC Report 2008 Review of Exploration Geological Work carried out at Bisie – Walikale, Kivu North, Eastern Democratic Republic of the Congo

Pearl, L M 2011 Report on the Bisie Tin Project (in compliance with NI 43-101) – Concessions: PR5266, PR5267, PR5270 and PR4246, Walikale Territory, North Kivu, Democratic Republic of the Congo for Alphamin Resources Corp.

The MSA Group 2019 J4086 Alphamin Bisie Tin Project NI 43-101 MRE June 2019-MSA Sections Final

The MSA Group 2013 Alphamin Resources Corporation, Bisie Tin Project, North Kivu Province, Democratic Republic of Congo, NI 43-101 Technical Report – Mineral Resource Estimate

Wimmer, S Z and Hilgert, F 2011 Bisie. A One-Year Snapshot of the DRC’s Principal Cassiterite Mine.

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28. DATE AND SIGNATURE PAGE This report titled ‘NI 43-101 Technical Report – 31 December 2019’ with an effective date of 31 December 2019 prepared by Sound Mining International (SMI) on behalf of Alphamin Resources Corp. (Alphamin), was prepared and signed by the following author:

Vaughn Glenn Duke Pr.Eng., PMP, B.Sc. Mining Engineering (Hons.), MBA, FSAIMM Director and Principal Consultant Sound Mining International Dated: 11 February 2020 at Johannesburg, South Africa. Effective Date of this Report: 31 December 2019.

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CERTIFICATE OF AUTHOR

This Certificate of Author has been prepared to meet the requirements of National Instrument 43-101 Standards of Disclosure for Mineral Projects, Part 8.1. I, Jeremy Charles Witley Pr. Sci. Nat. do hereby certify that:

1. I am Principal Mineral Resource Consultant of: The MSA Group (Pty) Ltd Henley House, Greenacres Office Park Victory Road Victory Park, Gauteng, South Africa, 2195

2. This certificate applies to the technical report titled “NI 43-101 Technical Report – 31 December 2019, Bisie Tin Mine, North Kivu Province, Democratic Republic of Congo, prepared for Alphamin Resources Corp.”, that has an effective date of 31 December 2019 and a report date of 31 December 2019 (the Technical Report).

3. I graduated with a degree in Mining Geology from The University of Leicester, UK in 1988. In addition, I have obtained a MSc (Eng.) from the University of Witwatersrand, South Africa in 2015.

4. I am registered with The South African Council for Natural Scientific Professions (SACNASP) and am a Fellow of the Geological Society of South Africa.

5. I have worked as a geologist for a total of 31 years. I have worked in a number of roles, including senior management, in mine geology, exploration projects and Mineral Resource management. I have conducted Mineral Resource estimates, audits and reviews for a wide range of commodities and styles of mineralization, including vein style deposits.

6. I have read the definition of “qualified person” set out in National Instrument 43-101 (NI 43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

7. I visited the Bisie Property during the period 18 – 20 July 2013, 20 – 22 May 2014 and 11 - 13 August 2015. 8. I am responsible for the preparation of items 1.3, 1.4, 1.5, 6 to 12, and 14, and co-responsible for the preparation of

items 1.12, 1.13, 2, 24.1, 25 and 27 of the Technical Report. 9. I have not had prior involvement with the property that is the subject of the Technical Report. 10. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that

is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading. 11. I am independent of the issuer according to the definition of independence described in section 1.5 of National

Instrument 43-101. 12. I have read National Instrument 43-101 and Form 43-101F1 and, as of the date of this certificate, to the best of my

knowledge, information and belief, those portions of the Technical Report for which I am responsible have been prepared in compliance with that instrument and form.

13. I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated 11 February 2020

/s/ Jeremy Witley ___________________________ Signature of Qualified Person Jeremy Charles Witley Principal Mineral Resource Consultant The MSA Group (Pty) Ltd

140

CERTIFICATE OF AUTHOR This Certificate of Author has been prepared to meet the requirements of National Instrument 43-101 Standards of Disclosure for Mineral Projects, Part 8.1. I, V.G. Duke do hereby certify that: 1. I am an employee of:

Sound Mining International (Pty) Ltd., 2A 5th Avenue, Rosebank, Gauteng, South Africa.

2. This certificate applies to NI 43-101 Technical Report, dated 31 December 2019, with an effective date of 31 December 2019 (the Technical Report).

3. I graduated with an Honours Degree in Mining Engineering from the University of the Witwatersrand in South Africa. 4. I am a Fellow of the South African Institute of Mining and Metallurgy. 5. I have worked as a mining engineer for a total of 33 years. I have worked in a number of roles, including mining

operations, mining projects and mining consulting. I have conducted mining reviews for a wide range of commodities, mining methods and styles of mineralisation.

6. I have read the definition of ‘Qualified Person’ set out in National Instrument 43-101 (NI 43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a ‘Qualified Person’ for the purposes of NI 43-101.

7. I have visited the Bisie Property from the 13 to 15 May 2019. 8. I am responsible for the preparation of items 1.1, 1.2, 1.6 to 1.13, 2 to 5, 13, and 15 to 27 inclusive, of this technical

report. 9. I have not had prior involvement with the property that is the subject of the Technical Report. 10. I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that

is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading. 11. I am independent of the issuer according to the definition of independence described in section 1.5 of National

Instrument 43-101. 12. I have read National Instrument 43-101 and Form 43-101F1 and, as of the date of this certificate, to the best of my

knowledge, information and belief, those portions of the Technical Report for which I am responsible have been prepared in compliance with that instrument and form.

13. I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated 11 February 2020

__________________________ Signature of Qualified Person Vaughn Glenn Duke Pr.Eng., PMP Principal Consultant Sound Mining International (Pty) Ltd DIRECTORATE SOUND MINING HOUSE

P POTGIETER, B. Ing. Mynbou, MSAIMM 2A Fifth Ave Rivonia 2128

D VAN BUREN, Pr.Sci.Nat., B.Sc. (Hons.) Geology, MGSSA, FFF PO Box 97194 Petervale 2151

M LOTRIET, Pr.Eng., B.Sc. Min. Eng. (Hons.), PMP, FSAIMM Tel No : +27 (0) 11 234 7152

G P STRIPP, B.Sc. Min. Eng. (Hons.), M.Sc. Ph.D., FSAIMM Reg No : 2002/002265/07

V G DUKE, Pr.Eng, PMP, B.Sc. Min. Eng. (Hons.), MBA, FSAIMM Website : www.soundmining.co.za

CONSULTANTS TO THE MINING INDUSTRY

http://www.soundmining.co.za/

141

29. APPENDIX 1: THREATS AND OPPORTUNITIES

Threats/Opportunities Impact Rating Comments

Planned production not achieved due to hauling constraints A shortfall in revenue and project cash flow. Threat to revenue and cost.

M Additional hauling capacity can be sourced. Waste hauling will be reduced by using waste as backfill underground. The complement of trackless equipment underground is relatively small which will minimise congestion. Roadways and hauling routes to be maintained to high standards.

Insufficient flexibility in the mining plan could affect the production rate

Negatively impact revenue and add to costs. H Sufficient time has been allowed in the mining plan for construction and development activities. Increasing development rates by introducing a third shift can be implemented if required.

Additional ore sources and mineral resource/reserves being brought into production from Mpama extensions and Mpama South

Production increases and life of mine extension improving revenue generation and lowering unit costs through increased volume throughput.

H Mpama South has a detailed exploration and development plan commenced in Q4 2019 and expected to bring early additional ore into production by Q1 2021, Mpama extension exploration programme can also be designed using underground development access or surface drilling.

Mine Call Factor in mining to date (YE2019) has repeatedly shown that mined grades are often higher than modelled grades

Increased revenues to forecast and profitability. M This effect could be a function of the resource modelling or the performance of mining. Although a producing asset, a longer production history will be required to adequately reconcile and understand this positive deviation from forecasts and plans.

New mining method performance (open stoping with hydraulic backfill)

Slow ramp up to full production thereby negatively affecting revenue generation.

H Implement detailed planning schedules and close supervision to ensure strict mining discipline.

New technology (trackless mining equipment) and downtime of trackless equipment

Planned production not achieved. M On the job training of operators and maintenance personnel with assistance from OEMs.

Intersecting old abandoned artisanal mining excavations Localised delays in production. M Cover drilling to check for water and old excavations simultaneously.

Backfill delays affecting production schedule Planned production not achieved. M Backfill logistics to be planned in detail with empty trucks being filled with waste on the return trip to fetch ore.

Poor quality access road and bridge collapses will impact the safety and efficiency of supply vehicles travelling to site as well as concentrate leaving site

Repairs to road will require higher expenditure and impact the cost as well as cashflow.

H Proactively monitor and manage the condition of the road over the Kumbel Pass, particularly during the winter season.

Force majeure and threats of closure/delays from surrounding communities

Production delays or curtailment. M Maintain good relations with communities and local authorities.

Fluctuating tin price Directly affects revenue. M This is both a threat and an opportunity.

High levels of company debt Increased interest costs and cashflow implications from repayments.

M Closely monitor company debt ratio and seek to repay expensive debt quickly and/or restructure debt.

High variability in quality of RoM ore sent to plant Reduced recovery in the process plant with lower revenue.

H Grade control skills with suitably trained and qualified personnel to monitor RoM ore quality.

Mining and processing operations have already commenced Revenue stream commenced with Sn sales. M Demonstrates commitment and progress of the project. Ramp up to steady state to be monitored strictly against the mining plan.

Additional exploration on strike likely to result in an increased Mineral Resource

Opportunity for mine life extension and production flexibility.

H The planned 2020 exploration programme should be progressed timeously so that the anticipated upside potential can be included in future mine planning to facilitate additional flexibility and the overall robustness of the business.

Ongoing in-fill and downdip exploration drilling Opportunity for mine life extension and additional flexibility in future mine planning.

H Additional information from exploration should be incorporated into the mine planning.

Dilution could be higher (>20%) or lower (<20%) depending on the orebody width and degree of grade control

Opportunity for cost saving, improved head feed grades and hence revenue.

H Identification of ore vs waste by means of continuous exploration and sampling, is likely to result in RoM ore that meets or exceeds the planned grade.

Increased tin (Sn) recovery Increased revenue and profitability. H Careful monitoring and management of RoM ore to ensure minimum inclusion of undesirable materials, particularly those that increase the consumption of reagents. Process plant upgrade and inclusion of a fine tin recovery circuit at nominal capital cost. Process Plant debottlenecking and operational improvement.

Source: SMI

142

30. APPENDIX 2: CHECKLIST OF ASSESSMENT AND REPORTING CRITERIA

Compliance Remarks

Drilling Techniques

All drillholes were diamond drill cored and drilled from surface (mostly NQ) at angles of between -60° and -75°. The drillholes were drilled from east to west along section lines spaced between approximately 25m and 50m apart. 21 PQ sized holes from a metallurgical drilling campaign were included that were drilled in three clusters approximately 25m apart.

Drill Sample Recovery

All of the drillholes were geologically logged by qualified geologists. The logging is of an appropriate standard for grade estimation.

Drill sample recovery

Core recovery in the mineralised zones was observed to be very good and is on average greater than 95%. Five of the shallow drillholes intersected artisanal workings and so recovery of the high-grade mineralisation was poor and therefore the data from these holes were not used for grade estimation.

Sampling methods

Half core samples were collected continuously through the mineralised zones after being cut longitudinally in half using a diamond saw. Drillhole samples were taken at nominal 1m intervals, which were adjusted to smaller intervals in order to target the vein zones. Lithological contacts were honoured during the sampling. The QP’s observations indicated that the routine sampling was performed to a reasonable standard and is suitable for evaluation purposes.

Quality of assay data and laboratory tests

The assays were conducted at ALS Chemex in Johannesburg where samples were analysed for tin using fused disc ME-XRF05 conducted on a pressed pellet with 10% precision and an upper limit of 10,000ppm. This was reduced to 5,000ppm from 2014 onwards. Over limit samples were sent to Vancouver for ME-XRF10 which uses a Lithium Borate 50:50 flux with an upper detection limit of 60% and precision of 5%. ME-ICP61, HF, HNO3, HCLO4 and HCL leach with ICP-AES finish was used for 33 elements including base metals. ME-OG62, a four-acid digestion, was used on ore grade samples for Pb, Zn, Cu and Ag. External quality assurance of the laboratory assays for the Alphamin samples was monitored. Blank samples, certified reference materials and duplicate samples were inserted with the field samples accounting for approximately 10% of the total sample set. The QA/QC measures used by Alphamin revealed the following:

• The high-grade CRM (31.42% Sn) assays by ALS prior to 2015 returned values approximately 8% higher than the certified mean value. 98 pulp rejects from this period of between 1.5% and 60% Sn were re-assayed by ALS in 2016 together with the high-grade CRM. The 2016 assays correlated well with those prior to 2015 and the high-grade CRM returned values within tolerance. Therefore, the pre-2015 assays were accepted for estimation without modification.

• The lower grade CRM assays (<2% Sn) indicated that the Sn and Cu assays were accurate and unbiased, consistently returning values within two standard deviations of the accepted CRM value.

• The field duplicates confirmed the nuggetty nature of the tin mineralisation. The majority of the duplicate assays were within 20% of the field sample.

• Blank samples indicated that no significant contamination occurred.

Verification of sampling and assaying

A selection of cores representative of the entire drilling programme at Mpama North have been visually verified during three site visits by the QP (July 2013, May 2014 and August 2015). The QP observed the mineralisation in the cores and compared it with the assay results. It was found that the assays generally agreed with the observations made on the core. The QP took ten quarter core field duplicates for independent check assay in 2013, which confirmed the original sample assays within reasonable limits for this style of mineralisation 150 pulp duplicates were sent to SGS (Johannesburg) in 2013 for confirmation assay and a further 173 were assayed in 2015. In 2015, 99 pulp duplicates were sent to Set point (Johannesburg) for confirmation assays.

• The pulp duplicates assayed by SGS in 2013 showed excellent correlation with the ALS assays at both high- and low-grade ranges.

• SGS assays were lower than ALS for grades above 20% for the 2014 data checked in 2015. SGS under-reported the grade of all the CRMs that were inserted. The high-grade CRM was under assayed by approximately 5%.

• SGS assays in 2016 in the < 5% Sn grade range confirmed the ALS assays. The SGS tin assays were significantly lower for the higher grades, however the titration method used by SGS results in a significantly low-grade bias compared with the CRM

Set point assays were lower than ALS for grades above 10% for the 2014 data checked in 2015. Set Point tended to under-report the grades of the CRMs.

143

Compliance Remarks

Location of data points

All except five of the Bisie surface drillhole collars used in the Mineral Resource estimate were surveyed by digital GPS. For those that were not surveyed, the hand-held GPS readings were used with the elevation being corrected to that of the LIDAR topographic survey. Down-hole surveys were completed for all of the holes drilled at Mpama North.

Tonnage factors (in situ bulk densities)

SG determinations were made for 2,698 drillhole samples using laboratory gas pycnometer. A regression formula of tin grade against SG was developed that was applied to the samples that did not have direct SG measurements. The assigned SG was interpolated into the block model using ordinary kriging. The laboratory pycnometry readings compared well with a number of SG measurements completed using the Archimedes principle of weight in air versus weight in water.

Data density and distribution

The holes were drilled from east to west along section lines spaced approximately 50m to 60m apart with infill drilling on 25m to 30m spaced sections in a portion of the shallower area. Along the section lines, the drillholes intersected the mineralisation between approximately 25m and 50m apart in most of the Mineral Resource area. The 21 PQ sized holes from a metallurgical drilling campaign were included that were drilled in three clusters approximately 25m apart. Within the clusters, the PQ holes were drilled approximately 5m apart. In the Mineral Resource area, 122 NQ drillholes were used for the grade estimate. A number of holes intersected mineralisation outside of the area currently defined as a Mineral Resource and five of the shallow drillholes intersected artisanal workings. The data from these holes were not used for grade estimation.

Database integrity

Data are stored in an Access database. The QP completed spot checks on the database and is confident that the Alphamin database is an accurate representation of the original data collected.

Database integrity

The area defined as a Mineral Resource extends approximately 700 m in the down-plunge direction. It extends for approximately 300 m in the plane of mineralisation perpendicular to the plunge. The main zone of the Mineral Resource, which accounts for 97% of the Mineral Resource, is on average approximately 9 m thick, although is narrower (less than 1 m) at the margins and up to 20 m thick in the central areas. The zones that occur several metres above and below the main zone are considerably narrower than the main zone and cover areas of between 100 m and 200 m in the dip and strike directions

Geological interpretation

The mineralised intersections in drill core are clearly discernible. The Mineral Resource is interpreted to occur as irregular tabular mineralised zones, dipping 65° to the east, containing several narrow veins, blocks and disseminations of cassiterite. The mineralised zones are hosted in chlorite schist that is the result of intense alteration and may originally have been a distinct stratigraphic interval. The main zone of the Mineral Resource is almost continuous for over 650 m although it has been affected by a number of faults causing local displacement. Several faults with throws in excess of 10 m have been modelled.

• The Main Vein mineralisation consists of a number of uncorrelated cassiterite veins within pervasively chloritised schist. This zone generally occurs over thicknesses of between 2m and 22m with an average thickness of approximately 9m. The Main Vein zone is generally the highest grade and most consistent overall.

• Hanging Wall Vein mineralisation occurs within partly chloritised schist and micaceous schist between 4m and 20m above the Main Vein. This zone of mineralisation is generally between 0.5m and 4m wide and occurs in the central area of the deposit and tapers out northwards. The middling between the Hanging Wall Vein and the Main Vein decreases in areas and it is possible that this vein merges into the Main Vein in some parts of the deposit.

• Footwall Vein (FW Vein) mineralisation occurs within the micaceous schist and amphibolite schist between 2m and 12m below the Main Vein. This zone is restricted to the southern areas, is very narrow (<50cm) and high-grade in its most northern occurrences but thickens to the south to several metres. It is possible that this vein merges into the Main Vein in the some parts of the deposit.

A three-dimensional wireframe model was created for the three zones of mineralisation based on a grade threshold of 0.35% Sn. The Main Vein zone is the most consistent zone and occurs within a persistent chlorite schist. Narrower less continuous zones occur above and below the main zone within chlorite-mica schists.

144

Compliance Remarks

Domains

The mineralisation was modelled as three tabular zones containing irregular vein style mineralisation. A hard boundary was used to select data for estimation in order to honour the sharp nature of vein boundaries.

Compositing

Sample lengths were composited to 1 m. Composites of less than 1 m occurred in the narrow vein areas, which were retained. Accumulations of Sn%-density-composite length were calculated for grade estimation so that narrow extremely high-grade composites did not excessively influence the estimate.

Statistics and variography

Two populations of Sn mineralisation occur, a high-grade population of cassiterite veins and a lower grade population containing disseminated cassiterite as vein fragments and blebs. The data were separated into the two statistical populations, which resulted in the coefficient of variation for the Sn accumulation composites in the high-grade population being 0.5 and for the lower grade population being 1.6. The histograms are positively skewed. Normal scores variograms were calculated in the plane of the mineralisation, down-hole and across strike. Variogram ranges for the Sn accumulation in the main zone were modelled with ranges in the order of 75m in the longest direction of continuity and 60m in the second direction. Reliable variograms could not be produced for both the Hanging Wall or Footwall zones, and the Main Vein zone variogram was used to estimate these areas.

Top or bottom cuts for grades

Top cuts were applied to outlier values that were above breaks in the cumulative probability plot for metals other than Sn. The high-grade Sn values occur as a statistically distinct population with a low coefficient of variation and no top-cuts were considered necessary, the high-grade distribution being estimated separately, with a restricted search.

Data clustering

21 PQ sized holes from a metallurgical drilling campaign were included that were drilled in three close clusters approximately 25m apart. Within the clusters the PQ holes were drilled approximately 5 m apart. Outside of the metallurgical sampling area the grid is approximately regular.

Block size

20 mN by 2 mE by 10 mRL three-dimensional block models were used. The blocks were divided into sub-cells to better represent the interpreted mineralisation extents. The blocks were rotated into the plane of mineralisation prior to estimation.

Grade estimation

The accumulation of tin grade, density and composite length were estimated using ordinary kriging. Copper, lead, zinc, silver, arsenic and sulphur grades were estimated directly. The Sn%-density-composite length accumulations were divided into a high-grade population (>80%t/m) and a lower grade population (<80%t/m). The probability of a block containing values above and below this threshold was estimated by indicator kriging. Outside of the indicator variogram range, estimates did not use the extreme high-grades (>80 %tm) in order to reduce the influence of these values on estimates further away from them. The high- and low-grade populations were estimated separately using ordinary kriging and the block model grade was then assigned based on the estimated grade of the high and low grade and their proportion in each block. A minimum number of four and a maximum of ten one metre composites were required for the high-grade Sn-accumulation population. A minimum number of eight and a maximum of 24 one-metre composites were required for the lower grade Sn-accumulation population and other variables. Search distances and orientations were aligned with the variogram range and mineralised trends. Estimates were extrapolated for a maximum distance of 20m up- or down-plunge from the nearest drillhole intersection. Extrapolation is minimal over most of the Mineral Resource as the up-and down dip limits have been well defined by the drilling.

Resource classification

Measured Mineral Resources were declared where the drillhole spacing is approximately 25m and where the geological model has low variability. The mineralisation was classified as Indicated Mineral Resources if block estimates occur within the 50m drilling grid, so that all Indicated estimates are informed by samples within the variogram range. The remainder of the interpreted model within the sparser drilled area was classified as Inferred Mineral Resources with a maximum extrapolation from a drillhole of 20m along plunge. The up-plunge extremity (S1 Block) is separated from the main area by a fault and the structural interpretation in this area is tenuous and it does not contain sufficient data to classify them as Indicated Mineral Resources. Consequently, this area was classified as Inferred Mineral Resources. The high-grade mineralisation of reasonable tonnage leads no doubts as to reasonable potential for economic extraction, it being one of the highest-grade undeveloped tin deposits in the world.

145

Compliance Remarks

Mining cuts

The thickness of the mineralisation was honoured in the estimate and as a result some areas will be more sensitive to dilution than others. The thickness, grade and steep dip implies that the Mineral Resource can be extracted using established underground mining methods.

Metallurgical factors or assumptions

The tin mineralisation occurs as cassiterite, an oxide of tin (SnO2). The Cu, Zn and Pb mineralisation occurs as sulphides. Each of these minerals is amenable to standard processing techniques for each metal. The mineralisation contains an average of 464ppm arsenic which is unlikely to have a material deleterious effect.

Legal aspects and tenure

Alphamin through its wholly owned DRC subsidiary, Alphamin Mining Bisie SA, has a Mining License PE 13155 which includes the Bisie Tin Mine. Alphamin has an 80.75% interest in ABM. The Government of the Democratic Republic of Congo (GDRC) has a non-dilutive, 5% share in ABM.

Audits, reviews and site inspection

The following review work was completed by the QP:

• Inspection of approximately 25% of the Alphamin cores used in the Mineral Resource estimate

• Database spot check

• Inspection of drill sites

• Independent check sampling

146

31. APPENDIX 3: ABBREVIATIONS, SI UNITS AND GLOSSARY OF TERMS

Term Explanation

% Percentage

%t/m Proportion of tonnes per metre

~ Approximately

‘ minute arc

“ second arc

° degree

°C degree centigrade

3D three dimensional

a Annum

ABM Alphamin Bisie Mining South Africa

Ag Silver

Alphamin Alphamin Resources Corp.

ALS ALS Chemex South Africa

AMD Acid Mine Drainage

As Arsenic

ASM Artisanal Small-scale Miner

Bahori Bahori Consulting (Pty) Ltd

BCS-CRM Bureau of Analysed Samples Ltd of Great Britain

BMA Bulk Modal Analysis

BOCO The Base of Complete Oxidation

CBSH Carbonaceous Shale

Ce Cerium

CIM The Canadian Institute of Mining, Metallurgy and Petroleum

CL Core Loss

cm centimetre

CO2 Carbon Dioxide

CRM Certified Reference Materials

Cu Copper

CV Coefficient of Variation

DCF Discounted Cash Flow

DD Drillholes

DFS Definitive Feasibility Study (or Bankable Feasibility Study)

DGPS Digital Global Positioning System

DMS Dense Media Separation

DPEM Department Responsible for the Protection of the Mining Environment

DRA DRA Projects (Pty) Ltd

DRC Democratic Republic of the Congo

DTM Digital Terrain Model

EIA Environmental Impact Assessment

EIS Environmental and Social Impact Study

EMPP Environmental Management Plan of the Project

EPCM Engineering, Procurement and Construction Management

ESIA Environmental and Social Impact Assessment

ESIS Environmental and Social Impact Study

EZ East Zone

FARDC Armed Forces of the Democratic Republic of the Congo

FEL Front End Loader

FeS2 Ferrous Disulfide

g Gram

G&A General and Administration

g/t grams per tonne

GDRC The Government of Democratic Republic of the Congo

GHG Green House Gas

GIS Geographical Information System

GPS Global Positioning System

GSSA Geological Society of South Africa

h hours

HCl Hydrochloric Acid

HClO4 Perchloric acid

HF Hydrogen Fluoride

HG High Grade

HQ Diamond Core Diameter of 63.5 mm

HNO3 Nitric acid

ICP-AES Inductively Coupled Plasma Atomic Emission Spectroscopy

ICP-MS Inductively Coupled Plasma Mass Spectrometry

ICT Information and Communications Technology

ID Identity

IDC Industrial Development Corporation of South Africa Limited

147

Term Explanation

JMP Computer programs for statistical analysis developed by the JMP business unit of SAS Institute

JSE-AltX Johannesburg Stock Exchange Alternative Exchange

KAl2 Aluminium Potassium

kg kilogram

kℓ kilolitres

km kilometre

km2 square kilometre

kph kilometre per hour

kt kilotonnes

ktpa kilotonnes per annum

ktpm kilotonnes per month

kV kilovolt

ℓ litre

ℓ/hr litre per hour

ℓ/s litre per second

ℓ/t litre per tonne

Latona Latona Consulting

LDV Light Delivery Vehicle

LENDEN product (accumulation) of length (m) and SG

LG Low Grade

LHD Load, Haul, Dump Machine

LHOS Long Hole Open Stoping

LIDAR Light Detection and Ranging

LoM Life-of-Mine

LME London Metals Exchange

m metres

M Million

m/s metres per second

m2 square meter

m3 cubic metres

Ma Million years ago

mamsl metres above sea level

MDM Master Data Management

METS Meta-Sediment

Mintek A division of the Council for Scientific and Industrial Research (CSIR)

Mℓ Million litres

Mm millimetres

MMS Maelgwyn Mineral Services

MPa Megapascal

MPC Mining and Processing Congo SPRL

MRM Mineral Resource Management

MRP Mitigation and Rehabilitation Plan

MSA The MSA Group

MSCH Mica Schist

Mt Million tonnes

Mtpa Million tonnes per annum

NASA National Aeronautics and Space Administration

ND Not Determined

NI 43-101 National Instrument 43-101 Standards of Disclosure of Mineral Projects

NPV8 Net Present Value at a Discount Rate of 8%

NQ Diamond Core Diameter of 47.6mm

NRG New Resolution Geophysics

number/h number per hour

OECD Organisation for Economic Cooperation and Development

OREAS Ore Research and Exploration of Australia

Pb Lead

PEs Permis d’Exploitation

PER Certificats d’Exploitation des Rejets

PGEP Environmental Management Plan of the Project

ppm Parts per Million

PQ Diamond Core Diameter of 85.0mm

PRs Permis de Recherches Minières

QA/QC Quality Assurance and Quality Control

QEMSCAN SMS QEMSCAN Specific Mineral Search

QP Qualified Person

QQ Quantile-Quantile

QSSH Quartz-Sericite Schist

Quantile-quantile QQ

RDI Resource Development Incorporated

REE Rare Earth Element

148

Term Explanation

RoM Run-of-Mine

ROW Rest of the World

SANAS South African National Accreditation System

SD Standard Deviation

SG Specific Gravity

SGS Société Générale de Surveillance Lakefield in Johannesburg

SiO2 Silicon Dioxide

SMI Sound Mining International (Pty) Ltd

Sn Tin

SnO₂ Cassiterite

SNPCTDL product (accumulation) of Sn %, length (m) and SG

Snowden Snowden Mining Industry Consultants (Pty) Ltd

Sound Mining Sound Mining Solution (Pty) Ltd

t tonne

t/a tonnes per annum

t/m3 tonne per cubic metre

The Mine Bisie Tin Mine

The Project The Development phase of Bisie Mine preceding 2020

TSF Tailings Storage Facility

TSX-V TSX Venture Exchange

μm micrometres

USD United States Dollars

USD/kWh United States Dollars per kilo Watt hour

USD/ℓ United States Dollars per litre

USD/t United States Dollars per tonne

UV Utility Vehicle

VAT Value Added Tax

XRD X-ray Powder Diffraction

XRF X-Ray Fluorescence

Zn Zinc

149

32. APPENDIX 4: MINERAL RESOURCE SPLIT OF ROM SCHEDULED FROM THE BLOCK MODEL

Source: SMI

Description Unit 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 2032

RoM ore from development

Measured (kt) 8.7 0.5 0.2 0.5 1.5 -

Indicated (kt) 140.1 87.0 108.3 87.4 47.6 69.1 45.0 20.1 24.5 19.7 0.2 -

Inferred (kt) 14.8 13.0 16.8 26.6 13.4 35.9 31.2 25.8 20.8 32.5 0.2 -

RoM ore from stoping

Measured (kt) 8.0 14.1 1.2 0.1 15.3

Indicated (kt) 139.6 249.7 247.5 263.7 309.1 244.0 270.2 269.3 190.1 172.6 165.0 84.2 21.6

Inferred (kt) 3.5 8.8 8.9 9.7 11.0 22.5 13.5 47.4 26.7 47.5 55.7 30.0 11.9

All RoM ore scheduled

Measured (kt) 16.7 14.6 0.2 1.7 0.1 16.8

Indicated (kt) 279.7 336.7 355.8 351.2 356.7 313.2 315.2 289.4 214.5 192.3 165.2 84.2 21.6

Inferred (kt) 18.3 21.8 25.7 36.3 24.4 58.4 44.7 73.2 47.5 81.0 55.9 30.0 11.9

RoM content from development ore

Measured (kt) 0.35 0.02 0.01 0.02 0.05 - -

Indicated (kt) 5.62 3.62 4.27 3.56 1.98 3.22 1.69 0.88 0.77 0.48 0.01 - -

Inferred (kt) 0.59 0.54 0.66 1.08 0.56 1.67 1.18 1.12 0.65 0.79 0.01 - -

RoM content from stoping ore

Measured (kt) 0.32 0.59 0.05

Indicated (kt) 5.60 10.39 9.76 10.74 12.87 11.38 10.17 11.72 5.97 4.22 6.05 5.23 1.30

Inferred (kt) 0.14 0.37 0.35 0.40 0.46 1.05 0.51 2.06 0.84 1.18 2.04 1.87 0.72

Total RoM content

Measured (kt) 0.67 0.61 0.01 0.07 0.53

Indicated (kt) 11.22 14.01 14.03 14.29 14.85 14.60 11.86 12.59 6.74 4.70 6.05 5.23 1.30

Inferred (kt) 0.73 0.91 1.01 1.48 1.02 2.72 1.68 3.19 1.49 1.98 2.05 1.87 0.72

Classification split from ore development

Measured % 2.8 0.1 0.2 0.6

Indicated % 44.5 23.3 28.4 22.6 12.4 18.6 12.5 5.5 8.8 7.2 0.1

Inferred % 4.7 3.5 4.4 6.9 3.5 9.6 8.7 7.1 7.4 11.9 0.1

Classification split of ore from stoping

Measured % 2.6 3.8 0.3 0.0 5.5

Indicated % 44.4 66.9 64.9 68.0 80.7 65.7 75.1 74.3 68.1 63.2 74.6 73.7 64.5

Inferred % 1.0 2.4 2.3 2.5 2.9 6.1 3.7 13.1 9.6 17.7 25.2 26.3 35.5

Classification split of all RoM ore

Measured % 5.3 3.9 0.5 6.1

Indicated % 88.9 90.2 93.3 90.6 93.1 84.3 87.6 79.8 76.9 70.4 74.7 73.7 64.5

Inferred % 5.7 5.9 6.7 9.4 6.4 15.7 12.4 20.2 17.0 29.6 25.3 26.3 35.5

150

33. APPENDIX 5: INVENTORY OF DRILLHOLE INTERSECTIONS WITHIN THE MINERAL RESOURCE Hole Name From

(m) To (m)

Easting (m)

Northing (m)

RL (m)

Vein Zone

Sn (%)

Cu (%)

Zn (%)

Pb (%)

Ag (g/t)

As (ppm)

S (%)

SG True Width (m)

BGC001 80.00 82.00 582850 9885756 720 FWVN 9.72 0.07 0.05 0.00 0.00 28 1.11 3.13 1.68

BGC002 80.00 80.50 582846 9885856 708 FWVN 0.47 0.00 0.04 0.00 0.00 19 0.05 3.14 0.41

BGC006 82.30 89.00 582868 9885708 712 FWVN 1.14 0.07 0.08 0.00 0.62 22 1.03 2.97 5.62

BGC007 80.00 80.40 582852 9885812 717 FWVN 16.95 0.19 0.12 0.00 1.10 137 0.57 4.27 0.34

BGC010 122.20 122.50 582891 9885855 641 FWVN 6.90 0.14 0.01 0.00 0.00 373 0.81 3.25 0.25

BGC017 112.50 114.00 582891 9885804 659 FWVN 0.79 0.06 0.04 0.01 0.24 39 1.14 2.95 1.27

BGC018 119.60 123.00 582893 9885704 659 FWVN 2.25 0.05 0.12 0.01 0.92 29 0.90 3.13 2.88

BGC021 147.00 147.50 582905 9885675 623 FWVN 0.66 0.28 0.04 0.04 10.30 44 0.39 3.20 0.42

BGC022 75.00 81.00 582864 9885681 720 FWVN 1.13 0.13 0.05 0.00 1.52 45 1.06 3.12 5.03

BGC023 116.00 118.00 582880 9885682 673 FWVN 1.20 0.36 0.20 0.01 3.74 62 1.32 3.21 1.70

BGC024 151.50 153.00 582912 9885808 605 FWVN 5.93 0.07 0.01 0.00 0.00 232 0.70 3.41 1.25

BGC052 110.00 113.00 582876 9885695 680 FWVN 1.92 0.03 0.02 0.00 0.00 18 0.92 2.91 2.13

BGC055 100.00 100.40 582880 9885744 687 FWVN 2.86 0.05 0.04 0.01 0.50 23 0.96 3.08 0.36

BGC057 91.60 93.00 582873 9885803 695 FWVN 1.48 0.09 0.15 0.01 0.59 17 0.56 3.02 1.16

BGC058 128.00 134.00 582900 9885803 635 FWVN 4.21 0.11 0.11 0.01 0.52 21 0.64 3.18 4.47

BGC058A 133.00 135.50 582902 9885800 631 FWVN 1.45 0.05 0.15 0.01 0.55 22 0.66 3.07 1.85

BGC063A 181.00 184.00 582922 9885809 567 FWVN 2.02 2.15 0.30 0.00 17.00 344 5.06 3.40 2.21

BGC076 121.00 121.50 582880 9885827 654 FWVN 52.90 0.36 0.03 0.00 2.10 87 0.79 5.22 0.42

BGC078 98.00 100.80 582887 9885734 683 FWVN 1.51 0.12 0.11 0.01 0.97 150 0.57 3.21 2.36

BGC079 115.10 115.65 582881 9885880 647 FWVN 11.80 0.38 0.09 0.01 3.20 31 3.34 3.46 0.46

BGC080 102.60 103.00 582865 9885829 683 FWVN 6.74 0.13 0.06 0.00 1.50 311 3.01 3.33 0.34

BGC081 113.00 113.50 582885 9885779 666 FWVN 0.94 0.03 0.01 0.01 0.90 39 1.81 3.20 0.42

BGC082 108.00 110.00 582852 9885877 668 FWVN 3.13 0.11 0.10 0.12 9.87 45 0.25 3.25 1.68

BGC085 103.50 104.80 582874 9885779 684 FWVN 15.08 0.05 0.05 0.01 0.79 332 0.14 3.55 1.09

BGC100 91.40 93.00 582868 9885723 702 FWVN 1.41 0.03 0.06 0.01 0.34 8 0.09 3.20 1.34

BGC102 67.00 70.00 582856 9885731 732 FWVN 1.96 0.08 0.04 0.01 0.73 29 1.90 3.01 2.47

BGC103 92.55 93.00 582869 9885754 700 FWVN 1.10 0.06 0.02 0.01 0.90 36 0.61 2.98 0.29

BGC105 81.00 82.25 582847 9885831 714 FWVN 10.59 0.04 0.35 0.22 8.67 117 0.47 3.47 1.05

METBGC012 85.60 86.00 582851 9885836 705 FWVN 18.55 0.39 0.25 0.02 7.20 124 3.17 3.66 0.33

METBGC013 103.00 104.00 582866 9885845 677 FWVN 0.46 0.14 0.06 0.00 1.00 134 0.58 3.19 0.84

METBGC014 117.60 120.00 582887 9885845 651 FWVN 4.34 0.04 0.02 0.00 0.00 1988 0.78 3.27 1.61

METBGC015 102.00 103.00 582865 9885850 677 FWVN 13.00 0.25 0.17 0.00 1.70 22 1.42 3.49 0.84

METBGC016 118.40 120.00 582885 9885841 651 FWVN 26.21 0.05 0.02 0.01 0.69 4787 1.76 3.93 1.07

METBGC017 101.00 102.10 582863 9885844 678 FWVN 26.83 0.30 0.23 0.00 2.22 11 2.46 3.98 0.90

METBGC018 117.00 118.00 582882 9885847 653 FWVN 12.68 0.33 0.02 0.01 1.58 715 1.34 3.50 0.67

METBGC019 97.00 97.66 582861 9885856 681 FWVN 28.16 0.02 0.06 0.07 10.57 61 0.13 4.01 0.56

METBGC020 117.00 118.00 582880 9885856 652 FWVN 0.92 0.13 0.05 0.00 0.60 40 3.87 3.20 0.67

METBGC021 98.00 100.00 582854 9885858 681 FWVN 2.24 0.13 0.68 0.22 19.19 634 0.77 3.23 1.71

METBGC022 116.30 116.80 582877 9885865 653 FWVN 51.50 0.50 0.21 0.02 3.90 10 1.63 5.14 0.33

METBGC023 101.00 104.45 582854 9885868 677 FWVN 2.30 0.06 0.16 0.11 7.39 84 0.28 3.23 2.89

METBGC026 129.00 129.40 582889 9885838 641 FWVN 8.77 0.13 0.06 0.01 0.70 143 0.55 3.38 0.26

METBGC027 100.00 101.00 582858 9885849 681 FWVN 34.47 0.63 0.16 0.03 14.97 53 1.94 4.25 0.86

BGC002 51.50 54.30 582859 9885857 732 HWVN 1.38 0.02 0.04 0.01 0.00 39 0.15 2.92 2.31

BGC011A 36.50 37.50 582851 9885915 746 HWVN 0.74 0.18 0.03 0.00 0.00 74 0.55 3.20 0.84

BGC025 72.80 73.40 582873 9885911 691 HWVN 2.37 0.04 0.03 0.01 0.00 7 0.23 3.23 0.51

BGC026 100.00 101.00 582906 9885911 645 HWVN 0.58 0.01 0.07 0.00 0.00 98 0.36 3.15 0.85

BGC028 25.00 26.00 582830 9885883 771 HWVN 1.74 0.02 0.00 0.01 0.00 12 0.07 3.22 0.84

151

Hole Name From (m)

To (m)

Easting (m)

Northing (m)

RL (m)

Vein Zone

Sn (%)

Cu (%)

Zn (%)

Pb (%)

Ag (g/t)

As (ppm)

S (%)

SG True Width (m)

BGC031 116.10 116.55 582897 9885961 647 HWVN 2.09 0.02 0.05 0.00 0.00 21 0.66 3.23 0.38

BGC064A 85.20 85.60 582886 9885913 673 HWVN 0.32 0.01 0.01 0.01 0.00 6 0.03 3.19 0.27

BGC067 49.00 50.00 582856 9885917 731 HWVN 0.13 0.04 0.04 0.00 0.00 38 0.23 3.19 0.69

BGC073 50.00 55.40 582860 9885878 727 HWVN 0.58 0.01 0.06 0.00 1.06 12 0.12 3.20 4.53

BGC076 100.00 102.00 582890 9885827 671 HWVN 6.79 0.10 1.49 0.05 8.27 62 1.47 3.34 1.70

BGC079 87.40 88.00 582895 9885880 671 HWVN 3.81 0.04 0.11 0.00 0.00 428 0.43 3.26 0.50

BGC080 74.00 74.40 582879 9885829 707 HWVN 9.66 0.06 1.86 0.00 0.00 8 2.78 3.40 0.34

BGC082 65.33 66.30 582874 9885878 705 HWVN 0.00 0.00 0.00 0.00 0.00 0 0.00 3.18 0.81

BGC105 51.50 52.10 582862 9885831 739 HWVN 3.82 0.04 0.05 0.02 0.00 5 0.34 2.97 0.50

BGC114 21.00 28.50 582834 9885873 773 HWVN 0.34 0.01 0.01 0.00 0.38 44 0.01 3.19 5.75

METBGC007 53.00 56.00 582859 9885852 732 HWVN 17.81 0.04 0.06 0.02 0.19 37 0.15 3.65 2.50

METBGC008 47.70 56.00 582863 9885846 734 HWVN 3.42 0.03 0.08 0.02 0.96 270 0.44 3.26 6.80

METBGC009 49.60 57.00 582863 9885844 733 HWVN 5.13 0.03 0.06 0.01 0.16 81 0.50 3.30 6.08

METBGC010 51.40 56.60 582864 9885842 732 HWVN 0.59 0.03 0.06 0.01 0.18 20 0.68 3.20 4.33

METBGC011 53.00 60.00 582864 9885842 730 HWVN 2.52 0.02 0.08 0.01 0.00 6 0.21 3.24 5.87

METBGC012 54.00 55.00 582866 9885839 732 HWVN 4.72 0.11 3.35 0.01 0.41 7 2.36 3.28 0.84

METBGC013 70.00 77.00 582881 9885845 703 HWVN 1.08 0.02 0.13 0.00 0.10 7 0.32 3.07 5.87

METBGC014 90.30 92.70 582895 9885845 678 HWVN 3.13 0.05 0.07 0.00 0.31 11 0.30 3.25 1.61

METBGC015 69.00 73.00 582881 9885850 704 HWVN 0.79 0.02 0.38 0.01 0.00 0 0.25 3.20 3.35

METBGC016 89.00 92.00 582893 9885843 679 HWVN 3.48 0.07 0.09 0.00 0.28 13 0.18 3.26 2.00

METBGC017 66.00 74.00 582878 9885848 705 HWVN 2.15 0.02 0.20 0.01 0.07 9 0.21 3.23 6.58

METBGC018 87.00 88.00 582890 9885849 681 HWVN 0.81 0.01 0.03 0.01 0.00 11 0.06 3.20 0.67

METBGC019 64.00 71.40 582876 9885856 707 HWVN 0.83 0.01 0.10 0.00 0.00 49 0.59 3.20 6.23

METBGC020 73.00 82.10 582890 9885857 691 HWVN 1.04 0.02 0.09 0.00 0.00 15 0.35 3.21 6.02

METBGC021 61.60 68.40 582872 9885860 710 HWVN 1.85 0.01 0.07 0.01 0.05 51 0.10 3.23 5.86

METBGC022 76.90 85.00 582886 9885864 687 HWVN 0.40 0.01 0.06 0.00 0.00 50 0.09 3.19 5.40

METBGC023 64.10 67.00 582872 9885866 709 HWVN 1.54 0.01 0.06 0.01 0.08 9 0.15 3.22 2.42

METBGC024 86.00 90.00 582884 9885871 680 HWVN 2.93 0.03 0.06 0.00 0.00 33 0.28 3.27 2.65

METBGC025 72.00 73.00 582886 9885839 703 HWVN 6.79 0.09 0.17 0.01 0.00 0 2.54 3.33 0.84

METBGC026 99.00 100.00 582896 9885839 670 HWVN 2.16 0.08 0.09 0.00 1.00 0 1.73 3.23 0.66

METBGC027 64.00 72.45 582876 9885850 708 HWVN 0.78 0.04 0.09 0.01 0.06 8 0.21 3.20 7.28

BGC001 53.00 72.00 582859 9885756 736 MVN 2.30 0.13 0.14 0.01 0.48 8 0.53 3.03 15.84

BGC002 64.20 72.50 582852 9885856 719 MVN 7.84 0.91 0.13 0.00 3.01 106 1.67 3.49 6.87

BGC004 127.00 131.50 582849 9886003 672 MVN 0.42 0.08 0.04 0.01 0.00 47 0.73 3.19 3.77

BGC005 113.40 127.50 582904 9885750 650 MVN 1.35 0.16 0.19 0.01 1.57 46 1.06 3.21 11.82

BGC006 65.70 72.00 582876 9885708 726 MVN 1.50 0.31 0.17 0.01 4.59 8 0.78 3.14 5.28

BGC007 53.00 70.40 582861 9885812 733 MVN 2.54 0.40 0.13 0.00 1.84 152 0.88 3.34 14.58

BGC010 111.00 115.63 582895 9885855 649 MVN 3.38 0.06 0.06 0.00 0.08 33 0.20 3.17 3.80

BGC011A 70.74 72.75 582835 9885914 716 MVN 1.52 0.62 0.12 0.01 3.04 13 3.00 3.22 1.64

BGC017 100.00 103.50 582896 9885804 669 MVN 0.47 0.29 0.10 0.00 1.39 13 0.57 3.13 2.96

BGC018 100.00 117.00 582899 9885704 670 MVN 7.75 0.52 0.17 0.00 3.82 3516 1.20 3.46 14.23

BGC019 132.50 146.70 582919 9885758 623 MVN 2.06 0.55 0.21 0.00 3.87 2423 3.01 3.28 12.03

BGC020 137.70 146.00 582920 9885705 621 MVN 1.05 1.18 0.24 0.01 11.08 384 3.27 3.24 7.05

BGC021 130.00 140.00 582911 9885675 634 MVN 0.79 0.60 0.21 0.00 5.85 364 1.64 3.05 8.45

BGC022 58.50 70.50 582871 9885681 732 MVN 1.61 0.30 0.19 0.01 3.25 13 0.87 3.08 10.08

BGC023 91.50 111.40 582888 9885682 686 MVN 5.06 0.46 0.22 0.00 4.06 4553 3.15 3.30 16.89

BGC024 133.00 146.50 582918 9885808 616 MVN 3.17 0.17 0.11 0.01 1.24 795 0.90 3.32 11.22

BGC025 97.45 108.00 582858 9885911 666 MVN 3.27 0.22 0.06 0.00 0.90 129 0.72 3.24 8.84

152

Hole Name From (m)

To (m)

Easting (m)

Northing (m)

RL (m)

Vein Zone

Sn (%)

Cu (%)

Zn (%)

Pb (%)

Ag (g/t)

As (ppm)

S (%)

SG True Width (m)

BGC026 125.00 140.00 582889 9885911 618 MVN 9.07 0.13 0.07 0.01 1.57 144 0.78 3.40 12.67

BGC027 125.00 135.00 582913 9885863 621 MVN 2.56 0.13 0.09 0.01 0.88 23 0.55 3.09 8.39

BGC028 50.00 54.50 582817 9885883 748 MVN 0.63 0.17 0.05 0.00 1.83 2 0.03 3.20 3.79

BGC029 97.00 98.00 582843 9885961 690 MVN 0.63 0.10 0.04 0.00 0.00 124 2.02 3.20 0.84

BGC030 112.50 115.00 582861 9885957 664 MVN 0.49 0.06 0.04 0.00 0.23 98 0.95 3.19 2.11

BGC031 140.50 147.00 582884 9885961 623 MVN 5.40 0.05 0.06 0.01 0.71 95 0.75 3.31 5.48

BGC034 175.20 199.65 582909 9885961 566 MVN 10.19 0.05 0.31 0.06 4.30 105 0.71 3.58 16.36

BGC035 165.00 190.80 582940 9885807 566 MVN 4.87 0.64 0.16 0.00 5.59 684 1.99 3.34 17.22

BGC036 133.00 148.20 582919 9885913 596 MVN 4.55 0.06 0.27 0.01 0.35 9 0.21 3.24 10.24

BGC037 185.00 200.00 582927 9885860 548 MVN 10.21 0.61 0.20 0.00 4.56 166 1.76 3.46 10.61

BGC038 247.00 268.35 582964 9885953 488 MVN 2.46 0.52 0.19 0.01 3.57 1141 1.41 3.27 16.81

BGC039 193.84 201.17 582930 9885739 557 MVN 0.44 0.18 0.14 0.00 1.76 68 0.57 3.08 5.52

BGC041 203.00 220.00 582907 9886012 574 MVN 1.97 0.40 0.08 0.02 2.79 296 1.47 3.23 13.98

BGC042 262.60 282.55 582955 9886001 497 MVN 3.05 0.08 0.08 0.01 0.39 147 0.36 3.25 17.49

BGC043 317.00 324.00 582985 9886053 442 MVN 3.97 0.15 0.13 0.01 0.81 417 0.87 3.22 5.87

BGC044 376.00 386.00 582989 9886095 392 MVN 1.03 0.29 0.08 0.01 1.96 68 1.04 3.18 8.39

BGC045 252.30 260.00 582912 9886055 542 MVN 2.06 0.13 0.10 0.04 3.38 113 0.44 3.23 6.46

BGC046 276.00 282.50 582961 9885853 508 MVN 1.74 0.68 0.35 0.00 4.81 890 2.87 3.37 5.38

BGC047 280.70 305.85 582942 9886053 494 MVN 2.98 0.07 0.14 0.06 4.29 101 0.32 3.29 21.46

BGC048 152.00 163.30 582918 9885895 589 MVN 2.47 0.12 0.04 0.01 0.63 17 0.49 3.24 9.48

BGC049 301.50 306.50 582967 9886105 470 MVN 0.24 0.04 0.08 0.06 2.69 91 0.18 3.19 4.19

BGC050 181.00 193.00 582948 9885893 545 MVN 4.15 0.40 0.31 0.01 2.51 861 2.26 3.36 8.03

BGC051 364.00 376.00 583003 9886059 407 MVN 2.73 0.22 0.15 0.01 0.83 139 0.67 3.25 10.06

BGC052 81.00 107.00 582881 9885697 697 MVN 1.64 0.32 0.30 0.01 3.17 42 1.17 3.19 18.45

BGC053 115.00 137.00 582908 9885704 647 MVN 11.32 0.69 0.25 0.00 4.87 1484 1.51 3.50 17.11

BGC054 288.90 293.00 582975 9885904 495 MVN 6.19 1.06 0.34 0.01 7.75 310 3.90 3.42 3.50

BGC055 88.00 98.00 582884 9885745 692 MVN 1.01 0.17 0.13 0.01 0.84 15 1.22 3.00 9.06

BGC056 397.60 409.00 583017 9886092 375 MVN 3.59 0.26 0.16 0.01 1.82 73 0.94 3.32 9.61

BGC057 86.00 88.60 582875 9885803 699 MVN 3.60 0.12 0.16 0.01 0.00 16 0.34 3.21 2.16

BGC058 122.00 128.00 582902 9885803 641 MVN 0.53 0.34 0.08 0.01 2.58 124 0.69 3.21 4.47

BGC058A 122.00 131.20 582905 9885801 638 MVN 1.82 0.22 0.18 0.01 1.25 118 0.79 3.22 6.72

BGC059A 154.00 163.50 582921 9885859 587 MVN 1.58 0.21 0.10 0.01 1.46 81 1.29 3.18 7.09

BGC060 204.18 214.50 582945 9885850 528 MVN 7.47 0.79 0.15 0.01 5.68 263 2.14 3.50 6.05

BGC061A 243.00 263.00 582938 9886004 521 MVN 5.66 0.09 0.09 0.02 2.31 25 0.40 3.33 16.91

BGC062 304.00 315.50 582991 9885949 452 MVN 15.37 0.66 0.24 0.00 5.19 190 1.86 3.68 9.72

BGC063A 159.50 178.50 582927 9885809 580 MVN 2.07 0.36 0.24 0.00 2.63 133 1.85 3.30 14.04

BGC064A 113.00 126.00 582876 9885913 640 MVN 2.30 0.40 0.16 0.02 2.82 160 0.78 3.34 8.87

BGC066 224.85 244.00 582951 9885961 518 MVN 0.93 0.19 0.09 0.01 1.37 94 0.80 3.20 15.68

BGC067 85.20 92.20 582845 9885918 693 MVN 1.19 0.06 0.05 0.00 0.15 165 0.74 3.09 4.79

BGC068 284.60 288.00 582960 9885804 498 MVN 2.02 0.15 0.34 0.01 1.67 1395 0.85 3.23 2.86

BGC069 128.00 134.00 582872 9885957 640 MVN 1.96 0.05 0.05 0.00 0.19 184 0.71 3.17 4.44

BGC070 146.00 155.00 582905 9885960 606 MVN 8.50 0.21 0.11 0.02 3.29 110 0.68 3.50 6.59

BGC071 180.00 185.00 582941 9885703 568 MVN 0.97 0.05 0.19 0.01 0.26 234 0.63 3.20 3.37

BGC072 427.00 438.00 583021 9886057 350 MVN 16.33 0.76 0.43 0.00 4.87 346 2.41 3.71 9.17

BGC073 69.00 78.00 582850 9885878 709 MVN 2.60 0.52 0.08 0.00 1.67 153 1.49 3.41 7.60

BGC074 317.00 330.00 582993 9886000 426 MVN 3.08 0.56 0.19 0.01 3.16 1310 2.95 3.42 11.10

BGC076 107.40 113.00 582886 9885827 664 MVN 1.33 0.13 0.07 0.01 0.06 33 0.85 3.21 4.75

BGC077 454.00 463.70 583046 9886098 309 MVN 9.60 0.57 0.15 0.00 2.57 1316 3.03 3.42 8.07

153

Hole Name From (m)

To (m)

Easting (m)

Northing (m)

RL (m)

Vein Zone

Sn (%)

Cu (%)

Zn (%)

Pb (%)

Ag (g/t)

As (ppm)

S (%)

SG True Width (m)

BGC078 86.00 93.00 582892 9885734 691 MVN 1.43 0.20 0.13 0.01 1.59 57 0.76 3.21 5.89

BGC079 102.40 109.00 582886 9885880 655 MVN 3.66 0.06 0.14 0.00 0.09 222 0.25 3.27 5.53

BGC080 82.50 93.25 582872 9885829 695 MVN 12.18 0.07 0.10 0.00 0.24 92 0.16 3.49 9.04

BGC081 105.00 108.30 582888 9885779 672 MVN 1.21 0.06 0.04 0.00 0.06 17 0.85 3.21 2.79

BGC082 82.00 97.00 582862 9885877 685 MVN 2.43 0.20 0.05 0.01 0.67 167 0.79 3.24 12.58

BGC083 86.00 88.50 582843 9885931 696 MVN 5.32 0.26 0.05 0.00 1.06 17 3.03 3.30 2.08

BGC084 170.60 185.80 582875 9886008 614 MVN 1.14 0.10 0.04 0.00 0.40 77 1.03 3.21 12.48

BGC085 76.40 94.10 582884 9885779 700 MVN 0.56 0.08 0.08 0.01 0.55 6 0.37 3.20 14.83

BGC086 83.00 102.00 582886 9885638 694 MVN 4.88 0.96 0.52 0.00 7.02 2422 3.04 3.29 15.90

BGC087A 218.50 230.00 582899 9886047 575 MVN 1.84 0.09 0.05 0.01 3.23 97 0.53 3.09 9.77

BGC088 379.00 384.50 583028 9886006 373 MVN 2.60 0.33 0.15 0.00 4.76 137 2.32 3.30 4.59

BGC089 362.00 364.00 582988 9885901 414 MVN 1.03 0.09 0.11 0.00 1.39 105 0.31 3.21 1.66

BGC090 116.80 121.70 582892 9885637 663 MVN 2.88 0.77 0.43 0.01 9.74 883 1.28 3.30 4.12

BGC093 516.00 524.00 583059 9886099 241 MVN 1.84 0.10 0.13 0.06 2.57 267 1.12 3.27 6.70

BGC097 423.00 426.45 583044 9886006 328 MVN 0.26 0.14 0.05 0.02 0.00 48 1.12 3.25 2.92

BGC098 116.40 129.50 582907 9885616 648 MVN 2.63 1.11 0.36 0.00 9.13 2168 2.28 3.24 11.06

BGC099 121.00 137.65 582912 9885640 640 MVN 6.69 0.89 0.23 0.01 9.30 3453 2.58 3.42 14.00

BGC100 78.60 83.23 582874 9885723 712 MVN 0.78 0.08 0.05 0.00 0.66 13 1.94 3.00 3.87

BGC101 361.00 366.25 583019 9885952 380 MVN 5.77 0.19 0.17 0.03 4.40 100 0.53 3.40 4.38

BGC102 50.50 58.00 582863 9885732 745 MVN 0.15 0.23 0.09 0.00 1.06 35 1.08 3.26 6.17

BGC103 72.50 87.00 582872 9885754 713 MVN 0.40 0.15 0.05 0.00 0.92 935 1.08 3.25 9.26

BGC104 473.90 483.00 583052 9886056 282 MVN 4.31 0.04 0.27 0.01 0.67 236 0.45 3.22 7.70

BGC105 55.80 75.00 582855 9885831 728 MVN 2.31 0.36 0.09 0.01 1.38 210 0.73 3.34 16.02

BGC106 143.00 148.50 582913 9885615 620 MVN 11.49 0.18 0.12 0.00 1.25 1264 1.20 3.49 4.62

BGC107 125.00 137.00 582911 9885587 626 MVN 6.44 0.82 0.14 0.00 4.78 2496 1.51 3.28 9.98

BGC109 146.00 153.00 582925 9885642 608 MVN 3.18 0.21 0.16 0.01 3.25 282 0.89 3.20 5.94

BGC112 145.75 148.50 582925 9885587 602 MVN 1.05 0.41 0.02 0.00 0.61 110 1.53 3.12 2.03

BGC113 169.20 170.00 582942 9885642 578 MVN 2.94 0.39 0.15 0.03 4.30 25 1.33 3.12 0.60

BGC114 45.00 47.00 582824 9885865 755 MVN 0.69 2.01 0.07 0.03 8.20 12 3.40 3.20 1.54

BGC122 46.00 66.00 582864 9885781 740 MVN 1.58 0.11 0.08 0.01 1.60 86 0.80 3.25 16.71

BGC125 110.00 116.50 582891 9885653 669 MVN 4.21 1.56 0.32 0.01 16.60 1417 5.22 3.42 4.95

BGC127 301.00 320.70 582937 9886054 476 MVN 8.95 0.11 0.20 0.03 3.07 430 0.40 3.41 16.68

BGC130 387.00 397.00 583007 9886027 383 MVN 7.74 0.65 0.39 0.00 4.98 229 3.07 3.48 8.40

BGC131 434.00 449.00 583002 9886140 346 MVN 2.90 0.45 0.11 0.01 2.92 268 1.41 3.19 12.70

BGC132 142.00 143.00 582927 9885567 597 MVN 0.31 0.00 0.02 0.00 0.00 19 9.03 2.92 0.73

BGC133 323.00 332.00 583005 9885951 424 MVN 7.91 0.54 0.08 0.00 3.19 607 2.73 3.44 7.11

BGC134 309.55 316.50 582965 9885855 468 MVN 1.66 0.11 0.13 0.00 0.90 80 0.66 3.22 5.45

BGC135 166.00 168.40 582930 9885614 589 MVN 0.40 0.20 0.12 0.01 1.74 25 0.31 2.92 1.75

BGC136 178.00 192.50 582929 9885754 569 MVN 2.16 0.38 0.23 0.00 4.26 1467 1.14 3.02 10.70

BGC137 333.50 342.50 582980 9885855 435 MVN 1.29 0.12 0.42 0.07 3.40 67 0.71 3.00 6.58

BGC138 374.00 383.20 583038 9885951 357 MVN 2.86 0.12 0.13 0.03 4.52 201 0.37 3.17 6.27

BGC139 250.00 255.35 582945 9885770 536 MVN 1.80 0.17 0.28 0.00 1.79 234 0.65 3.22 4.49

BGC140 534.40 546.00 583032 9886138 227 MVN 11.02 0.23 0.11 0.00 0.85 241 1.50 3.48 9.69

BGC141 456.35 469.55 583030 9886134 313 MVN 4.30 0.47 0.21 0.01 2.54 379 2.51 3.26 10.97

BGC143 167.00 175.00 582870 9886032 632 MVN 2.54 0.11 0.05 0.00 0.40 19 1.20 3.05 6.77

BGC148 399.50 414.65 583010 9886080 374 MVN 7.86 0.85 0.24 0.00 7.31 987 2.61 3.38 12.58

BGC150 489.20 497.50 583012 9886082 294 MVN 6.61 0.13 0.09 0.00 1.26 190 1.02 3.28 7.01

BGC159 402.70 406.70 583016 9885979 345 MVN 6.68 0.12 0.07 0.01 1.88 386 0.59 3.21 3.38

154

Hole Name From (m)

To (m)

Easting (m)

Northing (m)

RL (m)

Vein Zone

Sn (%)

Cu (%)

Zn (%)

Pb (%)

Ag (g/t)

As (ppm)

S (%)

SG True Width (m)

BGC161 346.55 355.20 582971 9886136 421 MVN 0.44 0.02 0.05 0.01 0.41 185 0.28 3.18 7.18

BGC162 448.30 454.65 583045 9886030 308 MVN 0.54 0.07 0.04 0.01 0.93 64 0.44 3.22 5.40

BGC164 290.15 295.67 582931 9886091 496 MVN 1.22 0.04 0.08 0.01 0.08 187 0.23 3.32 4.61

BGC165B 336.70 349.20 582964 9886083 428 MVN 15.96 0.17 0.29 0.01 1.71 669 1.45 3.62 10.54

BGC166 387.45 403.03 582990 9886140 390 MVN 29.07 0.25 0.12 0.01 1.19 380 0.59 4.13 13.04

BGC167 318.15 331.45 582976 9885933 452 MVN 9.66 1.25 0.26 0.00 9.05 418 3.91 3.67 11.18

BGC171 243.05 246.10 582902 9886141 541 MVN 0.66 0.10 0.03 0.00 0.28 145 1.49 3.11 2.54

METBGC007 62.00 71.50 582853 9885851 721 MVN 7.57 0.46 0.11 0.01 2.80 157 0.68 3.36 7.96

METBGC008 62.80 71.13 582855 9885844 721 MVN 8.64 0.23 0.09 0.01 1.22 714 0.40 3.39 6.82

METBGC009 63.65 72.63 582856 9885842 720 MVN 11.18 0.37 0.10 0.02 2.97 109 0.66 3.46 7.39

METBGC010 64.00 73.30 582857 9885840 720 MVN 12.73 0.37 0.11 0.02 2.48 459 0.61 3.51 7.76

METBGC011 66.00 73.50 582858 9885842 718 MVN 20.86 0.44 0.18 0.00 2.08 154 0.67 3.75 6.29

METBGC012 58.00 75.40 582860 9885838 721 MVN 7.68 0.25 0.18 0.00 1.60 196 0.61 3.37 14.58

METBGC013 79.00 93.00 582875 9885845 692 MVN 10.63 0.06 0.12 0.01 0.38 71 0.27 3.37 11.74

METBGC014 100.50 108.00 582891 9885845 665 MVN 1.31 0.06 0.10 0.01 0.16 9 0.28 3.21 5.02

METBGC015 81.60 93.00 582873 9885850 690 MVN 11.95 0.08 0.10 0.00 0.14 128 0.20 3.48 9.56

METBGC016 104.00 109.00 582889 9885842 663 MVN 0.89 0.02 0.08 0.01 0.10 6 0.44 3.20 3.33

METBGC017 75.00 92.45 582872 9885846 693 MVN 17.33 0.09 0.09 0.00 0.95 89 0.33 3.72 14.35

METBGC018 101.00 107.00 582885 9885848 666 MVN 0.35 0.05 0.09 0.02 0.81 12 0.61 3.19 4.02

METBGC019 77.00 86.00 582869 9885856 695 MVN 13.13 0.21 0.55 0.00 1.21 133 0.83 3.52 7.58

METBGC020 97.50 109.00 582884 9885856 666 MVN 4.40 0.16 0.10 0.00 1.25 30 0.29 3.28 7.63

METBGC021 76.00 86.00 582864 9885859 697 MVN 11.50 0.62 0.11 0.01 2.80 388 1.16 3.47 8.57

METBGC022 95.60 109.00 582880 9885864 667 MVN 20.50 0.05 0.08 0.00 0.33 242 0.20 3.75 8.92

METBGC023 76.60 88.00 582864 9885867 694 MVN 9.97 1.01 0.11 0.01 4.65 205 1.71 3.43 9.53

METBGC024 94.40 110.00 582881 9885871 667 MVN 7.07 0.05 0.09 0.00 0.89 151 0.20 3.41 10.35

METBGC025 77.50 93.40 582879 9885839 692 MVN 4.25 0.05 0.11 0.00 0.20 39 0.18 3.28 13.39

METBGC026 116.40 117.30 582892 9885838 653 MVN 5.68 0.04 0.03 0.01 0.00 3 0.18 3.30 0.59

METBGC027 76.00 95.00 582866 9885850 693 MVN 8.99 0.14 0.13 0.00 1.25 148 0.28 3.40 16.36

BGC092 123.15 137.10 582872 9885617 655 S1VN 9.72 0.04 0.01 0.00 0.26 53 3.15 3.29 11.78

BGC096 124.50 136.00 582886 9885589 637 S1VN 6.26 0.64 0.04 0.00 2.08 2775 4.84 3.40 9.69

BGC110 131.00 137.00 582908 9885568 615 S1VN 8.73 0.36 0.08 0.00 1.60 419 0.80 3.37 5.07

155

34. APPENDIX 6: HISTOGRAMS AND LOG PROBABILITY PLOTS OF THE COMPOSITE DATA

156

157

158

159

160

161

162

163

164

165

166

167

168

169

35. APPENDIX 7: SEMI VARIOGRAMS

170

171

172

173

174

175

176

Documents

NI 43-101 Technical Report 31 December 2019 Bisie Tin Mine