Technical Report on the Lace Diamond Mine Free State, Republic of South Africa for Diamondcorp Plc

TECHNICAL REPORT ON THE LACE DIAMOND MINE FREE STATE, REPUBLIC OF SOUTH AFRICA FOR DIAMONDCORP PLC

P. Sobie, P.Geo. JJ Ferreira PhD Pr. Sci. Nat. February 1, 2016 P.G. Allan, BSc Hons. Toronto, Canada P.Loudon, BSc (Hons),MSAIMM

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

TABLE OF CONTENTS SUMMARY ...... vi 1.0 INTRODUCTION ...... 1-1 1.1 Authorization and Terms of Reference ...... 1-1 1.2 Qualifications of Authors ...... 1-1 1.3 Scope of Work and Sources of Information ...... 1-2 2.0 RELIANCE ON OTHER EXPERTS ...... 2-1 3.0 PROPERTY DESCRIPTION AND LOCATION ...... 3-1 3.1 Mineral Policy South Africa ...... 3-1 3.2 Lace Diamond Mine Title, Location and Demarcation of Mining Permit ...... 3-1 3.3 Permitting Rights and Agreements Related to the Lace Diamond Mine ...... 3-4 3.3.1 Surface Rights ...... 3-4 3.3.2 Taxes and Royalties ...... 3-4 3.3.3 Obligations ...... 3-4 3.3.4 Environmental Liabilities...... 3-4 3.4 Permits ...... 3-6 4.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY...... 4-1 5.0 HISTORY ...... 5-1 5.1 Summary of Pre-Great Depression Period 1896-1931 ...... 5-1 5.2 De Beers Period 1931-1996 ...... 5-5 5.3 Christian Potgeiter Trust 1996-2005 ...... 5-6 5.3.1 Delineation Drilling and Sampling Programme ...... 5-8 5.3.2 Tailings Bulk Sampling Programme ...... 5-10 5.4 DiamondCorp plc 2005-2014 ...... 5-12 5.4.1 General ...... 5-12 5.4.2 Tailings Reprocessing and Plant Construction ...... 5-14 5.4.3 2009 Satellite and Main Pipe Bulk Sampling ...... 5-14 5.4.4 2011 Main Pipe Bulk Sampling ...... 5-15 5.4.5 2012-2014 Development and Underground Delineation Drilling ...... 5-19 6.0 GEOLOGICAL SETTING AND MINERALIZATION ...... 6-1 6.1 Regional and Local Geology ...... 6-1 6.2 Lace Kimberlite Geology ...... 6-1 6.2.1 K2 (Hypabyssal now called Coherent) Kimberlite ...... 6-1 6.2.2 K4 (Hypabyssal now called Coherent Kimberlite) ...... 6-2 6.2.3 K6 (TKB now called Fragmental) Kimberlite ...... 6-4 6.2.4 K8 (Lava-Rich Kimberlite Breccia) ...... 6-4 6.2.5 Country Rock Breccia (Lava Breccia, Shale Breccia, CRB) ...... 6-4 6.2.6 Ventersdorp Lava (LV) ...... 6-4 6.3 Internal and Pipe Margin Geometry ...... 6-5 6.4 Mineralization – Lace Kimberlites ...... 6-6 6.4.1 General ...... 6-6 6.4.2 The Satellite Pipe ...... 6-6 6.4.3 The Main Pipe ...... 6-7

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6.4.4 Commercial Potential...... 6-8 6.4.5 Main Pipe Upper K4 Mine -230m to -370m ...... 6-8 6.4.6 Block Caves 1, 2 and 3 -370m to -900m ...... 6-8 7.0 DEPOSIT TYPES ...... 7-1 8.0 EXPLORATION ...... 8-1 9.0 DRILLING...... 9-1 10.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY ...... 10-1 10.1 Delineation Core Microdiamond Sampling ...... 10-1 10.2 Bulk Sample Microdiamond Sampling ...... 10-3 10.3 Bulk Sample Macrodiamond Sampling ...... 10-4 10.3.1 Bulk Sample Mining and Stockpiling ...... 10-4 10.3.1 Bulk Sample Processing and Security ...... 10-5 11.0 DATA VERIFICATION ...... 11-1 11.1 Geological Model...... 11-1 11.2 Drill Core Logs ...... 11-2 11.3 Internal Dilution Data ...... 11-2 11.4 Bulk Density Data ...... 11-2 12.0 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 12-1 13.0 ADJACENT PROPERTIES ...... 13-1 14.0 MINERAL RESOURCE ESTIMATE...... 14-1 14.1 Evaluation Databases ...... 14-1 14.1.1 Drill Hole Database...... 14-1 14.1.1 Bulk Sample Database ...... 14-5 14.2 Geological Modelling ...... 14-5 14.2.1 Geological Continuity ...... 14-10 14.2.2 Volume and Tonnage Estimates ...... 14-13 14.2.1 Confidence Level of Geological Model ...... 14-13 14.3 Block Model...... 14-14 14.3.1 Bulk Density Data ...... 14-14 14.4 Bulk Dilution ...... 14-16 14.4.1 Bulk Dilution Data ...... 14-16 14.4.2 Bulk Density and Dilution Estimation ...... 14-17 14.4.3 Confidence Level of Bulk Density and Tonnage Model ...... 14-19 14.5 Diamond Grade ...... 14-20 14.5.1 Approach and Sampling ...... 14-20 14.5.1.1 Bulk Sample Macrodiamond Data ...... 14-21 14.5.1.2 Core Microdiamond Data ...... 14-22 14.5.2 Diamond Size Distribution from Bulk Sampling...... 14-23 14.5.3 Diamond Size Distribution from Microdiamonds ...... 14-27 14.5.3.1 Group 1 Size and Grade-Size Modelling ...... 14-31 14.5.3.2 Group 2 Size and Grade-Size Modelling ...... 14-34 14.5.3.3 Group 3 Size and Grade-Size Modelling ...... 14-37 14.6 Diamond Value Estimation ...... 14-40 14.6.1 Approach and Data ...... 14-40

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

14.7 Diamond Grades and Values ...... 14-41 14.8 Discussion on Grade Modelling Results ...... 14-43 14.9 Diamond Grade Model ...... 14-45 14.10 Mineral Resource Statement ...... 14-47 15.0 MINERAL RESERVE ESTIMATE ...... 15-1 16.0 MINING METHOD ...... 16-1 16.1 Geotechnical and Hydrological Aspects ...... 16-1 16.1.1 Field Strength Estimates ...... 16-1 16.1.2 Laboratory Rock Test Results ...... 16-2 16.1.3 Jointing ...... 16-3 16.1.3.1 Jointing in Ventersdorp Lavas and Shales ...... 16-3 16.1.3.2 Jointing in Kimberlite ...... 16-3 16.1.3.3 Rock Competency from Core Measurements ...... 16-4 16.1.4 In Situ Ground Stress ...... 16-7 16.1.5 Rock Mass Classifications ...... 16-8 16.1.5.1 Ventersdorp Rock Mass Conditions ...... 16-8 16.1.5.2 Kimberlite Rock Mass Conditions ...... 16-8 16.1.6 Design Rock Mass Strength ...... 16-8 16.2 Mining Method Option Study ...... 16-10 16.2.1 Mining Methods Considered...... 16-10 16.2.2 Mining Method Evaluations ...... 16-10 16.2.3 Mining Method Selection ...... 16-11 16.3 High Level Risks...... 16-13 16.3.1 Introduction ...... 16-13 16.3.2 Mud Rush Risk Potential - Conditions Required for Mud Rushes to Occur ... 16-13 16.3.2.1 Mud Forming Material ...... 16-14 16.3.2.2 Water ...... 16-14 16.3.2.3 Ground Disturbance (or mud disturbance) ...... 16-14 16.3.2.4 Discharge Points on an Underground Working Level ...... 16-14 16.3.3 Strategies for the Prevention of Mud Rushes ...... 16-15 16.3.3.1 Surface Storm Water Control...... 16-15 16.3.3.2 Mine Waste Disposal ...... 16-15 16.3.3.3 Mine Design ...... 16-15 16.3.3.4 Draw Control ...... 16-15 16.3.3.5 Mine De-watering ...... 16-16 16.3.3.6 Mandatory Code of Practise...... 16-17 16.4 Mine Design ...... 16-18 16.4.1 Geotechnical Aspects ...... 16-18 16.4.2 310 Level UK4 Mine Upper Block Layout ...... 16-19 16.4.2.1 Access and Ground handling ...... 16-19 16.4.2.2 General ...... 16-20 16.4.2.3 Doming Level (290m Level) ...... 16-21 16.4.2.1 Production Level (310m Level) ...... 16-23 16.4.3 360m Level UK4 Mine Lower Block Layout ...... 16-23 16.4.3.1 Doming Level (340m Level) ...... 16-23 16.4.3.2 Production Level (360m Level) ...... 16-23

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

16.5 Life of Mine and Production Rates ...... 16-25 16.6 Mining Schedule ...... 16-25 16.7 Mine Ventilation ...... 16-27 16.7.1 Ventilation during Development and Construction ...... 16-27 16.7.2 Production Ventilation ...... 16-27 16.8 Underground Mobile Equipment and Workshop ...... 16-27 16.9 Mine De-watering ...... 16-27 16.9.1 Dewatering Requirements during the Development Phase ...... 16-27 16.9.2 Dewatering Requirements during the Production Phase ...... 16-28 16.10 Labour Requirements ...... 16-29 16.11 Other Parameters and Infrastructure ...... 16-30 16.11.1 Surface Lamp Room and Access Control Facility ...... 16-30 16.11.2 Water Reticulation ...... 16-30 16.11.3 Compressed Air Reticulation ...... 16-30 16.11.4 Electricity Reticulation ...... 16-30 16.11.5 Shaft Head Explosives Delivery Bay ...... 16-30 16.11.6 Refuge Bays ...... 16-30 16.11.7 First Aid Room & First Aid ...... 16-30 16.11.8 Security ...... 16-31 16.11.9 Change House / Locker & Laundry Facility ...... 16-31 16.11.10 Sewage System ...... 16-31 17.0 RECOVERY METHODS ...... 17-1 17.1 Process Description ...... 17-1 17.1.1 Feed Preparation Plant ...... 17-1 17.1.2 Dense Media Separation (“DMS”) Circult ...... 17-2 17.1.3 Re-Crush Circuit ...... 17-3 17.1.4 Water Recovery Circuit ...... 17-4 17.1.5 Final Recovery Plant ...... 17-4 17.2 Plant Modifications During Bulk Sampling ...... 17-5 18.0 ENVIRONMENTAL CONSIDERATIONS ...... 18-1 18.1 Environmental Management ...... 18-1 18.1.1 Slimes Dam ...... 18-2 18.1.2 Processed Kimberlite Dump ...... 18-2 18.1.3 Waste Rock Dump ...... 18-2 18.2 Social and Community ...... 18-2 19.0 PROJECT INFRASTRUCTURE ...... 19-1 19.1 General Overall Underground Infrastructure ...... 19-1 19.2 Infrastructure Additions for UK4 Mine ...... 19-2 20.0 ECONOMIC ANALYSIS ...... 20-1 20.1 Introduction ...... 20-1 20.2 Capital Costs ...... 20-1 20.3 Operating Costs ...... 20-2 20.4 Revenue...... 20-3 20.5 Cashflow ...... 20-3 20.6 Sensitivities ...... 20-5

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

21.0 OTHER RELEVANT DATA AND INFORMATION ...... 21-1 21.1 Outlook for Rough Diamond Prices ...... 21-1 21.2 Current Contracts ...... 21-3 22.0 INTERPRETATIONS AND CONCLUSIONS ...... 22-1 23.0 RECOMMENDATIONS ...... 23-1 23.1 Further Delineation and Bulk Sampling for UK4 Mine Lower Block ...... 23-1 23.2 Further Delineation and Bulk Sampling for Block Cave 1 ...... 23-1 24.0 REFERENCES ...... 24-1 25.0 CERTIFICATES OF QUALIFICATION ...... 25-1

APPENDIX 1 – GLOSSARY OF TECHNICAL TERMS APPENDIX 2 – DASSAULT GEMCOM BLOCK MODELLING REPORT APPENDIX 3 – MICRODIAMOND DATABASE

LIST OF TABLES Table 5-1 Summary of Lace Diamond Mine Historical Production ...... 5-5 Table 5-2 Summary of 1997 - 1998 Exploration Programme Delineation Drilling ...... 5-9 Table 5-3 Summary of 1998 Tailings Resource Estimates ...... 5-11 Table 5-4 Summary of Lace Diamond Mines’ Operations 2005-2014 ...... 5-13 Table 5-5 Summary of 250m Level Bulk Sampling Results...... 5-18 Table 5-6 Summary of Delineation Drilling ...... 5-21 Table 6-1 Regional Stratigraphy (after Howarth, 2010) ...... 6-2 Table 9-1 Surface and Underground Delineation Drilling Summary ...... 9-1 Table 14-1 Lace Mine Main Pipe Delineation Microdiamond Samples ...... 14-2 Table 14-2a Lace Mine Main Pipe 290m Level Bulk Samples ...... 14-6 Table 14-2b Lace Mine Main Pipe 310m Level Bulk Samples ...... 14-7 Table 14-3 Lace Mine Main Pipe Volume and Tonnage Estimate ...... 14-13 Table 14-4 Lace Mine Main Pipe Raw Density Measurements ...... 14-15 Table 14-5 Summary of Lace Mine Bulk Sample Macrodiamonds ...... 14-22 Table 14-6 Summary of Lace Mine Delineation Microdiamonds with Depth...... 14-23 Table 14-7 Bulk Samples Showing Fine and Coarse Combined Samples ...... 14-27 Table 14-8 Microdiamond Samples Grouped With Respect to Distribution of Diamond Size 14-28 Table 14-9 Bulk Sample Valuation Summary Jan., 2016 ...... 14-40 Table 14-10 Average Diamond Value and Carat Contribution per 100 Carat Parcel ...... 14-42 Table 14-11 Alignment Factors from Total Recovery to +1.25mm Recovery ...... 14-42 Table 14-12 Average Grade and Dollar per carat Value per Lithology and Mine Level ...... 14-43 Table 14-13 Mineral Resource Statement ...... 14-48 Table 15-1 Mineral Reserve Statement ...... 15-2 Table 16-1 Summary of Rock Strength Tests on Kimberlite Core Samples...... 16-2

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Table 16-2 Typical Strength Values of Hypabyssal Kimberlite ...... 16-3 Table 16-3 Summary Statistics and Histograms for Core Competency Measurements ...... 16-5 Table 16-4 Mining Rock Mass Ratings for the Lace Kimberlites ...... 16-9 Table 16-5 Mining Method Comparison for UK4 Mine ...... 16-11 Table 16-6 LOSBU Mining Method Analysis ...... 16-12 Table 16-7 Production Schedule for UK4 Block ...... 16-26 Table 16-8 Labour Requirements for UK4 Mine ...... 16-29 Table 20-1 Capital Cost Summary for UK4 Mine Lift 1 ...... 20-1 Table 20-2 Capital Cost Summary for UK4 Mine Lift 2 ...... 20-2 Table 20-3 Mining Cost Breakdown for UK4 Mine (ZAR per month) at 35,000 tpm ...... 20-2 Table 20-4 Processing Cost Breakdown for UK4 Mine (ZAR per month) at 30,000 tpm ...... 20-3 Table 20-5 Cash Flow Analysis for UK4 Mine ...... 20-4

LIST OF FIGURES Figure 3-1 Location Map ...... 3-2 Figure 3-2 Property and Infrastructure Map...... 3-3 Figure 3-3a Property Map ...... 3-5 Figure 3-3b Property Google Earth Satellite Image ...... 3-6 Figure 5-1 Surface Plan of Lace Mine and Tailings Dumps with 1997-98 Work ...... 5-7 Figure 5-2 3-D Perspective of Lace Mine and Tailings Dumps from 1997-98 Work ...... 5-10 Figure 5-3 Development Work and Bulk Sampling on -55/-60m Levels ...... 5-15 Figure 5-4 Development Work and Bulk Sampling on -250/260m Levels ...... 5-16 Figure 5-5a Lace Mine Development to Dec.18, 2015 ...... 5-19 Figure 5-5b UK4 Mine Development to Dec.18, 2015 ...... 5-20 Figure 6-1 Local Geology ...... 6-3 Figure 6-2 3-D Perspective Lace Main Pipe Geology ...... 6-5 Figure 6-3 3-D Perspective Looking North of Lace Main Pipe Microdiamonds Sampling ...... 6-7 Figure 9-1 3-D Perspective Lace Underground Drilling with All Workings ...... 9-2 Figure 10-1 Bulk Sample Development on 310m Level ...... 10-4 Figure 10-2 Bulk Sample 290-01 Stockpile before Processing ...... 10-5 Figure 10-3 Bulk Sample 290-01 Processing Checklist ...... 10-6 Figure 10-4 Bulk Sample Coarse Tailings Stockpiles...... 10-6 Figure 10-5 DMS/Recovery Tailings Stockpiles ...... 10-7 Figure 14-1 Bulk Sampling on 290m Level ...... 14-8 Figure 14-2 Bulk Sampling on 310m Level ...... 14-9 Figure 14-3 3-D Perspectives Lace Main Pipe Geological Model ...... 14-10 Figure 14-4 Level Plan Compilation Lace Main Pipe Geological Model -250m to -900m ...... 14-11

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Figure 14-5 Section Compilation through Lace Main Pipe Geological Model ...... 14-12 Figure 14-6 3-D Perspective Lace Main Pipe K4 Facies Block Model Facing Northeast ...... 14-15 Figure 14-7 3-D Perspective Lace Main Pipe Density Measurements Facing North ...... 14-16 Figure 14-8 3-D Perspective Lace Main Pipe Dilution 3m Composites Facing North ...... 14-17 Figure 14-9a 3-D Perspective Lace Main Pipe K4 Dilution Blocks Facing Northeast ...... 14-18 Figure 14-9b 3-D Perspective Lace Main Pipe K6 Dilution Blocks Facing Northeast ...... 14-19 Figure 14-10 Log-Probability Graphs for Bulk Samplng by Lithology ...... 14-24 Figure 14-11 Facies Reflecting Fine and Coarse Diamond Size Distributions ...... 14-25 Figure 14-12 Indication of Coarser Size Distribution for K6 South Level 290 ...... 14-26 Figure 14-13 LP-Graphs for Combined Microdiamond Samples by Lithology & Depth Zone 14-28 Figure 14-14 LP-Graphs for Microdiamond Lithological and Depth Groupings ...... 14-29 Figure 14-15 LP-Graphs for Group1, 2, 3 and Group 2 & 3 Combined Populations ...... 14-30 Figure 14-16 LP-Graphs for Group 1 Microdiamonds with Bulk Sample...... 14-32 Figure 14-17 Grade-size Model Based on Group 1 Microdiamonds ...... 14-33 Figure 14-18 LP-Graph for Group 2 Microdiamonds with Bulk Sample ...... 14-35 Figure 14-19 Grade-size Model Based on Group 2 Microdiamonds ...... 14-36 Figure 14-20 LP-Graphs for Group 3 Diamond Recovery ...... 14-37 Figure 14-21 Grade-size Model Based on Group 3 Microdiamonds ...... 14-38 Figure 14-22 Grade-size Graphs for Typical Parcel and Bulk Sample +1.25mm Results ...... 14-39 Figure 14-23 Bulk Sample Diamond Parcel Prior to Sorting ...... 14-41 Figure 14-24 Sensitivity Graphs of Diamond Grade and Revenue ...... 14-45 Figure 14-25a 3-D Perspective Lace Main Pipe K4 Grade Model Facing Northeast ...... 14-46 Figure 14-25b 3-D Perspective Lace Main Pipe K6 Grade Model Facing Northeast ...... 14-47 Figure 15-1 3-D Perspective Lace Main Pipe K4 Grade Model Facing Northeast ...... 15-1 Figure 16-1 3-D Schematic of Trough and Dome Layout for UK4 Mine ...... 16-13 Figure 16-2 3-D Schematic of Development Layout for UK4 Block ...... 16-20 Figure 16-3 Section View of Trough and Dome Layout for UK4 Mine ...... 16-21 Figure 16-4 Mining Progression for UK4 Mine Lift 1 ...... 16-22 Figure 16-5 360m Level Plan showing UK4 Mine Lower Block Production Lay-out ...... 16-24 Figure 17-1 Lace Mine Process Flowsheet ...... 17-1 Figure 20-1 Sensitivity Analysis for UK4 Mine ...... 20-5

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

SUMMARY

Introduction At the request of the Board of Directors of DiamondCorp plc (“DiamondCorp”), the authors, Mr. Paul Allan of Durban, South Africa, Mr. Paul Sobie of Toronto, Canada and Dr. Johan Ferreira of Wells, United Kingdom (collectively “the independent authors”, with Mr. Sobie and Dr. Ferreira the “Competent Persons”) have completed an independent technical report in collaboration with senior staff personnel on the Lace Diamond Mine, Free State, Republic of South Africa (“LDM”). Mr. Paul Loudon of DiamondCorp has provided the life of mine economic model and contributed to the economic analysis section of this report. LDM’s mining team and mining consultants have provided the mining sections of this report, updated from their 2014 scoping study.

The Lace Diamond Mine is an active underground development operation, fully permitted and already in possession of key components including all manner of mining infrastructure, access, power and water supply. The present report deals primarily with the resource slice between -230m and -370m, designated the Upper K4 Mine Block, which has been identified as an early addition to the mine plan. Since early in 2014, the authors in collaboration with LDM management have designed and implemented the evaluation program while appropriate mining methods were investigated, and the development of the UK4 Mine infrastructure has been taking place.

As of the effective date of this report, February 1st, 2016, LDM is approximately four months from production ramp-up for full-scale UK4 Mine operations.

Property and Agreements The Lace project is held by Lace Diamond Mines (Pty) Ltd. (“LDM”), a company owned 74% by DiamondCorp Holdings Limited, 13% by Shanduka Resources (Pty) Ltd and 13% by Sphere Investments (Pty) Ltd. DiamondCorp Holdings is a wholly owned subsidiary of DiamondCorp plc. Shanduka and Sphere are corporations owned and controlled by historically disadvantaged South Africans as required by the Black Economic Empowerment Charter as it applies to the South African mining and minerals industry.

The Lace Mine property consists of 239.97 ha comprising Portion 1 of Farm Ruby 691 and a portion of the remainder of Farm Ruby 691 upon which LDM owns surface and sub-surface rights (Figure 1). The mine, tailings dumps and new slimes deposits are all within Subdivision 1. The main Lace Mine property comprises a New Order Mining Right over Portion 1 and a portion of the remainder of Farm Ruby 691 totalling 239.97 ha. The mine, tailings dumps and plant site are all located within Portion 1, and are included within the surface rights.

The Lace Diamond Mine property is situated approximately 200km southwest of Johannesburg, the main South African centre of transportation, commerce and industry, in the Free State Province. The property lies less than 20km off the N1 national highway, some 20km along a secondary highway northwest of the city of Kroonstad. Local access is by a 1km all-weather gravel road from the secondary highway, and thence a network of gravel roads on the property itself.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 1 Property Map

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

The nearest population centre is Kroonstad, 20 km southeast of the mine. Kroonstad is a city of 100,000 people, which provides all manner of support, including a labour force familiar with mining as many of the Free State gold mines as well as several small diamond fissure mines are some 40 km south in the Welkom area. Established mining centres are located in Klerksdorp and Welkom, all within a 100km radius of the mine. Major engineering services and mining equipment suppliers are located in Johannesburg. LDM buses their staff to and from the mine, or individual employees make their own way.

Lace is located at an elevation of 1,375m and the climate is considered local semi-arid cold steppe climate (BSk) in accordance with the Köppen-Geiger climate classification system. The temperature averages 16.6°C and the rainfall averages 604 mm per annum. The area surrounding the mine is primarily utilized for cattle ranching.

History The Lace Mine, also known as the Crown Mine, is a past-producer which operated intermittently between 1902 and 1932 after being discovered in 1896, and was purchased by De Beers after shutting down. A fairly complete record of production during that time is available from the records of the Crown Diamond Mining and Exploration Company Limited, and from various summaries prepared by De Beers personnel. In all, the Lace Mine is estimated by De Beers to have processed 4,947,203 tonnes during its evaluation and production history, yielding 727,675 carats. The grade of this material is 14.7cpht as calculated by De Beers.

The records indicate that some ~1.8Mt were mined from underground during the last four years at a grade of ~18.9cpht, as compared to the ~3.1Mt that were extracted primarily from the open pit at 12.3cpht. Historical records indicate that ore was extracted from the 105m level to the 212m by partial chambering, using first the 165m haulage level, then the 240m level. No exploitation below -240m was carried out, although some development work was. All of the historical level plans show that kimberlite was not mined to the margins of the pipe, even within portions of the open pit where ~30m of very dilute remnant material was mapped and sampled by MPH in historical adits on the north pit wall on the -30m level.

De Beers Period 1931-1996 “The Crown Mine was kept pumped out, and the machinery maintained until 1939, when the mine was auctioned off and sold to De Beers Consolidated Mines Ltd (Rossouw, 1997). Eventually maintenance of the mine was abandoned, the main shaft collapsed and the open cut was flooded. The project was shelved until March 1997, when it was sold to the Christian Potgieter Trust.” (Henderson, 1997)

Christian Potgeiter Trust 1996-2005 The author (Sobie) was contacted early in 1997 by representatives of CPT (the Christian Potgieter Trust), to elicit interest in the Lace tailings dumps for clients. Upon examining the dumps, and noticing large blocks of unprocessed, high-interest hypabyssal kimberlite, MPH sent several samples of this material in for microdiamond analysis, which were positive and suggestive of high-grade. MPH believed that both the kimberlite resource and

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tailings dump potential were of high interest, and attracted a Canadian junior mining company to the project, Rupert Resources Ltd. (“Rupert”), which negotiated a phased acquisition deal with CPT in 1997.

MPH carried out an extensive due-diligence bulk-sampling programme on the Lace Diamond Mine tailings dumps in 1997-98 on behalf of Rupert, which led to a positive feasibility study and Detailed Cost Estimate in 2000, all managed by the author and experienced diamond evaluation geologist, Mr. Tim Wilkes. Reserves and resources estimated at that time based on the positive study were as per Table 1 below, with the diamonds valued independently at $43/ct.

Table 1 - Tailings Re-treatment Project Resource Statement

Resource Category Tonnes (000’s) % Tonnes Grade (cpht) Proven Reserve 2 853.2 72 10.7 Probable Reserves 742.3 19 8.7 Total Reserves 3 595.5 91 *10.3 Inferred Resource 390.8 9 4.3 *weighted average. +1mm bottom cut-off screen size

MPH also carried out a 1997-98 kimberlite exploration programme at Lace, the first modern work ever carried out, the purpose of which was to begin to accurately delineate the kimberlites below the workings, and to gain an initial feeling as to their economic prospects. Mr. Paul Allan, an experienced kimberlite geologist, managed all of the core work on that project. The delineation and microdiamond sampling program was successful, and led to initial scoping studies with SRK Consulting Ltd. (“SRK”), for a conceptual block cave operation as well as an underground bulk sampling program utilizing a refurbished historical shaft for access (MPH, 1998c).

Rupert was unable to raise capital for the project in the depressed resource markets of 2000- 2001, and it reverted to the vendor, CPT. The project was dormant between 2001-2005 as CPT marketed the property to various diamond mining companies and ventures, with MPH variably providing technical data to the interested parties.

DiamondCorp plc 2005-2014 The 2000 tailings dump work was updated for DiamondCorp to 2006 (MPH and Crosston Technologies, 2006a), and along with the updated report on the potential of Lace underground kimberlite resource (MPH, 2005), formed the basis for capital raising and listing in late 2006. DiamondCorp also embarked on the historical shaft refurbishment programme recommended, as a first step to commencing underground kimberlite evaluation. Circumstances have dictated that DiamondCorp’s path to the present underground evaluation program and designing/commissioning of the UK4 Mine has been less direct than one would have hoped in 2007. A number of events have caused changes in plans including:

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i) Despite attaining steady-state tailings dump production at 120tph with their purpose-built plant in late 2007, the world-wide financial crisis in 2008 saw diamond prices drop precipitously, to the point that the tailings reprocessing operation was unprofitable and shut down in late 2008. ii) Shaft refurbishment was found to be impossible below -160m due to unstable ground conditions, and a decision was therefore made by LDM in 2008 to sink a decline ramp to access shallow Satellite and Main Pipe material left behind by historical miners. The shaft was later joined with the new underground infrastructure to provide ventilation. The plant was modified to be able to process fresh kimberlite material in addition to the tailings. iii) The remnant kimberlite at the -60m target level in both the Satellite and Main pipes was found to be heavily dilute (and unstable in the Main Pipe), and the decision was made in 2009 to continue the decline to the -250m level, below all historical production stopes. iv) Weakness in the capital markets in 2011-2012 made it difficult to advance the project following successful Main Pipe bulk sampling on the 250m level, however SRK was commissioned to design a higher confidence block cave mine. Their design included conveyor systems and underground crushing, with new twin declines from surface to the -470m level. v) Capital availability in 2013 saw the commencement of decline development and late in the year an underground core drilling rig was purchased, and installed on the 250m level in early 2014. vi) The delineation drilling initiated in early 2014, combined with detailed mapping of the 250m bulk sampling level allowed for MPH to identify the upper K4 resource potential, previously only hinted at from the shallowest 1998 borehole, Crown-7.

Tailings Reprocessing and Plant Construction LDM commissioned Consultmet (Pty.) Limited to design and construct the Lace Mine tailings retreatment plant in 2006, which was completed and commissioned in late 2007. The plant was designed to treat 1.6 million tonnes per annum, at a 220 tonnes per hour (“tph”) head feed. The plant utilizes a large double screening front end, 100 tph (coarse +6mm fraction) and 65 tph (fine -6+1mm fraction) DMS modules, a re-crush circuit and hands free grease belt and glove box recovery units for diamond recovery.

The company purchased its own fleet of Bell mobile mining equipment in 2007, and commenced site preparations, as well as the construction of a new tailings and slimes dam.

Production has been discontinuous, and has totalled some 1.58 million tonnes processed, recovering 101,904 carats for a grade of 6.45cpht, and a value per carat of $58.70. Problems with the re-crush circuit during 2007-2008, since rectified, were felt to be a major reason for the lower recovered grades during that first, major period of tailings retreatment.

2009 Satellite and Main Pipe Bulk Sampling With the tailings resource proving to be marginal in late 2008, the decision was made to put that operation on care and maintenance, whilst modifying the plant to handle kimberlite

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and proceeding with the 4m x 4m decline. A key development was the granting of the mining right for the Lace Underground Mine by the Department of Minerals and Energy (“DME”) in February 2009. Also the Company purchased second hand underground mining fleet equipment to carry out the development work in-house.

The decline accessed the Satellite Pipe and a remnant lens of heavily dilute K6 kimberlite material on the north edge of the open pit, on the -55m and -60m levels respectively. The Satellite Pipe kimberlite was also heavily dilute K6, as the 1997-98 drilling by MPH had intersected, which was postulated as the reason the early miners had chosen not to exploit this resource. In general it was felt that Satellite Pipe grades must have been significantly less than Main Pipe grades at the same elevations, otherwise it would have been mined historically.

The Satellite Pipe bulk sampling operation was quite successful in recovering some 808.21 carats from 11,380tonnes of K6 material, for a recovered grade of 7.1cpht. The parcel was valued at $108/ct. in 2009 at a tender conducted by Sadiamex (Pty) Limited in Johannesburg, an encouraging value in depressed diamond market conditions at the time. The Main Pipe attempted bulk sample encountered unstable conditions and no bulk sampling was completed.

2011 Main Pipe Bulk Sampling DiamondCorp spent 2010 and early 2011 advancing the decline some 1,800m to the -250m level, deciding to get beneath the lowest historical stopes which reached the -238m level. Kimberlite was reached on 10 May, 2011. Although plans had been for a systematic bulk sampling program comprising six 3m by 3m development drives across the entire kimberlite, poor ground conditions necessitated a more modest approach. The development drives encountered numerous historical tunnels and winzes, resulting in significant caving, as attempts were made to access the central portions of the pipe. The final orientation of bulk sampling drives, and individual samples is shown below in Figure 2 along with up-to-date geology and drilling. Results were very encouraging, with 2,157.41 carats recovered from 15,414 tonnes processed, and the diamond parcel valued at $160/ct. at that time.

None of the samples were limited to a single kimberlite facies type although K6 dominated, as they were oriented along major K4/K6 contacts in many cases. Results therefore were generally somewhat higher grading than expected for pure K6 material, with the plant using a lower bottom screen (+1.00mm) for this work also contributing to the higher grades. Prior to the plant upgrade in 2013, the plant had 0.5 mm dewatering screens on the feed preparation section of the plant, resulting in any carry over of small diamonds on the 1.00 mm trammel of the primary scrubber (as was common because the 1.00 mm panels easily blind) would report to the dense media separation (“DMS”) circuit and final recovery. The plant currently has 1.00 mm de-watering screens, so those small diamonds now report to coarse tails.

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2012-2014 Development and Underground Delineation Drilling Following on capital raising in 2012, DiamondCorp initiated the recommended twin decline infrastructure development from the -92m level, working both upwards toward the portal, and downwards toward the existing decline. The SRK report (SRK, 2012) recommended a continuous trough block cave on the -470m level, serviced by twin declines, with ore and waste material conveyed to surface. Critical to firming up much of the engineering work was a far better understanding of the internal geology and morphology of the pipe, and the company purchased an underground core rig and set it up initially on the 250m bulk sampling level.

The identification of the UK4 resource in 2014 necessitated redeploying equipment and personnel to construct the UK4 Mine underground infrastructure, described in detail in Section 16 of this report.

Figure 2a and 2b show the completed development work in orange as of the effective date of this report, as well as completed drill holes. Figure 2a shows all workings from surface to the bottom of the known geological model, planned to be mined by successive block caves (Block Caves 1, 2 and 3). Figure 2b is a more detailed rendition of the UK4 Mine developments and better shows the underground drill cubbies. For both figures, the gold coloured workings have been completed as of the report date.

Drilling on the project now totals 10,183.88 metres, of which 29 holes totalling 4,422m have been drilled from four drill stations or “cubbies”, underground. These have been established to best meet both geological and mining development objectives, and as a summary statement have allowed for concise modelling of the UK4 resource slice portion, of the overall Lace deposit.

Interestingly, the historical “Bulge” has now been located and defined between -310 and - 400m and has thus far been found to be limited to dilute K6 and CRB facies.

Geology The Lace Mine lies within the south central Archean Kaapvaal Craton within a subcluster of kimberlites referred to as the Kroonstad Group II Kimberlite Cluster (Howarth, 2010) of which only Lace and nearby Voorspoed have seen formal mining. Four other small kimberlite blows and several dykes are known within the cluster. The Lace kimberlites have been dated at 133.2 +/- 2.8 Ma by the 40Ar/39Ar technique on ground mass phologopite grains (Philips et al., 1999), similar to the early Cretaceous ages of the other Kroonstad kimberlites and to other Group II kimberlites in South Africa including Finsch.

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Figure 2 Geological Model and Lace Mine Development to Dec.18, 2015

The delineation drilling program has shown the geology of Lace to be relatively simple (Figure 2 above), with five main internal facies, within an irregular pipe-shaped body that varies in areal extent from ~1.7ha to ~2.6ha within the well-constrained -230m to -370m UK4 Mine depth slice. The delineation drilling to date has shown that a central SW-NE oriented “ridge” of K4/K2 coherent kimberlites (segregationary/massive hypabyssal kimberlites respectively) occupies the Lace Main Pipe, from the 230m level to the deepest intersections at ~875m, gradually flaring out to occupy greater portions of the pipe, with depth. This coherent kimberlite ridge is flanked in both the northwestern and southeastern quadrants by dilute K6 volcaniclastic material, generally with sharp contacts, which in turn grades at the margins to heavily dilute CRB units.

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Within the coherent kimberlite ridge, several “pods” of K8 basaltic kimberlite breccia have been intersected above the 350m level that range up to 30m in diameter and likely represent basaltic country rock material assimilated early in the intrusion history of Lace. Within the southeastern K6 unit, several large rafts of country rock up to 20m in width have been intersected to date, of both shale and lava, depending on depth. Above ~290m, intercalations of K6 with K4/K2 are evident and become more common upwards, suggesting this depth marks the zone of breaching of the K4/K2 intrusion. Generally the contacts of K4 with K6 are marked by a sharp decrease in country rock xenoliths and a narrow zone of flow banding.

The pipe morphology between 230m and 370m depth has been found to be complex, with the margins quite irregular, as now understood from the detailed drilling. Beneath 370m pipe shape is not well constrained from the few deep holes, however we expect that the irregular margins will continue as is typical of kimberlites at this relative exposure level, ie. Kimberley, and the model will be updated periodically going forward.

The Main Pipe The historical production records, modern microdiamond and bulk sampling have established that commercial grades and diamond values are present for the Main Pipe. The microdiamond sampling data for the project totals 168 samples weighing 3,615kg, recovering 5,397 diamonds. As a general statement, the microdiamond sampling has demonstrated continuously mineralized K4, K6 and K8 facies throughout the present geological model, with grade modelling, detailed in Section 14 of this report, providing high-confidence estimates of commercially interesting grades. The CRB facies has been shown to be significantly diamondiferous as well, but has not been sufficiently micro or bulk sampled thus far to estimate grade.

Bulk sampling on the 250m, 290m and 310m levels, also detailed in Section 14, has confirmed the grade modelling. Bulk sampling totalled 11,241 tonnes on 290m level, and 8,479 tonnes on 310m level, recovering 2,270.44cts and 554.96cts respectively, in addition to the 15,414 tonnes and 2,157.44cts recovered from the 250m level in 2011.

Commercial Potential Given its production history from 1896-1931, it is not surprising that Lace again is on the cusp of commercial production. The modern work on the project is now allowing for a better understanding of what was mined historically, and what future mining holds at depth. It is important to note however that the historical underground grade of 18.9cpht, which combined with the tailings dump grades achieved by DiamondCorp to date (6.45cpht which understates the dump grades somewhat), was commercially viable until the Great Depression. This in turn suggests that a considerable portion of higher grading K4 was exploited along with the lower grading K6 facies above -240m. The delineation drilling and mapping work, and derived geological modelling, has now mapped the K4 “ridge” to -230m, and it is reasonable to assume it reached higher levels. K4 becomes volumetrically more dominant with depth, emphasizing the commercial potential of the deeper underground mine.

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Mineral Resource and Reserve Estimates Geological, dilution, density and mineralization (grade) block models have been constructed for the UK4 Mine Block depth slice (-230m to -370m) of the Lace Main Pipe, and the updated geological interpretation extended to the base of the present model at -920m (410m absl). The grade model has been found to show high precision in predicting ultimate recovered grades achieved in the bulk sampling on -250m, -290m and -310m levels.

Mineral resource estimates have been prepared based on the new geological, dilution, density and mineralization data per Table 2 below. In particular, Lace has been demonstrated as a reliable microdiamond producer allowing for high-confidence grade estimates, and the valuation data from 4,982 carats recovered during bulk sampling similarly gives confidence to the valuation model

There is compelling geological evidence that K4 and K6 grade will improve within the block cave mining depths (and Lift 2 of the UK4 Mine), however much more evaluation work is needed to verify these trends. The microdiamond database is still small for this very large portion (~75%) of the deposit, and all of the caustic dissolution work is from 1997, such that a great deal of modern work is needed to allow for higher precision estimates.

Additionally, the CRB unit will be bulk and microdiamond sampled and will be assigned a grade as these data become available, such that carat content at Lace is highly likely to improve on these estimates.

At this time we advocate uniformity of grade estimates for all three block cave depth slices until such time as larger databases are available. Undoubtedly volumes and tonnages will also change as more delineation work is completed.

Similarly, the combined $/carat value of USD164/ct. is regarded as the best value for the overall Lace parcel at this time. It does not take into account the significant upside that can be achieved from the recovery of high value “special” stones, for which Lace was known during its previous production period pre-Great Depression, with historical recovery of high quality stones up to 122 carats in size.

When comparing to the previous mineral resource estimates prepared for the Lace mine, it is important to note that the estimated recovered grade for all facies has now been aligned with the Lace production plant utilizing 1.25mm slotted screens, compared to 1.00mm screens in its original design configuration. While the increase in bottom screen sizes means that recoveries will be lower in terms of grade, the carat value will increase as the smallest diamonds no longer being recovered are also the lowest value diamonds. So while the impact on recovered grade resulting from the change can be perceived as significant, the economic impact is minimal. The new resource statement should be considered a conservative base case from which there is compelling evidence that considerable grade and value per carat upside is likely, but still to be further refined with additional production and evaluation data.

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Table 2 Mineral Resource Statement

RESOURCE KIMBERLITE VOLUME % OF RECOVERED MINING BLOCK DENSITY TONNES CARATS USD/carat CLASSIFICATION FACIES (m3x1000) TOTAL GRADE (cpt) Upper K4 Mine 230-370m Levels Indicated K4 1,065.486 2.585 2,754,281 36.9% 0.365 1,005,313 $ 164.00 (1110-960m absl) Indicated K6 1,834.957 2.563 4,702,995 63.1% 0.090 422,329 $ 164.00 Total Indicated 2,900.443 7,457,276 100.0% 0.191 1,427,642 $ 164.00 Inferred K8 144.722 2.641 382,211 16.3% 0.160 61,154 $ 164.00 Inferred CRB 723.803 2.709 1,960,782 83.7% 0.000 - $ 164.00 Total Inferred 868.525 2,342,993 100.0% 0.026 61,154 $ 164.00

Block Cave 1 370-510m Levels Inferred K4 1,626.754 2.59 4,213,293 48.4% 0.400 1,685,317 $ 164.00 (960-820m) Inferred K6 1,262.561 2.56 3,232,157 37.1% 0.100 323,216 $ 164.00 Inferred K8 13.713 2.64 36,203 0.4% 0.160 5,793 $ 164.00 Inferred CRB 451.786 2.71 1,224,339 14.1% 0.000 - $ 164.00 Total 3,354.815 8,705,993 100.0% 0.231 2,014,326 $ 164.00 Block Cave 2 510-700m Levels Inferred K4 2,225.776 2.59 5,764,760 59.0% 0.400 2,305,904 $ 164.00 (820-630m) Inferred K6 1,484.048 2.56 3,799,164 38.9% 0.100 379,916 $ 164.00 Inferred K8 0.000 2.64 - 0.0% 0.160 - $ 164.00 Inferred CRB 74.018 2.71 200,589 2.1% 0.000 - $ 164.00 Total 3,783.842 9,764,513 100.0% 0.275 2,685,820 $ 164.00 Block Cave 3 700-920m Levels Inferred K4 2,800.965 2.59 7,254,499 71.0% 0.400 2,901,799 $ 164.00 (630-410m) Inferred K6 1,153.812 2.56 2,953,759 28.9% 0.100 295,376 $ 164.00 Inferred K8 0.000 2.64 - 0.0% 0.160 - $ 164.00 Inferred CRB 3.577 2.71 9,694 0.1% 0.000 - $ 164.00 Total 3,958.354 10,217,952 100.0% 0.313 3,197,175 $ 164.00 Block Caves 1,2,3 Inferred K4 6,653.495 2.59 17,232,552 60.1% 0.400 6,893,021 $ 164.00 -370m to -920m Levels Inferred K6 3,900.422 2.56 9,985,081 34.8% 0.100 998,508 $ 164.00 (960m-410m absl) Inferred K8 13.713 2.64 36,203 0.1% 0.160 5,793 $ 164.00 Inferred CRB 529.381 2.71 1,434,622 5.0% 0.000 - $ 164.00 Total Inferred 11,097.011 28,688,458 100.0% 0.275 7,897,321 $ 164.00

Lace Mine Totals Indicated 2,900.443 2.57 7,457,276 19.4% 0.191 1,427,642 $ 164.00  Based -230m on to a -920m recoverable Levels gradeInferred model for the current 11,965.54 Lace plant 2.59configuration 31,031,451 (+1.25mm80.6% bottom 0.256 cut-off 7,958,475screen size)$ 164.00 (1110m-410m absl) Totals 14,865.979 2.59 38,488,727 100.0% 0.244 9,386,117 $ 164.00  Diamond price based on bulk sample parcels and January 2016 price book  Effective date February 1st, 2016.

Reserves for the UK4 mining block have been estimated based on the detailed mining plan prepared by LDM, which should be noted is being updated constantly based on real costs incurred during development. Trade off studies by LDM have indicated that the optimum bottom cut-off screen size for the project should be +1.25mm with a de-grit circuit, as per the present configuration of the Lace production plant.

Grade control will be on-going, as blocks that include both K6 and K4 material will be mined at both the north and south ends of the open stopes, and some or all of the K6 material may be stockpiled on surface. The interior blocks will be entirely of K4 facies and will be processed in their entirety. The UK4 Mine is targeting approximately 2million tonnes of the ~7.5million tonne indicated resource. There may well be opportunities to optimize the mine plan as more K4 is found to be present as the mine progresses.

The reserve statement for the UK4 mining operation is provided below in Table 3.

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Table 3 Mineral Reserve Statement

Kimberlite Grade UK 4 Mine Classification Tonnes Carats USD/carat USD/tonne Facies (cpt) 250-370m Levels Probable K4 1,427,841 0.362 516,575 164.00$ 59.33$ (1110-960m absl) Probable K6 782,244 0.090 70,296 164.00$ 14.74$ Total Probable 2,210,086 0.266 586,870 164.00$ 43.55$  Based on a recoverable grade model for the current Lace plant configuration (+1.25mm bottom cut-off screen size)  Diamond price based on bulk sample parcels and January 2016 price book  Rounding has been applied  Plant recovery 100% of recoverable grade, mining recovery 100%  Effective date February 1st, 2016.

Mining Method DiamondCorp’s mining team and consultants spent a great deal of time in 2014 investigating mining method alternatives for the UK4 mining block at Lace, to bring forward cash flow by exploiting it ahead of Block Cave 1. The team, which includes underground kimberlite mining specialists Bob Harverson and Pat Bartlett, prepared a scoping study (DiamondCorp PLC, 2014), updated here-in. The report concluded that a continuous trough type doming mining method incorporating a sub level open stope mined from the bottom, best suits mining of the UK4 block. The greatest risk of mining any block close to old workings is the potential risk of inundation of water and mud from the old workings. The chosen mining method – called Longhole Open Stoping Bottom Up (LOSBU) - meets the requirements of the SIMRAC report on the guidelines to mitigate potential risk of mud rushes.

The LDM UK4 Upper Lift 1 block design layout comprises two main levels, namely a doming level (290m level) and the main production level (310m level) as shown schematically in Figure 3. Two open stopes are positioned to extract the required ore tonnage from the block. Two continuous troughs will be situated for ore collection below the Open stopes. The doming level will be situated on 290m level to enable doming long hole rings to blast ore down into the collection troughs. This ore will be loaded from draw points spaced along the troughs. One stope will be mined initially and once this has broken through to the old workings and the mud and water is dissipated and drawn down the second stope will be mined. Lift 2 on the 360m level will be similar, with the ultimate layout slightly larger.

Economic Analysis The economic analysis of mining the UK4 Block at the Lace Mine has been conducted by the discounted cash flow method. The projected mine cash flows have been modelled flat real based on actual costs capital already and currently being incurred, and revenue estimates as at January 2016.

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Figure 3 UK4 Mine Lift 1

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The UK4 Block has the potential to be mined at a rate of 35,000 tonnes per month for at least 60 months. It will be mined until the first block cave begins to propagate. Once this occurs, mining will relocate to the 500m level and any ore remaining in the UK4 Block will subsequently fall into the first cave. For modelling purposes, it is assumed the Block is mined for 60 months.

The cash flows have been modelled flat real (i.e. uninflated) with a 10% discount rate applied, and major parameters applied from LDM’s actual costs as follows:

USD:ZAR Exchange Rate 1:15.0

Mining Costs/Tonne ZAR 193.60

Processing Costs/Tonne ZAR 52.72

Average Recovered Grade 23cpht

Carat Value ($/carat) 164.00

The model generates an NPV of R133.3 million (US$8.9 million) and a robust IRR of 59%.

The UK4 Mine is most sensitive to grade, carat value and exchange rates. The mine is less sensitive to increases in mining and processing costs.

Interpretation and Conclusions The presently reported drilling and sampling programs within the UK4 Mine block have allowed for increased understanding and confidence in the geology, volume, tonnage, and mineralization within this depth slice of the overall Lace Main Pipe. This in turn has allowed for updated, higher confidence Mineral Resource estimates within that portion of the pipe, and greater confidence in interpretations of these same resource parameters at greater depth. The present methodology of systematic delineation core drilling from underground drill stations, and audited bulk sampling of development drives, is proving effective. As the Lace production plant is being utilized, the recovered grades from the bulk sampling represent true, non-factorized, commercially attained results based on the present plant flowsheet, and have been utilized for all resource estimates.

From a commercial perspective, the UK4 Mine plan which includes a portion of the indicated mineral resource, is deemed economically viable and robust based on real cost parameters, determined during DiamondCorp’s development period, still in progress and within four months of attaining name-plate production. The mineral resource for the UK4 Mine block along with these cost and revenue parameters is the basis for the first industry-standard Probable Mineral Reserve at Lace, dedicated solely at present to the UK4 Mine operation.

Further evaluation work of the same types will allow for progressively deeper depth slices of Lace to be moved into the Indicated Resource category, and thereafter into probable reserves for the pending, larger block cave mining operations.

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Recommendations It is recommended that LDM and DiamondCorp continue the present evaluation program, which can be improved upon slightly to increase precision. Specifically the usage of digital borehole directional survey equipment will assist the modelling by bringing an added measure of precision to the plotting of all geological and sampling borehole data. On-going delineation drilling, although primarily designed to provide information from below the -370m level, will serve to improve the geological knowledge and sampling database for the lower lift of the UK4 Mine depth slice, as the holes will be collared at either the 290 or 310m levels. Continued assimilation of this data is recommended.

Similarly the CP’s recommend that development drives for UK 4 Lift 2 on the -340m doming level, and the -360m production levels, be systematically bulk and microdiamond sampled as has been documented here-in on the Lift 1 development drives. As well, it is essential to attain micro and bulk sampling data into the CRB units on the northern quadrant of the pipe such that this large facies can be better understood and a reliable grade estimate ascribed to it. Deeper drilling has indicated that the CRB unit grades into K6 facies, with better grade potential.

Finally, continued detailed geological mapping, combined with resource management and grade reconciliation work on the UK4 Lift 1 operation will allow for updating and improving upon these initial resource and reserve estimates.

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1.0 INTRODUCTION

At the request of the Board of Directors of DiamondCorp plc (“DiamondCorp”), the authors, Mr. Paul Allan of Durban, South Africa, Mr. Paul Sobie of Toronto, Canada and Dr. Johan Ferreira of Wells, United Kingdom (collectively “the independent authors”, with Mr. Sobie and Dr. Ferreira the “Competent Persons”) have completed an independent technical report in collaboration with senior staff personnel on the Lace Diamond Mine, Free State, Republic of South Africa (“LDM”). Mr. Paul Loudon of DiamondCorp has provided the life of mine economic model and contributed to the economic analysis section of this report.

The Lace Diamond Mine is an active underground development operation, already in possession of key components including all manner of mining infrastructure, access, power and water supply. The present report deals primarily with the resource slice between -230m and -370m, designated the Upper K4 Mine Block, which has been identified as an early addition to the mine plan. Since early in 2014, the authors in collaboration with LDM management, have designed and implemented the evaluation program while appropriate mining methods were investigated, and the development of the UK4 Mine infrastructure has been taking place.

As of the date of this report, LDM is approximately four months from production ramp-up for full- scale UK4 Mine operations.

1.1 Authorization and Terms of Reference

DiamondCorp and subsidiary Lace Diamond Mines (“LDM”) retained the authors on October 1st, 2015, to prepare an Independent Technical Report to conform to SAMREC and all international mining code standards. The report was prepared in RSA, UK and Toronto, Canada during November 2015 through February 2016. All of the independent authors have been actively consulting to LDM and DiamondCorp since early 2014, and additionally have had extensive prior involvement with the project. LDM staff contributions were primarily made by Steve West, Chief Operations Officer, Andre Labuschagne, Mine Manager, and Paul Loudon, Chief Executive Officer.

1.2 Qualifications of Authors

Mr. Paul Allan, B.Sc. Hons., has over 25 years of experience in multi-commodity geology, predominantly in precious stones exploration and evaluation. Initial experience began with Anglo American Research Laboratory / De Beers in kimberlite mineral chemistry and petrography as well as Witwatersrand type gold research. Mr. Allan was project geologist for MPH during the exploration/evaluations at Lace during 1997-2000, while also lecturing in geology at the University of Fort Hare. From 2001-2005 he was employed by SouthernEra Diamonds Inc. where he became the Regional Exploration Manager for Southern Africa in 2005. In 2008 Mr. Allan joined Firestone Diamonds plc as Project Geologist (Modeling) and as Senior Project Geologist which included responsibility for the geological exploration and evaluation conducted on Firestone Botswana’s Projects. Paul also worked on Firestone’s Liqhobong Diamond Mine, before

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joining Gemfields in Mid-2012. Mr. Allan re-joined MPH and the Lace Project in early 2014 and has managed all drilling and bulk sampling described in this report.

Dr Johan Ferreira is a professional Geostatistician with over 30 years’ experience in the geostatistical modelling of diamond deposits worldwide, with 26 years’ experience with De Beers before becoming a private consultant. Dr Ferreira obtained his PhD titled “Sampling and Evaluation of Kimberlites Based on Microdiamond Sampling”, at the Ecole Des Mines. He is a member of the Canadian Institute of Mining, Metallurgy and Petroleum and is registered with SACNASP as a Professional Natural Scientist. Through a combination of his qualification, experience and professional registration Johan is a competent person according to the SAMREC code, and has consulted to Lace Diamond Mines since 2012.

Mr. Sobie, B.Sc., P.Geo. is a professional geologist with 28 years of diamond exploration, evaluation and development experience in Southern and West Africa, Canada and Russia. Mr. Sobie is president of MPH Consulting Limited, an international exploration and mining consultancy based in Toronto, and a non-executive director of Firestone Diamond plc, which is developing the Liqhobong diamond mine in Lesotho. Mr. Sobie has spent a considerable amount of his career focused on diamonds in Southern Africa, and led the team that identified and explored the Lace Diamond Mine in 1997, as well as evaluated the Lace tailings deposits, and also the Liqhobong kimberlites in Lesotho. Additionally Mr. Sobie authored the competent persons’ reports for DiamondCorp’s successful listings on both the AIM Exchange in London in 2006, and the Johannesburg Stock Exchange in 2007.

Mr. Sobie and Dr. Ferreira are the independent competent persons for this report.

1.3 Scope of Work and Sources of Information

DiamondCorp commissioned the authors to prepare the Technical Report on the Property and develop an ongoing evaluation program to allow for continuous updating of deeper resources to higher confidence levels going forward. As such this report is regarded as the first of a series, which will be updated as the continual evaluation work allows for resource and reserve updates for the next mining block, Block Cave 1 which will draw from the -500m level.

In preparing this report, the authors reviewed geological reports and maps, miscellaneous technical papers, company reports and announcements, memoranda and other public and private information as listed in the “Reference” section of this report.

The report is based in part on personal geological observations and supervision of work by Messers Allan and Sobie, together with expert geostatistical input from Dr. Ferreira. Extensive input for the Mining Method, Recovery Methods, Environmental Considerations, Project Infrastructure and Economic Analysis Sections comes from

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Messers Labuschagne, Mine Manager, and Loudon, Chief Executive Officer, as Lace is actively developing the mine as described here-in.

This report is based on geological, mine development and sampling information known to the authors as of January 13th, 2016.

Unless otherwise noted, all measurement units used in this report are metric, and currency is expressed in United States Dollars.

All co-ordinates utilized in this report are utilizing the South African Coordinate Reference System (SACRS) including the Hartesbeesthoek94 Datum (WG 27°).

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2.0 RELIANCE ON OTHER EXPERTS

The authors assumed that all of the information and technical documents reviewed and listed in the “References” are accurate and complete in all material aspects. While the authors carefully reviewed all of this information, and they have been working with this database and project since 1997 in some cases, it is not possible in all cases to verify its accuracy and completeness. Original reports from all analytical facilities that carried out work for the project have been acquired to ensure valid data is being utilized.

Diamond value data reported in Section 14.7 were provided either by Mr. Riaan Gloy of Distinctive Choice 1235, of Johannesburg, or by Mr. Samuel Scheffer, or Natural Diamond Corporation NV of Antwerp, Belgium. Both companies have been involved with the sale of Lace production since the tailings reprocessing operation commenced. The competent persons have relied on these data for the basis of the revenue estimates in Section 14.7.

The information and conclusions contained herein are based on the information available to the authors at the time of preparation of this report with assumptions, conditions and qualifications as set forth in the Report and data listed in the “References”.

DiamondCorp has reviewed draft copies of the report for factual errors. Any changes made as a result of these reviews did not involve any alteration to the conclusions made. Hence, the statement and opinions expressed in this document are given in good faith and in the belief that such statements and opinions are not false and misleading at the date of this report.

The authors reserve the right, but will not be obligated to, revise this report and conclusions thereto if additional information becomes known to them subsequent to the date of this report.

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3.0 PROPERTY DESCRIPTION AND LOCATION

3.1 Mineral Policy South Africa

New Order mineral rights were introduced by the South African Department of Minerals and Energy effective from 1 May 2005. In essence, the new regulations mean that surface right holders no longer have a prior right to lodge a claim for mineral rights on their land. The new law means that any individual or company has the right to lodge an application to acquire mineral rights regardless of who owns the surface rights.

Under the previous mining regime, it was possible for landowners to lodge an application for the mineral rights on their land and thereafter effectively sit on the rights preventing any other individual from conducting exploration or mining.

A period of time was granted for existing mineral rights holders to convert their rights to the New Order rights. The application had to be accompanied by a work programme, budget and schedule. The Department of Minerals and Energy has the right to check that the licence holder is conducting work as per the programme. Failure to do so can result in the license being withdrawn.

The changes in effect introduce a “use it or lose it” policy to mineral rights similar to the regulatory framework which has existed for many years in other large, mining countries such as Australia or Canada.

3.2 Lace Diamond Mine Title, Location and Demarcation of Mining Permit

The Lace project is held by Lace Diamond Mines (Pty) Ltd [“LDM”], a company owned 74% by DiamondCorp Holdings Limited, 13% by Shanduka Resources (Pty) Ltd and 13% by Sphere Investments (Pty) Ltd. DiamondCorp Holdings is a wholly owned subsidiary of DiamondCorp PLC. Shanduka and Sphere are corporations owned and controlled by historically disadvantaged South Africans as required by the Black Economic Empowerment Charter as it applies to the South African mining and minerals industry.

The Lace Diamond Mine property is situated approximately 200km southwest of Johannesburg, the main South African centre of transportation, commerce and industry, in the Free State Province (Fig. 1). The property lies less than 20km off the N1 national highway, some 20km along a secondary highway northwest of the city of Kroonstad. Local access is by a 1km all-weather gravel road from the secondary highway, and thence a network of gravel roads on the property itself.

The Lace Mine property consists of 239.97 ha comprising Portion 1 of Farm Ruby 691 and a portion of the remainder of Farm Ruby 691 upon which LDM owns surface and sub-surface rights. The mine, tailings dumps and new slimes deposits are all within

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Figure 3-1 Location Map

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Figure 3-2 Property and Infrastructure Map

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Subdivision 1. National grid electrical power is similarly close to the project.

Kroonstad is a city of 100,000 people, which provides all manner of support, including a labour force familiar with mining as many of the Free State gold mines as well as several small diamond fissure mines are some 40 km south in the Welkom area. LDM buses their staff to and from the mine, or individual employees make their own way.

3.3 Permitting Rights and Agreements Related to the Lace Diamond Mine

The main Lace Mine property comprises a New Order Mining Right over Portion 1 and a portion of the remainder of Farm Ruby 691 totalling 239.97 ha. The mine, tailings dumps and plant site are all located within Portion 1, and are included within the surface rights.

3.3.1 Surface Rights

The surface rights to Subdivision 1 were acquired outright from the previous owner, Christiaan Potgieter Trust, on 28 March 2005. The New Order Prospecting Right over Ruby 691 was acquired on 28 March 2005 from the previous owner, Papavangelo Trading 35 (Pty) Ltd., following Ministerial consent to the transfer of the prospecting right to LDM. The Mining Right was granted to LDM on 5 February 2009.

3.3.2 Taxes and Royalties

LDM is subjected to Government royalties on diamonds recovered and sold from the kimberlites on the property and usual corporate taxes. Government royalties are calculated on a sliding scale based on corporate earnings up to a maximum of 7% of gross revenue. Diamonds recovered from the tailings dumps are not subjected to royalties. Corporate taxes in South Africa are currently 28%.

3.3.3 Obligations

LDM is obligated to undertake its mining activities in accordance with the Mining Works Program submitted to the Department of Mineral Resources.

3.3.4 Environmental Liabilities

Environmental liabilities are calculated in accordance with an annual quantum submitted and approved by the Department of Mineral Resources. In the year ended 31 December 2015, LDM provided an amount of R10,471,777 as an environmental bond for mine shutdown and site clean-up.

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Figure 3-3a Property Map

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Figure 3-3b Property Google Earth Satellite Image

3.4 Permits

LDM has a Mining Right, an approved Environment Management Plan and a water usage licence.

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4.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The Lace Diamond Mine property is situated approximately 200km southwest of Johannesburg, the main South African centre of transportation, commerce and industry, in the Free State Province (Fig. 1). The property lies less than 20km off the N1 national highway, some 20km along a secondary highway northwest of the city of Kroonstad. Local access is by a 1km all-weather gravel road from the secondary highway, and thence a network of gravel roads on the property itself.

The nearest population centre is Kroonstad, 20 km southeast of the mine. Established mining centres are located in Klerksdorp and Welkom, all within a 100km radius of the mine. Major engineering services and mining equipment suppliers are located in Johannesburg, approximately 200km north of the mine.

The area surrounding the mine is primarily utilized for cattle ranching.

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5.0 HISTORY

The Lace Mine, also known as the Crown Mine, is a past-producer which operated intermittently between 1902 and 1932 after being discovered in 1896. A fairly complete record of production during that time is available from the records of the Crown Diamond Mining and Exploration Company Limited, and from various summaries prepared by De Beers personnel. The following section 5.1 on historical production and development is taken directly from past MPH reports.

5.1 Summary of Pre-Great Depression Period 1896-1931

1896-1907 “The main Crown kimberlite pipe was discovered in 1896 and was offered to De Beers Consolidated Mines Ltd., who over a several month time span conducted extensive testing of the pipe, concluding that it was unpayable. The pipe was eventually purchased by the Lace Diamond Mining Company (“Lace Mining”), who worked the pipe from 1902 to 1907. The Crown Mine was initially operated as an open cut mine. Total production between 1902 and 1907, won from the open cut operation, was 878,507 tonnes of kimberlite ore, yielding 139,799 carats of diamonds, for an average grade of 15.91 cpht (Kujawa, 1995, duToit, 1967). In 1907, Lace Mining, badly in debt, sold the mine to the Crown Diamond Mining and Exploration Company Limited (Dixon, 1979).”

1907-1918 “Crown Mining continued operating the Crown Mine as an open cut operation between 1908 and 1910, during which time an estimated 195,407 tonnes of ore was processed, yielding 27,357 carats at an average grade of 14.0 cpht (Kujawa, 1995). Production ceased between 1910 and August 1913 for unspecified reasons (difficulties with flooding, wall collapses and inadequate mining and processing facilities are cited (CM Reports)). These difficulties were overcome by August 1913, and production resumed, with 65,087 tonnes of kimberlite processed, yielding 9,051 carats of diamond at an average grade of 13.91 cpht, including one 72-carat fine quality stone. A 1919 CM Report cites the average grade for kimberlite mined from January to May 1914 as 13.78 cpht, and an estimated 65,318 tonnes of ore, yielding 9,000 carats was mined during this period. In May 1914 the Crown Mine was shut down for the duration of WWI, and production was resumed in 1918.”

1918-1924 “Very little production from the Crown Mine appears to have been achieved during the period 1918 to 1923. Notwithstanding, significant advancements in development of the mine were made during this time, in an attempt to determine the grade and limits of the kimberlite pipe at depth. A development shaft, established through ‘low-grade’ ore approximately in the centre of the kimberlite pipe was extended from the 43m level to the 135m level in 1918, and four drives were constructed. Kimberlitic material from these drives was said to consist of approximately 10% stones (waste material), a significant improvement over higher levels in the mine (CM Reports). While the

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average grade of this ore was disappointing (6.82 cpht), a number of diamonds of good size and quality were recovered (53¼ carats, 47¼ carats, 18½ carats) (CM Reports). In 1919, the prospecting shaft was sunk a further 30m to the 165m level. At this time, with a recovery in the price of diamonds, further development was put on hold and production was started up. In 1920, mining was carried out continuously throughout the year, with 122,444 tonnes of kimberlite processed, yielding 14,189 carats of diamond at an average grade of 11.59 cpht (CM Reports).”

“Exploratory work, concurrent with production, started up again in 1920, with test material from the 165m level yielding up to 20.34 cpht. During this development work, it was observed that western and southwestern sections of the pipe appeared to consist of cleaner, higher grade ground than other sections, but were blocked by a large mass of ‘hardibank’ (hard, low-grade ore). Three tests of this ‘better’ kimberlite, conducted in 1921, yielded a total of 312 carats of diamond from 1,516 tonnes of kimberlite, for an average grade of 20.58 cpht (CM Reports). An underground drill hole designed to test the western extent of this ‘better’ kimberlite encountered a large body of basalt at the western edge of the pipe. Suspicious that additional kimberlite may occur west of the basaltic body, a west-trending drive was initiated at the 135m level, and after passing through 35m of basalt, broke out into ‘blue ground’ (kimberlite ore), containing rounded pebbles and small boulders, but generally cleaner than the kimberlite in the main portion of the pipe (CM Reports). This new kimberlite area west of the basaltic body, was later referred to as the “Satellite Pipe”. Additional tests from the 165m level in 1921 and 1922 of a total of 5,706 tonnes of kimberlite yielded 1,501 carats at an average grade of 26.31 cpht. Anticipating increased production from the new area, as well as improving grades throughout the mine, plans were initiated to upgrade the processing facilities to be able to accommodate 36,000 tonnes (50,000 ‘loads’) per month. Production continued at the Crown Mine until April 1921, when the focus again shifted to exploratory work exclusively. The Crown Mine records indicate that exploratory work continued through to the end of 1923. During this time, the plant was upgraded, an electrical hoist was installed, new quarters and a hospital were built and the open cut was prepared for larger scale production.”

1924-1931 “In 1924, armed with a new 57,500 tonne per month plant, Crown Mining was able to increase its production, but could not be as selective about which ground to mine as it was able to be when lesser feed was required for the plant. This resulted in a generally lower yield in production, as more hardibank was being processed along with the blue ground. In addition, severe rains caused the mine operations to be severely handicapped. From January to August, 1924, 235,890 tonnes of material was processed, yielding 20,890 carats, at an average grade of 8.85 cpht. A total of 29,608 carats were produced for the year (CM Reports), presumably at the same average grade. In 1925, a major collapse of ‘reef’ material from the sides of the open cut caused a hiatus in production from May to September. Production for the year totalled 15,287 carats from 179,130 tonnes of kimberlite, at an average grade of 8.53 cpht.”

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“Crown Mining decided to conduct further testing of kimberlite at the 165m level, following positive results from the 1921/22 tests that resulted in an average grade of 26.31 cpht. In 1926, a test of 2,703 tonnes of blue ground from the 165m level yielded 722 carats of diamond, at an average grade of 26.71 cpht, by far the most high grade [sic] material encountered to date in the Crown Mine. This yield included one 13¼- carat diamond and one 12-carat diamond, both of which were ‘fine quality’ (CM Reports). The blue ground tested was described as “soft, true kimberlite”, and was compared favourably to the best ground at the Kimberley Mine with respect to its texture and softness (CM Reports). It was also noted that the area of good blue ground was increasing with depth, from 10,404 m2 at the 135m level to 14,154 m2 at the 165m level (CM Reports).”

“Following these positive indications in volume and grade, and realizing that the kimberlite at this level could not be won from the increasingly unstable open cut, Crown Mining decided to develop a new shaft to begin underground mining operations. By the end of 1926, the main rock shaft (the new shaft to be used for the underground operations) had been sunk to 256m and two cross cuts between the rock shaft and the prospecting shaft were initiated at the 165m and 240m levels. The prospecting shaft, originally sunk in hardibank kimberlite in the centre of the pipe, broke through into blue ground at 170m, the hardibank apparently being culminated.”

“While the underground workings were being developed, Crown Mining continued crushing and washing available blue ground from the open cut, and from January to June 1926, produced 11,552 carats of diamond from 88,759 tonnes of kimberlite, at an average grade of 13.01 cpht (CM Reports). The Company changed its year-end date to June 30 at this time, and adopted a forward-looking policy to maintain development in the mine significantly in advance of production.”

“From July 1926 to June 1927, 235,143 tonnes of kimberlite were processed, yielding 28,740 carats of diamond at an average grade of 12.22 cpht. During this time, the main rock shaft was sunk to 246m, with a main haulage tunnel driven at the 165m level, and the 105m, 120m, 135m, and 165m levels were opened up and ready for chambering. For the year ending June 1928, 65,488 carats of diamond were recovered from 418,903 tonnes of kimberlite, at an average grade of 15.63 cpht, an increase in yield of over 3 cpht.”

“From April 1928 onwards, all ore processed was taken from the underground workings, the open cut mine having ceased to operate. The schedule of production and development continued through the 1929 year, with 77,753 carats recovered from 428,773 tonnes of kimberlite, at an average grade of 18.13 cpht. All but 61,645 tonnes of this material was taken from above the 190m level. In keeping with Crown Mining expectations, the diamond yield increased as ore was obtained from progressively deeper levels. Also in 1929, the main haulage level was moved from the 165m level to the 240m level. The main shaft was sunk to 335m, and levels down to 287m were driven. Crown Mining reports that the size of the pipe, and the quality of the blue ground were consistent down to the 287m level.”

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“For the year ending June 1930, with the bulk of the output coming from the 165m level downwards, 113,299 carats of diamond were recovered from the processing of 562,892 tonnes of kimberlite, at an average grade of 20.13 cpht. With the extension of the main shaft and the prospecting shaft down to 362m and the construction of the 335m cross cut complete, Crown Mining’s development policy was put on hold, as it was many years ahead of production (CM Reports). 109,785 carats of diamond were recovered from 600,994 tonnes of kimberlite between July 1930 and June 1931. This represents an average yield of 18.27 cpht, a 1.86 cpht decrease from the previous year. This decrease was apparently anticipated, as the blue ground between the 212m and 240m levels was found, during the development stage, to be mixed with diabase and other ‘reef’ material to an unusual extent, thereby decreasing the grade (CM Reports).”

“In 1931 the 335m cross cut was completed. During the construction of this cross cut, in its approach from the main shaft to the south, the kimberlite pipe was encountered much earlier than expected (approximately 75 m south of where it was encountered on the 240m crosscut (Beetz, 1931). Crown Mining reports a “large and important extension of the pipe” towards the south and east, and reports the area of the pipe at the 335m level increased at least 50% over the area at the 240m level (CM Reports). Crown Mining did not have time to fully explore this additional area, and expected to do so in the next year. 5,589 tonnes of kimberlite from construction of the 335m level were hoisted and dumped on surface - there is no record of the grade of this material.”

“From July to October 1931, operations continued as normal, with 37,503 carats of diamond recovered from 200,307 tonnes of kimberlite (Kujawa, 1995). In October 1931, as a result of the Great Depression and a crash in the world diamond market, operations at the Crown Mine were suspended. The African and European Investment Company Ltd. (a company which controlled Crown Mining) advanced funds to maintain the mine in the hopes that diamond prices would shortly rebound and production could resume (CM Reports).”

In all, the Lace Mine is estimated by De Beers to have processed 4,947,203 tonnes during its evaluation and production history, yielding 727,675 carats. The grade of this material is 14.7cpht as calculated by De Beers, as per Table 5-1, below.

The records indicate that some ~1.8Mt were mined from underground during the last four years, as compared to the ~3.1Mt (Kujawa) that were extracted primarily from the open pit. Historical records indicate that ore was extracted from the 105m level to the 212m by partial chambering, using first the 165m haulage level, then the 240m level. No exploitation below -240m was carried out, although some development work was. All of the historical level plans show that kimberlite was not mined to the margins of the pipe, even within portions of the open pit where ~30m of remnant material was mapped and sampled by MPH in historical adits on the north pit wall on the -30m level.

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Table 5-1 Summary of Lace Diamond Mine Historical Production

Year Tonnes Carats Grade 1902-07 878,507 139,799 15.9 1908-10 195,407* 27,357 14.0 1913 39,226 5,127 13.1 1914 120,985* 16,333 13.5 1918-191 24,502 3,449 14.1 1919-20 4,296 293 6.8 1920-212 122,436 14,188 11.6 1,515 307 20.3 1921-223 46,655 6,205 16.1 5,115 1,125 22.0 1922-234 12,352 1,580 12.8 4,915 1,275 25.9 1923-24 235,880 20,896 8.9 1924-25 446,051* 29,608 6.6 1925-265 369,120 26,839 7.3 3,724 772 20.7 1926-27 235,143 28,737 12.2 1927-28 418,903 65,488 15.6 OP Total 3,154,652 389,378 12.3 1928-29 428,773 77,753 18.1 1929-30 562,456 113,299 20.1 1930-31 601,029 109,742 18.3 1931-32 200,293 37,503 18.7 UG Total 1,792,551 338,297 18.9 Lace Mine Totals 4,947,203 727,675 14.7

* Denotes tonnages estimated by Kujawa 1 Bulk sample results for development on -135m level 2 Bulk sample results for development on -165m level 3 Bulk sample results for development on -50m, -135m and -165m levels 4 Bulk sample results for development on -165m level 5 Bulk sample results for development on -165m level

5.2 De Beers Period 1931-1996

“The Crown Mine was kept pumped out, and the machinery maintained until 1939, when the mine was auctioned off and sold to De Beers Consolidated Mines Ltd (Rossouw, 1997). Eventually maintenance of the mine was abandoned, the main shaft collapsed and the open cut was flooded. In 1957, rights were granted to the S.A.R. & H. Railway to use the water in the open cut (Dixon, 1979) (as far as is known, this right still exists).

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A small number of reports outlining De Beers’ activities at the Crown Mine are available, including a fairly comprehensive evaluation of the tailings dumps. In a 1967 evaluation of the Crown Mine, du Toit states that based on revenues generated when the mine was operational, De Beers considered the operation marginal, and could not recommend further evaluation. The project was shelved until March 1997, when it was sold to the Christian Potgieter Trust.” (Henderson, 1997)

5.3 Christian Potgeiter Trust 1996-2005

The author (Sobie) was contacted early in 1997 by representatives of CPT (the Christian Potgeiter Trust), to elicit interest in the Lace tailings dumps for clients. Upon examining the dumps, and noticing large blocks of unprocessed, high-interest hypabyssal kimberlite, MPH became quite intrigued and sent several samples of this material in for microdiamond analysis, which were positive and suggestive of high-grade. MPH believed that both the kimberlite resource and tailings dump potential were of high interest, and attracted a Canadian junior mining company to the project, Rupert Resources Ltd. (“Rupert”), which negotiated a phased acquisition deal with CPT in 1997.

Rupert was unable to raise capital for the project in the depressed resource markets of 2000-2001, and it reverted to the vendor, CPT.

The project was dormant between 2001-2005 as CPT marketed the property to various diamond mining companies and ventures, with MPH variably providing technical data to the interested parties.

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Figure 5-1 Surface Plan of Lace Mine and Tailings Dumps with 1997-98 Work

The Rupert exploration program commenced during October, 1997 lasting until September 1998 in terms of the major field work, with sporadic activity until late 2000. Survey crews were brought to site to establish a control grid on the property, and establish tailings dump volumes through the construction of a digital terrain model (DTM). The work was subcontracted to Global Surveys (Pty) Limited of Florida, RSA.

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The grid incorporated a surveyed and levelled baseline and 50m-spaced tie-lines oriented at true azimuth 0o, with 25m-spaced perpendicular cross-lines over the 950m length. Stations were every 25m along all lines, and all were surveyed utilizing Global Position System (GPS) equipment such that the grid is easily referenced into National and International grids. Accuracy for the levelling was 10mm/km. Permanent bench marks were established along the baseline.

Geophysical surveys were contracted to Integrated Geophysical Services cc of Randburg, RSA. The magnetic survey was carried out using a GEM GSM-19 Overhauser instrument in walking mode linked to a Garmin SRVYII GPS system. With a cycling time of 1 second on the magnetometer, it ensured that a reading was taken every 1-2 metres. Both the GPS and magnetic base stations were located within the surveyed area. The gravity survey was carried out using a Scintrex CG-3 Autograv system, with stations read every 25m along all lines. The gravity base station was similarly located on the grid at station 200W/350S and was read every 3 hours during the survey.

5.3.1 Delineation Drilling and Sampling Programme

Following a tender process, the core drilling contract was awarded to Zaimann Exploration Drilling cc of Witbank, RSA, who provided a Boyles BBS 37 rig equipped with H, N and B-sized rods and all ancilliary equipment. A second smaller rig was on-site for approximately two months and drilled holes Crown-2, 4 and 5, all of which were short. Surveying of holes was performed by Borehole Survey (Pty.) Ltd., who initially were utilizing a Sperry-Sun Multishot system for the first five holes and thereafter utilized a state-of-the-art Reflex EMS Downhole Survey system which digitally probed the holes’ trace.

A summary of the delineation drilling program is provided below, which can also be referenced by Fig. 5-2, and all later plans and sections in this report.

The delineation programme included systematic petrographical, major and trace element as well as comprehensive microdiamond sampling, which was designed to determine fine-diamond distributions within the Lace kimberlites. As such, virtually every kimberlite intersection was sawed in half, utilizing a rotary concrete saw with synthetic diamond blades, to yield a split sample of nominally 25 kg weight represented by 6m of split material. In practice, more friable core, and the BQ core, had to be submitted whole as alternate metres over selected interval lengths.

Samples were placed in sealed pails and shipped under secure conditions directly by bonded expeditors to Lakefield Research Limited (now SGS Lakefield Research Limited), near Peterborough, Ontario, Canada.

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Table 5-2 Summary of 1997 - 1998 Exploration Programme Delineation Drilling

Hole Objective Collar Az. Dip Final Depth Crown-1 Test Bulge from south 100E/240S 345o -60o 905.00m Crown-2 Test Bulge from above 035E/015S 345o -85o 432.20m Crown-3 Test Main Pipe from N. 100W/430N 165o -57o 977.40m Crown-4 Test Sat. from above. 175W/075S 000o -85o 442.77m Crown-5 Test Sat. Pipe from S. 175W/075S 000o -45o 103.70m Crown-6 Test Sat/Main from W. 330W/055N 075o -55o 869.30m Crown-7 Test Main/Sat. from E. 200E/225N 255o -60o 652.90m Crown-8 Test Main Pipe from E. 212E/175N 255o -60o 893.90m Crown-9 Test Main Pipe from W. 250W/025N 075o -60o 484.00m Total Drilling 5,961.17m

De Beers Research Laboratories in Kimberley provided their services as the umpire facility. The second split of selected samples were therefore sent to De Beers for audit purposes. The two facilities make use of different methodologies; Lakefield utilizes caustic soda dissolution and De Beers hydrofluoric acid dissolution, however, in essence both facilities dissolve as much of the rock as possible to leave a concentrate, which is then examined for microdiamonds. Both methodologies have merits, namely the first is much safer and does not involve crushing, whereas the second can handle larger volumes and involves less steps, thereby lessening the chances of contamination.

In all, some 93 kimberlite samples totalling 2,156.95kg constituted the useable Main Pipe database at that time, with two basalt-breccia, xenolith samples and two hole Crown-2 samples (into the outer kimberlite termed the Bulge) not incorporated into the tallies. Much more limited drilling of the Satellite Pipe resulted in only 17 samples and 402.08kg constituting the database.

In general terms the drilling was very successful in outlining a Main Pipe kimberlite with high-interest K4 hypabyssal facies becoming more dominant over the dilute K6 volcaniclastic kimberlite with depth. The program was unsuccessful in confirming the precise position or nature of the southern “Bulge”, which has subsequently been delineated and will be discussed later in the report. The 1997-98 program was initial in scope, and allowed for only rudimentary modelling based on few margin intercepts between -345m and -855m below surface.

A ~8,000m program of drilling from surface was recommended (MPH 2005) to more concisely establish geology, volumes, density and tonnages above -400m.

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Figure 5-2 3-D Perspective of Lace Mine and Tailings Dumps from 1997-98 Work

5.3.2 Tailings Bulk Sampling Programme

The overall objective of the tailings sampling project, carried out in late 1997 to mid-1998 by MPH, was to obtain sufficient information as to the contained tonnages and grades of and value of the diamonds in the dumps so that the appropriate resource classifications could be assigned to the deposit, or various portions of it. Specific objectives were:

 To excavate and treat a large tonnage of the tailings that would be considered representative of the total resource.  To be able to assign grades (in carats per hundred tonnes) to each of the dumps treated.  To determine the variability of the diamond content in each of the dumps by batch sampling.  To obtain a minimum 5,000 carat parcel for valuation purposes, as diamond value is a key component in determining whether or not resources can subsequently be converted to reserves.

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A representative programme was subsequently designed, with deep slots extracted from all of the major dumps as per Figure 5-1, which in total represented approximately 2.5% of the surveyed resource volume, and therefore was comprehensive by industry standards.

A total of 6,217.45 diamond carats were recovered which indicated a grade of 6.9 carats per hundred tonnes from the primary recovery stage of the programme, through CPT’s pan plant and grease table recovery module. The tailings produced from the sampling of Dumps 1A, 1B and 2A (the largest dumps) were subsequently audited, i.e. re-processed, at an independent laboratory to determine the efficiencies of the pan plant and diamond recovery facilities utilized. Audit results indicated that only 65% of the diamond content of the dumps was being recovered, such that resulting grades were revised upwards by a factor of 1.5. All of the diamonds recovered during the sampling process were sold and diamond sales certificates indicate the average dollar per carat value to be $46 at that time.

The valuations and value models related to the size-frequency distribution of the bulk sample parcel, were updated for MPH in 2005 by expert consultants WWW International Diamond Consultants Limited, using their May 2005 price book, who concluded that an increase of 50% to $69/carat was justified. MPH further concluded that the value of $63/carat should be utilized in all studies, to account for the upward factorization of grade from the audit, combined with the downward factorization of value inherent in the recovery of the smaller stones missed by the inefficient primary processing plant.

Results of the programme allowed the tailings to be placed in the following resource categories, pursuant to the SAMREC Code for South Africa:

Table 5-3 Summary of 1998 Tailings Resource Estimates

Resource Category Tonnes % Tonnes Grade (cpht) Measured 2,853,000 72 10.7 Indicated 742,300 19 7.6 Measured plus Indicated 3,595,300 91 10.3 Inferred 391,000 9 4.3

*weighted average. +1mm bottom cut-off screen size

It was recommended, in 1998, that appropriate capital and operating costs be amassed and that a Feasibility Study be completed for this tailings material, and that bulk sampling of the above inferred resource be undertaken during the commissioning of any new plant facilities at Lace.

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5.4 DiamondCorp plc 2005-2014

The tailings dump work was updated for DiamondCorp to 2006 (MPH and Crosston Technologies, 2006a), and along with the updated report on the potential of Lace underground kimberlite resource (MPH, 2005), formed the basis for capital raising and listing in late 2006. DiamondCorp also embarked on the historical shaft refurbishment programme recommended, as a first step to commencing underground kimberlite evaluation.

5.4.1 General

Circumstances have dictated that DiamondCorp’s path to the present underground evaluation program and designing/commissioning of the UK4 Mine has been less direct than one would have hoped in 2007. A number of events have caused changes in plans including:

a. Despite attaining steady-state tailings dump production at 120tph with their purpose-built plant in late 2007, the world-wide financial crisis in 2008 saw diamond prices drop precipitously, to the point that the tailings reprocessing operation was unprofitable and shut down in late 2008. b. Shaft refurbishment was found to be impossible below -160m due to unstable ground conditions, and a decision was therefore made by LDM in 2008 to sink a decline ramp to access shallow Satellite and Main Pipe material left behind by historical miners. The shaft was later joined with the new underground infrastructure to provide ventilation. The plant was modified to be able to process fresh kimberlite material in addition to the tailings. c. The remnant kimberlite at the -60m target level in both the Satellite and Main pipes was found to be heavily dilute (and unstable in the Main Pipe), and the decision was made in 2009 to continue the decline to the -250m level, below all historical production stopes. d. Weakness in the capital markets in 2011-2012 made it difficult to advance the project following successful Main Pipe bulk sampling on the 250m level, however SRK was commissioned to design a higher confidence block cave mine. Their design included conveyor systems and underground crushing, with new twin declines from surface to the -470m level. e. Capital availability in 2013 saw the commencement of decline development and late in the year an underground core drilling rig was purchased, and installed on the 250m level in early 2014.

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f. The delineation drilling initiated in early 2014, combined with detailed mapping of the 250m bulk sampling level allowed for MPH to identify the upper K4 resource potential, previously only hinted at from the shallowest 1998 borehole, Crown-7.

A chronology of major milestones since LDM took over the Lace Mine is provided below in Table 5-4.

Table 5-4 Summary of Lace Diamond Mines’ Operations 2005-2014

U/G Bulk Bulk Tailings Tailings Tailings Tailings U/G Bulk U/G Carat Major Plant & Infrastructure Sample Sample Year Permits Obtained Tonnes Carats Grade Carat U/G Developments Sample Value Developments Carats Grade Processed Recovered (cpht) Value $/ct Tonnes $/Ct Recovered (cpht) Acquired property 1.6mTPA Tailings Plant designed and 2006 with surface construction commenced mining rights Surface mining fleet purchased 2007 Tailing plant commissioned 304,045 25,266 8.31 $ 63.00 Tailings reprocessing commenced Q4 Historical shaft refurbishment commenced Decision to ramp to upper Satellite and Decline portal sunk and 2008 UG Mining Lease Main pipes made, UG development 803,810 50,521 6.29 $ 53.52 reached Satellite Pipe at - 11,380 808.21 7.10 $ 108.00 designed 60m, Main Pipe at -55m U/G mining fleet purchased Portal and decline commenced Historical shaft refurbishment progressed Connection to new power line (2nd line) Plant modifications for kimberlite progressed Decline advanced towards 2009 Tailings operation suspended 01/2009 Main Pipe -250m level Plant kimberlite modifications completed "Upper" U/G Main Pipe mine (-260 to - 330m) designed Decline reached -250m level 2010 Plant recommissioned Q4/2010 24,145 1,321 5.47 $ 94.00 of Main Pipe Total of 1,800m of development U/G engineering study (SRK) for -470m Bulk sampling Main Pipe on - 2011 15,414 2,157.41 14.00 $ 160.00 block cave 250/260m level Conveyor system/UG crushing in design Heavy equipment workshop established

2012 Plant decommissioned No U/G activity U/G mine on care and maintenance Twin declines established Twin decline boxcut and portal 2013 from -92m, developed established upwards towards portal 400 tph conveyor system commenced U/G core rig bought and set- fabrication/installation up on -250m level 1130m of decline Plant recommissioned 138,745 6,442 4.64 $ 63.00 development completed Upper K4 ("UK4") resource base identified 1410m of decline 2014 307,953 18,354 5.96 $ 63.00 and scoping study completed development completed

Additional water storage dam constructed

517m of decline development Conveyor system completed from -250m 2015 completed to the end of level to surface. August 165m of tunnel development in the UK4 mining block completed to the end of August. Tailings Totals 1,578,698 101,904 6.45 58.70$

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5.4.2 Tailings Reprocessing and Plant Construction

LDM commissioned Consultmet (Pty.) Limited to design and construct the Lace Mine tailings retreatment plant in 2006, which was completed and commissioned in late 2007. The plant was designed to treat 1.6 million tonnes per annum, at a 220 tonnes per hour (“tph”) head feed. The plant utilizes a large double screening front end, 100 tph (coarse +6mm fraction) and 65 tph (fine -6+1mm fraction) DMS modules, a re-crush circuit and hands free grease belt and glove box recovery units for diamond recovery.

The company purchased its own fleet of Bell mobile mining equipment in 2007, and commenced site preparations, as well as the construction of a new tailings and slimes dam.

Production has been discontinuous as discussed previously, and as per Table 5-4, has totalled some 1.58 million tonnes processed, recovering 101,904 carats for a grade of 6.45cpht, and a value per carat of $58.70. Problems with the re-crush circuit during 2007-2008, since rectified, were felt to be a major reason for the lower recovered grades during that first, major period of tailings re-treatment. That notwithstanding, the marginal nature of the resource has meant that LDM now plans on blending tailings material with the UK4 Mine underground material going forward.

5.4.3 2009 Satellite and Main Pipe Bulk Sampling

With the tailings resource proving to be marginal in late 2008, the decision was made to put that operation on care and maintenance, whilst modifying the plant to handle kimberlite and proceeding with the 4m x 4m decline. A key development was the granting of the mining right for the Lace Underground Mine by the Department of Minerals and Energy (“DME”) in February 2009. Also the Company purchased second hand underground mining fleet equipment to carry out the development work in-house.

Primary and secondary crushing modules were installed and commissioned in early 2009, and as well the operation benefitted from the installation of a second power supply from a new power line servicing the adjacent Voorspoed Mine of De Beers.

The decline accessed the Satellite Pipe and a remnant lens of heavily dilute K6 kimberlite material on the north edge of the open pit, on the -55m and -60m levels respectively, as per figure 5-3 below. The Satellite Pipe kimberlite was also heavily dilute K6, as the 1997-98 drilling by MPH had intersected, which was postulated as the reason the early miners had chosen not to exploit this resource. In general it was felt that Satellite Pipe grades must have been significantly less than Main Pipe grades at the same elevations, otherwise it would have been mined historically.

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The Satellite Pipe operation was quite successful in recovering some 808.21 carats from 11,380tonnes of K6 material, for a recovered grade of 7.1cpht. The parcel was valued at $108/ct. in 2009 at a tender conducted by Sadiamex (Pty) Limited in Johannesburg, an encouraging value in depressed diamond market conditions at the time. The Main Pipe attempted bulk sample encountered unstable conditions and no bulk sampling was completed. MPH had previously investigated this lens of remnant kimberlite by rafting across the flooded open pit to the exposed historical tunnels, and noted very highly diluted K6/CRB.

Figure 5-3 Development Work and Bulk Sampling on -55/-60m Levels

5.4.4 2011 Main Pipe Bulk Sampling

DiamondCorp spent 2010 and early 2011 advancing the decline some 1,800m to the -250m level, deciding to get beneath the lowest historical stopes which reached the -238m level. Kimberlite was reached on 10 May, 2011. (DiamondCorp 2010 Annual Report)

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Figure 5-4 Development Work and Bulk Sampling on -250/260m Levels

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Although plans had been for a systematic bulk sampling program comprising six 3m by 3m development drives across the entire kimberlite, poor ground conditions allowed for a more modest approach. The development drives encountered numerous historical tunnels and winzes, resulting in significant caving, as attempts were made to access the central portions of the pipe. The final orientation of bulk sampling drives, and individual samples is shown below in Figure 5-4 along with up-to-date geology and drilling. Results were very encouraging, with 2,157.41 carats recovered from 15,414 tonnes processed. The Company’s consultants at the time had a differing geological interpretation than that of MPH, such that reports and press releases referred to Brown, Grey and Contact VK’s (volcaniclastic kimberlites).

Table 5-5 tabulates the 2011 results, with lithological determinations of the material sampled determined from detailed mapping. None of the samples were unfortunately limited to a single kimberlite facies type, as they were oriented along major K4/K6 contacts in many cases. Results therefore were generally somewhat higher than expected for pure K6 material, with the plant using a lower bottom screen (+1.00mm) for this work also contributing to the higher grades.

Prior to the plant upgrade in 2013, the plant used to have 0.5 mm dewatering screens on the feed preparation section of the plant, resulting in any carry over of small diamonds on the 1.00 mm trammel of the primary scrubber (as was common because the 1.00 mm panels easily blind) would report to the dense media separation (“DMS”) circuit and final recovery.

The plant now has 1.00 mm de-watering screens, so those small diamonds now report to coarse tails.

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Table 5-5 Summary of 250m Level Bulk Sampling Results

Lace Mine 250m Level Bulk Sample Results (+1mm Bottom Screen Size) Sample 250 - S1 250 - S2 250 - S3 250 - S4 250 - S5 250 - S6 250 - S7 250 - S8 250 - S8b 250 - S9 Totals Lithology Mixed K6/CRB Mixed ~50%K6/50%K4 Mixed ~85%K6/15%K4 Mixed ~85%K6/15%K4 Mixed K6/K4 Mixed K6/CRB Mixed ~95%K6/5%K4 Mixed ~95%K6/5%K8 Mixed K6/K4 Mixed ~88%K6/12%K4 ROM Tonnes 535 1086 761 2084 2087 864 2121 1850 2670 1356 15414 Total Carats 80.81 135.800 121.260 189.750 290.200 48.260 184.840 254.780 690.780 160.960 2157.44 Grade (cpht) 15.10 12.50 15.93 9.11 13.91 5.59 8.71 13.77 25.87 11.87 Parcel 8388, 8389 8391, 8392 8392, 8393 8394, 8395, 8412 8396 - 8399 8400 - 8401 8402 - 8405 8409 - 8415 8419 - 8426 8680 Sorting Date 09/06/2011 17/06/2011 30/06/2011 30/06/2011 06/07/2011 19/07/2011 29/07/2011 20/08/2011 26/08/2011 16/07/2015 Diamond Mesh Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats +150 CT +100 CT +60 CT +45 CT +30 CT +20 CT 1 20.58 +15 CT 1 16.08 +10 CT +9 CT +8 CT +7 CT 1 7.96 +6 CT 1 6.28 1 6.86 1 6.07 +5 CT 1 5.38 2 10.73 1 5.69 +4 CT 1 4.61 1 4.48 1 4.14 +3 CT 1 3.25 1 3.33 1 3.72 1 3.70 1 3.30 +2.5 CT 1 2.59 2 4.82 2 5.37 1 2.87 +2 CT 0.00 1 2.10 1 2.44 1 2.30 3 6.23 2 4.70 +1.5 CT 0.00 3 4.47 2 3.99 3 6.24 6 9.54 2 3.06 4 6.68 2 4.51 15 22.71 4 5.29 +1.25 CT 2 2.57 0 0.00 2 2.69 3 3.84 1 1.38 1 1.26 1 1.26 6 7.62 2 3.58 +21 (+1 CT) 3 3.08 5 5.17 4 4.62 7 7.22 12 12.90 3 3.24 7 7.28 9 8.59 29 30.30 6 6.13 +19 4 2.50 6 4.03 9 6.40 8 5.63 13 9.97 2 1.43 8 5.96 11 8.26 28 20.89 7 4.47 +15 20 8.89 36 15.69 33 14.59 45 19.91 68 30.13 11 5.01 44 19.28 63 27.66 163 71.73 38 16.71 Coarse Total 32 28.26 53 41.06 50 32.28 64 41.44 108 91.09 20 19.81 72 76.47 91 66.03 243 181.79 60 49.69 +11 56 12.11 109 23.71 102 22.17 175 38.09 233 50.66 38 8.23 128 27.82 230 50.48 623 135.59 129 28.10 +9 65 8.08 136 16.94 138 17.12 218 27.16 291 36.16 40 5.02 161 20.05 294 36.58 774 96.24 161 20.08 +8 110 9.53 171 14.78 165 14.20 266 22.92 366 31.60 49 4.25 233 20.12 323 27.85 896 77.29 210 18.08 +7 178 11.85 298 19.81 275 18.18 406 26.97 637 42.32 76 5.04 335 22.28 560 37.20 1514 100.53 329 21.87 +5 128 5.15 216 8.66 189 7.63 389 15.64 462 18.57 77 3.04 219 8.78 404 16.24 1099 44.03 255 10.26 +4 250 5.83 419 9.77 375 8.73 674 15.73 792 18.48 149 3.48 379 8.84 787 18.37 2152 49.80 497 11.60 -4 0 0.00 60 1.08 384 0.95 99 1.80 72 1.32 22 0.39 26 0.48 111 2.03 303 5.51 70 1.28 Fines Total 787 52.55 1409 94.74 1628 88.98 2227 148.31 2853 199.11 451 29.45 1481 108.37 2709 188.75 7361 508.99 1651 111.27 Total (Coarse + Fine) 819 80.81 1462 135.80 1678 121.26 2291 189.75 2961 290.20 471 49.26 1553 184.84 2800 254.78 7604 690.78 1711 160.96 Mida Sample U/G L-1812 L-1813 L-1814 L-1815 L-1816

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5.4.5 2012-2014 Development and Underground Delineation Drilling

Following on capital raising in 2012, DiamondCorp initiated the recommended twin decline infrastructure development from the -92m level, working both upwards toward the portal, and downwards toward the existing decline. The SRK report (SRK, 2012) recommended a continuous trough block cave on the -470m level, serviced by twin declines, with ore and waste material conveyed to surface. Critical to firming up much of the engineering work was a far better understanding of the internal geology and morphology of the pipe, and the company purchased an underground core rig and set it up initially on the 250m bulk sampling level.

The identification of the UK4 resource in 2014 necessitating redeploying equipment and personnel to construct the UK4 Mine underground infrastructure, described in detail in Section 16 of this report.

Figure 5-5a and 5-5b show the completed development work in orange as of the date of this report, as well as completed drill holes. Figure 5-5a shows all workings from surface to the bottom of the known geological model, planned to be mined by successively block caves (Block Caves 1, 2 and 3). Figure 5-5b is a more detailed rendition of the UK4 Mine developments and better shows the underground drill cubbies. For both figures, the gold coloured workings have been completed as of the report date.

Figure 5-5a Lace Mine Development to Dec.18, 2015

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Figure 5-5b UK4 Mine Development to Dec.18, 2015

Drilling on the project now totals 10,183.88 metres as per table 5-6 below, of which 29 holes totalling 4,422m have been drilled from four drill stations or “cubbies”, underground. These have been established to best meet both geological and mining development objectives, and as a summary statement have allowed for concise modelling of the UK4 resource slice portion, of the overall Lace deposit.

Interestingly, the historical “Bulge” has now been located and defined between - 310 and -400m and has thus far been found to be limited to dilute K6 and CRB facies.

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Table 5-6 Summary of Delineation Drilling

Hole ID Target Collar Location Collar X Collar Y Collar Z Length Azimuth Dip CROWN-1 Main Pipe / Bulge Surface 12400.0 -3037269.0 1338.0 905.5 345.0 -60 CROWN-2 Bulge Surface 12335.0 -3037044.0 1338.0 432.2 345.0 -85 CROWN-3 Main Pipe Surface 12200.0 -3036599.0 1333.0 977.4 165.0 -57 CROWN-4 Satellite Pipe Surface 12125.0 -3036954.0 1335.5 442.77 0.00 -85 CROWN-5 Satellite Pipe Surface 12125.0 -3036954.0 1335.5 103.7 0.0 -45 CROWN-6 Satellite/Main Pipe Surface 11970.0 -3036974.0 1333.0 869.3 75.0 -55 CROWN-7 Main Pipe Surface 12500.0 -3036804.0 1336.0 652.9 255.0 -60 CROWN-8 Main Pipe Surface 12512.5 -3036854.0 1337.0 893.9 252.7 -60 CROWN-9 Main Pipe Surface 12050.0 -3037004.0 1333.0 484.0 72.0 -60 BH-27 Main Pipe 250 Level Cubby 2 12239.91 -3036909.39 1081.05 45.00 30.00 0 BH-28 Main Pipe 250 Level Cubby 2 12239.91 -3036909.39 1081.05 135.40 65.00 0 BH-29 Main Pipe 250 Level Cubby 2 12239.91 -3036909.39 1081.05 132.40 50.55 -17 290-03 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 76.00 163.00 -57 335-05 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 100.31 121.00 -28 335-07 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 151.00 84.00 -33 335-08 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 112.00 74.00 -34 430-03 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 136.00 163.00 -66 430-05 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 124.40 121.00 -61 430-06 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 187.50 102.00 -54 430-07 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 178.30 84.00 -49 430-08 Main Pipe 250 Level Cubby 1 12237.63 -3036942.83 1081.89 148.23 74.00 -53 PH-01 Raise Bore Pilot Hole 12237.63 -3036942.83 1081.89 74.90 0.00 -90 BH-30 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 217.0 335 -20 BH-31 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 223.0 321 -20 BH-32 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 200.7 302 -21 BH-33 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 254.2 302 -35 BH-34 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 257.5 320 -30 BH-35 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 146.4 335 -33 BH-36 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 180.7 350 -45 BH-37 Main Pipe 250Level Cubby 3 12383.48 -3036994.20 1084.65 240.1 350 -15 BH-38 Main Pipe / Bulge 270 Level Cubby 12397.49 -3036894.10 1050.36 160.0 215 -42 BH-39 Main Pipe / Bulge 270 Level Cubby 12397.49 -3036894.10 1050.36 156.6 215 -69 BH-40 Main Pipe / Bulge 270 Level Cubby 12397.49 -3036894.10 1050.36 164.5 219 -33 BH-43 Main Pipe 270 Level Cubby 12397.49 -3036894.10 1050.36 216.7 252 -21 BH-46a Main Pipe 270 Level Cubby 12397.49 -3036894.10 1050.36 35.0 280 -21 BH-46b Main Pipe 270 Level Cubby 12397.49 -3036894.10 1050.36 32.8 280 -24 BH-48 Main Pipe 270 Level Cubby 12397.49 -3036894.10 1050.36 118.9 309 -25 BH-51 Main Pipe 270 Level Cubby 12397.49 -3036894.10 1050.36 216.7 316 -38 10 ,18 3.8 8 10,183.88

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6.0 GEOLOGICAL SETTING AND MINERALIZATION

6.1 Regional and Local Geology

The Lace Mine lies within the south central Archean Kaapvaal Craton within a subcluster of kimberlites referred to as the Kroonstad Group II Kimberlite Cluster (Howarth, 2010) of which only Lace and nearby Voorspoed have seen formal mining. Four other small kimberlite blows and several dykes are known within the cluster.

The Kroonstad kimberlites are reported by Howarth (2010) to be emplaced into stratigraphy composed of lower Ventersdorp Supergroup (2720-2650Ma) volcanics, unconformably overlying Transvaal Supergroup (2650-2050Ma) sediments (with subordinate volcanics), and through the uppermost unconformable Karoo Supergroup (300-180Ma) sediments as per Table 6-1 below. Local surface country rocks on the property are Ecca shales, with Transvaal lavas known to outcrop between Lace and Voorspoed (Figure 6-1).

Drilling from surface through the wallrocks into the kimberlites in 1997-98 outlined a relatively simple stratigraphy of Karoo shales to ~60m, Ventersdorp volcanics to ~300m, an intercalated shale/volcanic sequence to ~350m, clean lavas again to 380m, and then Kameeldoorns Formation shales again to ~550m, the deepest wall rock drilling to date, in hole Crown-1 (MPH 1998, 2005). We therefore do not believe that Transvaal Supergroup rocks are present at Lace based on drilling to date.

The Lace kimberlites have been dated at 133.2 +/- 2.8 Ma by the 40Ar/39Ar technique on ground mass phologopite grains (Philips et al., 1999), similar to the early Cretaceous ages of the other Kroonstad kimberlites and to other Group II kimberlites in South Africa including Finsch.

6.2 Lace Kimberlite Geology

The delineation drilling program has shown the geology of Lace to be relatively simple, with five main internal facies as follows, within an irregular pipe-shaped body that varies in areal extent from ~1.7ha to ~2.6ha within the well-constrained -230m to -370m UK4 Mine depth slice.

6.2.1 K2 (Hypabyssal now called Coherent) Kimberlite

Mauve to brown coloured, massive micaceous matrix kimberlite, not segregationary on a macroscopic scale. Up to 21% olivine macrocrysts (generally 10 to 15%). Generally <15% country rock xenoliths (CRX), mostly <5% CRX.

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Table 6-1 Regional Stratigraphy (after Howarth, 2010)

Supergroup Group Sub-group Formation Lithologies Drakensberg Dolerite, basalt Clarens Coarse sandstone Elliot Mudstone, fine sandstone Stormberg Molteno Coarse sandstone, mudstone Beaufort Tarkastad Driekoppen Mudstone Verkykerskop Sandstone Karoo (300-180Ma) Adelaide Normandien Sandstone, mudstone Ecca Volksrust Mudstone/shale, siltstone, sandstone Vryheid Sandstone, siltstone, shale Dwyka Diamictite, conglomerate, shale, mudstone

Unconformity Pretoria Daspoort Quartzite Strubenkop Shale, tuff Hekpoort Dominant lava, quartzite, shale Transvaal (2650-2050Ma) Timeball Hill Shale, quartzite Chuniespoort Malmani Monte Cristo Chert-rich dolomite Oaktree Chert-poor dolomite Blackreef Quartzite, shale, conglomerate Unconformity Allanridge Mafic lava, tuff Bothaville Conglomerate, quartzite Platberg Rietgat Mafic lava, quartzite, limestone, chert Makwassie Felsic lava Goedgenoeg Intermediate-mafic lava Ventersdorp (2720-2650Ma) Kameeldoorns Conglomerate, quartzite, shale Klipriviersberg Edenville Mafic lava, tuff Loraine Mafic lava, tuff Jeanette Mafic lava, tuff Orkney Mafic lava, tuff Alberton Mafic lava, tuff

6.2.2 K4 (Hypabyssal now called Coherent Kimberlite)

Mauve to brown coloured, massive micaceous matrix kimberlite with a generally segregationary or globular segregationary texture on a macroscopic scale. Up to 21% olivine macrocrysts (generally 10 to 15%). Generally 5 to 20% CRX with an average abundance around 15%. CRX are generally <3cm in size, but occasionally range up to 25cm in size. K2 and K4 are very similar except for the minor difference in texture and the abundance of CRX’s, and are combined for modelling purposes.

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Figure 6-1 Local Geology

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6.2.3 K6 (TKB now called Fragmental) Kimberlite

Grey to green coloured, fragmental kimberlite breccia. Up to 15% olivine macrocrysts (generally <10%). Generally 35 to 90% country rock xenoliths (CRX), mostly ~50% CRX. CRX generally much larger than in K4, and can range up to metres in size.

6.2.4 K8 (Lava-Rich Kimberlite Breccia)

Dark grey-brown coloured, fresh, competent lava-rich fragmental kimberlite breccia comprised largely of brecciated lava fragments of various types (very often amygdaloidal) in a very fine grained to aphanitic matrix. The specific percentage of kimberlite is difficult to estimate in macrosection in the fine grained variant, but typically is 15-25%. The coarser matrixed K8 is more obviously kimberlitic visually, confirmed by clearly defined micaceous rims and magmaclasts in places, and occasional garnet, where the matrix is coarser. Generally the percentage of kimberlite ranges from 20-40% in the more attractive K8 breccia. The breccia is also characterized by a variety of lava as well as shale fragments ranging up to 20cm in size, but generally less than 3cm.

This facies was termed “Hardebank” by the early miners and was often left behind as remnant pillars due to its competency and hardness.

6.2.5 Country Rock Breccia (Lava Breccia, Shale Breccia, CRB)

The country rock or lava breccia is comprised of 85 to 100% lava fragments with minor kimberlite in the matrix, and generally occurs at or near the margins of the pipe. Many of the intervals fine upwards and most probably represent talus breccias resulting from the mass deposits due to wall rock slumping or collapse of the lavas, following the explosive formation of peperite beneath Karoo cover.

6.2.6 Ventersdorp Lava (LV)

The code LV is reserved for clearly in-situ intersections of Ventersdorp lava wall rock consisting of a series of individual lava flows each characterized by massive finely crystalline bases, often slightly coarser grained intermediate sections and amygdaloidal flow tops which often show crackle top brecciation.

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Figure 6-2 3-D Perspective Lace Main Pipe Geology

Figure 6-2a – Lace Mine Geological Model Looking Figure 6-2b – Lace Mine Geological Model Looking Northwest South

6.3 Internal and Pipe Margin Geometry

The delineation drilling to date has shown that a central SW-NE oriented “ridge” of K4/K2 coherent kimberlite occupies the Lace Main Pipe, from the 230m level to the deepest intersections at ~875m, gradually flaring out to occupy greater portions of the pipe, with depth. This coherent kimberlite ridge is flanked in both the northwestern and southeastern quadrants by K6 volcaniclastic material, generally with sharp contacts, which in turn grades at the margins to heavily dilute CRB units.

Within the coherent kimberlite ridge, several “pods” of K8 have been intersected above the 350m level that range up to 20-25m in diameter and likely represent basaltic country rock material assimilated early in the intrusion history of Lace. Within the southeastern K6 unit, several large rafts of country rock up to 20m in width have been intersected to date, of both shale and lava, depending on depth.

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Above ~290m, intercalations of K6 with K4/K2 are evident and become more common upwards, suggesting this depth marks the zone of breaching of the K4/K2 intrusion. Following breaching, the sudden decrease in pressure fluidized the system resulting in reaming downwards of the volcaniclastic fragmental kimberlite breccia facies, K6. This reaming downwards appears to have favoured the western side of the Main Pipe, for as yet not understood reasons, as wall rock sediments and lavas appear undisturbed. As well, there are areas where transitional rocks exist at the higher elevations either in the form of more xenolith-rich K4, or K6 containing magmaclasts of K4 hypabyssal kimberlite. Generally the contacts of K4 with K6 are marked by a sharp decrease in country rock xenoliths and a narrow zone of flow banding.

The pipe morphology between 230 and 370m depth has been found to be complex, with the margins quite irregular, as now understood from the detailed drilling. Beneath 370m pipe shape is not well constrained from the few deep holes, however we expect that the irregular margins will continue as is typical of kimberlites at this relative exposure level, ie. Kimberley, and the model will be updated periodically going forward. At present it is modelled as gently tapering with depth, as per Figure 6-2, and 14-4, the level plan compilation, and 14-5, the section compilation.

A comparison of figure 5-2, the 1998 model, to figure 6-2 above shows how the increased drilling density is delineating the irregular margins as well as the internal geological relationships.

6.4 Mineralization – Lace Kimberlites

As the history section of this report describes, the Lace Main Pipe was a fairly significant producer in the past, and the modern evaluations to date have confirmed compelling commercial potential.

6.4.1 General

As documented above, the historical production and development records (along with the tailings dumps) attest to the mineralization above -240m in the Lace deposits. The lack of geological knowledge of that portion of the pipes does make it difficult to infer which facies were exploited, and in what proportion. Hence all modern work has focused on understanding and defining the geology and grade/value parameters of the pipe(s) beneath -230m.

6.4.2 The Satellite Pipe

The historical miners consistently included the Satellite Pipe in their evaluation drifting as their workings got deeper, but never decided to formally stope the deposit, suggesting sub-economic grades from surface to -240m. A small portion of the Satellite Pipe was mined during the open pitting, but not documented formally. MPH’s limited drilling and microdiamond sampling would tend to

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confirm low grade potential, with low microdiamond counts recovered thus far (MPH, 2005). The database is small however, and it would be worthwhile to drill a few more holes from underground, when next the drill is sited on the west side of the Main Pipe, close to the Satellite Pipe.

6.4.3 The Main Pipe

The historical records, modern microdiamond and bulk sampling have established that commercial grades and diamond values are present for the Main Pipe. Figure 6-3 below portrays the locations of the microdiamond sampling data. As a general statement, the microdiamond sampling has demonstrated continuously mineralized K4, K6 and K8 facies throughout the present geological model, with grade modelling, detailed in Section 14 of this report, providing high-confidence estimates of commercially interesting grades. Bulk sampling on the 250m, 290m and 310m levels, also detailed in Section 14, has confirmed the grade modelling.

The CRB facies has been shown to be significantly diamondiferous as well, but has not been sufficiently micro or bulk sampled thus far to estimate grade.

Figure 6-3 3-D Perspective Looking North of Lace Main Pipe Microdiamonds Sampling

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6.4.4 Commercial Potential

Given its production history from 1896-1931 as described in Section 5.1, it is not surprising that Lace again is on the cusp of commercial production. The modern work on the project is now allowing for a better understanding of what was mined historically, and what future mining holds at depth. It is important to note however that the historical underground grade of 18.9cpht (Table 5-1), which combined with the tailings dump grades achieved by DiamondCorp to date (6.45cpht which understates the dump grades somewhat, Section 5.4.2), was commercially viable until the Great Depression. This in turn suggests that a considerable portion of higher grading K4 was exploited along with the lower grading K6 facies above - 240m.

The delineation drilling and mapping work, and derived geological modelling, has now mapped the K4 “ridge” to -230m, and it is reasonable to assume it reached higher levels. K4 becomes volumetrically more dominant with depth, emphasizing the commercial potential of the deeper underground mine.

6.4.5 Main Pipe Upper K4 Mine -230m to -370m

This depth slice constitutes the upper portion of the unexploited Lace Mine deposit, and is being targeted for early mining by the company’s Upper K4 Mine in development. Microdiamond and bulk sampling results from within this depth slice have been positive, as reported in Section 14.

Impetus for the UK4 Mine operation came from the positive bulk sampling on the 250m level in 2012, combined with the underground delineation drilling and mapping commenced in 2014, which has shown that a compelling amount of the higher grading K4 kimberlite facies is present in the upper portions of the deposit. This depth slice had been previously intersected by very limited portions of three of the 1997-98 core holes drilled from surface, which were designed to avoid the historical workings extending to the -330m level, and hence could not be comprehensively modelled until the completion of the recent work.

6.4.6 Block Caves 1, 2 and 3 -370m to -900m

Data collected to date suggests that grade potential at depth should exceed that of the UK4 portion of the deposit, based on higher proportions of K4 being present, and on less dilution in K6 and less of the CRB facies. On-going drilling from underground will continue to build on the microdiamond, dilution and density databases, which will be combined with audited bulk sampling on the 470m development level. The combined geological, grade and value data, along with production from the UK4 Mine will allow for the Block Cave 1 depth slice, from - 370m to -510m, to next be moved into a higher confidence resource category.

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It is envisaged that the same process will be followed with depth over time, progressively increasing the knowledge and confidence in the resource for Block Cave 2 and 3.

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7.0 DEPOSIT TYPES

Lace as described previously is a diamondiferous Group II kimberlite, one of several in South Africa which have seen commercial mining operations. Group II kimberlites are a sub-set of typical or Group I kimberlites, based on their micaceous nature and geochemical characteristics.

Kimberlite intrusions are often composite, and made up of several different kimberlite lithologies or “facies” which have been intruded during successive eruptive episodes, so that one may cut into the other. The diamond grade and quality of diamonds may vary between different facies and lithologies, therefore a good geological understanding of the kimberlite bodies and lithologically controlled sampling are important in their evaluation.

Kimberlite facies may also vary in the proportion of country rock xenoliths they contain, which impacts on the grade. They may also vary in metallurgical properties, resulting in varying plant efficiencies.

Kjarsgaard (2007) provides the following description of Kimberlite Diamond Deposits:

“Simplified Definition of Deposit Type Primary deposits of diamonds occur in kimberlite as a sparsely dispersed xenocrystal mineral in pyroclastic, volcaniclastic, resedimented volcaniclastic, and subvolcanic (hypabyssal) rocks of mantle origin (kimberlite-hosted deposits). Secondary deposits, formed by the weathering of primary deposits, occur in unconsolidated and consolidated sediments (placers, paleo-placers, and volcanogenic sedimentary deposits).

Scientific Definition of Deposit Type Kimberlite is formed from alkali-poor, H2O- and CO2-rich ultrabasic magma that has an enriched incompatible (Ba, Zr, Hf, Ta, Nb, REE) and compatible (Ni, Co, Cr) trace element signature. These magmas give rise to rocks that form a wide variety of landforms and intrusions, similar to those exhibited by small-volume alkali basalt volcanic systems. Diamond xenocrysts are variably distributed throughout the host rocks at concentration levels of <0.01 to 2.0 ppm.

Deposit Subtypes Lamproite and orangeite (also called Group II kimberlite) also host primary magmatic diamond deposits. Placer and paleo-placer deposits can also be viewed as a subtype.

Global Distribution of Kimberlite Diamond Deposits Major producing or past-producing kimberlite-hosted diamond mines are known from southern Africa (South Africa, Botswana, Zimbabwe), south-central Africa (Tanzania, Democratic Republic of Congo, Angola), western Africa (Sierra Leone), Russia (Yakutia), China, Canada, and the United States. These deposits are all found in Precambrian terranes and specifically in Archean continental blocks. The absence of kimberlite-hosted diamond mines in countries or continents with significant Archean crustal blocks, such as Australia, India, and South America, is notable, especially given the formerly significant paleo-placer and placer deposits in India and Brazil.

Ages of Deposits

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A significant observation is that there are specific time periods in the Earth’s history when kimberlites with economically viable contents of diamonds erupted contemporaneously on different continents, for example, in Tanzania and Canada at ca. 52 to 56 Ma (Davis and Kjarsgaard, 1997) and in South Africa and Canada at ca. 535 to 542 Ma, whereas at other times kimberlite events are restricted to a specific region, e.g., Yakutia at ca. 360 Ma (Fig. 2; Heaman et al., 2003, 2004). The oldest kimberlite-hosted diamond mine is the Cullinan (or Premier) mine in South Africa, which has been dated by radiometric methods at ca. 1200 Ma (Allsopp et al., 1989). The Venetia kimberlite in South Africa is of similar age to the Gahcho Kue 5034 pipe and the Snap Lake sill in the Slave Province, Northwest Territories, Canada (all ca. 542-535 Ma; Heaman et al., 2003). Kimberlites that host diamond mines in China are dated at ca. 475 to 462 Ma (Dobbs et al., 1994). The Yakutian kimberlites that host the economically important diamond mines have been dated by radiometric methods at ca. 376 to 344 Ma (Davis, 1977, 1978; Kinny et al., 1997), whereas the Jwaneng kimberlite in Botswana is ca. 235 Ma in age (Kinny et al., 1989).

Numerous kimberlites hosting diamond mines in southern Africa, for example, Letseng, Lesotho; Jagersfontein, Koffiefontein, Du Toit’s Pan, Bultfontein, De Beers, Kimberley and Wesselton in South Africa and Orapa in Botswana, are of similar age (ca. 95-84 Ma; Davis, 1977). Slightly younger in age are kimberlites that host the Mbuji Maya mine in the Democratic Republic of Congo (71 Ma; Davis, 1977). The youngest kimberlites known to host diamond mines are found at the Ekati and Diavik mines in the Slave province, Northwest Territories (56-53 Ma; Graham et al.,1999; Kjarsgaard et al., 2002; Creaser et al., 2004) and in Tanzania at the Mwadui mine (52 Ma; Gobba, 1989).”

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8.0 EXPLORATION

Exploration work on the project has been summarized above in Section 5-3, and was previously described in detail in MPH (2007), and consisted of surface magnetic and gravity surveys in 1997, regional airborne magnetic surveying in 2006, and hydrographic geophysical surveying of the flooded open pit. As well of course, the surface drilling and tailings bulk sampling, along with compilations of historical production, all served to provide an initial understanding of geology, grade and size potential of the Lace kimberlites.

DiamondCorp’s advanced work, ie. the underground development, bulk sampling and delineation drilling are described in detail in subsequent sections of this report.

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9.0 DRILLING

DiamondCorp’s underground drilling campaign has utilized the Company’s Boart Longyear LM55 core rig, and also the LM30 on early holes. Both rigs utilize BW casing and BQ core barrels. Starting with BH-30 in Table 9-1 below, the larger LM55 has been utilized, with generally better hole length capability and core recovery.

Collar locations have been limited to four “cubbies” presently, with arrays of holes designed to delineate internal and margin geology at nominal 25m spacings (as per Figure 5-5b in previous section). The drilling pattern is further complicated by the still extensive network of historical workings portrayed in fig. 9-1, below. To date, 29 holes totalling 4,422.21m have been completed from underground, of the total 10,183.88m in total on the project in Table 9-1.

Table 9-1 Surface and Underground Delineation Drilling Summary

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Figure 9-1 3-D Perspective Lace Underground Drilling with All Workings

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10.0 SAMPLE PREPARATION, ANALYSIS AND SECURITY

The previous work carried out at Lace has followed normal industry best-practise standards by MPH, as reported in the 2005 and 2007 reports. Latest work standards and protocols are presented here-in.

10.1 Delineation Core Microdiamond Sampling

The 2014/2015 microdiamond sampling work has been processed using standard caustic fusion methods by the SGS Caustic Fusion Facility, and the residues examined by the MSA Microdiamond Laboratory, both in Johannesburg, with the diamonds sieved and weighed to modern mandated disclosure standards for diamond reporting. The bottom screen for this work is +0.075mm.

Samples are selected on geological grounds, with the BQ underground delineation hole samples taking the form of alternating rows from the core box, over the interval of the sample. These are designed to be nominally ~20kg. such that no more than three 8kg. aliquots are necessary at the caustic facility, per sample. No preferential or selective sampling has taken place, and the sample includes all internal country rock dilution inherent in that interval.

Large, mappable intervals of dilution or country rock rafts, are not sampled but are logged and contribute to the dilution database, discussed later in Section 14.

In general the sampling philosophy has been to demonstrate continuous mineralization within facies, with kimberlite intervals sampled from top to bottom in the Crown series holes. Some of the early 2014 underground delineation holes were near to adjacent Crown hole samples and were not sampled, and others have been sampled at selected locations to increase the sampling density per facies, in under-sampled quadrants of the pipe.

All core in the sample interval is collected in a 20 litre plastic bucket lined with a heavy duty plastic bag. A numbered sample tag is placed within each plastic bag; this sample number is also written on the bucket. Project geologists record the hole number, drilled interval, sample number and a brief sample description for the interval. At the end of the sample run the liner bag is sealed with a cable tie and the bucket is sealed with a plastic lid labeled with the sample number.

Lots of samples were shipped from the mine to SGS in Johannesburg periodically. Methodology as reported by MSA is as follows:

Processing and Microdiamond Recovery Procedures (Process 1) The processing and microdiamond analyses were carried out at the caustic fusion and micro-diamond sorting laboratory jointly developed and managed by SGS South Africa (SGS) and MSA. SGS is an approved service provider to MSA and has provided accurate results on several projects carried out in the past.

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The dissolution by caustic fusion (NaOH) of the kimberlite samples is performed by SGS. Diamond is highly resistant to caustic soda and therefore diamond etching and damage is eliminated. SGS provides a concentrate to MSA from which liberated diamonds can be readily extracted by microscopic examination. The caustic fusion residues of the samples were received from SGS from 05 May 2014 onwards (last residue received on 20 May 2014).

The sample weight and number of pours and re-fusions required to dissolve the samples from DiamondCorp are listed below in Table 3 1. The sample weight reduction after caustic fusion was greater than 99.98%. The sample log-in weight is the weight reported to MSA as recorded in the SGS LIMS on receipt of the sample prior to aliquot preparation for caustic fusion. This is the weight to be used in calculating diamond grade.

MSA monitored the quality control and assurance of the caustic fusion process by the addition of synthetic diamond tracers to the sample at the start of the dissolution process. The residue recovered from the caustic fusion process was sorted by MSA for both the synthetic diamond tracers and natural diamonds down to +75 microns.

The standard operating procedure is briefly described below.

• The carbonate content is assessed by testing with hydrochloric acid prior to aliquot preparation and caustic fusion to establish the reactivity of the sample.

• Other than simple breakage of the kimberlite core into +/- 20 mm sized pieces, no other sample preparation is performed prior to dissolution by caustic fusion. • Caustic soda is added to the kimberlite sample in each pot and the kiln is heated to 550ºC. This temperature is maintained for a minimum of fourteen hours.

• After the digestion in molten caustic soda, the sample residue is screened using a bottom screen of 75 microns (μm) as specified by the client.

• The residue, greater than 75 μm, is liberated from the NaOH by washing in hydrochloric acid leach and hot water baths. The bottom screen for the acid leach is 75 μm. The washed residue of each sample, enclosed in the 75 μm screen used during the leaching and washing process, is then dried.

• The dry residue of each sample, wrapped in the 75 μm screen used during the acid leach, is received by MSA. The quality control, sorting and diamond characterisation is done by MSA.

• Quality control throughout the process is monitored by spiking with sized synthetic diamonds that are easily identifiable. The synthetic diamond spikes are added to the sample at the start of the caustic fusion process.

• The synthetic diamonds used to monitor the process efficiency for each sample are selected from the following 2 size fractions: -212 μm to +150 μm and -150 μm to +106 μm. This is done at the discretion of MSA.

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• Natural and synthetic diamonds are recovered from the +75 μm residue using 60X magnification with a binocular microscope. The residue is examined a minimum of two times to ensure the total recovery of diamond. If required re- checking of the residue is undertaken to ensure the recovery of all diamonds from the sample.

• Recovery rate of the spikes (synthetic diamonds) is reported and the recovered spikes are stored on sample cards.

Synthetic diamonds, either client spikes or released into the sample from diamond drilling, are identified stored on sample cards and reported.

• The recovered natural diamonds are separated into 13 sieve classes by screening and stored on sample cards after weighing.

• Colour, clarity, and morphology of each diamond is determined and reported.

• X Y and Z dimensions of each diamond are measured in mm.

• All macrodiamonds and each microdiamond greater than 300 μm are weighed individually using a 7 place electronic microbalance and placed on sample cards. The microdiamonds smaller than 300 μm are weighed in groups and their combined weight is reported.

• The gram weight is converted to carats.

• Diamond data is tabulated in Excel spreadsheets. The Excel versions are supplied simultaneously to the client, together with a PDF version of the Test Report.

• All the residues and recovered diamonds are stored for a fixed period as arranged with the client.

Reports by MSA are comprehensive, providing both individual and totalled stone counts and weights, per CIM square sieve class. Additionally MSA provides a breakdown for the entire allotment of crystal shapes, resorption characteristics, colours and clarity.

10.2 Bulk Sample Microdiamond Sampling

Representative microdiamond samples are collected for each bulk sample underground, and as well are taken from the stockpile as random pieces of rock, before the bulk sample is processed.

Underground, the methodology generally consisted of Mr. Paul Allan mapping in detail the sidewall and hanging wall faces, and the development face. For each bulk sample, a microdiamond sample is collected underground from the sidewalls of the respective

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sample area. The samples are collected by extracting pieces of kimberlite by hand along the entire length of the sample (safety permitting) to make up as representative a sample as possible. The samples are then carefully labeled using sample tickets and sealed in two plastic bags for transport by vehicle to the surface. Once at the surface the samples were placed in 20 litre plastic seal-able drums for transport to SGS Lakefield in Johannesburg. Each sample weighed approximately 20 to 25 kg., ie. similar to the average core sample size.

10.3 Bulk Sample Macrodiamond Sampling

The mining method selected by DiamondCorp for the UK4 Mine, Bottom Up Longhole Open Stoping, discussed in detail in Section 16.0, includes development of a doming level on -290m and the production level on -310m, which have allowed for systematic bulk sampling, monitored by MPH, as the drives have been developed. A plan was put in place between DiamondCorp and MPH before kimberlite mining commenced, to break the drives down into manageable, representative bulk samples of ~500- 1,000tonnes, geologically controlled and to be processed separately as individual batches by the Lace plant.

10.3.1 Bulk Sample Mining and Stockpiling

DiamondCorp’s mining development tunnels are nominally 4.5m x 4.5m, such that the original layout of proposed bulk samples was for 16m lengths, delivering some 325 cubic metres of kimberlite to surface. In practise the lengths, widths and heights varied according to conditions and mining sequences, however all bulk samples slots are surveyed before the initial blast for a new sample takes place, such that volumes are precisely known.

Figure 10-1 Bulk Sample Development on 310m Level

Remedial Support in Cubby 4 (note 2mx2m historical tunnel) Blast Round in K6 Kimberlite

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All material within the bulk sample is hauled to surface, and stockpiled and labelled separately on a purpose demarcated pad near the plant’s feed bin as per the photos below for the first bulk sample 290-01. Stockpiles are surveyed to determine bulking factors.

Figure 10-2 Bulk Sample 290-01 Stockpile before Processing

The bulk sample drives are mapped in detail, including systematic dilution measurements, and all data recorded.

10.3.1 Bulk Sample Processing and Security

Before processing the plant is purged of any material, the only exception being the rock bed in the VSI (vertical spindle impact, ie. re-crush) crusher. Tracer tests are carried out for both the DMS and Final Recovery modules (grease belts).

The final recovery and diamond sort are monitored at all stages and signed off by all DiamondCorp managers and MPH. All of the bulk test diamonds are picked from final recovery concentrates inside secure glove boxes. Diamond parcels are weighed and counted using standard DTC sieve sizes. Weights and stone counts were checked and signed off by a member of DiamondCorp senior management and a representative of MPH Consulting before being deposited in a drop safe within the Lace mine sorthouse.

Figure 10-3 below presents an example of a bulk sample processing form.

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Figure 10-3 Bulk Sample 290-01 Processing Checklist

The Lace plant process description is described in detail in Section 17.0. All coarse tailings (-6mm) are stored separately at the mine site, appropriately labelled, such that they are available should further work be needed on this material as per Figure 10-4.

Figure 10-4 Bulk Sample Coarse Tailings Stockpiles

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Sorthouse tails are stored in a secure area for further reference (Figure 10-5).

Figure 10-5 DMS/Recovery Tailings Stockpiles

Security on the mine is contracted to a reputable firm which is used by many diamond producers. There are dedicated security officers assigned to the final recovery and when sorting, senior personnel are present. There is also a ‘roving’ security officer moving around the plant during processing. All screens, pumps and sumps handling concentrate are locked and security seals installed – these can only be removed in the presence of security who record the reason for opening the equipment. This applies to the final recovery as well.

There are 15 security cameras strategically positioned in the process plant and 22 cameras in the final recovery. There is also a central control room where these cameras are monitored.

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11.0 DATA VERIFICATION

All of the previous 1997-98 work discussed above has been verified to the extent possible, in the sense that original laboratory analytical reports have been utilized, and all of the laboratories and consulting groups used by MPH then, and DiamondCorp/MPH now, are reputable, well known and respected firms to the diamond industry. MPH was acting as independent consultant then for previous client Rupert Resources Ltd., a Canadian junior mining company listed on the TSX Venture exchange, and all reports were to Canadian regulatory standards at that time and verified accordingly.

11.1 Geological Model

The previous geological model has been completely superseded by the new work reported here-in. That being said though, original kimberlite facies identified in 1997 have been appropriately continued to be used, although nomenclature updated as follows:

Old Nomenclature New Nomenclature Reason TKB (Tuffisitic Fragmental Fragmental is a totally non genetic term for a Kimberlite / Breccia) Kimberlite / Breccia “mixed up kimberlite” ie it consists of various fragments some from the mantle and some from the country rock. Hypabyssal Coherent Again coherent is a totally non genetic term for a “crystalline looking kimberlite” ie it crystallized from a molten / magmatic state.

Summary of the original typical kimberlite designation characteristics.

Kimberlite Characteristics K2 Hypabyssal (coherent) Not segregationary on a macroscopic scale Generally < 15% CRX’s (Country Rock Xenoliths), mostly <5 % CRX’s. Mauve – brown micaceous matrix Up to 21% Olivine Macrocrysts (generally 10 to 15%) K4 Coherent Kimberlite with a generally segregationary or globular segregationary texture. 5 to 20% CRX’s, Average abundance around 15%. CRX’s are generally small (<3cm) but do range up to 25cm in size. K6 35 to 90% CRX’s, Average in the 50% range. CRX’s larger than K4

K2 and K4 are very similar except for the minor difference in texture and the abundance of CRX’s, and are combined as “K4” in the geological model.

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11.2 Drill Core Logs

All of the original Crown series holes have been re-examined to ensure continuity of interpretation by Mr. Allan, and as well to discern whether the new K8 facies (now known to correlate with historical “hardebank” plugs) had been overlooked in previous logging. It had not been intersected by these previous holes.

Mr. Sobie travels to the mine 3-4 times per year and spends several days reviewing all new core holes, and all underground development mapping, with Mr. Allan, before the geological logs and mapping are considered final and entered into the database.

11.3 Internal Dilution Data

For core holes, dilution is measured along a single straight line over the entire length of the borehole core. All waste rock fragments greater than 0.5cm in size are included in the measurement and recorded as the number of cm of waste per metre (ie. the percentage of dilution per metre of core). The results are stored in a Dilution Database that includes records of the various lithologies.

For the bulk samples, country rock dilution is measured as follows: country rock xenoliths (“CRX”) larger than 30cm in size are mapped on longwall maps of the sidewalls, hanging wall and development faces. An estimate of the dilution in the remaining matrix (i.e. a value of the dilution for all clasts less than 30cm) was either taken from the closest borehole intervals (measured on a line count) or from direct line counts on the development faces where possible. In each case fragments larger than 0.5cm were measured. The total dilution was then calculated by taking the large CRX dilution and adding the dilution in the remaining matrix.

This total CRX dilution value is recorded in the database over the entire length of the bulk sample, in the position of the corresponding microdiamond sample. If no microdiamond sample has been collected for that bulk sample, then the total dilution value is recorded in the database along the face that was mapped, at a relative elevation of 1.5m (ie. ~ chest height), above the floor of the tunnel.

11.4 Bulk Density Data

All core holes from the present delineation drilling program have had systematic density measurements taken, by the industry best-practice displaced water method, on average every 15m. Competent ~10cm pieces of whole core are precisely weighed before and after varnishing, and then immersed in a water filled beaker and the mass of displaced water also precisely weighed. Care is taken to ensure that representative populations of samples for all geological facies have been collected, and continue to be added to the database with each completed hole.

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As well, the density samples are dried overnight in an oven set at 120 degrees to discern whether material differences exist between wet and dry samples. Results are reported in Section 14.1.

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12.0 MINERAL PROCESSING AND METALLURGICAL TESTING

The Lace plant is described in detail in Section 17.0 of this report. It is important to note that this is a mine-scale production plant, not a bulk sample plant, which is being run in batch mode for the individual bulk samples. No tailings auditing has taken place, such that grades for the bulk samples are recovered grades, with no modifying factors or factorization.

Lace is carrying out test work on mixed ore-waste feed separation through optical sorting technology developed by Steinert Elektromagnetbau GmBh of Germany. “The Unisort equipment is based on the latest NIR (near infrared) camera technology, Hyper Spectral Imaging (“HIS”). Its advantages are due to the combination of extremely high spatial and spectral resolution.” (IMS Engineering Unisort PR Brochure).

The technology is based on the variation in the absorbance of NIR wavelengths (800 to 1700nm) between different lithologies, with the latest very significant improvement in the technology being the ability to measure the absorbance over 256 points (over a 25mm particle) as opposed to the initial 16 points. One is thus able to scan the whole belt as opposed to the earlier line sensors. The smallest particle size suitable is around 15mm and the throughput possible is in the order of 170tph.

As reported by DiamondCorp on 13 October, 2015, this technology has the potential to remove the majority of internal waste in especially the heavily dilute K6 facies, before it reaches the processing plant.

As also reported on 13 October, 2015, Lace introduced a de-grit circuit to the flowsheet during Q2 2014, with the bottom screen size having been increased from 1.00 mm to 1.25 mm. The de-grit circuit receives its product from the primary scrubber outlet trommel (double skin with 1.00 mm panels on the outer skin) – before it gets to the secondary scrubber and primary screen. All of the early 290m level bulk samples (to 290-09), and 310-01, were processed under this configuration, ie. with the bottom screen size at 1.25 mm.

The removal of this sand fraction results in up to a 50% reduction in water consumption in the processing plant, which is an important operational consideration at full production as water resources in the mine area are limited. While the sand fraction contains significant numbers of small diamonds, they are the lowest value stones and their recovery for sale is considered by management to be a break even exercise at best in the foreseeable future.

To confirm this, in October 2015 the plant was reconfigured to 1.00mm screen panels and the balance of the bulk samples were processed accordingly, with results reported in the resource estimate section of this report (Section 14).

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13.0 ADJACENT PROPERTIES

There is one nearby diamond mine to Lace, De Beers’ Voorspoed Mine which is an open pit operation that went into production in 2008 and is currently on Cut 4, with a planned ultimate pit depth of 420m and a mine life of 12-16 years. The mine was discovered in 1906 and exploited for several years before a large basalt xenolith (“floating reef”) was encountered which compromised most of the open pittable resource, and it closed in 1912. It sat dormant until the late 1990’s when De Beers instituted a detailed delineation drilling program, prompted by the successful drilling at Lace. The pipe is 12.5ha in diameter, with the dominant kimberlite a volcaniclastic facies similar to Lace’s K6. It is also a Group II kimberlite and has been dated at 131.06Ma using the 40Ar/39Ar technique of ground mass phologopite grains (M.Field et al., 2008).

The mine reportedly produced ~700,000 carats in 2014, and official probable reserves as at December 31, 2014 sat at 8Mt grading 23.7cpht with indicated resources of 9.1Mt grading 26.2cpht, for a total of 4.3 million carats in those categories. A further 20.3Mt grading 19.2cpht is categorized as inferred resources (Anglo American AR 2014).

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14.0 MINERAL RESOURCE ESTIMATE

This section summarizes the databases and methodologies used for updating the mineral resource estimate for the Lace Main Pipe diamond deposit. The previously reported mineral resource estimate (VP3, 2012) was based in large part on the drilling and early geological modelling and microdiamond grade estimation work of MPH in 1997-98, supplemented by VP3’s observations on the 2009 and 2012 bulk sampling. There was a large component of supposition in the MPH work, as there was only the historical production records, tailings dump bulk sampling results, and early expert interpretations of the microdiamond sampling database to work with. Additionally the geological model constructed by MPH in 1998 was not at all well constrained as discussed earlier in this report, therefore MPH considers that all earlier resource statements should be disregarded, and the present work regarded as the first modern estimate.

The present work has been designed to systematically evaluate the -230m to -370m UK4 depth slice of the deposit, and allows for diamond industry-standard estimation methodologies to be employed. Data is much sparser below -370m, however the geological model has been updated based on the better knowledge and interpretation above, and the resource work at depth also brought up to date. Resources will be updated periodically to higher confidence levels, as each successive depth slice is comprehensively delineated and bulk sampled.

14.1 Evaluation Databases

Databases for drill hole data, dilution measurements, density measurements, microdiamond sampling and macrodiamond bulk sampling have been constructed for the project. These incorporate all of the 1997-98 work, as well as all recent work. The microdiamond database utilized here-in is summarized in Table 14-1 following.

14.1.1 Drill Hole Database

Delineation drill holes utilized were summarized previously in Table 5-6. Geological logs, header, survey, core recovery, rock quality designation (RQD) and sampling data are all maintained in a Gemcom database, constantly updated and validated.

Density and dilution measurement protocols were described Section 11, and the databases are discussed in detail in in Sections 14.4 and 14.5 respectively.

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Table 14-1 Lace Mine Main Pipe Delineation Microdiamond Samples

+0.15mm +0.075mm +0.15st Mine HOLE-ID SAMPLE-ID FROM TO WEIGHT KG LITHOLOGY STONES CARATS X - Middle Y - Middle Z - Middle stones stones / 25kg. Level CROWN-1 2701 641.00 646.00 17.85 K6South 5 0.0006350 5 7 12320.56 -3036971.75 773.05 557 CROWN-1 2703 647.00 661.50 25.14 K4 26 0.0156050 26 26 12319.36 -3036967.32 763.33 567 CROWN-1 2707 713.50 721.70 15.76 K6South 3 0.0011700 3 5 12313.06 -3036940.86 706.15 624 CROWN-1 2708 721.70 731.00 18.98 K4 34 0.0138650 34 45 12312.30 -3036937.15 698.27 632 CROWN-1 2709 731.00 741.00 19.56 K4 22 0.0159650 22 28 12311.48 -3036933.06 689.57 640 CROWN-1 2710 741.00 751.00 21.06 K4 67 0.0926400 67 80 12310.67 -3036928.79 680.57 649 CROWN-1 2710B 741.00 751.00 20.00 K4 203 0.0114772 203 12310.67 -3036928.79 680.57 649 CROWN-1 2711 751.00 761.00 19.82 K4 30 0.0156000 30 38 12309.89 -3036924.51 671.57 658 CROWN-1 2712 761.00 771.00 17.77 K4/K6 15 0.0120350 15 21 12309.15 -3036920.22 662.57 667 CROWN-1 2713 771.00 794.40 27.13 K4/K6 15 0.0170850 15 14 12307.98 -3036913.01 647.56 682 CROWN-1 2714 794.40 813.40 25.93 K4 41 0.0098450 41 40 12306.62 -3036903.80 628.51 701 CROWN-1 2715 815.90 834.90 26.15 K4 50 0.0185550 50 48 12305.31 -3036894.44 609.20 721 CROWN-1 2716 839.50 858.40 25.14 K4 50 0.0285000 50 50 12303.86 -3036884.18 588.05 742 CROWN-1 2717 859.30 878.40 25.62 K4 17 0.0072050 17 17 12302.65 -3036875.51 570.18 760 CROWN-1 2717B 859.30 878.40 20.00 K4 54 0.0031790 54 12302.65 -3036875.51 570.18 760 CROWN-1 2718 879.30 898.70 26.11 K4/K2 43 0.0228150 43 41 12301.41 -3036866.73 552.09 778 CROWN-3 2721 476.30 487.50 17.90 K4 0 0.0000000 0 0 12262.66 -3036806.37 903.22 427 CROWN-3 2722 494.30 509.40 23.92 K4 33 0.0434450 33 34 12265.28 -3036813.51 884.78 445 CROWN-3 2723 529.00 543.00 26.61 K6North 10 0.0013000 10 9 12270.04 -3036825.69 853.24 477 CROWN-3 2723B 529.00 543.00 20.00 K6North 15 0.0001007 11 12270.04 -3036825.69 853.24 477 CROWN-3 2724 544.00 555.00 22.30 K6North 8 0.0025800 8 9 12271.94 -3036830.55 840.79 489 CROWN-3 2725 556.00 576.00 20.42 K6North 12 0.0024500 12 15 12274.25 -3036836.48 825.56 504 CROWN-3 2726 608.00 618.00 20.90 K2 8 0.0070050 8 10 12280.51 -3036853.22 782.09 548 CROWN-3 2727 618.00 628.00 19.30 K6North 3 0.0005600 3 4 12281.92 -3036856.77 772.85 557 CROWN-3 2728 634.50 646.00 22.10 K6North 10 0.0126700 10 11 12284.42 -3036862.77 756.87 573 CROWN-3 2729 654.00 667.00 22.79 K6North 4 0.0014400 4 4 12287.55 -3036869.66 738.09 592 CROWN-3 2730 667.00 682.00 24.63 K6North 5 0.0049050 5 5 12289.79 -3036874.44 725.12 605 CROWN-3 2731 683.00 692.00 19.76 K6North 0 0.0000000 0 0 12291.66 -3036879.01 713.10 617 CROWN-3 2732 722.50 730.20 19.07 K6North 6 0.0018150 6 8 12296.86 -3036892.52 677.05 653 CROWN-3 2733 730.20 743.00 24.05 K4 34 0.0476600 34 35 12298.26 -3036896.00 667.51 662 CROWN-3 2733B 730.20 743.00 20.00 K4 22 0.0005540 22 12298.26 -3036896.00 667.51 662 CROWN-3 2747 743.00 757.50 22.96 K4/K2 31 0.0074750 31 34 12300.30 -3036900.61 654.82 675 CROWN-3 2748 758.40 775.80 25.00 K2 35 0.0220150 35 35 12302.90 -3036906.29 639.18 691 CROWN-3 2754 830.80 849.70 24.83 K4/K2 28 0.0088050 28 28 12313.78 -3036932.10 571.61 758 CROWN-3 2755 864.40 885.60 28.40 K2/K4 43 0.1183450 43 38 12318.55 -3036944.26 539.41 791 CROWN-3 2756 941.00 958.10 17.70 K6South 1 0.0004600 1 1 12329.30 -3036970.88 470.61 859 CROWN-6 2764 549.00 558.00 21.89 K6North 44 0.0042200 44 50 12222.33 -3036893.43 848.87 481 CROWN-6 2765 559.00 568.00 22.67 K6North 6 0.0011550 6 7 12226.18 -3036892.09 839.74 490 CROWN-6 2766 569.00 578.00 22.19 K6North 6 0.0051800 6 7 12230.04 -3036890.79 830.61 499 CROWN-6 2767 581.00 590.00 21.70 K6North 0 0.0000000 0 0 12234.67 -3036889.26 819.65 510 CROWN-6 2768 599.00 608.00 21.76 K6North 22 0.0023500 22 25 12241.61 -3036887.04 803.20 527 CROWN-6 2769 613.00 623.00 22.68 K2 43 0.0136700 43 47 12247.20 -3036885.33 789.94 540 CROWN-6 2770 623.00 633.00 18.54 K2/K4 13 0.0223200 13 18 12251.05 -3036884.18 780.79 549 CROWN-6 2771 635.00 644.00 22.54 K6North 3 0.0014850 3 3 12255.48 -3036882.90 770.26 560 CROWN-6 2774 674.00 683.00 22.37 K2 18 0.0025550 18 20 12270.49 -3036878.84 734.52 595 CROWN-6 2775 684.00 693.00 20.78 K2 16 0.0036100 16 19 12274.34 -3036877.87 725.35 605 CROWN-6 2776 694.00 704.00 21.34 K2 22 0.0782850 22 26 12278.38 -3036876.89 715.71 614 CROWN-6 2777 705.00 714.00 22.97 K2 29 0.0101350 29 32 12282.42 -3036875.94 706.07 624 CROWN-6 2778 730.00 739.00 19.03 K6North 20 0.0231700 20 26 12292.04 -3036873.80 683.11 647 CROWN-6 2779 740.00 761.00 23.52 K6North 37 0.0145150 37 39 12298.19 -3036872.53 668.39 662 CROWN-6 2780 762.00 777.00 19.85 K4 41 0.0235300 41 52 12305.49 -3036871.07 650.91 679 CROWN-6 2781 778.00 799.00 25.39 K2 37 0.0115350 37 36 12312.77 -3036869.51 633.43 697 CROWN-6 2782 800.00 821.00 26.03 K4 23 0.0122150 23 22 12321.22 -3036867.72 613.20 717 CROWN-6 2804 822.00 832.00 23.41 K4 8 0.0022450 8 9 12327.58 -3036866.18 598.05 732 CROWN-7 2783 318.50 328.80 22.64 K6North 4 0.0007650 4 4 12357.84 -3036842.11 1048.01 282 CROWN-7 2784 328.80 343.00 25.52 K4 41 0.0202600 41 40 12353.02 -3036843.45 1036.83 293 CROWN-7 2785 343.00 356.00 24.60 K4 20 0.0205650 20 20 12347.58 -3036844.80 1024.44 306 CROWN-7 2786 356.00 366.70 22.48 K4/K2 37 0.1099550 37 41 12342.76 -3036845.88 1013.67 316 CROWN-7 2787 367.00 376.50 25.07 K6North 2 0.0006000 2 2 12338.47 -3036846.62 1004.22 326 CROWN-7 2788 377.20 391.00 24.90 K2/K4 17 0.0067400 17 17 12333.39 -3036847.73 993.02 337 CROWN-7 2789 391.00 404.00 24.96 K4 26 0.0097450 26 26 12327.79 -3036848.90 980.90 349

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+0.15mm +0.075mm +0.15st Mine HOLE-ID SAMPLE-ID FROM TO WEIGHT KG LITHOLOGY STONES CARATS X - Middle Y - Middle Z - Middle stones stones / 25kg. Level CROWN-7 2790 404.00 423.00 24.85 K4/K2 16 0.0186350 16 16 12321.07 -3036850.25 966.45 364 CROWN-7 2791 423.00 435.00 23.84 K4 35 0.0924250 35 37 12314.52 -3036851.55 952.46 378 CROWN-7 2792 435.00 446.80 23.05 K4 23 0.0214600 23 25 12309.37 -3036852.38 941.76 388 CROWN-7 2793 450.00 464.00 24.58 K4 41 0.0104600 41 42 12302.30 -3036853.26 927.33 403 CROWN-7 2794 464.00 478.00 25.78 K4 60 0.0376900 60 58 12296.11 -3036854.41 914.83 415 CROWN-7 2795 478.00 489.90 22.76 K4 60 0.1296800 60 66 12290.31 -3036855.27 903.28 427 CROWN-7 2796 498.00 512.50 23.96 K4 35 0.0231850 35 37 12280.70 -3036856.70 884.33 446 CROWN-7 2797 512.50 525.00 25.00 K2 17 0.0678300 17 17 12274.51 -3036857.62 872.37 458 CROWN-7 2798 525.00 538.00 24.68 K4 33 0.0085750 33 33 12268.62 -3036858.66 861.11 469 CROWN-7 2799 538.00 554.00 26.69 K4 48 0.0238100 48 45 12261.83 -3036859.81 848.35 482 CROWN-7 2800 559.00 570.00 25.73 K6North 4 0.0026100 4 4 12253.22 -3036859.91 832.10 498 CROWN-7 2801 571.00 582.00 24.02 K6North 4 0.0006250 4 4 12247.54 -3036860.67 821.56 508 CROWN-7 2802 583.00 597.00 24.02 K6North 14 0.0034850 14 15 12241.15 -3036861.75 809.71 520 CROWN-7 2803 598.00 618.00 25.74 K6North 3 0.0022800 3 3 12232.58 -3036863.13 793.95 536 CROWN-8 2805 326.00 337.00 25.60 K6South 0 0.0000000 0 0 12386.40 -3036904.08 1034.82 295 CROWN-8 2806 339.00 350.00 25.00 K6South 7 0.0025450 7 7 12381.89 -3036905.81 1022.75 307 CROWN-8 2807 351.00 360.00 21.07 K6South 4 0.0015250 4 5 12378.07 -3036907.27 1012.54 317 CROWN-8 2808 362.50 370.50 21.26 K6South 10 0.0039250 10 12 12374.29 -3036908.75 1002.32 328 CROWN-8 2809 503.80 518.00 27.11 K6South/K4 17 0.0077750 17 16 12325.03 -3036929.37 868.17 462 CROWN-8 2810 518.00 531.00 7.90 K2/K4/K6 29 0.0120950 29 92 12320.31 -3036931.38 855.57 474 CROWN-8 2811 539.00 554.50 20.61 K6South/K4 4 0.0027200 4 5 12312.55 -3036934.57 834.96 495 CROWN-8 2812 556.00 580.00 28.19 K4 87 0.0163450 87 77 12304.86 -3036937.90 815.44 515 CROWN-8 2813 581.00 605.00 22.64 K6South/K4 29 0.3657150 29 32 12295.76 -3036941.70 792.47 538 CROWN-8 2814 616.00 659.00 26.01 K6South/K4 25 0.0364800 25 24 12278.87 -3036948.61 751.88 578 CROWN-8 2815 678.00 697.00 24.60 K6South 18 0.0030100 18 18 12259.82 -3036955.84 706.23 624 CROWN-8 2816 698.00 720.00 27.35 K6South 13 0.0090550 13 12 12251.64 -3036958.76 686.56 643 CROWN-8 2817 726.00 773.00 26.31 K6South 10 0.0019200 10 10 12236.46 -3036964.25 649.42 681 CROWN-8 2818 774.00 795.00 26.88 K6South 7 0.0012000 7 7 12223.28 -3036968.97 617.34 713 CROWN-8 2819 803.00 823.00 24.75 K6South 4 0.0013850 4 4 12212.37 -3036972.87 591.30 739 CROWN-8 2820 827.00 846.00 25.34 K6South 8 0.0106250 8 8 12203.25 -3036975.91 569.86 760 CROWN-8 2821 847.00 866.00 24.04 K6South 9 0.0051700 9 9 12195.52 -3036978.55 551.60 778 CROWN-8 2822 867.00 887.00 25.18 K6South 16 0.0026500 16 16 12187.37 -3036981.16 532.98 797 290-03 B9 22.33 28.33 15.60 K6North 7 0.0015494 5 7 8 12241.66 -3036956.02 1060.64 269 335-05 B11 25.31 35.81 14.94 K2 38 0.0098095 20 38 33 12260.76 -3036956.72 1067.54 262 335-05 B12 37.75 51.20 16.44 K2 15 0.0092887 9 15 14 12271.29 -3036963.05 1061.01 269 335-07 B3 9.98 15.98 14.92 K6North 1 0.0029360 1 1 2 12248.46 -3036941.69 1074.82 255 335-07 B4 27.98 33.98 13.76 K6North 5 0.0002773 1 5 2 12263.47 -3036940.11 1065.01 265 335-07 B5 39.98 45.98 14.06 K4 18 0.0419530 14 18 25 12273.48 -3036939.06 1058.48 272 335-07 B6 87.98 93.98 15.16 K6South 4 0.0005249 4 4 7 12313.51 -3036934.85 1032.33 298 335-07 335-7-1 114.98 120.98 16.81 K6South 11 0.0055484 3 11 4 12336.03 -3036932.48 1017.63 312 335-08 B1 36.00 42.98 14.43 K4 35 0.0048465 20 35 35 12269.10 -3036933.80 1059.80 270 335-08 B2 48.98 54.98 15.64 K8 26 0.0361505 18 26 29 12279.05 -3036930.95 1052.82 277 430-03 B7 56.52 66.52 18.04 K4 7 0.0257508 7 7 10 12244.94 -3036966.76 1025.68 304 430-03 430-3-1 85.20 95.00 21.78 K6North 111 0.0123388 32 111 37 12248.34 -3036977.87 999.57 330 430-03 B8 121.58 127.58 16.04 K6North 8 0.0236213 6 8 9 12252.44 -3036991.28 968.08 362 430-05 B10 55.43 61.43 11.04 K4 20 0.0245541 12 20 27 12261.91 -3036957.42 1030.78 299 430-05 430-5-1 76.00 85.00 17.38 K4 149 0.0464922 39 149 56 12271.08 -3036962.93 1011.48 319 430-06 430-6-1 43.00 49.60 26.34 K4 22 0.0042751 14 22 13 12264.25 -3036948.48 1044.43 286 430-06 430-6-2 145.00 155.10 21.90 K6South 91 0.1458879 57 91 65 12323.90 -3036961.16 960.49 370 430-07 B13 31.28 41.48 15.16 K6North 2 0.0050685 2 2 3 12261.37 -3036940.33 1054.43 276 430-07 B14 42.90 52.70 15.10 K8 2 0.0007369 2 2 3 12268.82 -3036939.55 1045.81 284 430-07 B15 54.50 64.28 14.50 K8 8 0.0252097 8 8 14 12276.38 -3036938.75 1037.06 293 430-07 430-7-2 67.40 76.20 20.60 K8 60 0.0052500 24 60 29 12284.48 -3036937.90 1027.70 302 250-K4 250-K4 face rep sample 43.70 K4 68 0.0716241 42 68 24 12282.35 -3036872.29 1077.25 253 260-K4 260-K4 face rep sample 41.10 K4 77 0.1452658 48 77 29 12296.77 -3036907.00 1084.97 245 BH-30 L - 1914 98.1 108.7 15.30 K6South 5 0.0002613 2 5 3 12342.42 -3036906.14 1049.28 281 BH-30 L - 1915 110.2 120.7 14.90 K6South 8 0.5735700 4 8 7 12337.63 -3036895.88 1045.16 285 BH-30 L - 1916 126.6 135.5 14.04 K4 34 0.0064464 13 34 23 12331.44 -3036882.59 1039.82 290 BH-30 L - 1917 138.5 149.1 15.74 K6South 22 0.0180508 14 22 22 12326.38 -3036871.74 1035.46 295 BH-30 L - 1918 153.6 160.9 15.36 K2 5 0.0173350 3 5 5 12321.89 -3036862.11 1031.60 298 BH-30 L-1801 153.6 160.9 2.96 K2 31 0.0014147 8 31 68 12321.89 -3036862.11 1031.60 298 BH-30 L - 1919 165.4 175.8 15.22 K6North 2 0.0004421 2 2 3 12315.73 -3036848.91 1026.30 304 BH-31 L - 1913 211.6 221.5 12.60 CRBNorth 33 0.1459504 26 33 52 12255.42 -3036836.06 1010.58 319

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+0.15mm +0.075mm +0.15st Mine HOLE-ID SAMPLE-ID FROM TO WEIGHT KG LITHOLOGY STONES CARATS X - Middle Y - Middle Z - Middle stones stones / 25kg. Level BH-33 L - 1901 136.6 145.5 14.10 K2 79 0.1227033 40 79 71 12285.50 -3036932.98 1003.74 326 BH-33 L - 1902 147.3 157.4 13.58 K8 4 0.0122900 2 4 4 12277.65 -3036928.07 997.26 333 BH-33 L - 1903 158.8 174.1 21.54 K4 119 0.1116983 54 119 63 12267.85 -3036921.95 989.17 341 BH-33 L - 1904 175.5 198.3 20.28 K4 143 0.0488525 46 143 57 12253.65 -3036913.07 977.44 353 BH-33 L - 1905 200 216.2 20.22 K4/6/2/8 56 0.0294840 23 56 28 12238.92 -3036903.87 965.28 365 BH-33 L - 1906 219.1 234.6 23.86 K8 16 0.0297375 6 16 6 12225.89 -3036895.73 954.53 375 BH-33 L - 1907 237.7 248 11.66 Shale/CRBNorth 4 0.0000225 0 4 0 12214.78 -3036888.79 945.35 385 BH-34 L - 1908 138 156.4 20.62 K4 150 0.3233338 99 150 120 12301.54 -3036896.55 1011.05 319 BH-34 L - 1909 173 186.9 19.78 K4 78 0.0159951 32 78 40 12283.31 -3036874.82 994.68 335 BH-34 L - 1910 188.4 194 9.68 K8 3 0.0000399 0 3 0 12277.04 -3036867.36 989.05 341 BH-34 L - 1911 196 209.4 17.84 K6North 23 0.0146843 10 23 14 12270.64 -3036859.73 983.30 347 BH-34 L - 1912 211.1 254 12.56 CRBNorth 2 0.0000834 2 2 4 12254.03 -3036839.92 968.38 362 BH-29 L - 1920 19.00 29.40 13.88 LRKB / K8 3 0.0002922 1 3 2 12257.76 -3036894.70 1073.89 256 BH-29 L - 1921 35.30 45.90 14.08 K4 54 0.0270953 26 54 46 12269.85 -3036884.75 1069.03 261 BH-29 L - 1922 47.20 59.80 14.46 K4 21 0.0118825 19 21 33 12279.37 -3036876.92 1065.22 265 BH-29 L - 1923 61.30 74.60 17.78 K4 5 0.0140870 5 5 7 12290.03 -3036868.15 1060.94 269 BH-29 L-1802 61.30 74.60 20.92 K4 139 0.0405518 36 139 43 12290.03 -3036868.15 1060.94 269 BH-29 L - 1924 74.60 87.80 18.44 K2/K4 49 0.0191540 20 49 27 12299.80 -3036860.10 1057.02 273 BH-29 L - 1925 89.50 100.80 13.68 K2/K4 53 0.0433206 19 53 35 12310.09 -3036851.64 1052.89 277 BH-29 L - 1926 102.30 115.00 13.74 K6North 9 0.0017549 5 9 9 12320.05 -3036843.44 1048.90 281 BH-29 L - 1927 116.60 128.00 14.70 K6North 29 0.0022541 9 29 15 12330.12 -3036835.16 1044.86 285 Bulk Sample 290-01 L - 1929 face rep sample 19.94 K6South 18 0.0055412 8 18 10 12322.17 -3036944.71 1047.00 283 Bulk Sample 290-02 L - 1931 face rep sample 21.18 K6South 20 0.0125450 7 20 8 12314.38 -3036948.60 1047.00 283 Bulk Sample 290-03 L - 1933 face rep sample 21.06 K6South 31 0.0014557 11 31 13 12318.25 -3036932.73 1047.00 283 Bulk Sample 290-04 L - 1935 face rep sample 21.94 K6South 9 0.0013110 2 9 2 12314.62 -3036921.42 1047.00 283 Bulk Sample 290-05 L - 1937 face rep sample 22.46 K6South 12 0.0005298 5 12 6 12308.58 -3036902.99 1047.00 283 Bulk Sample 290-06 L - 1939 face rep sample 22.38 K4 150 0.1651253 58 150 65 12301.30 -3036880.78 1047.00 283 BS 290-08 East Drive East face rep sample 40.56 K4 133 0.0311100 54 133 33 12309.93 -3036879.79 1047.00 283 BS 290-09 West Drive West face rep sample 40.56 K4 248 0.1522130 98 248 60 12293.05 -3036885.20 1047.00 283 BS-290-10 Underground L-1818 face rep sample 19.00 K4 48 0.0124898 26 48 34 12296.62 -3036866.53 1047.00 280 BS-290-11 Underground L-1819 face rep sample 23.42 K4 95 0.0703278 56 95 60 12292.42 -3036853.71 1047.00 280 BS-250-03 Underground L-1812 face rep sample 28.20 85%K6South/15%K4 144 0.0220168 49 144 43 12284.81 -3036944.11 1085.00 245 BS-250-04 Underground L-1813 face rep sample 29.34 85%K6South/15%K4 24 0.0302558 11 24 9 12302.69 -3036933.20 1085.77 244 BS-250-07 Underground L-1814 face rep sample 29.30 95%K6North/5%K4 32 0.0019871 8 32 7 12239.92 -3036945.08 1082.14 248 BS-250-08 Underground L-1815 face rep sample 30.20 95%K6North/5%K8 11 0.0009485 6 11 5 12231.36 -3036901.32 1077.96 252 BS-250-09 Underground L-1816 face rep sample 26.66 88%KNorth/12%K4 40 0.0079244 18 40 17 12239.34 -3036867.23 1076.27 254 BS-310-01 Underground L-1803 face rep sample 22.24 K6South/CRB 7 0.0003310 1 7 1 12325.76 -3036956.00 1023.00 307 BS-310-Cubby1 U/G L-1807 face rep sample 22.78 K6South 14 0.0075063 10 14 11 12328.50 -3036939.00 1023.00 307 BS-310-02 Underground L-1805 face rep sample 20.68 K6South 14 0.0040341 6 14 7 12318.53 -3036933.82 1023.00 307 BS-310 Cubby2 U/G L-1809 face rep sample 24.20 K6South 31 0.0067968 13 31 13 12309.50 -3036931.00 1023.00 307 BS-310-03 Underground L-1811 face rep sample 27.00 K6South 26 0.0221938 16 26 15 12311.60 -3036912.55 1023.00 307 BS-310 Cubby3 U/G L-1810 face rep sample 22.86 K6South 16 0.0049127 6 16 7 12318.50 -3036915.25 1023.00 307 BS-310-04 Underground L-1820 face rep sample 28.26 92.5%K6S/7.5%K4 72 0.0099413 42 72 37 12304.59 -3036891.20 1023.00 307 BS-310 Cubby4 U/G L-1817 face rep sample 24.48 70%K6S/25%K8/5%K4 30 0.0305037 10 30 10 12301.25 -3036907.25 1023.00 307

168 Totals 3614.69 5397 4.7812038 Samples (excl. stockpiles)

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14.1.1 Bulk Sample Database

As of 11 December, 2015, on-going bulk sampling on the -290m and -310m levels had totalled 9,910 tonnes and 8,479 tonnes respectively. As well, the database includes the previous bulk sampling work on the -250m level. Results are provided in tables 14-2a and 2b following, with the level plans portrayed in Figs. 14-1 and 14-2 respectively. The 250m level work was presented earlier in Table 5-5 and Figure 5-4.

14.2 Geological Modelling

The geological model has been constructed at 1:500 scale using 3-D rings on 20m level plan intervals from 230m to 500m, and 40m level plan intervals from 500m to 900m. MPH subcontracted this work to consulting Gemcom specialist Ms. Hayley Manning of Dassault Systemes Canada Software Inc., who worked at the MPH office and used MPH’s software for the exercise, under the direction of Mr. Sobie.

The geological interpretation using detailed mapping and borehole traces on each level was digitized, and all 3-D rings were updated to adhere to clipped drill hole intersects for all levels. The model includes nine geological wireframe solids representing the four main geological domains as per the table below from Ms. Manning’s report, attached as Appendix II.

Geological Domain Solid Name # of Solids Rock Type Block Model Code Kimberlite 4 K4 2 K4 40 Kimberlite 6 K6 1 K6 60 Kimberlite 8 K8 1 K8 80 Country Rock Breccia CRB 5 CRB 100

Interim and final versions of each level plan were reviewed with project geologist Mr. Paul Allan before final inclusion in the model. The entire model was validated by preparing cross-sections on 25m intervals to ensure geological interpretations and contacts were consistent. The present model was presented in previous figures, and in Figure 14-3 following:

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Table 14-2a Lace Mine Main Pipe 290m Level Bulk Samples

Lace Mine 290m Level Bulk Sample Results (+1.25mm Bottom Screen Size) Bulk Sample Results (+1.00mm Bottom Screen Size) Sample 290-01 290-02 290-03 290-04 290-05 290-06/07 290-08East 290-09West COLLAPSE (290-08 and 09) 290-10 290-11 Totals Lithology Mixed K6/CRB K6 South K6 South K6 South K6 South K4 K4 K4 K4 K4 ROM Tonnes 1105 1280 417 440 1280 1707 497 1455 483 363 2214 11,241 Total Carats 34.887 96.283 24.952 19.217 59.806 501.939 162.468 388.168 175.76 121.925 719.923 2,270.44 Grade (cpht) 3.16 7.52 5.98 4.37 4.67 29.40 32.69 26.68 36.39 33.59 32.52 Parcel 8671 8672 8673 8674 8675, 8676 8677 to 8681 8684, 85, 87, 90 and 8690 8688, 89, 91 to 8695 8691, 8698, 8699 8721 - 8722 8727 - 8730 Sorting Date 16/04/2015 21/04/2015 28/04/2015 30/04/2015 18/06/2015 30/06/2015 10/07/2015 17/07/2015 22/07/2015 to 30/07/2015 30/07/2015 to 11/08/2015 31/7 to 15/9/2015 23/11 - 24/11/2015 30/11/2015-11/12/2015 Diamond Mesh Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats +150 CT +100 CT +60 CT +45 CT +30 CT +20 CT +15 CT +10 CT 2 31.246 1 10.498 +9 CT +8 CT +7 CT 1 7.862 2 14.796 +6 CT +5 CT 1 5.161 1 5.834 2 11.776 1 5.834 +4 CT 1 4.33 1 4.249 1 4.207 0 0.000 2 8.437 +3 CT 2 6.748 1 3.084 0 0.000 1 3.314 3 10.216 +2.5 CT 1 2.504 1 2.979 1 2.524 0 0.000 2 5.556 0 0.000 4 10.574 +2 CT 1 2.12 1 2.213 0 1 2.493 6 13.959 1 2.432 3 6.290 2 4.639 5 11.273 +1.5 CT 1 1.56 3 4.890 1 1.558 0 0 0.000 8 14.127 2 3.837 4 6.784 3 5.351 6 10.608 +1.25 CT 1 1.33 1 1.258 1 1.439 0 0.000 3 4.092 2 2.681 5 6.890 1 1.420 1 1.35 9 12.305 +21 (+1 CT) 2 2.43 2 2.488 0 0 4 3.555 8 8.889 1 1.077 5 7.704 2 2.203 3 3.13 18 19.976 +19 5 3.12 14 9.350 7 4.778 6 3.365 14 8.631 67 45.039 19 13.854 52 32.016 18 11.440 17 10.38 126 85.405 +15 11 5.25 26 11.140 8 3.322 8 3.045 20 7.633 166 63.643 43 11.799 135 50.060 43 17.837 52 20.05 221 83.639 Coarse Total 22 20.13 49 66.779 18 13.129 16 10.828 40 24.836 264 184.940 69 38.764 209 131.283 68 52.038 73 34.91 396 267.229 +11 39 8.47 80 17.475 28 6.034 18 3.978 71 15.516 599 125.215 159 32.014 520 99.260 211 39.559 149 29.12 925 180.459 +9 27 3.30 54 6.679 29 3.553 19 2.349 71 8.814 706 82.956 189 21.483 658 71.484 287 30.085 215 22.70 941 107.385 +8 17 1.43 31 2.710 13 1.143 7 0.626 35 3.016 403 33.957 89 6.892 317 24.929 165 13.117 137 10.57 532 44.424 +7 13 0.84 18 1.179 11 0.729 13 0.878 52 3.485 604 38.999 164 9.406 496 31.053 294 17.830 199 12.31 869 54.933 +5 17 0.67 35 1.415 9 0.364 12 0.478 77 3.110 757 30.790 390 12.287 633 25.169 457 18.046 253 10.83 984 42.725 +4 2 0.05 2 0.046 2 0.043 23 0.533 130 3.243 285 5.799 102 2.589 110 2.842 27 0.77 152 4.166 -4 0.00 0.000 2 0.037 27 0.496 100 1.839 1589 29.053 139 2.401 128 2.242 35 0.713 121 2.448 Fines Total 113 14.758 220 29.504 90 11.823 73 8.389 357 34.970 3299 316.999 2864 116.934 2865 256.885 1372 123.721 1015 87.014 4524 436.540 Total (Coarse + Fine) 135 34.887 269 96.283 108 24.952 89 19.217 397 59.806 3563 501.939 2933 155.698 3074 388.168 1440 175.759 1088 121.925 4920 316.273 Mida Sample U/G L-1929 L-1931 L-1933 L-1935 L-1937 L-1939 290EastDrive 290WestDrive L-1818 L-1819 Mida Sample Stockpile L-1928 L-1930 L-1932 L-1934 L-1936 L-1938 N/A N/A * Grit fraction processed

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Table 14-2b Lace Mine Main Pipe 310m Level Bulk Samples

Lace Mine 310m (+1.25mm Bottom Screen Size) Lace Mine 310m Level Bulk Sample Results (+1.00mm Bottom Screen Size) Sample 310 - Sample1 310 - Sample1 (1.00mm) 310-02 310-Cubby 1 310-Cubby 2 310-03 310-Cubby 3 310-04 310-Cubby 4 Totals Lithology Mixed K6/CRB Mixed K6/CRB K6 K6 K6 K6 K6 92.5% K6 / 7.5% K4 70% K6 / 5% K4 / 25% K8 ROM Tonnes 1654 919 868 500 514 1046 1002 1822 154 8,479 Total Carats 102.11 62.46 55.354 28.165 27.80 68.45 62.63 236.28 13.823 554.963 Grade (cpht) 6.17 6.80 6.38 5.63 5.41 6.54 6.25 12.97 8.98 Parcel 8700 to 8706 8707 to 8708 8709 - 8710 8711 - 8712 8713 - 8714 8715, 8716, 8719 8717 - 8718 8720, 8723, 8724, 8726 8725 Sorting Date 16/09/2015 to 03/10/2015 09/10/2015 to 16/10/2015 17/10 to 19/10/2015 20/10 to 21/10/2015 22/10 to 23/10/2015 2/11 to 19/11/2015 17/11 to 18/11/2015 20/11 to 30/11/2015 27/11/2015 Diamond Mesh Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats Total Stones Total Carats +150 CT +100 CT +60 CT +45 CT +30 CT +20 CT +15 CT +10 CT +9 CT +8 CT +7 CT +6 CT +5 CT +4 CT 1 4.476 +3 CT 1 3.053 +2.5 CT 1 2.741 1 2.82 2 5.442 +2 CT 1 2.027 0 0 2 4.344 3 6.766 5 10.825 1 2.169 +1.5 CT 3 5.198 1 1.678 0 0 1 1.5 +1.25 CT 0 0 1 1.403 0 0 3 4.077 1 1.385 2 2.641 3 3.998 +21 (+1 CT) 0 0 2 2.392 3 3.443 0 0 3 3.503 3 3.322 2 2.362 7 7.889 +19 18 11.794 21 12.972 11 6.654 7 4.619 3 2.295 9 8.921 11 7.478 31 20.01 +15 47 17.413 18 7.353 16 6.366 11 3.191 8 2.396 23 8.28 20 7.762 85 31.871 5 2.258 Coarse Total 69 36.432 44 28.539 33 23.627 21 11.887 15 9.579 42 23.163 37 28.844 135 84.588 6 4.427 +11 145 27.871 73 14.173 66 13.398 32 6.388 42 8.158 110 20.982 62 12.383 315 60.96 15 3.519 +9 131 14.408 89 9.599 69 7.835 36 4.013 36 4.074 95 10.357 79 8.681 350 37.422 28 2.857 +8 87 6.737 43 3.206 35 2.773 23 1.683 27 2.171 48 3.762 38 2.964 198 15.516 12 0.867 +7 138 8.105 63 3.749 62 3.868 39 2.414 32 1.97 93 5.642 74 4.584 304 18.671 16 1.02 +5 181 7.258 69 2.991 78 3.091 36 1.411 36 1.428 95 3.881 112 4.565 394 16.22 24 1.014 +4 34 0.863 6 0.141 21 0.581 6 0.185 11 0.279 14 0.385 8 0.219 72 1.927 3 0.093 -4 27 0.44 4 0.06 11 0.181 10 0.184 10 0.14 15 0.277 20 0.39 48 0.979 2 0.026 Fines Total 743 65.682 347 33.919 342 31.727 182 16.277 194 18.22 470 45.286 393 33.786 1681 151.695 100 9.396 Total (Coarse + Fine) 812 102.114 391 62.458 375 55.354 203 28.165 209 27.799 512 68.449 430 62.63 1816 236.283 106 13.823 Mida Sample U/G L-1803 L-1803 L-1805 L-1807 L-1809 L-1811 L-1810 L-1820 L-1817

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Figure 14-1 Bulk Sampling on 290m Level

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Figure 14-2 Bulk Sampling on 310m Level

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Figure 14-3 3-D Perspectives Lace Main Pipe Geological Model

Figure 14-3a – Lace Mine Geological Model Figure 14-3b – Lace Mine Geological Model Looking Northwest Looking South

14.2.1 Geological Continuity

Drilling to date is showing predictable continuity to the major facies, shown diagrammatically in Figure 14-3 to 14-5. At this juncture, confidence is very high in the geological continuity of facies, as well as pipe morphology, in the UK4 Mine depth slice of the deposit, and less so with depth where there is less drilling information.

That being said, logging thus far does not suggest major changes with depth, other than K4 becoming massive (ie. more K2) rather than globular-segregationary in texture, and some indications that K6 becomes less dilute. Also the very dilute CRB facies seems to not be present at depth. That being said, drilling density is low as per the various figures, such that geology, pipe shape and size will become better defined with time and more drilling.

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Figure 14-4 Level Plan Compilation Lace Main Pipe Geological Model -250m to -900m

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Figure 14-5 Section Compilation through Lace Main Pipe Geological Model

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14.2.2 Volume and Tonnage Estimates

The Gemcom geological and block model have established in-situ volumes and tonnages for the various units to an elevation of 430m ASL, or -900m below surface, broken down into the mine plan blocks of Lace Diamond Mines (Pty.) Limited. Densities utilized in Table 14-3 are as per the in-situ (wet) specific gravity database.

Table 14-3 Lace Mine Main Pipe Volume and Tonnage Estimate

KIMBERLITE VOLUME MINING BLOCK DENSITY TONNES % OF TOTAL FACIES (m3x1000) Upper K4 Mine 230-370m Levels K4 1,065.486 2.585 2,754,281 28.1% (1110-960m absl) K6 1,834.957 2.563 4,702,995 48.0% K8 144.722 2.641 382,211 3.9% CRB 723.803 2.709 1,960,782 20.0% Total 3,768.968 9,800,269 100.0% Block Cave 1 370-510m Levels K4 1,626.754 2.59 4,213,293 48.4% (960-820m) K6 1,262.561 2.56 3,232,157 37.1% K8 13.713 2.64 36,203 0.4% CRB 451.786 2.71 1,224,339 14.1% Total 3,354.815 8,705,993 100.0% Block Cave 2 510-700m Levels K4 2,225.776 2.59 5,764,760 59.0% (820-630m) K6 1,484.048 2.56 3,799,164 38.9% K8 0.000 2.64 - 0.0% CRB 74.018 2.71 200,589 2.1% Total 3,783.842 9,764,513 100.0% Block Cave 3 700-900m Levels K4 2,800.965 2.59 7,254,499 71.0% (630-430m) K6 1,153.812 2.56 2,953,759 28.9% K8 0.000 2.64 - 0.0% CRB 3.577 2.71 9,694 0.1% Total 3,958.354 10,217,952 100.0% Total 14,865.979 38,488,727

KIMBERLITE VOLUME HISTORICAL WORKINGS EXTRACTED DENSITY TONNES FACIES (m3x1000) Upper K4 Mine 230-370 Levels K4 19.538 2.585 50,506 (1110-960m absl) K6 15.364 2.563 39,378 K8 5.412 2.641 14,293 CRB 1.605 2.71 4,350 WASTE 2.697 - Total 44.615 108,526

Modelling suggests some 110,000 tonnes had been removed historically.

14.2.1 Confidence Level of Geological Model

The overall reliability of the geological model is considered by MPH to be moderate to high, for the UK4 Mine depth slice of the deposit (-230m to -370m, or 1110m to 960m absl), based on the density of drilling information, and bulk sampling on three levels. Geological contacts were found to be within 5m of modelled positions from the drilling data, when exposed by underground tunnelling. MPH is of the opinion

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that confidence is of the level of Indicated Mineral Resource for much of the UK4 Mine depth interval, with the exception of the Northwest quadrant of the pipe, where the CRB facies is not constrained to the same density of drilling, and is therefore appropriately at the Inferred Mineral Resource level of confidence.

Below -370m, to the bottom of the present geological model, confidence is lower and appropriately categorized as Inferred Mineral Resources.

14.3 Block Model

The Lace block model has been created between -230m and -370m as per Figure 14-6 following, with model parameters including the lateral extents and block size determined by the mining method and layout for the UK4 Mine. The blocks have been oriented as per the mine tunnel layout and are 20m x 20m x 20m as appropriate for the mining method. The blocks are centred about the -310m extraction level layout.

Appendix 2 provides the full report on modelling parameters.

Block properties are as follows:

Origin Rotation Block Size Number of Blocks X-12172 Columns – 20m Columns – 16 Y- -3037066 18 degrees Rows – 20m Rows – 14 Z- 1107 counter-clockwise Levels – 20m Levels – 8

A multiple folder format was used, such that each geological domain (K4, K6, K8 and CRB), as well as waste, have a folder. Model attributes include rock type, percent of block inside geology solid, interpolated grade, interpolated density and interpolated dilution (all with three variants using Nearest Neighbour, Inverse Distance and Ordinary Kriging methodologies), as well as confidence, number of holes used, number of points used and average distance to points used.

The rock type and percent attributes were updated using the ‘Update from Solids’ command in GEMS using an integration level of 25 with horizontal along columns needle orientation. The minimum percent entered to re-assign the block was 0.01%.

14.3.1 Bulk Density Data

A total of 434 density measurements comprise the current database, of which 422 are within the four main geological solids, above -370m and are shown schematically in Figure 14-7 below.

Statistics for the raw density values by geological domain are provided below in Table 14-4:

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Figure 14-6 3-D Perspective Lace Main Pipe K4 Facies Block Model Facing Northeast

Table 14-4 Lace Mine Main Pipe Raw Density Measurements

K4 K6 K8 CRB All Number of measurements 72 256 33 61 422 Mean 2.58 2.6 2.63 2.71 2.62 Standard Deviation 0.08 0.16 0.08 0.18 0.15 Coefficient of Deviation 0.031 0.062 0.031 0.067 0.059 Maximum 2.83 3.1 2.82 3.39 3.39 Upper Quartile 2.62 2.5 2.69 2.82 2.71 Median 2.57 2.58 2.64 2.71 2.6 Lower Quartile 2.54 2.72 2.57 2.61 2.52 Minimum 2.34 2.2 2.43 2.18 2.18

As described earlier in the report, drying of density samples took place and statistics recorded. The country rock lithologies show a very small decrease in SG after drying (0.2 to 0.4%) in contrast to the kimberlite facies which indicate decreases in SG ranging between 1.48 and 2.93%.

K2 showed the least decrease on drying possibly because of its highly competent olivine rich nature. This was followed by K6 with 1.94% and K8 with a loss of 2.51%. For both of these dilute lithologies, the SG is likely to be largely controlled by the high abundance of country rock lithologies (which showed the least decrease in SG on drying). K4 showed a slightly higher decrease in SG on drying of 2.93%

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possibly due to the larger amount of matrix (more weathered than K2) and the relatively low abundance of country rock xenoliths (especially in comparison to K6).

Figure 14-7 3-D Perspective Lace Main Pipe Density Measurements Facing North

Given the current water table and the degree of saturation of the old workings the raw method (taking SG as is without drying) is regarded as the most accurate reflection of the current in – situ SG.

14.4 Bulk Dilution

Compositing was carried out on the dilution data, which was collected every metre down most holes, although certain gaps were present where microdiamond samples had been collected from early holes. For the bulk samples, a single estimate of dilution along the entire length of the sample was calculated by Mr. Allan based on the methodology described earlier in Section 11.3

14.4.1 Bulk Dilution Data

Dilution drill hole data was examined at 3m, 4m and 5m composite lengths, and found to be best distributed relative to the raw data at the 3m composite length, with the others introducing smoothing. The composite intervals honoured the geological domain boundaries. Dilution values through compositing were found to vary from the raw values as follows (Appendix 2 - Table 5.4):

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Geological Domain Raw Mean Composited Mean Variance Dilution Dilution K4 29.1% 33.2% 14.1% K6 65.0% 65.9% 1.4% K8 57.6% 65.8% 14.3% CRB 91.8% 91.2% -0.7%

Figure 14-8 3-D Perspective Lace Main Pipe Dilution 3m Composites Facing North

14.4.2 Bulk Density and Dilution Estimation

Bulk densities interpolated through variography results (Appendix 2 – Section 6) and using ordinary kriging, made use of the following variogram and search ellipse orientations (Appendix 2, Tables 6.1 and 6.2):

Variogram Parameters: Grade Structure Nugget Sill Major Semi-Major Minor Co 0.002 Density Spherical 1 - 0.004 15.5 10.213 9.292 Spherical 2 - 0.018 58.2 38.348 34.890 Co 2.000 Dilution Spherical 1 - 169.00 21.0 16.835 11.293

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Spherical 2 - 301.00 98.0 78.565 52.701

Search Ellipse Parameters for Density and Dilution (rotations relative the block model):

Search Size Grade Rotn. Rotn. Rotn Range1 Range2 Range3 Confidence AboutZ aboutX aboutZ (m) (m) (m) Level

Medium Density 4.5 90 0 58 38 35 2 Medium Dilution -68 -45 0 98 78.5 53 2

Large Density 4.5 90 0 116 77 70 3 Dilution -68 -45 0 196 157 105 3

Average bulk densities applied to the blocks are in Table 14-3 are from the inverse density interpolation and are 2.585 for K4, 2.563 for K6, 2.641 for K8 and 2.709 for CRB.

Considerably more dilution has been noted in the southeastern quadrant of the pipe, which is most obvious in Figure 14-9b below, showing the K6 facies dilution block model.

Figure 14-9a 3-D Perspective Lace Main Pipe K4 Dilution Blocks Facing Northeast

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Figure 14-9b 3-D Perspective Lace Main Pipe K6 Dilution Blocks Facing Northeast

14.4.3 Confidence Level of Bulk Density and Tonnage Model

The overall reliability of the density, dilution and tonnage models is considered by MPH to be moderate to high, for the UK4 Mine depth slice of the deposit (-230m to -370m, or 1110m to 960m absl), based on the density of drilling information, and bulk sampling access on three levels. MPH is of the opinion that confidence is of the level of Indicated Mineral Resource for much of the UK4 Mine depth interval, with the exception of the Northwest quadrant of the pipe, where the CRB facies is not constrained to the same density of drilling, and is therefore more appropriately at the Inferred Mineral Resource level of confidence.

MPH/Dassault validated the model by comparing block and composite density and dilution values on level plans and sections and 3-D inspections, with the resultant estimated values (inverse distance for) corresponding very well with the composite values (Appendix 2 – Figure 9-2). Similarly, the results for dilution corresponded very well corresponded very well (Appendix 1 – Figure 9-3).

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Block model results for both the inverse distance and ordinary kriging methods were compared to declustered composite grades. Overall, for all domains, the estimated mean density value was 0.4% lower for the inverse distance method, and 0.8% lower for ordinary kriging, than the sample mean. The estimated mean dilution grade was 1.1% lower for the inverse density method and 1.2% lower for the ordinary kriging method, then the composite mean.

Below -370m, to the bottom of the present geological model, density confidence is lower and appropriately categorized as Inferred Mineral Resources.

14.5 Diamond Grade

Diamond grade within the block model has been estimated by a combination of macrodiamond recoveries from bulk samples, and microdiamond recoveries from drill core and the bulk samples, with a recovered diluted grade estimated for each block based on dilution. It is important to bear in mind that the Lace processing plant is a commercial scale production facility, such that macrodiamond results represent true recovered grades.

The estimated recovered grade includes inherent average matrix dilution, which values have been calculated from detailed measurements of bulk sample facies by Mr. Allan. This grade value is factorized by overall dilution in the block, which is further incorporating the large blocks of internal dilution. The model uses the following average matrix dilution values:

K6 Facies - 54.7% K4 Facies - 22.3% K8 Facies - insufficient data at present CRB Facies - insufficient data at present

Therefore block model grade is calculated as estimated grade * [100 – (total dilution – matrix dilution)/100] within that block. For K8 facies, block model grade is using the estimated microdiamond grade, which incorporates all dilution. CRB is presently being given a recovered grade of zero until bulk sampling has been carried out. Microdiamond data has demonstrated significant, but highly variable mineralization being locally present within the CRB facies.

14.5.1 Approach and Sampling

At the onset of this work the choice was between performing the calculations in terms of mining blocks or in terms of lithology. During the course of the analysis it became apparent the best approach would be to concentrate on individual lithology, but to include depth intervals to retain some of the changes observed with depth. It

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is felt that this approach provides all the information required to state the resource also in mining block format. The two basic components of diamond content are diamond size expressed in terms of a size frequency distribution and modelled by means of a Log probability graph, and diamond density that is expressed in terms of the average number of stones per kilogram or ton. With the aim being to obtain diamond content by lithology the entire sampling data base was split accordingly in order to model diamond size and density by lithology.

As a consequence of this approach it becomes clear which domains in the mine fall short in terms of data availability and the associated confidence in estimates.

With more than 3,000kg of material treated for diamond recovery by means of Caustic fusion and more than 4,500 stones actually used in this analysis, it is natural that more emphasis is placed on diamond content estimates based on microdiamond methodology.

Lace is showing itself as being ideal for this type of analysis with high stone counts and typical Lognormality in terms of the distribution of diamond size. Grade is directly related to stone density (stones/100kg) and different diamond size distributions are apparent in the lithological composition of the body. It is understood that there are geological reasons for it being an ideal source of microdiamonds.

14.5.1.1 Bulk Sample Macrodiamond Data

Little less than 35,000 tons of material has been treated for the recovery of diamonds for valuation and grade estimation, yielding 5,027 carats (+4 sieve). Table 14-5 is a summary of results from sampling at mine levels 250, 290, and 310. For the moment the table does not distinguish recovery between mine levels but this will be considered in the analysis section.

Different bottom cut-off sizes were used and for comparison recoveries are shown at +9 sieves, assuming that bottom screen effects are not present in this size fraction.

No macro diamond results are available for exclusively CRB and K8 bulk samples to date.

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Table 14-5 Summary of Lace Mine Bulk Sample Macrodiamonds

Bottom Carats/100t Carats/100t cut-off Lithology Note: Carats +9 Carats +4 ROM Tons +9 +4 mm

1.25mm K4 909.1 1,186 22 29 4,142

1mm K4 641.8 823 25 32 2,577

1mm K6K4 some K6 673.2 1,109 12 19 5,843

Sub total 2,224 3,117 18 25 12,562

1.25mm K6 CRB 110.6 137 4 5 2,759

1.25mm K6 South 180.0 200 5 6 3,417

1mm K6 some K4 929.9 1,382 7 10 13,924

1mm K6 CRB 133.8 192 6 8 2,318 Sub total 1,354 1,910 6 9 22,418 1mm K6K4K8 10.8 14 7 9 154

Total 3578.4 5,027 n/a n/a 34,980

14.5.1.2 Core Microdiamond Data

A total of 3.2tons of material has been treated to date for the recovery of diamonds by means of caustic fusion, yielding 4693 stones above 0.075 mm. Different bottom cut-off screen sizes were used during the sampling programs. In the analysis the main focus was on stone counts and diamond size distribution and where it was obviously necessary only +0.212mm recovery results were used. This was mostly the case with some of the older (pre-2000) microdiamond samples necessary. This size fraction is believed to be fully populated and without any effects caused by different bottom screens.

Microdiamonds sampling results were sorted by lithology and in mining blocks and is summarized in Table 14-6, showing recovery results above .075mm and 0.212 mm. Samples were combined by lithology and broken down into zones corresponding to the depths of mining blocks and block caves.

Analysis is based on stone counts only and grade is based on diamond density measured in stones per 25 kg. Stone densities are shown for both bottom cut-off sizes and inspection shows why it is necessary to restrict the use of microdiamonds to +0.212mm.

Stone densities in Table 14-6 give an indication of differences in diamond occurrence with depth and between facies, but must be combined with the

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associated diamond size distributions for a complete reflection of diamond content.

Table 14-6 Summary of Lace Mine Delineation Microdiamonds with Depth

Stones/25kg Sample Total Stones LITHOLOGY Weight +0.212 +0.075 (kg) +0.212 +0.075 CRB North 15 26 36.8 22 39 K2 230 to 700 21 38 254.9 214 388 K2/K4 230 to 700 23 46 111.9 102 204 K4 230 to 290 25 82 349.0 350 1142 K4 290 to 350 31 83 183.5 230 610 K4 350 to 510 33 50 257.4 339 511 K4 510 to 700 36 48 176.7 251 341 K4 700 to 900 25 31 152.3 151 189 K4/K2 290 to 900 20 32 186.4 152 241 K6 North 230 to 290 3 13 110.6 15 57 K6 North 290 to 510 6 18 278.5 69 197 K6 North 510 to 700 9 11 286.0 99 130 K6 South 230 to 290 3 19 136.8 16 103 K6 South 290 to 350 9 22 280.1 103 250 K6 South 350 to 900 5 9 255.8 47 94 K6 South/K4 12 19 96.4 46 75 K8 8 24 126.9 39 122 Grand Total 17 36 3279.7 2245 4693

14.5.2 Diamond Size Distribution from Bulk Sampling

The bulk sampling data was grouped by lithology and depth and Log probability graphs (LP-graphs) were plotted to investigate whether there are differences between facies or with depth within a lithology.

LP-graphs for all the samples are shown in Figure 14-10

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Figure 14-10 Log-Probability Graphs for Bulk Samplng by Lithology

The data shows two sets of distributions and are plotted separately in 14-11.

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Figure 14-11 Facies Reflecting Fine and Coarse Diamond Size Distributions

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The finer size distributions are shown on the top graph. On the lower graph the coarser size distributions are all on and above level 290. All graphs are shown at +9 sieve to eliminate bottom cut-off effects. No differences between the other groups are apparent, but when eliminating the plots for K6 CRB from Figure 14-11 it is shown that K6 South on level 290 has a significantly coarser size distribution in Figure 14-12 below.

Figure 14-12 Indication of Coarser Size Distribution for K6 South Level 290

The K6 South bulk sample comprises more than 3000 tons, suggesting that the coarser size distribution is unlikely to be due to sample size. The coarseness of material from K6 South on level 290 is confirmed by microdiamond results in the following section.

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Table 14-7 Bulk Samples Showing Fine and Coarse Combined Samples

Grade Lithology and Elevation ROM Tons Total Carats (carats/100t) K6 1mm 250 8172 911.59 11 K6 1mm 310 5752 478.68 8 K6K4 1mm 250 5997 1130.60 19 K4 1.25mm 290 4142 1228.33 30 K4 1mm 290 2577 841.85 33 K6 CRB 1.25mm 310 1654 102.11 6 K6 CRB 1mm 310 919 62.46 7 FINE COMBINED 29213 4755.63 K6 South 1.25mm 290 3417 200.26 6 K6 CRB 1mm 250 1399 129.07 9 K6 CRB 1.25mm 290 1105 34.89 3 COARSE COMBINED 5921 364.22 Total 35134 5119.84

14.5.3 Diamond Size Distribution from Microdiamonds

LP-graphs were plotted for the combined samples listed in Table 14-8 and shown in Figure 14-13 below.

Notably some combined samples are smaller and give a false impression of size, like CRB North, but generally the samples are expected to provide some indication of diamond size in the domain.

The graphs were systematically eliminated and regrouped and eventually came together into three groups that could be modelled together. (Although small the CRB North sample may well confirm the coarseness of the CRB bulk samples.)

The analysis shows that distribution of diamond size within a litho-facies is a major factor in the determination of diamond grade. The three groups of samples were analysed individually, followed by combining the two coarser groups and finally analysing the groups as one combined sample.

The combinations are shown in Table 14-8.

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Figure 14-13 LP-Graphs for Combined Microdiamond Samples by Lithology & Depth Zone

Table 14-8 Microdiamond Samples Grouped With Respect to Distribution of Diamond Size

Stones Stones Kg LITHOLOGY 2360 1700 1180 850 600 425 300 212 150 106 75 +0.075 / 100kg Weight

K2/K4 Above 350 2071 314 660 0 2 15 46 79 116 181 257 326 481 568 K6 South Below 260 70 374 1 1 1 9 18 20 40 47 72 29 22 350 GROUP 2 2331 225 1035 1 3 16 55 97 136 221 304 398 510 590 K2/K4 Below 350 1555 154 1012 0 6 5 15 68 138 312 549 315 71 76 K6 North Below 224 47 472 0 0 0 3 5 19 41 81 55 17 3 CRB350 North 39 106 37 0 0 2 8 6 1 2 3 6 4 7 GROUP 1 1818 120 1520 0 6 7 26 79 158 355 633 376 92 86 K8 122 96 127 0 0 3 2 1 4 12 17 22 28 33 K6 South above 350 262 66 395 0 0 0 3 8 11 17 36 57 56 74 K6 North above 350 203 98 207 0 0 1 1 4 5 14 15 39 53 71 GROUP 3 587 81 728 0 0 4 6 13 20 43 68 118 137 178 GROUP 2 & 3 2957 164 1800 1 3 22 69 116 157 266 375 522 651 775

GROUP 1 ,2 & 3 4736 144 3283 1 9 27 87 189 314 619 1005 892 739 854

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The table shows grouped sample stone densities as well as the mm size breakdown, which is used for size modelling. Samples in a group are meant to be modelled individually for grade on the basis of diamond density, but they will have the same size distribution model and consequently the same Dollar per carat value.

LP-graphs for the three groups are shown in Figure 14-14, each group compared with the total combined sample.

Figure 14-14 LP-Graphs for Microdiamond Lithological and Depth Groupings

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In view of the total combined sample mass in each group, the differences are significant. Diamond size was therefore modelled per group and used to represent the samples within the group. LP-graphs for the three groups of combined samples are shown in Figure 14-15.

Figure 14-15 LP-Graphs for Group1, 2, 3 and Group 2 & 3 Combined Populations

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The samples in Group 1 clearly indicate a finer size distribution.

The samples in Group 1 reflect the size distribution of the total combined sample, ie Groups 1, 2 and 3.

Group 3 samples are coarse, and even when combining Group 2 with Group 3 the combined group still reflects the size distribution of combined Group 3. As a consequence, the only benefit of considering combined Group 2 and 3 is to see the effect of the coarser size distribution on grade for the samples in Group 2 and is discussed in the results section.

14.5.3.1 Group 1 Size and Grade-Size Modelling

The linearity of the sample LP-graphs is an indication that diamond size is Lognormally distributed, therefore size modelling was focused on obtaining suitable Lognormal parameters per group.

A suitable Lognormal distribution for diamond size is obtained by simulating individual stones to form a diamond parcel with a specific size distribution. The simulated parcel is compared with the combined microdiamond sample and the procedure is repeated until the size distributions for simulated parcel and sample coincide exactly. Comparisons are done by means of LP-graphs as shown below. The simulated parcel is regarded as typical parcel with respect to diamond size and represents the samples in the Group.

The modelling procedure is completed by simulating individual samples containing stones at selected stone density, thus forming a sample with typical grade-size distribution. The average stone density for each domain in the group is used and a grade-size distribution is obtained for each domain within the group. The end result of this procedure is that the domains in each group has a common size distribution, but each has its own unique grade.

Figure 14-16 shows the LP-graph for Group 1 samples with the typical parcel for the group. The combined sample comprises 1,181 stones recovered from 1.52 tons of material.

The red graph on the left is the typical parcel at +0.075mm lying on top of the graph for the combined microdiamond sample, showing that the model represents the size distribution as reflected by the microdiamond sample.

To ensure that the model also represents diamond size in the larger size classes the typical parcel is compared with bulk sampling results, which is broken down into Christensen size classes. Since simulation is done in DTC carat-grainer size classes, the comparison is restricted to the +1 carat size

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Figure 14-16 LP-Graphs for Group 1 Microdiamonds with Bulk Sample

range to retain only size classes that coincide with the DTC carat-grainer size classes and are not affected by bottom cut-off losses. The comparison is illustrated by the graphs on the right in Figure 14-16.

The model and bulk sample coincide and suggests acceptance of the microdiamond size distribution model. The size distribution model is fully coherent with the microdiamonds and with a 1.5tonne sample the model is accepted as representative of the domain represented by the microdiamonds.

(Note the microdiamond points relate to the stone counts in mm size classes and in the simulation is extended upwards into DTC carat-grainer size classes. The bulk samples relate to a Christensen sieving breakdown used at the Mine, which is not suitable for revenue estimation. Therefore, in the simulation the DTC size breakdown is used for +0.85mm stones to coincide with the 2016 updated Dollar per carat class values, which makes it possible to use the typical parcel for revenue calculation.)

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The size distribution model was subsequently used to generate a typical sample comprising 1million x 100kg microdiamond ‘samples’, assuming the combined group microdiamond sample stone density. The typical sample reflects the distribution of grade with size and is depicted by means of a grade-size graph plotting size class grade expressed as stones/100t against diamond size in carats. Since size classes are not equal, grade is expressed per unit class interval and size is taken as the average stone size in the class.

The resulting grade-size plot is shown in Figure 14-17 and is compared with the grade-size plots for Group 1 microdiamonds and the fine sized combined bulk sample.

Figure 14-17 Grade-size Model Based on Group 1 Microdiamonds

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The microdiamond points lie on top of the model as expected, as the model is based on the average stone density for the combined Group 1 microdiamond sample. The bulk sample points lie on top of the model, suggesting the bulk sample grade is similar to the grade of the combined Group 1 microdiamonds. However, for the purpose of this exercise it does not matter if the bulk sample does not coincide with the microdiamond model, as the bulk- and microdiamond samples may not represent the same material.

Once a grade-size model is obtained, diamond grade is derived at given bottom cut-off for other samples with different average stone density, but with the same size distribution. Grade is derived from the grade-size distribution represented by the typical sample. With grade-size expressed in the DTC carat-grainer sieve breakdown it is possible to calculate the average Dollar per carat value for samples in the Group.

Exactly the same procedure was carried out for a mine production parcel recovered at +1.25mm bottom cut-off, but which is broken down into DTC size classes. The mine parcel contents in size classes below +7ds were adjusted to obtain total diamond recovery in the mine parcel in accordance with the typical sample. The factors that were required to achieve this are the alignment factors required to reflect +1.25mm recovery in the typical sample. These factors were then applied to the typical sample to reflect grade and average diamond value at a bottom cut-off of 1.25mm with a recovery profile similar to treatment at the mine.

Grade and factors are tabulated in the results section. The same procedure was carried out for each of the other two groups.

14.5.3.2 Group 2 Size and Grade-Size Modelling

The procedure was repeated for Group 2 microdiamond samples with their coarser size distribution. This combined sample comprises 587 stones recovered from 0. 728t of material.

For this procedure the combined bulk sample was used in all the comparisons. The total bulk sample was adjusted to contain the fine and coarse subsample in the same proportions to achieve a ‘mixed’ sample for comparison. The contribution from microdiamond sampling is substantial and modelling is based only on the microdiamond results, while the adjusted bulk sample is used for comparison only.

Figure 14-18 shows the LP-graph for the typical parcel generated in this case compared with the microdiamond graph at +0.075mm and with the bulk sample at +1carat.

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Figure 14-18 LP-Graph for Group 2 Microdiamonds with Bulk Sample

The model is coherent with the microdiamonds but slightly less so with the bulk sample.

The microdiamond points relate to the stone counts in mm size classes and the bulk sample relates to the size breakdown used at the Mine. The deviation of the model from the bulk sample is obvious, but not alarming in view of the size of the bulk sample and the deeper locations of the microdiamond samples. As mentioned, the bulk sample is shown for comparison and microdiamonds are used for actual modelling.

The final step in the modelling procedure was to generate a typical sample based on the microdiamond size distribution model and the average stone density for the combined Group 2 microdiamond sample. The product of this simulation is a grade-size breakdown in DTC size classes that is used to determine grade at given bottom cut-off and is suitable for diamond revenue calculation.

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The typical sample is based on a combined average stone density and the Group 2 size distribution model. The grade-size plot is shown in Figure 14- 19. The graph is accompanied by the Group 2 microdiamond points and the adjusted bulk sample above +1carat.

Figure 14-19 Grade-size Model Based on Group 2 Microdiamonds

The model shown is based on a Lognormal size model and an assumed stone density and, as before, the bulk sample grade-size points are shown only for comparison, not for a check on grade coherence between microdiamond and bulk sampling material.

For this comparison the microdiamonds above 0.85mm were allocated to DTC size classes. The model was derived from the microdiamond class frequencies as sieved into mm size classes at the time of their recovery and in both size class breakdowns there is highly satisfactory correspondence with the model as shown in Figure 14-19.

In addition, the LP-graphs in Figure 14-18 also show highly satisfactory correspondence with the size model. Therefore, the size and grade-size

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models for Group 2 microdiamonds were regarded as suitable to be used for the estimation of grade and revenue for the relevant sample combinations in the group.

14.5.3.3 Group 3 Size and Grade-Size Modelling

The samples in this group comprise recovery of 2,331 stones from 1.035t of material and the LP-graph reflecting diamond size distribution is shown in Figure 14-20.

The deviation of sample from model is peculiar. For a sample of this size one normally has little doubt about the nature of the size distribution, as there is little or no deviation of the LP-graph from linearity. In this case there are deviations but the fitted model is regarded to be the best fit possible for this group, taking into account the entire microdiamond parcel size range.

Retrospectively the grade-size plot in Figure 14-21 was also considered to decide on the size model to be used.

Figure 14-20 LP-Graphs for Group 3 Diamond Recovery

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The size distribution model was used to simulate a typical sample, assuming an average stone density for the Group. The typical sample was again simulated with diamonds reporting in DTC carat-grainer size classes for comparison with bulk sampling results and to facilitate revenue calculation.

The grade-size model for the Group was plotted with the graphs for microdiamonds and for the +1ct fraction of the coarse bulk sample, as shown in Figure 14-21.

The grade-size curve for the typical sample and microdiamond sample are coherent. Some of the microdiamond sample points deviate from the model (and typical sample), but this is understandable in view of the deviation of their deviation from the size model in Figure 14-20.

Figure 14-21 Grade-size Model Based on Group 3 Microdiamonds

The coarse bulk sample points above +1ct stone size are scattered around the model, indicating that the grade suggested by the model based on microdiamond sampling is ‘in touch’ with grade as observed in the bulk sample.

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The +1.25mm mine production parcel was used to determine alignment factors in the same way as described for Group 1. The grade-size points of the production parcel were plotted with the typical parcel and grade in production size classes below +7ds were adjusted to obtain total diamond recovery as indicated by the typical sample. The factors that were required to achieve this are the alignment factors required to reflect +1.25mm recovery in the typical sample. These factors were applied to the typical sample to reflect grade and average diamond value at a bottom cut-off of 1.25mm with a recovery profile similar to treatment at the mine.

This is illustrated in Figure 14-22.

Figure 14-22 Grade-size Graphs for Typical Parcel and Bulk Sample +1.25mm Results

The model represents total diamond content in the form of the distribution of grade with diamond size over a size range that includes the microdiamonds and macro diamonds from mine production at +1.25mm, adjusted to reflect total diamond recovery.

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The red markers show microdiamond class grade, the open circles show typical sample and the blue squares represent mine production. The latter is without losses in the bottom size classes, as these classes have been adjusted to total content to obtain the alignment factors required to achieve this. The typical sample is coherent with microdiamond sampling and mine production and is suitable for calculation of grade and revenue.

Similar analyses were carried out for Group 2&3 combined and for the total combined microdiamond sample. Results are summarised in Section 14.7.

14.6 Diamond Value Estimation

14.6.1 Approach and Data

Average diamond value per size class was provided by the Mine (Natural Diamonds, January 8th, 2016), based on the January 2016 Price Book. No modelling was required and class values are shown in Table 14-9.

Table 14-9 Bulk Sample Valuation Summary Jan., 2016

Size Class USD/carat 10+ CT 1,846 9 CT 1,606 8 CT 1,397 7 CT 1,214 6 CT 981 5 CT 854 4 CT 511 3 CT 427 10 GRN 371 8 GRN 335 6 GRN 245 5 GRN 192 4 GRN 167 3 GRN 86 2 GRN 59 -11+9 43 -9+7 37 -7+6 32 -6+5 26 -5+4 17 -4+3 12 -3+2 10 -2+1 9

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The parcel prior to sorting is photographed below in Figure 14-20.

Figure 14-23 Bulk Sample Diamond Parcel Prior to Sorting

14.7 Diamond Grades and Values

The diamond size distribution models were used to obtain the carat contribution per DTC size class corresponding to the valuation data as shown in Table 14-9.

The listing in Table 14-10 shows percentage carats, average diamond value per size class and class value for a 100carat parcel, with resulting overall average dollar per carat value for each group.

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Table 14-10 Average Diamond Value and Carat Contribution per 100 Carat Parcel

Average GROUP 1 GROUP 2 GROUP 3 Size class Price $/carat %CTS Dollar %CTS Dollar %CTS Dollar 150ct 1,846 0.017 32 0.056 104 0.045 82 100ct 1,846 0.040 74 0.118 219 0.096 176 60ct 1,846 0.094 173 0.247 456 0.204 377 45ct 1,846 0.089 164 0.211 390 0.178 330 30ct 1,846 0.194 358 0.422 779 0.362 668 20ct 1,846 0.316 584 0.621 1,147 0.544 1,004 15ct 1,846 0.335 618 0.605 1,116 0.537 991 10ct 1,846 0.694 1,281 1.153 2,128 1.038 1,916 9ct 1,606 0.237 381 0.371 596 0.337 542 8ct 1,397 0.298 416 0.454 634 0.415 579 7ct 1,214 0.385 467 0.570 692 0.522 634 6ct 981 0.514 504 0.737 723 0.679 666 5ct 854 0.720 614 0.993 847 0.920 785 4ct 511 1.142 584 1.503 768 1.403 717 3ct 427 1.588 678 1.977 844 1.859 793 10gr 371 1.349 501 1.601 594 1.515 563 8gr 335 2.223 745 2.518 844 2.397 804 6gr 245 2.600 636 2.797 684 2.678 655 5gr 192 2.767 531 2.840 545 2.735 525 4gr 167 5.425 907 5.265 880 5.099 853 3gr 86 4.905 423 4.490 387 4.373 377 +11 59 19.526 1,151 16.277 960 15.975 942 +9 43 21.963 947 16.266 701 16.084 694 +7 37 14.539 537 14.113 522 14.004 518 +6 32 6.779 215 10.044 318 9.981 316 +5 26 6.648 176 9.838 260 9.777 258 +4 17 3.530 59 2.964 50 4.905 82 +3 12 0.665 8 0.532 7 0.703 9 +2 10 0.372 4 0.374 4 0.554 5 +1 9.68 0.046 0 0.042 0 0.082 1 100.000 13,769 100.000 18,200 100.000 16,862 Average 137 182 169 Dollar/carat value: Recoverable grades are at +1.25mm bottom cut-off and the alignment factors from total diamond content to recoverable content are listed in Table 14-11 below.

Table 14-11 Alignment Factors from Total Recovery to +1.25mm Recovery +1.25mm Alignment factors DTC size class GROUP 1 GROUP 2 GROUP 3 +9 1 1 1 +7 0.7 1 1 +6 0.3 0.7 0.7 +5 0.2 0.5 0.5 +4 0.2 0.3 0.5 +3 0.02 0.03 0.04 +2 0.01 0.02 0.03 +1 0.0005 0.001 0.002

A summary of grade and value per individual combined sample in the groups is given in Table 14-12.

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Table 14-12 Average Grade and Dollar per carat Value per Lithology and Mine Level

1,2,3 individually 1, 2&3 combined 1,2,3 all combined Sample Stones/ Size USD/ Lithology Weight 100kg USD/ Grade USD/ Group Grade Grade carat (kg) +0.212mm carat +1.25 carat +1.25 mm +1.25 mm +1.25 +1.25 mm mm +1.25 mm mm

1 K2/K4 Below 350 1012 108 22 138 22 138 57 164 1 K6 North Below 350 472 32 7 138 7 138 17 164 1 CRB North 37 60 7 138 7 138 32 164 3/1 K2/K4 Above 350m 660 105 17 138 17 138 82 164

2 K8 127 31 19 182 22 159 16 164 2 K6 S above 350 395 19 13 182 15 159 10 164 2 K6 N above 350 207 19 19 182 22 159 10 164

3 K2/K4 Above 350 660 105 75 168 72 159 56 164 3 K6 S Below 350 374 37 28 168 27 159 19 164 1/3 CRB North 37 60 31 168 29 159 40 164

Grade is given for diamond recovery at +1.25mm as observed in the mine bulk sample parcel used in the analysis.

Diamond values are based on value per size class as provided by the Mine and listed previously. Recoverable grades are at +1.25mm bottom cut-off and the alignment factors from total diamond content to recoverable content are listed in Table 14-11.

14.8 Discussion on Grade Modelling Results

The following items are noted with respect to the values in Table 14-12:

 The values given in the combined columns are the most reasonable based on available microdiamond data. Note, bulk sample data was used comparatively and did not contribute to the values in the table.  The wide spread of LP-graphs for the combined samples shown is the reason for using three groups for grade and value estimation. A smaller breakdown of combinations did not result in materially different values while some smaller groupings do not contain sufficient stones for meaningful analysis.  CRB North is included with Group 1 and Group 3. If its diamond size distribution is fine, then Group 1 values will be valid, and if it is coarse, then Group 3 results will be more reasonable. The two sets of results for CRB North may therefore be regarded as estimates for lower and upper values.  Combining Groups 2 and 3 provides values that are little different from the previous option.

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 Combining all three groups has little effect on Group 2. This is understandable in view of Group2 having size distribution very similar to the total sample. This combination results in similar grade estimates for K4 above and below 350m level.  Grade estimates for K4 above 350m level are consistently higher than seen in bulk sampling. This is due to the coarseness of the size distribution suggested by microdiamonds. If a fine size distribution model is used for K4 above 350m level, then the values for K2/K4 at the bottom of Group 1 in Table 14-12 are obtained. This illustrates the impact of the diamond size distribution model on the estimates for the unit.  Average Dollar per carat value for samples in Group 3 is slightly lower than the average for Group 2. The LP-graphs suggest a coarser size distribution for Group 3 and consequently higher average diamond value in Dollar per carat. The reason for the ‘discrepancy’ is that the model that was eventually accepted for Group 3 reflects a size distribution with higher mean (average diamond size), but lower variance, resulting in a coarser size distribution but with smaller probability for large stones. This is in accordance with the sample size distribution. In this case the end result is a slightly lower average value. The Lace kimberlite contains a large proportion of microdiamonds, which makes it ideal for the application of microdiamond sampling for diamond content estimation.

As a consequence of the large proportion of small stones the balance between average grade and average diamond value must be examined to ensure that the correct bottom cut-off screen size is used. Grade quickly decreases with an increase in bottom cut-off, while average diamond value may increase more rapidly. Diamond revenue in terms of dollar per tonne must however be observed in order to appreciate the effects of the rates of change in the two entities.

This is illustrated in Figure 14-24, which uses combined K4 above 350m level sample from Group 3 with its somewhat problematic size distribution to show the effect of an increase in bottom cut-off.

Rapid changes are shown for grade and average diamond value, while the change in dollar per tonne revenue is relatively mild. The use of a high bottom cut-off size is suggested and high fluctuations in grade may be expected during production and sampling if small stone recovery is not tightly controlled.

For the illustration the Group 1 and Group 3 size distribution models are used and combined with sample stone density for the combined K4 sample above 350m level.

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Figure 14-24 Sensitivity Graphs of Diamond Grade and Revenue

The graphs on the right shows high grade and high average value, both associated with the coarse model. The graph on the left shows considerably lower grade and average value associated with the finer size distribution. Both graphs illustrate sharp changes in grade and value with mild change in Dollar per tonne revenue.

14.9 Diamond Grade Model

The robust microdiamond database and size frequency distribution data for the UK4 block allows for the usage of the following conservative, recoverable grades for K6 and K4 facies, to best represent diamond grades including inherent matrix dilution as discussed in Section 14.6.

UK4 Mine Block 230 to 370m

K6S – 10cpht Lowest value in the range of 10-15cpht indicated by grade models, supported by bulk sampling results which have recovered diluted grades of 4.4 to 7.5cpht in small samples K6N – 10cpht Lowest value in the range of 10-22cpht indicated by the grade models (no “pure” K6N bulk samples) K4 – 40cpht Low median value in the range of 17-82cpht indicated by the grade models, supported by bulk sampling results which have recovered diluted grades of 26.7 to 36.4cpht in small samples K8 – 16cpht Has not received bulk sampling to date, and therefore is assigned a grade of 16 cpht based on the microdiamond grade model, which included all internal dilution. CRB – 0cpht Has not received either bulk sampling or sufficient mida sampling to date, and is therefore assigned a grade of zero.

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The calculated block grades for the UK4 block are summarized below in Figures 14-25a and 14-25b, with more detail including validation provided in Appendix 2. The block model successfully predicted bulk sample recovered grades within 10% for both K6 and K4 facies.

The previous sections are suggesting that grades will improve in all facies, as the mining stopes move deeper into the block caves. This is seen geologically as the facies move away from the intermixing of K6 with K4 around and above the 290m level, in the microdiamond data, as well as in the dilution data reported in Section 14.4.

At this time we advocate using the same grades as above for these resources, which are felt to be reasonable, conservative grade estimates for diluted grades until such time as more data has been collected.

Figure 14-25a 3-D Perspective Lace Main Pipe K4 Grade Model Facing Northeast

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Figure 14-25b 3-D Perspective Lace Main Pipe K6 Grade Model Facing Northeast

14.10 Mineral Resource Statement

Detailed work on the UK4 depth slice of the Lace deposit has provided high confidence data for conservative resource estimation. Below -370m data is much sparser, and a prudent approach to grade estimation at this time utilizes the UK4 grade values, with the acknowledgment that the recommended continued delineation, microdiamond and bulk sampling work will allow for higher confidence updated estimates periodically. As well, continued mapping of all development drives will allow for refining of the block model, with more precision in terms of percentages of facies within individual blocks.

At this time we advocate uniformity of grade estimates for all three block cave depth slices until such time as larger databases are available to accurately discern variability.

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Undoubtedly volumes and tonnages will also change as more delineation work is completed.

When comparing to the previous mineral resource estimates prepared for the Lace mine, it is important to note that the estimated recovered grade for all facies has now been aligned with the Lace production plant utilizing 1.25mm slotted screens, compared to 1.00mm screens in its original design configuration. While the increase in bottom screen sizes means that recoveries will be lower in terms of grade, the carat value will increase as the smallest diamonds no longer being recovered are also the lowest value diamonds. So while the impact on recovered grade resulting from the change can be perceived as significant, the economic impact is minimal. The updated resource statement should be considered a conservative base case from which there is compelling evidence that considerable grade and value per carat upside is likely, but still to be further refined with additional production and evaluation data.

Table 14-13 Mineral Resource Statement

Lace Mine Totals Indicated 2,900.443 2.57 7,457,276 19.4% 0.191 1,427,642 $ 164.00  Based -230m on to a -920m recoverable Levels gradeInferred model for the current 11,965.54 Lace plant 2.59configuration 31,031,451 (+1.25mm80.6% bottom 0.256 cut-off 7,958,475screen size)$ 164.00 Totals 14,865.979 2.59 38,488,727 100.0% 0.244 9,386,117 $ 164.00  Diamond(1110m-410m price based absl) on bulk sample parcels and January 2016 price book  Effective date February 1st, 2016.

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15.0 MINERAL RESERVE ESTIMATE

Grade control will be on-going, as blocks that include both K6 and K4 material will be mined at both the north and south ends of the open stopes, and some or all of the K6 material may be stockpiled on surface. The interior blocks will be entirely of K4 facies and will be processed in their entirety. The UK4 Mine is targeting approximately 2million tonnes of the ~7.5million tonne indicated resource at this time. Reserve blocks and total block grades are shown schematically in Figure 15-1 below. The exterior blocks which contain a higher proportion of K6 (or CRB, or waste in one case) are obviously those with the lower block grades.

Figure 15-1 3-D Perspective Lace Main Pipe K4 Grade Model Facing Northeast

The reserve statement for the UK4 mining operation is provided below in Table 15-1.

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Table 15-1 Mineral Reserve Statement

There will be opportunities to optimize the mine plan as the mine progresses, focusing on additional K4 within the reserve, as per the recommendations of continual mapping and resource management as the UK4 Lift 1 is developed.

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16.0 MINING METHOD

Block caves are the best choice for mining massive ore deposits such as Lace in term of safety, percentage extraction, economics and efficiencies, but due to the capital and time required before any revenue can be generated, alternative revenue from a reserve which is readily accessible can bring forward cash flow. This can be done for example by:

 Mining a block of ground at the top of the block  Open sub level stoping ahead of the cave  High undercuts producing ore ahead of caving

DiamondCorp’s mining team and consultants spent a great deal of time in 2014 investigating mining method alternatives for the UK4 mining block at Lace, to bring forward cash flow by exploiting it ahead of Block Cave 1. The team, which includes underground kimberlite mining specialists Bob Harverson and Pat Bartlett, prepared a scoping study (DiamondCorp PLC, 2014), updated here-in.

The report concluded from the work done that a continuous trough type doming mining method incorporating a sub level open stope mined from the bottom, best suits mining of the UK4 block. The greatest risk of mining any block close to old workings is the potential risk of inundation of water and mud from the old workings. The chosen mining method – called Longhole Open Stoping Bottom Up (LOSBU) - will meet the requirements of the SIMRAC report on the guidelines to mitigate potential risk of mud rushes.

16.1 Geotechnical and Hydrological Aspects

Critical aspects to development include country rock transition from Ventersdorp lavas to shales at depths of ~300m, requiring support, as well as inundation of water and mud from historical workings. Also, drilling has indicated extensive brecciated zones along the margins of the pipe.

16.1.1 Field Strength Estimates

Field strength estimates are based on underground mapping during development of the main decline and bulk sampling on 250m and 260m levels. The Ventersdorp lavas higher up in the main decline typically fall in the very strong to extremely strong category (ISRM, 1978) and a Uniaxial Compressive Strength (UCS) of 250 MPa has typically been used for estimating required levels of rock support in the main decline. The kimberlites encountered in bulk sampling excavations on 250m and 260m levels, based on field estimates, are generally strong and a UCS of between 50 to 100 MPa was used for estimating required levels of support in bulk sampling tunnels and excavations on these levels.

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16.1.2 Laboratory Rock Test Results

A relatively short cored borehole (B-size) was drilled from the bulk sampling excavations on 250m level, with samples of this kimberlite core submitted to Rocklab in Pretoria for rock strength testing. A summary of the results is presented in 16-1.

Table 16-1 Summary of Rock Strength Tests on Kimberlite Core Samples

Strength Property Test Results UCS average 1 118.5 MPa range 1 22.31 1 to 166.1 MPa std. deviation 48.7 MPa no. of samples 7 Brazilian tensile strength average 10.9 MPa range 8.0 to 14.7 MPa std. deviation 2.6 MPa no. of samples 5 Corrected point load strength, Is(50) average 3.0 MPa range 0.4 to 6.0 MPa std. deviation 1.8 MPa no. of samples 13 Density average 2.65 t/m3 range 2.52 to 2.76 t/m3 std. deviation 0.1 t/m3 no. of samples 12 Notes: 1-The very low UCS value of 22.3 MPa is an outlier value which is a quarter or less of the other UCS results. If excluded from the data set, an average UCS of 134.5 MPa is obtained.

UCS values obtained from the laboratory test programme as summarised in 16-1 are significantly higher than initial field estimates presented in Section 16.1.1. These higher UCS values are compared with that of hypabyssal kimberlites encountered at other kimberlite pipe mines in South African in Table 16-2 and with that of TKB encountered at the nearby Voorspoed mine.

The strength of the Lace kimberlites is comparable with that of the competent hypabyssal kimberlites encountered at De Beers Mine in Kimberley and appears much more competent than the hypabyssal core found on 870m level at Du Toitspan mine in Kimberley. It is also significantly more competent than the TKBs encountered at the neighbouring Voorspoed mine.

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Table 16-2 Typical Strength Values of Hypabyssal Kimberlite

Name of Mine Type of Kimberlite UCS (MPa) De Beers, 720 m E. Hypabyssal core average: 115 range: 85 to 141 De Beers, 720 m E. Periphery, hypabyssal kimberlite average: 167 range: 163 to 171 Du Toitspan, 870 m Core, hypabyssal kimberlite average: - range: 60 to 80 Voorspoed, BH 30 m Tuffisitic kimberlite breccia average: 90 range: 60 to 118

16.1.3 Jointing

16.1.3.1 Jointing in Ventersdorp Lavas and Shales

Three to four joint sets with some random joints were typically noted in the Ventersdorp lavas in the main decline down to 250m level. Joint spacings varied from as little as 0.1m in places to up to 3m elsewhere. Estimates of RQD varied from 75 to 90% in more blocky ground up to 90 to 100% in ground with more widely spaced joints.

The two main joint sets encountered in the main decline are generally sub- vertical and through-going. Joints from the other joint sets are generally less persistent and flatter dipping. Joints are typically smooth, planar to undulating with striations evident on some of the joint surfaces indicating earlier shear deformation in the rock mass. Joints below the weathered zone are generally tightly closed with unweathered walls and no infill. A few areas of brecciation were also encountered in the main decline in the Ventersdorp lavas as well as kimberlite stringers / fissures near the pipe rim at the 250m level.

16.1.3.2 Jointing in Kimberlite

Joints or discontinuities in the kimberlite generally comprise some random jointing as well as internal contacts between different kimberlite facies, with these contact structures / discontinuities dipping steeply in places and shallowly in other areas. The kimberlite rock mass largely appears competent and stable, with overbreak largely restricted to tunnel intersections with old workings. The only rock fall of any significance during development of the bulk sampling tunnels on 250m and 260m level, occurred in the immediate face area of one of the ends where a rock beam in the roof of the tunnel, formed by a flat dipping internal contact, was not adequately supported.

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On the 290m level at the junction of the doming drive with the crosscut, ground stability issues developed and the roof caved in several metres. Whilst some minor ground stability issues were expected in these new development tunnels due to proximal historical tunnels, and were easily mitigated as encountered, the larger amount of caving at that junction was unexpected and could not have been foreseen from all drilling, and historical records in-hand.

The caving appears to have been caused by a shallow southerly dipping slip fault in the kimberlite, which let the material beneath it collapse as the void below became large enough. This fault structure continues beneath the 290m level at a shallow dip, but has not caused any issues on in the 310m development drive. Further, we expect that the impact of this fault structure will diminish with depth and should exit the kimberlite completely by the 350m level, more than 100m above the planned block cave level.

Nonetheless, in order to mitigate potential stresses at development cross cuts, Lace’s mining consultants have modified the tunnel layout on the 310m level to eliminate cross cuts entirely. Instead, the layout now incorporates staggered breakaways left and right without any full cross cuts. This modified layout has no impact on proposed mining rates and significantly reduces the potential for a repeat of the issue encountered on the 290m level.

16.1.3.3 Rock Competency from Core Measurements

Core Recovery (“SCR”) is measured systematically in all holes, as a percentage of the total (Whole) Core Recovery per drilling interval. Rock Quality Determination (“RQD”) is measured as a percentage of all sections of core greater than 10cm in length per drilling interval. All of these data are kept in the core drilling database, and summarized below.

KIMBERLITE: K4 is very stable with a high recovery percentage and rock quality. K8 is similar, again suggesting a similarity to the “Hardebank” noted in the historical mining plans. In contrast to K4, K6 generally shows good recovery although the rock quality is variable and unimodal reflecting the variable proportion of country rock xenoliths.

LAVA: The Country Rock Lava is extremely competent as demonstrated by a very high recovery and rock quality near 100%.

LAVA BRECCIA: Sections of the Country Rock Lavas showing original brecciation (i.e. to do with original lava flows / pyroclastic intervals etc.) has a similarly high recovery to the lava (close to 100%). However the RQD appears more spread out than the SCR and bimodal with a second peak

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around 30%. This is where intervals of the Lava Breccia have most probably been affected by the kimberlite intrusion (see similarity to the bimodality in the CRB below).

COUNTRY ROCK BRECCIA: The country rock breccia is an interval of country rock which has been brecciated by the forces associated with the kimberlite intrusion and impregnated by minimal amounts of kimberlite usually below 5%. Both the TCR and RQD show a large range with a bimodality due to a second peak around 30 to 40%.

SHALE: The shale wall rock is generally competent as shown by the high recovery percentage, although the RQD is highly variable.

SHALE BRECCIA: The shale breccia appears to be bimodal in terms of SCR although the RQD is highly variable and generally very poor. Intervals of shale breccia (which includes some of the wall rocks at depth) can thus be expected to be more challenging in terms of rock mechanics than the overlying lavas.

Table 16-3 Summary Statistics and Histograms for Core Competency Measurements

CRB-SCR K2-SCR K4-SCR K6-SCR K8-SCR L-SCR LB-SCR LRKB-SCR S-SCR SB-SCR Mean 62.4 84.8 92.7 89.5 92.3 93.1 89.6 84.7 87.5 82.3 Standard Error 3.9 4.4 1.9 1.0 2.8 1.5 2.3 10.1 3.7 9.7 Median 70.7 96.7 98.3 97.3 98.6 98.0 94.6 99.4 91.7 93.3 Mode 75.0 103.3 100.0 100.0 100.0 100.0 100.0 99.4 83.3 #N/A Standard Deviation 32.1 23.1 13.0 22.8 17.2 24.8 15.8 30.2 16.7 37.5 Sample Variance 1030.5 532.1 170.1 521.3 297.4 615.3 249.0 909.2 280.4 1408.1 Kurtosis -1.2 -0.4 10.8 14.8 3.5 64.9 2.7 0.7 0.5 1.6 Skewness -0.4 -1.2 -3.0 0.1 -2.1 4.8 -1.8 -1.6 -0.9 0.7 Range 108.3 64.7 70.8 295.0 68.7 378.4 72.7 69.8 65.4 148.4 Minimum 0.0 38.7 30.4 0.0 42.7 0.0 37.3 31.4 52.0 27.7 Maximum 108.3 103.3 101.2 295.0 111.3 378.4 110.0 101.2 117.4 176.1 Sum 4304.5 2288.8 4448.8 47994.3 3508.2 25589.3 4120.1 762.0 1836.6 1234.3 Count 69 27 48 536 38 275 46 9 21 15 CRB-RQD CRB-RQD K2-RQD K4-RQD K6-RQD K8-RQD L-RQD LB-RQD LRKB-RQD S-RQD SB-RQD Mean 40.5 70.7 80.2 68.6 88.2 81.4 57.4 91.8 64.2 48.6 Standard Error 2.9 5.7 3.0 1.0 2.0 1.3 4.5 1.7 4.4 8.2 Median 43.3 79.2 88.1 72.5 92.3 90.9 64.9 91.8 67.6 43.0 Mode 0.0 0.0 #N/A 100.0 100.0 100.0 0.0 #N/A #N/A #N/A Standard Deviation 24.0 29.8 20.7 23.2 12.3 21.5 30.6 5.2 20.2 31.9 Sample Variance 578.3 887.3 429.2 540.5 151.4 462.9 935.4 27.5 407.6 1020.6 Kurtosis -1.0 1.6 2.3 0.5 5.8 1.7 -1.1 -1.0 -1.2 -1.2 Skewness -0.1 -1.5 -1.6 -0.9 -2.1 -1.6 -0.5 0.0 -0.3 0.3 Range 86.3 99.0 86.5 100.0 60.9 100.0 100.0 14.3 65.5 100.0 Minimum 0.0 0.0 13.5 0.0 39.1 0.0 0.0 84.7 25.0 0.0 Maximum 86.3 99.0 100.0 100.0 100.0 100.0 100.0 99.0 90.5 100.0 Sum 2796.4 1910.2 3849.1 36784.7 3351.0 22394.2 2639.4 825.9 1347.7 729.4 Count 69 27 48 536 38 275 46 9 21 15

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K4 (SCR) RQD (K4) 40 30 20 20 10 0 0 Frequency Frequency

Bin Bin

K6 (SCR) K6 RQD 300 150 200 100 100 50 0 0 Frequency Frequency

Bin Bin

K2 (SCR) RQD (K2) 15 10 10 5 5 0 0 Frequency Frequency

Bin Bin

K8 (SCR) K8 RQD 30 30 20 20 10 10 0 0 Frequency Frequency

Bin Bin

LAVA (SCR) Lava RQD 200 200 100 100 0 0 Frequency Frequency

Bin Bin

Lava Breccia (SCR) Lava Breccia RQD 30 10 20 10 5 0 0 Frequency Frequency

Bin Bin

Shale (SCR) Shale (RQD) 15 6 10 4 5 2 0 0 Frequency Frequency

Bin Bin

Shale Breccia (SCR) Shale Breccia RQD 6 4 4 2 2 0 0 Frequency Frequency

Bin Bin

CRB (SCR) RQD (Country Rock Breccia) 20 15 10 10 5 0 0 Frequency Frequency

Bin Bin

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16.1.4 In Situ Ground Stress

No in situ stress measurement data is available for the immediate Lace project area. However, stress measurement data (Stacey et al, 1998) from the Free State goldfields in the Welkom area south of Lace and from the Klerksdorp area in the North West to the north of Lace, indicate a maximum horizontal stress to vertical overburden stress ratio, or K-ratio, of 0.92 for the region on average with ratio values varying from 0.59 to 1.37.

A K-ratio of 1.0 has therefore been used for estimating support requirements during development of the main decline in the Ventersdorp lava country rock. The vertical overburden pressure in kimberlite at 260m level has been estimated at 3 MPa for the purpose of support design in bulk sampling tunnels in the kimberlite (GeoStable SA, 2011), also taking cognisance of mined-out ground as well as pit sidewall collapse material in the pit bottom. The effect of raised foundation stresses below and around the footprint of the unmined block of low grade kimberlite (the so-called 'Hardebank') above 250m level was accounted for by adopting a K-ratio of 1.25 for bulk sampling support design purposes.

Stability of Old Workings in Kimberlite Underground workings visible from surface in the pit sidewalls as well as where encountered underground during bulk sampling, generally remained open without any support, albeit some spalling of kimberlite material in the roof and sidewalls did occur. The historical main drifts generally had a span of 3m while other working drifts had a smaller span of 2m.

Bieniawski (1989) presented a stand-up time chart for unsupported tunnels based on Rock Mass Rating (RMR) as presented below. Tunnels with unsupported spans of 2 to 3 m are plotted on this chart and as can be seen, some level of support would have been required to prevent the rock falls and spalling which did occur over the many tens of years that the mine has been closed. However, the old tunnel workings generally remained open as stated earlier and an RMR of the order of 70 to 80 and

Figure 1 : Stand-up Time Chart (Bieniawski, 1989)

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more (i.e. good to very good rock mass conditions) can therefore generally be assumed for the kimberlites above 250m level.

16.1.5 Rock Mass Classifications

16.1.5.1 Ventersdorp Rock Mass Conditions

Barton's NGI-Q rock mass classification values for the Ventersdorp lavas encountered in the 4.5 m wide main decline above 250m level typically varied from 10 to 20 (good tunnelling conditions) requiring only nominal support to ensure safe working conditions in decline development ends. Blocky ground (poor to fair tunnelling conditions) was encountered in a few places which required installation of additional support.

In tunnel junctions / at breakaways in Ventersdorp lavas, Q ratings typically reduced to 1 to 7 (poor to fair) due to the increased span, which necessitated installation of longer support. In areas of blocky ground, support spacing was also reduced.

16.1.5.2 Kimberlite Rock Mass Conditions

Laubscher's Mining Rock Mass Ratings (MRMR) (1990) for the Lace kimberlites, based on the above discussion of available data, are listed in Table 16-3. This table shows that the kimberlite rock mass is quite competent with basic RMR values of 75 to 85 (good to very good). However, when adjusted to take cognisance of weathering; adverse joint orientations; mining induced stresses; and blast damage, MRMR values ranging from 50 to 56 (fair) are generally obtained.

16.1.6 Design Rock Mass Strength

The Rock Mass Strength (RMS) for the kimberlite can be determined as follows (Laubscher, 1990):

RMS = (RMR – RUCS)/80 x UCS x 80/100, where: RUCS is the RMR component rating for the UCS. Therefore, the average RMS = (80 – 12)/80 x 118.5 MPa x 80/100 = 80 MPa. The Design Rock Mass Strength (DRMS) is then given by:

DRMS = RMS x adjusted for weathering, joint orientation and blasting That is, DRMS = 80 MPa x 0.88 x 0.8 x 0.94 = 50 MPa.

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Table 16-4 Mining Rock Mass Ratings for the Lace Kimberlites

Classification Typical Typical Parameter Values Rating Parameter Description Worst Average Best UCS very strong average: 118.5 MPa 8 12 16 rock worst case: 50 to 100 MPa best case: 150 MPa RQD Blocky 75 to 90% 12 to 14 Joint Spacing one set (or joint) only, 1.5 to 5 m 25 Joint Condition moist ground, not continuous, rough 40 x 1.0 x 0.75 x 1.0 x 1.0 = 30 undulating, no separation/infill, unweathered Basic RMR: 75 to 77 79 to 81 83 to 85 Adjustments Weathering: Slightly weathered in 6 months 0.88 Joint orientation: 60 degree plunge on joint intersections 0.8 to 1.0 Mining induced stress: Trough drives fully developed:  Once trough drives are fully developed: UCS / 1 = 118.5 MPa / 10 MPa = 11.9 o Approximately 15% extraction of the block cave production footprint and  / UCS = 20 MPa / 118.5 MPa = 0.17 the major principal stress can be estimated as 1 = v / 0.85 = 8.5 MPa / Medium stress conditions, 0.85 = 10 MPa. favourable for stability o Use an initial stress multiplier of 2 for K-ratio = 1 to determine raised 1.0 circumferential stress  in tunnel perimeter during trough drive development (see stress distributions around excavations published by Stacey et al, 1980). Therefore,  = 2 x 1 = 2 x 10 MPa = 20 MPa. Continuous trough development / undercutting, along caving front:  During continuous trough development / undercutting: Use a stress multiplier UCS / 1 = 118.5 MPa / 10 MPa = 11.9 of 4.5 for a K-ratio = 1.0 to estimate the raised abutment stress (see Stacey et pillar core / UCS = 45 MPa / 118.5 MPa al, 1980). Distinguish between abutment stress in the pillar core and = 0.38 abutment stress at the trough drive tunnel perimeter:  @ abutment / UCS = 90 MPa / 118.5 MPa = 0.76 In pillar core At pillar core: : High stress with tight structure, usually favourable for stability pillar core = 4.5 x  = 4.5 x 10 MPa = 45 MPa. At tunnel perimeter:  In trough drives   @ abutment = 4.5 x  = 4.5 x 20 MPa = 90 MPa. : Stress induced slabbing / squeezing, approaching a mild squeezing pressure  During production drive and drawpoint development assuming 0.8 pre-undercutting: o Approximately 50% extraction of the block cave production footprint, i.e. Production drive and drawpoint development the major principal stress as well as when in production: 1 = v / 0.5 = 8.5 MPa / 0.5 = 17 MPa. UCS / 1 = 118.5 MPa / 17 MPa = 7.0 o Use stress multiplier of 2 for K-ratio = 1 to determine raised  / UCS = 34 MPa / 118.5 MPa = 0.29 circumferential stress  in tunnel perimeter. Therefore, High stress with tight structure, usually  = = 2 x 1 = 2 x 17 MPa = 34 MPa. favourable for stability but may be unfavourable for wall stability 1.0

Blasting: Assume good conventional drill and blast control in all 0.94 kimberlite development MRMR (trough drives fully developed / during production): 50 53 56

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Based on the above assessments, it was concluded (again with reference to Laubscher, 1990) that:

. Drilling conditions will generally be good. . Brow support typically will comprise a concrete lining.

16.2 Mining Method Option Study

The 2014 scoping study reviewed all mining methods applicable to exploiting the UK4 block, and is summarized here-in.

16.2.1 Mining Methods Considered

The following mining methods were considered for the mining of the 310m level UK4 Upper Block:

• Longhole Open Stoping (LOS) • Longhole Open Stoping Bottom Up (LOSBU) • Vertical Crater Retreat (VCR) • Sub-Level Caving (SLC) • Front Caving (FC): During detailed discussions, this mining method was found to be totally impractical for LACE and it was therefore not considered in any further detail. • Block Caving (BC) • Inclined Caving (IC) • Cut & Fill (C&F)

16.2.2 Mining Method Evaluations

The different mining methods considered for mining of the 310m level mining block were compared in the Scoping Study and tabulated in Table 16-4 using ratings from 1 (excellent) to 10 (bad / totally unsuitable) for a wide range of aspects including access, ground handling, ventilation, etc.

A rating equal to or above 8 indicates an unacceptable risk and has been shaded in red.

Of the eight mining methods considered for mining of the 310m level UK4 mining block, Block Caving (BC), Inclined Caving (IC), Sub level Caving (SLC), Front Caving (FC) could not be used due to the small dimensions of the block.

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Table 16-5 Mining Method Comparison for UK4 Mine

Mining Method Access Ground handling Ventilation Complexities Geometry FOG Air Blast RushMud Potential Geotechnical Economics Risk Evaluation % Extraction LH-Open Stoping 3 5 3 2 7 4 0 10 8 4 10 90 LH-OS-Bottom Up 3 5 3 2 3 4 0 4 8 4 5 90 Vertical Crater 3 5 3 6 3 4 0 10 9 4 8 70 Retreat Sub-Level Caving Block too small Front Caving Block too small Block Caving Block too small Inclined Caving Block too small Cut & Fill 2 6 4 4 2 3 0 8 3 10 10 90

16.2.3 Mining Method Selection

Longhole Open Stoping Bottom Up (LOSBU) was selected as the mining method for the UK4 block based on extensive analysis. Strengths and weaknesses are tabulated in more detail in Table 16-5 below.

Based on a high level risk assessment and taking into account the considerations for all the listed mining methods above, the best option for the mining of the 310m level – UK4 upper block would be Longhole Open Stoping Bottom Up (LOSBU). This mining method meets safety, operational, design and financial criteria. The greatest risk to mining this block on the 310m level is the potential of water and mud in the old workings on 260m level. This method distances the draw points away from the potential water and mud with 40 metres of broken ore between them and the old workings when they are intersected.

Figure 16-1 schematically portrays the mining layout.

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Table 16-6 LOSBU Mining Method Analysis

Mining Aspect Advantages Disadvantages Considered Access options Short access with the ability to - produce early from ore reserve. Ground handling Minimal crusher requirements. - High rate loading. Ventilation Simple ventilation system. - Lesser air quantities required. Complexities Lay-out and operation simple. All material to be blasted. Suitability of mining Unsuitable for mining under an old method to intended mined area which has the potential block to have mud Fall of Ground (FOG) Less development and risks exposure to FOG’s. Mud rush potential and - High likelihood and potential of risks mud rush occurrence due to mud accumulated in pit bottom and old working levels. Mud derived from pit sidewall collapses from shales at surface. Large scale collapse Not a risk at this depth Air blast potential and Not subject to air blasts. - risks Impact of historic - Old development can impact on the mining excavations, i.e. positioning of the development and the Lace open pit, old long hole drilling and blasting underground workings and bulk sample excavations

Financials, working Comparatively low working - costs, capital, etc. cost and capital.

Conclusions:  High risk from mud rushes due to the probability of mud and water in the overlaying old workings at LACE migrating down into working stopes in the proposed 310 Lvl mining block.

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Figure 16-1 3-D Schematic of Trough and Dome Layout for UK4 Mine

16.3 High Level Risks

16.3.1 Introduction

Only one main risk was identified in the mining method selected for UK4 Upper level block mining options study, namely the risk of mud rushes with mud from the old working levels and pit bottom migrating down into and inundating the 310m Level production level.

16.3.2 Mud Rush Risk Potential - Conditions Required for Mud Rushes to Occur

The four elements required for a mud rush at any operation are:

• Mud forming materials. • Water • Ground disturbance (or mud disturbance) • Discharge points on an underground working level

All four these elements will exist at Lace; the proposed Lace mine must therefore be considered a mud rush prone operation. The four elements may be described as follows:

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16.3.2.1 Mud Forming Material

The Lace kimberlite is very similar to the kimberlite which was found above the main gabbro sill at Cullinan Diamond Mine. It is a competent kimberlite which does not readily decompose to form mud and no significant mud rushes occurred in the workings at Cullinan Diamond Mine in this area.

However, the mudstone, shale and overburden in the first 60m from surface which have slumped into the Lace open pit over the years as well as some kimberlite fines which formed in the pit bottom and during mining of the old working levels above the 310m level UK4 Upper Block, have resulted in the accumulation of a significant amount of mud in the pit bottom, some of which have migrated through open holes into the old underground workings down to 260m level.

16.3.2.2 Water

There is underground water at Lace and the inflow into the underground workings has been estimated to be approximately 22m3/hour. This was ascertained by first pumping the open pit empty of water and then pumping from the vertical shaft for 5 years in order to maintain a steady groundwater level below the 250m and 260m levels where bulk sampling was carried out. Inflow of water into the underground workings is mainly as a result of surface stormwater runoff into the open pit and natural groundwater recharge from adjacent groundwater aquifers.

16.3.2.3 Ground Disturbance (or mud disturbance)

Production from the 310m level UK4 Upper Block will take place 50m below the old workings and the cave will eventually develop into the existing old workings and the current pit bottom. The planned draw for the 310m level LOSBU is to draw off the ground at a rate of 1 tonne for every 3 tonnes blasted until the stope breaks through into the old workings. The remainder of the blasted ore will then be drawn off.

16.3.2.4 Discharge Points on an Underground Working Level

The production level and draw points of the 310m level UK4 Upper Block will be located 50m below the old workings with draw points 4.5m wide and high and spaced 15m apart along the continuous cone. The draw will be controlled to draw off swell caused by blasting so as to keep the stope partially full of ore.

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16.3.3 Strategies for the Prevention of Mud Rushes

The following implementations are in place for the UK4 operations.

16.3.3.1 Surface Storm Water Control

Surface drains and storm water control trenches will be constructed around the open pit and box cut / tunnel portals to the underground workings to prevent storm water runoff from running into the open mine / underground mine. These storm water drains and trenches will be regularly checked and cleaned and / or repaired as may be required.

16.3.3.2 Mine Waste Disposal

The mine waste will be deposited in properly constructed slimes dams designed by a competent professional engineer. The slimes dams will be located far enough away from the open pit and underground mine, the twin access / conveyor decline and ventilation shafts so as not to pose a mud rush risk in case of a slimes dam failure. The caving operation will similarly be too far away from the slimes dams to cause any settlement of slimes dam embankments or to induce ground instability / settlement underneath the impounded slimes which may culminate in a mud rush into the underground workings.

16.3.3.3 Mine Design

In designing the underground mining operation, the following will be implemented and made use of:

• Blast damage in key mine infrastructure such as the production level tunnels, crosscuts and draw points will be kept low by exercising smooth blasting techniques during development of these excavations. • Groundwater will be drained off the block cave well ahead of production. The mine from 240m down to 340m is inundated with numerous old development tunnels, tips and shafts which will facilitate this mine dewatering process. Sumps and mine dewatering infrastructure will be designed and installed as required to ensure this requirement is met before production starts. 16.3.3.4 Draw Control

The UK4 Upper Block planned draw is to maintain a partially full stope by drawing off the swell caused by blasting to create an air gap between the ore pile and the face to be blasted. A draw control strategy will be implemented with main objective to maintain this draw over the stope. This will substantially reduce the likelihood of a mud rush occurring as:

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• The stope will be almost partially full of broken ore at all times prior to the break through to the old workings. • Any water or mud which then enters the stope will be drawn down under controlled conditions A Draw Control Officer will be appointed to chair a Draw Control Committee comprising the Draw Control Officer, the Mining Engineer, Production Mine Overseer, Geotechnical Engineer and Health and Safety Officer. This Draw Control Committee will determine draw control strategies and production schedules will be developed accordingly. Any deviations from the production schedules issued by the Draw Control Meeting must be approved in writing by the Committee. Actual draw will be compared with the monthly draw control schedules and any unauthorised deviations will be investigated and mitigation measures implemented to prevent its recurrence. No manager or production official will have the authority to override decisions made by the Draw Control Committee.

16.3.3.5 Mine De-watering

Lace is located in a relatively low rainfall area with an average rainfall of approximately 650mm per annum with most rainfall occurring during mid- summer. Direct precipitation within the limits of the open pit and storm water runoff into the open pit also rapidly reports to the underground workings (i.e. within half an hour from such rainfall occurring), resulting in an almost immediate increase in pumping requirements following a rainfall event.

Work by SRK (2005) concluded that there are two groundwater aquifer systems present in the Lace area, namely: a localized perched aquifer, and an unconfined or confined fractured and / or weathered rock aquifer, as follows:

 A perched aquifer of limited and variable extent occurs at shallow depth. This aquifer may develop in areas underlain by shallow, relatively impervious horizons of residual clayey soils derived from Karoo Supergroup shales. This perched aquifer is recharged from surface runoff but is unlikely to have a significant impact on the mining operations.  Unconfined or confined aquifers exist in the weathered and / or fracture zones within the shales, at the contact of the weathered zone and bedrock, and, in the Ventersdorp lavas at depth, in radial fissures typical of kimberlite intrusions. Water-bearing zones are likely to be associated with fracturing and / or weathering adjacent to structural features such as faults, joints, intrusions (particularly dolerite and kimberlite dykes), contact zones (between shale and Ventersdorp lava as well as between weathered overburden and bedrock) and zones of preferential weathering. These aquifers are generally low yielding and occur as discrete zones.

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Higher inflows would be expected along fractures and joint planes and particularly at the contacts of dykes.

No major groundwater strikes have been reported during development of the main decline down to 310m level or during bulk sampling operations completed within the kimberlite pipe on 250m, 290m and 310m levels. However, minor groundwater inflows were noted occasionally on intersections of more persistent geological discontinuities with the decline tunnel.

16.3.3.6 Mandatory Code of Practise

A Mandatory Code of Practice (COP) to Prevent Rockmass Failure Accidents in a Massive Mining Operation will be compiled in accordance with the relevant DMR Guidelines. This COP will be compiled by a COP Drafting Committee who will consult with LDM’s Health and Safety Committee regarding the preparation, implementation and revision of this COP. The COP will include detailed risk controls and mitigation strategies to prevent accidents as a result of mud rushes, focussing on the 3-D principle as follows:

• Distance: Keep the mud material away from the mining operation on 310m level, 50m below the old workings.

• Drain: Use the existing tunnels as well as new dewatering strategies to drain off groundwater from the 310m level Mining Block.

• Draw: Manage draw-down of ore reserve to prevent isolated draw which can result in the sudden discharge of mud pockets and layers.

 Mitigation strategies and mud rush risk procedures and systems will include:

• Signing of a mud rush book when entering areas where a mud rush risk exist.

• Installation of audible alarms and warning lights on the production level which can be triggered in case of a mud rush, to alert all working on the level to evacuate.

• Clear demarcation of escape routes.

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• Training of personnel working in areas where a mud rush risk exists, to allow early identification of the onset of a mud rush.

• Reporting structures and procedures.

• Mud push and mud rush incidents will be recorded in a mud rush register.

• Draw control strategies will be adjusted as and where appropriate in order to reduce the likelihood of mud rushes occurring.

16.4 Mine Design

16.4.1 Geotechnical Aspects

Data to date shows that the main pipe is made up of two competent kimberlite types (UCS >100 MPa) that do not easily decompose in the presence of water. The Ventersdorp lava in which access, infrastructure and auxiliary mining excavations around the pipe will be sited is furthermore a very competent rock type.

The seismicity of the Lace project area has been reviewed and it is unlikely that seismicity will significantly impact on the mining operations at Lace.

Support required in excavations in kimberlite on the various working levels to be constructed for the 310m level UK4 Upper Block, has focused on:

• Mining layouts, methods and sequencing. • Tunnel sizes and shapes. • Support considerations. • Drilling and blasting control.

Initial support recommendations were based on the amount of geotechnical information available which was obtained when developing the sample tunnels on 250m level. This will be reviewed as more geotechnical information becomes available, with the required level of support reduced or increased as may be appropriate, based on such final design calculations.

Regular precision survey pit rim monitoring, especially in the area of the current decline box cut and portal, to timeously establish the onset of any major pit perimeter instability which may affect access into the underground mine, also needs to be implemented.

Although the occurrence of a significant air blast at Lace is not considered likely based on current information it will be re-evaluated in detail as more geotechnical information becomes available and appropriate draw control strategies

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implemented to avert an air blast accident. The required mitigation measures and control strategies that will be implemented have been discussed in detail in Section 16.3.

16.4.2 310 Level UK4 Mine Upper Block Layout

The UK4 Upper Block is situated in the North-Eastern side of the main pipe as seen in Figure 14-2 – 310m Level Plan, and extends from 230m Level downwards.

16.4.2.1 Access and Ground handling

Existing Infrastructure

Development of mining tunnels and levels in the Lace kimberlite pipe was carried out down to 340m level historically, although no mining as such was conducted on the deeper mining levels. The tunnels on these old working levels have also remained largely open which facilitate drainage of water from the old mining levels.

A small rectangular shaft was put in down to 360m level during the 1920’s with crosscuts to the pipe on 160m, 240m and 330m levels. The shaft timber has collapsed and debris is hung up about 30m above the 160m level station with complete closure of the shaft between 160m and 330m. A bypass is in on 160m level to function as a temporary return air pass (“RAP”) during the initial mine development stages until a new 4.5m diameter RAP from surface has been constructed.

Part of the existing bulk sampling workings on 250m and 260m levels as well as an old travelling way on 250m level will be re-supported and made safe to allow installation of a dewatering pump and infrastructure for dewatering of the mining block down to just above 340m level during access development for the UK4 - Upper Block.

Planned Access and Groundhandling

The UK4-310m Level Upper Block will be accessed via the twin ramp system from surface accessed with a 12o ramp down to the 290m and 310m levels from Leg 5.

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A ramp from the man and materials ramp above leg 5 will be developed to above the conveyor ramp to create an ore pass with a capacity of 1000 tonnes. This ore pass will feed directly onto the conveyor and will have a suitable grizzly and rock breaker installed at its top.

16.4.2.2 General

The LDM UK4 Upper Lift 1 block design layout comprises two main levels, namely a doming level (290m level) and the main production level (310m level) as shown schematically in Figure 16-2.

Figure 16-2 3-D Schematic of Development Layout for UK4 Block

Two open stopes are to be positioned to extract the required ore tonnage from the block as shown in Figure 16-3.

Two continuous troughs will be situated for ore collection below the Open stopes. The doming level will be situated on 290m level to enable doming long hole rings to blast ore down into the collection troughs. This ore will be loaded from draw points spaced along the troughs. One stope will be

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mined initially and once this has broken through to the old workings and the mud and water is dissipated and drawn down the second stope will be mined. The mining sequence is presented schematically in Figure 16-4, following.

Figure 16-3 Section View of Trough and Dome Layout for UK4 Mine

16.4.2.3 Doming Level (290m Level)

From the doming tunnel doming rings will be drilled over the draw troughs from 290m Level to 250m Level as per Figure 16-3. The longest hole will be 48m and the dimensions of the full ring for both stopes will be 60m wide and 40m high.

The level will have a combined RAP and second outlet situated at its end to up cast air to the 250m level which will be connected to the RAP to surface. It will be equipped with a ladder to enable persons working in the block availability to a second outlet. The intake air will come down the 120 ramp from the twin ramps.

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Figure 16-4 Mining Progression for UK4 Mine Lift 1

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16.4.2.1 Production Level (310m Level)

This level will be the production level at the base of the draw troughs and will have LHD’s loading ore from draw points spaced along their length into trucks which will transport the ore up the 120 ramp to the ore pass over the conveyor ramp.

The level will basically have three tunnels. Two cone cutting tunnels with a production tunnel and draw points situated between them. The doming level will be positioned 20m above the production level tunnels.

From the doming tunnel the first row of long holes will be blasted down into the first trough cone. This will be followed by the subsequent rows with sufficient ore being drawn off on 310m level to create a sufficient air gap for each successive row. This will continue until the stope breaks into the overlying old worked out area. If any water or mud exists this will fill up the existing air gap above the broken ore in the stope. This ore will be drawn down in a controlled manner.

16.4.3 360m Level UK4 Mine Lower Block Layout

The UK4 Mine Lower Block will essentially be a duplication of the upper block, but with modifications incorporated from upper block development and operations.

16.4.3.1 Doming Level (340m Level)

The 340m doming level will be very much as per the description of the 290m doming level above.

16.4.3.2 Production Level (360m Level)

The presently planned lay-out makes provision for mining in more K6 material, than in the Upper Block, as seen in Figure 16-5. The actual size and amount of x-cuts will be determined as the development is opened up on the 360m Level. All of the x-cuts will be developed to allow for mining of the K6 material if desired, but only the K4 portion fully developed with the loading and doming tunnels of the stope desired.

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Figure 16-5 360m Level Plan showing UK4 Mine Lower Block Production Lay-out

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16.5 Life of Mine and Production Rates

LDM is now well into the development of the UK Upper Block infrastructure, such that costs and schedules are well-constrained.

16.6 Mining Schedule

The following parameters were used in determining a development schedule (Table 16- 7) for the underground mine:

• Main access and associated development has been planned at an advance rate of 115m per month per crew on multi-blast conditions. Development operations will be carried out by crews working on three shifts per day

• All declines are developed at 12 degrees with 60m between crosscuts connecting the twin decline ramps.

• Longhole production drilling has been based on a rate of 360m per rig shift; giving 15,840 drill metres per rig per month.

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Table 16-7 Production Schedule for UK4 Block

Lace Diamond Mines (Pty) Ltd

Description Crew Jan-16 Feb-16 Mar-16 Apr-16 May-16 Jun-16 Jul-16 Aug-16 Sep-16 Oct-16 Nov-16 Dec-16 Jan-17 Feb-17 Mar-17 Apr-17 May-17 Jun-17 Leg 1

K4 Block - Ore Pass

250 Level

290 Level

310 Level

Total Development 112 80 145 95 50 60 90 100 100 100 100 100 100 100 100 90 90 90

Trough Drilling

Doming Drilling

Total Kimberlite Tonnage Build-Up 3281 2187 6781 13827 14734 18781 30000 30000 30000 30000 30000 30000 30000 30000 30000 30000 30000 30000 Waste Tonnage 2927 2252 2127 1407 0 0 5067 5630 5630 5630 4222 2252 2252 1126 2252 2815 2252 2252 Waste Development Kimberlite Development Block Drilling K4 Kimberlite Production UK4 LB Development

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16.7 Mine Ventilation

16.7.1 Ventilation during Development and Construction

The old shaft with its bypass is used as a Return Air Pass (“RAP”) from 160m level with large enough cross section to up cast 80m3 per second which will be sufficient for initial development of the block with the use of vent pipes and the ultimate production from the UK4 Upper Block, as well as the on-going twin ramp development down to the 330m Level.

A RAP from 310m level and 290m level will consist of a raise bored hole of 2.1m diameter connected to a tunnel developed on 250m level. This tunnel will be connected to the old shaft and used as the return airway for the block. The return air pass will be equipped with a ladder to be used as a second outlet.

16.7.2 Production Ventilation

The RAP at the back of the block will have sufficient negative pressure to down cast air from the twin ramps into the block. A quantity of 10 m3/sec on 310 level and 10 m3/sec on 290m level will be maintained by regulating at the RAP.

16.8 Underground Mobile Equipment and Workshop

A workshop has been developed on 230m level. The workshop has a repair bay, a store, office, cleaning bay, tyre bay, refuelling bay and maintenance bay.

The workshop exhausts the ventilating air directly into the RAP. The diesel will be piped from surface into a diesel tank and the pipe will not store diesel.

An estimate of the number of mobile machines working in the UK4 Upper Block will be one long hole drilling machine , two 10 t LHD’s and four 13 t Trucks .

16.9 Mine De-watering

16.9.1 Dewatering Requirements during the Development Phase

The average volume of water pumped for the 10 months for 2011 was 573 m3 per day. The planned 490m Level pump station with a pumping capacity of 90m3/hr will therefore provide adequate installed pumping capacity for future Block Cave 1 mining operations on 470 Lvl.

For the UK4 Block Mine, LDM has constructed pump stations at Cubby 8 on the 150m level, and at Cubby’s 16 and 17 on the 247m and 250m levels respectively. Two submersible pumps feed the Cubby 16 and 17 pump stations. The Cubby No.8

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main pump station is a permanent installation, equipped with two C5 Envirotech dirty water centrifugal pumps in series fitted with 110 kW motors. This pump station pumps water directly to surface via a 150 mm diameter sleeved borehole to save on piping costs.

LDM has installed a large submersible pump in one of the old vertical passes inside the kimberlite at 290m level, which pumps water from the shaft / old workings up to the inter-level pump station at Cubby No. 17 or 16 in the existing ramp way.

Also LDM has installed a 6” 26kW 13-stage submersible borehole pump in a 10” pilot hole from 250Lvl to 335Lvl with a 8” casing. The water is pumped up to the inter-level Cubby No. 17 or 16 pump station and the water level is currently kept between the 310m and 325m Levels, with the borehole pump able to deliver 60m3/hr.

The Cubby No. 16 pump station consists of a temporary buffer tank with two by two in series KSB 100-250 centrifugal pumps fitted with 45 kW motors which pumps ground water ingress from the two submersible pumps, as well as return water from exploration and development drilling activities, up to the ramp way Cubby No. 8 pump station.

The Cubby No. 17 pump station consists of a coffer dam with two parallel KSB Etanorm 65-50-315 centrifugal pumps fitted with 55 kW two pole motors which pump ground water ingress from the two submersible pumps as well as return water from exploration and development drilling activities, up to the ramp way Cubby No. 8 pump station.

The pumps are automatically controlled and also have adequate capacity to pump solids which may settle out in the dirty water dam compartment. All piping for water management control, on surface and underground, is HDPE piping.

Additional inter-level vertical spindle pump stations for mine dewatering will be installed as and where required as ramp development proceeds deeper.

16.9.2 Dewatering Requirements during the Production Phase

All levels will be graded where possible towards the south east corner of the pipe. A water drain hole is to be established to drain all water down to the pump chamber and keep the ramps clear of water. The pump chamber will have a holding dam directly above it.

The pump chamber will be equipped with a watertight door which can be closed at a moment’s notice in case of flooding. The pump chamber will also be accessible from 470m level via a separate travelling way into the top of the pump chamber to

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enable access to the pump chamber in emergency conditions when the main watertight door is closed in case of flooding or a major pump break down.

16.10 Labour Requirements

LDM’s table of the labour complement required is as set out in Table 16-8 for decline and production development and steady state production respectively.

Table 16-8 Labour Requirements for UK4 Mine

Labour Complement No. Personnel Mine Overseer 1 Shift Boss 3 Labour Complement(Development) No. Personnel Miner 6 LHD Operator 9 Dump Truck Operators 6 Drill Rig Operator 6 Emulsion Operator 6 Miner Assistant 6 General (PVG) 18 Labour Complement(Production) No. Personnel Miner 3 Long Hole Drill Operator 3 Emulsion Operator 3 LHD operators 4 Dump Truck Operators 11 Rock breaker Operator 3 Miner Assistant 6 General (PVG) 6 TOTAL 100

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16.11 Other Parameters and Infrastructure

16.11.1 Surface Lamp Room and Access Control Facility

A Lamp Room and Access Control Facility are to be constructed on surface at the top of the access decline which will cater for clearance and control of all underground employees. LDM use low power lightweight LED mine safety lamps and the Afrox Self-Contained Self-Rescuer (SCSR).

16.11.2 Water Reticulation

Service water will be supplied from the surface plant via a 100 mm water column down the IAP to underground storage dams. Drinking water will be supplied to underground drinking water facilities via a 1 inch water column.

16.11.3 Compressed Air Reticulation

Compressed air will be supplied to the underground workings, refuge chambers and other workings via a 150 mm steel column.

16.11.4 Electricity Reticulation

LACE’s electricity supply will be serviced from the grid of ESKOM. The main electrical supply underground will be established via the IAP supplying 6.6 kV to mini-substations and distributed as 525 V to underground workings as and where required.

16.11.5 Shaft Head Explosives Delivery Bay

Explosives and accessories will be delivered to an approved shaft head explosives delivery bay on surface and transported underground via explosive delivery vehicles and distributed to explosive storage boxes in the various underground sections.

16.11.6 Refuge Bays

Refuge bays will be established not further apart than 750 m in travelling distance in accordance with the relevant DMR Guidelines and Regulations.

16.11.7 First Aid Room & First Aid

A first aid room on surface, manned by an Occupational Health practitioner, will be situated at the processing plant.

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16.11.8 Security

The overall site security of the Lace mining area comprises a perimeter fence with manned main access control booms at the entrance to the mine. Main surface infrastructure areas are well lit and securely fenced and locked off.

16.11.9 Change House / Locker & Laundry Facility

Sufficient change house and laundry facilities will provide adequate ablution, showering, storage and clothes washing facilities to cater for all employees for the decline development and steady state production.

16.11.10 Sewage System

A sewage system will be installed capable of managing on-mine sewage for the number of employees working at LACE during decline development and steady state production later on.

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17.0 RECOVERY METHODS

The Lace processing plant was constructed in 2007 and modified as described in previous sections to be able to handle both kimberlite and tailings feed, at a combined 220 tonnes per hour with 2% DMS yield. The flowsheet is presented below in Figure 17-1.

Figure 17-1 Lace Mine Process Flowsheet

17.1 Process Description

17.1.1 Feed Preparation Plant

A front-end-loader (FEL) will feed 200tph run of mine material into a heavy duty scalping unit which have 500mm spacings.

All +500mm dump oversize will be removed by the scalping unit and broken up by a rock pecker. The -500mm ore is fed by a conveyor into a VGF (vibrating grizzly feeder). The VGF has 150mm spacings which drops out the -150mm onto a conveyor and thus bypasses the primary jaw crusher – the +150mm ore is fed into

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the jaw crusher which is set at 120mm (CCS). The crusher discharge drops onto the same conveyor belt as the -150mm ore and is transported to the primary scrubber. This conveyor is fitted with a weightometer for tonnage measurement and an over- band magnet for tramp iron removal.

The scrubber feed conveyor discharges into the primary 2.4m diameter by 6.0m length scrubber complete with a double skin trommel. The scrubber feed material is mixed with water at a ratio of 1:1 in the scrubber and the material is disagglomerated. Primary scrubber discharge is classified and dewatered with the trommel with -1mm material gravitating to the primary screen under pan; +80mm gravitating to the oversize conveyor and -80+1mm gravitating to the secondary scrubber. Secondary scrubber feed is mixed with fresh water at a ratio of 1:1 and disagglomerated in a 2.0m by 5.0m scrubber. This scrubber is fitted with a bell mouth discharge.

Secondary scrubber discharge gravitates to the 2.4m by 4.8m double deck primary screen. The top deck is fitted with 28mm aperture polyurethane panels and the bottom deck with 1mm slotted apertures polyurethane panels. The material is washed thoroughly on this screen. Screen oversize (+28mm) gravitates to the oversize transfer conveyor and is fed into a surge bin ahead of the secondary cone crusher. This feed bin also receives the +80mm oversize from the primary scrubber. Screen undersize (-1mm) is pumped to the water recovery circuit. Screen product (-28+1mm) is conveyed to the DMS classifying screen and surge bins.

The oversize surge bin (+28mm -150mm) is fitted with a variable speed feeder and feeds the Osborn secondary cone crusher by conveyor belt which is fitted with an over-band belt magnet for tramp iron removal. The cone crusher is choke fed and reduces the product to 28mm and feeds it back to the primary screen by conveyor belts.

17.1.2 Dense Media Separation (“DMS”) Circult

-28+1mm material is conveyed to the DMS classifying screen above the DMS surge bins. This material is classified on the screen into -10mm and +10mm fractions. The primary reason for this separation is that the DMS feed can be processed utilising smaller diameter cyclones, thereby improving DMS cyclone efficiencies and in turn assisting with the recovery of fine diamonds.

-10mm is gravitated to the 50T Fines DMS surge bin. +10mm is gravitated to the 50T Coarse DMS surge bin. Material is extracted from the bins and discharged to the DMS feed conveyors utilising variable speed vibrating feeders. The option exists for fines DMS feed to be bypassed to the Coarse DMS module thereby preventing bottlenecking in the event of excessive fines. The DMS feed conveyors are fitted with weightometers to regulate tonnage.

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The Fines DMS conveyor feeds a 100TPH standard Consulmet DMS module. The Coarse DMS conveyor feeds a 65TPH standard Consulmet DMS module.

Both the Fines and Coarse module utilise conventional DMS technology. In the case of both modules material is de-slimed on feed preparation screens, -1mm is gravitated to the effluent sump and clean screen product is discharged into a mixing box. The material is mixed with correct medium Ferrosilicon (FeSi) slurry in the mixing box and is pumped with the cyclone feed pump to the cast separating cyclones at the correct density. Cyclone inlet pressure is maintained by a variable speed drive on the cyclone feed pump. In the case of the 65TPH the separating cyclone is 1-off 510mm cyclone and the 100TPH module comprises 2-off 420mm cyclones complete with barrel extensions fed by a single distributor.

Cyclone underflow (Sinks concentrate and FeSi) gravitates over static drainage panels and discharges to the sinks screen. FeSi is drained on the static panels and the drain section of the sinks screen. Drained FeSi gravitates to the Correct Medium sump. The concentrate is then rinsed of the remaining FeSi on the second section of the sinks screen using spray water and magnetic separator effluent water. Dilute FeSi gravitates to the dilute medium sump. Clean concentrate from both DMS modules is stored in a 3T surge hopper.

Cyclone Floats gravitates over static panels to the Floats screen and FeSi is drained and rinsed as in the procedure above. The 100TPH fines DMS module is fitted with a double deck floats screen. The top deck aperture is 6mm. +6mm floats is conveyed to the 40T re-crush surge bin. -6+1mm floats is conveyed to the final tailings stockpile. Coarse module Floats is conveyed with Fines +6mm to the re- crush circuit.

Dilute FeSi is pumped to the magnetic separator header box and gravitated to a single stage counter current magnetic separator. This unit recovers FeSi from the dilute slurry. The recovered FeSi (over-dense) is gravitated through a de- magnetising coil to the correct medium sump. Magnetic separator effluent is split into drain and rinse flood box water and final effluent.

Correct medium is pumped to the mixing box with the correct medium pump. The medium density is controlled with an automatic density controller. Densification is achieved with correct medium densifiers. Correct medium from the correct medium sump is pumped with the densifier pump to 100mm tube densifiers (2 running, 1 standby on 100TPH and 1 running 1standby on the 65TPH module)

17.1.3 Re-Crush Circuit

The +6mm floats from the 100TPH Floats screen and the Coarse Floats are conveyed to the 40T re-crush surge bin. The surge bin ensures choke feed on the Oremaster VSI (vertical spindle impactor) crusher. Choke feed is critical to

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guarantee re-crush efficiencies. A variable speed feeder extracts material from the bin and feeds the wet flush crusher. VSI crusher discharge gravitates to a double deck re-crush sizing screen complete with spray bars. The top deck is fitted with 6mm panels, the bottom deck with 1mm panels. +6mm is re-circulated to the re- crush surge bin. -1mm is pumped to the primary screen under-pan. -6+1mm is conveyed to the secondary scrubber feed.

17.1.4 Water Recovery Circuit

Total plant -1mm is pumped to the 25m diameter Supaflo thickener located at or close to the slimes dam. Flocculent and coagulant is added to the dilute thickener feed and the slurry is clarified in this unit. Clean water overflow is gravitated to the process water dam and process water is re-cycled to the plant using the process water pump.

Thickener underflow (dense slurry) is pumped on to the slimes dam and allowed to settle. Clean return water from the slimes dam is pumped back to the plant process water tank.

The thickener control philosophy includes variable speed underflow pumps and slurry densitometers. Flocculent saving instrumentation may be included.

17.1.5 Final Recovery Plant

DMS concentrate is stored in a 20T storage hopper. A vibrating feeder extracts concentrate from the bin and feeds a 1.0m by 3.0m scrubber. This concentrate is scrubbed to polish the material before final concentration. Scrubber discharge is jet pumped to the final recovery building.

The jet pump delivers the material to a de-sliming and classifying screen. The concentrate is washed and drained and then classified into the following fractions - 3+1mm, -6+3mm, -12+6mm and +12mm. These size fractions are stored in secure hoppers of approximately 1m3 capacity each.

The size fractions are fed over grease belts and concentrate is collected in tamper- proof canisters.

Grease tailings are conveyed out of the final sort house building and stored in a secure area.

Final hand sorting, classification and de-falsification are carried out in the glove boxes.

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17.2 Plant Modifications During Bulk Sampling

Lace introduced a de-grit circuit to the flowsheet during Q2 2014, with the bottom screen size having been increased from 1.00 mm to 1.25 mm. The de-grit circuit receives its product from the primary scrubber outlet trommel (double skin with 1.00 mm panels on the outer skin) – before it gets to the secondary scrubber and primary screen. All of the early 290 m level bulk samples (to 290-09), and 310-01, were processed under this configuration, ie. with the bottom screen size at 1.25 mm.

To confirm this, in October 2015 the plant was reconfigured to 1.00 mm screen panels and the balance of the bulk samples were processed accordingly, with results reported in the resource estimate section of this report (Section 14).

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18.0 ENVIRONMENTAL CONSIDERATIONS

This section addresses the tailings handling and storage at the mine. With regard to tailings, the material earmarked for disposal exits the plant as two distinct products, namely a coarse fraction and a fine fraction. The coarse fraction consists of waste kimberlite fragments ranging in size from 6 mm to 1 mm and with a dust component of finer material making up about 10% of this product. This coarse product is known as “tailings”. The other product, which is known as “slimes”, is a much finer product, ranging in size from about 1 mm down to 2μm with the majority of the slimes being less than 100μm.

Lace Diamond Mine (PTY) Ltd has an approved Environmental Management Plan (“EMP”) in terms of the Minerals and Petroleum Resources Development Act 28 of 2002, approved on 5 February 2009. Locations of the slimes dam, processed kimberlite and waste dumps are shown in Figure 3-3b. Lace Diamond Mines is not aware of any other relevant environmental considerations to this report.

18.1 Environmental Management

The environmental management programme of Lace Diamond Mines is presented in chapter 7 of the approved EMP. Lace Diamond Mines recognises that this programme is legally binding, in terms of the Minerals and Petroleum Resources Development Act, 2002 (Act 28 of 2002).

Lace Diamond Mines management system complies with the objectives and principles set out in chapter 7 of the approved EMP. The system is be based on the “BATNEEC” philosophy which is the philosophy of managing potential environmental impacts using the Best (proven) Available Technology Not Entailing Excessive Cost. The philosophy implies that technology being utilised has been found to be effective by practical application and is not so expensive that it will seriously impair the economic viability of a development.

The rehabilitation financial provision is re-calculated in accordance with Regulation 54 of the Minerals and Petroleum Resources Development Act, 2002 (Act 28 of 2002) on an annual basis. The financial provision for the total extent of the outstanding rehabilitation in the event of sudden or premature closure is forwarded to the Regulators based on the yearly calculated quantum.

The rehabilitation is monitored on a monthly basis and an annual environmental awareness plan through the rehabilitation of areas previously disturbed by other mining companies/individuals as well as future disturbances by Lace Diamond Mine (Pty) Ltd is maintained. Lace Diamond mine is committed to responsible mining in the Free State Province. Lace Diamond Mine is not aware of any other relevant environmental considerations to this report.

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18.1.1 Slimes Dam

The slimes dam consists of a ring-dyke formation with the fine product (slimes) being impounded by a ring-dyke of coarse product (tailings).

The slimes dam is managed in accordance with Department of Minerals and Energy (“DME”) Guideline Ref DME 16/3/2/2 – A1 issued by the Chief Inspector of Mines (“CIM”). The Code of Practice (“COP”) is mandatory in terms of section 9 (2) and (3) of the Mine Health and Safety Act, 1996 (Act No. 29 of 1996) (“MHSA”). The COP may be assessed in an accident investigation/inquiry to ascertain compliance and to establish whether the code is effective and fit for purpose. All managerial instructions, recommended procedures and relevant topics comply with the COP and must be reviewed at least biannually or when a major accident or incident has occurred, or at request of a stakeholder. The review will be based on risk assessments, PTO’s (plan task observations), audits and inspections and will incorporate any new strategies, techniques or methods which have become known to the industry.

18.1.2 Processed Kimberlite Dump

Most processed kimberlite material is used on the slimes dam ring dyke, however a small dump has been established in the northwest corner of the mining lease. The processed kimberlite dump is managed in accordance with the approved EMP which allows for co-disposal of Kimberlite tailing on slimes dam which is managed in accordance with the DME Guideline Ref DME 16/3/2/2 – A1 issued by the CIM.

18.1.3 Waste Rock Dump

Most waste rock has been utilized as a dump protecting the dual ramp portal, or as retaining walls for water storage adjacent to the slimes dam as per Figure 3-3b. The dumps are managed in accordance with the approved EMP.

18.2 Social and Community

Lace Diamond Mine (Pty) Ltd. conducts their operation in terms of the approved Mining Right which incorporates their Social and Labour commitments.

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19.0 PROJECT INFRASTRUCTURE

SRK (2012) comprehensively describes the Lace Mine design for the 1.2million tonne per annum (“tpa”) block cave operation which remains the medium-term focus of LDM, while the near-term Upper K4 Mine is brought on stream. Major infrastructure related to the two operations, as well as status as at December 2015, have been presented in the various figures earlier in this report.

19.1 General Overall Underground Infrastructure

The SRK report summarizes the block cave mine as follows (with modifications to their text in brackets which describe actual vs. planned infrastructure):

“LDM has opted for a continuous trough block cave as opposed to a conventional drawbell layout. Continuous troughs can be utilised due to the relatively shallow depth of the cave and the rock strength. An undercut level is thus not utilised (LDM’s doming level is the undercut level).

The mine will be accessed by a conveyor decline in conjunction with a parallel men and materials decline as it will be quicker and cheaper than refurbishing the collapsed shaft system. SRK considers this a sound approach.

The twin ramp will consist of a 3m x 3m (3m x 5m) conveyor decline and a 4.5m x 4.5m men and materials ramp connected by cross cuts at 80m (60m) intervals. The ramps will be 20m (10m) apart to allow space for an airlock to be installed and provide space for a dump truck or LHD between the doors. The ramps will be at a 12° inclination.

The block cave infrastructure comprises three levels, the production level, the doming level and the slot drive, or anti-socket level.

The anti-socket level initially acts as cave footprint delineation level and is later utilised to enable physical observation of the blasted slots. It is located 30 m above the production level and is developed by breaking away from the ramp when the ramp reaches a position 35 m vertically above the production level.

LDM has opted for a high undercut (30m) in order to allow for a higher tonnage output as early as possible before undercutting is complete and the block production builds up. The doming level effectively enables 30m high trough rings to be blasted, thus enabling rapid opening up of the slot to obtain early high tonnage.

The drawpoints on production level will be loaded by LHDs. The west main (rim tunnel) of the production level will serve as a trucking loop with the LHDs loading in the tunnel drawpoints in the western side of the block and then loading into a truck backed into the mouth of the tunnel. These trucks will travel around the block in one direction tipping into the crusher on the east side. The ore from the drawpoints on the east side of the tunnels can be tipped directly into the crusher by the LHD or, if trucks are unavailable,

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the LHDs can tip into the crusher directly from any drawpoint. SRK considers this layout appropriate.

Caving / doming will start to take place as the undercutting span increases but a full collapse or full caving might only take place after 80 m of span is opened. If the end of the block is reached and the block has been fully undercut and caving has not yet taken place then drilling from the anti-socket level will be done to induce caving.

The old shaft with its bypass will be used as a return air pass (RAP) from 240 m level.”

19.2 Infrastructure Additions for UK4 Mine

Section 16 of this report comprehensively describes modifications to the Block Cave infrastructure that have been implemented, or are planned, for the 1,000 tonnes per day (“tpd”) Upper K4 Mine. In summary these are:

• Main Access Decline off of the 250m bulk sample drive for the upper block workings • Doming Levels on 290m and 340m levels • Production Levels on the 310m and 360m levels • Tipping Station/infrastructure for the upper block ore onto the conveyor at 160m level • Main Access Decline off of the twin conveyor ramp at the 340m level • Return Air Way raise bore tunnel from 340m to 250m connecting all UK4 mining drives • Tipping Station/Infrastructure for the lower block ore onto the conveyor at 340m level

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20.0 ECONOMIC ANALYSIS

The economic analysis of mining the UK4 Block at the Lace Mine has been conducted by the discounted cashflow method. The projected mine cashflows have been modelled flat real based on actual costs capital already and currently being incurred, and revenue estimates as at January 2016.

20.1 Introduction

Mining of the UK4 Block forms part of the overall Lace Mine development, where three subsequent block caves are planned over a 25-year mine life. The UK4 mining will share life of mine infrastructure which has been put in place and was devised to provide early positive cashflow from mining while development of the first block cave on the 500m level is completed. The UK4 Block has the potential to be mined at a rate of 35,000 tonnes per month for at least 60 months. It will be mined until the first block cave begins to propagate. Once this occurs, mining will relocate to the 500m level and any ore remaining in the UK4 Block will subsequently fall into the first cave. For modelling purposes, it is assumed the Block is mined for 60 months

20.2 Capital Costs

The major capital cost for the UK4 development has been 1,641 m of tunnel development in waste and kimberlite which would not have otherwise been incurred for the block cave development, and a 1,500 tonne ore pass and tipping arrangement for loading kimberlite mined from the UK4 Block onto the conveyor belt system for hoisting to surface. The total capital cost incurred for the tunnel development, includes drilling and blasting, load and haul, and the installation of water, electrical and ventilation reticulation. It also includes significant additional expenditure on secondary support to ensure tunnel stability in and around old workings.

In addition, one LHD and three 22 tonne dump trucks will be allocated to the project. The total capital cost of these trucks has been allocated to the project economic analysis.

Costs have been incurred in South African Rand (ZAR) and translated to US Dollars (USD) at and exchange rate of 15.0000 ZAR to 1 USD.

Table 20-1 Capital Cost Summary for UK4 Mine Lift 1

Capital item Cost ZAR Cost USD Tunnel development and infrastructure 103,006,620 6,867,066 Tipping arrangement, ore pass, belt 10,722,000 714,800 mechanicals 1 x LHD, and 3 x 22t dump trucks 16,585,700 1,105,713 TOTAL 130,314,320 8,686,869

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Table 20-2 Capital Cost Summary for UK4 Mine Lift 2

Capital item Cost ZAR Cost USD Tunnel development and infrastructure 98,490,000 6,566,000 Tipping arrangement, ore pass, belt 10,722,000 714,800 mechanicals Conveyor belt 6,000,000 400,000 Sustaining Capital 13,476,286 898,419 TOTAL 128,688,286 8,579,219

20.3 Operating Costs

Mining and processing costs for the UK Block have been forecast from actual operating cost data incurred during underground development and tailings retreatment operations at Lace, then factored for delivery of 35,000 tonnes per month to the plant.

Table 20-3 Mining Cost Breakdown for UK4 Mine (ZAR per month) at 35,000 tpm

Cost centre ZAR per month ZAR per tonne Labour 1,809,855 60.33 Fuel 884,949 29.50 Oil, R&M 768,155 25.61 Tyres 301,000 10.03 Underground electricity and vent 496,800 16.56 Conveyor R&M 86,215 2.87 Drill & Blast 472,200 15.74 Dewatering 130,000 4.33 PPE & Consumables 101,250 3.38 Contingency 15% 757,564 25.25

TOTAL 5,807,988 193.60

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Table 20-4 Processing Cost Breakdown for UK4 Mine (ZAR per month) at 30,000 tpm

Cost centre ZAR per month ZAR per tonne

Labour 1,244,636 22.63 Water & Electricity 425,000 7.73 Consumables (gas, welding rods etc) 26,800 0.49 Contracted Labour/consultants 18,000 0.33 Plant Oil & Greases 28,140 0.51 Plant R&M (incl electrical and crushing) 330,000 6.00 FeSi 194,810 3.54 Flocculent 89,375 1.63 Gypsum 27,500 0.50 Loose Tools 6,000 0.11 Protective Clothing 10,000 0.18 Buses - Plant 25,000 0.45 Diesel 129,980 2.36 Rolling stock R&M – Scheduled Service 116,400 2.12 Small vehicles 12,000 0.22 Contingency 15% 215,851 3.92 TOTAL 2,899,492 52.72

In addition to these mining and processing costs, a 50% share of Lace mines’ current general and administration costs have been incorporated as a cost applied to the UK4 Block mining. The remaining 50% are applied to ongoing development of the first block cave.

20.4 Revenue

Revenue forecasts are based on the independent valuation conducted by DiamondCorp’s Antwerp diamond valuation agents in January 2016 of Lace diamonds recovered from bulk testing together with the revenue forecast modelling based on micro and macrodiamond recoveries detailed in Section 14 herein. An average carat value of $164 per carat is used in the modelling.

20.5 Cashflow

Mining of the UK4 block has been modelled over 60 months at a mining rate of approximately 35,000 tonnes per month. The cashflows have been modelled flat real (i.e. uninflated) with a 10% discount rate applied.

The model generates an NPV of R133.3 million (US$8.9 million) and a robust IRR of 59%.

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Table 20-5 Cash Flow Analysis for UK4 Mine UK4 Block - Cashflow

Assumptions USD:ZAR FX 15.0000 Mining Cost (R/t) 193.60 Processing Cost (R/t) 52.72 Average grade (cpt) 0.23 Carat value 164.00

2014-2015 2016 2017 2018 2019 2020 2021 Total H1 H2 H1 H2 H1 H2 H1 H2 H1 H2 H1 H2 UK4 Tonnes 60,500 194,521 184,354 177,226 181,357 175,694 185,091 177,339 189,892 179,548 190,735 192,791 2,089,048 Recovered grade (cpt) 0.28 0.29 0.24 0.23 0.21 0.23 0.23 0.26 0.26 0.19 0.20 0.17 Diamonds recovered (cts) 17,182 56,217 43,323 41,471 37,904 40,761 42,201 45,221 48,992 34,653 37,575 32,774 478,274

Revenue 42,267,720 138,292,760 106,575,047 102,018,375 93,242,888 100,272,080 103,813,840 111,244,755 120,520,655 85,245,799 92,433,996 80,625,196 1,176,553,110 (less marketing costs & royalties) 4,564,914 14,935,618 11,510,105 11,017,984 10,070,232 10,829,385 11,211,895 12,014,434 13,016,231 9,206,546 9,982,872 8,707,521 127,067,736 Net revenue 123,357,142 95,064,942 91,000,390 83,172,656 89,442,695 92,601,945 99,230,321 107,504,424 76,039,253 82,451,124 71,917,675 1,011,782,568

Mining costs 37,659,187 35,690,860 34,310,882 35,110,642 34,014,287 35,833,543 34,332,759 36,763,014 34,760,420 36,926,219 37,324,259 392,726,071 Processing costs 10,254,766 9,718,781 9,343,007 9,560,785 9,262,243 9,757,635 9,348,964 10,010,734 9,465,418 10,055,175 10,163,563 106,941,072 General & Admin (50% share) 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 4,408,264 52,899,171 Total costs 4,408,264 52,322,217 49,817,905 48,062,153 49,079,691 47,684,795 49,999,441 48,089,987 51,182,012 48,634,103 51,389,658 51,896,087 552,566,314

Operating profit -4,408,264 71,034,925 45,247,037 42,938,237 34,092,965 41,757,900 42,602,504 51,140,334 56,322,412 27,405,150 31,061,466 20,021,588 459,216,254

Capital (including sustaining capital) 97,735,740 16,289,290 16,289,290 30,149,697 30,149,697 30,149,697 30,149,697 1,348,250 1,348,250 1,348,250 1,348,250 1,348,250 1,348,250 259,002,606 Credit from development revenue -31,251,448 Cashflow -97,735,740 10,553,893 54,745,635 15,097,340 12,788,540 3,943,268 11,608,204 41,254,254 49,792,084 54,974,162 26,056,900 29,713,216 18,673,338 200,213,648 Cumulative -97,735,740 -87,181,847 -32,436,212 -17,338,871 -4,550,331 -607,062 11,001,142 52,255,396 102,047,480 157,021,641 183,078,541 212,791,758 231,465,095 431,678,743

NPV @ 10% discount 133,256,616 IRR 59%

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20.6 Sensitivities

The UK4 mine cashflow was tested for sensitivity to the key variables of grade, carat value, exchange rates, mining cost and processing cost.

The mine is most sensitive to grade, carat value and exchange rates. These three factors are directly correlated, all of which can individually have a significant impact on NPV. A 20% fall in any of these three variables can reduced the NPV of the mine to zero.

The mine is less sensitive to mining costs, where a 50% increase would be required to reduce the NPV to zero.

The mine is not particularly sensitive to increases in processing costs, where a 50% increase would reduce the NPV from R133 million to R97 million.

Figure 20-1 Sensitivity Analysis for UK4 Mine

600

500

400

300

200

100

0 -50% -40% -30% -20% -10% 0% 10% 20% 30% 40% 50% -100

-200

-300 Mine NPV (R million)

Grade/$/ct/FX Mining Processing

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21.0 OTHER RELEVANT DATA AND INFORMATION

The authors are not aware of any other relevant data or information to this report, other than that below.

21.1 Outlook for Rough Diamond Prices

The 2015 Bain Report has recently been published and provides a useful overview of the rough market presently, and moving forward, as they conclude:

• The world rough-diamond demand in the next 15 years is fore-casted to grow at an average annual rate of about 3% to 4%, and the supply is projected to decline by 1% to 2%, causing the gap between supply and demand to widen starting in 2019. The forecast reflects fundamental supply and demand factors rather than short-term fluctuations or unforeseeable long-term macroeconomic shifts. • Our forecast of the rough-diamond supply is based on the analysis of existing mines, publicly announced plans and anticipated production at every expected new mine. We foresee the global supply of rough declining on average by 1% to 2% per year from 2015 to 2030 because of the aging and depletion of existing mines and relatively little new supply coming online.

Bain expect the current weakness in the market to be short-lived (maximum 18-24 months and probably shorter) based on past sector performance in addressing issues as per their Figure 6 below:

Bain sees the new mines coming on-stream including Lace to add considerable carats to the pipe line, but for overall supply to diminish starting around 2020 as per their figures 35 and 36 below:

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Bain expect the gap between supply and demand to begin to widen in 2019 as per their figure 39 below:

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21.2 Current Contracts

LDM (through DiamondCorp Holdings Limited) has an off-take agreement signed in November 2012 with Laurelton Diamonds Inc., a wholly owned subsidiary of Tiffany & Co. The off-take agreement allows Laurelton to purchase, on commercial terms related to fair market value, diamond production from Lace which meets their high quality and colour requirements for polishing. The off-take does not include large (+10.5 carat) and special (>$5,000/carat) stones. The off-take agreement was in exchange for a USD $6 million term loan that bears interest at 9% per annum and is due on the 8th anniversary of the second draw down, being the 10th of April, 2013. No interest or principle payments were due for the first three years, with interest accumulating during this period. LDM has just begun to make payments on the loan.

All production not purchased by Laurelton is marketed by LDM in Johannesburg and Antwerp on regular intervals.

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22.0 INTERPRETATIONS AND CONCLUSIONS

The presently reported drilling and sampling programs within the UK4 Mine block have allowed for increased understanding and confidence in the geology, volume, tonnage, and mineralization within this depth slice of the overall Lace Main Pipe. This in turn has allowed for updated, higher confidence Mineral Resource estimates within that portion of the pipe, and greater confidence in interpretations of these same resource parameters at greater depth. The present methodology of systematic delineation core drilling from underground drill stations, and audited bulk sampling of development drives, is proving effective. As the Lace production plant is being utilized, the recovered grades from the bulk sampling represent true, non-factorized, commercially attained results based on the present plant flowsheet, and have been utilized for all resource estimates. In areas where the delineation drilling has attained the desired density of ~25m spacings between boreholes, geologically reliability has been found to be high.

From a commercial perspective, the UK4 Mine plan which includes a portion of the indicated mineral resource, is deemed economically viable and robust based on real cost parameters, determined during DiamondCorp’s development period, still in progress and within four months of attaining name-plate production. The Indicated Mineral Resource for the UK4 Mine block along with these cost and revenue parameters is the basis for the first industry-standard Probable Mineral Reserve at Lace, dedicated solely at present to the UK4 Mine operation.

The cash flows have been modelled flat real (i.e. uninflated) with a 10% discount rate applied, and major parameters applied from LDM’s actual costs as follows:

USD:ZAR Exchange Rate 1:15.0

Mining Costs/Tonne ZAR 193.60

Processing Costs/Tonne ZAR 52.72

Average Recovered Grade 23cpht

Carat Value ($/carat) 164.00

The model generates an NPV of R133.3 million (US$8.9 million) and a robust IRR of 59%.

The UK4 Mine is most sensitive to grade, carat value and exchange rates. The mine is less sensitive to increases in mining and processing costs.

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Geological, dilution, density and mineralization (grade) models have been constructed for the UK4 Mine Block depth slice (-230m to -370m) of the Lace Main Pipe, and the updated geological interpretation extended to the base of the present model at -920m (410m absl). The grade model has been found to show high precision in predicting ultimate recovered grades achieved in the bulk sampling on -250m, -290m and -310m levels.

Further evaluation work of the same types will allow for progressively deeper depth slices of Lace to be moved into the Indicated Resource category, and thereafter into reserves for the pending, larger block cave mining operations.

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23.0 RECOMMENDATIONS

It is recommended that LDM and DiamondCorp continue the present evaluation program, which can be improved upon slightly to increase precision. Specifically the usage of digital borehole directional survey equipment will assist the modelling by bringing an added measure of precision to the plotting of all geological and sampling borehole data. Deviations of the latest round of boreholes are not known, and would undoubtedly be present in holes reaching +200m with relatively shallow dips.

As well the delineation program needs a dedicated drill cubby in the western quadrant of the pipe at either the -290 or -310m level which will allow for arrays of holes to be drilled into the western and northwestern portions of the pipe, beneath the majority of the historical workings.

Microdiamond work would benefit from the usage of an umpire facility for a representative selection of samples, as well as from spiking of samples before they are sealed and sent to the laboratories

23.1 Further Delineation and Bulk Sampling for UK4 Mine Lower Block

On-going delineation drilling, although primarily designed to provide information from below the -370m level, will serve to improve the geological knowledge and sampling database for the lower block of the UK4 Mine depth slice, as the holes will be collared at either the 290 or 310m levels. Continued assimilation of this data is recommended.

Similarly the CP’s recommend that development drives on the -340m doming level, and the -360m production levels, be systematically bulk and microdiamond sampled as has been documented here-in. As well, it is essential to attain delineation and bulk sampling data into the CRB units on the northern quadrant of the pipe such that this large facies can be better understood and a reliable grade estimate ascribed to it.

23.2 Further Delineation and Bulk Sampling for Block Cave 1

Systematic core drilling and microdiamond sampling of this block will be essential to higher confidence resource estimates, as well as bulk sampling of the initial development drives on the -480m level. As well we’d recommend a single development drive for bulk sampling purposes at the -420m level, ie. midway between the -360m UK4 mining level, and the -480m Block Cave 1 production developments. This will allow for a higher degree of correlation between the microdiamond and macrodiamond sampling estimates for this block.

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24.0 REFERENCES

Anglo American plc, 2014: Annual Report 2014.

Bain & Company Inc., 2015: The Global Diamond Industry 2015 – Growth Perspectives Amid Short-term Challenges.

Beetz, P.F.W., 1931: Crown Diamond Mining and Exploration Company Ltd. – Summary of the Crown Mine, Crown Diamond Mining and Exploration Company Ltd. memo summarizing information known about the Crown Mine.

Bruton, E., 1970; Diamonds (2nd Edition): Published by N.A.G. Press Ltd., London England. ISBN 0-8019-6789-9.

Clement, C.R., 1978: Lace Mine; DeBeers Consolidated Mines Ltd. memo summarizing information known about the Crown Mine.

CM Reports: 1913-1932: Minutes of Annual General Meetings, Mine Managers and Consulting Engineers Report, of the Crown Diamond Mining and Exploration Company Limited.

DiamondCorp plc, 2007: Annual Accounts and Letter to Shareholders, including Audited Results for the Period Ended 31 December, 2006.

DiamondCorp plc, 2008: Annual Report 2007 – Pipe to Prosperity, including Audited Results for the Period Ended 31 December, 2007.

DiamondCorp plc, 2009: Annual Report 2008, Audited Results for the Period Ended 31 December, 2008.

DiamondCorp plc, 2011a: Consolidated Financial Statements for the Year Ended 31 December, 2010.

DiamondCorp plc, 2011b: Operational Update. News Release dated 05 October, 2011.

DiamondCorp plc, 2012: Consolidated Financial Statements for the Year Ended 31 December, 2011.

DiamondCorp plc, 2013a: Audited Consolidated Annual Financial Statements for the Year Ended 31 December, 2012.

DiamondCorp plc, 2013b: Lace Mine Project Update. News Release dated 31 October, 2013.

DiamondCorp plc, 2014a: Lace Mine Project Update. News Release dated 30 January, 2014.

DiamondCorp plc, 2014b: Lace Mine Project Update. News Release dated 30 April, 2014.

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DiamondCorp plc, 2014c: Audited Consolidated And Separate Financial Statements for the Year Ended 31 December, 2013.

DiamondCorp plc, 2014d: Lace Diamond Mine (Pty) Ltd. Scoping Study, Mine Method Selection & Mine Design – 310 Level – Upper K4 Block Underground Mine. Report No LDM/July 14/01 Rev 01, July 2014.

DiamondCorp plc, 2015a: Audited Consolidated And Separate Financial Statements for the Year Ended 31 December, 2014.

DiamondCorp plc, 2015b: Lace Diamond Mine Project Update. News Release dated 30 April, 2015.

DiamondCorp plc, 2015c: Lace Diamond Mine Project Update. News Release dated 31 July, 2015.

DiamondCorp plc, 2015d: Conveyor Belt Installation Complete. Commissioning Underway. Diamond Sales and Valuation. News Release dated 13 October, 2015.

DiamondCorp plc, 2015e: Lace Diamond Mine Project Update. News Release dated 03 November, 2015.

DiamondCorp plc, 2015f: IDC Loan Reschedule Approval. Proposed Placing to Raise up to £4 million. Operational Outlook. News Release dated 02 December, 2015.

De Beers Annual Report 2000, Review of Operations.

Dixon, R., 1979: Crown Diamond Mine Summary, DeBeers Consolidated Mines Ltd. memo summarizing information known about the Crown Mine.

DuToit, D., 1967: Notes on Crown Mine, DeBeers Consolidated Mines Ltd. memo summarizing information known about the Crown Mine.

Field, M., Stiefenhofer, J., Robey, J. and S. Kurzlaukis, 2008: Kimberlite-hosted Diamond Deposits of Southern Africa: A Review. Ore Geology Reviews 34 (2008) 33-75.

Howarth, G.H., 2010: Geology of the Kroonstad Kimberlite Cluster, South Africa. M.Sc. Thesis – Rhodes University.

Kujawa, T., 1995: Lace Mine Dump Sampling Programme – Summary Report, DeBeers Consolidated Mines Ltd.

MPH Consulting Limited, 1997: Status Report on Tailings Re-Processing Project, Crown Diamond Mine December 1997; project reports for Rupert Resources Ltd.

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MPH Consulting Limited, 1998a: Update on Tailings Re-processing Project, Crown Diamond Mine May 31, 1998; project reports for Rupert Resources Ltd.

MPH Consulting Limited, 1998b: Final Report on Tailings Sampling Project Crown Diamond Mine for Rupert Resources Ltd. Republic of South Africa, September 28, 1998

MPH Consulting Limited, 1998c: Report on Delineation Drilling and Preliminary Scoping Studies on the Crown Diamond Mine for Rupert Resources Ltd., October, 1998;

MPH Consulting Limited, 2005: Report on the Lace Diamond Mine for DiamondCorp PLC, June 2005; MPH Consulting Limited, 2006a: Update to the Definitive Cost Estimate for the Dump Re- treatment Project, Lace Diamond Mine for DiamondCorp PLC, 31 March 2006.

MPH Consulting Limited, 2006b: Competent Person’s Report on the Lace Diamond Mine Project, Republic of South Africa, for DiamondCorp PLC, November 2006.

MPH Consulting Limited, 2007: Competent Person’s Report on the Lace Diamond Mine Project, Republic of South Africa, for DiamondCorp PLC, August 2007.

Natural Diamond Corporation, 2015: Market Valuation, Lace Diamonds. Letter to Paul Loudon, by Samuel Scheffer, Director Natural Diamond Corporation NV, 12 October, 2015.

Natural Diamond Corporation, 2016: Market Valuation, Lace Diamonds. Letter to Paul Sobie, by Samuel Scheffer, Director Natural Diamond Corporation NV, 8 January, 2016.

Snowden, 2006: Risk Assessment for Lace Mining – Removing the Plug from the Shaft, 05 October 2006.

SRK Consulting, MPH Consulting Limited and ETS, 2000: Crown Diamond Mine Definitive Cost Estimate for the Dump Re-treatment Project, for Rupert Resources Ltd., February 2000.

SRK Consulting, MPH Consulting Limited and Crosston Technologies cc, 2005: Lace Diamond Mine Definitive Cost Estimate Dump Re-treatment Project for Diamondcorp PLC, June 2005 Report 350849.

SRK Consulting, 2012: Independent Engineer’s Report on the Lace Diamond Mine, Free State, South Africa – Report Prepared for DiamondCorp plc. Report No 443261/Lace Mine IER 19 March 2012

Stewart, A.D.H., 1994: Crown Mine – Summary of Existing Data, DeBeers Consolidated Mines Ltd. memo summarizing information known about the Crown Mine.

VP3 Geoservices (Pty) Ltd., 2012: A Resource Review of the Lace Mine, Kroonstad, South Africa for DiamondCorp plc. 07 March, 2012.

Wikipedia, 2016: Köppen Climate Classification System

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25.0 CERTIFICATES OF QUALIFICATION

Paul A. Sobie 179 Guelph St., P.O. Box 278, Rockwood, Ontario, N0B 2K0 Tel: 416-365-0930 Fax: 416-365-1830 Email: [email protected]

I, Paul A. Sobie, B.Sc., am a Professional Geologist, residing at 179 Guelph Street, Rockwood, Ontario, and hereby certify:

1. I am a member in good standing of the Association of Professional Geoscientists of Ontario, Membership # 0374;

2. I am a graduate of Laurentian University, Sudbury, Ontario with a B.Sc. Honours degree in geology in 1987;

3. I have been employed continuously from 1987 as a geologist with MPH Consulting Limited, and during the period 1993-2006, as managing director of MPH Consulting Botswana (Pty) Limited, a subsidiary company, located in Gaborone, Botswana, MPH Consulting Lesotho (Pty.) Limited, in Maseru, Kingdom of Lesotho, and of MPH Consulting South Africa (Pty) Limited, in Johannesburg, Republic of South Africa;

4. I am President and managing director of MPH Consulting Limited (“MPH”) of Toronto, the parent company;

5. I have a total of twenty-eight years of direct experience with diamond projects in Canada, the Russian Federation and Western and Southern Africa, including managerial responsibilities for projects ranging from conceptual grassroots exploration through to feasibility studies on advanced deposits. Additional experience has included independent valuations and production audits on producing mines, as well as verification/audit work for parties completing due-diligence, and/or participating in joint ventures;

6. I have historical experience and knowledge of the Lace Diamond Mine project, having been the founder of the project in 1996, and having managed the initial exploration and evaluation work which started the modern technical database for the project;

7. I have been consulting with present owners DiamondCorp plc since 2005, and personally managed all facets of the UK4 Mine block evaluation program described in this report;

8. As a result of my education, professional experience and professional qualifications, I am a competent person as defined by SAMREC, and in the London Stock Exchange Guidance Note for Mining, Oil and Gas Companies dated March 2006;

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

9. I have visited the Lace Diamond Mine Project (the “Property”) numerous times, the most recent being during September 2015 at the bequest of DiamondCorp plc;

10. This report was prepared under my supervision, with maps and figures prepared by MPH and consulting geologists under my direction;

11. The sources of information and data not based on personal examinations were obtained from the references cited in the report in connection with such information, which to the best of my knowledge and experience is correct;

12. I am not aware of any material fact or material change with respect to the subject matter of this technical report which is not reflected in this report, the omission to disclose which would make this report misleading;

13. I have no interest, direct or indirect, in the Property nor do I have any beneficial interest, direct or indirect, in the securities of DiamondCorp plc, or any of the companies mentioned in this report;

Dated this 24th of February, 2016.

(Signed) Paul Sobie Paul A. Sobie, B.Sc., P.Geo.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

CERTIFICATE OF QUALIFICATION

Johannes J. Ferreira 11 Village Close, Wokingham, Berkshire, United Kingdom, RG41 1FZ Tel: +44 1189 783908 Mob: +44 7760343038 Email: [email protected]

I, Johannes Ferreira, PhD, am a Geo-Statistician specialising in primary diamond resources, residing at 11 Village close, Wokingham in the United Kingdom, and hereby certify:

1. I am a member in good standing of the South African Council for Natural Scientific Professions, Membership # 400047/06;

2. I am a graduate of the Paris Institute for Technology, France with a PhD, the title of the thesis is ‘Sampling and Estimation of diamond content in kimberlite based on microdiamonds’, completed in 2013;

3. I have been employed continuously from 1981 as a geo-statistician with De Beers and Anglo American in South Africa and England, and since my retirement in 2008 I have been Director and consultant with Johan Ferreira and Associates limited. I joined Kleingeld, Young and Partners as Director in 2015;

4. I have a total of thirty-five years of direct experience with diamond projects in Africa, Canada, Australia and the Russian Federation, including being responsible for resource aspects in projects ranging from exploration through to feasibility studies on advanced deposits.;

5. I have been involved with the Lace Diamond Mine project since 2012, having been requested initially to demonstrate that microdiamond sampling may be used for resource estimation at the mine;

6. As a result of my education, professional experience and professional qualifications, I am a competent person as defined by SAMREC;

7. The sources of information and data not based on personal examinations were obtained from the references cited in the report in connection with such information, which to the best of my knowledge and experience is correct;

8. I am not aware of any material fact or material change with respect to the subject matter of this technical report which is not reflected in this report, the omission to disclose which would make this report misleading;

9. I have no interest, direct or indirect, in the Property nor do I have any beneficial interest, direct or indirect, in the securities of Diamond Corp plc, or any of the companies mentioned in this report;

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Dated this 29th of February, 2016.

Johannes J. Ferreira, PhD, Pr. Sci. Nat.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

CERTIFICATE OF QUALIFICATION

Paul R. Loudon Glendree, Feakle Co. Clare, Ireland Tel: +27-216 1300 Fax: +27 56 216 1352 Email: [email protected]

I, Paul R. Loudon, B.Sc., am a professional mining executive with 30 years experience in mining corporate finance and company management, and hereby certify:

14. I am a member in good standing of the Southern Africa Institute of Mining and Metallurgy # 706114;

15. I am a graduate of Open University, Milton Keynes, United Kingdom with a B.Sc. first class honours degree in economics and engineering in 2015;

16. I have been employed continuously in the mining and mining corporate finance sectors since 1985, having previously held the positions of mining analyst Intersuisse Limited, senior vice president corporate development Indomin Limited, chairman, BDI Mining Limited, and head of equity finance, Loeb Aron & Company Limited;

17. I am currently chief executive officer and a director of DiamondCorp plc, a position I have held continuously since 2005;

18. I have a total of thirty years of direct experience with analysing and financing mining projects around the world, including financing and managing diamond mining projects in Zimbabwe, Indonesia and South Africa;

19. I have been directly involved in overseeing the sale of rough diamonds since 1996.

20. The sources of information and data used in the economic analysis of the Lace project are based on first principles, having access to all current and historical cost data on the project since its acquisition by DiamondCorp in 2005;

21. I have a direct and indirect interest in the securities of DiamondCorp plc.

Dated this 28th of February, 2016.

(Signed) Paul Loudon Paul R. Loudon, B.Sc. (Hons), MSAIMM.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

CERTIFICATE OF QUALIFICATION

Paul Guthrie Allan 46 Kingston Drive Umhlanga Rocks, South Africa Tel: +27 31 561 5585 Mob: +27 82 821 1003 Email: [email protected]

I, Paul Allan, am an independent precious stones geologist specializing in diamonds and rubies, residing at 46 Kingston Drive, Umhlanga Rocks, South Africa, and hereby certify:

1. I am a member in good standing of the South African Geological Society.

2. I am a graduate of the University of the Witwatersrand (BSc Hons, 1988)

3. I have a total of twenty eight years of experience in the exploration and evaluation of precious stone deposits, principally diamonds in kimberlite.

4. I was involved with the Lace Diamond Mine project during the period of deep drilling (1997 – 2000) and have again become involved as Project Geologist (as and when required) since 2014.

5. The sources of information and data not based on personal examinations were obtained from the references cited in the report in connection with such information, which to the best of my knowledge and experience is correct.

6. I am not aware of any material fact or material change with respect to the subject matter of this technical report which is not reflected in this report, the omission to disclose which would make this report misleading.

7. I have no interest, direct or indirect, in the Property nor do I have any beneficial interest, direct or indirect, in the securities of Diamond Corp plc, or any of the companies mentioned in this report.

Dated this 29th of February, 2016.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

APPENDIX 1 – GLOSSARY OF SELECTED GEOLOGICAL AND MINING TERMS

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

“adit” An adit is an entrance to an underground mine which is horizontal or nearly horizontal, by which the mine can be entered, drained of water, ventilated, and minerals extracted at the lowest convenient level. “Aeolian” Aeolian processes, also spelled eolian or æolian, pertain to wind activity in the study of geology and weather and specifically to the wind's ability to shape the surface of the Earth (or other planets). Winds may erode, transport, and deposit materials and are effective agents in regions with sparse vegetation, a lack of soil moisture and a large supply of unconsolidated sediments. “alluvium” Alluvium is loose, unconsolidated (not cemented together into a solid rock) soil or sediments, which has been eroded, reshaped by water in some form, and redeposited in a non- marine setting. Alluvium is typically made up of a variety of materials, including fine particles of silt and clay and larger particles of sand and gravel. When this loose alluvial material is deposited or cemented into a lithological unit, or lithified, it is called an alluvial deposit. “airborne magnetic survey” A survey conducted from the air for the purpose of recording the natural magnetic characteristics of rocks on and below the surface of the earth “amygdaloidal” Volcanic rock in which rounded cavities formed by the expansion of gas or steam “aphanitic” A fine-grained igneous rock having such compact texture that the constituent minerals cannot be detected with the naked eye “Archean” Refers to the early Precambrian period from 4billion to 2.5billion years ago “basalt” Dark colored, fine-grained volcanic igneous rock “block cave” A low-cost method of bulk mining in which large blocks of ore are undercut and the supporting pillars are blasted away, causing the ore to cave under its own weight. “blow” A thickening (usually more than 10m thick) of a dyke “blue ground” Colloquial term from early diamond mining in South Africa, blue ground refers to a layer of non-oxidized kimberlite, much harder than the weathered “yellow ground” nearer to surface. “breccia” Rock composed of sharp-angled foreign fragments embedded in a fine-grained matrix “calcite” The mineral calcium carbonate “carat” or “ct” A unit of weight for diamonds, equivalent to 0.2 of a gram

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

“care and maintenance” A term used in the mining industry to describe processes and conditions on a closed minesite where there is potential to recommence operations at a later date. During a care and maintenance phase, production is stopped but the site is managed to ensure it remains in a safe and stable condition. “chambering/partial chambering” An underground mining method where large stopes (“chambers”) are evenly spaced with pillars between on one level, with the stopes positioned beneath the pillars on the next level down, allowing for the upper pillars to collapse into the lower chambers. “cofferdam” Temporary enclosure built within a body of water and constructed to allow the enclosed area to be pumped out, creating a dry work environment for the major work to proceed. “competency/competent” Rock which, because of its physical and geological characteristics, is capable of sustaining openings without any structural support except pillars and walls left during mining (stalls, light props, and roof bolts are not considered structural support). “COP” Code of practise “country rock” The rock bodies which enclose an intrusive mass of igneous rock. “cpht” Carats per hundred tonnes “CRB” Country rock breccia “cyclone” A hydrocyclone (cyclone) is a device to classify, separate or sort particles in a liquid suspension based on the ratio of their centripetal force to fluid resistance. This ratio is high for dense (where separation by density is required) and coarse (where separation by size is required) particles, and low for light and fine particles. A hydrocyclone will normally have a cylindrical section at the top where liquid is being fed tangentially, and a conical base. The angle, and hence length of the conical section, plays a role in determining operating characteristics.

“crosscut” A cross-cut (“x-cut”) is a horizontal underground passageway that provides access to mining operations and is usually bored from the mining shaft at near right angles to the strike of a vein or orebody. “diabase” Diabase or dolerite or microgabbro is a mafic, holocrystalline, subvolcanic rock equivalent to volcanic basalt or plutonic gabbro. Diabase dikes and sills are typically shallow intrusive bodies and often exhibit fine grained to aphanitic chilled margins which may contain tachylite (dark mafic glass).

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

“diamond resource” A concentration or occurrence of diamonds of economic interest in or on the earth’s crust in such form, quantity (volume/tonnage), grade and value that there is reasonable and realistic prospects for eventual economic extraction. The location, quantity, grade, value, continuity and other geological characteristics of a Diamond Resource are known, or estimated from specific geological evidence, sampling and knowledge interpreted from an appropriately constrained geological model. “diopside” A crystallized variety of pyroxene, of a clear, grayish green colour. “DDH” Diamond drill hole. “DME” Department of Mines and Energy “DMS” Dense media separation. “DTM” Digital terrain model. “drive” Horizontal tunnels or excavations. “dyke or fissure” A sheet-like body of igneous rock which is discordant i.e. cuts across the bedding or structural planes of the host rock. “EMP” Environmental management plan “ESKOM” Eskom is a South African electricity public utility, established in 1923 as the Electricity Supply Commission by the government of South Africa in terms of the Electricity Act. “facies” Distinctive rock sub-type. “feasibility study” A comprehensive study, including final engineering, undertaken to determine the economic viability of a project; the conclusion will determine if a production decision can be made and is used for financing arrangements. Typically, the accuracy of these studies is in the +/- 10 per cent range. “friable” Easily crumbled or pulverized. “garnet/pyrope” Important constituent mineral of kimberlite. “geophysical survey” The exploration of an area in which physical properties relating to geology are used. Geophysical methods include seismic, magnetic, electromagnetic, gravity and induced polarization techniques. “Geological model/modelling” The applied science of creating computerized representations of portions of the Earth's crust based on geophysical and geological observations. Geologists involved in mining and mineral exploration use geologic modelling to determine the geometry and placement of mineral deposits in the subsurface of the earth. Geologic models help define the volume and

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

concentration of minerals, to which economic constraints are applied to determine the economic value of the mineralization. Mineral deposits that are deemed to be economic may be developed into a mine. “GPS” Global positioning system “grade” The relative mass of diamonds in a mass of rock “grizzly” Grid of iron bars that allows ore of the correct size to travel down the ore pass to the bottom of the mine, ready for hoisting to the surface. Or at the primary feed point to the processing plant, before primary crushing. “HDPE” High density polyethylene piping “hypabyssal” A magmatic intrusion which has solidified before reaching the earth’s surface “indicated mineral resource” That part of a diamond resource for which tonnage, densities, shape, physical characteristics, grade and diamond value can be estimated with a reasonable level of confidence. It is based on exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. The data must be of sufficient confidence to allow the application of modifying factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit. locations are too widely or inappropriately spaced to confirm geological and/or grade continuity but are spaced closely enough for continuity to be assumed, and sufficient diamonds have been recovered to allow a reasonable estimate of average diamond value. “inferred mineral resource” That part of a diamond resource for which tonnage, grade and diamond value can be estimated with a low level of confidence. Geological evidence is sufficient to imply but not verify geological and grade continuity. It is based on information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes which may be of limited or uncertain quality and reliability. “in-situ” Rock occurring as was originally emplaced with all associated after emplacement episodes that have tectonically and structurally influenced the rock as seen today. “interpolation” In the mathematical field of numerical analysis, interpolation is a method of constructing new data points within the range of a discrete set of known data points. “Inverse Distance” Inverse distance weighting is the simplest interpolation method. A neighborhood about the interpolated point is

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

identified and a weighted average is taken of the observation values within this neighborhood. The weights are a decreasing function of distance. The user has control over the mathematical form of the weighting function, the size of the neighborhood (expressed as a radius or a number of points), in addition to other options. “joint” A divisional plane or surface that divides a rock and along which there has been no visible movement parallel to the plane or surface. “kimberlite” An uneven grained, ultramafic, intrusive rock in which the visible minerals may include olivine, phlogopite, pyrope garnet, picroilmenite and chrome-diopside cemented by a groundmass, which may include serpentine, calcite and chromate. Kimberlite may be diamondiferous and, along with olivine lamproites, are the only know primary source of diamonds “kriging” An interpolation technique in which the surrounding measured values are weighted to derive a predicted value for an unmeasured location. Weights are based on the distance between the measured points, the prediction locations, and the overall spatial arrangement among the measured points. Kriging is unique among the interpolation methods in that it provides an easy method for characterizing the variance, or the precision, of predictions. Kriging is based on regionalized variable theory, which assumes that the spatial variation in the data being modeled is homogeneous across the surface. That is, the same pattern of variation can be observed at all locations on the surface. “LHD” Load, Haul, Dump machine – a low profile, underground mining front-end loader (“FEL”). “LOM” Life of Mine. “Ma” Millions of years ago “macrocrysts” A large crystal “macrodiamonds” Diamonds that are greater than 0.5mm in the longest axial dimension “magmaclasts” Also termed 'pelletal lapilli'—well-rounded clasts consisting of an inner 'seed' particle with a complex rim, thought to represent quenched early magma, which coincide with a transition from magmatic to pyroclastic behavior during kimberlite emplacement. “mantle” The region of the interior of the Earth between the core (on its inner surface) and the crust (on its outer). Note: The mantle is

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

more than two thousand miles thick and accounts for more than three-quarters of the volume of the Earth. “matrix” The matrix or groundmass of rock is the finer grained mass of material in which larger grains, crystals or clasts are embedded. The matrix of an igneous rock consists of finer grained, often microscopic, crystals in which larger crystals (phenocrysts) are embedded. “measured mineral resource” That part of a mineral resource for which tonnage, densities, shape, physical characteristics, grade and mineral content can be estimated with a high level of confidence. The data must be of sufficient confidence to allow the application of modifying factors in sufficient detail to support detailed mine planning and final evaluation of the economic viability of the deposit. It is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes. The locations are spaced closely enough to confirm geological and grade continuity “MHSA” Mine Health and Safety Act “micaceous” Consisting of, containing, or resembling mica “microdiamond” Diamonds that are less than 0.5mm in all axial dimensions “Mpa” Megapascal, a metric measurement unit of pressure “Mtpa” million tonnes per annum “MRMR/RMS/DRMS” Mining rock mass rating/rock mass strength/design rock mass strength, a rating system for calculating rock competency for underground tunnels “olivine” An olive-green, silicate mineral rich in magnesium and iron. It is a common rock-forming mineral in the lower part of the crust and the upper mantle “open pit mining” Open pit mining is the process of mining a near surface deposit by means of a surface pit excavated using one or more horizontal benches. The term open pit mining is usually used for metallic or non-metallic deposits and sparingly used for bedded deposits like coal. A quarry is a type of open pit mine used to mine building materials (construction aggregate, riprap, sand and gravel) and dimension stones usually at shallower depths. The term quarry has traditionally been used to mine stones. “ore” An ore is a type of rock that contains sufficient minerals with important elements including metals that can be economically extracted from the rock.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

“pan plant” Heavy mineral separation equipment where crushed ore, when mining kimberlite, or alluvial gravel and soil is mixed with water to create a liquid slurry called “puddle” which has a density in the 1.3 to 1.5 g/cm3 range. The mix is stirred in the pan (up to 16 feet in diameter) by angled rotating “teeth”. The heavier minerals, or “concentrate”, settle to the bottom and are pushed toward an extraction point, while lighter waste remains suspended and overflows out of the centre of the pan as a separate stream of material. The concentrate, representing just a small percentage of the original kimberlite ore or alluvial gravels, is drawn off for final recovery of the diamonds. “pipe” or “diatreme” The carrot shaped volcanic vent that has formed by explosive action and is characteristic of kimberlite eruptions. Diatremes typically cut through non-volcanic basement rocks “phenocryst” A conspicuous, usually large, crystal embedded in porphyritic igneous rock “phlogopite” A brown form of mica consisting of hydrous silicate of potassium and magnesium and aluminum “portal” The structure surrounding the immediate entrance to a mine; the mouth of an adit or tunnel. “prefeasibility study” Preliminary feasibility (prefeasibility) studies are the intermediate step in project evaluation. At this stage there is sufficient drilling, bulk sampling and process test work for preliminary engineering. Typically, the accuracy of these studies is in the +/- 15-25 per cent range. The goal of these studies is to determine the mining and milling extraction methods and rates, the product recoveries, environmental and permitting issues, preliminary capital and operating cost estimates. “probable mineral reserve” The economically mineable part of an Indicated, or in some circumstances, a Measured Mineral Resource. “raise” In underground mining a raise refers to a vertical or inclined excavation that leads from one level, or drift, to another. A raise may also extend to surface. “RAP” Return air pass “ROM” or “Run-of-mine” tonnage of rock as it comes out of a mine before crushing or any other form of treatment. “scrubber” Scrubbers are designed to break up alluvial gravels, clay and sand. They can handle stone washing, feeds with high clay content and difficult ore. Scrubbers process precious metals, base metal ores, minerals, aggregates, gravel and sand. This is done through rotation and the force of particles hitting each

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

other. As the scrubber rotates slowly, the fines are churned along wiith the oversize and water. As a result, the soil matrix is broken and the target material is liberated.

“scoping study” a preliminary study to define the scope of a project – often an internal document not for public disclosure. Now termed a Preliminary Economic Assessment (PEA) under some mining codes. “Search ellipsoid” These parameters determine how far out to search for data to support a particular kriged estimate. “SG” Specific gravity “shaft” A vertical or sloping passageway made in the earth for finding or mining ore and ventilating underground excavations . “shear deformation” The amount of stretch or compression along material line elements or fibers is the normal strain, and the amount of distortion associated with the sliding of plane layers over each other is the shear strain, within a deforming body. “SIMRAC” Safety in Mines Research Advisory Committee, a permanent committee of the Mine Health and Safety Council, was established in terms of the Mine Health and Safety Act (29/1996) to conduct research and surveys regarding, and for the promotion of, health and safety in the South African mining industry. “solid/wireframe” A skeletal three-dimensional model in which only lines and vertices are represented. “spalling” Spall are flakes of a material that are broken off a larger solid body and can be produced by a variety of mechanisms, including as a result of projectile impact, corrosion, weathering, cavitation. Spalling and spallation both describe the process of surface failure in which spall is shed. “stope” The openings made in the process of extracting ore from underground mines are called stopes or rooms. There are two steps involved in stoping. The first is development—that is, preparing the ore blocks for mining—and the second is production, or stoping, itself. “striations” Striations means a series of ridges, furrows or linear marks, and is used in several ways: Glacial striation. Striation (geology), a striation as a result of a geological fault. “surge bin” A compartment for temporary storage in front of crushing units etc., which will allow converting a variable rate of supply into a steady flow of the same average amount.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

“tailings/slimes” Tailings are the byproducts left over from mining and extracting resources, such as extracting bitumen from the oilsands or minerals such as copper or gold from ores. Slimes include finely ground rock particles – ranging from sand-sized to silt-sized, often stored in purpose built dams or reservoirs on the mine site. “talus” Talus slope or deposit, a slope formed by an accumulation of broken rock debris, as at the base of a cliff or other high place, also called scree “TCR” Total core recovery “TPA, TPM, TPD, TPH” tonnes per annum, month, day, hour generally as referred to processing equipment “trenching” The process of locating the position of a geological structure by digging trenches perpendicular to its expected strike “trommel” A trommel screen, also known as rotary screen, is an essential unit which is used mainly in the mineral and solid-waste processing industries. It consists of a perforated cylindrical drum which is normally elevated at an angle at the feed end. Physical size separation is achieved as the feed material spirals down the rotating drum, where the undersized material smaller than the screen apertures passes through the screen, while the oversized material exits at the other end of the drum “turbidite” Sedimentary rock formed from fine-grained to very fine- grained sediments in a marine environment “UCS” Uniaxial compressive strength “unconformity” An unconformity is a contact between two rock units in which the upper unit is usually much younger than the lower unit. Unconformities are typically buried erosional surfaces that can represent a break in the geologic record of hundreds of millions of years or more. “variogram/variography” In spatial statistics, a graph which relates the variance of the difference in value between pairs of samples, to the distance between them. Allows the weighting of a sample value in terms of its distance from the point where an estimate of sample value is required. “waste” Is the waste rock or other material that overlies an ore or mineral body and is displaced during mining without being processed. “winze” A minor connection between different levels in an underground mine. When worked upwards from a lower level it is usually called a raise; when sunk downward from a higher level it may be called a sump. The top of a winze is located underground

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

and it is not equipped with winding gear, in contrast to a shaft, which is a deeper connection between levels and does have winding gear, whether the top of the excavation is located on the surface or underground. “xenolith” An inclusion of a pre-existing rock in an igneous rock, often derived from the country rocks that have been intruded through.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

APPENDIX 2 – DASSAULT GEMCOM BLOCK MODELLING REPORT

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Technical Report MPH Consulting Project Assistance

3DS.COM/GEOVIA 3DS.COM/GEOVIA

Date:

Written by: Hayley Manning, Business Analyst Dassault Systèmes Canada Software Inc. (GEOVIA)

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Contents 1. Introduction ...... 25-4 2. Data Preparation ...... 25-4 3. Geological Model ...... 25-5 4. Exploratory Data Analysis ...... 25-5 5. Compositing ...... 25-6 6. Variography ...... 25-8 7. Block Model ...... 25-15 8. Block Model Estimation Methodology ...... 25-16 9. Block Model Validation ...... 25-18 10. Resource Estimate Reports ...... 25-23

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

1. Introduction Dassault Systèmes Canada Software Inc. (GEOVIA) has produced a geological and resource model in GEMS for internal use for engineering design, in-house consultants, and reporting estimates on reserves. The scope of work included data preparation, data analysis, geological modelling, resource modelling, and documentation.

2. Data Preparation The drillhole data was initially stored in multiple workspaces. These were combined into a single workspace titled DRL1511ALL. The amalgamated workspace contained data for 38 drillholes and 38 bulk samples. The bulk samples were entered into the database in the form of drillholes based on location data provided by MPH employees. All bulk sample Hole-IDs are prefixed by BS.

Four new tables were created and populated in the DRL1511ALL drillhole workspace (Table 2.1.). The assay and composite data were extracted from the Drillhole workspace tables and saved to the Point Area workspace COMP1511 (Table 2.2), which was used to conduct the sensitivity and geostatistical analysis.

Validation was performed on the Drillhole workspace HEADER, SURVEY, ASSAYS, DENSITY and DILUATION tables. Only one error was located in the DILUTION table, where one overlapping interval was identified in drillhole 430-03. Interval with the FROM-TO values of 6.02-7.72 was changed to 6.02-7.52.

Table 2.1: List of new drillhole workspace tables Drillhole Table Description SLDINT Updated using drillhole-solid intersects COMP_ASSY Assay composites within each interval defined in the SLDINT table COMP_DILU Dilution composites within each interval defined in the SLDINT table

Table 2.2: List of new workspaces Workspace Type Description DRL1511ALL Drillhole All current drillhole data compiled into single workspace GeoLin1511 Polyline All line work used to create solids in GeoSolid1511 workspace GeoSolid1511 Triangulation All current geology solids COMP2015 Point Area Data extracted from DRL1511ALL workspace GeoStats1511 GeoStatistics Contains results from statistical analysis Lace2016 Block Model Block Model populated with solids from GeoSolid1511 and COMP1511

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

3. Geological Model The geological model used in creating the resource model was produced by Dassault Systèmes Canada Software Inc. (GEOVIA). The model included nine geological solids representing four geological domains in the kimberlite pipe (Table 3.1).

Table 3.1: List of geology solids used to update the block model. Geological Domain Solid Name 1 Number of Solids Rock Type Block Model Code Kimberlite 4 K4 2 K4 40 Kimberlite 6 K6 1 K6 60 Kimberlite 8 K8 1 K8 80 Country Rock Breccia CRB 5 CRB 100

A geological model was produced using 3D-rings and ties constructed over 25 plan view sections. The top fifteen levels, 230 (1100m) to 510 (820m), were completed with 20 m spacing, while the bottom 10 levels, 540 (790m) to 900 (430m), were completed with 40 m spacing. The rings were digitized based on hand drawn interpretations made by MPH Consulting on level plans. The 3D- rings adhere to drillhole intersects for all levels unless otherwise specified by MPH Consulting. The lines used to construct the updated geology solids were stored in the workspace GeoLin1511. Each 3D-ring and tie line has been tagged with the geological domain for filtering purposes. To avoid gaps and overlaps between adjacent solids in the final geological model, the 3D-Rings were created so there was overlap between each unit. This allowed the solids to be clipped against one another at a later stage. The area of interest was well below the surface therefore no overburden or topographic surface was used to constrain the geological model. The final clipped versions on the solids were stored in the workspace GeoSolid1511 and have the Name 3 field populated with 20160111.

4. Exploratory Data Analysis

The CPHT, Density, and Dilution values of the geological domains defined by solids were analysed. The assay data was partitioned by location within the various domains solids and compared statistically in Tables 4.1, 4.2, and 4.3.

Table 4.1: CPHT grade distribution of assay samples inside domain solids Geological Domain Average CPHT Minimum Maximum # of Samples K4 153.2 0.0 1615.3 86 K6 31.9 0.0 553.2 81 K8 28.2 0.0 173.9 26 CRB 71.8 0.0 1158.3 20

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Table 4.2: Density distribution inside the domain solids Geological Domain Average Density Minimum Maximum # of Samples K4 2.58 2.34 2.83 75 K6 2.60 2.20 3.10 259 K8 2.64 2.43 2.82 36 CRB 2.71 2.18 3.39 64

Table 4.3: Dilution distribution inside the domain solids Geological Domain Average Dilution Minimum Maximum # of Samples K4 29.1 2.0 100.0 255 K6 65.0 2.5 100.0 1038 K8 57.6 2.5 100.0 159 CRB 91.8 38 100.0 239

The number of samples in each domain varied considerably, but the variations in values were considered significant enough to treat each domain separately during the interpolation process (Table 4.1, 4.2, and 4.3) using a hard boundary methodology.

5. Compositing CPHT: No compositing was performed on the CPHT assays because the block model CPHT values were to be calculated based on constant undiluted grade values pre-determined for each diamond barring domain based on previous studies. See section 8 ‘Block Model Estimation Methodology’ for more details. Density: No composting was required on the density values because it was collected as point data and not interval data. All density records from the DENSITY table that were found within the geology solids were extracted directly to the Point Area DENSITY stored in the COMP2015 workspace. Statistics for the raw Density values by domain are outlined in Table 5.1.

Table 5.1: Raw density values by geological domain K4 K6 K8 CRB All Number of Intervals 72 256 33 61 422 Mean 2.58 2.60 2.63 2.71 2.62 Standard Deviation 0.08 0.16 0.08 0.18 0.15 Coefficient of Variation 0.031 0.062 0.031 0.067 0.059 Maximum 2.83 3.10 2.82 3.39 3.39 Upper Quartile 2.62 2.50 2.69 2.82 2.71 Median 2.57 2.58 2.64 2.71 2.60 Lower Quartile 2.54 2.72 2.57 2.61 2.52 Minimum 2.34 2.20 2.43 2.18 2.18

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Dilution: Interval length statistics for dilution values within the geological solids (K4, K6, K8, and CRB) show that the intervals vary in length from 0.70 m to 72.6 m, with a median of 1.0 m, and an upper quartile length of 3.0 m. Based on this information, a minimum composite length of 3.0 m was determined. Table 5.2 outlines a comparison of the raw records to the composited values using lengths of three, four, and five meters. The 3 m composites exhibited a dilution distribution closer to the raw samples, while four and five meter composites introduced some data smoothing. This combined with the fact that >98% of all raw dilution samples were 3 m or greater a composite length of 3.0 m was selected. The composited intervals honoured the boundaries defined by the vein solids. All remnant intervals less than half the composite length (1.5 m) were discarded.

Table 5.2: Dilution percent values of raw data for all domains compared to varying composite lengths Raw Data 3m Comp. 4m Comp. 5m Comp. Number of Intervals 1691 945 713 569 % Samples > Comp. Length - 1.24 1.12 0.89 Mean 62.69 64.43 64.51 64.46 Standard Deviation 24.60 21.70 21.37 21.04 Coefficient of Variation 0.392 0.337 0.331 0.326 Maximum 100.00 100.00 100.00 100.00 Upper Quartile 81.00 78.82 79.15 78.01 Median 63.39 65.87 65.73 65.62 Lower Quartile 48.62 53.82 54.71 55.00 Minimum 2.00 2.52 4.80 4.56

The bulk sampling data was entered into the GEMS project as drillhole data, where a single dilution value was assigned to the entire length of the drillhole. The available bulk sample intervals ranged from 5.91 m to 31.74 m. A sensitively analysis was conducted in order to see if compositing the longer bulk sample intervals along with the drillholes would have a significant impact on the dilution values. Table 5.3 outlines the results based on a 3 m composite length. With only a 0.4% difference in the average Dilution values, the difference was not considered significant enough to treat the bulk samples differently.

Table 5.3: Dilution interval length Raw Dilution Interval Lengths Composite (3m) Dilution Data (%) No Bulk Samples With Bulk Samples No Bulk Samples With Bulk Samples Number of Samples 1677 1691 884 945 Mean 1.64 1.74 64.7 64.43 Standard Deviation 1.95 2.29 22.0 21.70 Coefficient of Variation 1.183 1.312 0.340 0.337 Maximum 72.58 72.58 100.0 100.00 Upper Quartile 3.00 3.00 79.9 78.82 Median 1.00 1.00 66.1 65.87 Lower Quartile 1.00 1.00 53.2 53.82 Minimum 0.70 0.70 2.5 2.52

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

All dilution intervals from the DILUTION table that were found within the geology solids were extracted directly to the Point Area DILUTION stored in the COMP2015 workspace. Table 5.4 compares the mean dilution values for the raw samples to the 3 m composite values for each for the geological domains.

Table 5.4: The mean grade values for the dilution assay samples compared to composited values by domain. Rock Mean DILUTION Difference Code Samples Composites 40 29.1 33.2 14.1% 60 65.0 65.9 1.4% 80 57.6 65.8 14.23 100 91.8 91.2 -0.7%

6. Variography Linear Variography A downhole analysis was conducted to define the magnitude of the nugget effect. A linear semi- variogram was created by analyzing the 945 dilution composites, and 422 density points inside all the geology solids and by geological domain. Figures 6.1 through 6.10 illustrate the down-hole experimental semi-variograms for Density and Dilution, on which the number of sample pairs at lag distances are plotted.

The linear variography for density was conducting using data from all domains combined and all domains separately. The nugget effect using all sample data was determined to be 0.0075, while the nugget effect conducted on each individual domain ranged from 0.0003 to 0.006. Due to the limited number of sample pairs available the nugget effect determined for Density was not considered reliable, and was only used as a guideline for the 3D semi-variogram analysis.

Figure 6.1: Downhole (Linear) Experimental Semi-Variograms for all density data using a lag of 7. A line projected through the first three lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 0.0075.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 6.2: Downhole (Linear) Experimental Semi-Variograms for K4 density data using a lag of 5. A line projected through the first three lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 0.0003.

Figure 6.3: Downhole (Linear) Experimental Semi-Variograms for K6 density using a lag of 5. A line projected through the 2nd and 3rd lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 0.0003.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 6.4: Downhole (Linear) Experimental Semi-Variograms for K8 density a lag of 5. A line projected through the 1st and 2nd lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 0.0003.

Figure 6.5: Downhole (Linear) Experimental Semi-Variograms for CRB density using a lag of 7. A line projected through the 1st and 2nd lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 0.006.

The linear variography for dilution was also conducting using data from all domains combined and all domains separately. The nugget effect using all sample data was determined to be around 2, while the nugget effect conducted on each individual domain ranged from 5 to 12. A lag of 3 was used in all cases.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 6.6: Downhole (Linear) Experimental Semi-Variograms for all dilution data. A line projected through the first 3 lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 2.

Figure 6.7: Downhole (Linear) Experimental Semi-Variograms for K4 dilution data. A line projected through the 2nd and 3rd lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 12.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 6.8: Downhole (Linear) Experimental Semi-Variograms for K6 dilution data. A line projected through the 1st and 2nd lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 5.

Figure 6.9: Downhole (Linear) Experimental Semi-Variograms for K8 dilution data. . A line projected through the 1st and 2nd lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 6.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 6.10: Downhole (Linear) Experimental Semi-Variograms for CRB dilution data. A line projected through the 1st, 2nd, and 3rd lag values to intersect the variogram (gamma) axis indicates a relative nugget effect of around 9.

3D Variography

Initial 3D variography was conducted on the deposit as a whole in order to define the spatial continuity within the dilution and density datasets. 3D experimental semi-variograms were calculated by analyzing the 945 dilution composites and 422 density points inside all of the geological domains, in the plane corresponding to the average dip direction and dip of the deposit (Azimuth 125°, Dip -90°). 3D variography was attempted on each domain separately; however, each unit did not contain enough points to perform robust variography. It was therefore assumed that the spatial continuity was comparable to the deposit as a whole.

Examination of the variogram map in the plane indicated a primary axis aligned horizontally roughly trending roughly NW-SE (Azimuth 140°, Dip 0°) for density. This orientation may not reflect the true direction of maximum continuity as it is likely influenced by the density point spacing which appears to be more continuous in this direction. Figure 6.11 illustrates the experimental semi-variogram for this orientation and the corresponding semi-major axis (Azimuth 230°, Dip 45°), and minor axis (Azimuth 230°, Dip -45°). A Lag distance of 20 m, a 30° angular tolerance and a 20 m bandwidth were used.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 6.11: 3D Experimental Semi-Variograms for Density

Examination of the variogram map in the plane indicated a primary axis aligned horizontal along strike (Azimuth 67.5°, Dip 0°) for Dilution. Figure 6.12 illustrates the experimental semi-variogram for this orientation and the corresponding down dip semi-major axis (Azimuth 247.5°, Dip -90°) and minor axis (Azimuth 337.5°, Dip 0°). A Lag distance of 20 m, a 20° angular tolerance and a 10 m bandwidth were used.

Figure 6.12: 3D Experimental Semi-Variograms for Dilution

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

The results of the variography were used in the interpolation to assign the appropriate weighting to the sample pairs being used to calculate the block model grade when Ordinary Kriging is applied. The total ranges modeled were used to help define the search ellipse dimensions used in the interpolation. The variogram parameters for Density and Dilution are shown in Table 6.1 and the search ellipse orientations are outlined in Table 6.2.

Table 6.1: Variogram Parameters for Density and Dilution. Grade Structure Nugget Sill Major Semi-Major Minor

C0 0.002 - - - Spherical - 0.004 15.5 10.213 9.292 Density 1 - 0.018 58.2 38.348 34.890 Spherical 2 C0 2.000 - - - Spherical - 169.00 21.000 16.835 11.293 Dilution 1 - 301.00 98.000 78.565 52.701 Spherical 2

Table 6.2: Search Ellipse Parameters for Density and Dilution (rotations relative the block model) Rotation Rotation Rotation Range 1 Range 2 Range 3 Confidence Search Size Grade about Z about X about Z (m) (m) (m) Level Density 4.5 90 0 58 38 35 Mediu Dilutio 2 m -68 -45 0 98 78.5 53 n Density 4.5 90 0 175 115 104 Large Dilutio 3 -68 -45 0 196 157 105 n

7. Block Model The block model Lace2016 was created in a multiple folder block model format. A folder was created for each geological domain, including K4, K6, K8, and CRB as well as a default Standard folder as a placeholder for waste material. The block model parameters, including the lateral extents and block size, were determined during the initial scoping meeting between Paul Sobie and Hayley Manning. The block model properties are outlined in the Table 7.1 and the model attributes are outlined in Table 7.2. The waste folder does not hold any grade attributes or special kriging values; however it does contain four additional attributes, ToMine, AveCPHT, TotCarats, and TotTonnage, which contain values which apply to the deposit as a whole. See Table 7.2 for more details.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Table 7.1: Block model properties Origin Rotation Block Size Number of Blocks X – 12172 Columns - 20 Columns - 16 18° Y – -3037066 Rows - 20 Rows - 14 counter clockwise Z – 1107 Levels - 20 Levels - 8

Table 7.2: Block model attributes Name Mapping Date type Default Description Rock Type Rock Type Integer 0 Rock type populated from geology solids Percent Percent Single 0 Percent of block inside geology solids CPHT_NN CPHT_NN Single 0 Interpolated CPHT using Nearest Neighbour CPHT_ID CPHT_ID Single 0 Interpolated CPHT using Inverse Distance CPHT_OK CPHT_OK Single 0 Interpolated CPHT using Ordinary Kriging DENS_NN DENS_NN Single 0 Interpolated Density using Nearest Neighbour DENS_ID DENS_ID Single 0 Interpolated Density using Inverse Distance DENS_OK DENS_OK Single 0 Interpolated Density using Ordinary Kriging DILU_NN DILU_NN Single 0 Interpolated Dilution using Nearest Neighbour DILU_ID DILU_ID Single 0 Interpolated Dilution using Inverse Distance DILU_OK DILU_OK Single 0 Interpolated Dilution using Ordinary Kriging Confidence Generic Single 0 Records search ellipse used during interpolation NumHoles Generic Single 0 Number of drillholes used in grade calculation NumPoints Generic Single 0 Number of points used in grade calculation PointDist Generic Single 0 Ave. distance to points in grade calculation Carats Generic Single 0 Weighted carats for single unit Tonnage Generic Single 0 Weighted tonnage for single unit TotCarats Generic Single 0 Total carats based on all units TotTonnage Generic Single 0 Total tonnage based on all units AveCPHT Generic Single 0 Weighted CPHT grade based on all units ToMine Generic Single 0 Reserve blocks coded with the value 1

The Rock Type and Percent attributes were updated using the ‘Update from solids’ command in GEMS using an integration level of 25 with horizontal along columns needle orientation. The minimum percent entered to re-assign the block was 0.01%. See section 8 ‘Block Model Estimation Methodology’ for details related to updating the block model with CPHT, Density, and Dilution attributes.

8. Block Model Estimation Methodology During the update of the CPHT, Density, and Dilution block model attributes, only samples contained within a vein solid were used to update blocks with the corresponding rock codes. All Density and Dilution values were estimated using Inverse Distance (Anisotropic, power of 2) and with Ordinary Kriging using composites within the respective solids. For both calculation methods a minimum of 2 samples and a maximum of 12 samples were used. A maximum of 3 samples per drillhole was applied. The interpolation was run in two passes. The first used the search ellipse profiles defined by the ranges determined in the 3D-semi-variograhy analysis. However, some blocks did not get

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

interpolated due to an insufficient number of samples in the search volume. A second pass was conducted using a search ellipse profile with larger ranges to populate the rest of the blocks coded as ore. All blocks interpolated in the first past were assigned a Confidence value of 2 and, all blocks interpolated in the second past were assigned a Confidence value of 3. See Table 6.2 to see which search ellipse properties used for Density and Dilution. The following special-model kriging results were saved during the interpolation runs for dilution:  Number of points used for the estimation (NumPoint)  Number of holes used (NumHoles)  Mean distance for samples used (PointDist)  Special profile value (Confidence)

The CPHT attribute was updated using known CPHT grades and matrix dilution values determined from previous analyses on the three diamond bearing domains, K4, K6, and K8. These values were chosen by MPH Consulting. The estimated undiluted CPHT values for each domain are outlined in Table 8.1. The dilution values for K4 and K6 domains were then applied to the constant undiluted grade values to calculate the diluted CPHT grades using the formula below. This was done for each of the estimation methods (NN, ID, and OK). A diluted grade of 16 CPHT was assigned to all K8 blocks. CPHT = Estimated Undiluted CPHT value x (1-(Total Dilution-Matrix Dilution)/100)

Table 8.1: Estimated undiluted CPHT values based on microdiamond analysis Domain Estimate matrix Estimated undiluted dilution CPHT value (%) K4 22.3 40 K6 54.7 10

To determine the weighted CPHT grade value for each block the tonnage and carats were calculated for each domain using block simple manipulation scripts. All block mode scripts used were stored in the GEMS project folder under the BlockData folder.

Tonnage and carat values were calculated for each domain using the expression below. The density and CPHT attributes updated using the Inverse Distance interpolation method were used. A default density of 2.6 was used for waste material. Tonnage = Block Volume * Density * (Percent/100) Carats = (CPHT/100)*Tonnage Total tonnage and total carats were calculated by adding the values from each of the K4, K6, K8, CRB, and Standard (Waste) block model folders. The results are stored in the block model attributes TotCarats and TotTonnage in the default folder. The total tonnage and total carat values were then used to compute the weighted average CPHT per block using the expression below. Weighted Average CPHT = Total Carats/(Total Tonnage/100)

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

9. Block Model Validation Two validation exercises were completed on the resource model:  Visual comparison of block and composite grades on sections and plans  Comparison of composites and bulk samples to block model values

The visual comparison indicates that in general the estimated values correspond reasonably well with the bulk sample and composite values. Cross section, plan views and 3D inspection for CPHT, Density, and Dilution were checked. In most all cases the model behaved as expected showing the same trends seen on the composites. Figure 9.1 shows the resulting CPHT values compared to the bulk sample results. Figures 9.2 and 9.3 illustrate level plans views with the composite points used in the interpolation for comparison with the Density and Dilution block model values.

A comparison of the block model results for both Inverse Distance (ID) and Ordinary Kriging (OK) were compared to declustered composite grades (nearest neighbour model – NN). Overall (for all domains) the estimated mean Density value was 0.4% lower (for ID2 method) and 0.8% lower (for OK method) than the sample mean of 2.62. The estimated mean Dilution grade was 1.1% lower (for ID2 method) and 1.2% lower (for OK method) than the composite mean of 64.4% (Table 9.1). Comparisons for each domain are tabulated in Table 9.2 and Table 9.3.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 9.1: Visual comparison of block model weighted CPHT values (ID2) against the bulk sample results for levels 290 and 310.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 9.2: An example of visual inspection on level plans 290 and 310 showing density blocks and raw data comparison for K4, K6, and K8.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Figure 9.3: An example of visual inspection on level plans 290 and 310 showing dilution blocks and composite comparison for K4, K6, and K8.

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Table 9.1: Comparison of block model Density and Dilution values to sample values for all geological domains Sample Mean Mean Mean % Dif. % Dif. % Dif. Mean NN ID OK Density 2.62 2.59 1.1 2.61 0.4 2.60 0.8 Dilution 64.4 62.5 2.9 63.7 1.1 63.6 1.2

Table 9.2: Comparison of Density values for all geological domains. K4 K6 Density Comp Block NN Block ID Block OK Comp Block NN Block ID Block OK No. Samples 72 297 297 297 256 508 508 508 Minimum 2.34 2.34 2.46 2.48 2.20 2.20 2.34 2.35 Maximum 2.83 2.83 2.78 2.67 3.10 3.10 2.91 2.91 Mean 2.58 2.57 2.59 2.58 2.60 2.55 2.57 2.57 Median 2.57 2.57 2.59 2.58 2.58 2.54 2.56 2.56 Variance 0.007 0.006 0.001 0.001 0.026 0.019 0.007 0.006 SD 0.082 0.083 0.035 0.034 0.163 0.136 0.081 0.077 CV 0.032 0.032 0.013 0.013 0.062 0.054 0.032 0.030 K8 CRB Density Comp Block NN Block ID Block OK Comp Block NN Block ID Block OK No. Samples 33 95 95 95 61 324 324 324 Minimum 2.43 2.43 2.45 2.55 2.18 2.18 2.35 2.45 Maximum 2.82 2.80 2.74 2.74 3.39 3.07 2.89 2.86 Mean 2.63 2.62 2.63 2.64 2.71 2.69 2.68 2.66 Median 2.64 2.63 2.63 2.65 2.71 2.68 2.71 2.66 Variance 0.007 0.006 0.002 0.001 0.03 0.02 0.01 0.005 SD 0.08 0.08 0.04 0.04 0.18 0.15 0.10 0.07 CV 0.031 0.030 0.017 0.014 0.067 0.054 0.037 0.026

Table 9.3: Comparison of Dilution values for all geological domains. K4 K6 Dilution Comp Block NN Block ID Block OK Comp Block NN Block ID Block OK No. Samples 139 297 297 297 599 508 508 508 Minimum 2.5 2.5 10.9 16.9 12.8 12.8 30.3 35.7 Maximum 95.0 72.9 66.2 64.6 100.0 100.0 100.0 100.0 Mean 33.2 30.0 32.0 32.2 65.9 64.9 64.0 64.3 Median 26.0 26.9 29.5 30.3 65.4 66.8 65.2 65.1 Variance 414.5 195.5 86.7 61.7 181.7 315.9 96.1 83.0 SD 20.36 13.98 9.31 7.85 13.48 17.77 9.80 9.11 CV 0.613 0.466 0.291 0.244 0.205 0.274 0.153 0.142 K8 CRB Dilution Comp Block NN Block ID Block OK Comp Block NN Block ID Block OK No. Samples 82 95 95 95 125 324 324 324 Minimum 22.5 32.1 43.4 42.5 48.0 48.0 68.6 69.8

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Maximum 98.7 98.7 92.0 86.7 100.0 100.0 99.7 98.0 Mean 65.8 63.0 63.2 61.2 91.4 88.5 92.2 92.0 Median 67.9 56.8 60. 59.4 95.0 94.1 94.3 94.0 Variance 449.4 324.2 173.2 119.4 112.0 192.7 36.8 26.6 SD 21.2 18.01 13.16 10.93 10.58 13.88 6.06 5.16 CV 0.322 0.286 0.208 0.178 0.116 0.157 0.066 0.056

10. Resource Estimate Reports The volumetrics profile BMVOLS was created to report resources on the new block model Lace2016. See Table 10.1 for volumetrics report by domain.

Table 10.1: Resource report sorted by vein and grade group. Volume Density_ID Tonnage DILU_ID DILU_OK Domain CPHT_ID CPHT_OK (m3x1000) (T/m3) (T) (%) (%) K4 1,065.486 2.585 2,754,428.434 36.725 36.450 30.486 31.176 K6 1,834.957 2.563 4,702,111.239 8.984 8.983 64.864 64.874 K8 144.722 2.642 382,351.218 16.000 16.000 66.429 64.261 CRB 723.803 2.687 1,945,025.602 0.000 0.000 93.292 89.261 Total 3,768.968 2.596 9,783,916.493 15.282 15.204 60.898 60.211

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

APPENDIX 3 – MICRODIAMOND DATABASE

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Sample Weight Mine Hole id Lithology Stones Mining Block 2360 1700 1180 850 600 425 300 212 150 106 75 id kg Level K4 250-K4 250-K4 43.70 77 253 UK4 Mine 0 0 1 1 5 11 13 7 10 12 17 260-K4 260-K4 41.10 K4 68 245 Upper Block 0 1 1 1 4 10 12 5 8 9 17 BH-29 L - 1921 14.08 K4 54 261 0 0 0 2 1 1 4 9 9 12 16 335-05 B11 14.94 K2 38 262 0 0 0 0 0 4 8 4 4 8 10 BH-29 L - 1922 14.46 K4 21 265 0 0 0 0 1 1 6 9 2 1 1 335-05 B12 16.44 K2 15 269 0 0 0 0 1 1 2 3 2 3 3 BH-29 L - 1923 FAIL QC 269 BH-29 L-1802 20.92 K4 139 269 0 0 0 1 3 2 2 10 18 48 55 335-08 B1 14.43 K4 35 270 0 0 0 0 0 0 5 8 7 8 7 335-07 B5 14.06 K4 18 272 0 0 1 1 3 2 2 3 2 2 2 (above 290m BH-29 K2/K4 L - 1924 18.44 49 273 level) 0 0 0 1 1 3 1 5 9 13 16 BH-29 L - 1925 13.68 K2/K4 53 277 0 0 1 0 2 6 0 4 6 16 18 BS 290-08 U/G East 40.56 K4 133 283 0 0 0 0 3 6 13 9 23 37 42 BS 290-09 U/G West 40.56 K4 248 283 0 0 1 6 8 11 9 29 34 70 80 BS-290-10 U/G L-1818 19.00 K4 48 280 0 0 0 1 0 1 5 7 12 15 7 BS-290-11 U/G L-1819 23.42 K4 95 280 0 0 1 1 4 6 9 10 25 22 17 430-06 430-6-1 26.34 K4 22 286 0 0 0 0 0 1 3 6 4 4 4 U/G (290-06) L - 1939 22.38 K4 150 283 0 0 2 6 8 9 6 12 15 37 55 BH-30 L - 1916 14.04 K4 34 290 0 0 0 0 0 2 3 3 5 7 14 CROWN-7 2784 25.52 K4 41 293 0 0 0 0 3 7 6 15 9 1 0 BH-30 L - 1918 FAIL QC 298 BH-30 L-1801 2.96 K2 31 298 (below 290m 0 0 0 0 0 0 2 1 5 11 12 430-05 B10 11.04 K4 20 299 level) 0 0 0 2 0 0 4 2 4 4 4 430-03 B7 18.04 K4 7 304 0 0 0 1 3 0 2 1 0 0 0 CROWN-7 2785 24.60 K4 20 306 0 0 0 1 2 3 4 9 1 0 0 27 Samples 494.71 1416 0 1 8 25 52 87 121 171 214 340 397 CROWN-7 2786 22.48 K4/K2 37 316 0 1 0 0 1 3 9 15 8 0 0 430-05 430-5-1 17.38 K4 149 319 0 0 0 2 3 4 8 5 17 33 77 BH-34 L - 1908 20.62 K4 150 319 0 0 3 11 16 10 18 16 25 28 23 BH-33 L - 1901 14.10 K2 79 326 0 0 2 5 3 2 5 10 13 17 22 BH-34 L - 1909 19.78 K4 78 335 0 0 0 1 0 2 4 8 17 24 22 UK4 Mine K2/K4 CROWN-7 2788 24.90 17 337 Low er Block 0 0 0 0 1 1 4 4 6 1 0 BH-33 L - 1903 21.54 K4 119 341 0 0 2 2 2 7 3 15 23 38 27 CROWN-7 2789 24.96 K4 26 349 0 0 0 0 1 0 9 13 3 0 0 BH-33 L - 1904 20.28 K4 143 353 0 0 0 1 5 4 6 11 19 41 56 CROWN-7 2790 24.85 K4/K2 16 364 0 0 0 1 2 2 5 3 3 0 0 BH-33 L - 1905 20.22 K4/6/2/8 56 365 0 0 0 0 6 3 2 2 10 13 20 11 Samples 231.11 870.00 0 1 7 23 40 38 73 102 144 195 247

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Sample Weight Mine Hole id Lithology Stones Mining Block 2360 1700 1180 850 600 425 300 212 150 106 75 id kg Level CROWN-7 2791 23.84 K4 35 378 0 1 0 1 7 8 5 7 5 1 0 CROWN-7 2792 23.05 K4 23 388 0 0 0 1 1 1 8 12 0 0 0 CROWN-7 2793 24.58 K4 41 403 0 0 0 0 0 5 9 13 13 1 0 CROWN-7 2794 25.78 K4 60 415 0 0 0 1 4 3 10 30 12 0 0 CROWN-7 2795 22.76 K4 60 427 0 1 0 1 3 7 20 27 1 0 0 CROWN-3 2721 17.90 K4 0 427 0 0 0 0 0 0 0 0 0 0 0 Block Cave 1 CROWN-3 2722 23.92 K4 33 445 0 0 1 0 1 8 5 13 5 0 0 CROWN-7 2796 23.96 K4 35 446 0 0 0 0 4 4 10 11 6 0 0 CROWN-7 2797 25.00 K2 17 458 0 1 0 0 1 4 2 7 2 0 0 CROWN-7 2798 24.68 K4 33 469 0 0 0 0 0 3 7 16 7 0 0 CROWN-8 2810 7.90 K2/K4/K6 29 474 0 0 0 0 0 5 8 10 6 0 0 CROWN-7 2799 26.69 K4 48 482 0 0 0 0 2 5 17 19 5 0 0 12 Samples 270.06 414 0 3 1 4 23 53 101 165 62 2 0

CROWN-8 2812 28.19 K4 87 515 0 0 0 0 1 2 12 38 32 2 0 CROWN-6 2769 22.68 K2 43 540 0 0 0 1 0 1 6 10 18 7 0 CROWN-3 2726 20.90 K2 8 548 0 0 0 0 1 1 1 1 4 0 0 CROWN-6 2770 18.54 K2/K4 13 549 0 0 1 0 1 1 3 3 2 2 0 CROWN-1 2703 25.14 K4 26 567 0 0 0 0 1 9 3 12 1 0 0 CROWN-6 2774 22.37 K2 18 595 0 0 0 0 0 0 2 8 8 0 0 CROWN-6 2775 20.78 K2 16 605 0 0 0 0 0 1 4 10 1 0 0 CROWN-6 2776 21.34 K2 22 614 0 1 0 1 1 3 3 10 3 0 0 CROWN-6 2777 22.97 K2 29 624 0 0 0 0 1 2 11 10 5 0 0 K4 CROWN-1 2708 18.98 34 632 Block Cave 2 0 0 0 1 0 1 6 15 11 0 0 CROWN-1 2709 19.56 K4 22 640 0 0 0 1 1 2 6 7 5 0 0 CROWN-1 2710 21.06 K4 67 649 0 1 0 1 4 9 13 20 17 2 0 CROWN-1 2711 19.82 K4 30 658 0 0 0 0 2 1 7 9 11 0 0 CROWN-3 2733 24.05 K4 34 662 0 0 1 0 4 5 13 7 4 0 0 CROWN-1 2712 17.77 K4/K6 15 667 0 0 0 0 2 2 2 3 6 0 0 CROWN-3 2747 22.96 K4/K2 31 675 0 0 0 0 0 1 10 11 9 0 0 CROWN-6 2780 19.85 K4 41 679 0 0 0 1 2 2 12 19 4 1 0 CROWN-1 2713 27.13 K4/K6 15 682 0 0 0 1 0 1 1 6 6 0 0 CROWN-3 2748 25.00 K2 35 691 0 0 1 0 0 2 2 18 12 0 0 CROWN-6 2781 25.39 K2 37 697 0 0 0 0 1 3 7 22 4 0 0 20 Samples 444.48 623 0 2 3 7 22 49 124 239 163 14 0

CROWN-1 2714 25.93 K4 41 701 0 0 0 0 0 3 8 20 9 1 0

CROWN-6 2782 26.03 K4 23 717 00001397300 CROWN-1 2715 26.15 K4 50 721 0 0 0 0 1 2 16 24 7 0 0 CROWN-6 2804 23.41 K4 8 732 00000035000 Block Cave 3 CROWN-1 2716 25.14 K4 50 742 0 0 0 2 0 3 9 21 15 0 0 CROWN-3 2754 24.83 K4/K2 28 758 0 0 0 0 0 5 3 13 7 0 0 CROWN-1 2717 25.62 K4 17 760 00001256300 CROWN-1 2718 26.11 K4/K2 43 778 0 0 0 0 3 4 10 19 7 0 0 CROWN-3 2755 28.40 K2/K4 43 791 0 1 1 0 4 5 11 14 7 0 0 9 Samples 231.62 303 0 1 1 2 10 27 74 129 58 1 0

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Sample Weight Mine Hole id Lithology Stones Mining Block 2360 1700 1180 850 600 425 300 212 150 106 75 id kg Level

BH-31 L - 1913 12.60 CRBNorth 33 319 0 0 2 8 6 1 2 3 4 3 4 UK4 Mine BH-34 L - 1912 12.56 CRBNorth 2 362 0 0 0 0 0 0 0 0 2 0 0

BH-33 L - 1907 11.66 Shale/CRBNorth 4 385 0 0 0 0 0 0 0 0 0 1 3

3 Samples 36.82 39 Block Cave 1 0 0 2 8 6 1 2 3 6 4 7

BH-29 L - 1920 13.88 LRKB / K8 3 256 UK4 Mine Upper Block0 0 0 0 0 0 0 1 0 1 1 335-08 B2 15.58 K8 26 277 0 0 1 0 1 1 7 4 4 4 4 (above 290m) 430-07 B14 15.14 K8 2 284 0 0 0 0 0 0 1 1 0 0 0 430-07 B15 14.54 K8 8 293 0 0 1 0 0 1 2 2 2 0 0 (below 290m) 430-07 430-7-2 20.60 K8 60 302 0 0 0 0 0 1 2 7 14 16 20 BH-33 L - 1902 13.58 K8 4 333 0 0 0 1 0 0 0 0 1 1 1 UK4 Mine K8 BH-34 L - 1910 9.68 3 341 Low er Block 0 0 0 0 0 0 0 0 0 3 0 BH-33 L - 1906 23.86 K8 16 375 0 0 1 1 0 1 0 2 1 3 7 8 Samples 126.86 122 Block Cave 1 0 0 3 2 1 4 12 17 22 28 33

BH-30 L - 1914 15.30 K6South 5 281 0 0 0 0 0 0 0 0 2 0 3 BH-30 L - 1915 14.90 K6South 8 285 0 0 0 0 1 0 0 0 3 3 1 U/G (290-01) L - 1929 19.94 K6South 18 283 UK4 Mine 0 0 0 0 0 2 1 2 3 4 6 U/G (290-02) L - 1931 21.18 K6South 20 283 Upper Block 0 0 0 1 0 0 0 1 5 6 7 (Above 290m) U/G (290-03) L - 1933 21.06 K6South 31 283 0 0 0 0 0 0 0 4 7 9 11 U/G (290-04) L - 1935 21.94 K6South 9 283 0 0 0 0 0 1 1 0 0 3 4 U/G (290-05) L - 1937 22.46 K6South 12 283 0 0 0 0 0 0 0 2 3 4 3 136.78 103 0 0 0 1 1 3 2 9 23 29 35

BH-30 L - 1917 15.74 K6South 22 295 0 0 0 1 1 3 1 4 4 4 4 CROWN-8 2805 25.60 K6South 0 295 0 0 0 0 0 0 0 0 0 0 0 335-07 B6 15.16 K6South 4 298 0 0 0 0 0 0 0 3 1 0 0 CROWN-8 2806 25.00 K6South 7 307 0 0 0 0 0 1 1 3 2 0 0 BS-310-Cubby1 U/GL-1807 22.78 K6South 14 307 (below 290m) 0 0 0 0 1 1 0 2 6 2 2 BS-310-02 U/G L-1805 20.68 K6South 14 307 0 0 0 0 0 1 1 1 3 2 6 BS-310 Cubby2 U/GL-1809 24.20 K6South 31 307 0 0 0 0 1 0 5 1 6 7 11 BS-310-03 U/G L-1811 27.00 K6South 26 307 0 0 0 1 1 1 2 5 6 5 5 BS-310 Cubby3 U/GL-1810 22.86 K6South 16 307 0 0 0 0 1 0 0 2 3 4 6 16 Samples 199.02 134 0 0 0 2 5 7 10 21 31 24 34 335-07 335-7-1 16.81 K6South 11 312 00001010135 UK 4 Mine CROWN-8 2807 21.07 K6South 4 317 00000120100 Low er Block CROWN-8 2808 21.26 K6South 10 328 (check 370) 00001026100 430-06 430-6-2 21.90 K6South 91 370 0 0 1 5 12 9 8 9 13 16 18 4 Samples 81.04 116 0 0 1 5 14 10 13 15 16 19 23 CROWN-8 2809 27.11 K6South/K4 17 462 00001413620 Block Cave 1 CROWN-8 2811 20.61 K6South/K4 4 495 00001020100 2 Samples 47.72 21 00002433720 CROWN-8 2813 22.64 K6South/K4 29 538 1 1 0 1 1 1 3 7 12 2 0 CROWN-1 2701 17.85 K6South 5 557 00000011210 CROWN-8 2814 26.01 K6South/K4 25 578 0 0 0 2 2 1 3 11 5 1 0 CROWN-8 2815 24.60 K6South 18 624 Block Cave 2 0 0 0 0 0 1 2 2 11 2 0 CROWN-1 2707 15.76 K6South 3 624 00000101100 CROWN-8 2816 27.35 K6South 13 643 00001152310 CROWN-8 2817 26.31 K6South 10 681 00000042400

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA

Sample Weight Mine Hole id Lithology Stones Mining Block 2360 1700 1180 850 600 425 300 212 150 106 75 id kg Level 7 Samples 160.52 103 1 1 0 3 4 5 18 26 38 7 0 CROWN-8 2818 26.88 K6South 7 713 0 0 0 0 0 0 1 3 2 0 1 CROWN-8 2819 24.75 K6South 4 739 0 0 0 0 0 0 3 1 0 0 0 CROWN-8 2820 25.34 K6South 8 760 0 0 0 1 0 0 0 2 4 1 0 Block Cave 3 CROWN-8 2821 24.04 K6South 9 778 0 0 0 0 0 2 2 1 4 0 0 CROWN-8 2822 25.18 K6South 16 797 0 0 0 0 0 0 4 2 4 3 3 CROWN-3 2756 17.70 K6South 1 859 0 0 0 0 0 0 1 0 0 0 0 6 Samples 143.89 45 0 0 0 1 0 2 11 9 14 4 4 335-07 B3 14.92 K6North 1 255 0 0 0 0 1 0 0 0 0 0 0 335-07 B4 13.76 K6North 5 265 0 0 0 0 0 0 0 1 0 1 3 290-03 B9 15.64 K6North 7 269 0 0 0 0 0 1 1 1 2 1 1 430-07 B13 15.16 K6North 2 276 UK4 Mine 0 0 0 0 0 0 1 1 0 0 0 BH-29 L - 1926 13.74 K6North 9 281 Upper Block 0 0 0 0 0 0 1 1 3 2 2 CROWN-7 2783 22.64 K6North 4 282 0 0 0 0 0 0 1 2 1 0 0 BH-29 L - 1927 14.70 K6North 29 285 0 0 0 0 0 1 0 2 6 9 11 BH-30 L - 1919 15.22 K6North 2 304 0 0 0 0 0 0 1 1 0 0 0 8 Samples 125.78 59 0 0 0 0 1 2 5 9 12 13 17

CROWN-7 2787 25.07 K6North 2 326 0 0 0 0 0 0 1 0 1 0 0 430-03 430-3-1 21.78 K6North 111 330 UK4 Mine 0 0 0 0 2 2 4 5 19 34 45 Low er Block BH-34 L - 1911 17.84 K6North 23 347 0 0 0 1 1 0 1 1 6 5 8 430-03 B8 16.04 K6North 8 362 0 0 1 0 0 1 3 0 1 1 1 4 Samples 80.73 144 0 0 1 1 3 3 9 6 27 40 54 CROWN-3 2723 26.61 K6North 10 477 0 0 0 0 0 0 1 4 5 0 0 CROWN-6 2764 21.89 K6North 44 481 0 0 0 0 0 0 4 12 15 11 2 CROWN-3 2724 22.30 K6North 8 489 0 0 0 0 0 1 2 4 1 0 0 CROWN-6 2765 22.67 K6North 6 490 0 0 0 0 0 0 2 1 2 0 1 CROWN-7 2800 25.73 K6North 4 498 Block Cave 1 0 0 0 0 0 2 1 1 0 0 0 CROWN-6 2766 22.19 K6North 6 499 0 0 0 0 1 1 2 1 1 0 0 CROWN-3 2725 20.42 K6North 12 504 0 0 0 0 0 0 4 3 4 1 0 CROWN-7 2801 24.02 K6North 4 508 0 0 0 0 0 0 0 3 1 0 0 CROWN-6 2767 21.70 K6North 0 510 0 0 0 0 0 0 0 0 0 0 0 9 Samples 207.53 94 0 0 0 0 1 4 16 29 29 12 3 CROWN-7 2802 24.02 K6North 17 520 0 0 0 0 0 1 4 2 10 0 0 CROWN-6 2768 21.76 K6North 22 527 0 0 0 0 0 0 1 9 9 3 0 CROWN-7 2803 25.74 K6North 3 536 0 0 0 0 0 1 1 1 0 0 0 CROWN-3 2727 19.30 K6North 3 557 0 0 0 0 0 0 1 2 0 0 0 CROWN-6 2771 22.54 K6North 3 560 0 0 0 0 0 1 0 1 0 1 0 CROWN-3 2728 22.10 K6North 10 573 0 0 0 1 1 1 5 1 1 0 0 Block Cave 2 CROWN-3 2729 22.79 K6North 4 592 0 0 0 0 0 1 1 2 0 0 0 CROWN-3 2730 24.63 K6North 5 605 0 0 0 0 1 0 1 3 0 0 0 CROWN-3 2731 19.76 K6North 0 617 0 0 0 0 0 0 0 0 0 0 0 CROWN-6 2778 19.03 K6North 20 647 0 0 0 2 1 4 1 11 1 0 0 CROWN-3 2732 19.07 K6North 6 653 0 0 0 0 0 1 0 3 1 1 0 CROWN-6 2779 23.52 K6North 37 662 0 0 0 0 1 5 10 17 4 0 0 12 Samples 264.26 130 0 0 0 3 4 15 25 52 26 5 0 3282.93 4736

2016 Technical Report LACE DIAMOND MINE, REPUBLIC OF SOUTH AFRICA