An Independent Technical Report on the Maseve Project

(WBJV Project areas 1 and 1A) located on the Western Limb

of the Bushveld Igneous Complex, South Africa

for

PLATINUM GROUP METALS (RSA) (PTY) LTD

REPUBLIC OF SOUTH AFRICA REGISTERED COMPANY

REGISTRATION NUMBER: 2000/025984/07

A WHOLLY-OWNED SUBSIDIARY OF

PLATINUM GROUP METALS LTD.

TSX – PTM; MKT - PLG

Report Date: 28 August 2015

Effective Date of Mineral Resources and Reserves: 15 July 2015

NI43-101 Technical Report

IMPORTANT NOTICE

This report includes results for Mineral Resources. The report communicates the updated Mineral Resource and Reserve estimate for the Project Areas 1 and 1A of the WBJV Project 1 Mine (Maseve). The reader is warned that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

No Inferred Mineral Resources are reported in this Estimate.

Cautionary Note to U.S. Investors
Estimates of mineralization and other technical information included or referenced in this technical report have been prepared in accordance with NI 43-101. The definitions of proven and probable reserves used in NI 43-101 differ from the definitions in SEC Industry Guide 7. Under SEC Industry Guide 7 standards, a "final" or "bankable" feasibility study is required to report reserves, the three-year historical average price is used in any reserve or cash flow analysis to designate reserves and the primary environmental analysis or report must be filed with the appropriate governmental authority. As a result, the reserves reported by the Company in accordance with NI 43-101 may not qualify as "reserves" under SEC standards. In addition, the terms "mineral resource", "measured mineral resource", "indicated mineral resource" and "inferred mineral resource" are defined in and required to be disclosed by NI 43-101; however, these terms are not defined terms under SEC Industry Guide 7 and normally are not permitted to be used in reports and registration statements filed with the SEC. Mineral resources that are not mineral reserves do not have demonstrated economic viability. Investors are cautioned not to assume that any part or all of the mineral deposits in these categories will ever be converted into reserves. "Inferred mineral resources" have a great amount of uncertainty as to their existence, and great uncertainty as to their economic and legal feasibility. It cannot be assumed that all or any part of an inferred mineral resource will ever be upgraded to a higher category. Under Canadian securities laws, estimates of inferred mineral resources may not form the basis of feasibility or pre-feasibility studies, except in rare cases. Additionally, disclosure of "contained ounces" in a resource is permitted disclosure under Canadian securities laws; however, the SEC normally only permits issuers to report mineralization that does not constitute "reserves" by SEC standards as in place tonnage and grade without reference to unit measurements. Accordingly, information contained or referenced in this technical report containing descriptions of the Company's mineral deposits may not be comparable to similar information made public by U.S. companies subject to the reporting and disclosure requirements of United States federal securities laws and the rules and regulations thereunder.

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QUALIFIED PERSONS
Independent geological qualified person:
	Mr Charles J Muller (B.Sc. (Hons) Geology) Pri. Sci. Nat.
	CJM Consulting (Pty) Ltd
	Mineral Resource Consultants
	Ruimsig Office Estate
	199 Hole-in-one Road
	Ruimsig
	Roodepoort
	1724
	Republic of South Africa
	Mobile: +27 83 230 8332
	Phone:  +27 11 958 2899
	Fax:       +27 11 958 2105
	E-mail: charles@cjmconsult.com

Independent engineering qualified person:
	Mr G. Roets (B.Eng Mining), Pr. Eng (ECSA), Professional association to AMMSA
	DRA Minerals Park
	3 Inyanga Close
	Sunninghill
	2157
	P O Box 3567
	Rivonia, 2128
	Gauteng
	Republic of South Africa
	Mobile: +27 72 472 7545
	Phone: +27 11 202 8686
	E-mail: Gert.Roets@draglobal.com

Independent engineering qualified person:
	Mr Gordon I. Cunningham, B. Eng. (Chemical), Pr. Eng.(ECSA),Professional association to
	FSAIMM
	Turnberry Projects (Pty) Ltd.
	PO Box 2199
	Rivonia, Sandton 2128
	Republic of South Africa
	Mobile: +27 83 263 9438
	Phone: +27 11 726 1590
	E-mail: turnberry@iafrica.com

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OPERATING COMPANIES
Local operating company:
	Platinum Group Metals (RSA) (Pty) Ltd
	1st Floor, Platinum House
	24 Sturdee Avenue
	Rosebank
	Johannesburg 2196
	Republic of South Africa
	Phone: +27.11.782.2186
	Fax: +27.11.447.1000
	E-mail: info@platinumgroupmetals.net

Parent and Canadian-resident Company:
	Platinum Group Metals Limited
	Suite 788-550 Burrard Street
	Vancouver, BC
	Canada V6C 2B5
	Phone:1.866.899.5450
	e-mail: info@platinumgroupmetals.net
	Website: www.platinumgroupmetals.net

For technical reports and news releases filed with SEDAR, see www.sedar.com

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TABLE OF CONTENTS

1.0	SUMMARY		13
	1.1	Introduction	13
	1.2	Ownership	13
	1.3	Geology	14
	1.4	Mineralisation	14
	1.5	Project Status	15
	1.6	Resources	15
	1.7	Mineral Reserves	16
	1.8	Mining Operations	21
	1.9	Metallurgical Testwork and Recovery	27
	1.10	Process Plant Design	28
	1.11	Infrastructure	29
	1.12	Environmental Studies, Permitting and Social or Community Impact	31
	1.13	Production schedule	32
	1.14	Operating Costs	35
	1.15	Capital Costs	35
	1.16	Economic Analysis	36
	1.17	Conclusion and Recommendations	38
2.0	INTRODUCTION		39
	2.1	Issuer	39
	2.2	Terms of Reference and Purpose of the Report	39
	2.3	Sources of Information	39
	2.4	Involvement of the Qualified Person: Personal Inspection	39
	2.5	Frequently used Acronyms, Abbreviations, Definitions and Units of Measure	39
	2.6	Specific Areas of Responsibility	43
3.0	RELIANCE ON OTHER EXPERTS		45
4.0	PROPERTY DESCRIPTION AND LOCATION		46
	4.1	Sufficiency of Surface Rights	46
	4.2	Extent of the Project	46
	4.3	Location of the Maseve Mine (Maseve) Project	46
	4.4	Status of Surface Rights and Mineral Title	49
	4.5	Royalties, Payments and other Agreements	49
	4.6	Environmental Liabilities	50
	4.7	Permitting Status	51
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		52
	5.1	Accessibility	52
	5.2	Climate	52
	5.3	Local Resources	54
	5.4	Regional Infrastructure	55
	5.5	Physiography	56
6.0	HISTORY		62

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	6.1	Prior Ownership	62
	6.2	Work Done by Previous Owners	62
	6.3	Historical Mineral Reserves and Resources	62
	6.4	Production from the Property	64
7.0	GEOLOGICAL SETTING AND MINERALISATION		65
	7.1	Regional Geology of the BIC	65
	7.2	Local Geology – Western Bushveld Limb	69
	7.3	Local Structure	74
	7.4	Property Geology	76
	7.5	Mineralisation	90
8.0	DEPOSIT TYPES		91
	8.1	Geological Modelling	91
9.0	EXPLORATION		97
	9.1	Survey (field observation) results, procedures and parameters	97
	9.2	Interpretation of survey (field observation) results	100
10.0	DRILLING		101
	10.1	Type and Extent of Drilling	101
	10.2	Procedures, Summary and Interpretation of Results	101
	10.3	Additional Information for Projects other than Advanced Projects	103
11.0	SAMPLE PREPARATION, ANALYSIS AND SECURITY		104
	11.1	Sample Preparation and Quality Control Prior to Dispatch	104
	11.2	Sample Preparation and Analytical Procedures Employed by Laboratory	107
	11.3	Quality Assurance and Quality Control Results	108
	11.4	Adequacy of Sample Preparation, Security and Analytical Procedures	114
12.0	DATA VERIFICATION		115
	12.1	Verification of Data by QP	115
	12.2	Nature of the Limitations of Data Verification Process	115
	12.3	Possible Reasons for not having Completed a Data Verification Process	115
13.0	MINERAL PROCESSING AND METALLURGICAL TESTING		116
	13.1	Historical Metallurgical Testwork Summary	116
	13.2	New Metallurgical Testwork	120
	13.3	Merensky Recovery Prediction and Benchmarking	125
	13.4	UG2 Recovery Prediction and Benchmarking	127
14.0	MINERAL RESOURCE ESTIMATES		129
	14.1	Key Assumptions and Parameters	130
	14.2	Data Analysis	134
	14.3	Data Verification	138
	14.4	Quality Assurance and Quality Control Data	138
	14.5	Population Statistics	138
	14.6	Mining Considerations	145
	14.7	Reef Compositing Definitions	145
	14.8	Mineral Resource Modelling	146
	14.9	Grade Estimation	150
	14.10	Mineral Resource Classification	151

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	14.11	Model Results	152
	14.12	Compliance with Disclosure Requirements for Mineral Resources	162
	14.13	Effect of Modifying Factors	163
15.0	MINERAL RESERVE ESTIMATES		164
	15.1	Resource to Reserve Calculation	164
	15.2	Basic Grade Equation	172
	15.3	Mineral Reserve Statement	175
	15.4	Material Changes to Mineral Reserves	177
16.0	MINING METHODS		179
	16.1	Introduction	179
	16.2	Geological and Geotechnical Features Affecting on Reef Mining	183
	16.3	Mining Method Selection	183
	16.4	Rock Engineering	184
	16.5	Mine Design Features	186
	16.6	Mine Ventilation and Cooling	206
	16.7	Basic Mining Equation	207
17.0	RECOVERY METHODS		246
	17.1	Process Design	246
	17.2	Process Description Summary	249
	17.3	Control System	250
	17.4	Recoverability	250
	17.5	Concentrator Status	250
	17.6	Plant Operations	251
18.0	PROJECT INFRASTRUCTURE		252
	18.1	Site Layout and Access Roads	252
	18.2	Underground Trackless Mobile Machinery and Logistics	256
	18.3	Power and Energy Infrastructure	262
	18.4	Fuel Storage	264
	18.5	Water Supply Pipelines and Services	265
	18.6	Tailings Storage Facility	267
	18.7	Fire Suppression Infrastructure	268
	18.8	First Aid Facilities	269
	18.9	Security Infrastructure	270
	18.10	Communication Systems	271
19.0	MARKET STUDIES AND CONTRACTS		273
	19.1	Market Review and Metal Prices	273
	19.2	Material Contracts	274
20.0	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT		277
	20.1	Environmental	277
	20.2	Permitting	281
	20.3	Social and Community Impact	282
21.0	CAPITAL AND OPERATING COSTS		288
	21.1	Capital Cost	288
	21.2	Operating Cost	291
22.0	ECONOMIC ANALYSIS		299

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	22.1	Production Profile	299
	22.2	Revenue and Revenue Assumptions	299
	22.3	Annual Cash Flow Results	301
	22.4	Net Present Value and Internal Rate of Return	302
	22.5	Taxes and Royalties	302
	22.6	Sensitivities	303
23.0	ADJACENT PROPERTIES		306
	23.1	Public Domain Information about Adjacent Properties	306
	23.2	Source of Adjacent Property Information	307
	23.3	Relevance of the Adjacent Property Information	307
	23.4	Application of the Adjacent Property Information	308
24.0	OTHER RELEVANT DATA AND INFORMATION		309
	24.1	Construction and Development	309
	24.2	International Reporting Standards	310
25.0	INTERPRETATION AND CONCLUSIONS		311
	25.1	Geology and Estimation	311
	25.2	Mining Methods	312
	25.3	Metallurgical Processing	312
	25.4	Economic Outcome	312
26.0	RECOMMENDATIONS		313
	26.1	Objectives to be Achieved in Future Work Programmes	313
	26.2	Detailed Future Work Programmes	313
	26.3	Declaration by QP with respect to the Project’s Warranting Further Work	313
27.0	REFERENCES		314
28.0	DATE AND SIGNATURE PAGE		316
	28.1	Certificates	316
	APPENDICES		322

LIST OF TABLES

Table 1-1: Mineral Resource Estimate for the Maseve Project	16
Table 1-2: Mineral Reserve 4E Statement	19
Table 1-3: MR and UG2 Prill Split	20
Table 1-4: Key Operating Cost Details	35
Table 1-5: Capital Cost Summary	36
Table 1-6: Average 3 year Trailing Metal Prices used in Financial Model	37
Table 1-7: Basket Prices	37
Table 1-8: Economic Evaluation Summary	37
Table 5-1: Meteorological Data from Rustenburg and Pilanesberg Meteorological Stations for period 2001 – 2010	53
Table 5-2: Local Community and Town Distances in Relation to Project Site	55
Table 5-3: Vegetation Species Evident in Project Area	58
Table 5-4: Bird Species (Red Data) that may occur in Project Area	60
Table 5-5: Snakes (Red Data) that Potentially occur in Project Area	60

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Table 5-6: Mammals that Potentially occur in Project Area	61
Table 6-1: Historical Mineral Resources Estimated for Project 1 and 1A – October 2009 (100% WBJV Area)	63
Table 8-1: Geological Losses Estimated for the MR and UG2	95
Table 11-1: Standards Currently in Use	111
Table 14-1: Maseve Mineral Resource Estimation Tabulation	129
Table 14-2: Average Specific Gravities per Lithology	133
Table 14-3: Drillhole Data	135
Table 14-4: Statistical Analysis – Drill Hole Data per Geo Zone: MR	139
Table 14-5: Statistical Analysis – Drill Hole Data per Geo Zone: UG2	141
Table 14-6: Descriptive Statistics for the MRFW Units	144
Table 14-7: Summary of Mining Widths	145
Table 14-8: Variogram Parameters	147
Table 15-1: Block Model per Mining Blocks - MR	165
Table 15-2: Block Model as per Mining Blocks – UG2 Reef	166
Table 15-3: Planning Pay Limit Basis	167
Table 15-4: MR Dilution % per Mining Method	171
Table 15-5: UG2 Dilution % per Mining Method	171
Table 15-6: Basic Grade Equation – MR	172
Table 15-7: Basic Grade Equation – UG2	174
Table 15-8: Mineral Reserve Statement	175
Table 15-9: MR and UG2 Prill Split	176
Table 16-1: Breakdown	185
Table 16-2: Bord and Pillar Design at Depth to Surface	199
Table 16-3: MR Support Recommendations in Conventional and Hybrid Stoping Sections	201
Table 16-4: UG2 Support Recommendations in Conventional and Hybrid Stoping Sections	201
Table 16-5: BME – MR and UG2	209
Table 16-6: Bord and Pillar BME (Basic Mining Equation)	209
Table 16-7: Conventional Stoping - MR BME	210
Table 16-8: Conventional Stoping – UG2 BME	211
Table 16-9: Hybrid Stoping – MR BME	211
Table 16-10: Hybrid Stoping – UG2 BME	212
Table 16-11: Underground Mine Design Criteria and LoM Scheduling Rates	222
Table 16-12: LHD Cleaning of 3.5m W x 2.1m H End	228
Table 16-13: LHD Cleaning of 5.0m W x 4.0m H End	229
Table 16-14: LHD Cleaning of 6.5m W x 3.8m H End	230
Table 16-15: 30t Trucks – Cycles	233
Table 16-16: WBJV1 Trackless Mobile Machinery Equipment List: Current vs Future	237
Table 18-1: Current Diesel Fleet – Diesel Consumption	265
Table 18-2: Future Diesel Fleet – Diesel Consumption	265
Table 19-1: 3 Year Average Trailing Metal Prices	274
Table 21-1: Capital Cost Summary	288
Table 21-2: Operating Cost Summary	292
Table 21-3: Total Net On Site Cost	297
Table 21-4: Total Net On Site Cost per Reef Type	298
Table 22-1: Metal Prices used in Financial Model	299
Table 22-2: Basket Prices	300
Table 22-3: Revenue net of Offtake Agreement per Reef Type	300
Table 22-4: Economic Evaluation Summary	302
Table 22-5: Sensitivity Analysis (From Start of Project)	304
Table 22-6: Sensitivity Analysis (July 2015 Onwards)	305

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LIST OF FIGURES

Figure 1-1: Relationship between Mineral Resource and Mineral Reserves	18
Figure 1-2: WBJV Project 1 Maseve LoM MR Design	24
Figure 1-3: WBJV Project 1 Maseve LoM UG2 Design	25
Figure 1-4: Overall Site Layout	30
Figure 1-5: Capital Expenditure Profile	36
Figure 4-1: Location of the Maseve Mine Area	47
Figure 4-2: Locality Plan of the Project Areas in the Maseve Mine	48
Figure 5-1: Monthly Average Rainfall Distribution (Rainfall Station 0548165)	54
Figure 7-1: Location of the WBJV in Relation to the Western Limb of the BIC	68
Figure 7-2: Detailed Stratigraphy of the Western Bushveld Sequence	72
Figure 7-3: Regional Structural Data	75
Figure 7-4: Major Faults and Dykes at the Project Area	77
Figure 7-5: Structural Blocks	78
Figure 7-6: Crystallisation and Setting Zones	79
Figure 7-7: Schematic Cross-Sections	81
Figure 7-8: Regional Facies Types	83
Figure 7-9: Facies Types at the Project	84
Figure 7-10: Location of the UG2 Facies Types in Maseve Project Area	86
Figure 7-11: Geological Domains – MR	88
Figure 8-1: Location of MR Geostatistical Zones pertaining to the Project Areas	92
Figure 8-2: Location of UG2 Geostatistical Zones pertaining to the Project Areas	93
Figure 8-3: Plan View of the MR Model illustrating Dip Variations	94
Figure 8-4: Map Showing Areas to which Geological Losses were applied	96
Figure 14-1: Drill Hole Data Utilised in the Estimation of Mineral Resources for the WBJV Project1 (Maseve)	131
Figure 14-2: Reef Definition Process	132
Figure 14-3: Drill Hole Data Utilised in the Estimation of Mineral Resources for the WBJV Project 1 (Maseve)	135
Figure 14-4: Location of MR Drill Holes Utilised in the Estimation of Mineral Resources for the WBJV Project 1 (Maseve)	136
Figure 14-5: Location of UG2 Drill Holes Utilised in the Estimation of Mineral Resources for the WBJV Project 1 (Maseve)	137
Figure 14-6: MR Search Volumes	153
Figure 14-7: UG2 Search Volumes	154
Figure 14-8: Minimum Samples Employed to Estimate the MR	155
Figure 14-9: Minimum Samples Employed to Estimate the UG2	156
Figure 14-10: MR Regression Slope Plot for 4E Content	157
Figure 14-11: UG2 Regression Slope Plot for 4E Content	158
Figure 14-12: Kriging Efficiency Plot for the MR	159
Figure 14-13: Kriging Efficiency Plot for the UG2	160
Figure 14-14: 2011 Mineral Resource Classifications for the MR	161
Figure 14-15: 2011 Mineral Resource Classifications for the UG2	162
Figure 15-1: MR Mining Blocks at 2.5g/t Stope Pay Limit	164
Figure 15-2: UG2 Mining Blocks at 2.5g/t Stope Pay Limit	166
Figure 15-3: MR Block Model – 4E Grade	172
Figure 15-4: Basic Grade Equation – MR	173
Figure 15-5: UG2 Block Model – 4E Grade	174
Figure 15-6: Basic Grade Equation – UG2 Reef	175

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Figure 15-7: MR Prill Splits	177
Figure 15-8: UG2 Prill Splits	177
Figure 16-1: North Shaft Current Workings	179
Figure 16-2: South Shaft Current Workings	181
Figure 16-3: WBJV Project 1 (Maseve) New LoM MR Design	187
Figure 16-4: WBJV Project 1 (Maseve) New LoM UG2 Design	188
Figure 16-5: Declines Configuration	189
Figure 16-6: TMM and Belt Decline Excavation Layouts	190
Figure 16-7: Primary Access Development	191
Figure 16-8: Ramps Depicting Various Equipment	192
Figure 16-9: Bord and Pillar Mining Section	193
Figure 16-10: Conventional Mining Section	194
Figure 16-11: Hybrid Mining Section	195
Figure 16-12: Bord and Pillar Layout	196
Figure 16-13: Typical Conventional Mining Section Layout	200
Figure 16-14: Conventional Stoping Panel Layout with Support	201
Figure 16-15: Loading Bay and Winch Layout in a Conventional Stoping Section	203
Figure 16-16: A Typical Hybrid Mining Section Layout	205
Figure 16-17: Hybrid Stoping Panel Layout with Support	206
Figure 16-18: Winch and Scraper Cleaning	227
Figure 16-19: HPE Reticulation Layout of a Stope	235
Figure 17-1: MF1 Rougher Rate Test: Grind-Recovery Curve	247
Figure 17-2: MF2 Flowsheet	248
Figure 17-3: Maseve MF1 Flowsheet	248
Figure 17-4: Maseve Plant Layout	251
Figure 18-1: Overall Site Layout	252
Figure 18-2: Plant Layout	253
Figure 18-3: North Shaft Layout	254
Figure 18-4: South Shaft Layout	255
Figure 18-5: North Portal Surface Workshop	257
Figure 18-6: South Portal Surface Workshop	257
Figure 18-7: Impofu Substation and Generator Layout	263
Figure 18-8: First Aid Facility Layout	270
Figure 23-1: Surface Owners	308

LIST OF GRAPHS

Graph 1-1: MR Prill Split	21
Graph 1-2: UG2 Prill Split	21
Graph 1-3: WBJV1 LoM Production Schedule per Reef	33
Graph 1-4: WBJV1 LoM Production Schedule per Mining Method	34
Graph 15-1: MR Grade Tonnage Curve	168
Graph 15-2: UG2 Reef Grade Tonnage Curve	169
Graph 16-1: North Shaft Monthly Production	180
Graph 16-2: South Shaft Monthly Production	181
Graph 16-3: Bord and Pillar DTS Distribution	198
Graph 16-4: Reserve Break-up - MR	208
Graph 16-5: Reserve Break-up – UG2 Reef	208
Graph 16-6: Total Labour Build-up – Total Mine: MR	214

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Graph 16-7: Stoping Crew Requirements – Total Mine: MR	215
Graph 16-8: Development Crew Requirements – Total Mine: MR	215
Graph 16-9: Labour Split – Total Mine: MR	216
Graph 16-10: Labour Efficiency per ton – Total Mine: MR	217
Graph 16-11: Labour Efficiency per m2– Total Mine: MR	217
Graph 16-12: Labour Efficiency per m2 – Bord and Pillar: MR	218
Graph 16-13: Bord and Pillar build-up - m2 – Total Mine: MR	218
Graph 16-14: Labour Efficiency per m2 – Conventional Stoping: MR	219
Graph 16-15: Conventional build-up - m2 – Total Mine: MR	219
Graph 16-16: Labour Efficiency per m2 – Hybrid Stoping: MR	220
Graph 16-17: Hybrid build-up - m2 – Total Mine: MR	220
Graph 16-18: WBJV1 LoM Total Development Meters	224
Graph 16-19: WBJV1 LoM Production Schedule per Reef Type	225
Graph 16-20: WBJV1 LoM Production Schedule per Mining Method	226
Graph 16-21: 30t Trucks – tonnes Trammed vs Distance Travelled	233
Graph 16-22: Total Fleet Deployment Composition	238
Graph 21-1: Capital Expenditure Profile	291
Graph 22-1: 4E Ounces Produced in Concentrate	301
Graph 22-2: Annual Cash Flow Results	302
Graph 22-3: IRR Sensitivity from Project Start	303
Graph 22-4: IRR Sensitivity from July 2015 Onwards	303

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1.0	SUMMARY
1.1	Introduction

Since the update on the Feasibility Study in 2009, changes to the mining methodology and method of access, on and off reef development have been reviewed. In the 2009 Updated Feasibility Study, access development was to be via 3 decline systems with trackless development and ore transport, with raise development, ledging, equipping and stoping being conventional hand held mining methodology.

Reason for Updated NI43-101 Report

The Mineral Resources and Mineral Reserves for Project 1 have been updated to account for the planned increased use of mechanized mining methods where the deposit is estimated to be thicker and accessible from nearby completed underground development. The updated Mineral Reserves have been calculated using current three year trailing metal prices and current cost estimates, updated detailed surface and underground drilling results and a revised mine plan.

Production guidance for 2016 is 116,000 ounces platinum, palladium, rhodium and gold (“4E”) (100% Project basis) and 185,000 ounces 4E in 2017 in concentrate. Steady state has been estimated to be 250,000 ounces 4E per year.

Exclusive of smelter discount, on site costs are estimated to be US$526 (12R/US$) per 4E ounce for the life of mine on the Merensky Reef including copper, nickel and other minor elements as a credit and US$774 per 4E ounce on the UG2 (12R/US$). The planned increased use of mechanized mining methods in areas near current development, and a slightly weaker Rand has resulted in similar cost guidance to earlier estimates despite increased labour and other cost escalation in Rand terms.

1.2 Ownership

The Western Bushveld Joint Venture (“WBJV”) Maseve property is located in the western limb of the Bushveld Igneous Complex (“BIC”), 110km West-NorthWest of Pretoria and 120km from Johannesburg. The WBJV is owned by Maseve Investments 11 (Pty.) Ltd. 82.9% owned by Platinum Group Metals (RSA) (Pty) Ltd a wholly owned subsidiary of Platinum Group Metals Ltd Canada the issuer. The resources of the WBJV Project 1 and 1A are located approximately 1km from the active Merensky reef (“MR”) mining face at the operating Bafokeng Rasimone Platinum Mine (“BRPM”) along strike. BRPM completed opencast mining on the UG2 Reef within 100m of the WBJV property boundary.

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The government of South Africa holds the mineral rights to the project properties under the new Act, No. 28 of 2002: Mineral and Petroleum Resources Development Act, 2002. The mineral rights are held through a mining right under the Mineral and Petroleum Resources Development Act, 2002.

1.3 Geology

The WBJV property is partly situated in a layered igneous complex known as the BIC and its surrounding sedimentary footwall rocks. The BIC is unique and well known for its layering and continuity of economic horizons mined for platinum, palladium and other platinum group elements (PGE’s), chrome and vanadium.

The area is structurally complex with numerous phases of faulting as well as soft-crystalline deformation within the MR and UG2 layers.

Major structures, which occur within the WBJV area, include the Caldera and Elands Faults, Chaneng dyke and a major North-South trending feature, which can be observed across the entire Pilanesberg Complex. These East-West trending structures dip steeply (between 80° and 90°). The magnetics indicate that the Chaneng Dyke dips steeply to the North. This is consistent with similar structures intersected underground on the neighbouring BRPM, which all dip steeply Northward.

1.4 Mineralisation

The potential economic horizons in the WBJV Maseve Project area are the MR and UG2 situated in the Critical Zone of the Rustenburg Layered Suite (“RLS”) of the BIC. These horizons are known for their continuity. The MR and UG2 are mined at the BRPM adjoining the WBJV property as well as on other contiguous platinum mine properties. In general, the layered package dips at less than 20° and local variations in the reef attitude have been modelled. The MR and UG2, in the Project Area, generally dip between 4° and 42°, averaging 22°.

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The precious metals occur in a variety of forms. One or more of the metals may be present in combination with sulphur, arsenic, selenium or tellurium metallic particles of PGE’s or of PGE’s alloyed with base metals are also found. Additional PGE’s are found in solid solution in base metal sulphide particles.

1.5 Project Status

The database for the Maseve Project available for mineral resource estimation comprises a total of 669 drill holes (comprising of original parent holes only, excluding deflections). The MR Mineral Resource estimate is based on 366 intercepts and the UG2 mineral resource estimate is based on 415 intercepts. An additional 213 drillholes were included in the resource estimate Update which comprise new holes drilled by PTM in the time elapsed since the previous resource estimate of 2009.

At the stage of this resource estimate the Project was in construction and mine building with reconnaissance underground development continuing and plant development in progress. This is planned to expand to a fully operational mine with commissioning in 2015.

1.6 Resources

Mineral resource estimation was conducted using Datamine StudioTM and Minesoft’s geostatistical package ‘RES’ adopting an ordinary kriging method of resource estimation. In keeping with industry best practice in mineral resource estimation, allowance is made for known and expected geological losses. From drill data and other known information areas with no reef have been delineated and excluded from mineral resource estimation. These areas comprise 35% of the project area. Within expected reef areas, further geological losses of up to 14% for the MR and 13% for the UG2 were applied to the area to accommodate for areas of potentially un-mineable structural and geological conditions, and this was considered in the Mineral Resource estimate. This geological loss considers losses for faults, dykes, potholes and areas of iron replacement pegmatite. Structural loss estimates are based on drilling, field mapping and remote sense data, which include a high resolution aeromagnetic survey.

Total measured and Indicated mineral resources amount to 6.63 million ounces (“Moz”) of 4E (platinum, palladium, rhodium and gold) for Maseve Project area. The mineral resource estimate for the Maseve Project area is shown in the following tables.

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Table 1-1: Mineral Resource Estimate for the Maseve Project

Merensky - Mining Cut 100% Project Basis
Resource Category	Cut-off	Tonnage	Grade					Metal		Reef Width
	4E		Pt	Pd	Rh	Au	4E		4E
	cmg/t	Mt	g/t	g/t	g/t	g/t	g/t	kg	Moz	cm
Measured	300	9.266	3.35	1.41	0.21	0.26	5.23	48,461	1.558	152
Indicated	300	12.552	3.65	1.54	0.23	0.29	5.71	71,672	2.304	141
Total	300	21.818	3.53	1.49	0.21	0.28	5.51	120,133	3.862	146
Inferred	300	0.196	2.32	0.98	0.14	0.18	3.62	710	0.023	118
UG2 - Mining Cut 100% Project Basis
Resource Category	Cut-off	Tonnage	Grade					Metal		Reef Width
	4E		Pt	Pd	Rh	Au	4E		4E
	cmg/t	Mt	g/t	g/t	g/t	g/t	g/t	kg	Moz	cm
Measured	300	8.496	2.29	0.94	0.36	0.04	3.63	30,841	0.992	140
Indicated	300	14.183	2.46	1.01	0.39	0.04	3.90	55,314	1.778	136
Total	300	22.679	2.39	0.99	0.38	0.04	3.80	86,155	2.770	137
Inferred	300	0	0	0	0	0	0	0	0	0

1.7 Mineral Reserves

The mineral reserves are fully included within the measured and indicated mineral resources, and are not in addition to them.

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The mineral reserve statement for the WBJV project 1 and 1A (Maseve) is based on the South African Code for the Reporting of Exploration Results, Mineral resource and Mineral reserves (SAMREC code). There is no material difference between the SAMREC and CIM code for mineral reserve estimation in this case.

The SAMREC code definition of a Mineral Reserve is:

“A ‘Mineral Reserve’ is the economically mineable material derived from a Measured or Indicated Mineral Resource or both. It includes diluting and contaminating materials and allows for losses that are expected to occur when the material is mined. Appropriate assessments to a minimum of a Pre-Feasibility Study for a project and a Life of Mine Plan for an operation must have been completed, including consideration of, and modification by, realistically assumed mining, metallurgical, economic, marketing, legal, environmental, social and governmental factors (the modifying factors). Such modifying factors must be disclosed.”

Mineral reserves are reported as inclusive of diluting and contaminating uneconomic and waste material delivered for treatment or dispatched from the mine without treatment.

The CIM code definition for a Mineral Reserve:

“A Mineral Reserve is the economically mineable part of a Measured and/or Indicated Mineral Resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at Pre-Feasibility or Feasibility level as appropriate that include application of Modifying Factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

The reference point at which mineral reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.”

For this technical report, the CIM mineral eserves for the WBJV project 1 and 1A has been stated under the SAMREC Code. The point of reference is ore delivery to the RoM silo at the processing plant.

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Mineral reserves are sub-divided in order of increasing confidence into probable mineral reserves and proven mineral reserves. A probable mineral reserve has a lower level of confidence than a proven mineral reserve.

A probable reserve is the economically mineable part of an indicated resource, and in some circumstances a measured resource. This is demonstrated by at least a Pre-Feasibility Study (“PFS”) including adequate information on mining, processing, metallurgical, economic and other factors that demonstrate, at the time of reporting, the economic extraction can be justified.

A proven reserve is the economically mineable part of a measured resource demonstrated by the same level and factors as above. A proven mineral reserve implies that there is a high degree of confidence. All mining and permit approvals need not be in place for the declaration of reserves.

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The conversion to mineral reserves was undertaken at 2.5g/t stope cut-off grade for both MR and UG2 reefs. From the mineral resource as estimated in this report, each stope has been fully diluted, comprising of a planned dilution and additional dilution for all aspects of the mining process. There are no inferred mineral resources included in the Reserves. The Qualified Person for the Statement of Reserves is Mr G. Roets (DRA Projects SA (Pty) Ltd) (“DRA”).

The conversion of the 2015 Updated Mineral Resource Estimate to Reserves differs from the 2009 Updated Feasibility Study Reserve calculation in the following aspects:

>	A lower planning face cut-off grade of 2.5g/t (vs 3.5g/t from 2009 Updated Feasibility Study (“FS”) was used,
>	The software used for reserve valuation is Datamine. Studio5 was used for mine design and EPS for production scheduling.

The Mineral Reserve statement has been calculated based on the outcome of the updated reserve calculation and economic evaluation and is detailed in Table 1-2.

Table 1-2: Mineral Reserve 4E Statement

Estimated Total Reserve 100% Project Basis
	Reserve tonnes - Mt	Pt g/t	Pd g/t	Rh g/t	Au g/t	Reserve 4E Grade - g/t	Reserve 4E Content – t	Reserve 4E Content - Moz
MR Proven and Probable	17.525	2.94	1.24	0.18	0.23	4.59	80.401	2.585
UG2 Proven and Probable	14.914	2.01	0.83	0.32	0.03	3.19	47.649	1.532
Total	32.439	2.51	1.05	0.25	0.14	3.95	128.050	4.117

Prill splits are calculated using the individual metal grades reported as a percentage of the total 4E grade.

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Table 1-3: MR and UG2 Prill Split

Merensky Reserve
	Reserve tonnes - Mt	Pt g/t	Pd g/t	Rh g/t	Au g/t	Reserve 4E Grade - g/t	Reserve 4E Content - t	Reserve 4E Content - Moz
Proven	7.082	2.89	1.22	0.18	0.22	4.51	31.905	1.025
Probable	10.443	2.98	1.26	0.18	0.23	4.65	48.496	1.560
Total	17.525	2.94	1.24	0.18	0.23	4.59	80.401	2.585

UG2 Reserve
	Reserve tonnes - Mt	Pt g/t	Pd g/t	Rh g/t	Au g/t	Reserve 4E Grade - g/t	Reserve 4E Content - t	Reserve 4E Content - Moz
Proven	5.452	1.95	0.80	0.31	0.03	3.09	16.821	0.540
Probable	9.462	2.05	0.85	0.33	0.03	3.26	30.828	0.992
Total	14.914	2.01	0.83	0.32	0.03	3.19	47.649	1.532

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1.8 Mining Operations

1.8.1 Geotechnical Factors

The main findings in the geological and rock engineering investigations that influenced on-reef mine design are discussed below:

> The MR has an average dip of 15.31° and an average stoping width of 142cm at a cut-off grade of 2.5g/t,

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>	The UG2 reef has an average dip of 16° and an average stoping width of the mine is 129cm at a cut-off grade of 2.5g/t,
>	Certain mining blocks have the potential for increased mechanization, while other blocks have a greater potential for mining methods more suitable to narrow, steep dipping ore bodies. Currently predicted dip, structure and width can further be confirmed by additional drilling, either from surface or from underground with stope definition drilling,
>	A complex geological structure with faults and dykes intersecting the ore body subdivides the deposit into a number of discrete mining blocks, each of which requires access development on different mining elevations. The resultant blocks of ground left un-mined add to the regional stability of the mine.

After application of appropriate pay limits, the MR reserve contains 40% more recoverable metal than the UG2 and is therefore the primary target. The parting between the MR and the UG2 reserve varies in thickness from contact in the West to 70m in the East and deeper part of the deposit. Mining of both reefs generally only occurs when the parting is greater than 20m as prescribed by the rock engineer.

1.8.2 Mining Methods Selected

The geological and structural models, in conjunction with geotechnical considerations, formed the basis of the mining methods that were selected to provide the best practical outcome under the given conditions. The selected methods had to be versatile and easily interchangeable with the lowest impact on production during transition between mining methods. The mining methods would also have to be able to integrate with the trackless environment that would supply access and service all of the mining areas.

1.8.3 Final Chosen Mining Methods:

> Bord and Pillar

• Bord and pillar mining was considered in flatter dipping areas with a maximum dip of 15º and where the actual seam height of the reef was conducive to the increased mining heights required for the bord and pillar equipment.

> Conventional Mining

• The conventional mining method was considered in the steepest dipping areas. Raises can be developed at a maximum dip (or apparent dip) of 34º, thus giving the best and most practical outcome conditions with the lowest replacement rates. under high variability

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> Hybrid Mining

• The hybrid mining method was considered in the moderate to steeper dipping areas where the bord and pillar method is not eligible. This method took precedence over the conventional method where applicable, due to the on- reef access which allows a quicker reef access and a faster production build- up. The method also requires a more favourable development replacement rate.

1.8.4 Mine Design

The updated Life of Mine (“LoM”) design, as illustrated in Figure 1-2 and Figure 1-3, makes use of the twin decline system to access the underground workings. Initially during the development phase, men, material and rock will be transported by trackless mobile machinery. At steady state production, men and rock will be transported by chairlift and conveyor systems respectively (Refer to Section 18 of this report).

Mining methods have been adapted from the Updated Feasibility Study to include geological, geotechnical, engineering and timing modifications. Maseve will be operated as a trackless development (access development is off-reef and production development is on-reef) and a partially conventional, hybrid, and bord and pillar mechanised mine using diesel mobile mining equipment. The overall on-reef / hand held methods applied in the conventional and hybrid mining methods have not changed significantly from the conventional methods described in the previous study. The thicker reef (partially due to the lower cut-off grade and additional drilling information) and a deliberate drive towards a higher degree of mechanization, allows for a bord and pillar mining method to be applied in the deeper, shallower dipping areas of the mine.

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The mine design is focused on reducing waste footwall development by replacing most of the previously footwall located off-reef production development in the earlier design with on-reef production development. This accommodates the trackless mining method approach and delivers a faster Run of Mine (“RoM”) production build-up. The approach also allows for a reduced overall mining cost but it does however result in a moderately higher dilution percentage.

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This report documents the modifying factors including updated geological model, optimized mining plan and updated financial and economic models in support of the updated mineral reserves. The cost analysis and accuracy estimates have been updated to reflect changes in capital and operating costs associated with adjustments in the mine plan and current prices.

This document provides details of the mine plan modifications and associated layouts related to the updated mineral reserves. These changes include:

>	Two decline systems at North Shaft and South Shaft to access the ore body (the 2009 Updated Feasibility Study included a three decline system),
>	Mining blocks have changed in geometry, position and strike,
>	Dimensions and cross sectional areas of ends for the material decline, material ramp decline and strike drives.

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Realised development rates have been higher than estimated in the 2009 Updated FS. Updated rates have a positive effect on the LoM scheduling. Realised development rates average:

>	Material decline 100m per month,
>	Material ramp decline 80m per month,
>	Strike ramps 60m per month,
>	Increase size of diesel equipment fleet from 52kW/ktpm to over 65kW/ktpm.

Ventilation requirements for the modified layout consider:

>	Intake airway capacity with two declines is less than with three declines,
>	Raise Bore Hole (“RBH”) size and location,
>	Main fan(s) operating point, location and phase-in.

1.8.5 Ventilation

Subsequent to the updated mine designs completed in March 2012, changes were made to mine design to improve the viability of the Maseve Project. Some of the proposed changes have impacted on the original ventilation designs and specifications. The ventilation section assesses the impact of the significant changes on ventilation designs and associated costs.

The ventilation strategy considers safety and health in accordance with the Mine Health and Safety Act (“MHSA, Act 29 of 1996”) and complies with Maseve health and safety requirements. The primary ventilation quantity for Maseve is 1 100m³/s; dictated by the need to dilute diesel emissions, remove heat and dilute blasting fumes (during re-entry period). The primary ventilation quantity satisfies the mine heat load without the need for refrigeration. It must be noted however that, wet-bulb temperatures will exceed 27.5°C and approach 29.0°C and heat tolerance screening of the underground work force will be required. Interactive computer simulation of heat and air flow was used to determine ventilation requirements over the LoM for maximum depth and strike.

The mining plan is based on steady state production of 160 000 reef tonnes per month. During the first phase of the project, the primary ore body will be MR and accordingly discussion in this report focusses on access of the MR. Later UG2 will provide replacement tonnes and will be ventilated utilising the ‘existing’ MR infrastructure by extending established intake and returns to UG2 as required (e.g. step raise bore holes (RBH’s), drop raises, horizontal intake and return airways). The strategy will mine UG2 within a specific mining block only after the ‘overlying’ MR block has mined out, i.e., the two reefs will not be mined simultaneously from the same area.

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In general, each mining block will be ventilated as a separate district while at the same time utilising as much common infrastructure as practically achievable. Fresh air will be introduced to mining blocks through a combination of the main North and South access decline systems and strategically located fresh air RBH’s. Air returns through return RBH’s equipped with fans. Generally returns will serve more than one block, but in some cases a blocks will require a dedicated return to surface. It should be noted that RBH’s were phased in to meet the production requirements as provided.

1.9 Metallurgical Testwork and Recovery

Three sets of testwork have been conducted on the WBJV Elandsfontein deposit by SGS Lakefield for the Pre-Feasibility Study (“PFS”) and Mintek for the 2009 Updated Feasibility Study (“UFS”). SGS Lakefield completed metallurgical testwork on UG2 and MR (December 2006 and January 2007 respectively) reefs to characterise the ores and evaluate metallurgical performance. From comminution testwork, the UG2 ore was classified as being of medium to hard hardness. The ore could be treated using a standard MF2 circuit and the predicted recoveries were 82% (4E) with a grade of 150g/t. The predicted PGM recovery and grade for MR were 94% and 179g/t (4E) respectively. Copper and Nickel recoveries achieved were 89% and 59.5% respectively at grades of 2.4% and 3.6% . The testwork was conducted at a fine grind of 90% passing 75µm. The testwork was conducted with limited samples (four cores samples).

The Mintek 2009 UFS metallurgical testwork was conducted on MR and UG2 samples collected across the target mining area. The mineralogy, grade and ore occurrence exhibited marked variability. The MR grade varied from 1.9g/t to 8.5g/t (4E) with an average of 5.3g/t. The testwork was conducted with a location composite with a grade of 2.5g/t. Both MF1 and MF2 circuit configurations were tested. The overall MR MF1 recovery and grade were 86% (4E) and 61g/t respectively. It was also demonstrated during the rougher rate tests that for MF1, recovery increased with grinds being 88.9%, 90.1% and 94% for grinds of 40%, 60% and 90% passing 75µm respectively. The overall copper and nickel recoveries were 86% and 57%. Respective grades for copper and nickel were 1.6% and 2.1% . For MF2 overall 4E recovery and grade were 91% and 85g/t respectively. Overall copper and nickel recoveries were 84% and 58% respectively at grades of 2% and 2.8% .

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The UG2 testwork gave overall recovery of 86% at a grade of 102g/t with chromite grade less than 4%.

Mintek conducted confirmatory testwork in 2012 from nine MR drill core samples collected across the mining area. The testwork demonstrated that the ore was not significantly different to the other ores tested previously. The samples also demonstrated marked variability with 4E grade varying between 1.6g/t and 6.6g/t. Copper and nickel grades were 0.14% and 0.17% . Comminution testwork classified the ore as being hard. Rougher rates for MF2 and MF1 were 94% and 91% respectively. The cleaner efficiency was 95%, giving predicted recoveries of 89.3% and 86.5% for MF2 and MF1 respectively. The testwork also demonstrated an increase in recovery for the MF1 configuration for increase in fineness from 40%, 60% and 80% passing 75µm, with the extended time recoveries for the 60% and 80% being close to each other. From the MF2 locked cycle testwork overall 4E recovery and grade were 87.1% and 135g/t respectively at a mass pull of 3.3% . Copper and nickel recoveries were 88% and 64% at grades of 1.9% and 3.5% respectively.

Tailings thickening tests conducted yielded a flux of 0.7m ²/hr tonne.

DFS metallurgical testwork for UG2 was conducted, with seven drill core samples at an overall grind of 80% passing 75µm applying an MF2 configuration. Metallurgical characterisation confirmed that the response was variable but similar to ores tested previously. Overall 4E recovery and grade were 79% and 109g/t respectively. The chromite grade in concentrate was 3.7% .

1.10 Process Plant Design

The process plant design utilises a standard mill-float-mill-float (MF2) circuit configuration that is applied to treat PGM ores of the BIC. The plant has been designed to treat ore at a rate of 165 000tpm. The design offers flexibility to treat a blend of MR and UG2 at a predetermined ratio. The MR is the target of initial mining. The mining ramp-up to steady state production of 165 000tpm is over a three year period. Construction of the concentrator is in two phases, initially with an MF1 circuit during the ramp-up period. The completion of the MF2 circuit will only be decided after reviewing the mining production ramp-up in 2017.

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The MF1 circuit offers lower start-up costs, early revenue and good ore stockpile management. Only equipment required for the MF1 circuit will be installed with all services (power, water and air) completed during this phase. Main equipment installed in phase 1 includes the primary mill, secondary flotation bank (redeployed to primary rougher flotation in this phase), tailings thickener and disposal system and concentrate thickener and filter. The concentrate filter was sized for the full plant capacity (60m²), but only enough plates (48m²) to handle the phase 1 throughput were installed. Sufficient civil work will be completed in phase 1 to minimise interrupting production and allow for safe construction during completion of the MF2 plant.

The MF1 circuit will treat between 80 000 and 115 000tpm at a grind of just over 60% passing 75µm. The MF1 testwork shows that for this grind MF1 recoveries are between 1% and 3% lower than MF2 for extended residence time. The secondary rougher installed in the initial phase offers long residence time. The MF1 recovery has been discounted by 1.5% for the initial phase.

Any confirmatory testwork required for the UG2 ore will be determined at a later stage.

1.11 Infrastructure

The PTM Maseve site is divided into four secure areas, the operational site is primarily focussed East of the R565 Provincial road namely North Shaft, Plant and South Shaft. The training and induction centre functions are primarily focused West of the R565 at the Training Centre. Each shaft and the plant is equipped with its own offices, change house, control room, maintenance, storage and general management facilities. Senior management offices are located near the concentrator, central to all operational areas.

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The main Eskom supply is located close to North Shaft and the generator station, bulk fuel storage and Power Factor Correction (“PFC”) equipment is strategically located close to the main consumer substation. The mine is not completely self-sufficient on generator power, however, is able to supply 13% of it planned maximum demand with the two generators currently on site.

All critical infrastructure from electrical to bulk materials handling has been designed to accommodate full production out of a single shaft, which provides for any variation in production scheduling that might occur between North Shaft and South Shaft over the LoM.

Other important facilities to note on site are:

>	Medical clinic facilities,
>	Potable water and waste water management facilities,
>	Sewerage treatment works,
>	Tailing Storage Facility (“TSF”),

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>	Security buildings,
>	Visitors centre,
>	Shaft head explosive delivery facilities,
>	Fire water pump stations.

1.12 Environmental Studies, Permitting and Social or Community Impact

Baseline studies have been undertaken within the WBJV Project 1 (Maseve) area, in support of an Environmental Impact Assessment (“EIA”) and Environmental Management Plan (“EMP”), which is part of the mining right application. These studies were conducted to comply with local legislation as well as international requirements and consisted of the following:

>	Soils, land use and land capability study,
>	Fauna and flora Report,
>	Hydrological study,
>	Groundwater specialist report,
>	Air quality impact assessment,
>	Ground vibration and air blast,
>	Visual impact,
>	Archaeological assessment,
>	Traffic assessment.

The EIA summarises relevant results of the environmental and social baseline of the WBJV Project 1 (Maseve) area.

Maseve holds the following governmental authorizations:

>	The Environmental Impact Assessment and Environmental Management Plan (EIA and EMP) was approved by the Department of Mineral Resources in terms of the Mineral and Petroleum Resources Development Act (No 28 of 2002) on the 15 May 2012.
>	The Department of Economic Development, Environmental, Conservation and Tourism (“DEDECT”) granted to Maseve an environmental authorisation in terms of the NEMA to commence with the construction of infrastructure and facilities on 13 September 2013, ref: 30/5/1/2/3/2/1/528EM.

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>	Waste management license 12/9/11/L628/7 (WML) issued by the Department of Environmental Affairs in terms of the National Environmental Management: Waste Act No. 59 of 2008 (NEMWA) for a treatment plant to be utilized during Phase 2 of the Project 1.
>	Water Use License 03/A22F/ABCGIJ/2596 (WUL) issued by the Department of Water and Sanitation in terms of Chapter 4 of the National Water Act, 1998 (Act No 36 of 1998) for various water uses on the Mining Right Area.

Maseve has a programme of work in place to comply with the necessary environmental, social and community requirements. Key work includes:

>	EIA / EMP in accordance with the MPRDA, the National Environmental Management Act (“NEMA”) as well as the Equator Principles (“EP”).
>	Stakeholder Engagement Process in accordance with the NEMA principles.
>	Specialist investigations in support of the EIA / EMP.
>	Integrated Water Use License Application (“IWULA”) in compliance with the National Water Act (“NWA”).
>	Integrated Waste Management License in compliance with the National Environmental Management Waste Act (“NEMWA”).

Maseve posted an environmental rehabilitation guarantee of R 58.5 million as a requirement of the mining right application.

There are several communities within the proposed project area whom are affected by the WBJV Project 1 (Maseve).

1.13 Production schedule

The WBJV Project 1 (Maseve) LoM for both the MR as well as the UG2 is just over 21 years (21.03), of which MR is 11.62 years and UG2 10.94 years. See Graph 1-3 below.

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The production profile follows a phased approach to the planned 160 000tpm at steady state. In the initial phase to October 2017, the MR production builds up to 115 000tpm. Production then builds up to 160 000tpm and reaches steady state in September 2018 where it remains for the LoM.

The tail of the production schedule for the MR starts in January 2025 and final reef tonnes for the MR is scheduled for May 2026. In order to keep the mill fed at 160 000tpm, the UG2 reef starts production in November 2024, where it builds up steadily to the required 160 000 tonnes of reef in May 2026, supplementing the tail of MR. Steady state of the UG2 reef lasts up to May 2032 from where the tail of the UG2 reef decreases to final reef tonnes schedule in July 2035.

The production for each of the individual mining methods selected for the WBJV Project 1 (Maseve) is as shown in Graph 1-4 below.

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The average mining height (on both reef horizons) for conventional mining is 142cm, 146cm for Hybrid mining and 190cm for bord and pillar.

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1.14 Operating Costs

Table 1-4 below sets out key operating cost details of the technical report.

Table 1-4: Key Operating Cost Details

ZAR:USD = 12.00

	ZAR per tonne milled	USD per tonne milled
On Mine Operating Cost	896	75
By- & Co-Product Credits	(106)	(9)
Total Net Mine Site Cash Cost	791	66

	ZAR per 4E oz in concentrate	USD per 4E oz in concentrate
On Mine Operating Cost	8 392	699
By- & Co-Product Credits	(989)	(82)
Total Net Mine Site Cash Cost	7 403	617

>	On-mine operating costs include all mining costs, milling costs, support services and on-mine overheads,
>	By-& co-product credits consist of the revenue derived from Ru, Ir, Ni and Cu.

The life of mine on-mine operating cost is estimated to be ZAR896 per tonne milled or ZAR8 392 per oz 4E (Pt, Pd, Au and Rh) produced in concentrate.

1.15 Capital Costs

The project capital cost is anticipated to be ZAR5 070m to the start of production. An approximate additional ZAR708m will be spent on capital to achieve peak production of 165 000tpm by mid-2018. LoM Capital expenditure including sustaining capex is estimated at ZAR6 418m.

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Table 1-5: Capital Cost Summary

	ZAR ‘million
Sunk to July 2015	4 505
Capital Expenditure to Production	265
Capitalised Operating Cost to Production	300
Total Capital Expenditure to Production	5 070
Remaining LoM Capital Excl. Sustaining Capital	860
Sustaining Capital LoM	487
Total LoM Capital Expenditure	6 418

Project peak funding is estimated at R5 625 million.

1.16 Economic Analysis

The production schedule is as shown above in Graph 1-3 and Graph 1-4. This schedule has been based on the mining plan as developed by the project team and metallurgical results as determined by test work. The capital and operating costs as determined by the project team are included. The toll refining contract has been ratified and the numbers have been incorporated into the financial model.

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The following prices, based on a 3 year trailing average in accordance with U.S. Securities and Exchange Commission ("SEC") guidance, was used for the assessment of resources and reserves.

Table 1-6: Average 3 year Trailing Metal Prices used in Financial Model

Metal	Price	Unit
Platinum	1 408	USD/oz
Palladium	744	USD/oz
Rhodium	1 126	USD/oz
Gold	1 374	USD/oz
Copper	3.18	USD/lb
Nickel	7.11	USD/lb
Iridium	731	USD/oz
Ruthenium	73	USD/oz

The exchange rate between the ZAR and the USD is fixed at ZAR12.00:USD1.00 in the financial model throughout the LoM. The pricing and exchange rates above results in the estimated basket prices shown in Table 1-7 below.

Table 1-7: Basket Prices

Basket Prices per 4E	ZAR/4E Oz	ZAR/4E kg	USD/4E Oz	USD/4E Kg
Merensky Reef	14 590	469 093	1 216	39 091
UG Reef	14 487	465 776	1 207	38 815
Combined in LoM	14 553	467 902	1 213	38 992

The economic evaluation is summarised in Table 1-8 below.

Table 1-8: Economic Evaluation Summary

	NPV @ 5%		NPV @ 10%
Economic Evaluation Summary	ZAR million	USD million	ZAR million	USD million	IRR
Cash flow since start of project	1 597	133	(368)	(31)	8.8%
Cash flow from July 2015 (reporting date)	6 340	528	4 226	352	49.8%

The IRR of the project is based on estimated cash flows from July 2015 onwards is 49.8% .

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1.17 Conclusion and Recommendations

The Project Area represents measured and Indicated mineral resources. The definitions of the mineral resource classification are in accordance with the definitions stated in the SAMREC Code and the CIM Mineral Resource Classifications and comply with disclosure standards of NI 43-101.

The updated geological model has increased confidence in the western areas where iron replacement bodies within the ore body have been delineated out of the mineral resource more accurately, based on known and interpreted localities of such bodies. Based on data spatial localities, the mineral resource estimate was computed from relatively large blocks, i.e. 100m X 100m blocks.

Rolls in the mineralized reefs, mainly along the marginal zone on the western portion of the deposit have been modelled to detail and the mine plan adjusted accordingly to optimize underground layout and haulage.

The Maseve WBJV Project 1 and 1A areas are at development level and thus have had sufficient exploration completed that any material exploration activities and engineering studies have been concluded for these areas. The site infrastructure and processing plant completed to date is appropriate for the mine plan ahead. The project completed and the investment to date allows the project to have an attractive profile, at 3 year average trailing metal prices and reserves, in the current environment.

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2.0 INTRODUCTION

2.1 Issuer

The report has been compiled for Platinum Group Metals RSA (Pty) Ltd (“PTM”), a wholly owned subsidiary of Platinum Group Metals Ltd (Canada) (“PTML”).

2.2 Terms of Reference and Purpose of the Report

This report has been prepared in terms of the Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects, Form 43 101F1 Technical Report and the Companion Policy 43 101CP (“NI 43-101”), which incorporates the Canadian Institute of Mining, Metallurgy and Petroleum (“CIM”) Definition Standards on mineral resources and minerals reserves. The information and status of the project is disclosed in the prescribed manner. The report pertains to the minerals resources for the Maseve Project area, a portion of the WBJV.

The intentions of the report are as follows:

>	Inform investors and shareholders of the progress of the project;
>	Make public and detail the mineral resource and reserve estimation for the Project.

2.3 Sources of Information

The Independent Author / Qualified Person (“QP”) of this report has used the data provided by the representative and internal experts of PTM. This data has been derived from historical records for the area as well as information currently compiled by the operating company, PTM.

2.4 Involvement of the Qualified Person: Personal Inspection

The independent QP’s (CJ Muller, G Cunningham and GC Roets) have visited the WBJV property during 2015 and have undertaken due diligence with respect to the PTM data.

2.5 Frequently used Acronyms, Abbreviations, Definitions and Units of Measure

Abbreviation	Definition
2D	Two Dimensional
3D	Three Dimensional
4E	Platinum, palladium, rhodium and gold
AMSL	Above Mean Sea Level

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Abbreviation	Definition
AP	Anglo Platinum Ltd
Au	Gold
BIC	Bushveld Igneous Complex
BRPM	Bafokeng Rasimone Platinum Mine
CIM	Canadian Institute of Mining
CLO	Community Liaison Officer
CoV	Coefficient of Variation
CSI	Corporate Social Investment
CW	Channel Width
CZ	Critical Zone of RLS
DEA	National Department of Environmental Affairs
DEDECT	Department of Economic Development, Environment, Conservation and Tourism
DFS	Definitive Feasibility Study
DME	South African Department of Minerals and Energy
DMR	Department of Mineral Resources
DRA	DRA Projects SA (Pty) Ltd
DTM	Digital Terrain Model
DWS	Department of Water and Sanitation
EBIT	Earnings Before Interest and Tax
EIA	Environmental Impact Assessment
EMP	Environmental Management Programme
EP	Equator Principles
FPP	Pegmatoidal Feldspathic Pyroxenite
FS	Feasibility Study
FW	Footwall
HDSA	Historically disadvantaged South African
HW	Hanging Wall
IRUP	Iron Replacement Ultramafic Pegmatoid
IWULA	Integrated Water Use License Application / National Water Act (NWA),

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Abbreviation	Definition
LG	Lower Group of CZ
LoM	Life of Mine
LZ	Lower Zone of RLS
MF1	Mill-float
MF2	Mill-float-mill-float
MG	Middle Group of CZ
MHSA	Mine Health and Safety Act 29 of 1996
MPRDA	Mineral and Petroleum Resources Development Act, No. 28 of 2002
MR	Merensky Reef
MRA	Mining Right Application
MZ	Main Zone of RLS
NEMA	National Environmental Management Act
NEMWA	National Environmental Management Waste Act
NFPA	National Fire Protection Association -America
NI 43-101	Canadian National Instrument 43-101 Standards of Disclosure for Mineral Projects
NWA	National Water Act
OK	Ordinary Kriging
PBA	Public Benefit Activities
Pd	Palladium
PFC	Power Factor Correction
PFS	Pre-Feasibility Study
PGE	Platinum Group Element
PR	Prospecting Right
Pt	Platinum
PTM	Platinum Group Metals RSA (Pty) Ltd
PTML	Platinum Group Metals Ltd (Canada)
Ptn	Portion
PXNT	Pyroxenite
QA/QC	Quality Assurance and Quality Control
QP	Qualified Person
RBH	Raise Bore Hole

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Abbreviation	Definition
RBN	Royal Bafokeng Nation
Re	Remaining Extent
Rh	Rhodium
RLS	Rustenburg Layered Suite
RoM	Run of Mine
RPM	Rustenburg Platinum Mines Ltd
SACNASP	South African Council for Natural Scientific Professionals
SAMREC Code	South African Code for the Reporting of Exploration Results, Mineral Resources
SCADA	Supervisory Control and Data Acquisition
SD / SDV	Standard Deviation
SG	Specific Gravity
SK	Simple Kriging
SLP	Social and Labour Plan
SMU	Selective Mining Unit
TSF	Tailings Storage Facility
UCZ	Upper Critical Zone of RLS
UG1	Upper Group No. 1 Chromitite layer
UG2	Upper Group No. 2 Chromitite layer
USD	United States Dollar
UZ	Upper Zone of RLS
WBJV	Western Bushveld Joint Venture
WML	Waste Management License 12/9/11/L628/7
WUL	Water Use License 03/A22F/ABCGIJ/2596
ZAR	South African Rand

Unit	Definition
°	degrees
°C	degrees Celsius
°F	degrees Fahrenheit
cm	centimetre
cm.g/t	centimetre grams per tonne

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Unit	Definition
g/t	grams per tonne
h	hour
ha	hectare
kl	kilolitre
km	kilometre
km2	square kilometres
ktpm	Kilo tonne per month
m	metre
Mbps	megabits per second
min	minute
Ml	Megalitre
m/mo	meter per month
Moz	million ounces
m/s	metre per second
MVA	Mega Volt Amperes
MW	Mega Watts
ppb	parts per billion
Sec	second
t	tonnes
tph	tonnes per hour
t/m3	tonnes per cubic metre
tpm	tonnes per month

2.6 Specific Areas of Responsibility

The QP’s accept overall responsibility for the entire report. The QP’s were reliant, with due diligence, on the information provided by PTM’s internal and non-independent experts. The QP’s have also relied upon the inputs of the PTM personnel in compiling this filing.

GC Roets is a Senior Mining Consultant with the firm DRA Projects SA (Pty) Ltd, of DRA Mineral Park, 3 Inyanga Close, Sunninghill, 2157, Johannesburg, South Africa. He is a coauthor of this report and is responsible for Sections 15, 16 and 18 and jointly responsible for sections 1, 2, 3, 21, 24, 25, 26 and 27.

GI Cunningham is a Senior Metallurgical Consultant with Turnberry Projects and has been involved with the Maseve Project for more than 10 years. He is a co-author of this report and is responsible for Sections 13, 17, 19, 20, 22 and 23 and jointly responsible for sections 1, 21, 24 and 25.

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C Muller is a Principal Geological Consultant with CJM and has been involved with the Maseve Project since 2012. He is a co-author of this report and is responsible for Sections 4, 5, 6, 7, 8, 9, 10, 11, 12 and 14 and jointly responsible for sections 1, 2, 3, 25, 26 and 27.

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3.0 RELIANCE ON OTHER EXPERTS

In preparing this report, the author relied upon:

>	Land title information, as provided by PTM,
>	Geological and assay information supplied by PTM,
>	Drill hole analytical and survey data compiled by PTM,
>	Financial and costing information supplied by PTM,
>	All other applicable information, and
>	Data supplied by other persons and organisations as referenced.

The sources of information were subjected to a reasonable level of inquiry and review. The QP’s have been granted access to all information. The QPs’ conclusion, based on diligence and investigation, is that the information is representative and accurate.

This report was prepared in the format of the Canadian NI 43-101 Technical Report by the QP’s; CJ Muller, G Cunningham and GC Roets. These individuals are considered Qualified Persons under NI 43–101 definitions. The QP’s have reported and made conclusions within this report with the sole purpose of providing information for PTM’s and PTML’s use subject to the terms and conditions of the contract between the QP’s and PTM. The contract permits PTML to file this report, or excerpts thereof, as a Technical Report with the Canadian Securities Regulatory Authorities or other regulators pursuant to provincial securities legislation, or other legislation, with the prior approval of the QP’s. Except for the purposes legislated for under provincial securities laws or any other securities laws, other use of this report by any third party is at that party’s sole risk and the QP’s bear no responsibility.

The QP’s are not qualified to offer legal opinion on title and offer no opinion as to the validity of the titles claimed. The description of the properties and ownership is provided for general purposes only and was supplied by PTM. The QP’s were satisfied with the title to the extent required for the statement of Resources and Reserves.

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4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 Sufficiency of Surface Rights

The WBJV Project 1 (Maseve) purchased the following properties, all of which are registered in the Deed Office Pretoria, these being. Portion 7 (a portion of Portion 2) of the farm Frischgewaagd 96 JQ.Measuring in extent 157.4103 Ha. Portion 14 of the farm Frischgewaagd 96 JQ.Measuring in extent 149.5992 Ha. Portion 17 (a portion of Portion 10) of the farm Frischgewaagd 96JQ. Measuring in extent 215.1301 Ha. Portion 8 of the farm Elandsfontein 102 JQ.Measuring in extent 35.3705 Ha.Remaining Extent of Portion 2 of the farm Elandsfontein 102 JQ. Measuring in extent 751.7458 Ha. Remaining Extent of Portion 9 of the farm Elandsfontein 102 JQ Measuring in extent 403.9876 Ha. Remainder of Portion 10 (a portion of Portion 4) of the farm Frischgewaagd 96 JQ Measuring in extent 216.2703 Ha and Portion 19 (a portion of Portion 2) of the farm, Frischgewaagd 96 JQ Measuring in extent 360.6676 Ha.

4.2 Extent of the Project

The total Maseve Project (Maseve) area includes certain portions of the farms Elandsfontein 102 JQ, Mimosa 81 JQ and Onderstepoort 98JQ, and Frischgewaagd 96 JQ. The mineral rights pertaining to the Maseve mine WBJV cover approximately 67km² or 6 700ha, and the Project Areas cover an area of 10.87km ² or 1 087ha in extent.

4.3 Location of the Maseve Mine (Maseve) Project

The Maseve Project (Maseve) properties are located on the South-Western limb of the BIC some 35km NorthWest of the town of Rustenburg, North West Province, South Africa. The property adjoins Anglo Platinum’s (“AP’s”) Bafokeng Rasimone Platinum Mine (BRPM) and the Styldrift Project to the South-East and East respectively. The Maseve Mine comprises three Project Areas; Project 1, Project 1A and Project 3. Only the Mineral Resources of the Project Areas 1 and 1A have been updated for the purposes of this report. The Project Areas 1 and 1A consist of a section of Portion (“Ptn”) 18, the Remaining Extent (“Re”), Ptn 13, Ptn 8, Re of Ptn 2, Ptn 7, Ptn 15 and Ptn 16 all of the farm Frischgewaagd 96JQ, sections of Ptn 2, Ptn 9 and Ptn 12 of the farm Elandsfontein 102JQ and a small section of the Re of the farm Mimosa 81JQ.

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These properties are centred on Longitude 27° 00’ 00’’ (E) and Latitude 25° 20’ 00’’ (S).

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4.4 Status of Surface Rights and Mineral Title

4.5 Royalties, Payments and other Agreements

The Investment in WBJV Project 1 has been via equity capital. The company has a working capital facility arranged with Sprott Lending for US$40m and is in the process of closing this facility.

The Maseve investments 11 (Pty) Ltd, the holder of the Mining Rights, has a life of mine concentrate offtake agreement with Rustenburg Platinum Mines, a subsidiary of Anglo Platinum. The offtake agreement terms have been considered in the financial model of the project. Project 1 and 1a, and Project 3, are owned by project operating company Maseve Investments 11 (Pty.) Ltd. (“Maseve”), in which the Company has an 82.9% working interest and the Company’s Black Economic Empowerment partner, Africa Wide Mineral Prospecting and Exploration (Pty) Ltd. (“Africa Wide”), a wholly owned subsidiary of Wesizwe Platinum Ltd., owns a 17.1% working interest. Project 3 is not currently being explored or developed. Maseve is governed by a shareholders’ agreement among PTM RSA, Africa Wide and Maseve. A formal mining right was granted for Project 1 on April 4, 2012 by the Government of South Africa.

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4.5.1 The Mineral and Petroleum Resources Royalty Act, 2008 “The Royalty Act” of South Africa

The Royalty Act came into effect on 1 March 2010. The Royalty Act gives effect to the MPRDA which requires that compensation be given to the State (as custodian) of the country’s Mineral and Petroleum Resources to the country’s “permanent loss of non-renewable resource”. The Act distinguishes between refined and unrefined Mineral Resources, where refined minerals have been refined beyond a condition specified by the Act, and unrefined minerals have undergone limited beneficiation as specified by the Act.

The royalty is determined by multiplying the gross sales value of the extractor in respect of that mineral resource in a specified year by the percentage determined in accordance with the royalty formula. Both operating and capital expenditure incurred is deductible for the determination of Earnings Before Interest and Tax (“EBIT”).

The royalty is determined by multiplying the gross sales value of the extractor in respect of that mineral resource in a specified year by the percentage determined in accordance with the royalty formula.

For Unrefined Mineral Resources:

The maximum percentage for unrefined minerals is 7% and the minimum royalty is 0.5% .

4.6 Environmental Liabilities

The WBJV Project 1 Maseve has, since the inception of the grant of its mining right, filed with the DMR all of the required regulatory environmental compliance reports being representative of two yearly monitoring and performance assessments of the environmental management plan and annual financial determination for environmental remediation in terms of regulation 55 and section 41 of the MPRDA respectively. The DMR has accepted and approved these reports.

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4.7 Permitting Status

The WBJV Project 1 (Maseve) was granted the following permits to operate the Maseve Mine.

>	Mining Right Number NW 30/5/1/1/2 528 granted on the 15 May 2012 for a period of 20 years which was registered in the Mineral Tiles Office on the 03 August 2012,
>	Environmental Management Plan and Environmental Impact Assessment approved by the DMR on the 15 May 2012,and authorised by DEDECT on the 13 September 2013,
>	Waste Management Licence Number 12/9/11/L628/7 granted by the DEA on the 06 November 2013, and
>	Water Use Licence Number 03/A22F/ABCGIJ/2596 granted by the DWS on the 16 July 2015.

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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Accessibility

South Africa has a large and well-developed mining industry. The Project is located in an area with a long history of mining activity and this, among other factors, means that the infrastructure in the area is well established, with well-maintained roads and highways as well as communication, electricity and water distribution networks.

The Project Area is located some 35km northwest of the town of Rustenburg. The town of Boshoek is situated 10km to the South along the paved regional highway R565 that links Rustenburg with Sun City and crosses the Project Area.

The WBJV properties are readily accessible from Johannesburg by travelling 120km NorthWest on Regional Road 24 (“R24”) to the town of Rustenburg and then a further 41km NorthWest along paved regional roads. The resort of Sun City is located approximately 7km NorthEast of Project Areas 1 and 1A. Both BRPM to the South of the Project Area and Styldrift, a joint venture between the Royal Bafokeng Nation (“RBN”) and AP, which lies directly to the East of the property, have modern access roads and services. Access across most of the property can be achieved by truck without the need for significant road building. Numerous gravel roads crossing the WBJV properties provide access to all portions.

The WBJV has sufficient surface rights under ownership for mining operations and mining personnel and the tailings area. Power and water are available and connected for the Project.

5.2 Climate

The Project area is considered semi-arid with an annual rainfall of 520mm. With low rainfall and high summer temperatures, the area is typical of the Highveld climatic zone. The rainy season is in the summer months from October to April with the highest rainfall in December and January. In summer (November to April), the days are warm to hot, with afternoon showers or thunderstorms; temperatures average 26ºC (79ºF) and can rise to 38ºC (100ºF), and night temperatures drop to around 15ºC (60ºF). During winter months (May to October), days are dry and sunny with moderate to cool temperatures, while evening temperatures drop sharply. Temperatures by day generally reach 20ºC (68ºF) and can drop to below 0ºC, with frost occurring in the early morning. The hottest months are generally December and January with June and July being the coldest. The climate of the area does not hinder the operating season, exploration and mining can continue all year long.

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Northerly and Easterly wind prevails over the study area with airflow infrequently occurring from the South-East and North-East. Diurnal shifts in the wind regime are apparent with night-time airflow, also being characterised by lower wind speeds and a greater frequency of calm periods. The highest wind speeds occur during spring months with a greater frequency of calms occurring in the winter and autumn seasons. Summer and spring is dominated mainly by Northerly and Easterly winds. During autumn and winter, South-Easterly winds along with Easterly winds prevail over the area. The average wind speed in the area is 3.5 – 4.5m/s.

Table 5-1: Meteorological Data from Rustenburg and Pilanesberg Meteorological Stations for period 2001 – 2010

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5.3 Local Resources

The closest major population centre to the project is the town of Rustenburg, located about 41km to the South-East of the project. Rustenburg is an established platinum mining centre with good parts and contractor availability. Pretoria lies approximately 100km to the East and Johannesburg about 120km to the South-East. A popular and large hotel and entertainment centre, Sun City, lies about 7km to the North-East of the Project Area. The Sundown Ranch Hotel which was purchased by Maseve, lies in close proximity to the Project Area and offers rooms and chalets as accommodation. The WBJV Project 1 and 1A Mine (Maseve) properties fall under the jurisdiction of the Moses Kotane Municipality.

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Table 5-2: Local Community and Town Distances in Relation to Project Site

Town / Community	Distance (km)	Direction from Project Site
Rustenburg	35	South East
Sun City	7	North North East
Phatsima	6.4	North West
Boschoek	6.2	South South East
Ledig	3.7	North North West
Rasimone	5.1	South East
Chaneng	3.3	East
Robega	2.9	East
Mafenya	2.3	South East

5.4 Regional Infrastructure

The WBJV is located in the heart of the Bushveld mining district close to major towns and settlements as a potential source of labour. Paved roads are abundant in the area for the transport of concentrate from surrounding operations to smelters. Power lines cross both Project areas. As several platinum mines are located adjacent to, and within 50km of the property, there is excellent access to materials and skilled labour. The nearest smelter complex of Anglo Platinum, to which the concentrate will be trucked, is located 50km SouthEast of the property at Rustenburg.

5.4.1 Power – Eskom Substation Completed on Site

The following power supplies are currently installed:

>	1.5MVA Eskom temporary power,
>	2.25MVA diesel generated power,
>	2 x 10MVA Eskom grid supplies.

The following supplies are planned for installation in the short term:

> 2.25MVA additional diesel generator power (October 2015),

The following supplies are planned for installation in the short to medium term:

>	Eskom is currently constructing Ngwedi substation to firm up the regional transmission network. (March 2016),
>	2 x 40MVA Eskom grid supplies will be phased into the Maseve supply and the 10MVA supplies removed. This will allow Maseve to use a total of 40MVA if required with a firm supply. (July 2016).

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5.4.2 Water

Magalies Water is the owner of a water treatment works at Vaalkop Dam and an agreement is in place with Maseve as well as other mining projects in the region to supply bulk water for industrial use. Under the agreement the pipelines for mine supply in the area are currently under construction and funded by the mining interests. The existing regional pipeline infrastructure is being used to supply water for underground development prior to production. Sufficient amounts of water are available in the reservoirs with a planned water infrastructure construction divided between stakeholder companies in the area. Once complete the maintenance will be done by Magalies Water Board.

5.4.3 Roads

A number of tarred access roads have been constructed around the property including:

>	Upgrade of the main road access from R565,
>	North Mine access road,
>	Plant access road shared by the South Mine access road.

The main access roads into the mine site have been completed. Access is directly off the main paved highway. The R565 is a tarred two lane regional highway that runs through the property.

Secondary access roads are compacted gravel surface roads. Dust mitigation is affected with the use of a water bowser filled up at the pollution control dams, and a grader maintains a good surface.

5.5 Physiography

5.5.1 Topography

The WBJV area is located on a central plateau. The Project has prominent hills in the Northern most portions, but generally variations in topography are minor and limited to low, gently sloped hills.

5.5.2 Elevation

The Elandsfontein and Frischgewaagd properties gently dip in a North-Easterly direction towards a tributary of the Elands River. Elevations range from 1 080m above mean sea level (“AMSL”) towards the Elands River in the North, to 1 156m AMSL towards the Onderstepoort farm in the South-West, with an average of 1 100m AMSL. On the Onderstepoort property to the West of the Project, the site elevation is approximately 1 050m AMSL, with the highest point at 1 105m AMSL.

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5.5.3 Vegetation

The area is characterised by extensive savannah with vegetation consisting of grasses and shrubs with few trees. Stony Clay Thornveld North of the Elands River, Black Turf Thornveld South of the Elands River and the Riparian Woodland along the Elands River. Several rocky ridges and boulder-strewn areas are present in the Black Turf Thornveld, and these contribute significantly to local biodiversity.

The project area’s vegetation is classified as Mixed Bushveld. The soil is mostly coarse, sandy and shallow, and overlies granite, quartzite, sandstone or shale, the vegetation varies from a dense, short bushveld to a fairly open tree savannah. On shallow soils Red Bushwillow Combretum apiculatum dominates the vegetation. Other trees and shrubs include Common Hook-thorn Acacia caffra, Sicklebush Dichrostachys cinerea, Live-long, Lannea discolor, Sclerocarya birrea and various Grewia species. Here the grazing is sweet, and grasses such as Fingergrass Digitaria eriantha, Kalahari Sand Quick Schmidtia pappophoroides, Wool Grass Anthephora pubescens, Stipagrostis uniplumis, and various Aristida and Eragrostis species dominate the herbaceous layer. On deeper and more sandy soils, Silver Clusterleaf Terminalia sericea becomes dominant, with Peeling Plane Ochna pulchra, Wild Raisin Grewia flava, Peltophorum africanum and Burkea Africana often prominent woody species, while Broom Grass Eragrostis pallens and Purple Spike Cat’s tail Perotis patens are characteristically present in the scanty grass sward.

Specifically, the project area is located in the Clay Thorn Bushveld – Bredenkamp and Van Rooyen (1996) – vegetation type in the Savannah Biome – Rutherford and Westfall (1994). The vegetation of the Eastern section of Elandsfontein is dominated by Acacia tortilis vegetation, which is typical of Clay Thorn Bushveld, with other species such as Rhus lancea, Ziziphus mucronata and Rhus pyroides adding to the species richness. The closed woodland areas occur along the main road where cattle kraals are located as well as along the drainage line. Some fallow lands occur in this area where a good grass layer dominated by species such as Themeda triandra, Cymbopogon contortis, Botriochloa bladhii and Sorghum versicolor has re-established as well as a sparse tree layer. The areas on the Western section of Elandsfontein consist of a fenced private game reserve as well as a natural area further to the North near the Elands River. The tree and herbaceous layer is more diverse in this area where the tree layer is dominated by Ziziphus mucronata, Acacia tortilis and the shrub Grewia flava. Table 5-3 shows flora species in this vegetation type (* exotic species).

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Table 5-3: Vegetation Species Evident in Project Area

Forbs	Grasses	Trees and Shrubs
Tephrosia capensis	Themeda triandra	Euclea divinorum
Commelina erecta	Eragrostis superba	Diospyros lycioides
Crotolaria eremicola	Cymbopogon plurinodus	Ziziphus mucronata
Chamaecrista comosa	Sorghum versicolor D	Vangueria infausta
Ruellia patula	Urelytrum agropyroides	Rhus lancea
Clematis brachiata	Aristida bipartita	Grewia flava
Eriosema cordatum	Pennisetum thunbergii	Acacia tortilis
Tithonia rotund i'folia*	Heteropogon contortis	Carissa bispinosa
Tagetes minuta*	Melinis repens	Olea europaea
Datura stramonium*	Panicum deustum	Asparagus laricinus
Schkuria pinnata*	Digitaria eriantha	Rhus pyroides
Vernonia oligocephala	Eragrostis rigidior	Acacia karroo
Sebaea grandis	Choris virgata	Gymnosporia buxifolia
Crabbea angustifolia	Urochloa mosambicensis	D Acacia karroo
Rhynchosia minima	Setaria sphacelata	Ziziphus mucronata
Tephrosia capensis	Eragrostis capensis	Rhus lancea
Hybiscus trionum	Cymbopogon excavatus	Rhus pyroides
Commelina erecta Cymbopogon sp.	Botriochloa insculpta	Dombeya rotundifolia
Tithonia rotundi'folia*	Urelytrum agropyroides	Grewia flava
Berkheya radula	Cymbopogon excavatus	Euclea divinorum
Pavonia burchelii	Botriochloa insculpta D	Diospyros lycioides

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Forbs	Grasses	Trees and Shrubs
Crinum bulbispermum	Botriochloa bladhii	Ziziphus mucronata
	Themeda triandra	Grewia flava
	Sorghum versicolor	Rhus lancea
	Echinochloa colona D
	Brachiaria brizantha D
	Paspalum urvillei

The area is typical of rural areas with varying degrees of disturbance evident throughout the study area. Existing impacts at the site include:

>	Collection of wood for fuel,
>	Poaching,
>	Roads,
>	Illegal waste dumps, especially in old quarries,
>	Current cultivation and livestock grazing,
>	Telecommunication and power lines,
>	Telecommunication and radar towers,
>	Quarries,
>	Rural infrastructure.

5.5.4 Fauna

The project area consists of natural habitats with operational ecosystems despite areas of disturbance within these habitats. No habitat of exceptional sensitivity or concern exists.

5.5.5 Birds

Approximately one third (328 species) of the roughly 900 bird species of South Africa occur in the Rustenburg / Pilanesberg area. The most characteristic of these include lilac-breasted rollers, African hoopoes and owls. The Red Data bird species that occur (*) or could potentially occur (**) in the study area are listed in Table 5-4.

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Table 5-4: Bird Species (Red Data) that may occur in Project Area

Species occurring in the area*	Habitat
Martial Eagle (V)	Tolerates a wide range of vegetation types found in open grassland, shrub, Karoo and woodland.
Species potentially occurring in the area**
African Whitebacked Vulture (V)	Nests in large trees, transmission and reticulation power lines.
Tawny eagle (V) areas.	Occurs mainly in woodlands as well as lightly wooded
Blue Crane (V)	Dry short grassland. Not very dependent on wetlands habitat for breeding. Preferred nesting sites are secluded open grasslands as well as agricultural fields.
Grass Owl (V)	Breeding in permanent and seasonal vleis. Vacates while hunting or post breeding.
Red Data status = V = vulnerable.

5.5.6 Herpetofauna

In total, 143 species of herpetofauna occur in the North West Province. This is considered high as it accounts for roughly one third of the total occurring in South Africa. Monitor lizards and certain snake and gecko species are found in the project area. Table 5-5 below shows Red Data species in the North West Province.

Table 5-5: Snakes (Red Data) that Potentially occur in Project Area

Scientific name	English name	Conservation Status
Python natalensis	Southern African Python	Vulnerable
Homoroselaps dorsalis	Striped Harlequin Snake	Rare
Dalophia pistillum	Blunt-tailed worm lizard	Data Deficient
Crocodylus niloticus	Nile crocodile	Vulnerable
Pyxicephalus adspersus	Giant Bullfrog	Near Threatened

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Habitats for all the above-named species, excluding the Nile crocodile, occur in the project area with the wetland patches along the stream potentially suitable habitats for Giant Bullfrog.

5.5.7 Mammals

The Southern Greater Kudu found in North West Province are among the biggest in the country. In the project area it is expected that larger antelope such as gemsbok, Cape eland, common waterbuck, impala, and red hartebeest may be kept on the farms, while smaller cats, viveriids, honey badgers and vervet monkeys should occur as free roaming game. The project area could potentially be a habitat for the following Red Data species.

Table 5-6: Mammals that Potentially occur in Project Area

Scientific name	English Name	Conservation Status
Atelerix frontalis	South African hedgehog	Rare
Proteles cristatus	Aardwolf	Rare
Hyaena brunnea	Brown hyena	Rare
Panthera pardus	Leopard	Rare
Mellivora capensis	Honey badger	Vulnerable

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6.0 HISTORY

6.1 Prior Ownership

Elandsfontein (PTM), Onderstepoort (Portions 4, 5 and 6), Onderstepoort (Portions 3 and 8) and Onderstepoort (Portions 14 and 15) were previously all privately owned. Previous work done on these properties has not been fully researched and is largely unpublished. Academic work such as that carried out by the Council for Geoscience (South African government agency) is generally not of an economic nature.

Elandsfontein (RPM), Frischgewaagd, Onderstepoort (RPM) and Koedoesfontein have been held by major mining groups resident in the Republic of South Africa. Portions of Frischgewaagd previously held by Impala Platinum Mines Limited were acquired by Johannesburg Consolidated Investment Company Limited, which in turn have since been acquired by AP through RPM and now contribute to the WBJV.

6.2 Work Done by Previous Owners

Previous geological exploration and mineral resource estimation assessments were done by AP as the original owner of some of the mineral rights. AP managed the exploration drilling programme for the Elandsfontein and Frischgewaagd drill hole series in the area of interest. Historical geological and sampling logs as well as the assay database were used by PTM to assess the geology of the area.

Prior to the establishment of the WBJV and commencement of drilling for the Pre-Feasibility study, PTM had drilled 36 drill holes on the Elandsfontein property.

Existing gravity and ground magnetic survey data were helpful in the interpretation of the regional and local geological setting of the reefs. A distinct increase in gravity values occurs from the southwest to the northwest, most probably reflecting the thickening of the Bushveld sequence in that direction. Low gravity trends in a south-eastern to north-western direction. The magnetic survey reflects the magnetite-rich Main Zone and some fault displacements and late-stage intrusive in the area.

6.3 Historical Mineral Reserves and Resources

Historical Mineral Resources for the Maseve Project area were estimated by Mr CJ Muller of Minxcon. These mineral resources are detailed in the Technical Report dated October 2009 (available at www.sedar.com). The mineral resources were estimated utilising the definitions set out in NI 43-101 and the SAMREC Code (2007). Table 6-1 summarises the historical mineral resources for Project 1 and 1A.

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Table 6-1: Historical Mineral Resources Estimated for Project 1 and 1A – October 2009 (100% WBJV Area)

Measured Mineral Resource (4E)	Cut-off	Tonnage	Grade 4E	Mining Width	Content 4E	Content 4E
	(cm.g/t)	(Mt)	(g/t)	(m)	(t)	(Moz)
Project 1 MR	300	6.603	8.38	1.33	55.333	1.779
Project 1 UG2	300	7.464	4.26	1.34	31.797	1.022
Total Measured	300	14.067	6.19	1.34	87.130	2.801

Prill Splits	Pt	Pt	Pd	Pd	Rh	Rh	Au	Au
	%	(g/t)	%	(g/t)	%	(g/t)	%	(g/t)
Project 1 MR	64%	5.36	27%	2.26	4%	0.34	5%	0.42
Project 1 UG2	63%	2.68	26%	1.11	10%	0.43	1%	0.04

Indicated Mineral Resource (4E)	Cut-off	Tonnage	Grade 4E	Mining Width	Content 4E	Content 4E
	(cm.g/t)	(Mt)	(g/t)	(m)	(t)	(Moz)
Project 1 & 1A MR	300	11.183	7.25	1.24	81.077	2.607
Project 1 & 1A UG2	300	19.209	4.46	1.39	85.672	2.754
Total Indicated	300	30.392	5.49	1.34	166.749	5.361

Prill Splits	Pt	Pt	Pd	Pd	Rh	Rh	Au	Au
	%	(g/t)	%	(g/t)	%	(g/t)	%	(g/t)
Project 1 MR	64%	4.46	27%	1.96	4%	0.29	5%	0.36
Project 1 UG2	63%	2.81	26%	1.16	10%	0.45	1%	0.04

Inferred Mineral Resource (4E)	Cut-off	Tonnage	Grade 4E	Mining Width	Content 4E	Content 4E
	(cm.g/t)	(Mt)	(g/t)	(m)	(t)	(Moz)
Project 1 & 1A MR	300	0.154	8.96	1.06	1.380	0.044
Project 1 & 1A UG2	300	0.022	3.91	0.83	0.086	0.003
Total Inferred	300	0.176	8.33	1.03	1.466	0.047

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Prill Splits	Pt	Pt	Pd	Pd	Rh	Rh	Au	Au
	%	(g/t)	%	(g/t)	%	(g/t)	%	(g/t)
Project 1 MR	64%	5.73	27%	2.42	4%	0.36	5%	0.45
Project 1 UG2	63%	2.46	26%	1.02	10%	0.39	1%	0.04

Resource Statement:

MR = Merensky Reef; UG2 = Upper Group No. 2 Chromitite seam; Conversion Factor used – kg to oz = 32.15076.

The cut-offs for mineral resources were established by a qualified person after a review of potential operating costs and other factors. PTM owns 37% of the WBJV. The mineral resources stated above are shown on a 100% basis, that is, for the WBJV as a whole entity. The mineral resource estimate was based on available drill hole results over the history of the Project Areas, including drill holes results obtained in 2009. The grade models were constructed from Ordinary Kriged and Simple kriged estimates. The MR was kriged over a minimum mining width of 0.80m and includes footwall mineralisation that is above 2g/t, within the first 60cm below the reef. The UG2 was kriged over a minimum mining width of 0.80m. The blocks were classified into measured, indicated and inferred resource categories using the following and not limited thereto: sampling quality assurance and quality control, geological confidence, number of samples used to inform a block, kriging variance, distance to sample (variogram range), lower confidence limit, kriging efficiency, regression slope, etc. The mineral resource is reported as inclusive of mineral reserves. No environmental, permitting, legal, taxation, socio-political, marketing or other issues are expected to materially affect the above mineral resource estimate and hence were not used to modify the mineral resource estimate.

6.4 Production from the Property

There has been no previous production from any of the WBJV properties.

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7.0 GEOLOGICAL SETTING AND MINERALISATION

7.1 Regional Geology of the BIC

The WBJV Project 1 (Maseve) project area is partly situated on the South-Western limb of the layered igneous complex known as the BIC and its surrounding sedimentary footwall rocks. The BIC is unique and well known for its layering and continuity of economic horizons mined for platinum, palladium and other PGE’s, chrome and vanadium.

The BIC is a large mafic to ultramafic layered complex, which was intruded into the rocks of the Transvaal Supergroup, approximately 2 060 million years ago. The mafic component of the BIC is collectively termed the Rustenburg Layered Suite (“RLS”) and can be divided into five zones, namely the Marginal, Lower, Critical (which hosts the MR and UG2 Reefs), Main and Upper Zones.

Regarding near surface geology of the Project Area, the Northern region of the Project Area is underlain by Pyramid Gabbro-Norite of the Main Zone of the RLS. Further South, underlying the Main Zone is the Mathlagama Norite-Anorthosite containing the MR and the upper (UG2) and middle Chromitite layers. The Critical Zone is underlain by the Kolobeng Norite of the Marginal Zone.

The Project Area is bounded in the West by the Rustenburg Fault and one of its NorthWest-SouthEast orientated branches can be seen to transect the central portion of the Project Area.

PGE mineralisation is hosted within the MR and UG2 Chromitite, located in the Critical Zone of the Rustenburg Layered Suite of the BIC. The MR and UG2 Chromitite strike NorthNorthWest-SouthSouthEast for approximately 7km within the Project area.

The MR is a well-developed seam, striking NorthWest-SouthEast and extending from the outcrop along the central part of the WBJV property and dipping North-East. This reef contains the most pronounced mineralisation associated with a 0.1 to 1.2m thick feldspathic pegmatoidal pyroxenite unit. The MR is generally also associated with thin Chromitite layers on either or both of the top and bottom contacts of the pyroxenite.

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The UG2 chromite layer is well developed towards the North-East of the Project Area, but deteriorates towards the South-West, where the reef is generally thin and discontinuous. In the North-West of the Project Area, the reef is generally well developed and occurs as a single prominent Chromitite layer varying in thickness from a few centimetres to ~2m.

The initial occurrence of payable reef is at 130m below surface. The payable reef then extends to a depth of approximately 630m below surface. There are no near-surface or outcropping reefs of sufficient width, geometry or value to provide an opportunity for opencast mining. The middling between the MR and UG2 Reefs increases from approximately 10m to as much as 80m towards the North-East. The MR and UG2 reefs, in the Maseve Project area, dip between 4° and 42°, averaging 14°.

A structural model was developed from geophysical information provided by the magnetic survey results, LANDSAT, aerial photography and the geological logs of drilled cores. At least three structural orientations have been identified. The SouthEast-NorthWest trend is predominantly down-thrown towards the West. The majority are normal faults, displaying listric characteristics.

The Rustenburg Fault trending SouthEast-NorthWest, exhibits down-throw to the East. Three major disruptions occur in an approximately East to West direction, namely the Chaneng Dyke and the Elands and Caldera faults.

The detailed magnetics data indicates a Northerly dip on the Chaneng Dyke, but appears to throw down towards the South. The dips on these Easterly trending structures are close to vertical. The Eland and Caldera Faults both throw down towards the North. Aerial photography of the Pilanesberg Complex into the WBJV Project 1 (Maseve) area indicates at least one major North to South trending fault that crosses the entire Pilanesberg Complex.

Several dolerite intrusives, mainly steeply-dipping dykes and bedding-parallel sills, were intersected in drill holes. These range in thickness from 0.5m to 30m. Most appear to be of a chilled nature and some are associated with faulted contacts. Dykes of Pilanesberg age intruding into pre-existing fractures are the most prominent dyke swarm with a NorthWest-SouthEast orientation.

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Replacement bodies exist within the Elandsfontein property and seem to be concentrated in the lower part of the Main Zone and HW5 of the Critical Zone. Reef packages to the South in the Elandsfontein (PTM) area are marginally affected. This has been taken into consideration in the resource estimation and geological losses for both MR and UG2 Chromitite seam.

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7.2 Local Geology – Western Bushveld Limb

Exposures of the BIC located on the Western Limb include the stratigraphic units of the RLS. The sequence comprises mostly gabbros, norites, anorthosites and pyroxenites. Viljoen (1999) originally proposed a structural interpretation based on geological and geophysical data for the Western lobe of the BIC. This study included gravity and vibrosis seismic data for the South-Western portion of the RLS NorthWest of Rustenburg (including the Boshoek section). It was concluded that the MR is present within much of this lobe, including the part further to the East below the Nebo granite sheet. The position of the MR is fairly closely defined by seismic reflectors associated with the cyclic units of the UCZ. The seismic data also portrayed an essentially sub-horizontal disposition of the layering within the BIC mafic rocks below the Nebo granite sheet. The gravity data indicates a gravity-high axis extending throughout the Western lobe following the upper contact of the mafic rocks with the overlying granitic rocks. A number of pronounced gravity highs occur on this axis. A gravity anomaly with a strike length of 9km is situated northeast of Rustenburg towards the East of the Boshoek section. The gravity highs have been interpreted as representing a thickening of the mafic rocks, reflecting feeder sites for the mafic magma of the Western BIC (Viljoen, 1999).

The Western Limb is interpreted by Viljoen as having two main arcuate feeder dykes which, closer to the surface, have given rise to arcuate, coalescing, boat-shaped keels containing saucer-shaped, inward-dipping layers, analogous to the Great Dyke of Zimbabwe.

In the Boshoek Section North of Rustenburg, the variable palaeo-topography of the Bushveld floor represented by the Transvaal Supergroup contact forms a natural unconformity with the overlying Bushveld layered sequence. Discontinuities due to structural interference of faults, sills and dykes are pronounced in the area and are ascribed to the presence of the Pilanesberg Alkaline Complex intrusion to the North of the property.

7.2.1 Stratigraphy of the UCZ

The UCZ of the RLS comprises mostly norites, leuconorites and anorthosites. Leeb-Du Toit (1986) assigned numbers to the various lithological units according to their position in relation to the Merensky Unit. The footwall (“FW”) layers range from FW14 below the UG1 Chromitite to FW1 directly below the MR. The hanging wall (“HW”) layers are those above the Bastard Reef and range from HW1 to HW5. The different layers within the Merensky unit are the Merensky feldspathic pyroxenite at the base, followed by a leuconorite (Middling 2) and a mottled anorthosite (Middling 3). The feldspathic pyroxenite layers (pyroxene cumulates) are named according to the reef hosted by them. These include (from the base upwards) the UG1, the UG2 (upper and lower), the Merensky and the Bastard pyroxenite.

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Schürmann (1993) subdivided the UCZ in the Boshoek section into six units based on lithological features and geochemical trends. These are the Bastard, the MR, the MR footwall, the Intermediate, the UG2 and the UG1 units. The Intermediate and MR footwall units were further subdivided based on modal-mineral proportions and whole-rock geochemical trends. The following is a detailed description of the subdivision of the UCZ in the Boshoek section (Schürmann).

7.2.2 Bastard Unit

The Bastard unit consists of a basal pyroxenite some 3m thick with a thin Chromitite developed on the lower contact. This Chromitite is the uppermost Chromitite layer in the CZ. A 6.5m -thick norite layer (HW1) overlies the pyroxenite. HW1 is separated from HW2 by two thin mottled anorthosite layers. HW3 is a 10m-thick mottled anorthosite and constitutes the base of the Giant Mottled Anorthosite. The mottled anorthosites of HW4 and HW5 are about 2m and 37m thick respectively. Distinction between HW3, 4 and 5 is based on the size of the mottles of the respective layers.

7.2.3 Merensky Unit

The Merensky unit, with the MR at its base, is the most consistent unit within the CZ.

7.2.4 Merensky Footwall Unit

This unit contains the succession between the FW7/FW6 and the FW1/MR contacts. Leeb-Du Toit (1986) indicated that where the FW6 layer is thicker than 3m, it usually consists of four well-defined rock types. The lowermost sublayer, FW6 (d), is a mottled anorthosite with mottles of between 30mm and 40mm in diameter. It is characterised by the presence of nodules or “boulders” and is commonly referred to as the Boulder Bed. The nodules are described as muffin-shaped, 5–25cm in diameter, with convex lower contacts and consisting of cumulus olivine and orthopyroxene with intercumulus plagioclase. A single 2–10mm Chromitite stringer is present at the base of the FW6 (d) sublayer. FW6 (c) is also a mottled anorthosite but not always developed. FW6 (b) is a leuconorite containing pyroxene oikocrysts 10–20mm in diameter. Two layers (both 2–3cm thick) consisting of fine-grained orthopyroxene and minor olivine define the upper and lower contacts. FW6 (a), the uppermost sublayer, is also a mottled anorthosite.

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FW6 is overlain by a uniform norite (FW5), with a thickness of 4.1m. It appears to thin towards the North to about 1m. FW4 is a mottled anorthosite 40cm thick, with distinct layering at its base. FW3 is an 11m-thick uniform leuconorite.

FW2 is subdivided into three sublayers. FW2 (b) is a 76cm-thick leuconorite and is overlain by a 33cm-thick layer of mottled anorthosite – FW2 (a). Where FW2 attains a maximum thickness of 2m, a third layer in the form of a 1–2cm-thick pyroxenite or pegmatite pyroxenite, FW2 (c), is developed at the base. FW2 (c) is absent in the Boshoek section area (Schürmann, 1993). FW1 is a norite layer about 7m thick.

Schürmann further subdivided the Merensky footwall unit into four subunits. The lowermost subunit consists of sublayers FW6 (d) and FW6 (b). Subunit 2, which overlies subunit 1, commences with FW6 (a) at the base and grades upwards into FW5. The FW5/FW4 contact is sharp and divides subunits 2 and 3. Subunit 3 consists of FW4, FW3 and sublayer FW2 (b). Subunit 4 consists of FW2 (a) and FW1 and forms the uppermost subunit of the MR footwall unit.

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7.2.5 Intermediate Unit

The Intermediate unit overlies the upper pyroxenite of the UG2 unit and extends to the FW7/FW6 contact. The lowermost unit is the 10m-thick mottled anorthosite of FW12, which overlies the UG2 upper pyroxenite with a sharp contact. FW11, a roughly 1m thick leuconorite, has gradational contacts with the under- and overlying layers. FW10 consists of a leuconorite layer of about 10m. Subdivision between these two units is based on the texture and subtle differences in the modal composition of the individual layers. Leeb-Du Toit (1986) termed FW11 a spotted anorthosite and FW10 an anorthositic norite. FW12, 11 and 10 constitute the first Intermediate subunit as identified by Schürmann (1993). The second Intermediate subunit consists of FW9, 8 and 7. The 2m-thick FW9 mottled anorthosite overlies the FW10 leuconorite with a sharp contact. The FW8 leuconorite and FW7 norite are respectively 3m and 37m thick. The FW9/FW8 and FW8/FW7 contacts are gradational but distinct. A 1.5m -thick highly contorted mottled anorthosite “flame bed” is present 15m above the FW8/FW7 contact.

7.2.6 UG2 Unit

The UG2 unit commences with a feldspathic pyroxenite (about 4m thick) at its base and is overlain by an orthopyroxene pegmatoidal layer (0.2 –2m thick) with a sharp contact. Disseminated chromite and Chromitite stringers are present within the pegmatoid. This unit in turn is overlain by the UG2 Chromitite (0.5 –0.8m thick) on an irregular contact. Poikilitic bronzite grains give the Chromitite layer a spotted appearance. A 9m feldspathic pyroxenite overlies the UG2 Chromitite. The upper and lower UG2 pyroxenites have sharp contacts with FW12 and FW13. The upper UG2 pyroxenite hosts the UG2 Leader seams, which occur between 0.2m and 3m above the main UG2 Chromitite.

7.2.7 UG1 Unit

The UG1 Chromitite layer is approximately 1m thick and forms the base of this unit. It is underlain by the 10m-thick FW14 mottled anorthosite. The UG1 Chromitite layer bifurcates and forms two or more layers within the footwall mottled anorthosite, while lenses of anorthosite also occur within the Chromitite layers. The overlying pyroxenite consists of cumulus orthopyroxene, oikocrysts of clinopyroxene and intercumulus plagioclase. The UG1 pyroxenite is separated from the overlying FW13 leuconorite (about 8m thick) by a thin Chromitite layer (1–10cm) with sharp top and bottom contacts.

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7.3 Local Structure

Floor rocks in the South-Western BIC display increasingly varied degrees of deformation towards the contact with the RLS. Structure within the floor rocks is dominated by the North-NorthWest trending post-Bushveld Rustenburg Fault. This normal fault with down-throw to the East extends Northwards towards the West of the Pilanesberg Alkaline Complex. A second set of smaller faults and joints, striking 70° and dipping very steeply South-SouthEast or North-NorthWest, are related to the Rustenburg Fault system. These structures were reactivated during the intrusion of the Pilanesberg Alkaline Complex. Dykes associated with the BIC intruded along these faults and joints.

Major structures, which occur within the WBJV area, include the Caldera and Elands Faults, Chaneng Dyke and a major North-South trending feature, which can be observed across the entire Pilanesberg Complex. These East-West trending structures dip steeply (between 80° and 90°). The magnetics indicate that the Chaneng Dyke dips steeply to the North. This is consistent with similar structures intersected underground on the neighbouring BRPM, which all dip steeply Northward.

Two stages of folding have been recognised within the area. The earliest folds are mainly confined to the Magaliesberg Quartzite Formation. The fold axes are parallel to the contact between the RLS and the Magaliesberg Formation. Quartzite xenoliths are present close to the contact with the RLS and the sedimentary floor. Examples of folding within the floor rocks are the Boekenhoutfontein, Rietvlei and Olifantsnek anticlines. The folding was initiated by compressional stresses generated by isostatic subsidence of the Transvaal Supergroup during sedimentation and the emplacement of the pre-Bushveld sills. The presence of an undulating contact between the floor rocks and the RLS, and in this instance the resultant formation of large-scale folds, substantiates a second stage of deformation. The fold axes trend at approximately orthogonal angles to the first folding event. Deformation during emplacement of the BIC was largely ductile and led to the formation of basins by sagging and folding of the floor rocks. This exerted a strong influence on the subsequent evolution of the LZ and CZ and associated Chromitite layers.

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The structural events that influenced the floor rocks played a major role during emplacement of the BIC. There is a distinct thinning of specific stratigraphic units from East to West as the BIC onlaps onto the Transvaal floor rocks, even to the extent that some of the normal stratigraphic units have been eliminated.

7.4 Property Geology

The sequence of the BIC within the WBJV area is confined to the lower part of the Main Zone (Porphyritic Gabbro Marker) and the Critical Zone (HW5–1 and Bastard Reef to UG1 footwall sequence). The rock sequence thins towards the SouthWest (subcrop) including the marker horizons with concomitant middling of the economic reefs or total elimination thereof. The UG2 and, more often, the UG1 are not developed in some areas owing to the irregular and elevated palaeo-floor of the Transvaal sediments.

7.4.1 Surface Geology

The WBJV is underlain by the lower portion of the RLS, comprising Critical Zone and lower portion of the Main Zone. The ultramafic Lower Critical Zone and the Mafic Upper Critical Zone and the Main Zone weather to dark, black clays with very little topography. The underlying Transvaal Supergroup comprises shale and quartzite of the Magaliesberg Formation, which creates a more undulating topography. Gravity, magnetic, LANDSAT, aerial photography and geochemistry have been used to map out lithological units.

The MR and UG2, in the Maseve Project area and 1A area, dip between 4° and 42°, averaging 22°.

7.4.2 Project Structure

A structural model was developed from data provided by the magnetic survey results and geological logs of drilled cores. Several generations of faults were identified on the property, however these were reduced somewhat from previous interpretations by defining the slope / terrace area close to outcrop more strictly.

The majority of the faults now trend 315 degrees Northwards respectively and appear to have consistent down-throws towards the West.

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A second phase of deformation may be related to a regional East-West striking dyke system causing discontinuity on adjacent structures, this is noted particularly on the Chaneng Dyke which almost bisects the project area. Several dolerite intrusives, mainly steep-dipping dykes and bedding-parallel sills, were intersected in boreholes. These range in thickness from 0.5 –30m and most appear to be of a chilled nature; some are associated with faulted contacts. Evident on the magnetic image is an east-west trending dyke, which appears to be of Pilanesberg-intrusion age. This dyke has a buffer effect on structural continuity as faulting and earlier stage intrusives are difficult to correlate on either side; and more work is required to understand the mechanics.

A reverse fault intersected in several drill holes was also included in this interpretation causing duplication of both Merensky and UG2 units. Figure 7-4 shows the major faults and dykes developed at the project area.

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The project area is divided into three main structural blocks (Figure 7-5) with the Chaneng dyke demarcating the boundary between structural Block 1 and structural Block 2. Each of these structural blocks is characterised by specific geological relationships and mineralisation patterns.

7.4.3 Crystallisation and Setting Zones

Figure 7-6 shows the location of different crystallisation and settling zones associated with the basin edge. The area marked “Basin Edge” represents the shallow edge where most of the Critical Zone stratigraphy has been eliminated or not developed. The Main Zone has eroded most of the underlying units in this area and there are only pockets of MR and UG2 present.

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The terrace / slope area represents an area where stratigraphic units have been thermally eroded, slumped down where palaeo dips were more than 7°, resulting in no reef development or isolated pockets of reef on the upper part of slopes and thickening of units at base of slopes with reefs well developed where the slope evens out. On these slopes, local or smaller potholes are typically formed. The normal basin area (including the horst or uplifted block) is where most of the stratigraphic units are developed with only minor erosion of specific units. The MR and UG2 reef units can be correlated over large distances with expected stratigraphic units with a stratigraphic thickness of more than 10m increasing towards the East.

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Figure 7-7 shows typical sections of the basin edge, terrace slope and normal basin area with expected geological relationships. Section A-A’ is to the North in structural block 1 and shows normal basin relationship where MR and UG2 stratigraphic thickness is constant and most of the stratigraphic units developed. The area has been subjected to later stage faulting resulting in specific smaller structural blocks with specific dip relationships. The dome structure is two faulted blocks and a paleo-high resulting in a dome type structure. Sections B-B’ and C-C’ sections of the middle portion of the project through are the basin edge and terraces slope areas. In the Western basin edge area the Main Zone has thermally eroded the lower units including the MR and in places the UG2 reef horizons. Section D-D’ is in the Southern portion of the project in structural block 3 and shows the normal basin relationships with no definite terrace and slope relationships.

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7.4.4 Facies Model

The most pronounced PGE mineralisation in the Project Area occurs within the MR and is generally associated with a 0.1 – 1.2m -thick feldspathic pegmatoidal pyroxenite unit. The MR is generally also associated with thin Chromitite layers on either/both the top and bottom contacts of the feldspathic pegmatoidal pyroxenite. The second important mineralised unit is the UG2 Chromitite layer, which is on average 1.28m thick and occurs within the Project Area.

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>	Regional and Local Facies Delineation,
>	MR Facies Types.

Within the Project Area, the emplacement of the MR is firstly controlled by the presence or absence of Chromitite seams and secondly by footwall stratigraphic units. The MR may be present immediately above either the FW3 or FW6 unit. This has given rise to the terms Abutment Terrace (FW3 thermal erosional level), Mid Terrace (FW3 or FW6 thermal erosional levels) and Deep Terrace (FW6 thermal erosional level). Within, and not necessarily confined to, each of the terraces, the morphology of the MR can change.

From logging and sampling information of holes on the WBJV property, it is evident that the footwall mineralization of MR below the main Chromitite layer occurs in reconstituted norite, which is the result of a high thermal gradient at the base of the mineralising Merensky cyclic unit. The upper Chromitite seam may form an upper thermal unconformity. Footwall control with respect to mineralization is in many cases more dominant than the actual facies (e.g. the presence of leucocratic footwall units) or a Chromitite (often with some pegmatoidal pyroxenite).

The specific facies on the project area fits well with the regional facies delineated by AP on the adjacent properties. The larger portion of the Western side of the project relates to the so-called “pothole facies” on the adjacent property. This pothole facies is a regional pothole area with typical basin edge and terrace slope areas, especially on the Western side, with larger portions of this facies related to normal basin relationships. The term pothole is therefore related to larger features of thermal erosion of stratigraphic units as is typically found on the Western portion of this facies or project area. For this project the facies are termed FPP facies (feldspathic pegmatoidal pyroxenite) and is associated with thin Chromitite layers on either or both of the top and bottom contacts of the pyroxenite, pegmatoidal or feldsphatic pegmatoidal pyroxenite unit.

Figure 7-8 shows the specific facies on the property. The FPP facies has been divided into “FPP Terrace Slope” and “FPP Normal Basin” facies.

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There are a number of different combinations or sub types of MR that can be found within the FPP facies type.

Sub Type A relates to the interface between the normal hanging wall of the MR and the footwall of the MR. There is no obvious chromite contact or any development of the normal feldspathic pegmatoidal pyroxenite. This may well be classified as hanging wall on footwall, but normally has a PGE value within the pyroxenite. Sub Type B is typified by the presence of a chromite seam, which separates the hanging wall pyroxenite from the footwall (which could be the FW3 or FW6 unit). Sub Type C can be found on any of the three terraces and has a characteristic top chromite seam overlying a pegmatoidal feldspathic pyroxenite. This facies has NO bottom chromite seam. Sub Type D is traditionally known throughout the BIC as Normal MR and has top and bottom chromite seams straddling the pegmatoidal feldspathic pyroxenite.

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The delineated IRUP areas link up with the abutment facies of the adjacent properties which consist of a mix of IRUP’s and poorly developed reefs areas.

On the Western side of the project is an additional facies termed “Contact Facies” and represents the basin edge where the stratigraphic units are drastically thinner and eliminated in places by thermal erosion, pothole effects, not developed and various other processes. The basin floor forms the controlling feature with a steep slope at the edge and in places creating slope and terraces before it evens out into the proper basin with fully developed stratigraphic units.

7.4.5 UG2 Facies Types

The facies model for the UG2 Reef has been developed mainly from drill hole exposure data in the North-East of the property. The integrity of the UG2 deteriorates towards the SouthWest of the Project Area, where it occurs as a thin chromite layer and/or pyroxenitic unit. It is thus unsuitable for the development of a reliable geological facies model. In the North-East of the Project Area, the UG2 is relatively well developed and usually has three thin chromite seams (Leaders) developed above the main seam. The UG2 Reef facies can also be explained in terms of four distinct facies types (Figure 7-10). Several factors appear to control the development of the UG2 package. Of these, the Digital Terrain Model (“DTM”) of the Transvaal Basement is likely to have the most significant impact. The distinct variance in the various facies is seen as directly related to the increasing isopach distance between the UG2 and MR. In this regard, the facies-types for the UG2 have been subdivided into the Abutment terrace facies, mid-slope terrace facies and the deep-slope terrace facies. They are described as follows:

The Abutment Terrace facies was identified in the area where the basement floor was elevated, perhaps as a result of footwall upliftment or an original palaeo-high. In this area, it appears that there was insufficient remaining volume for the crystallisation and mineralization of PGE’s. A reduced lithological sequence and thinning-out of layering is evident in the facies domain/s. In this environment there is an irregular and relatively thin (5 – 20cm) UG2 main seam developed with no evidence suggesting the presence of harzburgite footwall. No Leaders are present and there is a distinct absence of the normal overlying FW8–12 sequence.

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The intermediate area between the Abutment terrace facies and the mid-slope terrace facies has no UG2 development. The footwall is usually a thin feldspathic pyroxenite transgressing downwards to a medium-grained FW13 norite. The hanging wall generally occurs as either/both the FW7 and FW8 norites.

The Mid- and Deep-Slope Terrace facies environments that form the central and Northern boundaries of the Project Area are characterised by a thicker to well-developed UG2 main seam of about 0.5m to more than 3m respectively. Here, as with the Abutment terrace facies, the development of a robust UG2 is dependent on the MR / UG2 isopach. These facies are characterised by the fact that all Leaders are exposed at all times and Leader 3 (UG2L3) occurs as a pencil-line chromite seam. A prominent development of a harzburgite FW unit (5 – 30cm) is often present in this facies type.

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7.4.6 Geological Domains

A combination of structural features and facies types for the MR and UG2 were used to delineate geological domains (Figure 7-11) which each have a specific geological character related to structure, stratigraphic units and mineralisation. There are four larger geological domains, the IRUP areas, Contact Facies, FPP terrace and slope facies and the FPP normal basin facies. These have been further sub divided according to specific geological relationships which add up to a total of 11 geological domains. The geological domains are a result of information from extensive drilling in the project area. There is a good understanding of the specific geological controls within each geological domain and it is possible to predict what to expect for each reef horizon. The project area can be broadly subdivided into a zone of more complex geology and an area of stable geological relationships. The Eastern portion of the project area is characterised by Contact Facies and the Terrace and Slope Facies where stratigraphic units have been thermally eroded and eliminated over large areas. The Western side of the project area is characterised by far less elimination of stratigraphic units and the MR and UG2 reefs are continuous and correlate over large distances with less structural complexity.

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7.4.7 Reefs

The MR is a well-developed seam in the WBJV Project 1 (Maseve) area and along the central part and towards the North Eastern boundary of Maseve Project area. Islands of thin reef and relatively low-level mineralisation are present. The better-developed reef package, in which the intensity of Chromitite is generally combined with feldspathic pegmatoidal pyroxenite development, occurs as larger island domains along a wide central strip in a North-South orientation from subcrop to deeper portions.

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The UG2 is well developed towards the North-East of the Project Area, but deteriorates towards the South-West. Within the latter area, the reef is present as a thin discontinuous or disrupted Chromitite / pyroxenite layer. It also appears to be disrupted by the shear zone along the footwall alteration zone. Towards the NorthWest on Frischgewaagd, the reef is generally well developed and occurs as a single prominent Chromitite layer varying in thickness from a few centimetres to ~2m.

The thickness of the sequences between the UG2 CL and MR in the Project area increases from ~10m to 80m in a SouthWest-NorthEast direction. A similar situation exists in the North of the Project Area but with the thickness between the reefs ranging from 6m to 25m at depths of 200m below surface. In general, the thickness between the reefs appears to increase in a North-Easterly direction, sub-parallel to the strike of the BIC layered lithologies.

7.4.8 Disruptions

Identification of disruption intersections for the MR and UG2 Reefs are assisted with interpreted stratigraphic anomalies. Simply, the following factors may indicate disruptions:

>	Where footwall stratigraphic widths are wider,
>	Where the MR Pyroxenite or UG2 Chromitite is bifurcating, split or absent,
>	Where the MR width is anomalous with regard its normal facies widths.

MR disruptions have been identified within drill hole intersections and the 3D seismic survey conducted by AP. A clear understanding of normal reef facies behaviour has afforded their interpretation. These potholes are defined as areas where normal reef characteristics are destroyed. Pothole areas are hence believed to be un-mineable and are considered as a geological loss. The immediate footwall lithology underlying the MR Reef and UG2 Chromitite is often a key identifier of disruptions together with variations among deflections of the same drill hole. Disruptions appear to increase in frequency within the western most areas with the relative decrease in stability of the various lithologies in this area.

7.4.9 Replacement Pegmatite’s

Pseudo-form replacement bodies exist within the Project Area. A total of at least 12 drill holes of a population of 19 holes drilled in the WBJV Project 1A area have Intersected Iron Replacement Material (“IRUP”).

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It is evident from North-East orientated geological cross-sections constructed through the area that this IRUP mostly dominates the upper lithological sequence confined to the Main Zone and some upper lithological units of the Critical Zone. A typical down-dip section profile across Maseve Project area-A clearly indicates that the area of IRUP influence is sub parallel to the lithology pseudo layering.

The IRUP influence increases towards the MR subcrop environment and also further West where the Main Zone lies unconformable on Transvaal dolomites at shallow depths. The regional aeromagnetic image shows the surface expression of the IRUP influence and emphasizes the pseudo-morph shape of this anomaly.

Further South, and on the remainder of Maseve Project area, these IRUP anomalies occur as islands randomly spaced and are mostly recognizable on the aeromagnetic survey image. With the drilling grid averaging 250m Maseve Project area and in some geologically sensitive areas reduced to 130m, it was not always possible to further delineate the boundaries of these IRUP bodies.

7.5 Mineralisation

The potential economic horizons in the WBJV Maseve Project area and 1A are the MR and UG2 situated in the Critical Zone of the Rustenburg Layered Suite of the BIC; these horizons are known for their continuity. In general, the layered package dips at less than 20° and local variations in the reef attitude have been modelled. The MR and UG2 Reefs, in the Maseve Project area and 1A area, dip between 4° and 42° and occasionally are near vertical, averaging 22°.

The precious metals occur in a variety of forms. One or more of the metals may be present in combination with sulphur, arsenic, selenium or tellurium metallic particles of PGE’s or of PGE’s alloyed with base metals are also found. Additional PGE’s are found in solid solution in base metal sulphide particles.

Chromite crystals form a large part of the volume of UG2 and other chrome ores. Base-metal sulphides are much more prevalent in Merensky ores than in the UG2 ores. The grain size of mineral particles varies widely but is coarsest in MR.

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8.0 DEPOSIT TYPES

The most pronounced PGE mineralisation in the Project Area occurs within the MR and is generally associated with a 0.1 – 1.2m -thick feldspathic pegmatoidal pyroxenite unit. The MR is generally also associated with thin Chromitite layers on either/both the top and bottom contacts of the feldspathic pegmatoidal pyroxenite. The second important mineralised unit is the UG2, which is on average 1.28m thick and occurs within the Project Area.

8.1 Geological Modelling

8.1.1 Geostatistical Zones

The MR and UG2 were delineated into eleven and eight geostatistical zones respectively, each for the purposes of grade estimation. The Geo Zones were determined using the following criteria:

>	Major structural blocks,
>	The geological facies plan,
>	Sub-dividing the facies into smaller zones according to the grade and thickness variations.

The three facies type areas for the MR were then subdivided on grade trends resulting in 11 geostatistical zones, as shown in Figure 8-1 below. The areas not included in a domain represent areas where the reef is not developed, eroded or iron replacement took place, and these areas were excluded from the mineral resource estimation.

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The four UG2 facies types were sub-divided according to grade variations resulting in eight geostatistical zones for the UG2, as shown in Figure 8-2, with Domain 1 covering the largest area. The areas not included in a zone represent areas where iron replacement took place, and these areas are excluded from the mineral resource estimation.

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Estimation into the footwall units of the MR, FW1, FW2, FW3, and FW4 was also carried out. All four footwalls had one geostatistical zone each. The footwall units are twenty centimetre slices into the footwall of the MR. Footwall 1 represents the unit directly below the MR.

8.1.2 Geological Modelling and Development of a Dip Model

The geological modelling of the MR and UG2 was undertaken using the Datamine software to produce Three Dimensional (3D) base of reef of the two reefs. The base of reef model was developed from data provided by the magnetic survey results, seismic and geological logs of drilled cores to define the lower reef contact surface.

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Reef fault blocks were generated, bounded by faults traversing the reef. The wireframes were extended to cover the lease areas by honouring the dip trends at the edges. Figure 8-3 illustrates the plan view of the base of the MR showing dip changes.

8.1.3 Geological Losses

Table 8-1 details the estimated geological loss for the MR and UG2 in the Project Area, based on factors such as fault density and size, expected pothole development and areas with IRUP or no reef developed. With the close drill spacing in specific parts of the project area, it was possible to delineate areas of reef vs. areas with no reef. For these areas a lower geological loss has been applied as to areas where an estimated loss is determined. Figure 8-4 shows the areas that the losses have been applied to.

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For the project area the total excluded area is 35% of the total project area. Table 8-1 represents additional losses applied where reef is expected resulting in up to 48% geological losses per area.

Table 8-1: Geological Losses Estimated for the MR and UG2

MR Geological Loss		UG2 Geological Loss
Domain	Loss %	Domain	Loss %
1	14	1	14
2	14	2	14
3	20	3	15
4	11	4	15
5	11	5	16
6	14	6	16
7	11	7	16
8	14	8	15
9	14
10	11
11	35

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8.1.4 Hard and Soft Estimation Boundaries

The primary difference in soft and hard boundaries is the data that is available for any given process. For example, within kriging, hard boundaries are used against fault / intrusive structures where the data is unrelated to data on the opposite side of the boundary. Where the data represents a continuous un-interrupted area but has different facies / geostatistical characteristics, it is often prudent to include a thin area of the adjacent data in the estimation process in order to ensure grade continuity. Typically, the soft boundary is within 50m of the hard boundary.

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9.0 EXPLORATION

Exploration conducted at the WBJV Project 1 (Maseve) area includes geophysical surveys and diamond drilling, the details of which have been detailed in the Feasibility Report (July 2008). This report is available from www.sedar.com.

Since the feasibility report of 2008 additional exploration in the way of surface diamond drilling has been done to establish a more accurate 3 dimensional relationships of both the Merensky and UG2 reefs. The descriptions of the results of these additional drilling is part of this report.

9.1 Survey (field observation) results, procedures and parameters

Fieldwork in the form of soil sampling and surface mapping was initially done on the farm Onderstepoort, where various aspects of the lower Critical Zone, intrusive ultramafic bodies and structural features were identified. Efforts were later extended southwards to the farms Frischgewaagd and Elandsfontein. The above work contributed directly to the economic feasibility of the overall project, directing the main focus in the project area towards delineation of the subcrop position of the actual Merensky and UG2 economic reef horizons.

Geophysical information obtained from Anglo Platinum was very useful during the identification and extrapolation of major structural features as well as the lithological layering of the BIC. The aeromagnetic data alone made it possible to delineate magnetic units in the Main Zone, to recognise the strata strike and to identify the dykes and iron-replacements

BW Green was contracted to do ground geophysical measurements. Ground gravity measurements of 120.2km have been completed on 500m line spacing perpendicular to the strike across the deposit, together with 65.5km magnetic. The ground gravity data played a significant role in determining the hinge line where the BIC rocks start thickening down-dip, and this raised the possibility of more economic mineralisation. At the same time, the data shows where the Transvaal footwall causes the abutment or onlapping of the BIC rocks. Ground magnetic data helped to highlight faults and dykes as well as to delineate the IRUPs.

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9.1.1 Gravity Survey

The objective of the gravity survey was twofold:

>	To determine the structure of the subcropping mafic sheet on the sedimentary floor. This mafic sheet has a positive density contrast of 0.3 gram per cubic centimetre (Smit et al,) with the sediments.
>	To determine the thinning (or abutment) to the west of the mafic rocks on the floor sediments.

The instruments used for this survey are:

>	Gravity meter – Texas Instruments Worden Prospector Gravity Meter – This is a temperature-compensated zero length quartz spring relative gravimeter with a claimed resolution of 0.01mgal and an accuracy of 0.05mgal.
>	Position – Garmin GPS 12, Garmin GPS 72 and Magellan eXplorist 300 – These are 12- (Garmin’s) and 14-channel (Magellan) hand-held navigation GPSs; all with screens displaying the track, the ability to repeat and average each reading to a required level of accuracy and large internal memories. The GPSs were all set to the UTM projection (zone 35J) and WGS84 coordinate system. The X-Y positional accuracy was well within the specifications of this survey but the Z coordinate accuracy was inadequate.
>	Elevation – American Paulin System Surveying Micro Altimeter M 1-6 – This is a survey-standard barometric altimeter with a resolution of 30cm commonly used in regional gravity surveys. Although it does not meet the requirements of micro-gravity surveys, it is well up to the requirements of this survey.

9.1.2 Field Procedure

The survey was completed in two phases – a reconnaissance survey followed by a second detailed phase completed in four steps. The initial phase consisted of a gravity survey along the major public roads of the project area. All kilometre posts (as erected by the Roads department) were tied in as base stations through multiple loops to a principal base station. Readings were taken at 100m-intervals between the base stations, re-occupying the stations at less than hourly intervals. The instrument was only removed from its padded transport case for readings. The readings were taken on the standard gravimeter base plate and then used to determine the positions.

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At each station the gravimeter was read, the GPS X-Y position was taken until the claimed error was less than 5m and then stored along with the time on the instrument (All three GPS’s were used alternately during the survey with a short period of overlap to check for instrument error). The elevation was then determined using the Paulin altimeter. This exercise covered 55 line kilometres.

The second phase involved taking readings at every 100m along lines 500m apart with a direction of 51° true north. The GPSs played an important role in identifying gaps and ensuring that the lines being navigated were parallel to each other. Previously established base stations were re-occupied at least every hour. Where base stations were missing, additional stations were tied in with the original. This exercise covered 65km.

9.1.3 Post Processing

If drift on the altimeter and gravimeter were found to be excessive new readings were taken, otherwise drift corrections were applied to the readings. Using the gravimeters dial constant the raw readings were converted to raw gravity readings. The latitude, Bouguer and free-air corrections were then applied to the data. For the Bouguer correction, a density of 2.67 gram per cubic centimetre (g/cc) was used. The terrain-effect was calculated for the observation points closest to the Pilanesberg and was found to be insignificant in relation to the gravitational variations observed.

The resultant xyz positions were then gridded on a 25m grid using a cubic spline gridding algorithm. Filters were applied to this grid and the various products used in an interpretation which included information about the varying thickness of the mafic sheet, the presence of faults and the extent of the IRUPs.

9.1.4 Magnetic Survey

The purpose of the ground magnetic survey was to trace faults and dykes, determine the sense and magnitude of movement of such features and to delineate the highly magnetic IRUPs. It was decided to be consistent with the gravity survey and to use lines of a similar direction and spacing. In practise, however, this was not always possible owing to the magnetic survey’s susceptibility to interference from parallel fences, power lines and built-up areas in general. For these reasons as well as possible interference from gravity-related equipment, magnetic surveys are generally done after the gravity survey.

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The instruments used for this survey are:

>	Magnetics – Geometrics G 856 – This instrument is a proton-precession magnetometer used in this case as a total field instrument.
>	Position – Garmin GPS 12, Garmin GPS 72 and Magellan eXplorist 300

9.1.5 Field Procedure

The field procedure was similar to that of the second phase detailed gravity survey with the GPS used for guidance and covered 65km. With no equivalent to the gravity survey's first phase and no second magnetometer being used as a base station, a series of magnetic base stations also had to be tied in so that a base station was returned to every 30 minutes. Readings (including time) were taken at an average of 5m intervals. Position was determined by GPS every 100m and other positions interpolated through processing. Possible sources of interference such as fences and power lines were noted.

9.1.6 Post processing

All high-frequency signals associated with cultural effects were removed. The individual lines were then put through various filters and the results presented as stacked profiles and interpreted. Inversion modelling was also performed on specific anomalies and the results included in the interpretation compilation, together with information on faults, dykes and IRUPs.

9.2 Interpretation of survey (field observation) results

The structural features identified from the aeromagnetic data were interpreted in terms of a regional structural model shown. Major dyke features were easily recognised and these assisted in the compilation of a structural model for the Maseve project area. Exploration drilling later helped to identify a prominent east-west-trending linear feature as a south-dipping dyke. This dyke occurs along the northern boundary of the project area. A second dyke occurs along the northeastern boundary of the Elandsfontein and Frischgewaagd areas. Other major structural features include potential faults oriented at 345 degrees north in the deep environment of the Frischgewaagd south area.

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10.0	DRILLING
10.1	Type and Extent of Drilling

Since March 2005, PTM has drilled 242 264m of core from drill hole WBJV001 to WBJV454, as well as a total of 13 776m of core for the drill holes W001 to W041, drilled for geotechnical purposes. Additional drilling since 2009 necessitated the update of the Mineral Resources. The location and extent of the drill holes at the project are detailed and illustrated in Section 11 of this report.

The type of drilling conducted on the WBJV was a diamond-drilling core-recovery technique involving a BQ-size solid core extraction. The drilling was placed on an unbiased 500m x 500m grid and detailed when necessary to a 250m x 250m grid, and in mine development areas to a 100m x 100m or 50m x 50m grid.

10.2 Procedures, Summary and Interpretation of Results

10.2.1 Sampling Method, Location, Number, Type and Size of Sampling

The first step in the sampling of the diamond-drilled core was to mark the core from the distance below collar in one metre units and then for major stratigraphic units. Once the stratigraphic units were identified, the economic units – MR and UG2 – were marked. The top and bottom contacts of the reefs were clearly marked on the core. Thereafter, the core was rotated in such a manner that all lineation’s pertaining to stratification were aligned to produce a representative split. A centre cut line was then drawn lengthways for cutting. After cutting, the material was replaced in the core trays. The sample intervals were then marked as a line and a distance from collar.

The sample intervals were typically 15 – 25cm in length. In areas where no economic zones were expected, the sampling interval could have been as much as a metre. The sample intervals were allocated a sampling number, and this was written on the core for reference purposes. The half-core was then removed and placed into high-quality plastic bags together with a sampling tag containing the sampling number, which was entered onto a sample sheet. The start and end depths were marked on the core with a corresponding line. The duplicate tag stayed as a permanent record in the sample booklet, which was secured on site. The responsible Project Geologist then sealed the sampling bag. The sampling information was recorded on a specially designed sampling sheet that facilitated digital capture into the SABLE system. The sampling extended for about a metre (approximately four samples) into the hanging wall and footwall of the economic reefs.

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10.2.2 Drilling Recovery

All reef intersections that were sampled required a 100% core recovery. If less than 100% was recovered, the drilling company re-drilled, using a wedge / deflection to achieve the desired recovery.

10.2.3 Sample Quality and Sample Bias

The sampling methodology accords with PTM protocol based on industry-accepted best practice. The quality of the sampling is monitored and supervised by a Qualified Geologist. The sampling is done in a manner that includes the entire economic unit, together with hanging wall and footwall sampling. Sampling over-selection and sampling bias is eliminated by rotating the core so that the stratification is vertical and by inserting a cutline down the centre of the core and removing one side of the core only.

10.2.4 Widths of Mineralised Zones – Mining Cuts

The methodology in determining the resource cuts is derived from the core intersections. Generally, the higher grade positions of the reefs are about 60cm thick. For both the MR and UG2, the marker unit is the bottom reef contact, which is generally a chromite contact of less than a centimetre. The cut is taken from that chromite contact and extended vertically to accommodate most of the metal content. If this should result in a resource cut less than 80cm up from the bottom reef contact, it is extended further to 80cm. If the resource cut is thicker than the proposed 80cm, the last significant reported sample value (generally 2g/t and above) above 80cm is added to determine the top reef contact.

In the case of the UG2, the triplets or UG2 Leader (if and where developed and within 30cm from the top contact) are included in the resource cut.

The results of the drilling and the general geological interpretation were digitally captured in SABLE (commercially available logging software) and a GIS software package named ARCVIEW. The exact drill hole locations, together with the results of the economic evaluation, were plotted on plan. From the geographic location of the holes drilled, regularly spaced sections were drawn by hand and digitised. This information was useful for interpreting the sequence of the stratigraphy intersected as well as for verifying the drill hole information.

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10.3 Additional Information for Projects other than Advanced Projects

10.3.1 Comment on True and Apparent Widths of the Mineralised Zones

The geometry of the deposit has been clearly defined in the sections drawn through the Project Area. All drill holes were drilled vertically and the downhole surveys indicate very little deviation. A 3D surface – DTM – was created and used in the calculation of the average dip. This dip has been factored into the calculations on which Mineral Resource estimates are based.

10.3.2 Comment on the Orientation of the Mineralised Zones

The mineralised zones within the Project Area include the MR and the UG2, both of which are planar tabular ultramafic precipitants of a differentiated magma and therefore form a continuous sheet-like accumulate.

The stratigraphic markers above and below the economic horizons have been recognised and facilitate recognition of the MR and the UG2. There are a few exceptions to the quality of recognition of the stratigraphic sequences. These disruptions are generally of a structural nature and are to be expected within this type of deposit.

In some drill holes, no clear stratigraphic recognition was possible. These drill holes were excluded from the Mineral Resource estimation of grade.

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11.0	SAMPLE PREPARATION, ANALYSIS AND SECURITY
11.1	Sample Preparation and Quality Control Prior to Dispatch

Drilled core is cleaned, de-greased and packed into metal core boxes by the drilling company. The core is collected from the drilling site on a daily basis by a PTM Geologist and transported to the exploration office by PTM personnel. Before the core is taken off the drilling site, the depths are checked and entered on a daily drilling report, which is then signed off by PTM. The Core Yard Manager is responsible for checking all drilled core pieces and recording the following information:

>	Drillers’ depth markers (discrepancies were recorded),
>	Fitment and marking of core pieces,
>	Core losses and core gains,
>	Grinding of core,
>	One metre interval markings on core for sample referencing,
>	Re-checking of depth markings for accuracy.

Core logging is done by hand on a PTM pro-forma sheet by Qualified Geologists under supervision of the Project Geologist, who is responsible for timely delivery of the samples to the relevant laboratory. The supervising and Project Geologists ensure that samples are transported by PTM Contractors.

Samples are not removed from secured storage location without completion of a chain of custody document; this forms part of a continuous tracking system for the movement of the samples and persons responsible for their security. Ultimate responsibility for the secure and timely delivery of the samples to the chosen analytical facility rests with the Project Geologist and samples are not transported in any manner without the Project Geologist’s permission.

When samples were prepared for shipment to the analytical facility, the following steps are followed:

>	Samples are sequenced within the secure storage area and the sample sequences examined to determine if any samples were out of order or missing,
>	The sample sequences and numbers shipped are recorded both on the chain of custody form and on the analytical request form,

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>	The samples are placed according to sequence into large plastic bags (the numbers of the samples were enclosed on the outside of the bag with the shipment, waybill or order number and the number of bags included in the shipment),
>	The chain of custody form and analytical request sheet are completed, signed and dated by the Project Geologist before the samples are removed from secured storage. The Project Geologist keeps copies of the analytical request form and the chain of custody form on site,
>	Once the above is completed and the sample shipping bags are sealed, the samples may be removed from the secured area. The method by which the sample shipment bags have been secured must be recorded on the chain of custody document so that the recipient can inspect for tampering of the shipment.

During the process of transportation between the project site and analytical facility, the samples are inspected and signed for by each individual or company handling the samples. It is the mandate of both the supervising and Project Geologist to ensure secure transportation of the samples to the analytical facility. The original chain of custody document always accompanies the samples to their final destination.

The Supervising Geologist ensures that the analytical facility is aware of the PTM standards and requirements.

It is the responsibility of the analytical facility to inspect for evidence of possible contamination of, or tampering with, the shipment received from PTM. A photocopy of the chain-of-custody document, signed and dated by an official of the analytical facility, is faxed to PTM’s offices in Johannesburg upon receipt of the samples by the analytical facility and the original signed letter is returned to PTM along with the signed analytical certificate/s.

The analytical facility’s instructions are that if they suspect the sample shipment has been tampered with, they will immediately contact the Supervising Geologist, who will arrange for someone in the employment of PTM to examine the sample shipment and confirm its integrity prior to the start of the analytical process.

If, upon inspection, the Supervising Geologist has any concerns whatsoever that the sample shipment may have been tampered with or otherwise compromised, the Responsible Geologist will immediately notify the PTM management in writing and will decide, with the input of management, how to proceed.

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In most cases analysis may still be completed although the data must be treated, until proven otherwise, as suspect and unsuitable as a basis for a news release until additional sampling, quality control checks and examination prove their validity.

Should there be evidence or suspicions of tampering or contamination of the sampling, PTM will immediately undertake a security review of the entire operating procedure. The investigation will be conducted by an independent third party, whose report is to be delivered directly and solely to the directors of PTM, for their consideration and drafting of an action plan. All in-country exploration activities will be suspended until this review is complete and the findings have been conveyed to the directors of the company and acted upon.

11.1.1 Security

Samples are not removed from secured storage location without completion of a chain of custody document. This forms part of a continuous tracking system for the movement of the samples and persons responsible for their security. Ultimate responsibility for the secure and timely delivery of the samples to the chosen analytical facility rests with the Project Geologist and samples are not transported in any manner without the Project Geologist’s permission.

During the process of transportation between the project site and analytical facility, the samples are inspected and signed for by each individual or company handling them. It is the mandate of both the supervising and Project Geologist to ensure secure transportation of the samples to the analytical facility. The original chain of custody document always accompanies the samples to their final destination.

The Supervising Geologist ensures that the analytical facility is aware of the PTM standards and requirements. It is the responsibility of the analytical facility to inspect for evidence of possible contamination of, or tampering with, the shipment received from PTM. A photocopy of the chain of custody document, signed and dated by an official of the analytical facility, is faxed to PTM’s offices in Johannesburg upon receipt of the samples by the analytical facility and the original signed letter is returned to PTM along with the signed analytical certificate/s. The analytical facility’s instructions are that if they suspect the sample shipment has been tampered with, they will immediately contact the Supervising Geologist, who will arrange for someone in the employment of PTM to examine the sample shipment and confirm its integrity prior to the start of the analytical process.

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If, upon inspection, the Supervising Geologist has any concerns whatsoever that the sample shipment may have been tampered with or otherwise compromised, the Responsible Geologist will immediately notify the PTM management in writing and will decide, with the input of management, how to proceed. In most cases analyses may still be completed, although the data must be treated, until proven otherwise, as suspect and unsuitable as a basis for a news release until additional sampling, quality control checks and examination prove their validity.

Should there be evidence or suspicions of tampering or contamination of the sampling, PTM will immediately undertake a security review of the entire operating procedure. The investigation will be conducted by an independent third party, whose report is to be delivered directly and solely to the directors of PTM, for their consideration and drafting of an action plan. All in-country exploration activities will be suspended until this review is complete and the findings have been conveyed to the directors of the company and acted upon.

The QP of this report is satisfied with the level of security and procedures emplaced to ensure sample integrity.

11.2 Sample Preparation and Analytical Procedures Employed by Laboratory

The laboratories that have been used to date are Anglo American Research Laboratories, Genalysis (Perth, Western Australia), ALS Chemex (South Africa), Set Point Laboratories (South Africa) and (currently) Bureau Veritas / Inspectorate (South Africa).

The National Association of Testing Authorities Australia has accredited Genalysis Laboratory Services (Pty) Ltd, following demonstration of its technical competence, to operate in accordance with ISO/IEC 17025, which includes the management requirements of ISO 9001: 2000.

Anglo American Research Laboratories, Set Point Laboratories and ALS Chemex are accredited by the South African National Accreditation System (SANAS), testing laboratory numbers T0051, T0223 and T0387 respectively. Bureau Veritas/Inspectorate Rustenburg is also accredited by the South African National Accreditation System (SANAS), testing laboratory numbers T0551 as of July 2014.

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Samples are received, sorted, verified and checked for moisture and dried if necessary. Each sample is weighed and the results are recorded. Rocks, rock chips or lumps are crushed using a jaw crusher to less than 10mm, the samples are then split using a riffle splitter. The samples are then milled for five minutes to achieve a fineness of 85% less than 75μm.

Samples are analysed for Pt (ppb), Pd (ppb) and Au (ppb) by standard 25g lead fire-assay using silver as requested by a co-collector to facilitate easier handling of prills as well as to minimise losses during the cupellation process. The resulting prills are dissolved with aquaregia for ICP analysis.

After pre-concentration by fire assay and microwave dissolution, the resulting solutions are analysed for Au and PGM’s by the technique of ICP-OES (inductively coupled plasma–optical emission spectrometry).

11.3 Quality Assurance and Quality Control Results

PTM has a well-established and functional quality assurance or quality control (QAQC) procedure. PTM are the custodians of the QAQC results. Over the history of the Project CJM Consulting has reviewed the findings of QAQC results for the purposes of establishing validity of the data for inclusion into the Mineral Resource estimation, with particular focus on the results since the last Resource Statement. To this end, Data from Setpoint and Bureau Veritas was examined.

PTM has well established QAQC protocols. The following summarises the PTM protocols for quality control during sampling:

>	The Project Geologist oversees the sampling process,
>	The Core Yard Manager oversees the core quality control,
>	The Exploration Geologists and the sample technician are responsible for the actual sampling process,
>	The Project Geologist oversees the chain of custody,
>	The internal QP verifies both processes and receives the laboratory data,
>	The internal Resource Geologist and the Database Manager merge the data and produce the SABLE sampling log with assay values,

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>	Together with the Project Geologist, the Resource Geologist determines the initial mining cut,
>	An external auditor verifies the sampling process and signs off on the mining cut,
>	A second external database auditor verifies the SABLE database and highlights QAQC failures,
>	QAQC graphs (standards, blanks and duplicates) and anomalies and failures are reported to the internal QP, QAQC analysis is conducted by SABLE data warehouse,
>	The internal QP requests re-assay,
>	At this stage no check samples are sent to a second laboratory to verify the validity of data received from the first laboratory, only a single batch of overlap samples were submitted to both Setpoint and Burequ Veritas when the changeover happened.

Additional PTM QAQC procedures include examination of all core trays for correct number sequencing and labelling. Furthermore, the printed SABLE sampling logs (including all reef intersections per drill hole) are compared with the actual remaining drill hole core left in the core boxes. The following checklist forms the standard PTM checklist for verification:

>	Sampling procedure, contact plus 10cm, sample length 15 – 25cm,
>	Quality of core (core-loss) recorded,
>	Correct packing and orientation of core pieces,
>	Correct core sample numbering procedure,
>	Corresponding numbering procedure in sampling booklet,
>	Corresponding numbering procedure on printed SABLE log sheet,
>	Comparing SABLE log sheet with actual core markings,
>	Corresponding chain of custody forms completed correctly and signed off,
>	Corresponding sampling information in hardcopy drill hole files and safe storage,
>	Assay certificates filed in drill hole files,
>	Electronic data from laboratory checked with signed assay certificate,
>	Sign off each reef intersection (bottom reef contact and mining cut),
>	Sign off completed drill hole file,
>	Sign off on inclusion of mining cut into Mineral Resource database.

As part of the sampling protocol, PTM regularly inserts QC samples (i.e. standards and blanks) into the sample stream. It should be noted that PTM do not include field duplicates into the samples stream, and the analytical laboratory was asked to regularly assay split pulp samples as a duplicate sample to monitor analytical precision.

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11.3.1 Standards

Analytical standards were used to assess the accuracy and possible bias of assay values for Pt and Pd. Rh and Au were monitored where data for the standards were available, but standards were not failed on Rh and Au alone. Quality control data for the WBJV is managed by PTM using SABLE dataworks software.

A selection of standards including some made from BIC mineralisation were used from drill hole WBJV027 onwards. The standards currently in use are tabled below. The standards are stored in sealed containers and considerable care is taken to ensure that they are not contaminated in any manner (e.g. through storage in a dusty environment, being placed in a less than pristine sample bag or being in any way contaminated in the core saw process).

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Table 11-1: Standards Currently in Use

Std Ref	Std Type	PT Mean	PT 3Std Min	PT 3Std Max	PD Mean	PD 3Std Min	PD 3Std Max	AU Mean	AU 3Std Min	AU 3Std Max	CU Mean	CU 3Std Min	CU 3Std Max	NI Mean	NI 3Std Min	NI 3Std Max
SAND	Blank		-	0.08		-	0.08		-	0.08		-	100.00		-	100.00
AMIS0001	Standard	0.77	0.66	0.87	1.04	0.92	1.16	0.12	0.08	0.16
AMIS0089	Standard	1.09	0.91	1.27	0.70	0.61	0.79	0.04	0.02	0.06	59.00	50.00	68.00	452.00	374.00	530.00
AMIS0099	Standard	0.59	0.50	0.68	0.23	0.17	0.28	0.09	0.07	0.11	256.00	229.00	283.00	443.00	371.00	515.00
AMIS0151	Standard	4.64	4.10	5.18	3.15	2.73	3.57	0.07	0.05	0.09	150.00	129.00	171.00	1,281.00	988.50	1,573.50
AMIS0165	Standard	16.90	14.86	18.94	19.10	17.06	21.14	1.66	1.45	1.87	17,710.00	16,165.00	19,255.00	28,160.00	25,490.00	30,830.00
AMIS0207	Standard	2.28	1.95	2.61	1.26	1.14	1.38	0.09	0.07	0.10	85.00	71.50	98.50	1,059.00	871.50	1,246.50
AMIS0254	Standard	2.19	2.17	2.21	1.12	1.00	1.24	0.20	0.17	0.23	762.00	688.50	835.50	1,735.00	1,469.50	2,000.50
AMIS0256	Standard	4.86	4.53	5.19	2.50	2.32	2.68	0.34	0.28	0.40	1,252.00	1,148.50	1,355.50	2,913.00	2,641.50	3,184.50
AMIS0411	Standard	0.54	0.45	0.63	0.67	0.58	0.76	0.08	0.06	0.10	742.00	652.00	832.00	1,368.00	1,216.50	1,519.50

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Tolerance limits are set at two and three standard deviations from the certified mean value of the reference material:

>	A single batch is rejected when reference material assays are beyond the two standard deviations limit, and
>	Any two consecutive batches are rejected when reference material assays are beyond the two standard deviations limit on the same side of the mean.

A failed standard is considered cause for re-assay if it falls within a determined economic mining cut for either the MR or UG2 Reef (MRMC and UG2MC). Analysis of any fails is conducted and the appropriate action is undertaken. In general the compliance with the certified values is good, however there is some evidence of sample swapping, although this is minimal. This may be reduced or eradicated by a change in laboratory procedure regarding sample labelling and tray labelling.

11.3.2 Blanks

The insertion of blanks provides an important check on the laboratory practices, especially potential contamination or sample sequence mis-ordering. Blanks consist of a selection of Transvaal Quartzite pieces (devoid of Pt, Pd, copper, and nickel mineralisation) of a mass similar to that of a normal core sample. The blank being used is always noted to track its behaviour and trace metal content. The plotted graphs have a warning limit, which is equal to five times the blank background. These graphs and calculations are available on request in digital format. In general the failure rate is deemed not to have a material effect on the data.

11.3.3 Field Duplicates

The purpose of having field duplicates is to provide a check on possible sample over-selection. The field duplicate contains all levels of error – core or reverse-circulation cutting splitting, sample size reduction in the preparation laboratory, sub-sampling at the pulp and analytical error. Field duplicates were, however, not used on this project due to the assemblage of the core. Because of this problem, the laboratory was asked to regularly assay coarse reject samples as a duplicate sample to monitor analytical precision. The duplicate results graphs and calculations are available on request in digital format. The original analysis vs. the duplicate analysis showed no irregular values. This indicates no sample miss-ordering or nugget effect, which is of concern as the deposit is known to have a high nugget and be extremely variable at close range. These graphs and calculations are available on request in digital format.

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11.3.4 Third Party Review

Although samples are assayed with reference materials, an assay validation programme should typically be conducted to ensure that assays are repeatable within statistical limits for the styles of mineralisation being investigated. It should be noted that validation is different from verification; the latter implies 100% repeatability. The assay validation programme should entail:

>	A re-assay programme conducted on standards that failed the tolerance limits set at two and three standard deviations from the Round Robin mean value of the reference material,
>	Ongoing blind pulp duplicate assays,
>	Check assays conducted at an independent assaying facility.

Re-assay

This procedure entails re-submission and re-assaying of failed standard (2) together with the standard (1) submitted before and the standard (3) submitted after the particular failed standard (2) as well as all submitted field samples (pulps) in between standard (1) and standard (3) were re-submitted for re-assaying, generally this results in the entire batch being re-assayed due to the nature of the batch sizes submitted.

Laboratory Duplicates

The laboratory regularly assays pulp duplicates with each batch of data. The duplicate results graphs and calculations are available on request in digital format. The original analysis vs. the duplicate analysis showed no irregular values. This indicates no sample miss-ordering or nugget effect.

Check Assays

PTM has typically used Genalysis in Perth as the second laboratory for checks on the assay results from Set Point. There is one batch of data currently which was sent to both Set Point and Bureau Veritas when the laboratory service provider was changed. Umpire results confirm the satisfactory performance of the primary laboratory on the standards and duplicates.

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11.4 Adequacy of Sample Preparation, Security and Analytical Procedures

The QA and QC practice of PTM is a process beginning with the actual placement of the drill hole position (on the grid) and continuing through to the decision for the 3D economic intersection to be included in (passed into) the database. The values are also confirmed, as well as the correctness of correlation of reef / mining cut so that populations used in the geostatistical modelling are not mixed; this makes for a high degree of reliability in estimates of Mineral Resources / Mineral Reserves. In the opinion of CJM Consulting, the QA and QC procedures as well as the sample preparation and security procedures are adequate to allow the data to be used with confidence in the Resource Estimation.

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12.0	DATA VERIFICATION
12.1	Verification of Data by QP

Data verification was conducted by CJM Consulting as part of the Mineral Resource estimation for the WBJV Project 1 (Maseve) as detailed below:

>	With regard to missing Rh assay values within the assay data, a regression analysis was undertaken where the prill relationships to one another and to 3 PGE + Au (4E) were generated. The missing Rh values were infilled with values utilizing the regression equation (y=mx+c). The regression to back-calculate the Rh values was in excess of R=88%.
>	With regard to missing Specific Gravity (“SG”) values, the average was generated for each individual lithological type, and the missing SG values inserted according to the lithological unit. The SG is critical in the data import stage; during compositing of the drill hole intersections into single width intersections, the composting is weighted with SG.

12.2 Nature of the Limitations of Data Verification Process

As with all information, inherent bias and inaccuracies can and may be present. Given the verification process that has been carried out, however, should there be a bias or inconsistency in the data, the error would be of no material consequence in the interpretation of the model or evaluation.

The data is checked for errors and inconsistencies at each step of handling. The data is also rechecked at the stage where it is entered into the deposit-modelling software. In addition to ongoing data checks by project staff, the senior management and directors of PTM have completed spot audits of the data and processing procedures. Audits have also been carried out on the recording of drill hole information, the assay interpretation and final compilation of the information. The individuals in PTM’s senior management and certain directors of the company who completed the tests and designed the processes, are non-independent mining or geological experts.

12.3 Possible Reasons for not having Completed a Data Verification Process

All PTM data has been verified before being statistically processed. The QP states that the data is verified adequately for resource estimation.

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13.0	MINERAL PROCESSING AND METALLURGICAL TESTING
13.1	Historical Metallurgical Testwork Summary

The metallurgical test work that has been conducted on the WBJV Project 1 (Maseve) ore body was initially carried out by SGS Lakefield during the PFS Stage on both MR and UG2 ores (December 2006 to January 2007).

Mintek, South Africa conducted a more extensive testwork programme as part of the FS Stage during the first half of 2009 (Evaluation of the Elandsfontein / Frischgewaagd PGE Deposit for the WBJV). Most of this work was undertaken to establish flotation kinetics to input into simulation work to predict plant performance. Mintek also completed confirmatory testwork in 2012 for the DFS stage.

13.1.1 PFS Testwork on UG2 Ore at SGS Lakefield (December 2006)

During the second half of 2006, SGS Lakefield under the management of PTM completed the following testwork on UG2 core samples from the Frischgewaagd area:

>	Comminution testwork including Bond Abrasion Index test, Bond Rod Mill Work Index (“BRWi”) test, and a Bond Ball Mill Work Index (“BBWi”) test,
>	Optimisation of flotation conditions on a bulk composite from the individual core samples,
>	Determining the PGE grades and recoveries achievable from the composite sample,
>	Determining the flotation response variability of the individual core samples and composites of the Western Area cores vs the Eastern Area Cores.

The results from this testwork campaign can be summarised as follows:

>	The UG2 material tested reported a medium abrasion tendency, and a BRWi of 12.2kWh/t classifying the material as medium to hard. A BBWi of 17.3kWh/t was reported which classified the material as hard. It was noted that fully autogenous milling of this UG2 material would not be recommended,
>	Rougher rate tests at varying grinds demonstrated that a final grind of 80% passing 75µm gave optimum flotation performance,
>	Initial standard flotation tests on composites of the Western and Eastern areas resulted in 4E recoveries of 92% and 90.4%, respectively,

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>	Initial rougher / cleaner tests showed that multiple cleaning stages in a MF2 circuit with depressant addition during each cleaning stage would be required to produce concentrate grades of > 150g/t 4E,
>	From a final product grade vs recovery graph, a final 4E recovery of 82% at a grade of 150g/t was predicted,
>	The variability testing on the individual cores samples indicated small variations in the rougher circuit with larger variations in the final concentrate results.

13.1.2 PFS Testwork on MR at SGS Lakefield (January 2007)

In early 2007 SGS Lakefield, under the management of Turnberry Projects and PTM, completed testwork on MR core samples from the Frischgewaagd area in order to optimise the depressant type and dosage during the cleaning / recleaning stages to achieve higher grades whilst maintaining good recoveries. The target was 87.5% 4E recovery at a grade of 150g/t 4E.

The predicted MR 4E recovery and grade were 88% and 150g/t respectively in open circuit tests on a sample with a 6.3g/t 4E head grade. A locked cycle test using the same conditions resulted in a 4E recovery of 94% at a grade of 179g/t 4E. A final Cu recovery of 89% at a grade of 2.4% and Ni recovery of 59.5% at a grade of 3.6% was achieved. These results were achieved in a MF1 circuit utilising a grind of 90% - 75µm and a 21 minute rougher stage. The concentrate was cleaned in a 2 stage cleaning circuit (6 minute cleaner and a 4 minute recleaner). Collector and depressant were required in all the flotation stages.

13.1.3 Updated FS Testwork at Mintek (June 2009)

PTM requested through Eurus Mineral Consultants (EMC) that MINTEK perform a series of confirmatory metallurgical tests on MR and UG2 reef drill cores extracted from their Frischgewaagd property for the purpose of a bankable feasibility study.

Various samples were received from across the ore body. A total of eleven MR core samples from the central and Eastern mining sections were delivered but only four samples were used for the investigation. The other cores were rejected due to an anomalously low head grade and/or their location on the mining map. Results from composited samples of the adopted samples gave the following conclusions:

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>	The mineralogy, grade and occurrence of the reef in the deposit exhibited marked variability. The MR 4E head grade varied from 1.9g/t to 8.5g/t with an average of 5.3g/t. Extensive geological testwork pointed to an expected average 4E head grade of 6.8g/t. Of the cores received, only a few exhibited a head grade close to this and certain cores were excluded from the full investigation due to their anomalously low head grades.
>	The testwork was conducted with location composite sample with a 4E head grade of 2.5g/t. A new high-grade sample was compiled which exhibited a 4E head grade of 4.7g/t and underwent comparative testwork. There was a marginal difference in metallurgical response between the two samples.
>	MR Rougher rate tests performed at various grinds demonstrated predicted 4E recoveries to the rougher concentrate in excess of 85%, even at a coarse grind. 4E recovery increased with increased grind. 4E recoveries of 88.9%, 90.1% and 96.1% were achieved from grinds of 40%, 60% and 80% passing 75μm respectively, indicating that the PGM particles were easily liberated and that the ore exhibits fast flotation kinetics. The predicted base metal recoveries were 90% for copper and 78% for nickel.
>	From MF1 and MF2 rougher and cleaner rate tests performed at varying depressant dosages the flotation kinetics of the PGM’s and base metal sulphides could be modelled and the presence of floatable gangue determined.
>	The closest prediction of MF1 plant performance for the MR achieved from an MF1 Bryson test (at a grind of 40% passing 75μm) with the original location composite sample exhibiting a 4E head grade of 2.5g/t was predicted 4E grade and recovery of 82g/t and 81% respectively. The overall 4E grade and recovery was 61g/t and 86% respectively, with 8.5% recovery lost to the rougher tails. Overall, 86% of the copper and 57% of the nickel was recovered to the final concentrate at grades of 1.6% and 2.1%, respectively.
>	The closest prediction of MF2 Plant performance for the MR from an MF-2 Bryson test (at a primary and secondary grind of 40% and 80% passing 75μm) at a 4E head grade of 2.5g/t was predicted 4E grade and recovery of 122g/t and 87% respectively. The overall 4E grade and recovery was 85g/t and 91%, respectively, with 3% recovery lost to the rougher tails. The overall copper and nickel recoveries were 84% and 58% respectively at grades of 2% and 2.8% respectively.
>	Testwork performed on a new high-grade location composite show that, despite the higher 4E head grade, the base metal content was considerably lower and the flotation response is poorer than the lower grade location composite. It is evident that not all ore types have been identified and confirms the variability of the reef across the ore body.

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>	Additional flotation tests indicated that similar flotation performance was achieved from MF-1 and MF-2 circuits at similar grinds with excessive residence time.
>	The detailed concentrate assay indicated that the metal content of the final concentrates fell within the typical composition values for matte smelting.
>	A mineralogical investigation of flotation products indicated that the bulk, viz. 68% of the PGM’s in the final concentrate occurred, not as composites with base metal sulphide minerals but as liberated PGM minerals with the bismuthtelluride class predominating. The bulk of the PGM losses to the secondary cleaner tails (containing 6% of the 4E minerals at a grade of approximately 1g/t) also occurred as liberated bismuth-tellurides.
>	A Grind-mill test and simulation was conducted to assist in designing the milling circuit and tails thickening tests were done for thickener design purposes.

A total of 12 core samples were obtained from UG2 boreholes covering three mining sections, central, North Western and Eastern. Four core samples were used for the investigations. The other cores were rejected due to an anomalously low head grade and/or their location on the mining map. Various metallurgical and mineralogical testwork were conducted to characterise the ore and determine metallurgical performance. The following conclusions were drawn from the testwork:

>	The grade and occurrence were markedly variable. The 4E head grade varies from 0.7g/t to 7.2g/t with an average of 3.6g/t. Extensive geological testwork pointed to an expected 4E head grade of 3.8g/t.
>	Liberation influences metallurgical response as demonstrated by rougher rate 4E recoveries 87%, 88% and 92% at grinds of 40%, 60% and 80% passing 75µm respectively.
>	The 4E rougher rate test recovery for an MF1 circuit at a grind of 80% passing 75µm was 92% and for an MF2 circuit with secondary grind of 90% passing 75µm it was 94%. Grades were similar.
>	The predicted recovery and grade for an MF2 (40% and 90% passing 75µm primary and secondary grind respectively) test with open cleaner circuit were 85% and 75g/t, from a composite exhibiting a 4E head grade of 4.2g/t, with 6% lost to rougher tails.

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>	The closest approximation of the plant performance was achieved from a five-cycle MF2 locked cycle test, which predicted a 4E recovery of 86% at a grade of 102g/t, from the original location composite exhibiting a head grade of 3.6g/t (4E).
>	The chromite (Cr2O3) grade was less than 4%.
>	Mineralogical investigation of flotation products showed that the PGM’s in the final concentrate occurred mainly as sulphides and they were mostly, (63%), present associated with composite particles of base metal sulphides (predominantly pentlandite) and gangue particles (predominantly orthopyroxene) or associated with liberated base metal sulphides.
>	Some, (32%), of the PGM’s to the cleaner tails (containing 10% of the 4E elements at a grade of 5g/t) occurred as liberated sulphides with a high association in composites.

13.2 New Metallurgical Testwork

13.2.1 DFS Testwork on MR at Mintek (April 2012)

MINTEK was approached by DRA, on behalf of Maseve Investments 11 (Pty) Ltd to conduct confirmatory metallurgical testwork on 9 Merensky drill core samples from PTM’s Elandsfontein deposit for the Definitive Feasibility Study (DFS) and to compare the data from previous testwork conducted at Mintek. The scope was mainly to investigate ore variability:

>	Head Assay
>	Comminution
>	Grade-recovery relationship variability
>	Composite grade-recovery performance

The target results from the client were an overall flotation concentrate grade of 150g/t 4E at recovery of 88%. From the testwork it was concluded that the ore was not significantly different from ores tested previously. 4E recovery was 86% at a grade of 135g/t. Copper and nickel grades in concentrate were 1.9% and 3.5% respectively. There was a high variability in the 4E grades which varied from 1.8g/t to 13.11g/t. The composite grade was 4.89g/t (4E) at an average Pt/Pd ratio of 2.50. The Cu and Ni grades of the composite sample were 0.1% and 0.2% respectively.

Comminution testwork classified the Merensky ore as being hard, with a BBWi of 18.69 kWh/t at 75µm.

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Flotation testwork including both rougher and cleaner rate tests employing locked cycle testwork and open cleaner circuits (as required) were conducted to test the following:

>	Ore flotation variability
>	Effect of grind on recovery for an MF1 circuit
>	Comparison of MF1 and MF2 circuits
>	Effect of depressant dosage to the cleaner circuit

Recoveries for MF1 rougher rate tests conducted on 8 cores samples at a target grind of 60% -75μm demonstrated significant variability with the overall 4E recovery ranging between 79% and 93%. The base metal recoveries were also variable with Cu recoveries ranging between 57% and 95% and Ni recoveries ranging between 54 % and 81%.

Three MF1 rougher rate tests at grinds of 40%, 60% and 80% passing 75μm indicated that the ore is sensitive to grind, with overall (after 20min) 4E recoveries of 81,0%, 86,9% and 91,0% respectively. The 4E kinetics obtained for the MF2 circuit were better as 91% recovery of the PGM’s was after 10 minutes, compared to 20 minutes for the MF1 circuit. The overall 4E recovery after 20 minutes for the MF2 circuit was 94%. The overall rougher Cu and Ni recoveries were 71% and 72% respectively.

The cleaner rate tests conducted at various depressant dosages indicated that an increase in depressant increased the grade but did not adversely affect recovery. The results of these five tests showed that at the higher depressant dosages, there was a marginal improvement in 4E grade. Increasing the depressant dosage did not appear to have any significant impact on the overall 4E recovery to the cleaner concentrate, as the overall 4E recovery to the cleaner concentrate ranged from 87% to 90%. The cleaner efficiency was high at ~95%. A poor grade recovery curve was obtained for the MF2 circuit which was a result of the slow floating material from the secondary rougher as the primary rougher concentrate was not included. The secondary rougher concentrate was cleaned at a depressant dosage of 30g/t. The highest overall 4E recovery was 94% at a low 4E grade of 22g/t.

Two Bryson tests were performed to more accurately predict the overall concentration grade and recovery to be achieved from a full-scale plant (in the absence of a circulating load in the flotation circuit). The first Bryson test was for standard conditions. In the second Byron test primary cleaner residence time was doubled, SIBX dosage to the secondary cleaner increased and depressant dosage of the secondary cleaner stage was increased, in an effort to improve the primary circuit recovery, reduce the PGM losses to the secondary cleaner tails and improve the grade of the secondary cleaner concentrate.

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For the first Bryson test (standard), 4E recovery was 87% at a grade of 103g/t. The PGM losses were 6% to the secondary cleaner tails and 5% to the rougher tails. The second Bryson tests had improved PGM flotation performance, with 4E to recovery of 88% at a grade of 133g/t. The overall Cu recovery was 88.6% at a grade of 2.9%, while the Ni recovery was 65% at a grade of 3.9% . The losses to the secondary cleaner tailings were reduced from 6% to 3%.

The Bryson test performed in the 2009 testwork phase predicted a 4E grade and recovery of 122g/t and 87%, respectively. The composite worked on was from a similar location to the current cores and exhibited a 4E head grade of 2.6g/t. When comparing the Bryson test results of the current ore with the one obtained in the previous phase of testwork conducted at MINTEK; the grade recovery curve of the current ore was observed to be better than that of the previous ore; however this may be due to the low 4E head grade of 2.6g/t of the latter as compared to the current ore (5.3g/t) .

13.2.2 DFS Testwork on UG2 Ore at Mintek (August 2012)

Maseve Investments 11 (Pty) Ltd requested that MINTEK perform a series of confirmatory metallurgical testwork on a different set of UG2 drill cores extracted from PTM’s Elandsfontein deposit to provide information on the overall expected flotation performance.

The closest approximation of MF2 plant performance was achieved from a MF2 locked cycle test, which indicated that it was possible to achieve a 4E grade and recovery of 109g/t and 79%, respectively, at a mass pull of 3.2% . The Cr₂O₃ grade was 3.7% .

Seven UG2 drill core samples (each containing 1-3 deflections) were supplied and the brief given was to assess the upgrading response using a MF2 circuit at an overall grind of 80% -75µm. There was a large variation in terms of the 4E head grades of the individual drill cores extracted from the deposit. The 4E grades varied between 2.6 and 4.8g/t at an average Pt/Pd ratio of 2.2.

Grindmill tests were conducted at two feed sizes of -6.7 mm and -19.0 mm. The cumulative percent passing 75μm obtained at the highest energy of 40 kWh/t was found to be 92% and 90% using -6.7 mm and -19.0 mm feed respectively. The Bond ball mill test was conducted at 75μm limiting screen. The Bond work index was found to be 15.6 kWh/t, indicating that the UG2 sample could be classified as being hard.

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Variability test were performed on each of the individual cores using the conditions below: Theses variability test showed that there was a high degree of variability (i.e. in terms of grade and recovery) across the entire ore body. The flotation kinetics of the cores were variable with the overall 4E recovery ranging between 72% and 93%.

Four rougher rate tests conducted on the composite sample at various grinds to determine the effect of grind on the 4E kinetics demonstrated that the kinetics and recovery of the valuables improved with an increase in fineness of grind. For the MF1 rougher rates tests, an increase in the fineness of grind resulted in an increase in the 4E kinetics and recovery. The overall 20 minute 4E recovery increased from 84% at a grind of 40%-75μm to 87% at a grind of 80%-75μm. However, the 4E kinetic obtained for the MF2 circuit was much better as it was possible to recover 85% of the PGM’s after 10 minutes as opposed to 20 minutes using the MF1 circuit. The overall 4E recovery obtained for the MF2 circuit was 93%.

A total four MF2 cleaner rate tests were conducted to determine the effect of secondary grind and depressant dosage on the 4E kinetics:

>	MF2 (40-60%-75μm) – depressant dosage of 20g/t and 40g/t Sendep 369.
>	MF2 (40-80%-75μm) – depressant dosage of 20g/t and 40g/t Sendep 369.

The cleaner rate tests performed at various secondary grinds and depressant dosage indicated that the 4E recovery at finer grinds was sensitive to gangue depressant dosage. For the coarser secondary grind (40-60%-75μm), at higher depressant dosages, there was a marginal improvement in 4E grade. However this resulted in a negative effect on the overall 4E recovery (3% less) to the cleaner concentrate.

At a finer secondary grind (40-80%-75μm), increasing the depressant dosage resulted in a marginal improvement of the 4E grade and does not appear to have any significant impact on the overall 4E recovery to the cleaner concentrate. The poor grade recovery curves obtained for all the tests were a result of the slow floating material from the secondary rougher and that the primary rougher concentrates were not cleaned. The most favourable results were obtained using finer secondary grind at a higher depressant dosage of 40g/t which resulted in an overall 4E grade and recovery of 28g/t and 86% respectively.

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Bryson tests (open circuit cleaner test) were performed to more accurately predict the overall concentration grade and recovery to be achieved from a full-scale plant (in the absence of a circulating load in the flotation circuit). The target results from the client were an overall flotation concentrate grade of 102g/t 4E at a recovery of 86%. A total of three Bryson tests were conducted using the conditions outlined below:

For the first Bryson test (Test 1), it was possible to recover 76% of the PGM’s at a 4E grade of 89g/t (at a mass pull of 3.7%) at a head grade of 4.3g/t 4E, which was below the target conditions. The PGM losses were 7% to the secondary cleaner tails and 9% to the rougher tails.

For the second Bryson test (Test 2), the following conditions were changed in order to improve overall grade and recovery of the final concentrate:

>	The depressant dosage of the primary re-cleaner stage and secondary re-cleaner stage was reduced to improve the recovery of the primary and secondary re-cleaner concentrate.
>	The depressant dosage of the secondary cleaner stage was increased in order to improve the selectivity of the secondary cleaner concentrate.

A final concentrate with an overall concentrate 4E grade and recovery of 112g/t and 75% respectively at a mass pull of 2.9% was obtained. The PGM losses were 9% to the secondary cleaner tails and 9% to the rougher tails.

For the third Bryson test (Test 3), the following float conditions were changed in order to improve overall grade and recovery of the final concentrate:

>	The depressant dosage of the secondary cleaner stage was increased in order to improve the selectivity of the secondary cleaner concentrate.
>	The depressant dosage of the secondary re-cleaner stage was reduced to improve the recovery of the secondary re-cleaner concentrate.

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A final concentrate with an improved overall concentrate 4E grade and a marginal recovery of 117g/t and 74% respectively at a mass pull of 2.8% was obtained. The PGM losses were 10% to the secondary cleaner tails and 10% to the rougher tails.

The three Bryson Tests performed at different conditions yielded the highest overall PGM grade and recovery of 117g/t and 74% respectively and a Cr₂O₃ grade of 3.7% . This was below the targeted 4E grades and recoveries of 102g/t and 86% respectively. However it was expected that performing a locked cycle test on the sample would improve the overall 4E grades and recoveries (as the primary and secondary re-cleaner tails would be recycled).

Finer grinding of the secondary rougher material and secondary cleaner tail (SCT) to improve the liberation (as the PGM species are associated/ interlocked with gangue composites) and introducing a secondary cleaner scavenger may assist in improving the overall 4E grade and recovery.

13.3 Merensky Recovery Prediction and Benchmarking

13.3.1 MF1 Circuit

After the testwork conducted at SGS Lakefield during 2006 – 2007, the testwork information was used by Eurus Mineral Consultants (EMC) during May 2008 in order to predict the expected flotation circuit performance in terms of final grades and recoveries. Simulations were conducted at an average composite head grade of 2.67g/t 4E, 0.073% Cu and 0.151% Ni on MF1 and MF2 circuits.

The following grades and recoveries were predicted by the MF1 model:

>	In a MF1 Plant at a grind of 81-88% -75µm, a recovery of 89% 4E at a grade of 152g/t was predicted; while a Cu recovery of 84.3% at 3.94% grade and a Ni recovery of 70% at 6.76% grade was predicted,
>	In a MF1 Plant at a grind of 61% -75µm, a recovery of 86.2% 4E at a grade of 91g/t was predicted; while a Cu recovery of 90% at 2.6% grade and a Ni recovery of 67% at 4% grade was predicted,
>	In a MF1 Plant at a coarse grind of 33% -75µm, a recovery of 86.4% 4E at a grade of 78g/t was predicted; while a Cu recovery of 84% at 2.07% grade and a Ni recovery of 63% at 3.21% grade was predicted.

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Expected plant feed grades for the Merensky ore ranges from 3.7g/t 4E to 5.2g/t 4E with an average of 4.6g/t 4E, which is higher that the head grade used for the above EMC simulation work.

From the above information it is reasonable to expect a 4E recovery of 86% when operating as a MF1 circuit at a grind of 60% - 75µm, at a grade higher than 91g/t (due to a higher head grade). The 2009 testwork at MINTEK reported a MF1 Bryson test, at a grind of 40% - 75µm, on a 2.5g/t head grade sample achieving a 4E recovery of 86% at a grade of 60g/t. It was also proven that finer grinds resulted in higher recoveries. Further testwork by Mintek in 2012 on a MF1 circuit with a grind of 60% - 75µm reported 4E recoveries of 88% at a grade of 85g/t after a single cleaning stage, on a head grade of 5.6g/t 4E.

13.3.2 MF2 Circuit

The EMC MF2 model, at a primary grind of 40% -75µm and a final grind of 90%-75µm, predicted final 4E recoveries ranging from 80.0% to 85.1% at corresponding grades of 131g/t 4E to 110g/t 4E, depending on the depressant dosages.

A benchmarking exercise on the adjacent Styldrift deposit suggested that 87.5% 4E recovery could be expected at 150g/t 4E grade using a MF2 circuit with a final grind of 80% -75µm. 50% recovery of the nickel and 60% recovery of the copper were also predicted.

The neighbouring BRPM mine had published in 2005 an achieved head grade of 4.47g/t delivered to the processing plant with a recovery of 85.83% for MR alone for the production year 2004. Prior to this production year recoveries were poor, as the UG2 open pit was in operation and this downgraded the recovery and is therefore not applicable to the only MR operation. The Wesizwe project has a predicted recovery for MR of 88%, although with a number of cautionary statements.

The BRPM published results are regarded as an indication of the concentrator performance that can be achieved. The anticipated head grade to be delivered to the WBJV mine will average 4.65g/t (after MCF correction). Analysing the BRPM performance, the tailings value will be approximately 0.65g/t. There is a strong relationship between head grade and recovery on platinum operations such that the tailings values remain relatively constant – this means that as the head grade increases, the recovery increases. Applying this metallurgical knowledge to the WBJV MR feed, it can be expected that the recovery will be between 86% and 88% with tailings values of approximately 0.73 and 0.657g/t. The Mintek test work has indicated that an even lower tailings value may be achieved, thus offering upside recovery potential. The 2012 Mintek results reported a 4E recovery of 87% at a grade of 135g/t at a mass pull of 3.3%, with the copper grade 1.9% and the nickel grade 3.5% .

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13.4 UG2 Recovery Prediction and Benchmarking

The expected UG2 plant feed grades ranges from 2.3g/t 4E to 4g/t 4E with an average of 3.2g/t 4E.

13.4.1 MF1 Circuit

Not applicable.

13.4.2 MF2 Circuit

After the testwork conducted at SGS Lakefield during 2006 – 2007, the testwork information was used by Eurus Mineral Consultants (EMC) during May 2008 in order to predict the expected flotation circuit performance in terms of final grades and recoveries. Simulations were conducted at an average composite head grade of 3.71g/t 4E and 25.5% Cr₂O₃. The MF2 simulations predicted a final 4E recovery of 76.9% at a grade of 223g/t and 3.25% Cr₂O₃. However, work done by EMC in October 2008, on UG2 material predicted the following:

> Simulation based on a location composite sample with a 3.73g/t 4E head grade predicted a 4E recovery of 86% at a lower product grade of 113g/t 4E (3.48% Cr₂O₃) or a 4E recovery of 82.4% at a high product grade of 244g/t 4E (2.96% Cr₂O₃).

During the 2012 Mintek work, the closest approximation of MF2 Plant performance on the UG2 sample was achieved from a MF2 locked cycle test, which indicated that it was possible to achieve a 4E grade and recovery of 109g/t and 79%, respectively, at a mass pull of 3.2% . The Cr₂O₃ grade was 3.7%. The lower recovery obtained in the 2012 phase of work was possible caused by the coarser secondary rougher grind (40% and 80%-75μm) as compared to the finer secondary grind (40% and 90%- 75μm) used in the previous phase of work.

The current circuit design caters for a final grind of 80% – 75µm when treating only UG2 material.

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A benchmarking exercise performed on the neighbouring Styldrift deposit suggested that a final concentrate 4E grade of 150g/t could be achieved at a recovery of 83% and a Cr₂O₃ grade of less than 4%.

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14.0 MINERAL RESOURCE ESTIMATES

Table 14-1 summarises the Mineral Resource estimate for the Maseve Project at a 250cm 4E cut-off.

Table 14-1: Maseve Mineral Resource Estimation Tabulation

Merensky - Mining Cut
Resource Category	Cut-off	Tonnage	Grade					Metal		Reef Width
	4E	4E	Pt	Pd	Rh	Au	4E	kg	4E	4E
	cm.g/t	Mt	g/t	g/t	g/t	g/t	g/t	kg	Moz	cm
Measured	300	9.266	3.35	1.41	0.21	0.26	5.23	48,461	1.558	152
Indicated	300	12.552	3.65	1.54	0.23	0.29	5.71	71,672	2.304	141
Total	300	21.818	3.53	1.49	0.21	0.28	5.51	120,133	3.862	146
Inferred	300	0.196	2.32	0.98	0.14	0.18	3.62	710	0.023	118
UG2 - Mining Cut
Resource Category	Cut-off	Tonnage	Grade					Metal		Reef Width
	4E	4E	Pt	Pd	Rh	Au	4E	kg	4E	4E
	cm.g/t	Mt	g/t	g/t	g/t	g/t	g/t	kg	Moz	cm
Measured	300	8.496	2.29	0.94	0.36	0.04	3.63	30,841	0.992	140
Indicated	300	14.183	2.46	1.01	0.39	0.04	3.90	55,314	1.778	136
Total	300	22.679	2.39	0.99	0.38	0.04	3.80	86,155	2.770	137
Inferred	300	0	0	0	0	0	0	0	0	0

MR = Merensky Reef; UG2 = Upper Group No. 2 chromitite seam; 4E = Platinum Group Elements (Pd+Pt+Rh) and Au. The cut-offs for Mineral Resources have been established by a qualified person after a review of potential operating costs and other factors. The Mineral Resources stated above are shown on a 100% basis, that is, for the WBJV as a whole entity. Conversion Factor used – kg to oz = 32.15076.

The estimated SG values were used for the tonnage calculations; the average for the MR is 3.23t/m ³ and 3.60t/m ³ for the UG2. The SG represents the mining width intersections and not solely Chromitite for the UG2 and the pyroxenitic material for the MR.

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14.1 Key Assumptions and Parameters

Generation of the Maseve Resource was conducted using the following steps:

>	Coding of the drill holes to reflect the mining cut,
>	Statistical analysis to provide a basis for data verification and to establish specific information on population distributions and checks for anomalous values,
>	Variogram modelling for the grade values,
>	Kriging.

14.1.1 Data Used

The drill hole file received by CJM Consulting from the Client underwent six main steps before Mineral Resource estimation could be carried out, namely:

>	Reef coding,
>	Determination of Rh values and regression analysis,
>	Compositing with SG weighting,
>	Determination of reef cuts,
>	Conversion to true widths.

Figure 14-1 illustrates the location of the drill holes utilised in the estimation of Mineral Resources for the Project. For the 2015 update, an additional 213 drill holes were added to the project data.

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14.1.2 Reef Coding Methodology

The data was coded and then composited in Datamine. The reef coding was done under the following key guidelines:

>	The unit coded as MR and UG2 by the Client,
>	The sample length – where a sample was taken, the composites were not split,
>	The lithology with respect to the MR – any non-MR lithologies (norites, anorthosites etc.) coded as reef were failed unless the reef and non-MR lithology were collected as one sample.

After effecting the above four criteria to the reef intersections, the intersections underwent a “pass” or “fail” procedure where the intersection was failed as a valid MR intersection if:

a. Iron replacement occurred (an IRUP),

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b. The intersection was severely faulted.

Note that reference has been made to the MR only. However, the same coding strategy was applied to the UG2. Figure 14-2 shows the reef definition process.

14.1.3 Treatment of Holes Drilled through Area without Reef Development

A total of 366 MR drill hole intersections, and 415 UG2 drill hole intersections were available for this Resource update. To avoid extrapolating reef over those areas where reef was not intersected, it was attempted to domain such areas out of the model.

14.1.4 Determination of SG Values at Locations where SG Measurements were Incomplete

Since the compositing strategy required an SG weighting, the samples whose SG values were not measured were assigned an average SG for the relevant lithology. The average SG used per lithology is shown in Table 14-2 below. The tabulation is only for the units affected.

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Table 14-2: Average Specific Gravities per Lithology

Unit	FPYXCR	PEGFPXCR	CR	PEGFPYX	FPYX	LN	N	POIKAN	DOL
SG g/cc	3.60	3.69	3.80	3.25	3.28	2.95	3.00	2.92	3.21

Notes:	FPYXCR = Feldspathic pyroxenite with associated Chromitite layer
	PEGFPXCR = Feldspathic pegmatoidal pyroxenite with associated Chromitite layer
	CR = Chromitite
	PEGFPYX = Pegmatoidal feldspathic pyroxenite
	FPYX = Feldspathic pyroxenite
	LN = Leuconorite
	N = Norite
	POIKAN = Poikilitic Anorthosite
	DOL = Dolerite

Once all the samples had been assigned an SG, the compositing could be done as accurately as possible using the SG weighting.

14.1.5 Compositing Strategy

The samples utilized in the Mineral Resource estimation represent full reef composites, as estimation is conducted in Two Dimensions (2D). During the compositing phase, the drill hole intercepts were weighted with SG values.

14.1.6 Widths of Mineralised Zones – Resource Cuts

The methodology in determining the resource cuts is derived from the core intersections. Generally, the high grade portions of the reefs are about 60cm thick. For both the MR and UG2, the marker unit is the bottom reef contact, which is a chromite contact of less than a centimetre. The cut is taken from that chromite contact and extended vertically to accommodate most of the metal content. If this should result in a resource cut less than 100cm up from the bottom reef contact, it is extended further to 100cm.

In the case of the UG2, the triplets (if and where developed and within 30cm from the top contact) are included in the resource cut.

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14.2 Data Analysis

The drill hole file received by CJM Consulting from the Client underwent six main steps before Mineral Resource estimation could be carried out, namely:

>	Reef coding,
>	Determination of Rh values and regression analysis,
>	Compositing with SG weighting,
>	Determination of reef cuts,
>	Conversion to true widths.

A total number of 669 drill holes were supplied by the Client, of which 213 holes were drilled since the previous Resource Estimate. These additional holes were employed to update the Mineral Resources for the Project. The data was supplied as unsecured Microsoft Excel files which were then imported into Datamine. Figure 14-3 illustrates the drill hole data available for the Mineral Resource update.

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The QP’s of this report have used the data provided by the representative and internal experts of PTM. This data is derived from historical records for the area as well as information currently compiled by the operating company, which is PTM. The PTM-generated information is under the control and care of Mr T Botha, who is an employee of PTM and is not independent. Table 14-3 summarises the available data that was used for the geostatistical modelling and evaluation of the project.

Table 14-3: Drillhole Data

Data	Merensky Reef	UG2 Reef
No. of Drill hole intersections	366	415

The locations of the drill holes are illustrated in Figure 14-4 and Figure 14-5.

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14.3 Data Verification

Data verification was conducted by CJM Consulting as part of the Mineral Resource estimation for the WBJV Project 1 (Maseve).

14.3.1 Personal Data Verification by the Qualified Persons who Carried out Estimation

The drill holes used in the estimation of the Project Mineral Resources were personally inspected by the QP of this Report and thus the reliability, consistency and representativeness of the data was established. This verification included a visit to the PTM core yard and visual inspection of the drill hole core as well as verification of the hand written and digital copies of the logs. There were three noted discrepancies where the handwritten logs did not correlate with the digital copies. In this case there appears to have been a reclogging exercise which had not yet been captured into the database. CJM Consulting carried out validation of the data by checking for overlaps, missing values etc. as detailed in this report.

14.4 Quality Assurance and Quality Control Data

The Quality Assurance and Quality Control data and results are detailed in another Section of this report. QA and QC was verified and is sufficient for resource estimation.

14.5 Population Statistics

Statistical analyses were performed to develop an understanding of the statistical characteristics and sample population distribution relationships. Descriptive statistics in the form of histograms (frequency distributions) and probability plots (used to evaluate the normality of the distribution of a variable) were used to develop an understanding of such statistical relationships.

14.5.1 Descriptive Statistics

The descriptive statistics tables are illustrated per Geo Zone. The MR was subdivided into eleven Geo Zones.

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Table 14-4: Statistical Analysis – Drill Hole Data per Geo Zone: MR

MR Geo Zone 1 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	132	2.87	3.51	3.2	0	0.09	0.03
CM4E	132	3.83	2 000.00	555.78	222544	471.75	0.85
CW	132	53.65	300	106.02	1569	39.61	0.37
PGE4	132	0.04	16	5.14	14	3.8	0.74
MR Geo Zone 2 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	14	2.9	3.24	3.13	0	0.1	0.03
CM4E	14	3.87	300	196.79	10769	103.77	0.53
CW	14	93.31	160.97	108.21	587	24.23	0.22
PGE4	14	0.04	8.34	2.92	7	2.63	0.9
MR Geo Zone 3 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	8	3.06	3.33	3.21	0	0.09	0.03
CM4E	8	16.78	192.76	98.32	3795	61.61	0.63
CW	8	98.19	233.64	115.8	2267	47.62	0.41
PGE4	8	0.07	1.93	0.98	0	0.64	0.65
MR Geo Zone 4 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	22	3.02	3.2	3.17	0	0.05	0.02
CM4E	22	3.19	3 000.00	773.36	929211	963.96	1.25
CW	22	69.93	300	124.49	5791	76.1	0.61
PGE4	22	0.04	17.98	5.38	25	5.02	0.93

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MR Geo Zone 5 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	7	2.99	3.2	3.11	0	0.09	0.03
CM4E	7	5.29	1 029.98	384.46	115239	339.47	0.88
CW	7	87.18	165.85	100.85	830	28.82	0.29
PGE4	7	0.06	6.21	3.55	6	2.38	0.67
MR Geo Zone 6 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	64	2.89	3.35	3.14	0	0.13	0.04
CM4E	64	19.85	2 200.00	761.1	347933	589.86	0.78
CW	64	81.03	300	126.75	3709	60.9	0.48
PGE4	64	0.2	15.37	5.79	13	3.67	0.63
MR Geo Zone 7 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	15	2.86	3.28	3.16	0	0.11	0.03
CM4E	15	6.17	744.17	259.07	48369	219.93	0.85
CW	15	89.78	142.89	98.5	215	14.68	0.15
PGE4	15	0.07	7.66	2.51	4	1.98	0.79
MR Geo Zone 8 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	40	2.89	3.26	3.15	0	0.1	0.03
CM4E	40	4	800	338.6	61195	247.38	0.73
CW	40	44.51	165.75	94.22	237	15.39	0.16
PGE4	40	0.04	13.74	3.82	10	3.12	0.82

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MR Geo Zone 9 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	37	2.89	3.3	3.14	0	0.11	0.03
CM4E	37	5.28	3 000.00	770.03	617742	785.97	1.02
CW	37	73.99	300	130.57	3696	60.79	0.47
PGE4	37	0.06	14	5.31	22	4.71	0.89
MR Geo Zone 10 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	14	3.14	3.34	3.22	0	0.05	0.02
CM4E	14	177.9	1 500.00	890.95	249018	499.02	0.56
CW	14	95.54	200	140.45	1888	43.45	0.31
PGE4	14	1.82	8	5.96	6	2.41	0.4

The descriptive statistics tables are illustrated per Geo Zone. The UG2 Reef was subdivided into eight Geo Zones. Table 14-5 summarises the descriptive statistics for UG2 Reef.

Table 14-5: Statistical Analysis – Drill Hole Data per Geo Zone: UG2

UG2 Geo Zone 1 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	82	2.99	4.17	3.77	0.05	0.23	0.06
CM4E	82	2.49	1,889.70	455.06	130,204.19	360.84	0.79
CW	82	1.88	300.00	127.21	3,142.65	56.06	0.44
PGE4	82	0.04	7.93	3.33	2.90	1.70	0.51

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UG2 Geo Zone 2 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	48	2.97	4.24	3.75	0.06	0.25	0.07
CM4E	48	19.85	1,001.38	383.22	68,026.29	260.82	0.68
CW	48	56.60	242.20	118.36	1,239.93	35.21	0.30
PGE4	48	0.21	6.28	3.01	2.67	1.63	0.54
UG2 Geo Zone 3 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	51	3.09	3.99	3.67	0.05	0.23	0.06
CM4E	51	16.99	1,100.00	445.02	107,653.66	328.11	0.74
CW	51	69.11	250.00	129.97	2,803.11	52.94	0.41
PGE4	51	0.18	6.47	3.12	2.29	1.51	0.49
UG2 Geo Zone 4 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	72	3.21	4.09	3.80	0.01	0.12	0.03
CM4E	72	34.31	1,500.00	623.12	159,248.68	399.06	0.64
CW	72	83.05	300.00	156.14	4,429.25	66.55	0.43
PGE4	72	0.41	7.37	3.80	2.42	1.56	0.41
UG2 Geo Zone 5 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	78	2.83	4.15	3.60	0.10	0.32	0.09
CM4E	78	9.06	2,808.48	269.52	129,266.54	359.54	1.33
CW	78	72.85	236.89	99.62	561.97	23.71	0.24
PGE4	78	0.10	32.57	2.61	14.45	3.80	1.46

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UG2 Geo Zone 6 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	43	2.82	4.00	3.60	0.09	0.31	0.09
CM4E	43	27.31	900.00	333.64	72,801.98	269.82	0.81
CW	43	62.16	450.60	125.08	6,213.68	78.83	0.63
PGE4	43	0.30	6.09	2.59	1.77	1.33	0.51
UG2 Geo Zone 7 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	9	2.99	3.80	3.67	0.08	0.29	0.08
CM4E	9	22.02	61.39	35.91	209.91	14.49	0.40
CW	9	42.84	99.99	84.28	290.23	17.04	0.20
PGE4	9	0.26	0.71	0.43	0.02	0.15	0.35
UG2 Geo Zone 8 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	36	2.82	3.98	3.56	0.12	0.35	0.10
CM4E	36	0.37	974.41	189.57	54,065.10	232.52	1.23
CW	36	3.69	240.76	100.10	1,472.02	38.37	0.38
PGE4	36	0.04	7.63	1.72	3.80	1.95	1.13

Table 14-6 details the descriptive statistics for the footwall (“FW”) units that were also modelled to achieve the best mining cut width.

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Table 14-6: Descriptive Statistics for the MRFW Units

MR Footwall 1 Geo Zone 1 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	407	2.61	3.43	2.87	0.01	0.11	0.04
CM4E	405	0.07	1,095.38	48.52	8,998.88	94.86	1.96
CW	407	14.47	19.98	18.64	1.50	1.22	0.07
PGE4	405	0.00	65.46	2.62	28.03	5.29	2.02
MR Footwall 2 Geo Zone 1 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	407	2.75	3.50	2.87	0.01	0.11	0.04
CM4E	405	0.07	451.20	34.03	3,871.99	62.23	1.83
CW	407	14.47	19.98	18.58	1.42	1.19	0.06
PGE4	405	0.00	26.96	1.82	11.31	3.36	1.84
MR Footwall 3 Geo Zone 1 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	407	2.74	3.62	2.87	0.01	0.11	0.04
CM4E	405	0.06	322.73	22.89	2,592.50	50.92	2.22
CW	407	14.47	19.98	18.57	1.36	1.17	0.06
PGE4	405	0.00	16.87	1.22	7.31	2.70	2.21
MR Footwall 4 Geo Zone 1 Descriptive Statistics
Parameter	Number of Records	Minimum	Maximum	Average	Variance	Standard Deviation	CoV
SG	407	2.73	3.56	2.86	0.01	0.11	0.04
CM4E	405	0.16	404.60	17.49	2,144.67	46.31	2.65
CW	407	14.47	19.98	18.57	1.35	1.16	0.06
PGE4	405	0.01	20.33	0.93	5.91	2.43	2.60

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14.5.2 Histograms and Probability Plots for Holes within Lease Area

Histograms were derived to develop an understanding of the sample population distribution relationships. Probability plots were used to evaluate the normality of the distribution of the variables estimated, refer to 0 APPENDICES.

14.5.3 Capping and Cutting

No corrections were made (top cut, etc.) to the data as all fall within acceptable limits. The statistical analyses show the expected relationships for these types of reef.

14.6 Mining Considerations

As only mineral resources were evaluated, no mine plan was taken into consideration. The reefs were, however, evaluated over a realistic average diluted mining width as follows.

Table 14-7: Summary of Mining Widths

Reef	Project	Resource Category	Mining Width (cm)
MR	Maseve	Measured	152
		Indicated	141
		Inferred	118
UG2	Maseve	Measured	140
		Indicated	136
		Inferred	-

The initial occurrence of payable reef is at 130m below surface. The payable reef then extends to a depth of approximately 630m below surface. There are no near-surface or outcropping reefs of sufficient width, geometry or value to provide an opportunity for opencast mining.

A cut-off grade of 300cm.g/t was selected as a resource cut-off. The reason for using the 300cm.g/t cut-off is in compliance with responsible engineering practice to simulate probable working cost and flow of ore parameters, in order to report potentially economical resources.

14.7 Reef Compositing Definitions

For the Resource cut width model, the borehole samples were composited into single drillhole intersections according to the pertinent Resource width. Density was incorporated in the compositing phase. The boreholes were subsequently corrected for dip (3D dip model constructed), resulting in true width intersection values. The data was composited according to reef horizon.

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14.8 Mineral Resource Modelling

14.8.1 Modelling Methodology

Modelling was carried out using Datamine Studio^TM and Minesoft’s geostatistical package ‘RES’.

Full reef composite data – mining cut width (cm), 4E content (cm.g/t), and SG (t/m³) were estimated for both the MR and UG2, and three MR Footwall Units. Ordinary Kriging (“OK”) techniques were used.

The 4E grade concentration (g/t) was calculated from the interpolated kriged 4E content and reef width values. Detailed checks were carried out to validate kriging outputs, including input data, kriged estimates and kriging efficiency checks.

14.8.2 Variography

Variograms are a useful tool for investigating the spatial relationships of samples. Variograms for channel width (cm), 4E (4E Grade), SG and CM4E (4E Content) were modelled during the estimation process. All variograms are omni-directional spherical semi-variograms.

Table 14-8 summarises the 4E content variograms which have a modelled grade continuity range of ~150 – 300m for the MR and a modelled grade continuity range of ~200 - 420m for the UG2 Reef. The nugget effect is on average 37% of the sill or population variance for the MR and 45% for the UG2 Reef. This is slightly high, but it is expected that these ranges will improve as more data becomes available. No top-cuts were used for the generation of the experimental variograms. Parameters for the remaining contents and elements are available but were omitted from the report.

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Table 14-8: Variogram Parameters

					Structure1				Structure2
Reef	Parameter	Domain	TotalSill	Nugget Percentage (%)	Ran gel (m)	Ran ge2 (m)	Ran ge3 (m)	Sill1	Ran ge1 (m)	Ran ge2 (m)	Ran ge3 (m)	Sill2
M R	CMPt	1	120 916.52	45.18	120	120	1	47 213.45	350	450	1	19 076.48
M R	CMPt	2	71 258.54	45.18	101	101	1	27 823.84	320	320	1	11 242.15
M R	CMPt	3	1 806.97	45.18	101	101	1	705.55	320	320	1	285.08
M R	CMPt	4	888 218.22	45.18	101	101	1	346 816.54	320	320	1	140 130.34
M R	CMPt	5	43 581.11	45.18	101	101	1	17 016.82	320	320	1	6 875.60
M R	CMPt	6	209 499.09	45.18	120	120	1	81 801.69	320	450	1	33 051.76
M R	CMPt	7	20 750.18	45.18	101	101	1	8 102.18	320	320	1	3 273.67
M R	CMPt	8	36 154.33	45.18	101	101	1	14 116.94	320	320	1	5 703.91
M R	CMPt	9	323 885.04	45.18	101	101	1	126 465.19	320	320	1	51 097.94
M R	CMPt	10	360 361.31	45.18	120	120	1	140 707.83	350	450	1	56 852.64
M R	CW	1	703.31	45.18	120	120	1	274.62	320	320	1	110.96
M R	CW	2	586.92	45.18	101	101	1	229.17	320	320	1	92.6
M R	CW	3	2 267.38	45.18	101	101	1	885.33	320	320	1	357.71
M R	CW	4	812.27	45.18	101	101	1	317.16	320	320	1	128.15
M R	CW	5	830.47	45.18	101	101	1	324.27	320	320	1	131.02
M R	CW	6	2 319.39	45.18	120	120	1	905.64	320	450	1	365.92
M R	CW	7	215.4	45.18	101	101	1	84.1	320	320	1	33.98
M R	CW	8	236.77	45.18	101	101	1	92.45	320	320	1	37.35
M R	CW	9	2 128.07	45.18	101	101	1	830.93	320	320	1	335.74
M R	CW	10	1 100.14	45.18	120	120	1	429.56	350	350	1	173.56
M R	SG	1	0.01	45.17	101	101	1	0	320	320	1	0
M R	SG	2	0.01	45.17	101	101	1	0	320	320	1	0
M R	SG	3	0.01	45.17	101	101	1	0	320	320	1	0
M R	SG	4	0	45.18	101	101	1	0	320	320	1	0
M R	SG	5	0.01	45.18	101	101	1	0	320	320	1	0
M R	SG	6	0.02	45.18	120	120	1	0.01	320	450	1	0
M R	SG	7	0.01	45.17	101	101	1	0	320	320	1	0

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					Structure1				Structure2
Reef	Parameter	Domain	TotalSill	Nugget Percentage (%)	Ran ge1 (m)	Ran ge2 (m)	Ran ge3 (m)	Sill1	Ran ge1 (m)	Ran ge2 (m)	Ran ge3 (m)	Sill2
M R	SG	8	0.01	45.17	101	101	1	0	320	320	1	0
M R	SG	9	0.01	45.18	101	101	1	0	320	320	1	0
M R	SG	10	0	45.17	120	120	1	0	320	450	1	0
M R	CMAu	1	448.4	45.18	120	120	1	175.08	350	450	1	70.74
M R	CMAu	2	392.07	45.18	101	101	1	153.09	320	320	1	61.85
M R	CMAu	3	43.05	45.18	101	101	1	16.81	320	320	1	6.79
M R	CMAu	4	3 305.29	45.18	101	101	1	1 290.59	320	320	1	521.46
M R	CMAu	5	645.99	45.18	101	101	1	252.24	320	320	1	101.92
M R	CMAu	6	599.36	45.18	120	120	1	234.03	320	450	1	94.56
M R	CMAu	7	133.84	45.18	101	101	1	52.26	320	320	1	21.12
M R	CMAu	8	338.26	45.18	101	101	1	132.08	320	320	1	53.37
M R	CMAu	9	1 483.21	45.18	101	101	1	579.14	320	320	1	234
M R	CMAu	10	1 022.05	45.18	120	120	1	399.07	350	450	1	161.24
M R	CMR h	1	454.39	45.18	120	120	1	177.42	350	450	1	71.69
M R	CMR h	2	312.29	45.18	101	101	1	121.94	320	320	1	49.27
M R	CMR h	3	5.72	45.18	101	101	1	2.23	320	320	1	0.9
M R	CMR h	4	1 276.59	45.18	101	101	1	498.46	320	320	1	201.4
M R	CMR h	5	89.33	45.18	101	101	1	34.88	320	320	1	14.09
M R	CMR h	6	881.97	45.18	120	120	1	344.38	320	450	1	139.14
M R	CMR h	7	60.83	45.18	101	101	1	23.75	320	320	1	9.6
M R	CMR h	8	104.15	45.18	101	101	1	40.67	320	320	1	16.43
M R	CMR h	9	820.5	45.18	101	101	1	320.37	320	320	1	129.45
M R	CMR h	10	2 103.07	45.18	120	120	1	821.17	350	450	1	331.79
M R	CMPd	1	22 606.20	45.18	120	120	1	8 826.89	350	450	1	3 566.48
M R	CMPd	2	11 566.33	45.18	101	101	1	4 516.23	320	320	1	1 824.77
M R	CMPd	3	207.35	45.18	101	101	1	80.96	320	320	1	32.71
M R	CMPd	4	149 176.09	45.18	101	101	1	58 247.78	320	320	1	23 534.87
M R	CMPd	5	9 489.31	45.18	101	101	1	3 705.23	320	320	1	1 497.09

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					Structure1				Structure2
Reef	Parameter	Domain	TotalSill	Nugget Percentage (%)	Ran ge1 (m)	Ran ge2 (m)	Ran ge3 (m)	Sill1	Ran ge1 (m)	Ran ge2 (m)	Ran ge3 (m)	Sill2
M R	CMPd	6	39 385.69	45.18	120	120	1	15 378.66	320	450	1	6 213.71
M R	CMPd	7	3 437.95	45.18	101	101	1	1 342.39	320	320	1	542.39
M R	CMPd	8	6 390.09	45.18	101	101	1	2 495.09	320	320	1	1 008.14
M R	CMPd	9	52 342.46	45.18	101	101	1	20 437.80	320	320	1	8 257.84
M R	CMPd	10	57 413.84	45.18	120	120	1	22 417.99	350	450	1	9 057.93
M R	CMP GE4E	1	161 806.67	45.18	120	120	1	63 179.55	350	450	1	25 527.54
M R	C0MP GE4E	2	166 907.08	45.18	101	101	1	65 171.08	320	320	1	26 332.21
M R	CMP GE4E	3	3 795.26	45.18	101	101	1	1 481.91	320	320	1	598.76
M R	CMP GE4E	4	452 946.22	45.18	101	101	1	176 858.84	320	320	1	71 459.36
M R	CMP GE4E	5	115 238.72	45.18	101	101	1	44 996.48	320	320	1	18 180.71
M R	CMP GE4E	6	289 267.38	45.18	120	120	1	112 948.27	320	450	1	45 636.46
M R	CMP GE4E	7	48 368.93	45.18	101	101	1	18 886.29	320	320	1	7 630.96
M R	CMP GE4E	8	47 115.99	45.18	101	101	1	18 397.06	320	320	1	7 433.29
M R	CMP GE4E	9	489 366.36	45.18	101	101	1	191 079.56	320	320	1	77 205.21
M R	CMP GE4E	10	182 089.73	45.18	120	120	1	71 099.34	350	450	1	28 727.51

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14.9 Grade Estimation

14.9.1 Modelling Methodology

The following applies to the Mineral Resource area and was undertaken using Minesoft (Pty) Ltd’s ‘RES’ geostatistical program. The following parameters were used in the kriging process for both Project Areas:

>	Full reef composite data – Mining width (cm), content (4E) and SG,
>	100m x 100m x 1m block size,
>	2D estimation was conducted as the tabular nature of the ore body suited this type of estimation rather than 3D estimation,
>	Discretisation – 5 x 5 x 1 for each 100m x 100m x 1m block,
>	First search volume – on average 1.5 times variogram range up to a maximum of 1 100m:
>	Minimum number of samples 8,
>	Maximum number of samples 40.
>	Second search volume:
>	Minimum number of samples 4,
>	Maximum number of samples 40.
>	Third search volume:

• Minimum number of samples 1,

• Maximum number of samples 20.

>	Interpolation methods – ordinary kriging,
>	Metal grade (4E g/t) was calculated from metal content and mining width/thickness,

The following explains the terminology of certain of the parameters that were used in the kriging process:

• Search range – As range of variogram decreases to approach zero (pure nugget), the required neighbourhood configuration for good estimation will become progressively larger, and vice versa. A limited search range will result in a block estimate that is progressively uncorrelated to the true grade as the variogram range tends to zero. Using the variogram range or slightly larger than variogram range allows the search volume to have a long range relative to the block dimensions, thereby accessing samples particularly in areas of data scarcity.

• Discretisation – Used to divide the block area into many points to allow improved block estimates from point data. The block is divided into many points and then individual point estimates are averaged to get an average over the block. Spatial locality of point data relative to the block to be estimated is hence entertained.

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• Parent cell estimation – When the block model is created, sub-celling of the parent cells is used to allow for an improved representation of the volume. The grade of the parent cell is estimated and that value is assigned to all the (sub) cells inside the parent.

• Negative kriging weights – at the edges of the ore body / domains, the kriging weights will be small, even negative. The distance required to search before negative weights are encountered progressively increases as the nugget increases. In general, negative weights are not problematic in an estimation model if the number of negative weights is a small proportion, typically less than 5%. Re-setting the negative weights to zero allows conditional bias to be incorporated in the estimation exercise.

The first 50m of the ore body is considered to represent a weathered zone and is discarded in the modelling and estimation procedures. The kriged estimates were not post-processed to calculate the information effect, dispersion variance and grade tonnage intervals. The 4E cutoff values used ranged from 100 – 600cm.g/t.

14.10 Mineral Resource Classification

The Mineral Resource classification is a function of the confidence of the whole process from drilling, sampling, geological understanding and geostatistical relationships. The following aspects or parameters were considered for resource classification:

>	Sampling – QA and QC,
>	Measured: high confidence, no problem areas,
>	Indicated: high confidence, some problem areas with low risk,
>	Inferred: some aspects might be of medium to high risk,
>	Geological Confidence,
>	Measured: high confidence in the understanding of geological relationships, continuity of geological trends and sufficient data,
>	Indicated: good understanding of geological relationships,
>	Inferred: geological continuity not established,
>	Number of samples used to estimate a specific block,
>	Measured: at least eight drill holes within semi-variogram range and minimum of twenty 1m composited samples,

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>	Indicated: at least four drill holes within semi-variogram range and a minimum of twelve 1m composite samples,
>	Inferred: less than three drill holes within the semi-variogram range,
>	Distance to Sample (Semi-variogram Range),
>	Measured: at least within 60% of semi-variogram range,
>	Indicated: within semi-variogram range,
>	Inferred: further than semi-variogram range,
>	Lower confidence limit (Blocks),
>	Measured: < 20% from mean (80% confidence),
>	Indicated: 20% – 40% from mean (60% – 80% confidence),
>	Inferred: more than 40% (less than 60% confidence),
>	Kriging Efficiency,
>	Measured: > 40%,
>	Indicated: 20 – 40%,
>	Inferred: < 20%,
>	Deviation from Lower 90% Confidence Limit (Data Distribution within Resource Area Considered for Classification),
>	< 10% deviation from the mean – Measured Resource,
>	10 – 20% deviation from the mean – Indicated Resource,
>	> 20% deviation from the mean – Inferred Resource.

Using the above criteria, the MR and UG2 within the Project Area was classified as Inferred, Indicated and Measured Mineral Resources.

14.11 Model Results

14.11.1 Model Plans and Sections

The MR and UG2 block models were plotted to show which search volumes were used for the estimation. As seen in Figure 14-6 and Figure 14-7, the entire Project Area was estimated using the first search volume, i.e. the ranges used were within 1.5 times the variogram ranges. Only to the shallow weathered area towards where the reef outcrops and also towards where the ore is weathered, is the second search volume employed. The number of samples used for an estimation of a block indicates that the majority of the blocks have sufficient data for estimation purposes. The regression slope plots are taken from the ordinary kriging exercise; essentially, for a Measured classification.

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With regard to kriging efficiency, a measure of the accuracy and precision of the kriged estimates, a block with a kriging efficiency of greater than 40% is considered to qualify for the classification of measured status.

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The mineral resource categories for the MR and UG2 are Figure 14-14 and Figure 14-15 shown in respectively.

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The mineral resource categories have changed as a consequence of improved and more detailed drill hole coding and more confidence in the treatment of faults and IRUP’s.

14.12 Compliance with Disclosure Requirements for Mineral Resources

This Technical Report utilizes definitions within the NI 43-101 and the Resource classifications set out in the South African Code for the Reporting of Exploration Results, Mineral Resource and Mineral Reserves (“SAMREC Code”), relating to Maseve Project area.

The mineral resource pertains to 4E Content at a 300cm.g/t cut-off. A cut-off content of 300cm.g/t was selected as a Mineral Resource cut-off. The reason for using the 300cm.g/t cut-off is in compliance with responsible engineering practice to simulate probable working cost and flow of ore parameters, in order to report potentially economical resources.

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Only mineral resources have been estimated for this report. All mineral resources have been classified as measured, indicated and inferred mineral resources, according to the definitions of the SAMREC code.

Inferred mineral resources have been classified. However, no addition of the inferred mineral resources to other mineral resource categories has taken place.

There are differences between SAMREC and CIM standards for mineral resources to be used for NI43-101 standard reporting. However in this case the differences are not material and can be considered the same.

14.13 Effect of Modifying Factors

No modifying factors such as taxation, socio-economic, marketing or political factors have been taken into account. No environmental, permitting, legal or title factors will affect the estimated mineral resource. The mine is in development with the necessary permits.

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15.0	MINERAL RESERVE ESTIMATES
15.1	Resource to Reserve Calculation

15.1.1 Methods Used

During the planning, mine design and scheduling phase Datamine’s Studio-5D planner and Enhanced Production Scheduler (“EPS”) were used in order to obtain the mine design and production schedule.

Below is the summary of the mining resource for the MR broken up per mining block as per block model.

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Table 15-1: Block Model per Mining Blocks - MR

Cut-off Grade	2.50			Resource
Block Number	Average Seam Height - Total	Average Seam Height - Footwall	Average Dip	Volume	Tonnes	Content	Grade	Geo- loss %
0	119	3	28.4	50 268	158 628	661 555	4.17	14%
1	108	-	18.4	268 548	838 804	2 349 426	2.80	11%
2	129	11	24.9	970 717	3 107 672	18 762 482	6.04	13%
3	119	-	10.5	179 884	572 886	3 456 236	6.03	14%
4	131	22	20.5	182 021	574 177	2 825 947	4.92	14%
5	136	17	10.7	851 282	2 677 629	14 063 475	5.25	14%
6	111	7	28.5	77 517	243 974	1 112 723	4.56	14%
7	105	-	19.1	41 193	129 709	406 078	3.13	14%
8	99	-	19.4	17 174	54 029	160 971	2.98	14%
9	197	47	23.3	272 787	842 093	4 131 061	4.91	14%
10	138	36	16.9	882 506	2 724 144	13 228 089	4.86	15%
11	157	28	9.6	973 496	3 066 512	16 957 039	5.53	13%
12	197	53	33.2	172 697	530 978	2 804 512	5.28	11%
13	167	61	11.8	238 091	725 868	3 887 005	5.35	14%
14	141	12	15.3	236 745	741 600	3 794 707	5.12	14%
15	143	31	19.9	109 949	340 540	2 326 942	6.83	14%
16	146	10	20.2	435 279	1 345 314	8 246 772	6.13	14%
17	152	13	15.5	504 301	1 593 481	8 038 918	5.04	14%
18	147	15	16.2	250 068	788 815	3 936 617	4.99	14%
19	133	-	7.9	821 428	2 627 560	13 079 951	4.98	14%
20	166	23	5.9	752 744	2 385 128	16 622 753	6.97	14%
21	148	40	19.4	137 830	421 374	1 793 144	4.26	12%
Total	130		16.9	8 428 944	26 498 700	142 700 602	5.39	14%

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Below is the summary of the mining resource for the UG2 reef broken up per mining block as per block model.

Table 15-2: Block Model as per Mining Blocks – UG2 Reef

Cut-off Grade	2.50		Resource @ 2.5 g/t Cut-off
Block Number	Average Seam Height	Average Dip	Volume	Tonnes	Content	Grade	Geo- loss %
0	112	20.0	145 021	545 985	2 929 824	5.37	16%
1	143	5.7	1 403 831	5 353 728	21 550 016	4.03	11%
2	120	12.2	458 891	1 757 609	6 859 742	3.90	14%
3	106	9.8	84 341	322 407	1 191 237	3.69	14%
4	111	16.0	277 686	1 046 711	3 424 994	3.27	15%
5	105	7.9	37 355	140 128	365 797	2.61	16%
6	134	24.5	223 626	842 580	2 641 019	3.13	16%
7	134	22.0	48 888	181 563	561 691	3.09	16%
8	131	18.2	302 386	1 046 063	3 163 243	3.02	16%
9	122	11.0	515 655	1 969 072	7 014 911	3.56	11%
10	128	17.2	288 422	1 093 633	3 573 303	3.27	14%

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Cut-off Grade	2.50		Resource @ 2.5 g/t Cut-off
Block Number	Average Seam Height	Average Dip	Volume	Tonnes	Content	Grade	Geo- loss %
11	156	8.0	543 416	2 086 245	8 262 884	3.96	14%
12	145	11.4	300 130	1 105 174	3 947 317	3.57	14%
13	144	21.8	248 878	932 958	3 725 020	3.99	16%
14	132	23.9	259 265	940 290	3 231 807	3.44	16%
15	136	21.9	655 295	2 467 805	10 019 483	4.06	16%
16	131	20.7	301 888	1 148 300	4 950 079	4.31	15%
17	125	26.5	827 049	3 042 507	11 636 235	3.82	14%
18	160	12.9	164 684	615 047	2 423 328	3.94	14%
Total	130	16.9	7 086 706	26 637 804	101 471 929	3.81	14%

15.1.1.1 Grade Cut-off

A revised planning pay limit was calculated using estimates and benchmarked working costs and revenue projections based on the three year trailing price. This resulted in a planning pay limit of 2.5g/t for both the MR and UG2 reef. The planning pay limit calculation was not based on a full financial model but a preliminary estimate based on three year trading prices above and known mining and processing costs.

Table 15-3: Planning Pay Limit Basis

WBJV - Project 1 Cut off used to delineate mining blocks
	Merensky	UG2
Cost per tonne milled	ZAR     936.00	ZAR     844.00
Revenue R/g 4E	ZAR     502.04	ZAR     496.24
Metal Discount	Merensky	UG2	Resource Height cm @ 2.5g/t
			Merensky	UG2
Concentrator Recovery	88%	83%	145	134
Smelter Playability	86%	86%
Penalty’s (UG2 Chrome)	100%	99%
Mine Call Factor	98%	98%
Total Discount	74%	69%	Calculated cm.gt Cut Off
	Calculated g/t planning cut off		Merensky	UG2
Required (g/t)	2.51	2.46	365	329

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Applying the above cut off grades, mining blocks (on both reefs) for inclusion in the mine plan were identified. This resulted in the exclusion of 2.588 million tonnes at a grade 1.50g/t for MR and 8.852 million tonnes at a grade of 1.65g/t for UG2.

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15.1.1.2 Geological Losses

Based on the micro geological environment of each mining block, a geological loss factor was calculated. This factor was then applied to each individual mining block to calculate the known geological losses. This resulted in the exclusion 2.473 million tonnes at a grade of 5.45g/t for MR and 3.655 million tonnes at a grade of 3.80g/t for UG2.

In addition to the geological losses as calculated above, a further geological structure loss is applied. This is due to the physical macro geological features such as major faults and intrusions that are defined in the structural model. The result is a loss of 0.814 million tonnes at a grade of 5.25g/t for the MR and 1.036 million tonnes at a grade of 3.79g/t for the UG2 reef.

15.1.1.3 Merensky UG2 Reef Vertical Spacing

The average vertical spacing between the MR and UG2 reef is 75m for Maseve. There are areas where the vertical spacing between the two reefs is less than 20m. As per the rock engineering design parameters the UG2 reef is excluded in the areas where the vertical spacing is less than 20m. The UG2 reef excluded from the resource as a result of this is 5.357 million tonnes at a grade of 3.87g/t.

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15.1.1.4 Design Pillar and Other Losses

The production access methodology requires protection pillars for the on reef access production drives. The required systematic pillar support strategy results in the design pillar lockup of 3.698 million tonnes at a grade of 5.24g/t for MR and 3.191 million tonnes at a grade of 3.71g/t for UG2.

Additional design pillar losses are incurred for leaving protection pillars around priority excavation, or where as a result of close proximity, reef sterilisation occurred in order to protect priority excavations. Typically these pillars include a 5m protection pillar on either side of each known geological structure as well as the legally required 9m boundary pillar from the lease boundary. This has a resultant exclusion of 1.110 million tonnes at a grade of 5.30g/t for the MR with an additional 1.286 million tonnes at a grade of 3.84g/t for the UG2 reef.

15.1.1.5 Best Cut

In order to minimise the dilution, a minimum drive height of 1.8m (excluding planned dilution of 7.8%) was used for the ASD’s and ledging drives in the hybrid mining sections and 1.9m for the bord and strike drives in the bord and pillar sections. although possible, an increased height would not represent the most economically beneficial mining cut.

A mining cut of 1.8m is the proven minimum standard for the selected low profile trackless mechanised equipment. This consideration takes into account the minimum legislative required clearances for underground machinery.

15.1.1.6 Dilution and Planned Overbreak

Given the selected mining methods (bord and pillar, conventional and hybrid) for both reefs, it stands to reason that dilution varies considerably across the individual mining methods. The average dilution for each mining method on each of the two reefs respectively is tabled below.

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Sources of waste dilution will mainly come from two areas. Firstly, from the over break encountered on the reef horizon in the stoping section, and secondly, from the waste portion in the on-reef development and bord and pillar drives. A planned overbreak of 10% for stoping and 7.8% for on-reef development and bord and Pillar sections where applied over the LoM for both the MR and UG2 sections.

Table 15-4: MR Dilution % per Mining Method

	Resource tonnes (t)	Resource Content (g)	Resource Grade (g/t)	Dilution tonnes (t)	Dilution Content (g)	Dilution Grade (g/t)	Dilution %
Bord and Pillar Section	3 763 865	22 540 772	5.99	1 236 253	328 484	0.27	33%
Conventional Section	5 337 782	28 506 632	5.34	740 539	206 221	0.28	14%
Hybrid Section	5 409 064	29 266 608	5.41	1 296 328	147 660	0.11	24%

Table 15-5: UG2 Dilution % per Mining Method

	Resource tonnes (t)	Resource Content (g)	Resource Grade (g/t)	Dilution tonnes (t)	Dilution Content (g)	Dilution Grade (g/t)	Dilution %
Bord and Pillar Section	-	-	-	-	-	-	0%
Conventional Section	5 645 278	20 972 067	3.71	763 076	217 385	0.28	14%
Hybrid Section	6 915 972	26 249 327	3.80	1 589 519	210 111	0.13	23%

The average sundry tonnage diluting effect on resource tonnes for MR is 22.56% and for UG2 18.73%, with an average grade dilution of 17% for MR and 16.3% for the UG2.

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15.2 Basic Grade Equation

15.2.1 Merensky Reef

Table 15-6: Basic Grade Equation – MR

	tonnes (t)	4E Content (g)	Grade (g/t)
Resource @ 0g/t Cut-off	29 087 005	146 594 503	5.04
Resource @ 2.5g/t Cut-off	26 498 700	142 700 602	5.39
Geo-Structure losses	814 104	4 271 379	5.25
Geo-loss	3 634 420	19 547 177	5.38
RESCAT Losses - Inferred	965 822	2 804 707	2.90
Total Potential Resource for Design	21 084 354	116 077 339	5.51
Design Pillars - Bord and Pillar	1 764 145	10 503 179	5.95
Design Pillars - ORD	3 698 599	19 371 783	5.24
Design Loses - boundary Pillars & Structural Pillars & Losses	1 110 898	5 888 365	5.30
RoM Balance	14 510 711	80 314 012	5.53
Difference	-0.07%	-0.06%
RoM Schedule Resource tonnes	14 520 494	80 361 953	5.53
Footwall mineralization < 3.0 < Jan 2018	258 853	595 624	2.30
Dilution tonnes	3 273 120	682 366	0.21
RoM tonnes	17 524 979	80 400 753	4.59

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15.2.2 UG2 Reef

Table 15-7: Basic Grade Equation – UG2

	tonnes (t)	4E Content (g)	Grade (g/t)
Resource @ 0g/t Cut-off	35 489 936	116 083 451	3.27
Resource @ 2.5 g/t Cut-off	26 637 804	101 471 929	3.81
Geo-Structure losses	1 036 492	3 930 765	3.79
Geo-loss	3 655 811	13 900 793	3.80
RESCAT Losses - Inferred	-	-	-
Total Potential Resource for Design	21 945 501	83 640 372	3.81
UG2 vs MR - Z Diff < 20m	5 357 579	20 727 235	3.87
Design Pillars - ORD	1 286 204	4 937 569	3.84
Design Loses - Boundary Pillars & Structural Pillars & Losses	3 191 785	11 848 480	3.71
RoM Balance	12 109 932	46 127 088	3.81
Difference	-3.73%	-2.37%
RoM Schedule Resource tonnes	12 561 250	47 221 395	3.76
Footwall mineralization < 3.0 < Jan 2018	-	-	-
Dilution tonnes	2 352 596	427 496	0.18
RoM tonnes	14 913 846	47 648 890	3.19

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Mineral reserves are reported as inclusive of uneconomic material and dilution material delivered for treatment or dispatched from the mine without treatment. Modifying factors are applied as indicated in the basic grade equation.

15.3 Mineral Reserve Statement

Table 15-8: Mineral Reserve Statement

Estimated Total Reserve Project 1 100% Project Basis - July, 2015
	Reserve tonnes - Mt	Pt  g/t	Pd g/t	Rh g/t	Au g/t	Reserve 4E Grade - g/t	Reserve 4E Content - t	Reserve 4E Content - Moz
MR Proven and Probable	17.525 2.94		1.24	0.18	0.23	4.59	80.401	2.585
UG2 Proven and Probable	14.914 2.01		0.83	0.32	0.03	3.19	47.649	1.532
Total	32.439 2.51		1.05	0.25	0.14	3.95	128.05	4.117

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Table 15-9: MR and UG2 Prill Split

Merensky Reserve
	Reserve tonnes - Mt	Pt g/t	Pd g/t	Rh g/t	Au g/t	Reserve 4E Grade - g/t	Reserve 4E Content - t	Reserve 4E Content - Moz
Proven	7.082	2.89	1.22	0.18	0.22	4.51	31.905	1.025
Probable	10.443	2.98	1.26	0.18	0.23	4.65	48.496	1.560
Total	17.525	2.94	1.24	0.18	0.23	4.59	80.401	2.585

UG2 Reserve
	Reserve tonnes - Mt	Pt g/t	Pd g/t	Rh g/t	Au g/t	Reserve 4E Grade - g/t	Reserve 4E Content - t	Reserve 4E Content - Moz
Proven	5.452	1.95	0.80	0.31	0.03	3.09	16.821	0.540
Probable	9.462	2.05	0.85	0.33	0.03	3.26	30.828	0.992
Total	14.914	2.01	0.83	0.32	0.03	3.19	47.649	1.532

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Prill splits are calculated using the individual metal grades reported as a percentage of the total 4E grade. No one Prill split constitutes a mineral reserve and only the 4E grade is used in the Reserve calculation.

15.4 Material Changes to Mineral Reserves

The mining, metallurgical, infrastructure, permitting and other relevant factors are at a Feasibility level of confidence and thus the QP does not foresee any factors not disclosed that may materially affect the reserve estimate. The construction of the mine infrastructure is

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generally congruent with the feasibility plan. Metal prices on a longer term or spot basis provide margin for economic consideration.

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16.0	MINING METHODS
16.1	Introduction

A FS on this project was completed and published in July 2008. Subsequent to that study an Updated FS was completed and published in October, 2009. Work has been completed on phasing the implementation of the project as per the undated FS, 2009. Development work commenced on the North Decline in November 2011. Up to the effective date of this report underground development completed includes:

Decline waste development	- 9 562m
Underground Infrastructure development	- 2 868m
Reef drives development metres	- 1 050m
Reef Raise development metres	- 647m

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Development work commenced on the South Decline in April 2013. Up to the effective date of this report underground development completed includes:

Decline waste development	- 9 562m
Underground Access Waste Development	- 2 868m

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In parallel with the development work a reassessment of the Mineral Resource has been completed. Based on this updated resource as well as an improved confidence geological model and experience gained during development, a revised LoM Plan has been completed. The Mineral Reserve is based on the revised LoM Plan for the WBJV Project 1 and 1A. Significant changes to the LoM Plan from the Updated FS 2009 include:

>	Introduction of twin declines to outlying blocks,
>	Access method change from predetermined levels to optimal access points using ramps,
>	Inclusion of Underground Workshops and Silos,
>	Three mining methods have been planned for, based on the dip and mining height of the resource estimate,
>	Pay limit change from 3.5 g/t to 2.5g/t for MR and from 3.5g/t to 2.5g/t for UG2,
>	Inclusion of Project 1A.

The overall on-reef mining methods have changed from the previous study. The mine will be operated as a trackless and conventional operation. Personal, materials, equipment and rock will be moved between surface and underground by diesel vehicles operating in North and the South Decline during the development phase.

A Main conveyor installed in the north conveyor decline will transport ore and waste from underground to the waste and ore stockpile / silos on surface during production and personal will be transported by a chairlift. Underground silos are being constructed at the bottom of the first leg of the north decline system. Initial ore and waste will be transported by means of trucks to designated stockpile facilities on surface until the conveyors are installed. The assumption made in this mining section is that all development necessary to access reef is completed in time to meet the mine production schedule.

In the updated Mine Plan, early mine development has been scheduled to achieve full production of 160 000tpm of ore in the shortest time possible. Development is commencing simultaneously on both shafts and will proceed at the planned rates.

Initial mine development will focus on:

>	Providing access for personal, materials, and equipment to high-grade, shallow areas of the orebody,
>	Providing an early means of ventilating the workings,

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>	Providing two routes for emergency egress at all times,
>	Installing mine services (electric power distribution, communications, compressed air, water supply, mine dewatering).

Assumptions made for the mine planning areas are that all requirements as defined by this mining section (men, material, ventilation and services), are delivered to the workings as required and that all rock broken on the reef horizon, is removed immediately.

16.2 Geological and Geotechnical Features Affecting on Reef Mining

The main findings in the geological and rock engineering investigations that influenced on-reef mine design are discussed below:

>	The MR has an average dip of 15.31 degrees and an average stoping width of 142cm at a cut-off grade of 2.5g/t,
>	The UG2 reef has an average dip of 16 degrees and an average stoping width of the mine is 129cm at a cut-off grade of 2.5g/t,
>	Certain mining blocks have the potential for increased mechanization while other blocks has a greater potential for mining methods more suitable with narrow steep dipping ore bodies if the currently predicted dip, structure and width are confirmed by additional drilling, either from surface or from underground with stope definition drilling,
>	Significant faults and dykes intersect the orebody and subdivide the deposit into a number of discrete mining blocks, each of which requires access development on different mining levels. The resultant blocks of ground left un-mined add to the regional stability of the mine.

After application of appropriate pay limits, the MR contains 40% more recoverable metal than the UG2 and is therefore the primary target. The parting between the MR and the UG2 varies from contact in the South to 70m in the North. Mining of both reefs generally only occurs when the parting is greater than 20m as prescribed by the rock engineer.

16.3 Mining Method Selection

The geological and structural models in conjunction with geotechnical considerations formed the basis of the mining methods that was selected to provide the best practical outcome under the given conditions. The selected methods had to be versatile and easily interchangeable with the lowest impact on production during translation between the mining methods. The mining methods would also have to be able to integrate with the trackless environment that would supply access and service all of the mining areas.

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The selected mining methods are:

> Bord and pillar

• Bord and pillar mining was considered in flatter dipping areas with a maximum dip of 15º and where the actual seam height of the reef was conducive to the increased mining heights needed for the bord and pillar equipment.

> Conventional Mining

• The conventional mining method was considered in the steeper dipping areas due to raises being able to be developed at dips of up to 34º. This results in the best and most practical outcome under high dip variability conditions and delivers the most favourable development replacement rates.

> Hybrid Mining

• Hybrid mining was considered in the moderate to steeper dipping areas where the bord and pillar method was not eligible. This method took precedence over the conventional method where applicable, due to the on-reef access which allows a quicker reef access and faster production build-up.

The final selected mining methods are well proven throughout the region with positive results under similar conditions.

16.4 Rock Engineering

16.4.1 Mine Accesses

Access into the mining area is via a two barrel decline system. One system at North mine and a second duplicated at South mine. Primary support consist of 2,4m long resin bolts spaced 2m x 2m on dip and strike respectively. The maximum span allowable span is 8m. Additional support in the form of a thin spray liner, mesh and lace and long anchor bolts are required in areas where geotechnical conditions dictate.

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16.4.2 Ventilation Shafts

Ventilation shafts will be established via raise boring. Surface material (mainly turf) will be removed and replaced by mass concrete and the weathered Saprolites between the mass concrete and the competent ground will be consolidated via curtain grouting.

16.4.3 Merensky Reef Horizon

Narrow Reef:

The rock mass ratings indicate that the MR hanging wall rock mass is far superior to the UG2 hanging wall rock mass.

The potential for parting along the bastard reef contact and the variations in the distance between the MR and the bastard reef is incorporated into support design by the standardisation of grout packs with timber elongates. Maximum panel spans of 25m inclusive of ASG is planned.

Table 16-1: Breakdown

Reef type	Elongate support	Elongate spacing	Grout packs	Grout pack spacing
MR dip <45	160mm diameter Mine poles	1.5m x 2m	Required	5m x 3.5m skin to skin
MR dip >45	160mm diameter Mine poles	1.5m x 2m	To be assessed	To be assessed
UG2 Reef horizon	160mm diameter Mine poles	1.5m x 2m	Required	5m x 3.5m skin to skin

Wide reefs:

Bord and pillar mining method:

>	Pillar dimensions as per depth,
>	Bords will be supported with resin bolts.

16.4.3.1 UG2 Reef Horizons

The extraction of UG2 reef will be supported with elongates spaced 1.5m x 2.0m and in addition, grout packs will be installed to prevent cantilevering of Pyroxinite hanging wall. In-stope bolts measuring 1,2m long, will be required to keep Pyroxinite layers intact. Maximum panel lengths of 25m will be planned and this can likely increase if strong stiff support can be installed.

> In-stope Pillars:

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• In-stope pillars in the conventional and hybrid mining sections on both MR and UG2 reef horizons will be 6m wide along strike and 4m wide along dip.

> Seismicity:

• Due to the depth of the project no seismic activity is envisaged.

> Surface subsidence:

• No surface subsidence expected and all areas of importance specified by law will be monitored.

The following pillar loss percentages have been calculated:

>	MR – Conventional Mining (narrow reef) = 11%,
>	MR – Bord and Pillar Mining ( Wide reef) = 22-30%,
>	UG2 Mining = 13%.

The losses likely to be incurred by systematically leaving regional pillars will not exceed 1%. These pillars will only be required in areas where the depth to span ratio (between unmined ground and geological structures) exceeds 0.5.

1.	No bracketing of fault losses will be necessary unless local rock mass conditions dictate otherwise.
2.	Bracket pillars on dykes will be dictated by the condition of the rock mass influenced by the chill margins rather than improving average pillar stresses on regional pillars. Chill margins are not expected to exceed 3 - 5m.
3.	No mining in permitted closer than 20m from any major decline or excavation.

16.5 Mine Design Features

16.5.1 Mine Development Philosophy

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The updated LoM design as illustrated in Figure 16-3 and Figure 16-4, consist of the twin decline system as initially adopted in the execution plan to access the underground workings. Men, material and rock will be transported by trackless mobile machinery during the development phase of the mine up until the point where the conveyor belt and chairlift systems are installed and operational.

This design concentrates on reducing the waste footwall development by moving production access development to an on reef position. This is more suitable for the opted trackless layouts and allows a faster RoM build-up. It does result in a moderately higher dilution percentage.

16.5.2 Waste Development: Declines and Ramp Access Development

The orebody is being accessed by two sets of declines (twin decline system on North and a twin decline system for South), which connect surface to underground production levels. Both North and South declines have the same configuration.

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The decline system consist of two parallel declines – a material decline 5.0 m(w) x 4.5 m(h) for rubber-tired vehicles (trackless mobile machinery) to initially transport rock, men and material and a belt decline 6.5 m(w) x 3.8 m(h) for moving men and rock once the chairlift and conveyor is installed. Twin decline system is developed at a 9º dip below the horizontal. Connecting crosscuts between declines will be developed at a 100m spacing. Muck bays, 12m deep will be developed on the material decline and the belt decline’s side at an angle that will facilitate the mucking and loading process.

Haulage decline cross-sections and turning radii are sized to accommodate the largest trackless mining equipment (30t truck) that will be deployed underground. Declines were developed from the 25m high boxcut wall. The typical declines excavation sections are illustrated in Figure 16-6.

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Both the material and belt declines will be mined by mechanized methods, comprising of trackless drill rigs, roof-bolters, Load Haul Dumpers (“LHD’s”) and 30t underground trucks.

Broken rock will be loaded by LHD’s from the face and dumped in muck bays spaced at 100m intervals. An LHD will load waste rock into the 30t trucks to be hauled to the surface waste dump. Muck bays will be used as traffic passing bays, as temporary dam sumps, electrical substations, material storage, vehicle parking, and the construction of refuge bays.

The material and belt declines will have a minimum 10m intervening solid pillar to facilitate their use as independent means of emergency egress (as per statutory definition). The pillars also serves to protect the excavations for the LoM. Material and belt declines will be developed at the same rate of advance.

From the decline systems, a ramp access drive is developed in waste to connect the reef horizon to the main declines. The ramps serves as main accesses to the various mining blocks. Where the ramps access the reef horizon, a twin on-reef access drive system is developed on strike to the extent of the specific mining block. The Access drives are developed with a 10m skin to skin pillar spacing between them. The twin access drive system is required to establish an air intake and return system. This facilitates the early development and stoping of ore.

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Permanent pipelines, ventilation ducts, and power cables will be installed concurrent with the advance of the twin decline system and ramp access drives.

The twin decline system and all on-reef development will have separate dedicated crews and suits of equipment in order to ensure that the development rates are maintained throughout the LoM and that no particular system takes precedence over the other.

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All on-reef development will be mined by the same type of trackless mining equipment and methodology used in the twin decline system. The only difference being the 30t low profile trucks instead of the 30t trucks. Footwall development dimensions are be sized to accommodate the 30t diesel haul trucks, selected for this project. Construction activities will follow behind as the development headings are advanced.

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16.5.2.1 On-reef Development

Bord and Pillar Section

In general the on-reef access drives are broken away directly from the declines. Block 17 is connected to the declines by means of a ramp access. The twin on-reef access drives unlocks the reserves on strike. The expected inclination due to roll in the reef and localised changes in strike will be limited to a maximum of 9 degrees. Ledging drives are developed on either side of the twin on-reef access drives. A connecting crosscut will connect the outer ledging drives to the access drives at each mining section strike drive (See Figure 16-9 below for clarification).

Conventional Mining Section

In the conventional mining sections, a centre raise will be developed at a dip or apparent dip of less than 34 degrees from horizontal. The centre raise line originates in the on-reef access drive to the next on-reef access drive above the away position. At the bottom of the raise, a muck bay is developed to accommodate the broken rock created in that raise. The muck bay also and acts as a storage area form where the trackless machinery will be able to load and haul rock at will.

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Hybrid Mining Section

In the hybrid mining sections, access drives are developed on strike form a ramp access drive in order to expose the ore reserves. Twin decline / incline drive system are developed from these access drives at an inclination of 9 degrees to maximise the unlocking rate of ore reserves. The twin incline / decline drives will be accompanied by a ledging drive on either side of the system. A minimum protection pillar of 10m for each of the incline / decline drives is left to protect the twin incline / decline system from the ledging and stoping operations for the duration of mining operations in that block or mining area. The lateral connections between the twin drives and ledging drives will be developed in such a way that both sides are connected in line as to accommodate practical crossover of the trackless mobile machinery and conveyor belt systems where used.

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16.5.3 Stoping

The stoping process is divided up in two main processes. Firstly the ledging and equipping phase and then the actual stoping or reef extraction phase.

During the ledging and equipping phase, a newly established mining area will be prepared by slowly profiling the stope face. Conventional and hybrid stopes will be preceded by the ASG and ASD respectively in order to create a free breaking point. Mining stopes are advanced in the ledging mode of operation for a distance of 10m at which point the stope be equipped and prepared by installing the permanent services required for stope panels to operate at full capacity.

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16.5.3.1 Bord and Pillar Section

The bord and pillar mining method represents 26% and 0% of the total mined tonnes for the MR and UG2 reefs respectively. The bord and pillar mining method is used on the MR in areas where the dip is less than 15 degrees and the seam has an average thickness of 156cm. The method however was excluded on the UG2 reef mainly due to the reduced seam height of 133cm and lower grade. Mining dilution will render the UG2 uneconomical with the bord and pillar method applied on MR.

Mining Layout

Mining sections will consist of seven boards, four bords up-dip of the conveyor road, and two bords down-dip. This will minimise up dip tramming of fully loaded LHD units resulting is a more energy efficient loading regime. The layout is benchmarked in the area a based on best practice. The total section width is 84m, skin to skin. Figure 16-12 illustrates the designed layout.

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Although the bord and pillar mining method is a self-supporting method, in-bord local hanging wall support will be in the form of rock bolts installed by mechanised bolter. All blast holes will be drilled by low profile, single boom drill rigs. The effective drill hole length is 3.2m, with a programmed advance of 3.0m. Face cleaning and removal of broken rock will be done by low profile LHD’s loading onto strike conveyors.

All strike roadways are developed from the declines at a direction that will yield an overall inclination of 1:200 above strike; this is to produce effective drilling and ground water control, as well to cater for undulations in the reef and localised changes in strike. Service water will be pumped (using small movable electric face pumps) to the section dam in close proximity of the decline.

The depth of the areas earmarked for bord and pillar mining varies from 414m to 628m below surface. Graph 16-3 illustrates the Depth to Seam (“DTS”) distribution for the bord and pillar stopes. 15% of the bord and pillar stopes falls in the below 460m DTS category which requires a 8m bord width and a 5m (w) x 5m (l) pillar and another 12% of the stopes falls in the above 550m DTS category which requires a 6m bord width and a 5m (w) x 6m (l) pillar as per the geotechnical design pillar table. The majority or 73% of bord and pillar stopes lies between 460m and 550m and requires a 7m wide bord with a 5m (w) x 6m (l) pillar.

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A working section is equipped and resourced to advance at 30.5m per month, yielding 1 850m² or 11 180t per section. The maximum bord and pillar sections in operation at any time over the LoM is eleven.

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Table 16-2: Bord and Pillar Design at Depth to Surface

Rectangular pillar for bord and pillar layout at Maseve between 360m and 600m deep.
Depth	Pillar Width	Pillar Length	Weff	Board width	Vent Cut	Mining Height	W:H Ratio	Factor of Safety	Pencent. Extraction
360	5	5	5.00	8	4	3	1.67	1.76	78.6%
370	5	5	5.00	8	4	3	1.67	1.71	78.6%
380	5	5	5.00	8	4	3	1.67	1.67	78.6%
390	5	5	5.00	8	4	3	1.67	1.63	78.6%
400	5	5	5.00	8	4	3	1.67	1.59	78.6%
410	5	5	5.00	8	3.5	3	1.67	1.64	77.4%
420	5	5	5.00	8	3.5	3	1.67	1.60	77.4%
430	5	5	5.00	8	3	3	1.67	1.66	76.0%
440	5	5	5.00	8	3	3	1.67	1.62	76.0%
450	5	5	5.00	8	3	3	1.67	1.59	76.0%
460	5	5	5.00	7	3	3	1.67	1.68	74.0%
470	5	5	5.00	7	3	3	1.67	1.64	74.0%
480	5	5	5.00	7	3	3	1.67	1.61	74.0%
490	5	5	5.00	7	3	3	1.67	1.58	74.0%
500	5	6	5.45	7	3	3	1.82	1.72	72.2%
510	5	6	5.45	7	3	3	1.82	1.69	72.2%
520	5	6	5.45	7	3	3	1.82	1.66	72.2%
530	5	6	5.45	7	3	3	1.82	1.63	72.2%
540	5	6	5.45	7	3	3	1.82	1.59	72.2%
550	5	6	5.45	6	3	3	1.82	1.71	69.7%
560	5	6	5.45	6	3	3	1.82	1.68	69.7%
570	5	6	5.45	6	3	3	1.82	1.65	69.7%
580	5	6	5.45	6	3	3	1.82	1.62	69.7%
590	5	6	5.45	6	3	3	1.82	1.59	69.7%
600	5	6	5.45	6	3	3	1.82	1.57	69.7%

16.5.3.2 Conventional Mining Section

The conventional mining method represents 35% and 43% of the total mined tonnes for the MR and UG2 reefs respectively. The method is mainly applied in the areas of steeper dip throughout both the MR and UG2 reefs. The mining sections is serviced with trackless equipment. Reef extraction will be done by means of the conventional drilling and blasting methods with scraper winch cleaning tipping ore into a loading bay, form where a LHD will load and haul the generated material.

Mine Layout

Top and bottom access to the conventional stoping section is effected by trackless on-reef access drives. A conventional raise (centre gully) is developed on dip or at an apparent dip angle not exceeding 34 degrees. Raise (centre gully) spacing is 150m on strike.

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Raise back lengths were kept to practical lengths to best accommodate efficient scraper and winch cleaning of broken rock into the loading bays. Where back lengths exceeded 200m (mainly on the UG2 reef) footwall crosscuts were extended to install a cleaning orepass.

All the conventional mining panels are 25m long for both MR and UG2 reef. The mining panels advance in a strike direction away from the centre raise line. See Figure 16-13.

Advance strike gullies are mined on strike ahead of the conventional panels in order to create a free breaking point into which the panel can be blasted. Support pillars are left on the down dip side of the advance strike gullies. The pillars are not continuous on strike allowing for the flow of ventilation between panels. These ventilation holings are 2m wide and are blasted in an up dip direction towards the top ASG.

Additional support in the form roof bolts and 160mm diameter elongates (pencil sticks) for the MR and 180mm diameter elongates for the UG2 reef are required and is included in the mine design. The in stope roof bolts are drilled using customized remote controlled support drill rigs to ensure vertical drilling of roof bolt holes from a position of safety.

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Table 16-3: MR Support Recommendations in Conventional and Hybrid Stoping Sections

Reef Type	Elongate Support	Elongate Spacing	Grout Packs	Grout Pack Spacing
MR dip <45	160mm diameter pencil stick	1.5m x 2m	Required	5m x 3.5m skin to skin
MR dip >45	160mm diameter pencil stick	1.5m x 2m	To be assessed	To be assessed

Table 16-4: UG2 Support Recommendations in Conventional and Hybrid Stoping Sections

Reef type	Elongate support	Elongate spacing	Grout packs	Grout pack spacing
UG2 Reef horizon	180mm diameter Mine poles	1.5m x 2m	Required	5m x 3.5m skin to skin

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The reef extraction percentage for the support layout as illustrated above in Figure 16-14, is 88% (including the ASG tonnes) and 74% for the panel only (excluding the ASG tonnes).

All panels and gullies in the conventional stoping sections will be drilled by means of Hydro Powered Equipment (“HPE”) rather than with conventional pneumatic rock drills. Whilst this technology is relatively new, a study was done by Snowden for the WBJV1 project as per report: 130614_D_JR###-##-####_PTM_Maseve drilling report. Please see Section 2.7.10. Hydraulic rockdrills can also be considered for the mine in preference to reverting to compressed air. The blast holes are charged with pumpable emulsion type explosives and are detonated using shock tubes.

Broken rock is cleaned by scraper winch to the Advance Strike Gullies (“ASG”) and then scraped on strike, to the centre raise. The reef is scraped down the centre raise and is either tipped directly into an ore pass to the level below, or in the loading bay from where a LHD will load and haul the broken rock. See layout of the Loading Bay in Figure 16-15.

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Each conventional stoping crew will be equipped and resourced to advance at 17.5m and 19m per month for MR and UG2 reef respectively. An average raise line would contain 10 - 14 panels. For every 12 panels that are equipped for mining, 50% will be considered as available at any one time. At full production a maximum of 40 conventional stoping crews will be deployed. Each with the capacity to produce 350m².

16.5.3.3 Hybrid Mining Section

The hybrid mining method represents 37% and 57% of the total mined tonnes for the MR and UG2 reefs respectively. The hybrid mining method was used in areas where the dip was too steep for the bord and pillar mining method and the seam height too low for the 1.9m minimum required mining height needed for the selected low profile trackless mobile machinery of the bord and pillar method. The trackless mining approach was still applied in these sections, but Low Profile (“LP”) machinery with its capability to operate in low seam height environments is to be used in order to reduce the overall maximum mining height to 1.8m in the ledging drives and ASD’s. The reduced heights contributes to limiting the overall dilution incurred in the hybrid mining method.

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Mine Layout

Access to the hybrid stoping or reef extraction areas are given by the twin incline / decline drive system. ASD’s are developed on strike, outwards from the ledging drives. The ASD is positioned on the down dip side of the hybrid stoping panel, where it will act as a free breaking point for the panel for the life of that panel.

The ASD has the same purpose as the ASG in the conventional mining section, with the difference that it will be developed by LP trackless mobile machinery.

As in the conventional stoping sections strike pillars are not continuous allowing for the flow of ventilation between panels. The ventilation holings are also 2m wide and are blasted from the panel towards the top ASD until holing is affected. The instope mining equipment to be used in the hybrid stoping section will be the same as that of the conventional sections. Please see Section 2.7.10. for HPE equipment to be used.

All the instope equipment to be used in the hybrid stoping section will be the same that of the conventional sections. Please see Section 2.7.10. for HPE equipment to be used.

Support systems in the Hybrid stopes are exactly the same as that of the Conventional stopes in that 160mm and 180mm elongates (Pencil sticks), on the MR and UG2 reefs respectively, will be used in conjunction with grout packs and instope roof bolting. The blast holes will also be charged with pumpable emulsion type explosives and detonated by means of shock tubes.

The broken rock generated on the panel will be scraped from the panel face into the ASD by using a winch and scraper cleaning method. The broken rock in the ASD will be loaded by the LP LHD, and hauled to the twin decline / incline, where it will be tipped into 30t trucks at dedicated loading points. See Figure 16-16 for clarification.

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Each stoping crew will be planned, equipped and resourced to advance at 19m per month. At full production a maximum of 41 hybrid mining stoping crews will be operational each with the capacity to produce 350m².

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16.6 Mine Ventilation and Cooling

16.6.1 Summary

Subsequent to the initial detailed designs-completed in March 2012, changes were made to mine design to improve the viability of the Maseve Project. Some of the proposed changes have impacted on the original ventilation designs and specifications. The ventilation section assesses the impact of the significant changes on ventilation designs and associated costs.

The ventilation strategy considers safety and health in accordance with the Mine Health and Safety Act (MHSA, Act 29 of 1996) and complies with Maseve Mine health and safety requirements. The primary ventilation quantity for Maseve is 1 100 m³/s; dictated by the need to dilute diesel emissions, remove heat and dilute blasting fumes (during re-entry period). The primary ventilation quantity satisfies the mine heat load without the need for refrigeration. It must be noted however that, wet-bulb temperatures will exceed 27.5°C and approach 29.0°C and Heat Tolerance Screening of the underground work force will be required. Interactive computer simulation of heat and air flow was used to determine ventilation requirements over the LoM for maximum depth and strike.

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The mining plan is based on steady state production of 160 000 reef tonnes per month. During the first phase of the project, the primary ore body will be MR and accordingly discussion in this report focusses on access of the MR. Later UG2 will provide replacement tons and will be ventilated utilising the ‘existing’ MR infrastructure by extending established intake and returns to UG2 as required (e.g. step RBH’s, drop raises, horizontal intake and return airways). The strategy will mine UG2 within a specific mining block only after the ‘overlying’ MR block has mined out, in other words the two reefs will not mine simultaneously from the same area.

In general, each mining block will be ventilated as a separate district while at the same time utilising as much common infrastructure as practically achievable. Fresh air will be introduced to mining blocks through a combination of the main North and South access decline systems and strategically located fresh air RBH. Air returns through return RBH’s equipped with fans. Generally returns will serve more than one block but in some cases a blocks will require a dedicated return to surface. It should be noted that RBH’s were phased-in to meet the production requirements as provided.

Absorbed Power

The peak absorbed power for ventilation will be approximately 7.8MW; divided between 5.3MW for the main surface fans and 2.5MW for auxiliary underground fans.

16.7 Basic Mining Equation

An appropriate steady state production rate of 160 000 reef tonnes per month were used as basis and planned for in the LoM schedule. This 160 000 reef tonnes includes stoping, ledging and On-reef Development tonnes, but excludes all waste development tonnes.

Based on the reserve break-up (in terms of mining method), Graph 16-4 and Graph 16-5 for MR and UG2 reefs respectively, the following Basic Mining Equations, (Table 16-5), where drawn up.

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Table 16-5: BME – MR and UG2

Mining Method	Reserve tonnes (t)	Reserve Content (g)	Reserve Grade (g/t)	%	tonnes per Month needed
Total Merensky Reef	17 524 979	80 400 753	4.59		160 000
Bord and Pillar	5 000 118	22 869 256	4.57	29%	45 650
Conventional	6 078 321	28 712 853	4.72	35%	55 494
Hybrid	6 446 540	28 818 644	4.47	37%	58 856
Total UG2 Reef	14 913 846	47 648 890	3.19		160 000
Bord and Pillar	-	-	-	-	-
Conventional	6 408 355	21 189 452	3.31	43%	68 751
Hybrid	8 505 491	26 459 439	3.11	57%	91 249

Derived from Table 16-5, each individual mining method’s BME was calculated as below.

16.7.1 Bord and Pillar Sections

In the bord and pillar sections, the main contributors towards reef tonnes are the strike drives, bords as well as the ventilation holings. The on-reef development in bord and pillar sections are limited to the access drives only. The Bord and Pillar BME is tabled below in Table 16-6.

Table 16-6: Bord and Pillar BME (Basic Mining Equation)

Description	Unit	Rates
Bord width	m	7.00
Rooms per section		6.00
Strike belt width	m	7.00
Section width	m	84
Face length per section	m	49
Pillar Length	m	6.00
Pillar Width	m	5.00
Mining height	m	1.90
Rock bulk density	t/m3	3.18
Vent holing length	m	5.00
Vent holing width	m	3.00
Advance per blast	m	3.00
Tons per blast per room	t	127
Average face advance per month	m	30.50

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Description	Unit	Rates
Total area mined in rooms per month	m²	1 495
No. of ventilation holings per section per month		23.72
Vent holing area	m²	356
Total area per month	m²	1 850
Tons per section per month	t	11 180
Number of operating sections		4.08
Production per month from Bord and Pillar	t	45 650

The bord and pillar mining method will produce 45 650 reef tonnes per month and based on the calculation above, there needs to be an average of just over four operational crews with a maximum deployment of 11 crews throughout the life of the WBJV Project 1 (Maseve) mine.

16.7.2 Conventional Sections

Traditionally conventional narrow reef mining methods are development intensive and dependant thereon to establish infrastructure required to enable effective mining. At the WBJV Project 1 (Maseve) on-reef mechanised access development approach was selected. This allows for quicker reef access and faster production build up with the additional benefit of reduced waste development. Applying this approach the following BME was calculated for the conventional mining sections on the MR and the UG2 reefs respectively.

Table 16-7: Conventional Stoping - MR BME

Description	Unit	Rates
Panel width	m	19.50
Pillar Length	m	6.00
Pillar Width	m	4.00
Mining height - Ave	m	1.46
Rock bulk density - Ave	t/m3	3.12
Vent holing length	m	4.00
Vent holing width	m	2.00
Advance per blast	m	1.20
Tons per blast per panel	t	107
Average face advance per month	m	17.50
Total area mined in panel per month	m²	341
No. of ventilation holings per section per month		1.75
Vent holing area	m²	14
Total area per month	m²	355

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Description	Unit	Rates
Tons per section per month	t	1 618
Number of operating crews		34.29
Production per month from Conventional Stopes - Merensky	t	55 494

Table 16-8: Conventional Stoping – UG2 BME

Description	Unit	Rates
Panel width	m	19.50
Pillar Length	m	6.00
Pillar Width	m	4.00
Mining height - Ave	m	1.40
Rock bulk density - Ave	t/m3	3.74
Vent holing length	m	4.00
Vent holing width	m	2.00
Advance per blast	m	1.20
Tons per blast per panel	t	123
Average face advance per month	m	17.50
Total area mined in panel per month	m²	341
No. of ventilation holings per section per month		1.75
Vent holing area	m²	14
Total area per month	m²	355
Tons per section per month	t	1 860
Number of operating crews		36.96
Production per month from Conventional Stopes - UG2	t	68 751

16.7.3 Hybrid Sections

The Hybrid mining reserves contains the actual hybrid mining stopes and the on-reef development sections that caters not only for the hybrid mining sections but also for the bord and pillar and conventional sections respectively. The fraction of the on-reef development sections of the total hybrid reserves is 35% for the MR and 25% for the UG2 reefs. Thus the total stoping tonnes that needs to be produced only makes up for 65% and 75% of the two respective reef reserves.

Table 16-9: Hybrid Stoping – MR BME

Description	Unit	Rates
Panel width	m	17.50
Pillar Length	m	6.00
Pillar Width	m	4.00

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Description	Unit	Rates
Mining height - Ave	m	1.42
Rock bulk density - Ave	t/m3	3.11
Vent holing length	m	4.00
Vent holing width	m	2.00
Advance per blast	m	1.20
Tons per blast per panel	t	93
Average face advance per month	m	19.00
Total area mined in panel per month	m²	333
No. of ventilation holings per section per month		1.90
Vent holing area	m²	15
Total area per month	m²	348
Tons per section per month	t	1 536
Number of operating crews		38.33
Production per month from Hybrid Stopes - Merensky	t	58 856

Table 16-10: Hybrid Stoping – UG2 BME

Description	Unit	Rates
Panel width	m	17.50
Pillar Length	m	6.00
Pillar Width	m	4.00
Mining height - Ave	m	1.32
Rock bulk density - Ave	t/m3	3.71
Vent holing length	m	4.00
Vent holing width	m	2.00
Advance per blast	m	1.20
Tons per blast per panel	t	103
Average face advance per month	m	19.00
Total area mined in panel per month	m²	333
No. of ventilation holings per section per month		1.90
Vent holing area	m²	15
Total area per month	m²	348
Tons per section per month	t	1 703
Number of operating crews		53.59
Production per month from Hybrid Stopes - UG2	t	91 249

Although the stoping reserves only makes up fraction of the total Hybrid reserves (On-reef development makes up the other fraction of the Hybrid reserve), provision was made during calculations for the probability of the stoping sections required to produce the full production required for that mining method.

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Additional Information and Provisions

To allow for and accommodate panel re-establishment and equipping, a redundancy factor of 15% is applied to the mining labour. Production crew capability is stated at 350m² per crew per month but mine scheduling is done at 300m² per month.

At any one time stoping labour allocated to production will be involved in re-establishment of panels. For this reason, an additional redundancy factor of 15% where applied on the Conventional and Hybrid stoping sections over the LoM schedule. Production rates per crew were ascertained at 350m² per month per crew for stoping. The rate applied in the LoM schedule is 300m² per month per crew. This creates the required additional 15% redundancy for re-establishment at any given stage during the LoM schedule as there will be 15% more panels available than being mined.

Mining panels are only mined when they are ventilated (raises holed) and resources are available Sufficient timing has been allowed in the production schedule for ledging and equipping. The following is applicable:

>	For every panel planned to produce in any month, an additional panel, fully equipped is considered unavailable for geological or other reasons,
>	A stoping crew with compliment of 15 (excluding supervision), is allocated two producing panels which is worked over two shifts,
>	A raise line will generally have 4 or 5 operating mining crews (depending on back length),
>	Conventional crews will blasts 15 times per month at a scheduled face advance per blast of 1.2m. The Hybrid crews will blast 16 times per month and also at a face advance per blast of 1.2m,
>	A Maximum of 77 stoping crews are deployed at full production with an average of 50 crews operating during MR extraction,
>	77 stope crews have the potential to produce 121 500 reef tonnes or 27 000m² per month.

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For every ten stoping crews there will be one vamping and reclamation crew, which is catered for in the labour plan. These crews will consist of 15 employees spread out over 2 shifts and excludes supervision.

16.7.4 Labour

The total labour requirements for the WBJV Project 1 (Maseve) MR is as shown below in Graph 16-6 to Graph 16-9, this shows the stoping and development crew requirements respectively.

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The total staff requirement for the WBJV Project 1 (Maseve) peaks at 3 650 employees of which 3 098 employees will be on mine. A further 552 employees makes up the 16% unavailable category. The staffing composition is shown in Graph 16-9.

The labour productivity is determined by the square metres per total reef employee costed and is broken up into each individual mining method. The average labour efficiency is 57m², 22m² and 19m² excluding the 16% unavailable or relief category and 48m², 19m² and 16m² including the 16% unavailable or relief category for Bord and Pillar, Conventional and Hybrid stoping respectively.

The relief percentage of 16% (8.3% leave relief and 7.7% unavailability’s) is based on historical numbers and benchmarked for the industry in the area. Continuous operations have not been included in the production plan.

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16.7.5 Operating Cycles

Maintaining and sustaining full production is based on the following mining cycle. Each stoping crew under the day shift control of a qualified miner, is allocated two adjacent (or close to each other) stope panels. The two panels are mined on a rotational basis. Each day one panel is supported and prepared for drilling and the other is drilled and blasted.

Cleaning is done mainly on night shift but winch operators are also allocated to day shift. This allows for catch-up cleaning of any remaining reef, and ensures cleaning capability on day shift to make up for lost blasts. The day shift winch operators are also required to prepare the cleaning rigs for night shift.

In the face preparation cycle, permanent support (elongates) are installed as per mine standard. Each crew is allocated a remotely operated stope support drill rig. Roof bolts are installed on a predetermined systematic pattern during the preparation cycle and installed according to the rock engineering recommendation and mine standard. When support is complete, the face can be marked off by the miner with time available to ensure that the marking off is done correctly and according to mine standard.

The drilling crew drills the face on the second panel that has been prepared on the previous shift. This ensures that no time is lost waiting for the completion of the cleaning, marking of the holes and supporting of the face. This mitigates the risk of drilling being commenced before a face has been adequately prepared. The effective drilling time per shift is significantly improved. The system aims to reduce lost blast by ensuring a more consistent face availability. In addition, more attention given to marking the shotholes contributes to a better face advance per blast in a generally safer environment. Better stoping width control is also possible thus adhering to the best stoping cut as recommended by the Geologists.

The Opex budget makes provision for underground exploration drilling and well as development cover drilling for geological, structural and evaluation purposes. The cycle mining and panel redundancy affords the production crew the opportunity to make up for lost blast. A maximum of 77 stope crews are planned at steady state production.

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16.7.6 Underground Development Excavation Sizes and Rates

Table 16-11: Underground Mine Design Criteria and LoM Scheduling Rates

Description	Design Colour	Width	Height	Rate
Main Development
South main infrastructure	1	4.2	4.5	60m/mo
Belt Decline	2	6.5	3.8	100m/mo
Material Decline	5	5.0	4.5	100m/mo
Decline Ramp Access	3	5.0	4.5	100m/mo
Decline Lateral Connecting	4	6.0	4.5	100m/mo
Decline Muck bay	6	6.0	4.5	100m/mo
Turn Bay	8	5.0	4.5	100m/mo
Loading Bay	7	6.0	4.5	100m/mo
Safety Cubby	18	5.0	4.5	60m/mo
Material Transfer Decline	43	5.0	4.5	100m/mo
Silo Connecting	56	5.0	2.1	60m/mo
Silo	57	5.0	-	60m/mo
Ventilation Raise bore Down Cast - 3.5m Ø	19	3.5	-	100m/mo
Ventilation Raise bore Up cast - 3.5m Ø	49	3.5	-	100m/mo
Waste Development	33	4.0	2.1	60m/mo
Hybrid Section
Hybrid Twin Decline/Incline	25	4.0	1.9	60m/mo
Hybrid Twin Decline/Incline Lateral	27	4.0	1.9	60m/mo
Hybrid Ledging Drive	17	4.0	1.8	60m/mo
Hybrid On-reef Access Drive	11	4.0	2.1	60m/mo
Hybrid ASD – On-reef	10	3.5	1.8	30m/mo
Hybrid ASD – Off-reef	14	3.5	1.8	30m/mo
Hybrid stope	39			300ca/mo
Hybrid Ventilation Connecting	20	4.0	2.1	60m/mo
Bord and Pillar Section
Bord and Pillar Strike Drive	40	7.0	1.9	30m/mo
Bord and Pillar Bord	37	7.0	1.9	30m/mo
Bord and Pillar Ventilation Holing	35	3.0	1.9	30m/mo
Bord and Pillar Sacrificial Belt	53	5.0	2.1	30m/mo
Conventional Section
Conventional Raise – On-reef	9	1.5	1.8	45m/mo
Conventional Raise – Off-reef	31	1.5	1.8	45m/mo
Conventional ASG – On-reef	30	1.5	1.8	30m/mo
Conventional ASG – Off-reef	13	1.5	1.8	30m/mo

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Description	Design Colour	Width	Height	Rate
Main Development
Conventional Stope	20			300ca/mo
Overhand Stope	41			200ca/mo
Conventional Waste Development	34	1.5	1.8	60m/mo
Conventional Ventilation Connecting	21	1.5	1.8	60m/mo
Conventional On-reef strike development	47	1.5	1.8	60m/mo

Decline and decline ramp development rates significantly impact project economics and project schedule. Maintaining development rates as per LoM schedule, especially in the early years of mine development prior to commencing ore production is paramount. The nature of the orebody is such that continued high development rates needs to be maintained in order to ensure that the replacement rate is remained as required. Planning achievable and realistic advance rates for underground development has been the subject of much research in this study.

16.7.7 Underground Development Schedule

Early mine development has been scheduled to achieve full production of 160 000 tonnes of reef per month in the shortest most practical time possible, given that the planned development rates are achieved and maintained as per schedule. Both declines on the North Mine are developed concurrent and is scheduled at the planning criteria as per the Table 16-11 above. The South Mine declines were started 251 days after the North Mine declines.

Initial mine development focus:

>	Providing access for personal, materials, and equipment to high-grade, shallow areas of the orebody,
>	Providing an early means of ventilating stopes,
>	Providing two routes for emergency egress at all times,
>	Installing mine services (electric power distribution, communications, compressed air, water supply, mine dewatering).

Graph 16-18 below indicates the individual totals of the development meters per type of excavation over the LoM.

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16.7.8 Production Schedule

This section describes the possibilities according to the production profile and schedule if the scheduling parameters are strictly adhered to. The profile and dates used in this section may differ from what may have been used in the rest of the document.

The WBJV Project 1 (Maseve) LoM inclusive of both the MR and the UG reef is just over 21 years (21.03) . Initially MR is mined for 11.62 years and then UG2 for another 10.94 years. See below Graph 16-19.

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Production on the MR initially builds up to reach 115 000t in May 2016. Production levels are then maintained at 115 000tpm to October 2017 when it builds up again to steady state production of 160 000tpm. Steady state production is reached in September 2018 where it remains for the LoM. The MR production schedule starts to tail down in January 2025 and final reef tonnes is scheduled for May 2026.

In order to keep the mill fed at 160 000 reef tonnes per month, the UG2 reef starts production in November 2024 and builds up steadily to the required 160 000 tonnes of reef in May 2026. During this phase care must be taken in managing the changeover from MR to UG2 reef the plant. Steady state of the UG2 reef lasts up to May 2032 from where UG2 reef production tails down to depletion in in July 2035.

The production profile for each of the individual mining methods applied on the WBJV Project 1 (Maseve) is as shown below in Graph 16-20.

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16.7.9 Drilling

Typical drill patterns are indicative designs as the final designs will be done once a better understanding of the rock types and conditions have been established by practical means. Drilling times and cycles have been based on the number of holes for the various end sizes.

16.7.10 Multi-Blasting

Although it is planned that the mine will operate on timed blasting, in the course of development, it is imperative to provide multi-blast conditions, in order to maintain the rate of development in the main decline development sections. The mines approved code of practise is to be strictly adhered to. Continuous flammable gas monitoring instruments must be used at all drilling locations.

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16.7.11 Cleaning

16.7.11.1 Winch and Scraper Cleaning

Winch and scraper cleaning will be the main cleaning method in the conventional and hybrid productions stopes and Raises. In the conventional mining sections, a series of scraper winches will scrape the broken ore down the face into the ASG and down the centre gully (main raise) and deposit the ore directly into either an ore pass or a LHD loading bay. In the hybrid stoping sections, a face winch will scrape the broken rock from the stope face into the ASD. A LHD will load the broken rock and haul it to the dedicated area to be loaded onto LP 30t trucks for transportation to truck tips.

16.7.11.2 LHD Cleaning

Cleaning calculations have been based on 8t, 10t, 12.5t and a 14t LHD for all ends. Hauling distances for the all on-reef development are further than that of the declines as material and loading bays will be used for mucking. No additional muckbays will be developed in the on-reef development sections.

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Table 16-12: LHD Cleaning of 3.5m W x 2.1m H End

Name	Sandvik LH208L	Cat 1600G	Cat 1700G	Sandvik LH514	Unit	Comments
Description	LHD 8 T	LHD 10 T	LHD 12.5 T	LHD 14 T
3.5m W x 2.1m H face	Value	Value	Value	Value	Unit
Shift Time(10hrs)	600	600	600	600	min
Travel time	30	30	30	30	min
Tool Box Talk and Work allocation	20	20	20	20	min
Making Safe(Water down &bar) - Competent person	30	30	30	30	min
Team Briefing/Deployment	10	10	10	10	min
Mine Clearance and blasting, after end of shift	60	60	60	60	min
Mandatory in shift break	60	60	60	60	min
Available Face Time(after travel in/out & w/place procedures)	390	390	390	390	min
LHD Pre-checks and refuelling - done during make safe	0	0	0	0	min	done during making safe
Move LHD into face	5	5	5	5	min
Operator work place inspection.	10	10	10	10	min
Face width	3.5	3.5	3.5	3.5	m
Height	2.1	2.1	2.1	2.1	m
Advance per blast	3.2	3.2	3.2	3.2	m
In -Situ density	3.00	3.00	3.00	3.00	t/m3
tonnes Blasted per Panel	70.56	70.56	70.56	70.56	t
Swell Factor	1.6	1.6	1.6	1.6	ratio
Bucket capacity	5.35	6.80	8.17	9.35	m³
Bucket fill factor	80%	80%	80%	80%
Effective bucket capacity	4.28	5.44	6.532	7.48	m³
Broken density	1.88	1.88	1.88	1.88	t/m³
Payload	8.03	10.20	12.25	14.03	t
Working shifts per day	2	2	2	2	shifts
Working days per month	25	25	25	25	d
Normal shift duration	10	10	10	10	h
Available working hours per shift, excl. prechecks and preparations	6.25	6.25	6.25	6.25	h
Trip distance - average (Return trip)	200	200	200	200	m
Traveling speed	2.22	2.22	2.22	2.22	m/s	8 Km/h
Time travelled	90.00	90.00	90.00	90.00	sec
Loading /offloading time	180	180	180	180	sec

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Name	Sandvik LH208L	Cat 1600G	Cat 1700G	Sandvik LH514	Unit	Comments
Description	LHD 8 T	LHD 10 T	LHD 12.5 T	LHD 14 T
Trips / hour	13.00	13.00	13.00	13.00	trips
Tons / hour	104.33	132.60	159.22	182.33	tph	No breakdowns and 100% utilisation
LHD In-Shift Availability	85%	85%	85%	85%	%
Required LHD utilisation of available time	80%	80%	80%	80%	%
Effective Possible LHD hrs per shift	5.31	5.31	5.31	5.31	h
tonnes allocated per shift	70.56	70.56	70.56	70.56	t
Overall LHD productivity	70.941	90.168	108.267 9	123.981	tph	Assuming 85% avail & 80% Util.
LHD hours to clean allocated tonnes per shift	1.0	0.8	0.7	0.6	h
LHD tonnes per month	22 169	28 178	33 834	38 744		tonnes

Table 16-13: LHD Cleaning of 5.0m W x 4.0m H End

Name	Sandvik LH208L	Cat 1600G	Cat 1700G	Sandvik LH514	Unit	Comments
Description	LHD 8 T	LHD 10 T	LHD 12.5 T	LHD 14 T
5m W x 4m H face	Value	Value	Value	Value	Unit
Shift Time(10hrs)	600	600	600	600	min
Travel time	30	30	30	30	min
Tool Box Talk and Work allocation	20	20	20	20	min
Making Safe(Water down &bar) - Competent person	30	30	30	30	min
Team Briefing/Deployment	10	10	10	10	min
Mine Clearance and blasting, after end of shift	60	60	60	60	min
Mandatory in shift break	60	60	60	60	min
Available Face Time(after travel in/out & w/place procedures)	390	390	390	390	min
LHD Pre-checks and refuelling - done during make safe	0	0	0	0	min	done during making safe
Move LHD into face	5	5	5	5	min
Operator work place inspection.	10	10	10	10	min
Face width	5	5	5	5	m
Height	4	4	4	4	m
Advance per blast	3.2	3.2	3.2	3.2	m
In -Situ density	3.00	3.00	3.00	3.00	t/m3
tonnes Blasted per Panel	192	192	192	192	t

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Name	Sandvik LH208L	Cat 1600G	Cat 1700G	Sandvik LH514	Unit	Comments
Description	LHD 8 T	LHD 10 T	LHD 12.5 T	LHD 14 T
Swell Factor	1.6	1.6	1.6	1.6	ratio
Bucket capacity	5.35	6.80	8.17	9.35	m³
Bucket fill factor	80%	80%	80%	80%
Effective bucket capacity	4.28	5.44	6.532	7.48	m³
Broken density	1.88	1.88	1.88	1.88	t/m³
Payload	8.03	10.20	12.25	14.03	t
Working shifts per day	2	2	2	2	shift s
Working days per month	25	25	25	25	d
Normal shift duration	10	10	10	10	h
Available working hours per shift, excl. prechecks and preparations	6.25	6.25	6.25	6.25	h
Trip distance - average (Return trip)	200	200	200	200	m
Traveling speed	2.22	2.22	2.22	2.22	m/s	8 Km/h
Time travelled	90.00	90.00	90.00	90.00	sec
Loading /offloading time	180	180	180	180	sec
Trips / hour	13.00	13.00	13.00	13.00	trips
Tons / hour	104.33	132.60	159.22	182.33	tph	No breakdowns and 100% utilisation
LHD In-Shift Availability	85%	85%	85%	85%	%
Required LHD utilisation of available time	80%	80%	80%	80%	%
Effective Possible LHD hrs per shift	5.31	5.31	5.31	5.31	h
tonnes allocated per shift	192	192	192	192	t
Overall LHD productivity	70.941	90.168	108.2679	123.981	tph	Assuming 85% avail & 80% Util.
LHD hours to clean allocated tonnes per shift	2.7	2.1	1.8	1.5	h
LHD tonnes per month	22 169	28 178	33 834	38 744		tonnes

Table 16-14: LHD Cleaning of 6.5m W x 3.8m H End

Name	Sandvik LH208L	Cat 1600G	Cat 1700G	Sandvik LH514	Unit	Comments
Description	LHD 8 T	LHD 10 T	LHD 12.5 T	LHD 14 T
6.5m W x 3.8m H face	Value	Value	Value	Value	Unit
Shift Time(10hrs)	600	600	600	600	min
Travel time	30	30	30	30	min

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Name	Sandvik LH208L	Cat 1600G	Cat 1700G	Sandvik LH514	Unit	Comments
Description	LHD 8 T	LHD 10 T	LHD 12.5 T	LHD 14 T
Tool Box Talk and Work allocation	20	20	20	20	min
Making Safe(Water down &bar) - Competent person	30	30	30	30	min
Team Briefing/Deployment	10	10	10	10	min
Mine Clearance and blasting, after end of shift	60	60	60	60	min
Mandatory in shift break	60	60	60	60	min
Available Face Time(after travel in/out & w/place procedures)	390	390	390	390	min
LHD Pre-checks and refuelling - done during make safe	0	0	0	0	min	done during making safe
Move LHD into face	5	5	5	5	min
Operator work place inspection.	10	10	10	10	min
Face width	6.5	6.5	6.5	6.5	m
Height	3.8	3.8	3.8	3.8	m
Advance per blast	3.2	3.2	3.2	3.2	m
In -Situ density	3.00	3.00	3.00	3.00	t/m3
tonnes Blasted per Panel	237.12	237.12	237.12	237.12	t
Swell Factor	1.6	1.6	1.6	1.6	ratio
Bucket capacity	5.35	6.80	8.17	9.35	m³
Bucket fill factor	80%	80%	80%	80%
Effective bucket capacity	4.28	5.44	6.532	7.48	m³
Broken density	1.88	1.88	1.88	1.88	t/m³
Payload	8.03	10.20	12.25	14.03	t
Working shifts per day	2	2	2	2	shifts
Working days per month	25	25	25	25	d
Normal shift duration	10	10	10	10	h
Available working hours per shift, excl. prechecks and preparations	6.25	6.25	6.25	6.25	h
Trip distance - average (Return trip)	200	200	200	200	m
Traveling speed	2.22	2.22	2.22	2.22	m/s	8 Km/h
Time travelled	90.00	90.00	90.00	90.00	sec
Loading /offloading time	180	180	180	180	sec
Trips / hour	13.00	13.00	13.00	13.00	trips
Tons / hour	104.33	132.60	159.22	182.33	tph	No breakdowns and 100% utilisation
LHD In-Shift Availability	85%	85%	85%	85%	%

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Name	Sandvik LH208L	Cat 1600G	Cat 1700G	Sandvik LH514	Unit	Comments
Description	LHD 8 T	LHD 10 T	LHD 12.5 T	LHD 14 T
Required LHD utilisation of available time	80%	80%	80%	80%	%
Effective Possible LHD hrs per shift	5.31	5.31	5.31	5.31	h
tonnes allocated per shift	237.12	237.12	237.12 108.267	237.12	t
Overall LHD productivity	70.941	90.168	9	123.981	tph	Assuming 85% avail & 80% Util.
LHD hours to clean allocated tonnes per shift	3.3	2.6	2.2	1.9	h
LHD tonnes per month	22 169	28 178	33 834	38 744		tonnes

One LHD will be able to clean multiple ends per shift. The distance between these ends will determine the number of ends that a single LHD can clean.

The main ramp decline system will be cleaned by one LHD. Broken rock will be loaded at the blasted face and tipped into the muck bays excavated for that purpose. A second LHD will load the broken rock from the muck bays onto the trucks. The LHD loading the face will also load directly into the truck when it is available and in position.

16.7.12 Hauling

The average RoM tonnes derived from the LoM schedule is 170 000t. This amounts to a required minimum hauling rate of 231tph. In order to maintain this minimum rate over variable distances, the number of operating trucks will have to be adjusted accordingly. See Graph 16-21.

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Hauling cycles for a 30t truck were calculated for the various end sizes and also for hauling at various distances. Development truck cycles are not part of the critical on-face cycles; however the hauling from the stoping sections and loading bays is critical as to ensure efficient cleaning of the stopes and raises.

Table 16-15: 30t Trucks – Cycles

Name	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Unit	Comments
Description	Truck 30T	Truck 30T	Truck 30T	Truck 30T	Truck 30T
	Value	Value	Value	Value	Value	Unit
Shift Time(10hrs)	600	600	600	600	600	min
Travel time	0	0	0	0	0	min
Tool Box Talk and Work allocation	10	10	10	10	10	min
Making Safe(Water down &bar) - Competent person	0	0	0	0	0	min
Team Briefing/Deployment	10	10	10	10	10	min

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Name	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Unit	Comments
Description	Truck 30T	Truck 30T	Truck 30T	Truck 30T	Truck 30T
	Value	Value	Value	Value	Value	Unit
Mine Clearance and blasting, after end of shift	60	60	60	60	60	min
Mandatory in shift break	60	60	60	60	60
Available Face Time(after travel in/out & w/place procedures)	460	460	460	460	460	min
Pre-checks and refuelling - done during make safe	10	10	10	10	10	min	done during making safe
Move Truck into loading area	5	5	5	5	5	min
Operator work place inspection	0	0	0	0	0	min
Face width	3.5	5	6.5	6.5	6.5	m
Height	2.1	4	3.8	3.8	3.8	m
Advance per blast	3.2	3.2	3.2	3.2	3.2	m
In -Situ density	3.00	3.00	3.00	3.00	3.00	t/m3
tonnes Blasted per Panel	70.56	192	237.12	237.12	237.12	t
Swell Factor	1.6	1.6	1.6	1.6	1.6	ratio
Bucket capacity	14.5	14.5	14.5	14.5	14.5	m³
Bucket fill factor	110%	110%	110%	110%	110%
Effective bucket capacity	15.95	15.95	15.95	15.95	15.95	m³
Broken density	1.88	1.88	1.88	1.88	1.88	t/m³
Payload	30	30	30	30	30	t
Working shifts per day	2	2	2	2	2	shifts
Working days per month	25	25	25	25	25	d
Normal shift duration	10	10	10	10	10	h
Available working hours per shift, excl. prechecks and preparations	7.42	7.42	7.42	7.42	7.42	h
Trip distance - average (Return trip)	2000	2000	2000	3000	4000	m
Traveling speed	2.22	2.22	2.22	2.22	2.22	m/s	8 Km/h
Time travelled	900.00	900.00	900.00	1 350.00	1 800.00	sec
Loading /offloading time/Turning	223.4	223.4	223.4	223.4	223.4	sec
Trips / hour	3.00	3.00	3.00	2.00	1.00	trips
Tons / hour	90.00	90.00	90.00	60.00	30.00	tph	No breakdowns and 100% utilisation
LHD In-Shift Availability	85%	85%	85%	85%	85%	%
Required Truck utilisation of available time	80%	80%	80%	80%	80%	%

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Name	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Cat AD30 - Sandvik EJC533	Unit	Comments
Description	Truck 30T	Truck 30T	Truck 30T	Truck 30T	Truck 30T
	Value	Value	Value	Value	Value	Unit
Effective Possible Truck hrs per shift	5.93	5.93	5.93	5.93	5.93	h
tonnes allocated per shift	70.56	192	237.12	237.12	237.12	t
Overall Truck productivity	61.20	61.20	61.20	40.80	20.40	tph	Assuming 85% avail & 80% Util.
Truck hours to clean allocated tonnes per shift	1.2	3.1	3.9	5.8	11.6	h	h
Truck tonnes per month	22 695	22 695	22 695	15 130	7 565		tonnes

16.7.13 Underground Hydro Powered Equipment

A study on the use of HPE equipment was done by Snowden (report: 130614_D_JR###-##-####_PTM_Maseve drilling report). Based on the outcome of the study the decision was taken to pursue and implement the use of HPE equipment on the WBJV Project 1 (Maseve). Typical HPE stope layout.

Localised high-pressure water is generated within the hydropower pack which is stationed at dedicated cubbies in the access or ledging drives. Every four panels is equipped with two centralised power pack and reticulation system. Pressurised water delivered at 18MPa is distributed by piping from the dedicated cubby to the producing stopes. When mining is completed, the piping infrastructure may be dismantled and together with the power pack, be moved to a new position where it can provide pressurized water for new stopes.

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The centralised power pack consists of a series of pumps operated in parallel. The pumps are mounted on a flat car, and can be moved either by means of trackbound or trackless mobile equipment. The water is supplied to the working areas using high pressure piping, with pressure reduction and take-offs feeding the individual panels.

The high-pressure water is used to power both primary and ancillary HPE equipment (primary equipment: HPE rock drill, roofbolter. ancillary equipment: HPE chainsaw, HPE jackpot gun, HPE explosive hole desludger, HPE Sweeping tool, HPE jet pump) deployed in the stope. One or more manifolds are used to supply high pressure water to the operating equipment via hoses and can be operated with, or without a water return line. Water can be dumped, or collected via a return hose into a sump where it can be filtered and recirculated to power the drill. During drilling approximately 70% of water can be recycled, with the difference required as make-up water. The explosives emulsion pump in not HPE equipment but is an electric pump used to charge the shot holes.

The hydropower equipment incorporates excess flow control safety devices, such as hydraulic fuses that instantly “trip” and isolate the flow if the flow exceeds a safe limit. The stored energy in the small-diameter localized piping systems is relatively low and positive displacement pumps loose pressure when a significant “bleed” appears in the system. The power packs have individual earth leakage protection and many other protection devices.

16.7.14 Underground Mobile Equipment

The mining equipment is based on the latest technology. Sizing and selection is done on a combination of equipment capabilities and operating experience in mines with similar conditions.

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Table 16-16: WBJV1 Trackless Mobile Machinery Equipment List: Current vs Future

Equipment	Type	Payload	Rate per hour	Rate per month	Eng Kw	Current Comp	Future Comp	Delta
Cat AD30	Truck	30T	122	45 390	304	13	-	-13
Sandvik EJC533	Truck	30T	122	45 390	298	3	11	8
Cat 1600G	LHD	10T	90	28 178	200	4	-	-4
Sandvik LH410m	LHD	10T	90	28 178	204	1	-	-1
Sandvik LH208/9L	LP LHD	10T	90	28 178	204	0	21	21
Cat 1700G	LHD	12.5T	108	33 834	263	2	-	-2
Sandvik LH514	LHD	14T	124	38 744	256	1	6	5
Sandvik DD320- 26	Drill Rig - Twin Boom		41	14 541	74	1	5	4
Sandvik DD321- 40	Drill Rig - Twin Boom		41	14 541	110	9	-	-9
Sandvik D310	Drill Rig - Single Boom		20	8 156	62	1	-	-1
Sandvik DD211L	LP Drill Rig - Single Boom		20	8 156	55	0	9	9
Maclean Scissor Bolter	Bolter		41	14 541	110	1	7	6
Sandvik DS210L	LP Bolter		41	14 541	74	0	9	9
Fermel UV	UV		-	40 000	85	7	6	-1
Fermel RoRo	UV		-	40 000	85	9	6	-3
Fermel Grader	Grader		-	170 000	85	1	1	-
Cat 120K LP	Grader		-	170 000	93	1	-	-1
Fermel Tmech	Transporter		-	-	85	1	-	-1
Kubota	Transporter - Supervision 1		-	-	18.5	14	9	-5
Cat Bobcat	Construction		-	-	53.8	2	-	-2
Toyota Landcruiser	Transporter - Supervision 2		-	-	96	4	4	0

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>	Underground haul trucks and LHD’s will be equipped with engines using DDEC technology to optimize efficiency and to minimize emissions,
>	All articulating equipment will be fitted with articulation slings, rotating lights, red and green lights and tape to indicate safe and dangers zones as per site’s Man/Machine interface plan,
>	Collision avoidance and people detection devices will be fitted should the requirement be proven by means of a detailed risk assessment,
>	Steering and break interlocking safety systems, on-board fire suppression systems and engine protection systems will be standard features.

16.7.15 Underground Services

Underground services will include:

>	Mine dewatering,
>	Drilling potable and fire-fighting water,
>	Compressed air,
>	Electric power,
>	Communications.

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Mine services will be installed progressively as headings are advanced.

16.7.16 Compressed Air

Compressed air will be used underground for:

>	Refuge bay fresh air pressurisation (Primary use),
>	Secondary mine water pumping using pneumatic pumps during sinking operations,
>	Explosive emulsion loading during sinking operations,
>	Blasthole cleaning in designated areas,
>	Miscellaneous pneumatic small tools during sinking operations.

16.7.17 Water Supply

Industrial-quality water will be distributed by steel pipelines throughout the underground workings for drilling equipment, dust suppression, and cleaning of areas. Flexible hoses will be used to connect water pipelines to drilling equipment at working faces. A potable water tank located on surface will provide an uninterrupted supply of fresh water and fire-fighting water. Water pressure of potable and industrial water will be controlled by pressure reducing stations strategically placed underground to ensure water pressure is controlled and can be used by personnel and machinery respectively underground..

16.7.18 Mine Dewatering

Water inflows to the underground mine derive from groundwater inflows, drilling water, HPE equipment (hydro mining), drinking water, potable water used for firefighting and dust suppression water.

Mine water on strike haulages and cross-cuts will be directed by gravity to pump niches located at no more than 600 m intervals, from where it will be pumped horizontally to a dirty water pumping station. Dirty water pumping stations will be located on every second strike haulage, providing a 120 m vertical interval from one dirty water pump station to the next. Mine water will be pumped progressively from one pump station to the next until arrival at surface. Main water pipeline can dewater at no less than 250m³per hour.

Ultimately, a clean water pumping station will be installed on the 45 Level, allowing more efficient and cost effective clean-water pumping to be performed from 45 Level to surface. A cascade settling system, using muckbays excavated at 9 degrees, will be used to remove a large portion of suspended solids that can be dug out using an LHD when required.

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16.7.19 Underground Communication System

A leaky feeder communication system will be installed as the communication system for the mine and surface operations. Telephones will be located at key infrastructure locations such as chairlift stations, electrical substations, conveyor drives, refuge areas, lunchrooms, and pumping stations.

Key personnel (such as mobile mechanics, crew leaders, and shift bosses) and mobile equipment operators (such as loader, truck, and utility vehicle operators) will be supplied with underground radios for contact with the leaky feeder network.

A fire network will be established underground consisting of process network control and communications systems to surface, the fire network will be designed using single mode technology allowing for long transmission distances. Systems allowed for on the fire network will include. Standby electrical protection network, pump station control, access and security control, tracking and tagging control, sub decline chairlift SCADA, environmental systems, bulk materials handing control, paging systems, fire system monitoring and other system capacity for further expansion.

16.7.20 Underground Construction and Mine Maintenance

Supervised crews will perform general mine maintenance and construction work, ground support control and scaling, road checking and maintenance, construction of ventilation doors, bulkheads, and concrete work, mine dewatering, services movement, and safety work. Underground graders and scissor lifts will be used to maintain haulages travelled by mobile equipment.

16.7.21 Underground Mobile Equipment Maintenance

Initially, mobile underground equipment will be maintained in a shop located on surface close to decline portals. Ultimately, it will become economical to save vehicle travel time by establishing underground workshops and service bays. Surface and underground workshops are described fully in Section 18.2.

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A maintenance supervisor will provide a daily maintenance work schedule, ensure the availability of spare parts and supplies, and provide management and supervision of maintenance crews. He will also manage training of the maintenance workforce.

A maintenance planner will schedule maintenance and repair work and provide statistics on equipment availability, utilization and life cycle, mine efficiency, and personnel utilization. A computerized system using commercially available software will facilitate planning.

Equipment operators will provide equipment inspection at the beginning of the shift and perform small maintenance and repairs as required throughout the shift.

Mechanics using a mechanic’s truck will perform minor repairs and emergency repairs on equipment and vehicles where these cannot make their way to a workshop.

Major rebuild work will be conducted off site.

16.7.22 Underground Workshops:

>	The workshops have been designed to be connected to the main return airways and will be well ventilated and to be kept clean of waste and oil spillage.
>	The dirty water will be drained away to an oil separator. The recovered oil will be returned to surface and the dirty water will be pumped to the surface dams.
>	A brake test ramp will form part of the workshop exit as to ensure compliance in terms of brake testing after each service or repair.

The Maintenance Supervisor will provide a daily maintenance work schedule, ensure the availability of spare parts and supplies, and provide management and supervision of maintenance crews. He/she will also manage the training needs of the maintenance workforce.

A Maintenance Planner will schedule maintenance and repair work and provide statistics on equipment availability, utilization, life cycle, mine efficiency, and personnel utilization. A computerized system using commercially available software will facilitate planning.

Equipment Operators will do equipment inspection at the beginning of the shift and perform small maintenance and repairs as required throughout the shift. Mechanics using a mechanic’s vehicle will perform minor repairs and emergency repairs on equipment and vehicles where these cannot make their way to a workshop. Major rebuild work will be conducted off site.

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Maintenance, servicing and workshop facilities will be provided throughout the mine where deemed necessary. These workshops will provide the necessary resources to maintain the mining equipment and allow for weekly and monthly scheduled minor services and minor repairs. All major services and breakdowns will be done in the surface main workshop.

Equipment will be scheduled by the Engineering Planner to report to the underground workshops for servicing and maintenance. Refuelling, washing, daily maintenance and lubrication of mechanised machinery will take place at these designated diesel / lubrication bays located on the relevant production level. The underground diesel / lubrication bay will be provided with lighting; compressed air, service, and potable water and contain firefighting equipment.

The maintenance schedule will consist of a one hour daily service every 24 hours. The inspection, refuelling and lubrication of the mining equipment will take place during the six hours of no activity. Services are all hourly based services which are conducted at 125 hour, 250 hour, 500 hour, 1 000 hour and 2 000 hour. A 125 hour service takes four to six hours, a 250 hour service takes six hours, a 500 hour service takes about eight hours, a 1 000 hour service takes approximately 10 hours and a 2 000 hour service can last between 10-12 hours. Percussion, power pack and compressor services on the drill rig and the rock bolter are scheduled at intervals of 60 hours, 125 hours, 250 hours and 500 hours.

On-reef machinery consisting largely of electric drills, pumps, and scraper winches will adhere to a planned maintenance schedule that will consist of a daily inspection by the Operator as well as the use of a pre-use inspection checklist at the beginning of the shift and which will require the signing of the checklist by the Operator. Weekly inspection and servicing of the equipment will be performed by the artisan responsible for that section.

Conveyor maintenance will consist of belt splices, drive and return maintenance, and replacement of worn idlers. It is anticipated that any planned maintenance on the conveyors will be scheduled to have the lowest effect during production periods.

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16.7.23 Underground Logistics

The efficient movement of rock to surface, and delivery of consumable supplies and equipment to where it is needed for development and production underground, together with the recovery of equipment for refurbishment, are key to achieving the planned production rate within the planned cost structure.

To this end, state-of-the-art diesel trackless vehicles comprising 20t haul trucks, cassette carriers, and flat-bed trucks will be used to physically move rock, materials and equipment underground. To complement this equipment, appropriate levels of logistics management, supervision, and materials control hardware and software are allowed for.

Waiting places, refuge bays, mining equipment bay and toilet facilities will be provided per mining section and will be located in lay-bays and footwall drives in close proximity to the working faces. All major services on equipment, belt movement or extensions will normally take place during weekends to prevent production stoppages and delays. Service pipes together with ventilation columns, ventilation walls and all electrical cables extensions will be planned as part of the normal weekly mining cycle. Stoping material will be transported to the material bay located at the end of each production crosscut where the cassette will be offloaded and empty cassettes loaded for transport back to surface.

16.7.24 Underground Ore and Waste Handling

Refer to Section 18.2.2 for bulk materials handling details.

16.7.24.1 Reef Flow

Reef produced in the mining sections will be transported from the ASD’s and or bords/strike drives (hybrid and bord and pillar sections respectively) or Loading bays (Conventional sections) with LHD’s via the dedicated access drives to dedicated tipping bays. Here the ore will be transferred to 30 ton trucks, which will transport the broken rock to the main underground silos or onto the conveyor belts where equipped. A hydraulic hammer will break the oversized rock to fit through the 300 x 300 grizzly. The ore will be store in an underground silo.

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16.7.24.2 Waste Flow

Waste will be mined in development areas by means of electric drill rigs and LHD’s. The waste will be transferred to 30t Dump trucks. The truck will haul the waste to the silo waste tip. A Hydraulic hammer will break the oversized rock to fit through the 300 x 300 grizzly.

16.7.24.3 Silo Feed

A single electrical vibrating pan feeders per silo will control the flow of ore from each silo to the Dip 2 conveyor belt. Dip 2 conveyor will transfer the ore to Dip 1 conveyor which will convey the ore to the surface transfer tower. The underground portion of Dip1 and Dip 2 conveyors will be suspended from the hanging wall. Unless otherwise recommended by Rock Engineer.

16.7.25 Underground Transport

16.7.25.1 Transportation of Underground Workers

The majority of underground workers will make use of chairlift transport between surface and their workplaces. Workers will walk from where they get off the chairlifts to the stopes in the ramps. No truck haulage will take place during major shift movements.

Chairlifts convey at up to 900 persons per hour, and will allow a complete shift-change to be accomplished in less than 2 hours at the furthest distance of stopes from surface. Maximum walking distance will be generally be 1 000m.

Some parts of the deposit are small and isolated and walking distances may exceed 1 000m. Trackless personnel carriers will be used in these cases.

16.7.25.2 Transportation of Supervisory and Technical Personnel Underground

Supervisors and technical personnel will have convenient and rapid access to underground operations by means of light diesel supervision vehicles throughout the LoM.

16.7.25.3 Transportation of Materials and Supplies Underground

Utility vehicles (UV’s) and Roll-on Roll-off (RoRo) vehicles will transport mining consumables, material and equipment to the underground working places. RoRo vehicles will predominantly be 8 tonne cassette carrier diesel vehicles having the capability of offloading cassettes containing supplies and equipment at the destination point within the mine, and self-loading empty cassettes for removal to surface.

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Consumables and materials will be bundled, palletized on standard wooden pallets, or supplied in standard bulk bags, for ease of movement and on and off loading. Stoping ground support supplies will be assembled on surface and delivered in unit loads. The unit loads will contain all the materials and supplies required by an operating raise-line.

Ground support materials cassettes will be off-loaded in cross-cut storage bays (timber bays) from where it will be transported into the raise-lines and stopes by means of mono-rope winches.

16.7.25.4 Transportation and Storage of Explosives Underground

Explosives for all main and footwall development will be of the emulsion type with electric initiation. Emulsion explosives will be delivered to bulk silos on surface. From there it will be loaded into special explosives vehicles for delivery on a daily basis to underground development headings. Packaged explosive will be transported as palletized unit loads and offloaded direct from the delivery truck at an explosives receiving area on surface from where it is loaded into cassettes and delivered to underground explosive stores.

16.7.25.5 Fuel Storage and Distribution

See Section 18.4 Fuel Storage.

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17.0 RECOVERY METHODS

17.1 Process Design

The recovery of saleable product from the WBJV Project 1 (Maseve) will take place on site at the Maseve Concentrator. Moist flotation concentrate as a filter cake containing recoverable platinum, palladium, rhodium and gold will be produced from the concentrator plant for shipping to the smelter. The concentrate will also contain ruthenium and iridium in small amounts. Two ore types will be mined, MR and UG2 during the life of the mine. The MR concentrate will contain copper and nickel in small proportions, whilst UG2 concentrates will be significantly less in these metals.

The MR ore will be mined initially followed by UG2. The flowsheet selected for the processing plant applies a standard MF2 circuit configuration, widely used on both the South African Eastern and Western Bushveld operations.

Based on previous metallurgical testwork conducted by SGS Lakefield and Mintek, South Africa DRA were commissioned to complete detail design and construct a concentrator plant with a nameplate capacity of 165 000tpm to treat both MR and UG2 ores. The design approach is for a plant that is capable of treating either pure MR, blend of MR and UG2 ores and pure UG2 ore. The facility will treat pure MR in the early years of the mine life. Towards the depletion of the MR the plant will be capable of treating a blend of MR and UG2 in an 80:20 ratio and eventually pure UG2 ore.

The plant is designed to operate at an average overall availability of 92%. The Maseve Concentrator will be constructed in two phases. This approach offers an opportunity for early production to generate revenue at a lower initial capital outlay, while accommodating the slow mining ramp-up to steady state production. In phase 1 only equipment required to treat between 80 000 and 115 000tpm will be installed and configured in an MF1 flowsheet. All the civil and structural scope required for the rest of the equipment will be completed in phase 1 in order to minimise interrupting with production when implementing phase 2. Consideration was also given to constructability on an operating plant when deciding the civil and structural scope to be completed for phase 1. All the services required for the full phase 2 plant (namely bulk power supply, water supply, air and reagents) will be completed in phase 1.

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Testwork demonstrated that the MR ore can be treated efficiently applying an MF1 flowsheet configuration at slight loss in recovery compared to an MF2 circuit (1 to 3% as per testwork) with extended rougher flotation residence time and extended cleaner flotation residence time and additional collector. Figure 17-1 below shows the grind-recovery relationship for Merensky MF1 rougher rate tests.

The overall MF2 flowsheet includes the following key areas:

>	Ore receiving, crushing and storage,
>	Primary milling,
>	Primary rougher flotation,
>	Secondary milling,
>	Secondary rougher flotation,
>	Integrated cleaner flotation consisting of primary and secondary cleaner flotation,
>	Tailings thickening and disposal,
>	Concentrate thickening and filtration.

For the phase 1 (MF1 flowsheet Figure 17-3 below, with equipment for future in dotted lines), the secondary mill circuit and primary rougher flotation equipment will not be installed. The secondary rougher flotation circuit will be deployed to primary rougher duty in phase 1 because of the higher residence time it offers and its routing to final tails. Figure 17-2 below illustrates the overall MF2 flowsheet

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17.2 Process Description Summary

RoM ore will be conveyed from underground using an overland conveyor belt to a transfer station on surface. The transfer station as an option to feed material into the plant or an emergency ground stockpile. A re-load facility is also provided. From the transfer station the ore is conveyed via an overland conveyor to a crushing station. Belt magnets are installed at the head pulleys of the overland conveyor and crusher feed conveyor. Crusher product is conveyed to a 10 000t (live) Mill Feed Silo.

Crushed ore is extracted from the Mill Feed Silo at a controlled rate using belt feeders to the primary mill ball mill in closed circuit with a classification cyclone cluster. Water is introduced to the mill to produce a slurry which is pumped to the cyclones. Cyclone underflow gravitates to the mill and overflow gravitates to an agitated rougher feed surge tank. Slurry is pumped from the rougher feed surge tank to the primary rougher flotation circuit. Two concentrates are drawn from the primary rougher flotation circuit. The first concentrate (high grade) forms feed to the primary recleaners and the second concentrate forms feed to the primary cleaners.

Primary rougher tails is pumped to the secondary mill (in phase1 / MF1 circuit the tails is pumped to the tailings disposal area). Secondary mill product is pumped to the secondary rougher flotation circuit. Two concentrates are drawn from the secondary rougher flotation circuit. The first concentrate is routed to the primary recleaners and the second concentrate to the primary cleaners. Secondary rougher tails is pumped to the tailings disposal area for dewatering in the guard cyclones and thickener before being pumped to a TSF.

The primary cleaner flotation section consist of a three stages of cleaning (a fourth stage of final concentrate was included) and the two banks of secondary cleaner. The primary cleaner tails form feed to the secondary cleaners. Two concentrates are produced. High grade from the primary cleaner section and low grade from the secondary cleaner flotation section are pumped to the concentrate thickener and the concentrate thickener underflow is filtered on a continuous moving horizontal belt filter press. The filter cake is stored in bulk on site under cover for onward transportation to a smelter.

Overflow from the tailings thickener and concentrate thickener is stored on the plant and used as process water. Return water from the TSF is also used as process water.

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17.3 Control System

Sufficient level of automation has been provided for to allow the plant to be operated safely and offer good protection to equipment as well as process control for stabilising metallurgical performance. Field instrumentation will be installed to allow for control and data logging via a PLC / Supervisory Control and Data Acquisition (“SCADA”) system. Normal plant start-up will be by sequence control from the PLC. Remote monitoring of all services directly impacting on the operation of the installed major items of equipment will be included in the control solution. No in-stream analysing equipment has been installed at the plant during phase 1 development.

17.4 Recoverability

At 3.1% and 3.3% mass pull between 2 500t and 3 300tpm of concentrate are expected to be produced in phase 1. The phase 1 plant for MR concentrate grade is expected to be greater than 135g/t (4E) and up to 150g/t (4E) at a recovery of 86% (4E). Expected UG2 recovery is greater than 82% at a grade greater than 100g/t (4E) but as UG2 ore is not anticipated to be processed through the phase 1 plant, this estimate is not relevant.

With the implementation of phase 2, the concentrate grade and metal recoveries are expected to improve to that presented in the Feasibility Studies and as reported in Section 13 of this Technical Report.

17.5 Concentrator Status

As at the date of this Technical Report, the concentrator plant is a few weeks away from being hot commissioned. Cold commissioning has commenced and water testing was to be conducted. There is a stockpile of development reef on surface, adequate to commence the hot commissioning process.

Major equipment is depicted in the following site photographs, as at the effective date of this report.

The Tailings Storage Facility is ready to accept tailings from the hot commission process and early production.

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17.6 Plant Operations

The plant will be commissioned with the team from DRA Projects and the assistance of a contract plant operator. The contract operator will then take responsibility for plant operations, reporting the Mine Manager.

A contract tailings operator will manage the deposition of tailings upon the TSF.

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18.0 PROJECT INFRASTRUCTURE

18.1 Site Layout and Access Roads

The Main access route onto site, running from South to North, is the R565 which is a tarred double lane Provincial road, splitting the property in two.

The site is divided into four secure areas, the operational site is primarily focused east of the R565 namely North Shaft, Plant, and South Shaft. The training and induction centre functions are primarily focused West of the R565 at the training centre. Each shaft and the plant is equipped with its own offices, change house, control room, maintenance, storage, and general management facilities. Senior management offices are located near the Plant central to all operational areas.

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North Shaft is responsible for 60% of the mines production and will have a larger number of staff working out of it, the workshops are larger to accommodate more vehicles and discussed in the workshop section below. The change house fitted with seven large bays is capable of accommodating 1 200 people.

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South Shaft responsible for the remaining production and has a smaller staff compliment working out of it, workshop are also smaller and the change house with five large bays is capable of accommodating 800 people.

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A number of tarred access roads have been constructed around the property including:

>	Main provincial road access upgrade from the R565,
>	North Mine access road,
>	Plant access road shared by the South Mine access road.

Secondary access and service roads have a compacted gravel surface. Dust mitigation is affected with the use of a water bowser filled up at the pollution control dams, and a grader maintains a good surface.

The R565 will be the main route for concentrate transport. The daily ore concentrate tonnage to be hauled once steady-state production is achieved is estimated at 160t per day. The concentrated ore is expected to be transported by means of 2-axle trucks with a capacity of 25t per vehicle. Therefore the estimated truck trips per day are seven trips. A complete traffic impact assessment was completed and is part of the EIA.

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18.2 Underground Trackless Mobile Machinery and Logistics

18.2.1 Trackless Mobile Machinery (TMM) and Workshops

A Maintenance Supervisor will provide a daily maintenance work schedule, ensure the availability of spare parts and supplies, and provide management and supervision of maintenance crews. He will also manage training of the maintenance workforce.

A Maintenance Planner will schedule maintenance and repair work and provide statistics on equipment availability, utilization and life cycle, mine efficiency, and personnel utilization. A computerized system using commercially available software will facilitate planning.

Equipment operators will provide equipment inspection at the beginning of the shift and perform small maintenance and repairs as required throughout the shift. Mechanics using a mechanic’s truck will perform minor repairs and emergency repairs on equipment and vehicles where these cannot make their way to a workshop. Major rebuild work will be conducted off site.

Initially, TMM will solely be maintained in a workshop located on surface close to decline portals. The surface workshops at North and South are capable of executing minor and major services and minor and major repairs to all types of TMM equipment. The North and South portal workshops have the following facilities:

>	4 wash bays (North),
>	1 wash bay (South),
>	5 service bays (North),
>	4 service bays (South),
>	3 outdoor lube bays (North),
>	3 service pits (North),
>	2 service pits (South),
>	A sump with dirty water processing capabilities through a drizit to skim off oil and other hydrocarbons (Both),
>	Bulk lubrication dispensing system (Both),
>	Overhead crane (Both),
>	Spares and tool storage (Both),
>	Utilities – power, water, and compressed air (Both),
>	A waste management area (Both),
>	Toilet facilities (Both).

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Ultimately, it will become economical to save vehicle travel time and establish underground TMM workshops. The first underground workshop has been established at Gillooly’s and is currently in construction. The workshop will be capable of executing minor services and minor repairs to all types of TMM equipment.

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The Gillooly’s workshop will have the following facilities when complete:

>	2 wash bays,
>	4 service bays,
>	A single office with a small break room,
>	A sump with dirty water processing capabilities through a drizit to skim off oil and other hydrocarbons,
>	A refuel facility,
>	A low storage capacity lubrication dispensing system to two of the service bays,
>	Spares and tool storage,
>	Utilities – power, water, and compressed air,
>	A waste management area.

18.2.2 Underground Logistics and Rock Handling

The efficient movement of rock to surface, and delivery of consumable supplies and equipment to where these are needed for development and production underground, together with the recovery of equipment for refurbishment, are key to achieving the planned production rate within the planned cost structure.

To this end, state-of-the-art diesel trackless vehicles comprising 30t haul trucks, multipurpose cassette carriers, and flat-bed trucks will be used to physically move rock, materials and equipment underground respectively. To complement this equipment, appropriate levels of logistics management, supervision, and materials control hardware and software are allowed for.

Ore and Waste Rock Movement

Waste Rock (North Mine)

Waste rock from decline, footwall, and stope development will be hauled to waste dumps on surface by 30t haul trucks via two haulage declines until a bulk material handling facility is established underground.

The North Mine Decline conveyor system will be commissioned and waste rock produced on or below the 21 Level will be hauled to an underground waste rock bin on the 21 Level, which will feed onto the North Mine decline conveyor system. At the transfer onto the surface overland belt leading the plant a throw out conveyor will be built dumping waste rock in a conical stockpile, at a -300mm nominal size. The stockpile will be re-handled and dumped onto the waste stockpile using surface or underground mobile machinery as required. Waste rock originating from above the 21 Level waste will continue to be hauled directly to surface because of the safety-based decision made not to perform loaded down-hill hauling.

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Ore (North Mine)

Ore will be loaded from the ore passes in each raise line directly into 30t haul trucks, tonnage and size of haul trucks will be dependent on excavation height and will reduce in size and carrying capacity as excavation height varies.

Early in the LoM, ore from stopes and from reef development will be hauled in 30t haul trucks directly to an ore stockpile on surface. Ore will be crushed down to a minus 300mm product. The crushed product will be loaded via surface LHD into a reload hopper discharging onto the overland conveyor, leading to the process plant ore bin on surface.

The conveyor system feeding the plant silo directly via the crusher has full Variable Speed Drive (“VSD”) control and is capable of slowing down feed belts if the crusher is overloaded. Underground transfer belts feeding the main belt, fitted with belt scales, are fed by VSD controlled feeders. The feeders draw from the underground silos, to ensure fast adjustment of feed to the main belts.

Once the North Mine Decline conveyor system is commissioned, ore produced on or below the 21 Level will continue to be hauled by truck to two ore bins on the 21 Level, which will feed onto the North Mine Decline conveyor, which will convey it to the surface ore bin. Ore produced above the 21 Level will continue to be hauled directly to surface by haul trucks as a result of the safety-based decision made not to perform loaded down-hill hauling.

Waste Rock (South Mine)

Waste rock from decline, footwall, and stope development will be hauled to waste dumps on surface by 30t haul trucks via two Haulage Declines until a bulk material handling facility is established underground.

The South Mine decline conveyor system will be commissioned and waste rock produced on or below the 21 Level will be hauled to an underground waste rock bin on the 21 Level, which will feed onto the South Mine decline conveyor system. At the transfer onto the surface overland belt leading the plant a throw out conveyor will be built dumping waste rock in a conical stockpile, at a -300mm nominal size. The stockpile will be re-handled and dumped onto the waste stockpile using surface or underground mobile machinery as required. Waste rock originating from above the 21 Level Waste will continue to be hauled directly to surface because of the safety-based decision made not to perform loaded down-hill hauling.

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Ore (South Mine)

Ore will be loaded from the ore passes in each raise line directly into 30t haul trucks, tonnage and size of haul trucks will be dependent on excavation height and will reduce in size and carrying capacity as excavation heights vary.

Early in the LoM, ore from stopes and from reef development will be hauled in 30t haul trucks directly to an ore stockpile on surface. Ore will be crushed down to a minus 300mm product. The crushed product will be loaded via surface LHD into a reload hopper discharging onto the overland conveyor, leading to the process plant ore bin on surface.

The conveyor system feeding the plant silo directly via the crusher has full VSD control and is capable of slowing down feed belts if the crusher is overloaded. Underground transfer belts feeding the main belt, fitted with belt scales, are fed by VSD controlled feeders. The feeders draw from the underground silos, to ensure fast adjustment of feed to the main belts.

Once the South Mine decline conveyor system is commissioned, ore produced on or below the 21 Level will continue to be hauled by truck to one ore bin on the 21 Level, which will feed onto the South Mine Decline conveyor, which will convey it to the surface ore bin. Ore produced above the 21 Level will continue to be hauled directly to surface by haul trucks as a result of the safety-based decision made not to perform loaded down-hill hauling.

Switch-over from Truck Hauling to Conveying Ore and Waste

While the North and South Mine decline conveyor system is being installed and commissioned, it will not be possible for haul trucks to use the left-side decline as a travel route to surface. During this period, all underground vehicles will have to operate in the right-side decline. This switchover period is conservatively estimated at 12 months.

Transportation of Underground Workers

Until a chairlift is established in each of the respective areas, workers will be transported via trackless personnel vehicles from designated taxi stops.

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The majority of underground workers will be moved between surface and their workplaces by chairlifts installed in the conveyor decline parallel to the haulage decline. Workers will walk from where they get off the chairlifts to the stopes in the strike haulages. No truck haulage will take place during major shift movements.

Chairlifts convey at up to 900 persons per hour, and will allow a complete shift-change to be accomplished in less than 2 hours at the greatest distance of stopes from surface. Maximum walking distance will be generally 1 000m, although this may be greater in a small number of locations in the mine.

Some parts of the deposit are small and isolated and at greater walking distances than 1 000m. It will not be economic to install chairlifts to access these, and trackless personnel vehicles will be used in these cases, typically of the cassette personnel carrier type.

Transportation of Supervisory and Technical Personnel Underground

Supervisors and technical personnel will have convenient and rapid access to underground operations by light diesel vehicles throughout the LoM.

Transportation of Materials and Supplies Underground

Consumable supplies such as ground support materials will be delivered underground by diesel supply vehicles. Supply vehicle type will predominantly be 8t cassette carrier diesel trucks having the capability of offloading cassettes containing supplies and equipment at the destination point within the mine, and self-loading empty cassettes for removal to surface. A lesser amount of supply vehicles will be the flat-bed type, with or without a crane for self-loading.

To the greatest extent possible, supply materials will be bundled, palletized on standard wooden pallets, or supplied in standard bulk bags, for ease of movement on and off flat-bed vehicles and in cassettes. Stoping ground support supplies will be assembled on surface and delivered in unit loads. Unit loads, which will contain the correct mix of the various materials and supplies required by an operating raise-line.

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Ground support materials cassettes will be off-loaded in cross-cut storage bays (timber bays) from where materials will be taken from the cassette as required and moved to the raise-lines and stopes by mono-line winches.

Transportation and Storage of Explosives Underground

Explosives for Access Development and Footwall Development will be of the emulsion type with electric initiation. Emulsion explosives will be delivered by the manufacturer to bulk silos on surface. From there it will be loaded into special explosives vehicles for delivery on a daily basis to underground development headings. Blast holes will be charged pneumatically. Packaged explosive will be transported as palletized unit loads and offloaded direct from the delivery truck at an explosives receiving area on surface from where it is loaded into cassettes and delivered to underground explosive stores.

18.3 Power and Energy Infrastructure

Grid power will be required at the WBJV Project 1 (Maseve) mining site. It is estimated that the decline systems envisaged will require about 28MVA maintaining full production. It is expected that 16MVA will be needed for the concentrator and the remainder for the mining and other surface infrastructure.

The only authorized supplier of bulk power in South Africa is the parastatal organisation, Eskom. South Africa has recently been subjected to shortages of bulk power due to a number of reasons. The WBJV project team is assuming that bulk power will be available from Eskom. However as a result of this situation, back up diesel generators for self-generation of electric power have been included into the overall power infrastructure design. There is one 2.25MVA diesel generator installed at the site. The second 2.25MVA generator is currently being installed and will be online by September 2015.

The following power supplies are currently installed:

>	1.5MVA Eskom Temporary Power,
>	2.25MVA diesel generated power,
>	2 x 10MVA Eskom grid supply.

The following supplies are planned for installation in the short term:

> 2.25MVA additional diesel generator power (October 2015),

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Once completed the mine will have a 20MVA total supply. The following supplies are planned for installation in the short to medium term:

>	Eskom is currently constructing Ngwedi substation to firm up the regional transmission network. (March 2016),
>	2 x 40MVA Eskom grid supplies will be phased into the Maseve supply and the 10MVA supplies removed. This will allow Maseve to use a total of 40MVA, if required, with a firm supply. (July 2016).

The final system will be setup as follows, the full mine supply will be fed via a firm Eskom supply in entirety. The generator system will be setup with the following capabilities. Both generators operating will be capable to supplying 3.6MVA prime power. In the case for power being lost to the mine the generators will automatically start and using a power management system and will supply power to critical areas e.g. ventilation, thickeners, dewatering pumps, chairlifts, etc. When power is restored from the Eskom side, the generator will automatically synchronise with the Eskom supply and engage. The power management system will automatically re-engage the non-critical circuits and the generators will shut down.

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In the case that the mine is requested to reduce load, the generators will be capable to load clipping up to 3.6MVA prime power to the mine, at a maximum demand of 28MVA this equates to 13% of load. Generally in the past mines have been requested to shed 10% of their demand for certain periods while the grid is under pressure to meet demand. This will ensure PTM Maseve is able to shed 13% or upward of its immediate demand without impact to the operation.

Grid power will be required at the WBJV Project 1 (Maseve) mining site. The decline system envisaged will require about 12MVA maintaining full production. It is expected that 16MVA will be needed for the concentrator and the remainder for the mining and other surface infrastructure.

18.4 Fuel Storage

The bulk fuel storage facility is located close to the North Shaft facility between the North Shaft parking area and Impofu substation. The generator units are located on the west side of the bulk storage tanks to ensure a large supply of fuel is available during unforeseen power outages or load clipping. The delivery point is located off the North Shaft main access route to ensure that bulk fuel delivery vehicles do not enter the operational areas unnecessarily.

The total fuel storage being installed in phase one is 120 000l of bulk fuel storage and 5 000l of generator day tank storage. Civil bases and mechanical links are available to increase the bulk storage capacity by 60 000l totalling 180 000l. The system is designed to drain and fill the generator day tank on a weekly basis to ensure that Diesel does not stagnate in the day tank which can lead to issues starting up the generators.

The diesel fleet for the mine is shown below with the estimated diesel consumption for the operation. With the current fleet the expected diesel consumption is 174 045l per month. The future fleets expected consumption is 155 803l per month. At the current storage capacity this equates to 28 and 31 calendar days of storage respectively. When upgraded to 180 000l of bulk fuel storage this equates to 31 and 35 calendar days of storage respectively.

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Table 18-1: Current Diesel Fleet – Diesel Consumption

	Equipment	Type	Payload	Rate per hour	Rate per month	Eng Kw	Current Fleet Size	Operating hours	BSFC @ 80% - L/h	Shift Consumption - L	Daily Consumption - L	Monthly Consumption - L
Current Fleet Comsumption	Cat AD30	Truck	30T	122	45 390	304	13	5.93	43	1 051	2 102	48 341
	Sandvik EJC533	Truck	30T	122	45 390	298	3	5.93	42	238	475	10 936
	Cat 1600G	LHD	10T	90	28 178	200	4	5.31	28	578	1 155	26 567
	Sandvik LH410m	LHD	10T	90	28 178	204	1	5.31	29	147	295	6 775
	Cat 1700G	LHD	12.5T	108	33 834	263	2	5.31	37	316	633	14 548
	Sandvik LH514	LHD	14T	124	38 744	256	1	5.31	36	134	269	6 183
	Sandvik DD320-26	Drill Rig - Twin Boom		41	14 541	74	1	1.00	10	10	21	479
	Sandvik DD321-40	Drill Rig - Twin Boom		41	14 541	110	9	1.00	15	139	278	6 404
	Sandvik D310	Drill Rig - Single Boom		20	8 156	62	1	1.00	9	9	17	401
	Maclean Scissor Bolter	Bolter		41	14 541	110	1	1.00	15	15	31	712
	Fermel UV	UV			40 000	85	7	3.19	12	267	533	12 268
	Fermel RoRo	UV			40 000	85	9	3.19	12	343	686	15 774
	Fermel Grader	Grader			170 000	85	1	5.93	12	71	142	3 262
	Cat 120K LP	Grader			170 000	93	1	5.93	13	78	155	3 569
	Fermel Tmech	Transporter				85	1	3.00	12	36	72	1 650
	Kubota	Transporter				18.5	14	3.00	3	109	219	5 026
	Cat Bobcat	Construction				53.8	2	5.31	8	80	161	3 698
	Toyota Landcruiser	Transporter				96	4	3.00	14	162	324	7 452
	Total								349	3 784	7 567	174 045

Table 18-2: Future Diesel Fleet – Diesel Consumption

	Equipment	Type	Payload	Rate per hour	Rate per month	Eng Kw	Future Fleet Size	Operating hours	BSFC @ 0% - L/h	Shift Consumption - L	Daily Consumption - L	Monthly Consumption - L
Future Fleet Comsumption	Cat AD30	Truck	30T	122	45 390	304	-	5.93	43	-	-	-
	Sandvik EJC533	Truck	30T	122	45 390	298	11	5.93	42	1 268	2 536	58 323
	Cat 1600G	LHD	10T	90	28 178	200	-	5.31	28	-	-	-
	Sandvik LH410m	LHD	10T	90	28 178	204	-	5.31	29	-	-	-
	Sandvik LH208/9L	LP LHD	10T	90	28 178	170	21	5.31	24	754	1 508	34 691
	Cat 1700G	LHD	12.5T	108	33 834	263	-	5.31	37	-	-	-
	Sandvik LH514	LHD	14T	124	38 744	256	6	5.31	36	249	499	11 471
	Sandvik DD320-26	Drill Rig - Twin Boom		41	14 541	74	5	1.00	10	52	104	2 393
	Sandvik DD321-40	Drill Rig - Twin Boom		41	14 541	110	-	1.00	15	-	-	-
	Sandvik D310	Drill Rig - Single Boom		20	8 156	62	-	1.00	9	-	-	-
	Sandvik DD211L	LP Drill Rig - Single Boom		20	8 156	55	9	1.00	8	73	147	3 371
	Maclean Scissor Bolter	Bolter		41	14 541	110	7	1.00	15	108	217	4 981
	Sandvik DS210L	LP Bolter		41	14 541	74	9	1.00	10	99	197	4 536
	Fermel UV	UV		-	40 000	85	6	3.19	12	237	475	10 918
	Fermel RoRo	UV		-	40 000	85	6	3.19	12	237	475	10 918
	Fermel Grader	Grader		-	170 000	85	1	5.93	12	71	142	3 262
	Cat 120K LP	Grader		-	170 000	93	-	5.93	13	-	-	-
	Fermel Tmech	Transporter		-	-	85	-	3.00	12	-	-	-
	Kubota	Transporter		-	-	18.5	9	3.00	3	74	148	3 393
	Cat Bobcat	Construction		-	-	53.8	-	5.31	8	-	-	-
	Toyota Landcruiser	Transporter		-	-	96	4	3.00	14	164	328	7 545
	Total								391	3 387	6 774	155 803

18.5 Water Supply Pipelines and Services

The Bulk Water supply will be sourced from Magalies Water Board, a government institution supplying bulk-water to the regional communities. The source of bulk water is the Vaalkop dam, some 60km by road from the WBJV site. Magalies Water has stated to PTM that they have enough water from the dam to supply the Mine requirements, but not the infrastructure to deliver the water to the mine since all available delivery capacity for water supply is fully utilized. Thus a new higher capacity pipeline system for mine supply was necessary to supply water to the WBJV site as well as several other mining sites in the vicinity. The route of the pipeline mainly follows existing municipal supply pipeline corridors.

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To deliver water from Magalies Water at Vaalkop Dam, additional power for pumping will be required from Eskom. Whilst Magalies Water is of the opinion that the power will be supplied, the country power risk cannot be ignored, as detailed in the Power Report.

A temporary water supply was installed for construction purposes capable of delivering approximately 0.6Ml per day the permanent supply line from Magalies Water will be capable of delivering 6Ml/day. The water consumption for the total mine during steady state operations has been calculated at 2.5Ml/day based which is in line with a generally accepted rule of thumb of approximately 1kL water per tonne of rock mined. This is confirmed by the water balance equation.

The November 2012 agreement between Magalies Water and Maseve to supply bulk water for mining purposes via pipeline and appurtenant works to be built stipulates that the Mining Ventures, including Maseve, to be supplied with bulk water, fund, construct and commission the approved pipeline design. At the effective date of this report the water pipeline infrastructure to supply water to the WBJV is complete.

18.5.1 Underground

Industrial-quality water will be distributed by steel pipelines throughout the underground workings for drilling equipment, dust suppression, and cleaning of areas. Flexible hoses will be used to connect water pipelines to drilling equipment at working faces. A potable water tank located on surface will provide an uninterrupted supply of fresh water and fire-fighting water. Water pressure of potable and industrial water will be controlled by pressure reducing stations strategically placed underground to ensure water pressure is controlled and can be used by personnel and machinery respectively underground.

Water inflows to the underground mine derive from groundwater inflows, drilling water, HPE equipment (hydro mining), drinking water, potable water used for firefighting and dust suppression water. Mine water on strike haulages and cross-cuts will be directed by gravity to pump niches located at no more than 600m intervals, from where it will be pumped horizontally to a dirty water pumping station. Dirty water pumping stations will be located on every second strike haulage, providing a 120m vertical interval from one dirty water pump station to the next. Mine water will be pumped progressively from one pump station to the next until arrival at surface. Main water pipeline can dewater at no less than 250m³ per hour.

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Ultimately, a clean water pumping station will be installed on the 45 Level, allowing more efficient and cost effective clean-water pumping to be performed from 45 Level to surface. A cascade settling system, using muck bays excavated at 9°, will be used to remove a large portion of suspended solids that can be dug out using an LHD when required.

18.6 Tailings Storage Facility

The tailings dam is located to the east of the project site and to the south of the drainage channel. As such, the access from the concentrator to the TSF will be directly by a service road under the ESKOM power lines.

The facility was originally designed for 25 million tonnes over an area of 75 hectares. The TSF will be capable of holding 22 million tonnes over an area of 70 hectares with a total enclosed area of about 120 hectares, including return water dams.

> A three phase construction approach has been adopted with the fourth phase being an operational approach.

• Phase 1:

o Construction completion by September 2015

o Capacity 80 000 tonnes.

o Basin area +-70 000m2.

o Wall elevation 1061m.

• Phase 2:

o Construction completion by May 2016.

o Capacity 1 520 000 tonnes.

o Basin area +-300 000m2.

o Wall elevation 1059m.

• Phase 3:

o Construction start not determined yet.

o Capacity 560 000 tonnes.

o Basin area +-250 000m2.

o Wall elevation 1059m.

• Phase 4:

o No construction requirements, only operations.

o Capacity 22 160 000 ton.

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o Basin area +-700 000m2.

o Final tailings elevation 1083m.

18.7 Fire Suppression Infrastructure

A comprehensive fire suppression strategy was implemented across the mine site to ensure interconnectivity of fire suppression system information and management. This approach provided a complete system and a guideline for current infrastructure and new infrastructure in the future. The fire suppression pump stations have been constructed at each of the major infrastructure sites, namely North Shaft, Concentrator and South Shaft. Water is provided by a dedicated portion of the potable water tanks and each pump station is fitted with a primary electrical pump and a secondary diesel pump if power is not available. Pressure reducing stations regulated water pressure at all points on surface and underground.

Fire hydrants are placed around the mine where water is not suitable for fire extinguishing or fire water pipes are not fitted. Electrical switchgear and electrical motor control centres are fitted with dry power canisters inside the panels to automatically deploy if a fire or arc is detected.

18.7.1 Standards

The following items comply with the relevant National Fire Protection Association – America (“NFPA”) and manufacturers’ standards for the application:

>	Linear heat cable will be installed between the main and return belts on conveyors with the Infra-Red detection system and air purge,
>	Foam water sprinklers deluge systems,
>	Bund foam pourer protection for bunds, including medium and high expansion foam systems,
>	Medium velocity water spray fixed systems and high velocity water spray fixed systems for fire protection.

The following items comply with the relevant South African Nation Standards (“SANS”).

>	Fire hose reels and piping sizing,
>	Signage for fire protection equipment,
>	Hydrants,
>	Fire detection installation,
>	All electrical work.

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The fire pumps, fire tanks, piping and valve configuration shall conform to Automatic Sprinkler Inspection Bureau (“ASIB”) 11^th Edition sprinkler rule requirements.

18.7.2 General Requirements

>	All fire hose reels, and hydrants shall have signage indicating location,
>	Fire hose cupboards containing two 30m lengths of hose with instantaneous couplings and an A.W.G. (Specific European Manufacturer) style nozzle shall be located at every hydrant,
>	Detection shall activate the audible and visual local alarms at each area of installation,
>	All fire protection systems and medium and high velocity water spray systems, shall be capable of being activated manually and remotely,
>	All openings between risks such as cable entries and exits into substations shall be fire stopped,
>	All substation exits shall be fitted with manual call points and needs to be protected by smoke detection systems,
>	MCC substation shall be protected by fire alarm systems in accordance with the applicable SANS Standard,
>	All electrical equipment shown outside of buildings or within wet or spray protected areas shall be Weather Proof (IP 65),
>	Suitable lightning protection shall be implemented to mitigate the lighting risks posed to the equipment. The system will be properly earthed to the applicable earthing standards.

18.8 First Aid Facilities

A small first aid building has been constructed at the main entrance of the Maseve site, from this facility minor injuries and general ailments can be dealt with. It has been designed to be a mobilisation point for medical assistance and an emergency room is available within the building to stabilise a patient before transport. To ensure the facility is comprehensive, it has been constructed with a waiting area, consultation room, toilets, kitchen and rest area for staff. The proto team mobilisation point is located behind the clinic ensuring that if both the medical and proto team need to be mobilised all planning from a rescue and a medical treatment point of view can be executed from a single point.

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18.9 Security Infrastructure

Due to security concerns within the South African environment, multiple levels of security have been implemented within and around the mine site.

The first level is hard barricading using industrial razor wire topped with razor coil around the outer perimeter of the mine site, diamond mesh fencing is installed around secure areas within the site and small areas outside of the main site, topped with barbed wire where deemed necessary, to demarcate road 5 strand barb wire fencing is used. Manned boom gates are used for vehicle access and turnstiles or magnetic lock door are used for regular personnel access.

The second level is security presence and security guards are posed around the site in fixed locations as well as roaming security. The permanent guard houses and security buildings provide deployment facilities for security personnel around the site.

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The third level is access control which also includes time and attendance. The G4S system has been deployed around the site providing biometric and card reader access control. This system is amalgamated with training and medical validation to ensure that personnel are not allowed into work without valid certificates.

The fourth level is non evasive security systems.

18.10 Communication Systems

18.10.1 Surface

Surface communications are divided into four major networks

> Telecommunication Network:

• Direct Connection,

• External Service providers.

>	IT Network,
>	Control Network,
>	Radio Network.

Telecommunications network is an external communications supplier connecting the operation to telephony and data connections through a secure firewall on site. Cellular phone network providers provide an alternative communications across the operation to persons. The IT network is distributed around the site using a 24 core fibre network connecting major centres, training centre, North Shaft, Concentrator and South Shaft, of the mine to a single network. The control network is a 24 core fibre network connecting the operations SCADA systems to all areas of the operation. A licensed radio network is available over majority of the site as an additional communications system.

The site is directly connected to the corporate offices in Rosebank via a 25Mbps microwave link supplied by Vodacom.

18.10.2 Underground

A leaky feeder communication system is installed as the communication system for the mine. Telephones will be located at key infrastructure locations such as chairlift stations, electrical substations, conveyor drives, refuge areas, lunchrooms, and pumping stations.

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Key personnel, (such as Mobile Mechanics, Crew Leaders and Shift Bosses), and Mobile Equipment Operators, (such as loader, truck, and utility vehicle operators), will be supplied with underground radios for contact with the leaky feeder network.

A fire network is established underground consisting of process network control and communications systems to surface. The fibre network is designed using single mode technology allowing for long transmission distances. Systems allowed for on the fire network will include. Standby electrical protection network, pump station control, access and security control, tracking and tagging control, sub decline chairlift SCADA, environmental systems, bulk materials handing control, paging systems, fire system monitoring and other system capacity for further expansion.

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19.0 MARKET STUDIES AND CONTRACTS

19.1 Market Review and Metal Prices

Concentrates produced by the WBJV Maseve Mine will be sold to Rustenburg Platinum Mines, a subsidiary of Anglo Platinum Limited. As such the WBJV will not be sourcing ‘off takers’ to purchase end products. The emphasis applied in this marketing report is therefore on future PGM supply and demand scenarios and potential impacts on commodity prices.

The major products contributing to revenue are platinum, palladium and rhodium. Minor revenue contributors are gold, copper, nickel, ruthenium and iridium.

Commodity prices are influenced by the levels of supply and demand which are considered in detail for each commodity. Major economic parameters including the rand dollar exchange rate are also considered.

The WBJV study is complicated by the long time frames of the mine life involved. Generally short term price forecasts cover the ensuing 12 months with long term forecasts being for 2 to 3 years into the future. The WBJV Technical Report requires forecasts some 10 to 15 years into the future. Changes in supply and demand can and will be significant in such a time frame making accurate forecasting very difficult if not impossible. One has only to consider the effect of technological changes and the influence that they will have on demand. On the other hand supply could be impacted by the shuttering of high cost operations and successful geological exploration in new areas.

Global demand for PGM metals has been relatively stable. Consumption from the automobile sector, primarily in the form of catalytic converters, has reflected general growth in global automobile consumption. Jewellery demand has been driven by the growth in the Chinese market while general industrial consumption has reflected global GDP growth.

On the supply side the South African PGM mining industry has historically accounted for the bulk of global platinum supply. Recent events including cost escalation, labour unrest and metal price weakness has made many of these operations marginal to loss making. Supply destruction is occurring and will likely have an impact of metal prices at some point in the future. Recycled metal is also playing a larger role in the PGM supply story.

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The only logical approach is to use available knowledge on mine developments in projecting supply figures and available knowledge on technological research and economic development in different countries in projecting demand.

The Company is listed on the NYSE-MKT exchange and requires that economic studies consider trailing average prices over a 3 year period. These prices have been calculated and updated.

In summary the metal prices considered in this Technical Report are summarised in the table below.

Table 19-1: 3 Year Average Trailing Metal Prices

3 year Average Trailing Metal Prices as at 15 June 2015
Platinum	$     1 408	USD/ounce
Palladium	$        744	USD/ounce
Rhodium	$     1 126	USD/ounce
Gold	$     1 374	USD/ounce
Iridium	$        731	USD/ounce
Ruthenium	$          73	USD/ounce
Copper	$       3.18	USD/ pound
Nickel	$       7.11	USD/ pound
ZAR/USD	R     12.00

19.2 Material Contracts

19.2.1 Concentrate Sale Agreement

The off-take agreement for 100% of produced metals in concentrates for the LoM has been signed between Maseve Investments 11 (Proprietary) Limited and Rustenburg Platinum Mines Limited, a subsidiary of Anglo-Platinum Limited, dated 23^rd April 2013 and to come into effect upon the delivery of first concentrate from Maseve to Rustenburg Mines smelter complex. The payable metals are; Pt, Pd, Rh, Au, Ir, Ru, Ni and Cu. The contract is within the industry norms.

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The financial model for the consideration of resources and reserves has considered the confidential terms of the off-take agreement in place for the project. The contract is within the industry norms.

During the first year of production a number of considerations are included in the agreement and model to allow Maseve to stabilise and optimise the metallurgical plant.

Other industry normal contracts are in place with normal rates:

19.2.2 Mining Contract

JIC Mining Services, subsidiary of Westdawn Investments (Pty) Ltd, is contracted to perform the mining activities (waste and reef development, stope production, rock haulage and engineering support).

The mining contractor scope includes:

>	The provision of all labour required to perform the mining activities, inclusive of all supervisory, production, engineering, procurement and human resource management,
>	The procurement and supply of all materials required to perform the mining activities,
>	Provision of all spare parts and operating cost of all TMM equipment,
>	Haulage of broken rock to surface or to underground conveyor installations,
>	Overhead costs associated with the contractor’s on-site management, facilities, administrative and head office costs.

19.2.3 EPCM Contract

DRA Projects SA (Pty) Ltd is contracted to perform Engineering, Procurement and Construction Management (EPCM) for the construction of the Maseve concentrator, surface infrastructure and permanent underground installations as reported in Section 18.0 of this report.

19.2.4 ESKOM Contract

ESKOM, the national electricity provider in South Africa, has provided the project a formal budget quotation and undertaking for the conversion of the currently installed 20MVA (on a 88kV/11kV feed network) to 40MVA (on a 132kV/11kV network) at the Impofu substation.

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The document is dated 17 August 2015 and has been accepted by the project. The target energising date is currently the 3rd quarter of 2016.

19.2.5 Magalies Water

Magalies Water is the regional government water utility management company. The project entered into a bulk water off-take agreement with Magalies Water on 22 November 2012. The agreement commits the project to make a capital contribution to the expansion of the Magalies water scheme that will supply the project and secures bulk water off-take in line with the project’s future requirements.

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20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1 Environmental

20.1.1 Legislative Requirements

Prior to construction and operation of an underground mine, the following local legislative authorisations are required:

>	In support of a Mining Right Application (“MRA”), authorisation in terms of Section 22 of the Mineral and Petroleum Resources Development Act, 2002 (Act No. 28 of 2002) (MPRD Act) by the Department of Mineral Resources (“DMR”),
>	Environmental Authorisation as per the National Environmental Management Act, 1998 (Act No. 107 of 1998) (NEM Act) and EIA Regulations (GNR. 543, 544 and 545 of 18 June 2010,
>	A Water Use License in terms of Section 21 of the National Water Act, 1998 (Act No. 36 of 1998) from the Department of Water and Sanitation (“DWS”),
>	A Waste Management License for categorised waste activities in terms of the National Environmental Management Waste Act, 2008 (Act No. 59 of 2008) (NEMWA) from the National Department of Environmental Affairs (“DEA”).

Environmental regulation of mining in South Africa is currently in a transitional period, as the governing legislation is being transitioned from the MPRDA to the National Environmental Management Act, No. 107 of 1998 (NEMA). The MPRDA has been amended to delete all the environmental regulation provisions. While subject to academic debate, certain of these provisions (in particular the requirement for the DMR to approve an EMP for a mining right or environmental management plan (EMPlan) for a prospecting right) and requirements for compliance with approved EMP’s remain in force until certain pending provisions in the MPRDA and NEMA commence.

Under the prior environmental provisions of the MPRDA, the Company, and Maseve as the holders of mining and prospecting rights, must (a) conduct environmental monitoring on a continuous basis, (b) conduct performance assessments of the EMP or EMPlan as required, (c) conduct an annual financial determination in respect of the financial guarantees filed with the relevant prospecting right and mining right, and (d) compile and submit a performance assessment report to the Minister in which compliance with the EMP or EMPlan is demonstrated.

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20.1.2 Environmental Baseline Studies

Environmental baseline studies that formed part of the EIA include:

>	Soils, Land Use and Land Capability Study,
>	Fauna and Flora Report,
>	Hydrological Study,
>	Groundwater Specialist Report,
>	Air Quality Impact Assessment,
>	Ground Vibration and Air Blast,
>	Visual Impact,
>	Archaeological assessment,
>	Traffic assessment.

20.1.3 Environmental Monitoring

All environmental monitoring at the WBJV Project 1 (Maseve) and 1A site are conducted within the conditions set out in the EMP, EA and Waste Management License. Environmental Compliance Audits are conducted annually.

From 2010 to 2013, independent inspections of compliance with the EMP's were conducted annually over each prospecting right of Maseve in accordance with Regulation 55 of the MPRDA. The Environmental Compliance Report reported that all of the drilling sites and the immediate areas that were affected by prospecting had adequately complied with commitments as recorded in the EMPs. It was further reported that all rehabilitation of prospecting boreholes and related infrastructure was completed and that no further rehabilitation is necessary. Independent rehabilitation certificates were issued for each borehole and are on record at the site offices.

Subsequent Performance Assessment Reports were compiled to measure performance against the approved mining right EMP. Environmental monitoring forms an integral part of the compliance. The following environmental aspects in the subsections below are monitored.

20.1.3.1 Surface Water

The surface water sampling sites comprise natural rivers and streams, as well as development site water control infrastructure including pollution control dams, both North and South settling ponds and service water tanks. Surface water sampling is conducted on a monthly basis.

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20.1.3.2 Ground Water

There are 11 groundwater sampling points covered in the monitoring plan. Some of these have been destroyed during construction and others are currently dry. Six groundwater monitoring points are currently being monitored. The groundwater monitoring programme will be evaluated with new conditions of the WUL and new boreholes will be sited and drilled if necessary.

20.1.3.3 Water Use

As part of the environmental policy, water use is monitored closely. Magalies Water Board water consumption figures are recorded and captured in a data base and updated weekly to calculate the usage of water in the mine.

20.1.3.4 Air Quality

The monitoring of the air quality is currently conducted by Maseve's Environmental Assistant. Gravimetric analysis of the exposed dust fallout buckets for dust deposition rates was completed by an accredited laboratory – Quality Laboratories (Pty) Ltd and sent through to Bateleur Environmental and Monitoring Services for interpretation. Exposed dust fallout buckets are collected every two weeks and the dust filtered through a sub-micronic, pre-weighed filter using a vacuum filter bench. Dust fall out concentrations are measured in mg/m²/day from samples collected at 6 sites around the mining site as stipulated in the approved EMP. Gravel roads are sprayed with water as dust suppression method.

20.1.3.5 Noise

The ambient noise monitoring was conducted by JH Consulting; Acoustics, Noise and Vibration Control Consultants; this is conducted in quarterly intervals. The noise levels at the nearest residential areas and isolated dwellings are unlikely to exceed the day or night-time limit levels for suburban areas.

Ambient noise measurements were carried out according to SANS Code of Practice 10103:2008 at three relevant positions on or near the property boundary and a 4^th position to obtain a representative measurement of road noise on the R565. The noise levels at the nearest residential areas and isolated dwellings are unlikely to exceed the day or night-time limit levels for suburban areas.

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20.1.3.6 Fauna and Flora

Fauna and flora assessments are done once during summer and once during winter. Nyengere Wildlife and Environmental Solutions has been appointed to conduct the fauna and flora monitoring. The main vegetation characteristics show little to no variation when compared to the baseline data recorded for the vegetation classification of the Maseve study area in 2007.

20.1.3.7 Soils

The availability of soil during the rehabilitation phase of the mine is very important, hence the monitoring of the fertility of the soil. Not only will stripped and stockpiled soil be monitored, but also the general quality of the soil in and around the mining site. This will be conducted twice yearly.

20.1.4 Rehabilitation and Closure

In terms of the MPRDA, Regulation 54, the EMP outlines the specifications for the closure and rehabilitation of the mining area after operations cease and requires that a financial guarantee be held. The quantum of pecuniary provision is updated annually and the rehabilitation cost determination is reflected in the financial guarantee required. Calculations are based on the DMR methodology with current disturbed areas surveyed and used in the guideline. The EMP states that for the first year of construction 30% of the estimated rehabilitation costs be posted.

The environmental closure cost assessment is for the entire mining area and involves the quantification of mining and infrastructure components and applying rates to rehabilitate each component. The environmental liability is described in monetary terms in order for a financial provision to be set aside in a dedicated fund for closure and rehabilitation. Bateleur Environmental and Monitoring Services is the consultancy responsible for the determination of the rehabilitation costs and environmental monitoring of the WBJV Project 1 (Maseve) and 1A.

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In April 2012, Maseve posted an environmental rehabilitation guarantee of ZAR58.5 million as a requirement of Maseve’s mining right application. In October 2012, Maseve entered into an agreement with a third party insurer whereby a bond would be posted to the credit of the DMR against the Company’s ZAR58.5 million environmental guarantee for its Mining Right and the Company’s posted guarantee would be released back to the Company. The process was completed in 2013 fiscal year and the posted guarantee has been returned to the Company. As a term of the agreement with the third party insurer, in October 2012, Maseve posted ZAR12 million on deposit with the Standard Bank of South Africa against its environmental guarantee obligation and will make further annual deposits of approximately ZAR12 million per annum until the full amount of the environmental guarantee is returned and the third party arrangement will be wound up or renewed at Maseve’s election. Interest on deposits will accrue to Maseve and Maseve will pay an annual fee of approximately ZAR600 000 to the Insurer.

Project 1 platinum mine’s financial estimates include an accumulated charge per tonne for the costs of eventual closure of the mine.

20.2 Permitting

Applicable permit applications and granting are completed under the Maseve Investments 11 (Pty) Ltd holding company which is recognized by the issuing agencies as the owner of the WBJV Project 1 and 1A described as Portion 7 of the farm Frischgewaagd 96 JQ Portion 14 of the farm Frischewaagd 96 JQ Portion 17 of the farm Frischewaagd 96 JQ Remaining Extent of Portion 17 of the farm Elandsfonttein 102 JQ and Remaining Extent of Portion 2 of the farm Elandsfontein 102 JQ. Approved and issued permits are held under Maseve Investments.

As part of the requirements of the Mining Right the company must annually submit an implementation plan of the Social and Labour Plan to give effect to Regulation 46 (i),(ii) and (iii).

Subsequent to the issue of the Mining Right, the Company applied for land use and environmental permits for WBJV Project 1 (Maseve) and 1A. Maseve holds the following governmental authorizations:

> The EIA and EMP was approved by the DMR in terms of the Mineral and Petroleum Resources Development Act (No 28 of 2002) on the 15 May 2012,

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>	The Department of Economic Development, Environmental, Conservation and Tourism (“DEDECT”) granted to Maseve an Environmental Authorisation in terms of the NEMA to commence with the construction of infrastructure and facilities of the 13 September 2013, ref: 30/5/1/2/3/2/1/528EM,
>	The DEA approved Maseve’s application for Waste Management License in terms of the National Environmental Waste Management Act (No 59 of 2008),
>	Waste Management License 12/9/11/L628/7 (“WML”) issued by the DEA in terms of the National Environmental Management: Waste Act No. 59 of 2008 (NEMWA) for a treatment plant to be utilized during Phase 2 of the Project 1,
>	Water Use License 03/A22F/ABCGIJ/2596 (“WUL”) issued by the Department of Water and Sanitation in terms of Chapter 4 of the National Water Act, 1998 (Act No 36 of 1998) for various water uses on the Mining Right Area.

Delays or a failure to obtain or amend the required permits, or a failure to comply with the conditions of any such licenses that the Company has obtained, could have an adverse material impact on the Company.

20.3 Social and Community Impact

20.3.1 Social and Labour Plan

Maseve Mine‘s Social and Labour Plan was approved as part of the mining right. The commitments of the Social and Labour Plan include Human Resource Development; Mine Economic Development or Local Economic Development and management of downscaling and retrenchment. The Mine commenced with the implementation of the commitments from 2012 after the issuance of the mining right.

20.3.2 Employment and Contracting

Local employment is key under the Social and Labour Plan (“SLP”) and Maseve Mine has always ensured that the company maintained a 30% average of locally sourced employment from the communities that surround the mine area. The Company Human Resource (“HR”) policy is also detailed in the SLP.

Contractor compliance to HR policy of the WBJV Project 1 (Maseve) is part of the SLP. The Mineral and Petroleum Resources Development Act (MPRDA), 2002 (Act 28 of 2002), as well as the Mining Charter as specific requirements are included in any tender process and contract documentation for all contractors. Each contractual tender includes certain criteria as listed in the Act, as well as the Charter. It is therefore expected that signatories will attach the Charter’s provisions to specifications, which will be enforceable through contract compliance. In addition, Contractor is required to comply with the WBJV Project 1 (Maseve) requirements. The primary purpose of the policy is to assure that the spirit of the Mining Charter is addressed by all stakeholders concerned, namely development and empowerment of the local communities, and to protect the interests of the Employer, WBJV Project 1 (Maseve) in terms of their statutory obligations.

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20.3.3 Training

The mine capital development plan includes a significant investment in training through the LoM, allocated to a social and labour plan to ensure maximum value from the project for all stakeholders, including local residents. Based on interaction with the community, the completion of a skill and needs assessment, and the Company’s training plans; the project is planning for 2 700 jobs with a target of at least 30% from the local communities. Maseve Mine implements all the required HRD programme entailed in the approved SLP. The programmes have specific targets and timeframes and include ABET; Learnerships; Internships; Portable skills and bursaries.

20.3.4 Occupational Health and Safety

Regulations for mine operation fall under the Occupational Health and Safety Act, No. 85 of 1993 and the Mine Health and Safety Act, No. 29 of 1996.

Platinum Group Metals Ltd via Maseve is actively engaged with stakeholders, including employees, contractors and regulators, at Project 1 (Maseve). Employees and Contractors benefit from continuous training programs, such as SafeMap, and management interaction that develop and reinforce a culture of safety and performance. The Company has maintains a 20% to 30% local labour participation rate and has a “SLP” in place to deliver programmes focussed on attracting, training and retaining the best local talent. Additionally, the Company officially invites mine inspectors from the DMR, a key mining regulatory agency, to perform regular mine site inspections in an effort to achieve the highest environmental, health and safety standards.

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Collectively, these policies and programmes have allowed Maseve to foster an open, honest and team oriented relationship with its workforce and to have reached a current five million man hours of work at the Project 1 site with a low lost time incidence frequency of 1.1 per million man hours.

20.3.5 Local Communities

Maseve mine sits on the border of two local municipalities – Rustenburg Local Municipality (“RLM”) and Moses Kotane Local Municipality (“MKLM”). According to the 2001 Census, the total population for the RLM is 395 761, while the MKLM has a total population of 237 175. The MKLM is a rural area that is characterized by low accessibility and large distances between settlements. Generally, infrastructure in RLM is of a higher standard than the MKLM.

The 2007 Community Survey reveals that approximately 99% of people living in RLM have access to some sort of toilet facility (either a flush toilet with sewage system or a pit latrine), but only 85% of people in MKLM have access to toilet facilities. Access to piped water is between 30 - 35% for both municipalities, but whereas 60% of people living in RLM have access to government refuse removal, only 8% of people living in MKLM have access to this service.

Road infrastructure also differs between the two municipalities. The MKLM is characterized by poor roads, especially during the rainy season, which further limits the accessibility of services to people living in this municipality. PTM Foundation primarily targets the communities close to Maseve Mine.

20.3.6 Community Engagement

A Community Engagement Policy was implemented in 2011. The key objective of this policy is to put in place an engagement structure, the necessary systems, communication channels and procedures that will enable the most effective management of the interaction with, and the facilitation of the socio-economic development in Mine communities. These objectives are based on legislation and best practice to ensure standardisation across the operation regarding community engagement. The facilitation of sustainable relationships and partnerships with Maseve’s Mine communities is of the utmost importance in order to, not only secure Maseve’s licence to operate, but as well as to gain the communities’ social consent to operate the Mine on a day-to-day basis. Maseve Mine holds quarterly stakeholder engagement meetings with local communities to ensure that they updated on the status of the mine and the progress on the SLP. The company also has a complaint procedure in place. The procedure allows for external communities to lodge any complaint they have against the mine. The complaint gets referred to the relevant section and gets investigated and addressed accordingly.

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Maseve also has a full time Community Liaison Officer (“CLO”) who interacts with the communities on a daily basis.

20.3.7 Education, Health and Sport

The WBJV Project 1 (Maseve) and 1A has a dedicated Sports Health and Education (“SHE”) plan as part of the Corporate Social Investment (“CSI”). As such, Platinum Group Metals (PSA) (Pty) Ltd. (PTM) established the PTM Foundation, an Association not for Gain Company (AGC), in early 2014 to augment PTM’s CSI projects surrounding the WBJV Project 1 (Maseve) platinum mine (Maseve Mine) near Rustenburg, South Africa.

As mandated by the South African government, Maseve Mine created an SLP to implement local economic development and human resource development projects. To complement these projects, PTM resolved to expand its involvement in the local communities through a CSI program oriented at sports development, education, health, and employment skills training for local youths. These Public Benefit Activities (“PBA’s”) are supported by the PTM Foundation and, while consistent with the SLP's development concept, are completely voluntary and will be funded by donations from various sources. All SLP activities as committed to by Maseve are funded by Maseve Mine and under no circumstances will donations be diverted from the Foundation to fund Maseve’s own SLP commitments or other corporate social responsibilities.

The PTM Foundation seeks to develop leadership skills in youths through its sports development and education programmes. The Foundation’s local community development projects are currently focussed on:

20.3.7.1 Sports Development

In 2012, Maseve Mine began funding Sundown Ranch Sport School to create a sports development programme for the local schools around Maseve Mine. During its first year, the Sports Academy programme ran in eight local schools. The following year, the programme expanded to cover 16 local schools.

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In 2014, plans began to bring the programme under PTM Foundation and re-brand the programme as “Sports for Success.” The Sports for Success programme has been very successful introducing all learners to basic physical education. A competitive sports league would focus more on talent identification and increase the chances of the local learners qualifying for university scholarships. All the coaches and current management of the programme will be paid by the Foundation. Once the programme is completely under the Foundation, Maseve Mine will begin funding the programme through donations to the Foundation.

20.3.7.2 Education

Maseve Mine plays a vital role in the development of local schools both in infrastructure and school programmes. Through the SLP the mine refurbished the science laboratories of two high schools. The mine and its contractors also made a contribution towards the grade 12 additional classes for local schools. The aim of the programme was to assist in improving the Maths and Science subjects to ensure access to the scarce and critical skills. The mine also engages schools in terms of career guidance to assist learners to choose skills which guarantee employment and development. The other programme aims to bring more knowledge about mining and its related fields to local learners. For the programme’s initial year in 2015, Grade Three learners in local community schools were the only grade targeted. This gave the Foundation a chance to run the programme with learners who are old enough to focus on the workshop activities, but still young enough that the technical learning of the workshops could be minimal. The overall goal of the programme is to give young learners an opportunity to learn about mining through fun, hands-on and interactive workshops. Learners should be able to come away from the workshops with the idea that mining is interesting and fun.

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20.3.7.3 Health Awareness

In the past, Maseve Mine has hosted health awareness campaigns at its local community events, such as sport tournaments. At these events, Royal Bafokeng’s health team setup a booth and gives presentations and distributes health information pamphlets about health concerns common in the local communities. Previous health campaigns have focused on TB, HIV/AIDS, and non-communicable diseases, such as cardiovascular disease and diabetes. There has been a significant focus on prevention in most of the campaigns with information shared about safer sexual practices and nutrition.

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21.0 CAPITAL AND OPERATING COSTS

21.1 Capital Cost

The capital cost estimate has been developed from mining and infrastructure designs as developed by DRA and costed, based on quotations, the cost of recently completed work at the project and on the existing contracts for mine and infrastructure development.

The project capital cost is anticipated to be R5 070 million to the start of production, inclusive of R4 404m sunk to the date of this report. An approximate additional R708 million will be spent on capital to peak funding and beyond to achieve peak production of 165,000tpm by mid-2018. LoM Capital expenditure including sustaining capex is estimated at R6,418m.

Table 21-1: Capital Cost Summary

	ZAR ‘million
Sunk to July 2015	4 505
Capital Expenditure to Production	265
Capitalised Operating Cost to Production	300
Total Capital Expenditure to Production	5 070
Remaining LoM Capital Excl. Sustaining Capital	860
Sustaining Capital LoM	487
Total LoM Capital Expenditure	6 418

21.1.1 Sunk Capital Cost

Sunken cost to July 2015 of R4 505 million represents the cost of:

>	establishing the project infrastructure reported as complete in Section 18.0,
>	underground development reported as current workings in Section 16.0,
>	trackless mining machinery reported as current complement in Section 16.7.14,
>	other underground equipment and construction work to date,
>	construction of the Maseve concentrator that is approximately 90% complete at the effective date of this report,
>	capitalised operating cost related mainly to owners cost and specialised services as part of the construction activities, and
>	purchasing of land holdings.

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21.1.2 Capital Expenditure to Production

It is estimated that a further R265 million will be spent on capital to 1^st concentrate production. This cost relates mainly to the completion of the Maseve concentrator and the continuing work on surface infrastructure and underground construction and equipment.

21.1.3 Capitalised Operating Cost

R300 million is to be spent on cost of an operating nature to 1^st concentrate production. About R205 million of this cost relates to underground development, stope establishment and related equipment cost. The remainder of the cost includes support services, shared services, utility costs and pre-cost for the concentrator commissioning and start-up.

21.1.4 Remaining LoM Capital Expenditure (excluding sustaining capital)

It is estimated that approximately R860 million will be spent on capital post the date of 1^st concentrate production for the remainder of LoM. The main categories of spend are:

> Underground mining equipment (R378 million) comprising of:

• Conveyor and chairlift expansion installations,

• Electrical installations,

• Civil work,

• Pumping stations,

• Equipping of ventilation shafts, and

• Equipment requirements for the establishment production areas.

> Conversion of the current MF1 concentrator to a MF2 circuit (R189 million)

• The addition of an identical mill, mill motor and associated electrical and mechanical components to the front-end of the current concentrator. The civil work for the additional mill was completed during the construction of the current concentrator.

• The addition of flotation cells to the current concentrator circuit. All civil and steelwork for the positioning of the additional float cells was completed during the construction of the current concentrator.

• Minor modifications to piping, electrical and instrumentation components on the current concentrator.

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> Tailings Storage Facility (R145 million)

• As reported in Section 18.6 the tailings storage facility will be constructed in 3 phases. This mount is the required expenditure for the ongoing construction of the facility post 1^st concentrate production.

> Other (R148 million)

• Completion of commitments to Eskom and Magalies Water for energy and water projects,

• Other remaining surface infrastructure,

• Project indirects, and

• Other minor contractual closures.

21.1.5 LoM Sustaining Capital

An allowance of approximately R487 million is made in the LoM plan for sustaining capital. Sustaining capital will mostly be applied to the rebuilding of TMM equipment and major component replacements in the concentrator during the production years of the mine. On average, about R33 million per annum is estimated to be spent on this.

21.1.6 Capital Expenditure Profile

The graph below shows the capital spending profile over time measured against key milestones during the LoM.

>	1st concentrate production is planned to occur by end October 2015.
>	165ktpm concentrator capacity and associated mining production is planned for mid- 2018.

21.1.7 Capital Expenditure Escalation

The base date for the capital estimate is 30 June 2015. All costs are real and no-escalation factors have been applied in the capital cost model. No allowances were made for fluctuations in exchange rates.

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Graph 21-1: Capital Expenditure Profile

21.2 Operating Cost

Operating costs are estimated based on a combination of the current mining contract in place, previously established estimates based on quotations, new quotations, current and historic costs of services (both internal and external) and, where required, estimates where done using zero-based budgeting.

The operating cost estimate is based on the mining schedules reported in Section 16.7 of this report, concentrator requirements to process the mined ore and required support services.

The mine operating model used for operating cost estimates is that of “Owner Managed and Contractor Operated” as opposed to “Owner Managed and Operated”. Underground mining activities, trackless machinery maintenance and the operation of the concentrator are all currently planned to be performed by contracted organisations. Some support services such as security services, operating of change houses, cleaning services, survey, rock engineering and human resource training and administration are also planned to be performed by contracted organisations.

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The Company is planned to have a limited number of full time employees that will manage the operations and perform functions such as financial management, operating of permanent underground infrastructure, maintenance of surface infrastructure not related to the concentrator and all safety management.

21.2.1 On-mine Operating Cost Summary

The table below summarises the on-mine operating cost per area.

Table 21-2: Operating Cost Summary

	LoM Cost	R/tonne
LoM OPERATING COST	R 'million	mine/milled
Mining Cost	22 773	702*
Contractor Cost	15 227	470
Salary Cost	1 026	32
TMM Maintenance Contractor	2 300	71
TMM Machine Rental Cost	2 051	63
Contingency	461	14
Support Services	1 059	33
Sampling Cost	85	3
Utility Cost	565	17
Concentrator Cost	3 754	116
Salaries	735	23
Fixed Cost	215	7
Variable Cost	1 604	49
Utility Cost	1 200	37
Shared Services and Other	2 618	80
Salaries	1 210	37
Other	1 407	43
Total	29 145	898*

*Equivalent R/tonne milled is R896/tonne milled. Mining cost is shown in R/tonne mined.

The explanation and basis of estimate of each area are discussed below.

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21.2.2	Mining Operating Cost (R22 733 million)
21.2.2.1	Contractor Cost (R15 227 million)

Mining activities (waste and reef development, stope production, rock haulage and engineering support) are planned to be performed by a mining contractor. The operating cost estimate is based on rates contained in a current mining contract in use at the property.

The mining contractor scope includes:

>	The provision of all labour required to perform the mining activities, inclusive of all supervisory, production, engineering, procurement and human resource management,
>	The procurement and supply of all materials required to perform the mining activities,
>	Provision of all spare parts and operating cost of all TMM equipment,
>	Haulage of broken rock to surface or to underground conveyor installations,
>	Overhead costs associated with the contractors’ on-site management, facilities, administrative and head office costs.

21.2.2.2 Salary Cost (R1 026 million)

Salary cost in the mining area consists of:

>	Six (6) management level employees in mining who will manage and oversee the work of the mining contractors.
>	Two hundred and fifteen (215) engineering employees performing the following functions:
	•	Engineering management,
	•	Planned maintenance planning,
	•	Maintenance of surface infrastructure excluding the concentrator facilities, and
	•	Operation and maintenance of the permanent underground infrastructure, e.g. conveyor belts, ore handling facilities, electrical installations and water pumping installations.

21.2.2.3 TMM Maintenance Contractors (R2 300 million)

All trackless mobile machinery maintenance is planned to be done by contractors. The cost of this function is for labour only and is based on recent quotations received from original equipment manufacturers and contractors currently involved in the maintenance of equipment at the project. The estimate is further based on the labour requirements to service the TMM fleet shown in Section 16.7.

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21.2.2.4 TMM Machine Rental Cost (R2 051 million)

It is planned that all future trackless mobile machines will be procured on a rental basis. Rentals will be for periods of 3 to 4 years, whereafter the fleet will be replaced and new rentals entered into. The estimated amount is based on proposals received from rental companies in the South African market and is further based on the TMM fleet requirements shown in Section 16.7 of this report.

21.2.2.5 Contingency (R461 million)

An allowance of R461 million is made to cover incidental costs over the LoM not specifically covered in the mining area. These cost would cover inter alia items such as unforeseen poor ground conditions, water ingress and specialized work.

21.2.2.6 Support Services (R1 059 million)

Support services include the following:

>	Mining consultancy fees (R44.3 million)
>	Contracted rock engineering services (R60.5 million)
>	Contracted survey services (R69.8 million)
>	Contracted ventilation services & equipment, rescue team equipment and minor ventilation construction (R93.6 million)
>	Underground diamond (exploration) drilling (R388.0 million)
>	Geological consultancy fees (R39.9 million)
>	Engineering consumables associated with the maintenance activities mentioned in Section 21.2.2.2 above (R151.7 million)
>	Health and safety consumables (R24.8 million)
>	Contracted ambulance and paramedic services (R42.1 million)
>	Contracted medical surveillance and clinic services (R144.1 million)

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21.2.2.7 Sampling Cost (R85 million)

Sampling cost is based on an estimated cost of R2.62 per tonne mined. This is based on approximately 575 samples per month in mining block definition (±200), grade control (±300) and RoM checks (±75).

21.2.2.8 Utility Cost (R565 million)

21.2.2.8.1 Electricity (R494 million)

It is estimated that the mining operations, inclusive of surface installations, will consume a total of 19.683 kWh/t milled at an effective cost of 77.2c (Rand cents) per kWh giving a cost of R15.20 / tonne milled.

21.2.2.8.2 Water (R71 million)

It is estimated that the mining operations will consume 0.677m³ per tonne milled. Water is priced at R3.20 / m³, as per the Magalies water off-take agreement, giving a cost of R2.17 / tonne milled.

21.2.3 Concentrator Operating Cost (R3 754 million)

21.2.3.1 Salaries (R735 million)

Salary cost is determined on staffing the concentrator with 126 operational and engineering personnel. The intent is to contract these services from a reputable concentrator operating company. Initial proposals received from a company is comparable to the planned cost.

21.2.3.2 Fixed Cost (R215 million)

Fixed costs are based on the production profile and the GRD Minproc cost estimate from the 2009 study. Updated costs have been applied and the cost estimate is thus based on first principles. Fixed cost is estimated at about R896 000 per month.

21.2.3.3 Variable Cost (R1 604 million)

Variable costs are based on the production profile and the GRD Minproc cost estimate from the 2009 study. Updated costs have been applied to the consumables and the cost estimate is thus based on first principles. Variable cost is estimated at about R49 / tonne milled. Costs included are:

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>	Variable Stores/Consumables at R42/tonne milled (R1 355 million),
>	Stockpile handling (R54 million),
>	Contracted services for tailings disposal oversight (R44 million),
>	Sample assays and sample transport to Rustenburg (R74 million),
>	Concentrate transport (R77 million),
	•	An estimated 876kt of concentrate (including 15% moisture) will be transported 50km to the smelter at R1.76 tonne.km.

21.2.3.4 Utility Cost (R1 200 million)

21.2.3.4.1 Electricity (R1 120 million)

It is estimated that the concentrator will consume a total of 44.594 kWh/t milled at an effective cost of 77.2c (Rand cents) per kWh giving a cost of R34.43 / tonne milled.

21.2.3.4.2 Water (R80 million)

It is estimated that the mining operations will consume 0.770m³ per tonne milled. Water is priced at R3.20 / m³, as per the Magalies water off-take agreement, giving a cost of R2.46 / tonne milled.

21.2.4	Shared Services and Other Operating Cost (R2 618 million)
21.2.4.1	Salary Cost (R1 210 million)

103 personnel will be employed in areas not directly associated with either the mining or concentrator activities, or whose services will be shared.

>	Management and heads of department (10),
>	Financial and administration, inclusive of procurement and warehousing (33),
>	Human Resources (6),
>	Technical services (35),
>	Safety (18),
>	Security (1).

21.2.4.2 Other (R1 407 million)

R1.4 billion is the estimated cost for all services not directly associated with mining or concentrator activities. This includes:

	>	Materials to perform shared services functions,
.	>	Consultancy cost,

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>	Services and material to execute the Social and Labour Plan, Local Economic Development and the Environmental Management Program,
>	Legal costs,
>	Administrative and statutory costs,
>	Operating of shared infrastructure, e.g. change houses and other facilities,
>	Training cost,
>	Information technology,
>	Waste management,
>	General office running cost.

21.2.5 Co- and By- Product Credits

By-& Co-product credits consist of the net revenue derived from Ru, Ir, Ni and Cu as per the Concentrate Offtake Agreement.

21.2.6 Operating Cost Summary – Inclusive of Off-mine Costs

The table below summarises the total net cash cost of the project based on the LoM plan and the main operating cost components.

Table 21-3: Total Net On Site Cost

ZAR:USD = 12.00
	ZAR per tonne milled	USD per tonne milled
On Mine Operating Cost	896	75
By- & Co-Product Credits	(106)	(9)
Total Net On Site Cost	791	66

	ZAR per 4E oz in concentrate	USD per 4E oz in concentrate
On Mine Operating Cost	8 392	699
By- & Co-Product Credits	(989)	(82)
Total Net On Site Cost	7 403	617

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Table 21-4: Total Net On Site Cost per Reef Type

ZAR:USD = 12.00
	ZAR per tonne milled	USD per tonne milled
Merensky	796	66
UG2	778	65
Total Net On Site Cost	791	66

	ZAR per 4E oz in concentrate	USD per 4E oz in concentrate
Merensky	6 312	526
UG2	9 288	774
Total Net On Site Cost	7 403	617

21.2.1 Escalation and Inflation for Operational Cost

The base date for the operating costs estimate is 30 June 2015. All costs are real and no-escalation factors have been applied in the operating cost model. Operating costs are primarily sourced from South African suppliers and the impact of foreign suppliers is limited to the mechanised plant spares. No allowances were made for fluctuations in exchange rates.

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22.0 ECONOMIC ANALYSIS

The Maseve Project 1 financial model has been based on a number of inputs from a number of different persons and organisations, namely:

>	Platinum Group Metals (RSA) and Maseve management,
>	Turnberry Projects – reviewed concentrator opex, metallurgical performance with recovery and grade model and market studies,
>	DRA Mining – project implementation, mining infrastructure assistance and costing – EPCM Contractor managing implementation – detailed mine design and mining schedules,
>	DRA Minerals – concentrator – provided comparison concentrator capital.

Each of these persons or organisations had input into the financial model in some aspect, either for capital or operating costs. These experts, within their specific field, reported to one of the project QP’s for review of the inputs into the production or financial models.

The financial model is calculated on a monthly schedule and stope by stope basis, keeping Merensky and UG2 revenue analysis separate due to different metal ratio splits and recoveries between the two.

The financial model does not include price escalation nor does it include inflation.

22.1 Production Profile

The production profile used in the financial model is as per Section Error! Reference source not found. of this report.

22.2 Revenue and Revenue Assumptions

The following prices, based on a 3 year trailing average in accordance with U.S. Securities and Exchange Commission ("SEC") guidance, was used for the assessment of Resources and Reserves.

Table 22-1: Metal Prices used in Financial Model

Metal	Price	Unit
Platinum	1 408	USD/oz
Palladium	744	USD/oz

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Metal	Price	Unit
Rhodium	1 126	USD/oz
Gold	1 374	USD/oz
Copper	3.18	USD/lb
Nickel	7.11	USD/lb
Iridium	731	USD/oz
Ruthenium	73	USD/oz

The exchange rate between the ZAR and the USD is fixed at ZAR12.00:USD1.00 in the financial model throughout the LoM. The pricing and exchange rates above results in the estimated basket prices shown in the table below.

Table 22-2: Basket Prices

Basket Prices per 4E	ZAR/4E Oz	ZAR/4E kg	USD/4E Oz	USD/4E Kg
Merensky Reef	14 590	469 093	1 216	39 091
UG Reef	14 487	465 776	1 207	38 815
Combined in LoM	14 553	467 902	1 213	38 992

Revenue is recognised in the financial model net of smelter returns as per the Concentrate off-take agreement.

Table 22-3: Revenue net of Offtake Agreement per Reef Type

	ZAR per tonne milled	USD per tonne milled
Merensky	1 533	128
UG2	988	82
Total	1 284	107

	ZAR per 4E oz in	USD per 4E oz in
	concentrate	concentrate
Merensky	12 151	1 013
UG2	11 783	982
Total	12 019	1 002

Metallurgical recovery are assumed at the same rates as the 2009 updated feasibility from the project at 87.5% on the Merensky Reef and 82.5% on the UG2 Reef.

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Average PGM recovery in the financial model, inclusive of the initial 33 months of production in the concentrator MF1 circuit, is 86.6% on Merensky.

The graph below shows the resultant annual 4E ounces produced in concentrate.

Graph 22-1: 4E Ounces Produced in Concentrate

22.3 Annual Cash Flow Results

The graph below shows the cash flow results from the financial model.

The graph shows capital cost, operating cost, smelter pipeline financing cost and royalties & taxes as total cash outflows. Revenue shown is net revenue receipts as per the Concentrate Offtake Agreement. Net cash flow is net revenue above less all the cost items.

Pipeline financing is the interest charges levied in terms of the Concentrate Offtake Agreement against payment advances made on concentrate sales ahead of the settlement dates contained in the agreement. For purposes of the financial model, it is assumed that the option of taking payment advances will be exercised.

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Graph 22-2: Annual Cash Flow Results

22.4 Net Present Value and Internal Rate of Return

The economic evaluation is summarised in the table below.

Table 22-4: Economic Evaluation Summary

	NPV @ 5%		NPV @ 10%
Economic Evaluation Summary	ZAR million	USD million	ZAR million	USD million	IRR
Cash flow since start of project	1 597	133	(368)	(31)	8.8%
Cash flow from July 2015 (reporting date)	6 340	528	4 226	352	49.8%

The IRR of the project is based on estimated cash flows from July 2015 onwards is 49.8% .

The peak funding requirement of the project is R5 625 million, inclusive of all sunken cost.

22.5 Taxes and Royalties

The company tax rate in South Africa is 28% on normal income. Under current legislation, all historic capital expenditure on the project will be allowed as tax deductions in future.

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Royalties are South African government royalties and are calculated in terms of the Mineral and Petroleum Resources Royalty Act 28 of 2008 – refer to Section 4.5.1.

All value-added tax (VAT) raised on invoices from suppliers are claimed as input VAT from The South African Revenue Services (SARS) and refunded by SARS.

22.6 Sensitivities

The project sensitivity has been conducted for metal price, head grade, capital cost and operating cost, with variations of plus and minus 10% and 20% in each input. The data is presented in graphs and tabled in the tables below for post-tax cash flows.

Graph 22-3: IRR Sensitivity from Project Start

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Graph 22-4: IRR Sensitivity from July 2015 Onwards

Table 22-5: Sensitivity Analysis (From Start of Project)

SENSITIVITY ANALYSIS (Start of Project)
POST-TAX

	Change in	Change in	Change in	Change in	Change in
Parameter	Parameter	Parameter	Parameter	Parameter	Parameter
Metal Prices	-20%	-10%	0%	10%	20%
IRR (post-tax)	-0.7%	4.8%	8.8%	12.1%	15.1%
NPV (5% Discount) (R'000)	-R 1 856 627	-R 89 108	R 1 597 013	R 3 243 520	R 4 915 795
NPV (10% Discount) (R'000)	-R 2 554 609	-R 1 431 621	-R 367 563	R 664 101	R 1 705 762
Head Grade	-20%	-10%	0%	10%	20%
IRR (post-tax)	-6.8%	3.6%	8.8%	12.4%	15.3%
NPV (5% Discount) (R'000)	-R 2 805 680	-R 510 288	R 1 597 013	R 3 451 028	R 5 086 124
NPV (10% Discount) (R'000)	-R 3 103 425	-R 1 678 547	-R 367 563	R 786 317	R 1 798 989
Capex	-20%	-10%	0%	10%	20%
IRR (post-tax)	9.3%	9.0%	8.8%	8.5%	8.2%
NPV (5% Discount) (R'000)	R 1 812 091	R 1 704 552	R 1 597 013	R 1 489 474	R 1 381 936
NPV (10% Discount) (R'000)	-R 207 648	-R 287 605	-R 367 563	-R 447 521	-R 527 479
Opex	-20%	-10%	0%	10%	20%
IRR (post-tax)	13.2%	11.1%	8.8%	6.1%	3.0%
NPV (5% Discount) (R'000)	3 806 973	2 694 516	R 1 597 013	R 454 655	-R 723 468
NPV (10% Discount) (R'000)	1 030 041	326 383	-R 367 563	-R 1 092 669	-R 1 841 200

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Table 22-6: Sensitivity Analysis (July 2015 Onwards)

SENSITIVITY ANALYSIS (July 2015 Onwards)
POST-TAX

	Change in	Change in	Change in	Change in	Change in
Parameter	Parameter	Parameter	Parameter	Parameter	Parameter
Metal Prices	-20%	-10%	0%	10%	20%
IRR (post-tax)	22.7%	36.1%	49.8%	64.5%	80.4%
NPV (5% Discount) (R'000)	R 2 342 235	R 4 388 360	R 6 340 256	R 8 246 293	R 10 182 160
NPV (10% Discount) (R'000)	R 1 315 159	R 2 809 857	R 4 226 118	R 5 599 263	R 6 985 714
Head Grade	-20%	-10%	0%	10%	20%
IRR (post-tax)	16.6%	33.7%	49.8%	65.5%	80.8%
NPV (5% Discount) (R'000)	R 1 243 588	R 3 900 791	R 6 340 256	R 8 486 510	R 10 379 338
NPV (10% Discount) (R'000)	R 584 686	R 2 481 199	R 4 226 118	R 5 761 932	R 7 109 799
Capex	-20%	-10%	0%	10%	20%
IRR (post-tax)	58.2%	53.7%	49.8%	46.4%	43.3%
NPV (5% Discount) (R'000)	R 6 673 707	R 6 506 981	R 6 340 256	R 6 173 531	R 6 006 805
NPV (10% Discount) (R'000)	R 4 523 437	R 4 374 777	R 4 226 118	R 4 077 458	R 3 928 799
Opex	-20%	-10%	0%	10%	20%
IRR (post-tax)	75.8%	61.7%	49.8%	39.3%	30.0%
NPV (5% Discount) (R'000)	8 898 561	7 610 753	R 6 340 256	R 5 017 833	R 3 654 010
NPV (10% Discount) (R'000)	6 086 329	5 149 761	R 4 226 118	R 3 261 001	R 2 264 707

It is noted that the project is most sensitive to metal prices, head grades and to a lesser extent operating cost. The project is not significantly sensitive to capital cost as approximately 70% of LoM capital is sunken cost at the effective date of this report.

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23.0	ADJACENT PROPERTIES
23.1	Public Domain Information about Adjacent Properties

23.1.1 Producing Mine

The BRPM Joint Venture consists of the Bafokeng Rasimone Platinum Mine (‘BRPM’), which has been in operation since 1998, a concentrator and the Styldrift I project currently under construction. The BRPM operation lies directly to the South of the Project Area and operating stopes are essentially along strike and within 1 500m of the WBJV mine area. BRPM is an operational mine and the additional information is published in Royal Bafokeng Platinum 2014 Annual Report, which can be accessed on the www.bafokengplatinum.co.za website.

Current Operations:

>	Mining - Two decline shaft complex’s, conventional mining activities, 500m average mining depth,
>	Processing - Conventional mill/float concentrator − MF2 configuration,
>	LoM 35 years including MR and UG2,
>	MR in LoM 74% (up to 2028) and 30% for LoM with balance UG2.

Key Production Statistics

The Royal Bafokeng Platinum website includes the following points for their BRPM operations (Annual Report, 2014):

>	Annual 4E production 2014: 294 000 ounces
	o	Forecast till 2017: 280 000 ounces per annum.
	o	Forecast for LoM: 207 000 ounces per annum.
>	Concentrator Milling 2014: 2 400 000 tonnes
	o	Forecast till 2018: 2 400 000 tonnes per annum
	o	Forecast for LoM: 1 900 000 tonnes per annum excluding new project
>	Concentrator Recovery (4E) 2014: 87.0%
	o	Forecast MR: 87.9%
	o	Forecast UG2: 81.9%
>	Mill grade (4E) 2014: 4.29g/t
	o	Forecast LoM: 4.16g/t
>	Annual 4E total cash cost (4E) 2014: ZAR8 040 per ounce
	o	Forecast LoM: ZAR9 051 per ounce (2014 real)
>	Cost per tonne milled 2014: ZAR957

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o Forecast LoM: ZAR1 036 (2014 real)  

23.1.2 Developing Mines

Down-dip and to the east of the Maseve mine is BRPM Project Styldrift I, which is an underground mine, lies about 5km from the North shaft of BRPM and is located within the boundaries of the Styldrift and Frischgewaagd farms. In the 2014 annual report the project was expected to commence with stoping ramp-up in the third quarter of 2015 and to reach steady state in the first quarter of 2019 but in a press release dated 4 August 2015, construction activities have been reduced and production will be delayed. At steady state the mine production will be 230 000 tonnes a month.

The adjacent property to the North of the Maseve mine is Wesizwe Platinum Limited (“Wesizwe”). The Pilanesberg Project’s Bafokeng Platinum Mine (“BPM”) of Wesizwe is situated on the farms Frischgewaagd 96JQ, Ledig 909JQ, Mimosa 81JQ and Zandrivierpoort 210JP.

Wesizwe’s website includes the following points for their BPM operations (Annual Report, 2014):
BPM is in the development phase of a new underground mine to access the Merensky and UG2 PGM ore bodies. Wesizwe also has a 17.1% interest in neighbouring Projects 1 and 3 of Maseve Investments 11 (Pty) Limited (Maseve), held through a subsidiary company, Africa Wide Mineral Prospecting and Exploration (Pty) Limited. Maseve is operated by Platinum Group Metals Limited (PTM) Canada.

23.2 Source of Adjacent Property Information

Data from adjacent properties is obtained from publicly available documents published by the operating company of the adjacent properties. The BRPM operations information is available on website www.bafokengplatinum.co.za and the Wesizwe Platinum Limited information is accessible on website www.wesizwe.co.za.

23.3 Relevance of the Adjacent Property Information

The WBJV deposit is a continuation of the deposit concerned in the BRPM operations and the Wesizwe Project, and the information obtained from BRPM and Wesizwe is thus of major significance and appropriate in making decisions about the WBJV Project 1 (Maseve).

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The technical information on adjoining properties has been sourced from public domain information and has not been verified by the QP of this report.

23.4 Application of the Adjacent Property Information

BRPM and Styldrift information from AP was verified and used for geological context in the estimation of Mineral Resources. However, the details of this information cannot be disclosed as a confidentiality agreement exists between PTM and AP.

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24.0	OTHER RELEVANT DATA AND INFORMATION
24.1	Construction and Development

A substantial portion of the initial underground development and surface construction work has been already been completed therefore those costs are already known and the production schedule has been updated to correspond to timing of the completed work. At the effective date the main construction milestones remaining on the critical path to production are:

>	TSF,
>	Plant testing and commissioning,
>	Access development to blocks 10, 11 and 16,
>	Ventilation infrastructure,
>	Operational readiness:
	•	Mining labour build-up,
	•	TMM equipment availability.

Items completed as of effective date (from annual report)

>	Mill foundations and civil works for the primary mill,
>	Water infrastructure for construction and development,
>	Mill steel and bearings in place and mill delivered and lifted into position,
>	Structural steel for mill and flotation circuits complete,
>	Stands and civil works for the conveyor from North Mine to mills completed,
>	North change house for 1 200 persons complete,
>	South change house for 500 persons complete,
>	Electrical substation and connections for North and South Mines in place,
>	Pollution control dams completed,
>	Underground development along the MR and first underground raises,
>	North Mine development to five blocks advancing as planned,
>	Chairlift and conveyor in the North Mine commenced,
>	Ore silo on surface and conveyor transfer points completed,
>	Equipment maintenance shop completed,
>	Two raise bore holes developed and fans installed for ventilation,
>	South Mine declines advanced to the first turn and 150m from ore,
>	Space for underground ore silos and underground workshop completed,
>	Computer access control system installed and operational,

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>   Security control centre and mill control buildings completed. 

24.2 International Reporting Standards

The SAMREC Code sets out minimum standards, recommendations and guidelines for public reporting of Mineral Resources and Mineral Reserves in South Africa, some of which are outlined below.

Documentation for public release must be prepared by or under the direction of, and signed by, a Competent / Qualified Person. A Competent Person (CP) (equivalent to a QP) is a person who is a member of the South African Council for Natural Scientific Professions (SACNASP) or the Engineering Council of South Africa (ECSA) or any other statutory South African or international body that is recognised by the SAMREC Code. A CP should have a minimum of five years of experience relevant to the style of mineralisation and type of deposit under consideration.

A Mineral Resource is a concentration (or occurrence) of material of economic interest in or on the earth’s crust in such form, quality and quantity that there are, in the opinion of the CP, reasonable and realistic prospects for eventual economic extraction.

The definitions of the Measured, Indicated and Inferred Mineral Resources can be found under Item 14.2 of this document.

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25.0	INTERPRETATION AND CONCLUSIONS
25.1	Geology and Estimation

25.1.1 Results

The Mineral Resource tabulations for the Maseve Project demonstrated an increase in the tonnage of the Measured and Indicated categories from the 2009 estimates, with all areas previously classified as Inferred upgraded to higher confidence levels.

25.1.2 Interpretation of the Geological Model

The stratigraphy of the Project Areas is well understood and specific stratigraphic units could be identified in the drill hole core. The MR and UG2 were recognised in the core, and these correlate across the Project Areas. It was possible to interpret major structural features from the drill hole intersections as well as from geophysical information.

IRUP bodies can cause disruption of the stratiform mineralization by replacement, physical displacement and redistribution of the PGE. They have lower grades and recoveries than the unaffected surrounding lithologies. Implications of the IRUPs for mining and processing imply that they cannot be included in the Mineral Resource estimates declared by CJM Consulting.

25.1.3 Estimation Technique

The estimation of the Project was undertaken using best geostatistical practices. Ordinary kriging was selected as the best estimate for the specific drill hole distribution.

25.1.4 Reliability of the Data

The PTM data was specifically inspected by the QP and found to be reliable and consistent.

25.1.5 Strengths and Weaknesses with Respect to the Data

The regular QA and QC process carried out by PTM is of a high standard and applies to the full audit trail from field data to Resource modelling. The data have been found to be accurate, consistent and well structured. The system of support for the digital data by paper originals and chain of custody and drilling records is well developed.

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25.1.6 Objectives of Adherence to the Scope of Study

The intention of this phase of the work programme was to update the Mineral Resources for Project Areas 1 and 1A based on new diamond drill holes for Mineral Resource estimation and several technical drill holes. The Updated resource and geological model were used to update the Mine plan and financial models as described in this technical report.

25.2 Mining Methods

The geological and structural models in conjunction with geotechnical considerations formed the basis of the mining methods that was chosen. The mine design is based on the UG2 geological block model received on 12 May 2015 and the MR Geological Block model received on 13 May 2015. The mine design criteria and any assumptions takes cognisance of the information supplied by PTM.

The mining methods selected present the best practical outcome under the given conditions. The selected methods had to be versatile and easily interchangeable with each other with the lowest impact on production during conversion of a mining section, yielding the lowest degree of grade dilution. They would also have to be able to integrate with the trackless environment that would supply access to all of the mining areas.

25.3 Metallurgical Processing

The concentrator plant has been constructed in the MF1 configuration and will be commissioned in the near future. The MF2 plant configuration is to be implemented and commissioned by 2018 to improve throughput, concentrate grade and overall recovery – this implementation programme is to be reviewed during the next 2 years, as the production profile from the mine is better understood.

25.4 Economic Outcome

The IRR of the project is based on estimated cash flows from July 2015 onwards is 49.8%, with a peak funding requirement of ZAR5 625m. The project is most sensitive to metal prices, head grades and to a lesser extent operating cost.

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26.0	RECOMMENDATIONS
26.1	Objectives to be Achieved in Future Work Programmes

The Projects 1 and 1A are at the Feasibility and Development Level and thus have had sufficient exploration completed that any material exploration activities and engineering studies have for the most part been concluded for these areas. The project is recommended to move to operations.

26.2 Detailed Future Work Programmes

Due to the advanced nature of Projects 1 and 1A, the QP for the Mineral Resource Estimate does not expect a material change in Resources with further exploration work.

26.3 Declaration by QP with respect to the Project’s Warranting Further Work

The project based on the forward financial model is recommended to proceed to operations provided the metal prices, exchange rate, geological, mining, environmental and socioeconomic risks are acknowledged and accepted.

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27.0 REFERENCES

Assibey-Bonsu, W. and Krige, D.G. (1999). Use of Direct and Indirect Distributions of Selective Mining Units for estimation of Recoverable Resources/Reserves for new Mining Projects. Proc. APCOM 1999, Colorado, USA.

Bredenkamp, G. and Van Rooyen, N. (1996). Clay thorn bushveld. In: Low AB and Rebelo AG (1996) Vegetation of South Africa, Lesotho and Swaziland. Department of Environmental Affairs and Tourism, Pretoria.

Cawthorn, R.G. (1996). Re-evaluation of magma composition and processes in the uppermost Critical Zone of the Bushveld Complex. Mineralogy Magazine 60, pp. 131–148.

Cawthorn, R.G. (1999). The platinum and palladium resources of the Bushveld Complex. South African Journal of Science 95, November/December 1999, pp. 481-489.

Leeb-Du Toit, A. (1986). The Impala Platinum Mines. Mineral Deposits of South Africa, Volume 2, pp. 1091–1106. Edited by Anhaeusser, CR and Maske, S.

Maxwell GeoServices, (October 2009), QA/QC Report Done for Western Bushveld Joint Venture (Project 1, Project 1A and WVJ Decline), available from PTM.
PTM, July 2008,

Rutherford, M.C. and Westfall, R.H. (1994). Biomes of southern Africa: an objective categorization. National Botanical Institute, Pretoria.

SAMREC (2007). South African Code for Reporting of Exploration Results, Mineral Resources and Mineral Reserves. Available from/at: www.samcode.co.za

Schürmann, L.W. (1993). The Geochemistry and Petrology of the upper Critical Zone of the Boshoek Section of the Western Bushveld Complex. Bulletin 113 of the Geological Survey of South Africa, Pretoria.

Schürmann, L.W. (1993). The Geochemistry and Petrology of the upper Critical Zone of the Boshoek Section of the Western Bushveld Complex. Bulletin 113 of the Geological Survey of South Africa, Pretoria.

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Smit, P.J. and Maree, B.D. (1966). Densities of South African rocks for the interpretation of gravity anomalies. Bulletin 48 of the Geological Survey of South Africa, Pretoria.

Vermaak, C.F. (1995). The Platinum-Group Metals – a global perspective. Mintek, Randburg, pp. 247.

Viljoen, M.J. and Hieber, R. (1986). The Rustenburg section of the Rustenburg Platinum Mines Limited, with reference to the MR. Mineral Deposits of South Africa, Volume 2, pp. 1107–1134. Edited by Anhaeusser, CR and Maske, S.

Viljoen, M.J. (1999). The nature and origin of the MR of the western Bushveld Complex, based on geological facies and geophysical data. South African Journal of Geology, 102, pp. 221–239.

Wagner, P.A. (1926). The preliminary report on the platinum deposits in the south-eastern portion of the Rustenburg district, Transvaal. Memoirs of the Geological Survey of South Africa, 24, pp. 37.

Young, D. (2005). Competent Persons’ Report on the Exploration Assets held by Wesizwe Platinum Limited.

Preliminary Assessment – Scoping Study Report and Resource Update – Western Bushveld Joint Venture – Elandsfontein Project (Project 1) (2005)

Pre-Feasibility Study – Western Bushveld Joint Venture – Project 1 (2006)

Feasibility Study (Tecnhical Report) - Western Bushveld Joint Venture – Project 1 (Elandsfontein & Frischgewaagd) (2008)

Updated Technical Report (Updated Feasibility Study) Western Bushveld Joint Venture Project 1 (Elandsfontein and Frischgewaagd) (2009)

Maseve Drilling Report [Snowden Mining Industry Consultants (Pty) Ltd]. Report No 130614_D_JR###-##-####_PTM_Maseve drilling Report.docx

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28.0	DATE AND SIGNATURE PAGE
28.1	Certificates

Certificate of Qualified Person

As an author of the technical report titled “An Independent Technical Report on the Maseve Project (WBJV Project areas 1 and 1A) located on the Western Limb of the Bushveld Igneous Complex, South Africa” dated effective 15 July, 2015 (the “Report”) I, Gert Christoffel Roets, Pr Eng. residing at Johannesburg, South Africa hereby certify:

1.	I am a senior mining consultant with the firm of DRA Projects of DRA Mineral Park, 3 Inyanga Close Sunninghill, 2157, Johannesburg, South Africa
2.	I am a practising mining consultant and a registered professional engineer with the Engineering Council of South Africa (ECSA). I am also an associate member of the Association of mine managers of South Africa (AMMSA).
3.	I am a graduate of the University of Pretoria, and have a B.Eng in Mining Engineering (1990).
	I obtained my mine managers certificate of competence in 1994.
4.	I have practiced my profession continuously since 1990. I have more than twenty five years of mining experience including,
	a.	1990 – 1995, Mine planning and mining production
	b.	1995 – 1999, Various positions of mine production management (unit manager, production manager and assistant manager mining).
	c.	1999 – 2005, Various management positions (including directorship) in a mine contracting environment which included capital infrastructure projects and toll mining.
	d.	2005 – 2010, Managing director, mine contracting (mine development and toll mining).
	e.	2010 – Present, Senior Mining Consultant, DRA Projects, RSA
5.	I have read the definition of Qualified Person set out in National Instrument 43-101 (the “Instrument”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfil the requirements to be a Qualified Person for the purposes of the Instrument.
6.	I have not had prior involvement, with the property that is the subject of the Technical Report.
7.	I have visited the Maseve Project for personal inspection on various occasions since 2012.
8.	I am the co-author of this report and am responsible for reviewing Sections 15, 16 and 18 and jointly responsible for sections 1, 2, 3, 21, 24, 25, 26 and 27.
9.	I am not aware of any material fact or material change with respect to the subject matter of the Report, which is not reflected in the Report, the omission of which would make the Report misleading.
10.	I am independent of Platinum Group Metals Ltd. pursuant to section 1.5 of the Instrument.

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11.	I have read the Instrument and Form 43-101F1 (the “Form) and confirm that the parts of the Report that I am responsible for have been prepared in compliance with the Instrument and the Form.
12.	I do not have nor do I expect to receive a direct or indirect interest in the mineral properties of Platinum Group Metals Ltd. and I do not beneficially own, directly or indirectly, any securities of Platinum Group Metals Ltd. or any associate or affiliate of such company.
13.	At the effective date of the Report, to the best of my knowledge, information and belief, the parts of the Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Report not misleading.
14.	Dated at Johannesburg, South Africa, on 28 August, 2015

 

Gert C Roets Pr. Eng

B.Eng ( Mining) University of Pretoria

Registered professional Engineer with the Engineering Council of South Africa

(ECSA)

Mine Manager Certificate of competence

Associate of the Mine Manager Association of South Africa.

Senior Mining Consultant, DRA Projects, RSA

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Certificate of Qualified Person

As an author of the technical report titled “An Independent Technical Report on the Maseve Project (WBJV Project areas 1 and 1A) located on the Western Limb of the Bushveld Igneous Complex, South Africa” dated effective July 15, 2015 (the “Report”), I, Gordon Ian Cunningham, B. Eng. (Chemical), Pr. Eng., FSAIMM, residing at Sixth Avenue, Melville, South Africa do hereby certify that:

1.	I am currently employed as a Director and Principal Metallurgical Engineer by:
		Turnberry Projects (Pty) Ltd. PO Box 2199 Rivonia, Sandton 2128 South Africa
2.	I graduated from the University of Queensland (B. Eng. (Chemical) (1975)).
3.	I am a member in good standing of the Engineering Council of South Africa and am registered as a Professional Engineer – Registration No. 920082. I am a Fellow in good standing of the South African Institute of Mining and Metallurgy – Membership No. 19584.
4.	I have worked as a Metallurgist in production for a total of 20 years since my graduation from university. I have worked as a Consulting Metallurgist for 5 years since graduation and I have been working for Turnberry Projects for 15 years as a Project and Principal Engineer and Director, primarily associated with mining and metallurgical projects.
5.	I have read the definition of “qualified person” set out in National Instrument 43-101 (the “Instrument”) and certify that by reason of my education, affiliation to a professional association and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purposes of the Instrument.
6.	I visited the property, viewed the core, project development and more recently the process plant during construction on numerous occasions during the period June 2005 to August 2015.
7.	I am co-author of the Report and am responsible for Items 13, 17, 19, 20, 22 and 23 of the Report plus sections of Items 1, 21, 24 and 25. I have relied upon outside sources of information used in the completion of these Items from both Platinum Group Metals (RSA) (Pty) Ltd. and DRA Projects. I am aware of no reason to believe the dataset is not accurate and reliable.
8.	I am not aware of any material fact or material change with respect to the subject matter of the Report that is not reflected in the Report, the omission of which would make the Report misleading.
9.	I am independent of Platinum Group Metals Ltd. pursuant to Section 1.5 of NI 43-101.
10.	I am familiar with the specific type of deposit found in the property area and its metallurgical aspects and have been involved in similar evaluations and technical compilations.

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11.	I have read the Instrument and Form 43-101F1 (the “Form”) and confirm that the parts of the Report that I am responsible for have been prepared in compliance with the Instrument and the Form.
12.	I do not have nor do I expect to receive a direct or indirect interest in the mineral properties of Platinum Group Metals Ltd. and I do not beneficially own, directly or indirectly, any securities of Platinum Group Metals Ltd. or any associate or affiliate of such company.
13.	At the effective date of the Report, to the best of my knowledge, information and belief, the parts of the Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Report not misleading.

	Dated the 28th August 2015
Gordon Ian Cunningham
B Eng. (Chemical), Pr. Eng. FSAIMM

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Certificate of Qualified Person

As an author of the technical report titled “An Independent Technical Report on the Maseve Project (WBJV Project areas 1 and 1A) located on the Western Limb of the Bushveld Igneous Complex, South Africa” dated effective 15 July, 2015 (the “Report”) I, Charles Johannes Muller, Pr. Sci. Nat residing at Roodepoort, South Africa hereby certify:

1.	I am a senior Mineral Resource consultant with the firm of CJM Consulting of Ruimsig Office Estate, 199 Hole-in-one Road, Ruimsig, Roodepoort, South Africa.
2.	I am a practising Mineral Resource consultant and a registered professional scientist with the South African Council for Natural Scientific professions, Pr. Sci. Nat. Reg. No. 400201/04.
3.	I graduated with a B.Sc. (Geology) degree from the Rand Afrikaans University - Johannesburg in 1988. In addition I have obtained a B.Sc. Hons (Geology) from Rand Afrikaans University in 1994 and attended courses in geostatistics through the University of the Witwatersrand. I have worked as a Mineral Resource geologist for 22 years since my graduation from university.
4.	I have read the definition of Qualified Person set out in National Instrument 43-101 (the “Instrument”) and certify that by virtue of my education, affiliation to a professional association and past relevant work experience, I fulfil the requirements to be a Qualified Person for the purposes of the Instrument.
5.	I have not had prior involvement, with the property that is the subject of the Technical Report.
6.	I have visited the Maseve Project for personal inspection on various occasions since 2012.
7.	I am the co-author of this report and am responsible for Sections 4, 5, 6, 7, 8, 9, 10, 11, 12 and 14, and jointly responsible for sections 1, 2, 3, 25, 26 and 27.
8.	I am not aware of any material fact or material change with respect to the subject matter of the Report, which is not reflected in the Report, the omission of which would make the Report misleading.
9.	I am independent of Platinum Group Metals Ltd. pursuant to section 1.5 of the Instrument.
10.	I have read the Instrument and Form 43-101F1 (the “Form) and confirm that the parts of the Report that I am responsible for have been prepared in compliance with the Instrument and the Form.
11.	I do not have nor do I expect to receive a direct or indirect interest in the mineral properties of Platinum Group Metals Ltd. and I do not beneficially own, directly or indirectly, any securities of Platinum Group Metals Ltd. or any associate or affiliate of such company.

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12.	At the effective date of the Report, to the best of my knowledge, information and belief, the parts of the Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the Report not misleading.
13.	Dated at Roodepoort, South Africa, on 28 August, 2015

 

Charles J Muller, Pr. Sci. Nat.

Director – Mineral Resources

CJM Consulting (Pty) Ltd

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APPENDICES
stats appendix.pdf
qaqc appendix.pdf

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Bureau Veritas Coarse duplicates

Pt

Good correlation - with no bias shown in the q-q plots- ranked hard plots show that 85% of the data is within 10% difference.

	ORIG_PT	DUP_PT
Mean	0.587894737	0.524342105
Standard Error	0.104254902	0.087260054
Median	0.03	0.03
Mode	0.005	0.005
Standard Deviation	2.03230218	1.7010116
Sample Variance	4.130252152	2.893440463
Cov	0.289275258	0.308253104
Kurtosis	49.56229564	50.44642963
Skewness	6.486125497	6.373126248
Range	20.095	17.895
Minimum	0.005	0.005
Maximum	20.1	17.9
Sum	223.4	199.25
Count	380	380
Confidence Level (95.0%)	0.204990469	0.171574468

Bureau Veritas Coarse duplicates

Pd

Good correlation - with no bias shown in the q-q plots- ranked hard plots show that 90% of the data is within 10% difference.

	ORIG_PD	DUP_PD
Mean	0.237986842	0.234026316
Standard Error	0.036082775	0.036372397
Median	0.01	0.01
Mode	0.005	0.005
Standard Deviation	0.70338277	0.709028544
Sample Variance	0.494747322	0.502721477
Cov	0.338346135	0.330066141
Kurtosis	26.10975539	27.81681515
Skewness	4.785294204	4.942430624
Range	5.405	5.745
Minimum	0.005	0.005
Maximum	5.41	5.75
Sum	90.435	88.93
Count	380	380
Confidence Level (95.0%)	0.070947502	0.07151697

Bureau Veritas Coarse duplicates

Au

Poor correlation - with distinct low bias in original samples shown in the q-q plots- ranked hard plots show that 90% of the data is within 10% difference.

	ORIG_AU	DUP_AU
Mean	0.049552632	0.058763158
Standard Error	0.004851942	0.00934366
Median	0.01	0.01
Mode	0.01	0.01
Standard Deviation	0.094581762	0.182141459
Sample Variance	0.00894571	0.033175511
Cov	0.523913181	0.322623736
Kurtosis	24.37702528	177.7557745
Skewness	4.361958809	11.79068893
Range	0.87	2.97
Minimum	0.01	0.01
Maximum	0.88	2.98
Sum	18.83	22.33
Count	380	380
Confidence Level (95.0%)	0.009540097	0.018371905

Bureau Veritas Coarse duplicates

Cu

good correlation - with no bias shown in the q-q plots- ranked hard plots show that 95% of the data is within 10% difference.

	ORIG_CU	DUP_CU
Mean	226.8842105	225.5657895
Standard Error	19.34147584	19.17729379
Median	67.5	68
Mode	2	2
Standard Deviation	377.0347748	373.8342773
Sample Variance	142155.2214	139752.0669
Cov	0.601759375	0.603384449
Kurtosis	13.18021271	12.65638748
Skewness	3.287199284	3.238913648
Range	2648	2558
Minimum	2	2
Maximum	2650	2560
Sum	86216	85715
Count	380	380
Confidence Level (95.0%)	38.03004095	37.70721914

Bureau Veritas Coarse duplicates

Ni

good correlation - with no bias shown in the q-q plots- ranked hard plots show that 99% of the data is within 10% difference.

	ORIG_NI	DUP_NI
Mean	654.1973684	654.9131579
Standard Error	36.93821846	36.5989079
Median	453.5	461.5
Mode	713	1220
Standard Deviation	720.0584376	713.444057
Sample Variance	518484.1536	509002.4225
Cov	0.908533717	0.917960072
Kurtosis	13.61782683	12.97491658
Skewness	2.954683939	2.903411662
Range	6238	6058
Minimum	2	2
Maximum	6240	6060
Sum	248595	248867
Count	380	380
Confidence Level (95.0%)	72.62951245	71.96234544

Bureau Veritas Pulp duplicates

Pt

Good correlation - with no bias shown in the q-q plots- ranked hard plots show that 90% of the data is within 10% difference.

	ORIG_PT	DUP_PT
Mean	0.758648325	0.749196643
Standard Error	0.095942924	0.097402503
Median	0.05	0.05
Mode	0.005	0.005
Standard Deviation	1.961557713	1.989015391
Sample Variance	3.84770866	3.956182226
Cov	0.386758096	0.376667092
Kurtosis	64.41827159	72.13926058
Skewness	6.360900348	6.762930619
Range	25.495	26.595
Minimum	0.005	0.005
Maximum	25.5	26.6
Sum	317.115	312.415
Count	418	417
Confidence Level (95.0%)	0.188592046	0.191462435

Bureau Veritas Pulp duplicates

Pd

Good correlation - with no bias shown in the q-q plots- ranked hard plots show that 87% of the data is within 10% difference.

	ORIG_PD	DUP_PD
Mean	0.333502392	0.327597122
Standard Error	0.037142641	0.036790643
Median	0.02	0.02
Mode	0.005	0.005
Standard Deviation	0.759383096	0.751286183
Sample Variance	0.576662687	0.564430929
Cov	0.439175423	0.436048379
Kurtosis	11.29962812	11.07675689
Skewness	3.177058784	3.17495377
Range	5.185	4.965
Minimum	0.005	0.005
Maximum	5.19	4.97
Sum	139.404	136.608
Count	418	417
Confidence Level (95.0%)	0.073010144	0.072318737

Bureau Veritas Pulp duplicates

Au

Good correlation - with no bias shown in the q-q plots- ranked hard plots show that 90% of the data is within 10% difference.

	ORIG_AU	DUP_AU
Mean	0.061937799	0.061455635
Standard Error	0.005171332	0.005197407
Median	0.02	0.02
Mode	0.01	0.01
Standard Deviation	0.10572813	0.106134048
Sample Variance	0.011178437	0.011264436
Cov	0.585821383	0.579037893
Kurtosis	17.40229217	18.35229834
Skewness	3.516813117	3.615471304
Range	0.96	0.97
Minimum	0.01	0.01
Maximum	0.97	0.98
Sum	25.89	25.627
Count	418	417
Confidence Level (95.0%)	0.010165127	0.010216453

Bureau Veritas Pulp duplicates

Cu

Good correlation - with no bias shown in the q-q plots- ranked hard plots show that >95% of the data is within 10% difference.

	ORIG_CU	DUP_CU
Mean	252.3684211	254.6337349
Standard Error	17.76934786	17.881425
Median	85.5	83
Mode	2	2
Standard Deviation	363.2951752	364.2723218
Sample Variance	131983.3843	132694.3245
Cov	0.69466494	0.699020265
Kurtosis	6.981216365	6.815466053
Skewness	2.413455812	2.385964769
Range	2518	2518
Minimum	2	2
Maximum	2520	2520
Sum	105490	105673
Count	418	415
Confidence Level (95.0%)	34.92865892	35.14970689

Bureau Veritas Pulp duplicates

Ni

Good correlation - with no bias shown in the q-q plots- ranked hard plots show that >95% of the data is within 10% difference.

	ORIG_NI	DUP_NI
Mean	732.0956938	731.1301205
Standard Error	36.64716981	36.71140174
Median	522	526
Mode	1190	1210
Standard Deviation	749.2531568	747.8681117
Sample Variance	561380.293	559306.7125
Cov	0.97710058	0.97761906
Kurtosis	6.407014885	6.468622971
Skewness	2.06227429	2.064847656
Range	5778	5788
Minimum	2	2
Maximum	5780	5790
Sum	306016	303419
Count	418	415
Confidence Level (95.0%)	72.03621117	72.16399199

Standards

In general the Standards perform reasonably well- there is a distinct deviation on the Au readings for most of the standards indicating that the Au certified value is too close to detection limit- if required another standard should be considered for Au analysis.

STD_REF	STANDARD _TYPE	PT_ME AN	PT_3 S TD_MI N	PT_3S TD_M A X	PD_ME AN	PD_3S TD_MI N	PD_3S TD_M A X	AU_ME AN	AU_3S TD_MI N	AU_3S TD_M A X	CU_MEA N	CU_3STD _MIN	CU_3STD _MAX	NI_MEAN	NI_3STD_ MIN	NI_3STD_ MAX
SAND	BLANK		-	0.08		-	0.08		-	0.08		-	100.00		-	100.00
AMIS0001	STANDARD	0.77	0.66	0.87	1.04	0.92	1.16	0.12	0.08	0.16
AMIS0089	STANDARD	1.09	0.91	1.27	0.70	0.61	0.79	0.04	0.02	0.06	59.00	50.00	68.00	452.00	374.00	530.00
AMIS0099	STANDARD	0.59	0.50	0.68	0.23	0.17	0.28	0.09	0.07	0.11	256.00	229.00	283.00	443.00	371.00	515.00
AMIS0151	STANDARD	4.64	4.10	5.18	3.15	2.73	3.57	0.07	0.05	0.09	150.00	129.00	171.00	1,281.00	988.50	1,573.50
AMIS0165	STANDARD	16.90	14.86	18.94	19.10	17.06	21.14	1.66	1.45	1.87	17,710.00	16,165.00	19,255.00	28,160.00	25,490.00	30,830.00
AMIS0207	STANDARD	2.28	1.95	2.61	1.26	1.14	1.38	0.09	0.07	0.10	85.00	71.50	98.50	1,059.00	871.50	1,246.50
AMIS0254	STANDARD	2.19	2.17	2.21	1.12	1.00	1.24	0.20	0.17	0.23	762.00	688.50	835.50	1,735.00	1,469.50	2,000.50
AMIS0256	STANDARD	4.86	4.53	5.19	2.50	2.32	2.68	0.34	0.28	0.40	1,252.00	1,148.50	1,355.50	2,913.00	2,641.50	3,184.50
AMIS0411	STANDARD	0.54	0.45	0.63	0.67	0.58	0.76	0.08	0.06	0.10	742.00	652.00	832.00	1,368.00	1,216.50	1,519.50