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EASTERN PLATINUM LTD

Technical Report Update on the Crocodile River Mine, Eastern Platinum Limited, North West Province, South Africa, December 2010

Effective Date: November 30, 2010

Brian Montpellier, P. Eng.

TABLE OF CONTENTS

1. SUMMARY ............................................................................................................. 1

1.1 General description of project ............................................................................ 1

1.2 Geology and mineralization ................................................................................ 1

1.3 Resources and reserves ..................................................................................... 2

2. INTRODUCTION AND TERMS OF REFERENCE .................................................................. 4

2.1 TERMS OF REFERENCE ....................................................................................... 4

2.2 SOURCES OF INFORMATION ................................................................................ 4

3. RELIANCE ON OTHER EXPERTS .................................................................................. 5

4. PROPERTY DESCRIPTION AND LOCATION ...................................................................... 6

5. ACCESSABILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ........... 8

5.1 ACCESS .......................................................................................................... 8

5.2 CLIMATE ........................................................................................................ 8

5.3 LOCAL RESOURCES, INFRASTRUCTURE .................................................................. 9

5.4 PHYSIOGRAPHY ............................................................................................... 9

6. HISTORY ............................................................................................................ 10

6.1 PRIOR OWNERSHIP ......................................................................................... 10

6.2 HISTORICAL MINERAL RESOURCES AND MINERAL RESERVES ...................................... 10

7. GEOLOGICAL SETTING ........................................................................................... 12

7.1 PROJECT GEOLOGY ........................................................................................ 12

7.2 CORELATION BETWEEN PLATINUM ELEMENTS ....................................................... 16

8. DEPOSIT TYPE ..................................................................................................... 17

9. MINERALIZATION .................................................................................................. 18

10. EXPLORATION ..................................................................................................... 19

11. DRILLING ............................................................................................................ 20

11.1 SURFACE DRILLING INFORMATION ...................................................................... 20

11.2 SURFACE DRILLING – STATISTICAL ANALYSES ........................................................ 22

12. SAMPLING METHOD AND APPROACH .......................................................................... 31

12.1 CORE DRILLING SAMPLING PROCEDURE ................................................................ 31

12.2 RC SAMPLING ................................................................................................ 31

13. SAMPLE PREPARATION, ANALYSIS AND SECURITY ......................................................... 32

13.1 TESTING LABORATORIES .................................................................................. 32



14. DATA VERIFICATION .............................................................................................. 35

14.1 QUALITY CONTROL MEASURES & VERIFICATION (“QA/QC”) ...................................... 35

14.2 LIMITATIONS ON VERIFICATION .......................................................................... 35

15. ADJACENT PROPERTIES ......................................................................................... 36

16. MINERAL PROCESSING AND METALLURGICAL TESTING ................................................... 37

17. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ................................................ 38

17.1 STANDARD RESOURCE AND RESERVE REPORTING SYSTEM ........................................ 38

17.5 DETAILED MINERAL RESOURCE TABULATION ......................................................... 53

17.6 PRILL SPLITS ................................................................................................. 63

17.7 RECONCILIATION ............................................................................................ 63

17.8 EFFECT OF MODIFYING FACTORS ....................................................................... 64

18. OTHER RELEVANT DATA AND INFORMATION ............................................................... 69

19. INTERPRETATION AND CONCLUSIONS ........................................................................ 70

20. RECOMMENDATIONS .............................................................................................. 70

21. REFERENCES ....................................................................................................... 71

22. DATE AND SIGNATURE PAGE ................................................................................... 69

23. ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND

PRODUCTION PROPERTIES ............................................................................................ 70

23.1 MINE DESIGN .................................................................................................... 70

23.2 PRODUCTION SCHEDULE ..................................................................................... 74

23.3 INFRASTRUCTURE ............................................................................................. 75

23.4 CURRENT OPERATIONS ....................................................................................... 76

23.5 MINE SERVICES ................................................................................................. 82

23.6 PLANT DESIGN ................................................................................................. 83

23.7 CROCODILE RIVER MINE CAPITAL .......................................................................... 84

23.8 CROCODILE RIVER WORKING COST ........................................................................ 84

23.9 ENVIRONMENTAL CONSIDERATIONS .................................................................... 86

23.10 TAXES AND ROYALTIES .................................................................................... 87

23.11 MARKETS ..................................................................................................... 87

LIST OF TABLES

Table 1-1 Classified Resources Table 1-2 Classified Reserves Table 6-1 History of CRM Table 6-2 Mineral Resources Estimated for CRM as at November 2006 Table 6-3 Mineral Reserves Estimated for CRM as at November 2006 Table 7-1 Geology of the UG2 CL at the Project Area Table 11-1 Total project Area Drillhlole Data Table 11-2 Statistical Analysis – Total Project Area Total Drillhole Data Resource Cut Table 11-3 2009 In-Situ Statistical Analysis – Drillhole Data per Domain 2009 Table 11-4 2009 Resource Cut Statistical Analysis – Drillhole Data per Domain 2009 Table 11-5 Drillhole Data Distribution Table 13-1 Details of Accreditation of Testing Laboratories Used for Sample Analyses Table 13-2 Analytical Procedures used by Impala Laboratory Table 17-1 Variogram Parameters – In-Situ Model Table 17-2 Variogram Parameters – resource Cut Model Table 17-3 2009 Means – Simple Kriging Table 17-4 Geological Loss Factor Table 17-5 30th May 2009 In-Situ Mineral Resources: Cut-off of 200 cm.g/t Table 17-6 30th May 2009 Resource Cut Mineral Resources: Cut-off 200cm.g/t Table 17-7 Prill Splits Table 17-8 Kriged Estimates versus Input Point Data Table 17-9 Crocodile River Mine modifying factors Table 17-10 Resource Categories in life of mine Table 17-11 Crocodile River Mine Modifying Factors Table 17-12 CRM Overall Reserves in life of mine plan Table 17-13 Classified Reserves Table 23-1 Mine Design Criteria Table 23-2 Ventilation Planning Parameters Table 23-3 Equivalent stoping square meters 2008 – 2013 Table 23-4 Maroelabult equipment Table 23-5 Zandfontein Equipment Table 23-6 CRM Capital Table 23-7 Actual and Budget cash cost summary to 2013 (Rm) Table 23-8 Working Cost

LIST OF FIGURES

Figure 4-1 Schematic Layout of CRM Figure 5-1 Location of the Crocodile River Mine Figure 7-1 Generalized Stratigraphic Section of the UG2 Chromitite Layer Figure 7-2 Location of Eastplats Project Area in Relation to the Bushveld Igneous

Complex Figure 7-3 Location of CRM Project Area in Relation to the Western Limb of the

Bushveld Igneous Complex Figure 7-4 Generalised Grade Histogram of the UG2 CL Figure 11-1 Location of Drillholes Figure 11-2 Vertical grade distribution of the UG2 Main Band Figure 11-3 Contouring of UG2CL Horizon Figure 11-4 Location of Domains pertaining to the UG2W Project Area Figure 17-1 In-Situ Pooint Variograms – Per Domain Figure 17-2 Resource Cut Point Variograms – Per Domain Figure 17-3 4E Grade Tonnage Curves for In-Situ UG2 CL – Zandfontein Figure 17-4 4E Grade Tonnage Curves for In-Situ UG2 CL – Crocette Figure 17-5 4E Grade Tonnage Curves for In-Situ UG2 CL – Kareespruit Figure 17-6 4E Grade Tonnage Curves for In-Situ UG2CL – Maroelabult Figure 17-7 Resource Classification Plot Figure 17-8 Resource Classification Plot – Farm Boundaries Figure 17-9 Resource Classification Plot – Farm Boundaries on Topographic Map Figure 23-1 Life of Mine M2 Figure 23-2 Life of Mine – Tons Figure 23-3 Zandfontein Mine Figure 23-4 Crocette Mine Plan

Technical Report Update - Crocodile River Mine - December 1, 2010 1

1. SUMMARY

The purpose of this report is to compile a Technical Report of the Mineral Resource estimate of the

UG2 Chromitite Layer (“UG2 CL”) at the Crocodile River Mine (“the Mine” or “CRM” or “the Project”),

Eastern Platinum (“East Plats”), compliant with the definitions embodied in the definitions set out by

the Canadian Institute of Mining, Metallurgy and Petroleum Code (“CIM”) Instrument NI 43-101 for

reporting of Resources and Reserves. The work was done by outside consultants under the supervision

of Eastplats. Mr Brian Montpellier, P.Eng. Vice President of Eastplats, who reviewed the work and takes

responsibility for this report.

1.1 GENERAL DESCRIPTION OF PROJECT

CRM, incorporates mining of the UG2 CL, and the Project is located south west of the town of Brits, in

the North West Province of South Africa. The Project is located on the Western Limb of the Bushveld

Igneous Complex (“BIC”). The mineralization of interest at CRM is the platiniferous UG2 CL.

1.2 GEOLOGY AND MINERALIZATION

The Project mineralization comprises platinum (Pt), palladium (Pd), rhodium (Rh) and gold (Au) hosted

by the UG2 CL, a chromitite layer comprised of a basal feldspathic pyroxenite pegmatoid and overlying

chromitite layers termed leader and triplets.

A key component of the data utilized in the compilation of the estimation models was the coding of the

data for the In-Situ and Resource width models. The data is comprised of two datasets, the most

recent incorporating the boreholes drilled by East Plats during 2007 to 2009, and the historical

boreholes drilled by previous owners of the mineral rights. With regards to the stratigraphic coding of

the economic UG2 unit, the historical data on the whole was coded as UG2 for the entire geological

unit and not coded clearly as to the main band (hereafter termed the UG2MB), triplets and leaders.

Work undertaken by the East Plats geologists have attempted to code the data according to the sub-

units, such as the UG2MB, although this process is not complete. This more detailed coding was used

together with the lithological rock types and vertical grade profiles to code the remainder of the

drillhole intersections that constitute the database used for the Mineral Resource estimation. The data

was coded as “UG2CW” for the In-Situ Mineral Resource estimation and “UG2MC” for the Resource cut

estimation. The majority of the drillholes once coded were imported into Datamine‟s Downhole

Explorer package and the coding checked by means of plotting the vertical grade profile of the UG2CW

unit against the rock types and intersection depths of the UG2.

With regards to the calculation of the Resource cut widths, after consultation with the Mine a Resource

cut width of 1.5m was determined in 2008. A Resource cut width is the width of the mineralised unit

calculated on optimised geological cut based on historical and envisaged mining criteria. The width of

1.5m was determined using the average channel width of the UG2 Main Band (~1.4m) and the average

historical mining widths achieved (1.52m). The data was coded accordingly from the base sample

(across contact) upwards and coded “UG2MC”. A problem with the historical data is the frequent lack

2

of assay values above the UG2MB. Where this occurred, a zero grade (0g/t) was input into the

database. Although this will penalise the ore body it can be seen to represent the minimum value that

will be obtained at that mining width. Where the lengths of the drillholes exceeded 1.5m cut, the full

width intersection was maintained.

1.3 RESOURCES AND RESERVES

Resources and Reserves are given in the following tables. The Resources are effective as of 30 May

2009. The Reserves are effective as of 30 November 2010.

Table 1-1 Classified Resources

3

Table 1-2: Classified Reserves

Reserve Area Tonnage

(Kt)

Metal Concentration Contained Metal

4E (g/t) Pt (g/t) Pd (g/t) Rh (g/t) Au (g/t) 4E

Proven Kg K Oz

Crocette 0

Maroelabult 132 3.91 2.42 1.06 0.4 .04 515 17

Zandfontein 842 4.08 2.53 1.10 0.4 .04 3,435 110

Sub Total Proven 974 4.06 2.52 1.09 0.4 .04 3,950 127

Probable

Crocette 3,530 3.78 2.41 0.95 0.4 .02 13,340 429

Maroelabult 2,173 3.91 2.48 1.00 0.4 .02 8,487 272

Zandfontein 22,748 4.08 2.56 1.06 0.4 .03 92,896 2,987

Sub Total

Probable

28,451 4.03 2.54 1.04 0.4 .02 114,723 3,688

Total 29,425 4.03 2.54 1.04 0.42 0.03 118,673 3,815


2. INTRODUCTION AND TERMS OF REFERENCE

2.1 TERMS OF REFERENCE

The purpose of this report is to compile a Technical Report of the Mineral Resource and Reserve

estimate of the UG2 Chromitite Layer (“UG2 CL”) at the Crocodile River Mine (“the Mine” or “CRM” or

“the Project”), Eastern Platinum Limited (“East Plats”), compliant with the definitions embodied in

the Canadian Institute of Mining, Metallurgy and Petroleum Code (“CIM”) Instrument NI 43-101 for

reporting of Resources and Reserves..

2.2 SOURCES OF INFORMATION

The following people provided information that was used to compile this report:-

Geology and Database: Mrs P Naidoo & Mr M Ruygrok (Previously Eastplats); Mr T Mosholi (Eastplats);

Mineral Resources: Minxcon, Mrs HL King; Mineral Reserves: Mr Richard Skinner, Eastplats.

5

3. RELIANCE ON OTHER EXPERTS

Not Applicable

6

4. PROPERTY DESCRIPTION AND LOCATION CRM is located approximately 7km south of the town of Brits, North West Province, South Africa, and

approximately 70km from the city of Johannesburg. The Project is located at Latitude 25°42‟30”S and

Longitude 24°47‟30”E. CRM is easily accessible via the N4 tarred national road (Figure 1). The mine

covers an area of 11,256.24ha. The Project Area is located on the Western Limb of the Bushveld

Igneous Complex (“BIC”), a large layered igneous province world renowned for its platinum group metal

(“PGM”) / platinum group elements (“PGE”) content. The ore body of interest at CRM is the UG2

Chromitite Layer (“UG2 CL”).

The Project comprises four mining blocks (Figure 4-1):-

• Maroelabult (currently operational);

• Zandfontein (currently operational);

• Crocette (under development); and

• Kareespruit (under exploration).

Figure 4-1: Schematic Layout of CRM

7

The operating sections of the mine consist of the farms De Kroon 444, De Kroon 444 JQ, Zandfontein

447 and Krokodildrift 446 JQ. These are covered under New Order Mining Rights numbers 78, 151, 307,

332 and 363 issued between December 2006 and January 2009. All other areas of interest are covered

by Prospecting Rights which can be converted to New Order Mining Rights should any exploration work

be positive.

8

5. ACCESSABILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY

5.1 ACCESS

CRM is located approximately 7km south of the town of Brits, North West Province, South Africa, and

approximately 70km from the city of Johannesburg. CRM is easily accessible via the N4 tarred national

road. The mine covers an area of 11,256.24ha. The Project Area is located on the Western Limb of the

Bushveld Igneous Complex, a large layered igneous province world renowned for its platinum group

metal/platinum group elements content. The ore body of interest at CRM is the UG2 Chromitite Layer.

Figure 5-1: Location of the Crocodile River Mine

5.2 CLIMATE

9

Climate in the area is warm to hot in summer, with dry, mild, generally frost-free winters. The local

area can be described as sub-humid with a mean annual rainfall of 540 mm per annum as measured at

the Brits weather station. Rainfall is almost exclusively restricted to the summer months between

October and March. The average daily temperature in June is 19°C, and in January 29°C. There are

occasional short excursions to -2°C and occasional maximum temperatures of 44.5°C are experienced.

5.3 LOCAL RESOURCES, INFRASTRUCTURE

The area is significantly developed with PGM and chromite mining, heavy agriculture and light industry

all around the town of Brits. One ferrochrome mine and smelter is nearby, and several platinum mines

are in operation.

The town of Brits, to which the mine is adjacent, is well developed with many businesses offering mine

support services. Local infrastructure in terms of airports, road, rail, power and water is excellent,

with a knowledgeable, skilled local labour force. Highway N4 which traverses the northern part of

South Africa from Mozambique to Botswana via Nelspruit, Pretoria and Rustenburg crosses the southern

portion of the property.

5.4 PHYSIOGRAPHY

The property is bounded to the east by Xstrata‟s Eland mine and to the west be properties held by

Impala Platinum, Anglo Platinum and Lonmin. To the north and south is farmland. The Crocodile River

runs across the southern portion of the property. The land in the area is generally flat but to the south

(off the property) are the Magaliesberg hills and nature reserve.

10

6. HISTORY

6.1 PRIOR OWNERSHIP

The prior ownership and history of the Project is summarised in the table which follows:-

Table 6-1: History of CRM

Date Activities 1987 CRM (then Lefkochrysos) was listed in the Johannesburg Securities Exchange (“JSE”) and

development of decline at Zandfontein Section. The mine design was based on the concept of using a mechanised mining system to produce 160,000tpm.

1988 Controlling interest acquired by Rand Mines Limited through its subsidiary, Barplats Mines Ltd. 1991 Impala Platinum Holdings Limited (“Implats”) acquired a 38% interest in Barplats and the contract to manage CRM. Mining operations were suspended in late 1991 due to the decline in metal prices.

1998 Implats increased its interest in Barplats to 83%.

2000 After further exploration, the mine was brought back into production. The plant was also refurbished at this time.

2003 Mining operations ceased, and CRM was put on care and maintenance.

2004 Limited mining activities. A consortium of investors acquired a majority shareholding in the mine, and in December 2004 CRM began to refurbish its stoping sections.

2005 Drilling and blasting operations recommenced at CRM.

2006 East Plats acquired a 69% stake in Barplats.

2007 East Plats acquired a further 5% interest (total of 74%) in Barplats.

6.2 HISTORICAL MINERAL RESOURCES AND MINERAL RESERVES

Table 6- 2: Mineral Resources Estimated for CRM as at 10 November 2006

Mineral Resource Category

Tonnage (‘000t)

Reef Thickness (m)

Grade PGM+Au (g/t)

Content 3PGM+Au (oz)

Measured 6,894 1.44 4.19 928,000 Indicated 30,324 1.40 4.41 4,303,000 Total Measured and Indicated

37,218 1.41 4.38 5,231,000

Inferred 52,482 1.40 4.41 7,449,000 Note: Mineral Resources are inclusive of Mineral Reserves The following table summarises the Mineral Reserves estimated by RSG Global:- Table 6-3: Mineral Reserves Estimated for CRM as at 10 November 2006

Mineral Reserve Category)

Section

Tonnage (‘000t)

3PGM+Au (g/t

PGM (‘000oz)

Proven Zandfontein 3,546 3.48 397

Maroelabult 1,962 3.49 217

Total Proven 5,508 3.48 614

Probable Zandfontein 16,314 3.44 1,802

Maroelabult 2,076 3.69 247

11

Total Probable 18,390 3.47 2,049


7. GEOLOGICAL SETTING

7.1 PROJECT GEOLOGY

The Project is located on the Western Limb of the BIC (Figure 4) in a structurally complex area close to

the town of Brits. The geology of the Project Area has been described in the report compiled by RSG

Global as part of the Mineral Resource and Mineral Reserve Estimate of CRM, dated November 2006.

The following are extracts, comprising the „Project Geology‟ section, from this report.

CRM is located within and adjacent to the Brits Graben and mines the UG2 CL. The Merensky Reef in

this area is not currently considered economically viable. The graben and associated faults result in the

mine being split into four main sections/mining blocks (Figure 2).

The Upper Critical Zone (“UCZ”) of the Rustenburg Layered Suite (“RLS”) outcrops extensively, striking

inan east-west direction and dipping to the north from about 15° to 25°. Both the Merensky Reef and

theUG2 CL outcrop in the area, with a middling of approximately 200m. The UG2 CL occurs from

outcrop down to an estimated depth of at least 2,000m below surface.

The UG2 CL at CRM typically consists of a single chromitite layer some 1.35m to 1.5m thick. The

chromatitic leader hangingwall layers are generally absent or have coalesced with the main band to

form a virtually homogenous chromitite. A very thin chromitite stringer (~1mm thick) occurs at varying

heights above the top of the chromitite in the immediate hangingwall. However, variations in the

thickness of specific lithotypes do exist between individual drillholes, and several disturbances to the

layering by pothole structures and mafic pegmatites have been encountered during mining and in some

drillholes. A comparison between the features of the UG2 CL at the Zandfontein and Maroelabult

sections is presented in the table below:-

Table 7-1: Geology of the UG2 CL at the Project Area

UG2 CL Zandfontein Section Maroelabult Section Reef Thickness 1.3m – 1.4m 17°

Average Dip 17° °

Faulting Significant faulting with some scissor effects noted

Large dykes

Potholding Present Present

Other Features Reef horizon undulates Reef horizon undulates

The Bottom of Reef Contact (“BRC”) is undulating. The footwall is typically a pegmatoidal pyroxenite

but where the BRC transgresses through the pegmatoidal pyroxenite the footwall is a norite. The Top

ReefContact (“TRC”) is generally sharp and stable. The UG2 CL is predominantly impure chromitite

with muchinterstitial silicate, comprising pyroxene (bronzite) and feldspar (anorthite). It is typically

comprised of60% to 90% chromite, which is consistent with other localities in the BIC. Disseminated

sulphides areconcentrated in the lower part of the reef, which is always bottom to middle loaded with

respect to PGMconcentrations. The PGMs are associated with sulphides that are interstitial to the

chromite grains. Themineralised reef zone itself is defined by the sharp basal contact into the footwall

13

pyroxenite. The3PGE+Au (“4E”) concentration of the UG2 CL ranges from 2.5ppm to 6.6ppm, and is

generally dominatedby Pt-Pd sulphides

Other features of the UG2 CL are the undulating nature of the BRC, the presence of mafic/ultramafic

pegmatites and potholes.

Figure 7-1: Generalized Stratigraphic Section of the UG2 Chromitite Layer


Figure 7-2: Location of East Plats Project Area in Relation to the Bushveld Igneous Complex


Figure 7-3: Location of CRM Project Area in Relation to the Western Limb of the Bushveld Igneous

Complex

Figure 7-4: Generalised Grade Histogram of the UG2 CL

16

7.2 CORELATION BETWEEN PLATINUM ELEMENTS

This section summarises the correlation between the platinum elements of the UG2 CL at CRM. Not all

drillholes have individual prill elements assayed, hence only 136 units versus 411 assayed for 4E. The

UG2 CL at CRM has on average a 4E prill ratio of 62.6:26.2:10.4:0.8 in 2008 and 61.1:28.5:9.7:0.7 in

2009, derived from the drillholes with individual prill analyses. The rhodium component appears to

correlate with that of satellite BIC ore bodies and not generally the typical Western Limb BIC prill

ratio. There is a strong correlation between Pt and Rh on the UG2 CL at CRM. In 2008, the regression

relationship between the PGE was determined using a linear trend line against 4E grade. These

regression relationships are used to convert the 4E Mineral Resource to a Mineral Resource per prill

split. However, in the case of CRM, the historical assaying was typically undertaken on a single sample

that was regressed; therefore to apply this concept on this data set will introduce significant bias.


8. DEPOSIT TYPE

The mineralised units of the Project Area form part of the BIC, a large layered igneous intrusive body.

In a large layered intrusion such as the BIC, the sulphide droplets that segregate out of the parental

magma will eventually settle out of the magma, and once magma convection ceases are deposited on

already consolidated layers of the magma chamber to form sulphide-rich zones.

The Merensky Reef has traditionally been the most important PGM producing horizon in the BIC. In

addition to PGM mineralisation associated with the Merensky Reef, all chromitite layers in the Critical

Zone of the BIC contain elevated concentrations of PGM. The UG2 CL is the only chromitite layer where

significant mining of PGM takes place. The UG2 CL of the Western Bushveld Limb is generally less than

1m thick, and is comprised of 60% to 90% chromite; however in the Project Area the thickness of the

UG2 CL tends to average 1.3m to in excess of 1.5m. The PGMs are interstitial to the chromite grains,

and are concentrated at the base and middle of the chromitite layer. The PGM concentration of the

UG2 CL ranges from 3ppm to 19ppm, and is generally dominated by Pt-Pd sulphides.

18

9. MINERALIZATION

The broad stratigraphy of the UG2 sequence comprises three mineralized chromitite layers,

referenced(from bottom to top) as the main UG2, the UG2A and the UG2B units, which vary quite

significantly in their chromite content and separation distance between one another. This separation

distance will be important in any future mining operation as it will determine the underground mining

cut/stoping width and the inclusion or otherwise of low-grade waste. All the chromitite layers often

contain a variable amount of inter-reef pyroxenite, which is the cause of dilution of grade in some drill

intersections.

The lowermost main chromitite layer is the UG2 as identified elsewhere in the BC, whereas the UG2A

and UG2B are probably manifestations of the UG2 leader seams.

PGM and base metal sulphide mineralization is typically bimodal in its distribution within the main UG2

chromite with the higher PGE and Cu-Ni values being found close to the upper and lower portions of

the reef, with a distinct “bottom loading.” The central part tends to be moderately lower grade.

19

10. EXPLORATION

This report incorporates certain drilling work carried out on the property in 2008 and 2009 as described

in the next section.

20

11. DRILLING

11.1 SURFACE DRILLING INFORMATION

Several drilling campaigns have been undertaken on CRM by the pervious and current owners. The

majority of the drilling has been diamond drilling. RC and percussion drilling has been undertaken but

has not formed part of the current database used for the Mineral Resource estimation process.

11.1.1 DRILLING DATABASE & VERIFICATION

The drilling database is divided into two parts, the historical and current data. The current data refers

to the 2007 to current drilling campaign undertaken by East Plats. The historical drilling campaigns

were undertaken under the ownership of Lefkochrysos, Rand Mines, Impala and Barplats. The drilling,

logging and assaying techniques involved in each of the historical drilling campaigns have not been fully

captured, nor the procedures identical. The result being varying levels of information is available and

captured into the East Plats in-house Fusion database. With regards to the drilling, logging and assaying

procedures the differing techniques applied are considered not to have a major material effect on the

data. The majority of the historical data however lacks the support of the original drilling logs, assay

sheets and quality assurance and quality control (“QA/QC”). The data available from CRM does not

represent the entire database available for the ore body; data collection is still in progress. The data is

undergoing extensive gathering and verification. The data used from CRM represents data supplied by

the mine that is deemed valid and has undergone forms of data verification.

Data extracted from the Fusion database for use in the Mineral Resource estimation process included

collar co-ordinates, geological and sampling logs together with assay values. The assay data available

for the historical data is a combination of single 3PGE+Au (Pt, Pd, Rh & Au or 4E) assay results and prill

split analytical techniques. This is described in detail in the Assay and Sampling section of this report.

In

addition, the assays are captured as a combination of g/t and ppb. Assay sheets are difficult to come

by in order to establish the unit of measurement, e.g., 20 Pt could represent either 20g/t or 20ppb

(0.02g/t).The values that fall within this range that could represent either g/t or ppb were reviewed in

context of the over and underlying values and/or the adjacent mother hole or deflections. Once the

unit of measurement had being recalculated into g/t, the data was verified by graphical plotting of the

suspicious data and comparison to the neighbouring drillholes, as well as review of the vertical grade

profiles of the economic unit in Downhole ExplorerTM. No data as to the regression factors applied to

the 3PGE+Au single analyses were available.

With regards to the stratigraphic coding of the economic UG2 unit, the historical data on the whole was

coded as UG2 for the entire geological unit and not coded clearly as to the Main Band (hereafter

termed the UG2MB), triplets and leaders. Work undertaken by the East Plats geologists have attempted

to code the data according to the sub-units, such as the main band (UG2MB), although this process is

not complete. This more detailed coding was used together with the lithological rock types and vertical

grade profiles to code the remainder of the drillhole intersections that constitute the database used for

21

the Mineral Resource estimation. The data was coded as “UG2CW” for the in-situ Mineral Resource

estimation and “UG2MC” for the Resource cut estimation. The majority of the drillholes once coded

were imported into Datamine‟s Downhole Explorer package and the coding checked by means of

plotting the vertical grade profile of the UG2CW unit against the rock types and intersection depths of

the UG2.

A detailed log was maintained during the coding exercise as to whether the drillhole should be included

or excluded from the estimation database.

With regards to the calculation of the Resource cut widths, after consultation in November 2008 with

the Mine a Resource cut width of 1.5m was determined. A Resource cut width is the width of the

mineralised unit calculated on optimised geological cut based on historical and envisaged mining

criteria. The width of 1.5m was determined using the average channel width of the UG2 Main Band

(~1.4m) and the average historical mining widths achieved (1.52m). The data was coded accordingly

from the base sample (across contact) upwards and coded “UG2MC”. A problem with the historical data

is the frequent lack of assay values above the UG2MB. Where this occurred, a zero grade (0g/t) was

input into the database. Although this will penalise the ore body it can be seen to represent the

minimum value that will be obtained at that mining width. Where the lengths of the drillholes

exceeded 1.5m cut, the full width intersection was maintained.

CRM drillholes were typically drilled as vertical holes. Previous work indicated that the maximum

inclination away from vertical was 6%. All drillholes were assumed to be vertical for the Mineral

Resource estimation.

All drillhole data available up to and including April 2009 were included in this Mineral Resource

update.

The drillhole intersections were adjusted to represent corrected width for the Mineral Resource

estimation. A concern is the data was generated by various mining companies and hence may represent

a data set with varying levels of information

11.1.2 DRILLING TECHNIQUE

The typical drilling procedure conducted within the last two years on CRM involved the drilling of a

mother hole drilled to approximately 20-30m into the UG2 footwall. Non-directional deflections are

then drilled out of the mother hole through the UG2 CL and Merensky Reef. Sampling was conducted

over the economic horizons. Typically three non-directional deflections are drilled per hole. Additional

deflections were drilled if the core recovery in the economic unit was deemed sub-optimal. As

mentioned in the previous section, complete data per drillhole is often not available and hence within

the data used for the Mineral Resource estimation exercise data is missing with regards to either the

mother hole or deflections.

Due to the number of historical drilling campaigns and the extremely faulted nature of the ore body,

the drilling did not adhere to a uniform grid spacing. For the current Mineral Resource estimation the

data deemed to fulfil the requirements of being valid data typically had a 100 to 350m variable

spacing.

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11.1.3 FINAL DRILLING DATABASE USED FOR THE ESTIMATION PROCESS

411 drillholes (mother and deflections) were available for the Mineral Resource estimation process.

Figure 7 below illustrates the composition of the dataset. The 2007 Mineral Resource estimation was

based on 188 boreholes. The Mineral Resource has a substantial greater coverage of data, especially in

the areas of Kareespruit and Crocette.

Figure 11-1: Location of Drillholes

11.2 SURFACE DRILLING – STATISTICAL ANALYSES

Statistical analyses were performed on the 4E grade, 4E content (cm.g/t), SG and channel width (CW)

of the dataset used in the In-Situ and Resource Cut Mineral Resource models.

23

Descriptive statistics in the form of histograms (frequency distributions), probability plots (used to

evaluate the normality of the distribution of a variable) and trend plots (linear and quadratic) were

usedto develop an understanding of such statistical relationships.

11.2.1 SUMMARY OF STATISTICAL COMPARISON OF DATA SETS – CHANNEL WIDTH MODEL

Drillhole Data Analyses The tables below summarise the descriptive statistics for the total borehole dataset:-

Table 11-1 Total project Area Drillhole Data

Descriptive Statistics (compall_stats) Variable Valid N Mean Minimum Maximum Variance Std.Dev. Skewness Kurtosis

4E 411 4.5881 0.41168 8.328 0.92 0.9566 0.40964 2.31990 In 4E 411 1.4985 -0.88751 2.120 0.06 0.2437 -3.13338 28.11260 CW 411 132.9078 57.37800 224.732 657.28 25.6375 0.45190 0.74321 CMGT 411 609.5915 60.62864 1290.160 31193.86 176.6178 0.81604 1.55214 In cmgt 411 6.3694 4.10477 7.163 0.09 0.3078 -1.25817 8.27431 Coefficient of Variation (“CoV”) is 0.29 for 4E accumulation and 0.21 for 4E grade.

The histograms indicate a log-normal verging on normal distribution, with relatively few outliers.

Outliers will typically skew the data towards higher variances and affect the grade estimation in the

areas of the outlier values. The normal and log probability plots also indicated a log-normal verging on

normal distribution, with limited outliers. An aspect of the normal and log probability plots is the

limited distribution of the data along the probability line – this may be the result of the historical

values been assayed for total 4E with a regression value applied. The CoV is low, thereby indicating

that the data set has a low variance.

Total Drillhole Data Analyses – Resource Width

The tables below summarise the descriptive statistics for the entire borehole dataset, as coded into a Resource width of 1.5m:-

Table 11-2: Statistical Analysis – Total Project Area Total Drillhole Data Resource Cut


4E 408 4.508 0.4382 8.328 0.95 0.9752 0.29003 1.40246 Ln 4E 408 1.3660 -0.8251 2.120 0.08 0.2753 -2.03585 12.69572 CW 408 132.9078 146.3150 224.732 127.64 11.2980 3.08309 9.95321 CMGT 408 609.5915 65.7893 1290.160 29847.49 172.7643 0.72800 1.60733 Ln cmgt 408 6.3694 4.1865 7.163 0.09 0.2976 -1.48541 9.48553

Coefficient of Variation (“CoV”) is 0.27 for 4E accumulation and 0.24 for 4E grade. Three drillholes

were excluded from the Resource cut estimation based primarily on faulting in the hanging wall and

anomalous data.

The grade and content histograms indicate a unimodal distribution, albeit a very small high grade

portion. Plotting the data in log space and analysing the probability plots indicates that the data

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population does not form a true normal distribution, rather a log-normal distribution. The log

probability plots confirm this.

11.2.2 VERTICAL GRADE DISTRIBUTION OF THE UG2 CL

Review of the vertical grade distribution across the UG2 CL indicates at least a bimodal grade profile.

In the majority of cases the UG2 CL is bottom and middle loaded decreasing to the top contact of the

layer. This bimodal distribution hold true irrespective of the footwall unit, being the usual pegmatoidal

pyroxenite or norite, etc.

Figure 11-2 Vertical grade distribution of the UG2 Main Band

11.2.3 DOMAINING

The domains delineated in November 2008 were retained for the Mineral Resource modelling update,

although they were extended where necessary to cover all the available drillhole data. The following is

an extract from the 2008 Minxcon report dated the 30th November 2008.

The data was analysed extensively to establish any geostatistical domaining of the ore body.

Techniques employed were Cumulative Sum Analysis (“CUSUM”), reviewing loading of the UG2 channel

width profile, review of the footwall lithologies, trend analysis, inverse distance squared modelling on

CW and grade, structure, and correlation between elements (e.g., Pt vs Rh, etc.)

CUSUM, grade profiles and footwall lithological controls did not render any distinct patterns in the

data. CRM Geology Department undertook contouring of several aspects of the UG2 unit which are

described below. The following diagrams illustrate the relationship contoured for the UG2 CL as well as

the leaders, footwall unit thickness, associated chromitite layers, etc, on the Project Area. Although

there is an area in the middle of the contour plots with limited borehole representation, as only the

current boreholes were used, the contour plots are indicating variation which broadly corresponds to

the domains created for the deposit (Figure 9). A similar exercise will be conducted on the historical

drillholes. A remaining aspect which should be investigated is the relationship of the Transvaal

Supergroup footwall units on the characteristics of the UCZ. It is well known that in this area the

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Transvaal Supergroup footwall units have had a major impact on the characteristics of the LG and MG

chromitite seams. The Transvaal Supergroup footwall units in the Brits section of the Western BIC have

extensively influenced the development of the LCZ, and therefore may have influenced the formation

of the UCZ of the BIC, with localised changes away from the norm of the chromitite layers. Findings

from this modelling may possibly refine and add value to the current models available for this area.

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Figure 11-3: Contouring of UG2CL Horizon

UG2 Leaders UG2 Main Band Width

UG2 Triplets Width UG2 Footwall Widths

27

UG2 Main Waste Band Width Boreholes Utilized

The following domains/geozones were applied to the data. The domains were derived essentially from

the major structural interruptions on the UG2 CL horizon, essentially as summarised above because

there currently exists no definitive facies model for the area.

Figure 11-4: Location of Domains pertaining to the UG2CW Project Area

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Trend analysis In 2008, a trend analysis based on the 2008 dataset was undertaken in order to establish if the four domains represented homogeneous domains. Linear trend analysis was undertaken. The results indicate that Domains 1 to 4 each represent relatively homogeneous domains. The statistical analyses of the drillhole dataset, per domain, are illustrated in the tables below:- Table 11-3: 2009 In-Situ Statistical Analysis –Drillhole Data per Domain 2009 Domain 1 – Crocette


UG2CW_4E 72 4.3322 2.6515 6.595 0.52 0.7233 0.647945 1,216620 UG2CW_CW 72 139.4877 77.4610 224.732 857.64 29.2855 0.668634 0.209009 UG2CW_CMGT 72 611.5367 245.9571 1261.278 39509.76 198.7706 1.310440 1.951538 Ln_UG2CW_4E 72 1.4526 0.9751 1.886 0.03 0.1651 -0.004093 0.757081 Ln_UG2CW_CW 72 4.9169 4.3498 5.415 0.04 0.2057 0.130138 -060293 Ln_UG2CW_CMGT 72 6.3696 5.5052 7.140 0.09 0.3019 0.308056 0.669907 CoV for 4E accumulation is 0.33 for content and 0.16 for 4E grade.

Domain 2 - Zandfontein

Variable Valid N Mean Minimum Maximum Variance Std.Dev. Skewness Kurtosis

UG2CW_4E 135 4.6306 2.5937 8.328 0.78 0.8847 0.789535 2.166808 UG2CW_CW 135 131.0421 57.3780 212.300 538.01 23.1950 0.625855 1.709967 UG2CW_CMGT 135 609.6685 235.6335 1290.160 29050.68 170.4426 0.867545 1.513303 Ln_UG2CW_4E 135 1.5150 0.9531 2.120 0.04 0.1887 -0.055175 0.888205 Ln_UG2CW_CW 135 4.8600 4.0497 5.358 0.03 0.1783 -0.399447 2.913747 Ln_UG2CW_CMGT 135 6.3750 5.4623 7.163 0.08 0.2783 -0.213196 0.913265 CoV for 4E accumulation is 0.28 for content and 0.20 for 4E grade. Domain 3 – Kareespruit


UG2CW_4E 118 4.6801 0.41168 7.607 1.03 1.0154 -0.74062 3.70861 UG2CW_CW 118 128.0962 71.72300 196.042 538.11 23.1972 0.16824 0.58919 UG2CW_CMGT 118 597.9248 60.62864 1062.582 28204.97 167.9434 -0.08150 0.77222 Ln_UG2CW_4E 118 1.5057 -0.88751 2.029 0.11 0.3329 -4.50648 28.65387 Ln_UG2CW_CW 118 4.8359 4.27281 5.278 0.04 0.1876 -0.54188 0.97224 Ln_UG2CW_CMGT 118 6.3416 4.10477 6.968 0.13 0.3659 -267231 13.31077 CoV for 4E accumulation is 0.28 for content and 0.22 for 4E grade. Domain 4 – Maroelabult


UG2CW_4E 86 4.6097 2.4213 8.128 1.25 1.1195 1.051569 1.085720 UG2CW_CW 86 136.9295 65.9850 196.999 774.47 27.8293 -0.011697 0.073215 UG2CW_CMGT 86 623.8497 347.3298 1257.349 32443.53 180.1209 1.198504 1.661739 Ln_UG2CW_4E 86 1.5013 0.8843 2.095 0.05 0.2298 0.382499 0.282088 Ln_UG2CW_CW 86 4.8975 4.1894 5.283 0.05 0.2162 -0.765777 1.287093 Ln_UG2CW_CMGT 86 6.3988 5.8503 7.137 0.07 0.2689 0.441520 0.33764 CoV for 4E accumulation is 0.29 for content and 0.26 for 4E grade.

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Review of the descriptive statistics per domain illustrates the close similarity between the domains

with regards to grade (g/t) and content (cm.g/t). The ore body appears to be fairly consistent across

the Mine. With regards to CW, Maroelabult and Crocette have thicker seam widths in comparison to

Kareespruit and Zandfontein.

Table 11-4: 2009 Resource Cut Statistical Analysis –Drillhole Data per Domain 2009 Domain 1 – Crocette


UG2CW_4E 72 3.9245 1.8774 6.595 0.74 0.8629 0.697581 1.148671 UG2CW_CW 72 164.6139 148.5000 235.000 300.38 17.3314 2.001479 4.468651 UG2CW_CMGT 72 653.4929 281.6169 1318.908 39315.35 198.2810 1.418546 2.358904 Ln_UG2CW_4E 72 1.3438 0.6299 1.886 0.05 0.2191 -0.168852 0.999632 Ln_UG2CW_CW 72 5.0987 5.0006 5.460 0.01 0.0969 1.697191 3.003188 Ln_UG2CW_CMGT 72 6.4425 5.6405 7.185 0.08 0.2784 0.425518 0.976084 CoV for 4E accumulation is 0.30 for content and 0.22 for 4E grade. Domain 2 - Zandfontein


UG2CW_4E 135 4.1625 1.7672 58.328 0.95 0.9721 0.685796 1.874759 UG2CW_CW 135 156.3259 134.0000 222.000 166.89 12.9187 2.571051 7.839355 UG2CW_CMGT 135 652.9329 265.0820 1349.110 29941.34 173.0357 0.922464 1.570114 Ln_UG2CW_4E 135 13991 0.5694 2.120 0.06 0.2360 -0.322654 1.040197 1Ln_UG2CW_CW 135 5.0489 4.8978 5.403 0.01 0.0757 2.240062 5.990005 Ln_UG2CW_CMGT 135 6.4480 5.5800 7.207 0.07 0.2601 -0.077063 0.840194 CoV for 4E accumulation is 0.26 for content and 0.23 for 4E grade. Domain 3 – Kareespruit


UG2CW_4E 118 4.6801 0.41168 7.607 1.03 1.0154 -0.74062 3.70861 UG2CW_CW 118 128.0962 71.72300 196.042 538.11 23.1972 0.16824 0.58919 UG2CW_CMGT 118 597.9248 60.62864 1062.582 28204.97 167.9434 -0.08150 0.77222 Ln_UG2CW_4E 118 1.5057 -0.88751 2.029 0.11 0.3329 -4.50648 28.65387 Ln_UG2CW_CW 118 4.8359 4.27281 5.278 0.04 0.1876 -0.54188 0.97224 Ln_UG2CW_CMGT 118 6.3416 4.10477 6.968 0.13 0.3659 -267231 13.31077 CoV for 4E accumulation is 0.26 for content and 0.24 for 4E grade. Domain 4 – Maroelabult


UG2CW_4E 90 4.1597 1.1470 7.040 1.02 1.0080 0.57329 1.211692 UG2CW_CW 90 160.7889 146.0000 206.000 223.07 14.9354 1.52787 1.315838 UG2CW_CMGT 90 672.7329 172.0467 1314.800 36645.88 191.4311 0.84893 1.363288 Ln_UG2CW_4E 90 1.3951 0.1371 1.952 0.07 0.2562 -1.03969 5.567842 Ln_UG2CW_CW 90 5.0762 4.9836 5.328 0.01 0.0875 1.40770 0.929456 Ln_UG2CW_CMGT 90 6.4712 5.1478 7181 0.09 0.2916 0.69314 3.892141 CoV for 4E accumulation is 0.28 for content and 0.24 for 4E grade.

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11.2.4 DATA DISTRIBUTION

The drillholes have varying spatial distribution across the Project Area, varying between 100m and over

350m. The table below summarises the average drillhole spacing in each Resource area.

Table 11-5: Drillhole Data Distribution

Resource Area Average Drillhole Spacing (metres Average Drillhole Spacing (metres)

Crocette 100m 100m Zandfontein 350m 350m Kareespruit 350m 350m Maroelabult 100m 100m Typically, the historical and current drilling campaigns are limited to specific areas, for example,

Crocette and Kareespruit are predominantly covered by current drilling, whereas, Zandfontein and

Maroelabult are covered by historical drilling. No data is available for this Resource modelling that is

located close to the lease boundary in the north and northwest. Drillhole data from the Xstrata Mine

(Eland) on the eastern boundary has been included in the Mineral Resource estimation.

11.2.5 TREATMENT OF MISSING VALUES AND „OUTLIERS‟

Outlier values per domain were determined from the probability plots. At CRM, outliers were capped

during the point variography stage but remained uncapped in the kriging/estimation stage. Values are

capped in the variography stage to ensure that the population variance relationship was modelled and

not unduly influenced by outlier values. In the kriging stage, the outliers not being capped would affect

the final estimate value, but grid overriding would be applied if necessary, rather than altering the

dataset population. Grid overrides were not undertaken in the Mineral Resource estimation.

11.2.6 HARD AND SOFT ESTIMATION BOUNDARIES

When the domains were delineated, the data along the domain boundaries were reviewed and the

domain soft boundaries adjusted accordingly. The soft boundaries affect the estimation/kriging

process, in that data lying between the hard and soft boundaries for a specific domain are allowed to

be incorporated in the kriging of the relevant blocks. This ensures continuity of the grade relationships

across domains. The soft boundaries extend for approximately 10m beyond the domain hard

perimeters.


12. SAMPLING METHOD AND APPROACH

The sampling method, analysis and data verification for CRM exploration programmes have been

discussed in the Technical Report compiled by RSG Global, titled “Restated Technical Report” dated 10

November 2006 and filed on SEDAR 21 March 2007. 15 drillholes were examined and physical drillholes

were compared to the drill logs. RSG Global concluded that the sampling methods and approach

applied are considered to be to industry standard and appropriate for the deposit. The following

summaries have been extracted from the aforementioned report.

12.1 CORE DRILLING SAMPLING PROCEDURE

The core was laid out in the core trays with the low point of the layering in the middle of the core on

the visible side. The sample positions were marked off to ensure that appropriate and representative

samples were taken. The samples at the base and top of the UG2 CL overlap by 1cm to 2cm into the

adjacent footwall and hanging wall, respectively. The borehole depth of the top and bottom of the

sample are measured and recorded and a unique sample number assigned to each sample. Core dips of

contacts were measured and recorded to enable a true sample width to be calculated for each sample.

The number of breaks present within a sample is recorded and an assessment

The sample is then split along its length through the low point of the contact dip by an appropriately

trained assistant using a diamond saw. Half core samples were cut to length (15cm to 25cm), and then

half of the core bagged and sealed with a sample ticket number both inside the bag and attached to

the outside. The remaining half core was marked and kept in the core tray for future reference.

Suitable sample bags of heavy-duty plastic were used and discarded. Sample bags were not recycled.

Samples were dispatched with the necessary documentation to the respective laboratories for analysis.

These laboratories included Rio Tinto, SGS Lakefield Research Africa (Pty) Ltd (“SGS”), the Analytical

Services Division at Mintek (“Mintek”), Inspectorate M&L (Pty) Ltd, and Impala Mineral Process

Laboratories(“Impala Laboratory”).

12.2 RC SAMPLING

No detail of RC sampling is available and therefore none of these drillholes was used in the Resource

estimate.

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13. SAMPLE PREPARATION, ANALYSIS AND SECURITY

13.1 TESTING LABORATORIES

Sample preparation and analysis has historically been conducted by various laboratories, including Rio

Tinto, SGS, Mintek, Inspectorate M&L (Pty) Ltd, Impala Platinum Limited, Impala Laboratory and

Golden Dumps.

Approximately 80% of the cored intersections‟ analyses were undertaken at Impala Laboratory. A

record is kept of the specific assay technique used for each sample.

It should be noted that the majority of CRM samples were tested prior to accreditation of the

laboratories used. The following laboratories are currently accredited by the South African National

Accreditation System (“SANAS”) as detailed in Table 18.

The current boreholes including all the 2008 and 2009 boreholes were typically prepared and analysed

by Set Point Laboratory (SANAS T0223), which is accredited for PGE analyses.

Table 13-1: Details of Accreditation for Testing Laboratories Used for Sample Analyses

Testing Laboratory

Testing Laboratory Number

Complies With

Original Date of Accreditation

Type of Tests Accredited for

Impala Lab T0177 ISO/IEC 17025:2005

October 2002 The determination of Pt., Pd, Rh,Ru, Ir and Au by Direct Leach / Microwave Dissolution (ICP/Wet Chemistry) in converter matte. The determination of Pt, Pd, Rh, Ru, Ir and Au byNickel Sulphide collection / Microwave by dissolution (ICP/Wet Chemistry) for concentrate. The determination of Nickel, Copper, Cobalt and Chromium in concentrate type samples by ICP spectroscopy in concentrate. Analysis of Platinum, Palladium and Rhodium (GFAA/Wet Chemistry) including moisture

SGS T0169 ISO/IEC 17025:2005

December 2002 Determination of Au by Lead Fusion followed by Atomic Absorption Analysis or Gravimetry for carbon, ores, rocks, soils and metallurgical products. Determination of Au, Pt and Pd by Lead Fusion followed by ICP-OES and ICP-MS for ores and rocks.

33

Determination of Pt, Pd, Rh, Ru and Ir by Nickel Sulphide Fusion followed by ICP-OES for soils and metallurgical products.

Mintek T0042 ISO/IEC 17025:2005

May 1995 The collection and determination of PGM’s (Pt, Pd, Rh, Ru, Ir, Au) using Nickel Sulphide as the collector for ores and concentrates. The collection and determination of gold and the total PGM +Au using Lead as the collector for ores and concentrates. The determination of Individual PGM’s (Pt, Pd Rh, Ru, Ir, Au i.e. 6E) after Nickel sulphide collection by Fire Assay and measurement by ICP-OES for ores and concentrates. The determination of Pt, Pd, and Au (3E) by ICPOES after lead collection by Fire Assay for ores and concentrates. The determination of Pt, Pd, Rh and Au (4E) after lead collection by Fire Assay followed by high pressure sealed tube dissolution and measurement by ICP-OES for ores and concentrates.

Set Point T0223 ISO/IEC 17025:2005

August 2003 The determination of Pt, Pd and Au by Fire Assay and ICP

Notes: Only accredited tests applicable to the testing for PGE’s and gold have been reported in the above table.


Two of the assay procedures used by Impala Laboratory are described in the table which follows. The

other laboratories would have used similar techniques. The most important difference would be the

cupellation time used in the analytical procedure. A comprehensive comparison of all the data is

required to ensure that all the data used for the Mineral Resource estimates are comparable.

Table 13-2: Analytical Procedures used by Impala Laboratory

Fire Assay Method Individual Platinum Group Metal (IPGM) Method Borehole samples were routinely assayed only using the fire assay method and the 3PGE+Au concentration (g/t) determined. The pulps were dried under a space heater for 90 minutes, and then crushed and pulverised using a spindle pulveriser. Size analysis indicates that the pulverising results in 60% passing 100 mesh (150μm). The samples were then fluxed using a twin stream. The twin stream ensures that duplicate samples are processed in different batches. Most laboratory duplicate samples are analysed in the same batch. The laboratory uses ready-made flux from a supplier. 10mg of silver nitrate was added as a co-collector, which is removed at the high temperatures. The samples were fused in a fusion furnace at 1,200°C for 55 minutes. The buttons were then deslagged followed by low cupellation - ±1

The IPGM (Individual Platinum Group Metal) method of analysis was used to determine the prill split, i.e. Pt, Pd, Rh and Au. However, it is not routinely performed due to the cost (some 10 times more expensive than fire assay). IPGM concentrations are derived by fluxing the sample with nickel-sulphide, digesting it in a microwave and performing a direct leach. Concentrations are read on the Inductive Coupled Plasma Spectrometer (ICP-OES). With this technique round robin analysis of smelter plant, concentrator and monthly composites are routinely performed by other laboratories. Borehole composite samples (the entire UG2 CL) are also analysed using the technique for crosscorrelation of 3PGE+Au derived from the fire assay technique.

34

hour at 1,100°C in a muffle furnace. The beads were then hammered and placed in the high temperature (1,360°C) furnace for 6 hours (that is since June 2001 and between 1978 and December 1985). Boreholes drilled between December 1985 and May 2001 were assayed using a 20 minute fire assay technique and a correction factor was applied to bring them in line with the 6 hour procedure. Lead washes were done at two-hour intervals (the adding of lead shot). The prills were then weighed on a microbalance and the g/t (ppm) calculated. The methodology for the fire assaying of underground samples and borehole samples was the same. Each section sampled underground could be considered a ‘borehole’ section.

The methods of analysis, sample handling and preparation procedures are considered appropriate for

the Mineral Resource estimate. However, insufficient validation and verification have been undertaken

to demonstrate that the recorded assay values from the various assay techniques and laboratories are

equivalent. It was previously recommended by RSG Global that this should be undertaken by CRM and

consolidated into a single database.


In order to reconcile the drillhole sample and underground section sample analyses, there has been an

effort to consolidate the Impala drillhole assay database into a uniform format based on 5PGE+Au. This

has not yet been conducted on the East Plats drilled boreholes. The 5PGE+Au format was not used in

the Resource estimation.

This is a complex task as various assay techniques have been applied to samples from the several

drilling campaigns (e.g. 20 minute lead cupellation fire assay as opposed to the standard 6-hour fire

assay). Few laboratories perform a 6-hour high-temperature lead cupellation fire-assay. Impala Lab and

CRM laboratory have been consistently using this procedure for some 30 years. Other laboratories

routinely perform 2 hour high temperature cupellation. For this reason inter-laboratory check assays

show poor comparison between the various assay methods.

Comparison of 5PGE+Au and 3PGE+Au assay results is conducted by means of weighted compositing of

drillhole pulps (i.e. motherhole and deflections) followed by replicated analyses (i.e. individual PGM

grades derived from assaying drill composite samples as mill samples and assaying in duplicate and

triplicate). This value is then compared with the fire assayed 3PGE+Au value. A limitation was

identified with respect to the older data where neither core nor pulps were available for re-assay. The

conversion is based on older data available for the individual PGM proportions.

Various factors were determined based on the applicable drilling campaign. The factors to convert

3PGE+Au to 5PGE+Au ranged from 1.18 to 1.4.

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14. DATA VERIFICATION

14.1 QUALITY CONTROL MEASURES & VERIFICATION (“QA/QC”)

The following section details the QA/QC for the fire assay method used by Impala Laboratory.

The Impala Laboratory is well designed with dust extractors furnaces, and timers. Samples are logged

for each step of the process on a laboratory information management system (“LIMS”). An accredited

company checks balances every three months.

The twin stream batch process resulted in two assays for each borehole sample being submitted.

Samples for which PGM values do not agree within 10% are re-assayed. This method provides some

confidence that the variability in the assay technique has been minimised. However, it does not

necessarily provide adequate assurance or control for the assays. Round Robins, in which samples are

sent to a number of different laboratories for analysis, have been undertaken to check laboratory

accuracy. The laboratory standard samples were assayed before and after the borehole samples and

used to identify assay problems. Issues identified using standards indicate problems in certain sample

batches. In those cases, the drillhole samples are re-assayed. Although this method of QA/QC may not

be ideal, it should be adequate for in-mine geological purposes. In a typical month Impala Laboratory

would process 8,000 samples and on average 25,000 determinations - from twin stream, repeats, etc.

The precision and accuracy of the laboratories was not monitored. The analyses undertaken did not

include the field insertion of internal blanks, standards or duplicates. The data therefore rely on the

QA/QC work undertaken by the laboratories which was considered appropriate at the time.

14.2 LIMITATIONS ON VERIFICATION

It is our opinion that acceptable verification was undertaken on the drillhole database by CRM staff.

However, little original borehole logs and core and assay laboratory sheets are available to verify the

data. In addition, the level of detailed stratigraphic coding is limited on the historical drillhole logs. As

little underground data has been digitally captured to date, the underground data was not reviewed

and used in the Mineral Resource estimation.

Due to the fact that the deflections were non-directional, the co-ordinates of the deflections were

moved 0.5m away from the mother hole progressively.

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15. ADJACENT PROPERTIES

The property is bounded to the east by Xstrata‟s Eland PGM mine and to the west by development

properties held by Impala Platinum, Anglo Platinum and Lonmin. The Eland mine is in operation.

Lonmin‟s operating Marikana mine is approximately 15 km to the west.

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16. MINERAL PROCESSING AND METALLURGICAL TESTING

The Crocodile River Mine is in operation. See section 23 for a description of the processing plant.

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17. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1 STANDARD RESOURCE AND RESERVE REPORTING SYSTEM

All the aspects of these estimates are compliant with the specifications embodied in SAMREC and CIM

Instrument NI 43-101. The Mineral Resource pertains to 4E grades (g/t) at a cut-off of 200cm.g/t. The

cut-off was determined using the current PGE commodity prices, as well as optimistic improvements on

current operating costs and mining factors and forecasts. The Mineral Reserve is based on current

operating parameters.

17.2 COMMENT ON RESOURCE AND RESERVE SUBSETS The Resources reported are inclusive of Reserves.

17.3 COMMENT ON INFERRED RESOURCES Based on the data spacing and kriging efficiencies of the estimates, and a number of other parameters,

the Mineral Resource estimate for the Resource Area was deemed to fulfil the requirements of being

Measured, Indicated and Inferred Mineral Resources. The Inferred Resources were limited to

approximately the variogram range and not the lease area limits.

17.4 KEY ASSUMPTIONS, PARAMETERS AND METHODS OF RESOURCE CALCULATION

Although the structure of the area is relatively complex, the UG2 CL at CRM is defined as a tabular ore

body across the Project Area, which allowed a two dimensional (“2D”) approach to the Mineral

Resource estimation. Simple kriging method of geostatistical estimation was employed based on the

data spacing for the borehole data. The estimation process utilised the 4E metal content (cm.g/t) and

thickness (CW). A total of 411 accepted UG2 CL intersections were available for the Mineral Resource

estimation. CRM Project was subdivided into four domains. The average distance between boreholes is

between 100m and 350m. The metal grade for 4E was estimated from metal content and thickness. The

tonnages were adjusted for dip and geological losses (faults, dykes, potholes, IRUPs).

17.4.1 VARIOGRAPHY

Variograms are an essential tool for investigating the spatial relationships of samples. Variograms for

metal content 4E (cm.g/t) and channel width (“CW”), Resource Width and metal grade (4E g/t) were

modelled. Anisotropy in each area was investigated, although all variograms are deemed best

represented by omni-directional models. The tables below summarise the variogram model parameters

for the different domains areas, used in the kriging process. In 2008, the variograms compiled for the

Channel Width / In-Situ model were used in the Resource Cut model, as data with zero grade has to be

added to the Resource Cut database, which would skew the data away from its population

characteristics. In 2009, the variograms for the Resource cut were computed and modelled with the

39

Resource cut data and not borrowed from the In-Situ variograms. This resulted in minor changes to the

Mineral Resource estimates as stated in November 2008.

Variograms were modelled using point data with outliers removed. The nugget value was modelled

between 40 and 55% of the total sill/population variance for each domain and parameter. This value,

although high for a platiniferous reef, demonstrates the short-range variability of the data and/or

represents the effect that the mother-to-deflections have on the modelling of the variability. The

nugget value applied to the data was deemed not representative of the UG2 CL of the Western Limb of

the BIC, so further investigations were undertaken to establish more representative nugget values for

the domains. This entailed generating point variograms using a lag length of 1m and number of lags 5 in

order to counteract the effect of the mother hole and deflection holes. The average nugget:sill ratio

was calculated. The data was further regularised onto a 50X50m (X,Y) grid in order to decluster the

mother to deflections holes. Variograms were computed for each of the parameters (CW, cm.g/t and

g/t) for the entire data set (no domaining) and the nugget:sill ratios computed for this declustered

dataset. This nugget:sill ratio of the final variograms to be used in kriging was aligned to the average

nugget:sill ratio obtained from the point variograms on the 1m lag lengths and the declustered

nugget:sill ratio. The derived ratios were more in line with what is expected of the UG2 CL of the

Western Limb of the BIC. If the nugget:sill ratio demonstrated by the original experimental point

variograms was deemed representative, the nugget:sill ratio was maintained and not overridden.


Figure 17-1: In-Situ Point Variograms – Per Domain

Domain 1 cm.g/t Domain 2 cm g/t

41

Domain 3 cm.g/t Domain 4 cm g/t

42

Domain 1 CW Domain 2 CW

43


44

Figure 17-2: Resource Cut Point Variograms – Per Domain

Domain 1 cm.g/t Domain 2 cm.g/t

45

Domain 3 cm.g/t Domain 4 cm.g/t

46


47


48

Domain 1 SG Domain 2 SG

49

Domain 3 SG Domain 4 SG


Table 17-1: Variogram Parameters – In-Situ Model

Reef Parameter Domain Sill Nugget Percentage

Sill 1 Range X Str 1

Range Y Str 1

Range Z Str 1

Range X Str 2

Range Y Str 2

Range Z Str 2

% m m m m m m

UG2CW CMGT 1 19283.95 20.87 79.65 190 190 1 307 307 1

UG2CW CMGT 2 29066.45 24.94 71.71 286 286 1 541 541 1 UG2CW CMGT 3 25338.99 25.09 62.54 215 215 1 604 604 1 UG2CW CMGT 4 280456.86 25.12 83.73 116 116 1 398 398 1 UG2CW CW 1 661.13 22.52 78.36 108 108 1 391 391 1 UG2CW CW 2 4325.65 28.66 81.79 181 181 1 551 551 1 UG2CW CW 3 538.11 30.16 48.52 402 402 1 1317 1317 1 UG2CW CW 4 773.77 15.93 76.41 218 218 1 542 542 1

Table17- 2: Variogram Parameters – Resource Cut Model

Reef Parameter Domain Sill Nugget Percentage

Sill 1 Range X Str 1

Range Y Str 1

Range Z Str 1

Range X Str 2

Range Y Str 2

Range Z Str 2

% m m m m m m

UG2MC CMGT 1 17275.72 29.38 82.25 325 325 1 627 627 1

UG2MC CMGT 2 24345.52 30.19 83.91 263 263 1 1192 1192 1 UG2MC CMGT 3 23739.76 34.36 83.94 205 205 1 723 723 1 UG2MC CMGT 4 26417.30 30.28 85.56 187 187 1 578 578 1 UG2MC CW 1 142.10 20.07 74.43 322 322 1 621 621 1 UG2MC CW 2 56.96 30.29 89.32 236 236 1 1053 1053 1 UG2MC CW 3 22.91 19.49 55.00 386 386 1 644 644 1 UG2MC CW 4 112.09 25.89 67.45 187 187 1 595 595 1 UG2MC SG 1 0.12 19.94 63.18 509 509 1 799 799 1

UG2MC SG 2 0.01 32.96 56.90 343 343 1 664 664 1

UG2MC SG 3 0.09 19.64 55.00 212 212 1 1022 1022 1

UG2MC SG 4 0.03 7.89 55.00 586 586 1 665 665 1

17.4.2 LOCAL MEAN

The simple kriging process uses a local or global mean as a weighting factor in the kriging process. For

this estimation, all blocks within a specific domain have been assigned a global mean for that domain.

The means were determined from a 400X400m block model size. The data is regularised into 400X400m

blocks and the arithmetic average is calculated, hence when reviewing the table below, the values are

all individual averages and not derived from one another.

Ordinary kriging balances the kriging weights to one without the use of a local/global mean. Due to the

continuous nature of the mineralization and the close relationship of the mean to the domains grade

profiles, ordinary kriging would not derive any greater benefit over simple kriging. Inverse distance

squared is a rudimentary method of estimation which does not take into effect the spatial variability of

the domain, and the power used is subjective, and is hence not used in this estimation process.

The means determined for CRM domains are tabulated below:- Table17-3: 2009 Means – Simple Kriging

Domain In-situ Content

Channel Width

Resource Width

Resource Width 4E Content (cm.g/t)

Resource Width SG (t/m3)

1- Crocette 575 133 156 591 3.48

2- Zandfontein 599 130 154 618 3.73

3- Kareespruit 608 129 152 630 3.49

4- Maroelabult 622 136 156 612 3.75

Technical Report Update 51

17.4.3 BLOCK MODEL DEVELOPMENT

CRM UG2 CL block size was selected after consideration of the sample spacing and long-term overview

of the ore body. The block size utilized was 100m X 100m X 1m in the X, Y and Z dimensions

respectively. Typically the drillhole spacing varied between 100m to 350m; therefore the block model

size of 100X100 represents a stable block size for the data. A 2D seam model was used and corrected

for dip.

17.4.4 GRADE ESTIMATION

The following applies to CRM Resource area and was undertaken using Minesoft‟s „MT-EST‟

geostatistical program. Full reef composite mining cut and channel width values (4E content (cm.g/t),

grade (4E g/t) and CW (cm) have been interpolated into a 2D block model. The block model size of

100X100m was determined from the data spacing, variance relationships („stationarity‟ principal) and

the continuous nature of the ore body. Simple kriging estimation technique was used, based primarily

on data spacing. SG was estimated via kriging for the Resource Cut model based on the fact that the

Resource Cut included hangingwall lithologies. Detailed checks were carried out to validate kriging

outputs, including input data, kriged estimates and kriging efficiency checks.

The following parameters were used in the kriging process:

Point data – metal content (4E cm.g/t) and grade (4E g/t) and thickness (Resource width and

channel width (cm));

100m x 100 x 1m block size (X,Y,Z);

Parent cell estimation;

Discretisation 5 x 5 x 1 for each 100m x 100m x 1m block;

First search volume – approximately 1.5 times the variogram range (m);

Minimum number of samples 8 in search volume one (SVOL1) (in order to incorporate data

outside the deflection range);

Maximum number of samples 40 in search volume one (SVOL1);

Second search volume 1.5 times first search volume;

Minimum number of samples: 4 (SVOL 2);

Maximum number of samples: 40 (SVOL 2);

Third search volume 2 times second search volume;

Minimum number of samples: 1 (SVOL 3);

Maximum number of samples: 20 (SVOL 3);

Interpolation method – Simple kriging.

Metal Grade (4Eg/t) was calculated from metal content and mining width/channel width

thickness;

The kriged estimates was post-processed (log-normal post-processing) in order to align the

model with the likely mining unit dimensions of 25X25 smallest mining unit (“SMU”) (X & Y),

after log normal distributions were determined for the metals.

In order to confirm the estimation strategy of using simple kriging over ordinary kriging, the ordinary

andsimple kriged estimates were reviewed. Based primarily on the wide spacing of the drillholes the

simple kriged estimates were used to estimate the Mineral Resource.


17.4.5 DIP DETERMINATION

Analysis of the regional UG2 CL horizon and drillholes renders a regional dip value of 17o for the

estimation domains. This dip value was used to correct the tonnage calculation, as estimation was

undertaken in 2D space

17.4.6 GEOLIGICAL LOSS

Geological losses apply to faults, dykes, potholes and IRUPs, etc, that eliminate the reef. The following

geological losses, supplied by CRM Geology Department in November 2008, were applied to the

tonnage:-

Table 17-4: Geological Loss Factor

Domain Geological Loss Value

Zandfontein 25%

Crocette 28%

Maroelabult 25%

Kareespruit 25%

17.4.7 SPECIFIC GRAVITY

The SG value utilized for the In-Situ estimation was supplied by the mine and is 3.85t/m3. Review of

the drillhole information which is some instances contained SG values rendered a value of 3.80t/m3 for

chromitite, although the range of values varied considerably. Where samples were missing SG

information a default value of 3.85t/m3 was inserted. With regards to the Resource Cut estimates, the

SG value was estimated via kriging and the estimated value per block was applied to the tonnage

calculations. The average of this estimated SG for the Resource cut model equates to 3.62t/m3, due to

dilution from the hanging wall lithologies.

On import generation of the reef composites, the samples were weighted with SG values.

17.4.8 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 Mineral Resource classification:

1. Sampling – quality assurance & quality control (QA/QC):

a. Measured: high confidence, no problem areas;

b. Indicated: high confidence, some problem areas with low risk;

c. Inferred: some aspects might be of medium to high risk.

2. Geological confidence:

a. Measured: high confidence in the understanding of geological relationships,

continuity of geological trends and sufficient data;

b. Indicated: Good understanding of geological relationships;

c. Inferred: geological continuity not established.


3. Number of samples used to estimate a specific block:

a. Measured: at least 4 data points within variogram range and minimum of twenty

one-metre composited samples;

b. Indicated: at least 3 data points within variogram range and a minimum of twelve

one-metre composite samples;

c. Inferred: less than 3 data points within the variogram range.

4. Kriged variance:

a. This is a relative parameter and is only an indication and used in conjunction with

the other parameters.

5. Distance to sample (variogram range):

a. Measured: at least within 60% of variogram range;

b. Indicated: within variogram range;

c. Inferred: further than variogram range.

6. Lower confidence limit (blocks):

a. Measured: less than 20% from mean (80% confidence);

b. Indicated: 20%–40% from mean (80%–60% confidence);

c. Inferred: more than 40% (less than 60% confidence).

7. Kriging efficiency:

a. Measured: more than 40%;

b. Indicated: 10–40%;

c. Inferred: less than 10%.

8. Deviation from lower 90% confidence limit (data distribution within Resource area

considered for classification):

a. Measured: less than 10% deviation from mean;

b. Indicated: 10–20%;

c. Inferred: more than 20%.

Specifically with regards to Crocette, the south western corner of the Resource area, bordering the

graben structure with Zandfontein, is classified as Indicated albeit a paucity of data. The Indicated

classification is based primarily that it satisfies the conditions where an understanding of the geological

relationship exists; based on sufficient data from the remaining geostatistical domain and the

behaviour of the UG2 CL, that is, continuity has been established. The Crocette mining area represents

a single geostatistical homogeneous domain and the south western corner is bordered by data, hence

with reasonable confidence the geological nature of the ore body can be assumed.

17.5 DETAILED MINERAL RESOURCE TABULATION

CRM Project Area comprises Measured, Indicated and Inferred Mineral Resources, according to the

specifications of the SAMREC Code and CIM Instrument NI 43-101. The Mineral Resources area stated at

a cut-off of 200cm.g/t 4E. Only the UG2 CL has been estimated. The Resource Cut Mineral Resources

are stated at a mining width of 1.5m, a realistic mining/stope width, based on production records.

Increases in the Mineral Resource could result from significant changes to the commodity prices and

modifying factors which impact on the cut-off grade currently used to state the Mineral Resource.

The Mineral Resource of the Project Area is classified into confidence categories, based on a number of

parameters, such as data validation and reliability, estimation parameters and geology. The Measured

Mineral Resource component was based on the aforementioned parameters in conjunction with being


extended approximately 100m from the mined out areas. The Inferred Mineral Resource was limited to

approximately the maximum variogram range, which equated to ~1.2km, or approximately 1km from

last line of drillhole data. The Mineral Resources as at the 30th April 2009 are summarised in the

following tables.


17.5.1 PGE+AU IN-SITU MINERAL RESOURCE:

The following table details the 3PGE+Au CRM In-Situ Mineral Resources estimated as at 30th May 2009:-

Table 17-5: 30th May 2009 In-Situ Mineral Resources: Cut-off of 200cm.g/t

Mineral Resource Estimate for CRM (30 May 2009) – In-Situ Resource – Cut-off 200cm.g/t

Measured Mineral Resource

Metal Concentration Metal Content

Dip Tonnes

(„000)

True Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)

Au (g/t) Pt:Pd:Rh:Au

(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 2,534 141 4.21 2.63 1.11 0.44 0.03 62.6 : 26.2 : 10.4 :

0.8

0.3430 0.2143 0.0904 0.0358 2,444

Maroelabult 17 1,835 137 4.55 2.81 1.24 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

0.2684 0.1658 0.0732 0.0277 1,770

Zandfontein 17 3,793 132 4.61 2.84 1.26 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

0.5622 0.3463 0.1537 0.0585 3,658

Total 17 8,162 136 4.47 2.77 1.21 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

1.1736 0.7264 0.3172 0.1221 7,872

Indicated Mineral Resource

Dip Tonnes

(„000)

True Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)


(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 7,425 135 4.34 2.70 1.16 0.45 0.03 62.6 : 26.2 : 10.4 :

0.8

1.0360 0.6445 0.2769 0.1074 7,162

Maroelabult 17 2,834 137 4.53 2.80 1.23 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

0.4128 0.2551 0.1121 0.0428 2,733

Zandfontein 17 21,646 131 4.60 2.83 1.26 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

3.2013 1.9695 0.8769 0.3340 20,878

Kareespruit 17 15,764 132 4.63 2.85 1.27 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

2.3466 1.4445 0.6437 0.2433 15,205

Total 17 47,669 132 4.57 2.81 1.25 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

6.9967 4.3136 1.9095 0.7276 45,978

Measured and Indicated Mineral Resource

Dip Tonnes

(„000)

True Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)


(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 9,959 136 4.31 2.68 1.15 0.45 0.03 62.6 : 26.2 : 10.4 :

0.8

1.3790 0.8588 0.3673 0.1433 9,606

Maroelabult 17 4,669 137 4.54 2.80 1.23 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

0.6812 0.4209 0.1852 0.0706 4,503

Zandfontein 17 25,439 131 4.60 2.83 1.26 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

3.7635 2.3158 1.0305 0.3926 24,536

Kareespruit 17 15,764 132 4.63 2.85 1.27 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

2.3466 1.4445 0.6437 0.2433 15,205

Total 17 55,831 133 4.55 2.81 1.24 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

8.1703 5.0400 2.2268 0.8497 53,850

Inferred Mineral Resource

Dip Tonnes

(„000)

True Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)


(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 4,976 133 4.34 2.70 1.16 0.45 0.03 62.6 : 26.2 : 10.4 :

0.8

0.6943 0.4320 0.1856 0.0720 4,799

Maroelabult 17 0 0 0.00 0.00 0.00 0.00 0.00 - 0.0000 0.0000 0.0000 0.0000 0


Zandfontein 17 43,605 130 4.61 2.84 1.26 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

6.4629 3.9815 1.7664 0.6729 42,058

Kareespruit 17 35,370 131 4.71 2.89 1.30 0.49 0.03 62.6 : 26.2 : 10.4 :

0.8

5.3561 3.2864 1.4783 0.5572 34,115

Total 17 83,951 131 4.64 2.85 1.27 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

12.5133 7.6999 3.4303 1.3021 80,973

Total Mineral Resource

Measured 17 8,162 136 4.47 2.77 1.21 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

1.1736 0.7264 0.3172 0.1221 7,872

Indicated 17 47,669 132 4.57 2.81 1.25 0.47 0.03 62.6 : 26.2 : 10.4 :

0.8

6.9967 4.3136 1.9095 0.7276 45,978

Inferred 17 83,951 131 4.64 2.85 1.27 0.48 0.03 62.6 : 26.2 : 10.4 :

0.8

12.5133 7.6999 3.4303 1.3021 80,973


The Mineral Resource tonnage grade profiles areas were tabulated at various cm.g/t cut-offs. The

tonnage grade profiles indicate that CRM UG2 CL ore body is very stable at a variety of cut-offs but

radically decreases in both tonnage and grade at high cut-offs (e.g., 500cm.g/t and above).

Figure 17-3: 4E- Grade Tonnage Curves for In-Situ UG2 CL – Zandfontein

4.64.654.74.754.84.854.94.9555.055.15.155.25.255.35.355.4

1,800,000

2,000,000

2,200,000

2,400,000

2,600,000

2,800,000

3,000,000

3,200,000

3,400,000

3,600,000

3,800,000

4,000,000

0 100 200 300 400 500 600

Gra

de

Ab

ove

Cu

t-o

ff (

g/t)

Ton

nag

e (

t)

Cut-off Value (cm.g/t)

Zandfontein Measured Resource Cut GTC SMU 25*25

Tonnage Grade

Figure 17-4: 4E- Grade Tonnage Curves for In-Situ UG2 CL – Crocette

4.2

4.3

4.4

4.5

4.6

4.7

4.8

4.9

5

1,000,0001,100,0001,200,0001,300,0001,400,0001,500,0001,600,0001,700,0001,800,0001,900,0002,000,0002,100,0002,200,0002,300,0002,400,0002,500,0002,600,000

0 100 200 300 400 500 600

Gra

de

Ab

ove

Cu

t-o

ff (

g/t)

Ton

nag

e (

t)


Crocette Measured Resource Cut GTC SMU 25*25

Tonnage Grade

Figure17-5: 4E- Grade Tonnage Curves for In-Situ UG2 CL – Kareespruit


4.64.654.74.754.84.854.94.9555.055.15.155.25.255.35.355.4

7,500,0008,000,0008,500,0009,000,0009,500,000

10,000,00010,500,00011,000,00011,500,00012,000,00012,500,00013,000,00013,500,00014,000,00014,500,00015,000,00015,500,000

0 100 200 300 400 500 600

Grad

e (g

/t)

Tonn

age

(t)

Cut-off value (cm.g/t)

Kareespruit Resource Cut Indicated GTCSMU 25*25

Tonnage Grade

Figure 17-6 4E- Grade Tonnage Curves for In-Situ UG2 CL - Maroelabult

4.54.554.64.654.74.754.84.854.94.9555.055.15.155.25.255.3

900,000

1,000,000

1,100,000

1,200,000

1,300,000

1,400,000

1,500,000

1,600,000

1,700,000

1,800,000

1,900,000

2,000,000

0 100 200 300 400 500 600

Gra

de A

bove

Cut

-off

(g/

t)

Tonn

age

(t)


Maroelabult Measured Resource Cut GTC SMU 25*25

Tonnage Grade

17.5.2 3PGE+AU RESOURCE CUT MINERAL RESOURCE:


The following table details the 3PGE+Au CRM Resource Cut Mineral Resources estimated as at 30th May 2009:-

Table17-6: 30th May 2009 Resource Cut Mineral Resources: Cut-off of 200cm.g/t Mineral Resource Estimate for CRM (30 May 2009) – Resource Cut – Cut-off 200cm.g/t

Measured Mineral Resource

Metal Concentration Metal Content

Dip Tonnes („000) True

Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)


(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 2,662 158 3.89 2.47 0.99 0.41 0.02 62.6 : 26.2 : 10.4 :

0.8

0.3329 0.2114 0.0847 0.0351 1,712

Maroelabult 17 2,074 156 3.99 2.52 1.03 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

0.2661 0.1680 0.0687 0.0280 1,334

Zandfontein 17 4,318 154 4.08 2.56 1.06 0.43 0.03 62.6 : 26.2 : 10.4 :

0.8

0.5664 0.3554 0.1472 0.0597 4,165

Total 17 9,054 156 4.00 2.52 1.03 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

1.1654 0.7348 0.3006 0.1228 7,210

Indicated Mineral Resource


Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)


(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 7,677 156 3.89 2.47 0.99 0.41 0.02 62.6 : 26.2 : 10.4 :

0.8

0.9601 0.6096 0.2444 0.1012 4,936

Maroelabult 17 3,133 158 3.83 2.44 0.97 0.40 0.02 62.6 : 26.2 : 10.4 :

0.8

0.3858 0.2458 0.0977 0.0403 2,015

Zandfontein 17 24,922 154 4.01 2.53 1.04 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

3.2131 2.0272 0.8333 0.3365 16,025

Kareespruit 17 16,198 152 4.13 2.59 1.08 0.43 0.03 62.6 : 26.2 : 10.4 :

0.8

2.1508 1.3488 0.5624 0.2239 15,623

Total 17 51,930 154 4.02 2.53 1.04 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

6.7098 4.2314 1.7378 0.7020 38,600

Measured and Indicated Mineral Resource


Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)


(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 10,339 157 3.89 2.47 0.99 0.41 0.02 62.6 : 26.2 : 10.4 :

0.8

1.2931 0.8210 0.3291 0.1363 6,648

Maroelabult 17 5,207 157 3.89 2.47 0.99 0.41 0.02 62.6 : 26.2 : 10.4 :

0.8

0.6518 0.4138 0.1664 0.0683 3,348

Zandfontein 17 29,240 154 4.02 2.53 1.04 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

3.7795 2.3826 0.9805 0.3962 20,190

Kareespruit 17 16,198 152 4.13 2.59 1.08 0.43 0.03 62.6 : 26.2 : 10.4 :

0.8

2.1508 1.3488 0.5624 0.2239 15,623

Total 17 60,984 154 4.02 2.53 1.04 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

7.8752 4.9663 2.0384 0.8247 45,810

Inferred Mineral Resource


Thickness

(cm)

3PGE+Au

(g/t)

Pt

(g/t)

Pd

(g/t)

Rh

(g/t)


(%)

3PGE+Au

Moz

Pt

Moz

Pd

Moz

Rh

Moz

Au

oz Crocette 17 5,227 155 3.80 2.42 0.96 0.40 0.02 62.6 : 26.2 : 10.4 :

0.8

0.6386 0.4067 0.1613 0.0672 3,361

Maroelabult 17 - - - - - - - - - - - - -


Zandfontein 17 50,210 154 4.01 2.53 1.04 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

6.4733 4.0842 1.6789 0.6780 32,286

Kareespruit 17 38,020 153 4.13 2.59 1.08 0.43 0.03 62.6 : 26.2 : 10.4 :

0.8

5.0484 3.1659 1.3202 0.5256 36,671

Total 17 93,457 154 4.05 2.55 1.05 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

12.1603 7.6568 3.1604 1.2708 72,318

Total Mineral Resource

Measured 17 9,054 156 4.00 2.52 1.03 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

1.1654 0.7348 0.3006 0.1228 7,210

Indicated 17 51,930 154 4.02 2.53 1.04 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

6.7098 4.2314 1.7378 0.7020 38,600

Inferred 17 93,457 154 4.05 2.55 1.05 0.42 0.02 62.6 : 26.2 : 10.4 :

0.8

12.1603 7.6568 3.1604 1.2708 72,318


Figure 17-7: Resource Classification Plot


Figure 17-8: Resource Classification Plot – Farm Boundaries

Figure 17-9: Resource Classification Plot – Farm Boundaries overlain on Topographic Map


17.6 PRILL SPLITS

In 2003, prill splits for 3PGE+Au were computed from available drillhole data. The prill split values

were re-estimated based on the drillhole data set utilised in this modelling exercise, and compare

favourably with the 2003 prill splits declared.

Table 17-7: Prill Splits

Prill 2003/2007 2008 2009

Pt 63.3% 62.6% 61.1%

Pd 27.7% 26.2% 28.5%

Rh 9.3% 10.4% 9.7%

Au 0.7% 0.8% 0.7%

Total 100% 100% 100%

17.7 RECONCILIATION

The descriptive statistics of the kriged grades were reviewed to ascertain that the distributions and

estimates correspond with the input data. This was achieved, in that the two datasets correspond to

each other.

Table 17-8: Kriged Estimates versus Input Point Data

Resource Area Block Model 4E Grade (g/t) Simple Krig

Drillholes 4E Grade Percentage Difference %

Crocette 4.32 4.33 -0.23

Zandfontein 4.61 4.63 -0.43

Kareespruit 4.71 4.68 +0.64

Maroelabult 4.56 4.61 -1.08

The descriptive statistics of the kriged grades and means used in the simple kriging equations were

reviewed to ascertain that the distributions and estimates correspond with the input data. This was

achieved, in that the two datasets correspond to each other.

When reviewing the above descriptive statistics in comparison to the original input data the population

distribution of the kriged estimates are smoothed, a characteristic of the kriging process. High-grades

are smoothed lower and low grades are smoothed higher. When reviewing the post-processed

estimates, the discrepancy noticed in the post processed results in Maroelabult was investigated and it

was established that the area comprises two grade areas, namely a low grade in the central portion

and a high grade area surrounding the low grade area. The low grade area constitutes the greater

percentage of the estimated area. The same applies to Crocette post processed results – the boreholes

suggest a high grade zone across the middle of the Resource area, which is carried through to the

kriged estimates from the drillholes and the global mean.


17.8 EFFECT OF MODIFYING FACTORS

No account of any modifying factors such as taxation, socio-economic, marketing or political factors have been taken into account.

17.9 TECHNICAL PARAMETERS AFFECTING THE RESOURCE DECLARATION

No mining or plant recoveries factors have been built into the Mineral Resource model. However, the

Resource cut Mineral Resource is calculated on a mining width of 1.5m.

17.10 METAL SPLITS FOR DECLARED RESOURCE

The UG2 CL at CRM is a polymetallic ore body. See section concerning Prill Splits. The metal splits have

been calculated from regression analysis on the original borehole data set. The regression of the

individual metals to the 4E grade was computed from the current borehole data set. Typically the

current boreholes were analysed for the four metals in comparison to the historical data which

contained very few prill analyses.

17.11 Overview of Modifying factors

Crocodile River mine uses certain factors that are applied to calculate the planned tons called for in

the budget. The following modifying factors indicated in the table were applied to the planning done

for the 2009 budget.

Table 17-9 : Crocodile River Mine modifying factors

17.12 Mine Call Factor

Crocodile River mine assumes 100% as a mine call factor for planning purposes. Data from the mine

indicate that the mine call factor calculated for August 2008 to December 2008 is at an average of

90.63% for CRM with Zandfontein at 91.28% and Maroelabult at 89.87%.

17.13 Resource Cut

With regards to the calculation of the Resource cut widths, after consultation with the Mine a Resource

cut of 1.5m was determined in 2008. A Resource cut is the width of the mineralised unit calculated on


ptimised geological cut based on historical and envisaged mining criteria. The width of 1.5m was

determined using the average channel width of the UG2 Main Band (~1.4m) and the average historical

mining widths achieved (1.52m).

17.14 Resource Classification

CRM Project Area comprises Measured, Indicated and Inferred Mineral Resources, according to the

specifications of the SAMREC Code and CIM Instrument NI 43-101. The Mineral Resources area stated at

a cut-off of 200cm.g/t 4E. Only the UG2 CL has been estimated. The Resource Cut Mineral Resources

are stated at a mining width of 1.5m, a realistic mining/stope width, based on production records.

Increases in the Mineral Resource could result from significant changes to the commodity prices and

modifying factors which impact on the cut-off grade currently used to state the Mineral Resource.

A detailed breakdown in terms of the resource categories in the life of mine plan was done for the

Crocodile River mine. The following table gives a summary of the resources in the mine plan.

Table 17 -10: Resource categories in Life of Mine Plan

The following step to calculate reserves is to apply the modifying factors as discussed. The following

table again gives a summary of the modifying factors. Firstly the major geological losses and unknown

geological losses and then the mining and pillar losses are applied which will give the resultant

reserves.


Table 17 -11: Crocodile River Mine modifying factors

17.15 Reserve Classification

The modifying factors discussed in Table 17-11 and the financial data in Section 23 were applied to the

Resources to calculate the total Reserves for CRM. The results after the factors have been applied to

the Resources is summarised in the table below.

Table 17-12: CRM Overall Reserves in life of mine plan

CRM carries forward an operating reserve of 9 months in total (10 months at the Zandfontein section

and 5 months at the Maroelabult section). This reserve is defined as the ore for which adequate

development work has been done (stope access, panel development, infrastructure such as gulleys and

haulways) such that it is available on short notice to mine as feed to the mill. It is defined as the

Proven Reserve. The balance of the material in Table 17-15 is a Probable Reserve as defined by the

drill data and kriging. Therefore the reserve classification is:


Table 17-13 Classified Reserves

Reserve Area Tonnage

(Kt)

Metal Concentration Contained Metal

4E (g/t) Pt (g/t) Pd (g/t) Rh (g/t) Au (g/t) 4E

Proven Kg K Oz

Crocette 0

Maroelabult 132 3.91 2.42 1.06 0.4 .04 515 17

Zandfontein 842 4.08 2.53 1.10 0.4 .04 3,435 110

Sub Total Proven 974 4.06 2.52 1.09 0.4 .04 3,950 127

Probable

Crocette 3,530 3.78 2.41 0.95 0.4 .02 13,340 429

Maroelabult 2,173 3.91 2.48 1.00 0.4 .02 8,487 272

Zandfontein 22,748 4.08 2.56 1.06 0.4 .03 92,896 2,987

Sub Total

Probable

28,451 4.03 2.54 1.04 0.4 .02 114,723 3,688

Total 29,425 4.03 2.54 1.04 0.42 0.03 118,673 3,815


18. OTHER RELEVANT DATA AND INFORMATION

Not Applicable


19. INTERPRETATION AND CONCLUSIONS

The Mineral Resources of CRM are classified as Measured, Indicated and Inferred Mineral Resources, the

Mineral Resources being tabulated at an average density (or specific gravity (“SG”)) of 3.85t/m3 for the

In-Situ estimation and at an average of 3.62t/m3 for the Resource cut and Reserve estimations. The

Mineral Reserves are classified as Proven and Probable and include dilution and other factors so that

the grades represent ore fed to the concentrator. Similarly Reserves are stated as mill feed grade. The

density for the UG2 on the Western BIC has a range of between 3.5 and in excess of 4.0t/m3. The SG

used for the tonnages for the Resource width was based on kriged SG values for each block. The

Resource Width includes hangingwall lithologies with different SG‟s to chromitite, hence the SG for the

Resource width is representative of the lithological units intersected. The Mineral Resource and

Reserve of the Project Area are classified into confidence categories, based on a number of

parameters, such as data validation and reliability, estimation parameters and geology.


20. RECOMMENDATIONS

A technical review of the data used in the Mineral Resource estimation has shown the quality of the

data to be adequate for the intended purpose. The historical drill results have limited Quality

Assurance and Quality Control (“QA/QC”) monitoring as does the database in which the data is stored.

It is recommended that CRM locate as many of the original logs and assay reports and check the data in

the database. Also the majority of the historical core should be located and centralised into one core

shed. Review of the stratigraphic coding of the historical drillholes should be undertaken to comply

with the latest stratigraphic coding of the UG2. All mother and deflection holes should be logged in

detail and assayed.

A review of the SG values should be undertaken and site specific and lithological specific SGs should be

determined. According to the geology department at the mine, the drillhole information from the

databases is stated as intersected width and needs to be corrected for dip. The drillhole intersections

were adjusted to represent corrected width for the Mineral Resource estimation. A concern is the data

was generated by various mining companies and hence may represent a data set with varying levels of

information.

It is recommended that CRM undertake extensive data gathering and verification of the historical

drillholes. Detailed stratigraphic coding must be undertaken on all available data. Research should be

undertaken on the effect the Transvaal Supergroup footwall units have had on the deposition of the

Upper Critical Zone (“UCZ”). The Transvaal Supergroup footwall units in the Brits section of the

Western BIC have extensively influenced the development of the Lower Critical Zone (“LCZ”), and

therefore may have influenced the formation of the UCZ of the BIC, with localised changes away from

the norm of the chromitite layers. Findings from this modelling may possibly refine and add value to

the current models available for this area. Further research as to the characteristics of the UCZ should

be conducted in order to compile a detailed stratigraphic/facies plan.


21. REFERENCES

“Crocodile River Mine – Restated Independent Technical Report – Effective date 10 November 2006” –

RSG Global. Posted on SEDAR 3 May 2007.

“Report on the Estimation of the Mineral Resources, Crocodile River Mine, Eastern Platinum Limited,

North West Province, South Africa, 2009”. Done for CRM by Minxcon.


22. DATE AND SIGNATURE PAGE

Signed

Brian Montpellier, P.Eng.

6 December 2010


23. ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON

DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

23.1 MINE DESIGN

The following table lists the mine design criteria for Crocodile River mine.

Table 23-1 : Mine design criteria


Rock Engineering

At CRM the current stoping is a depth of approximately 80m - 260m. The maximum stress is less than

8MPa. This is in a rock mass that has strengths of 190MPa - Norite, 133MPa - Pyroxenite and 55MPa -

Chromites referred to as the UG2 reef. Thus at the current depths at CRM the stress regime has very

little influence on the mining. As the mining deepens clamping forces will start to generate in the

hanging wall. This will assist in stabilizing the hanging wall.

Regional Support


In accordance with the mandatory Code of Practice for rock fall and rock burst the regional support

was designed according to the guidelines set out in the guideline for compilation of that specific code

of practice. A system of stable non yielding pillars is employed to carry the entire weight of the

overburden.

Barrier Pillars

The barrier pillars are designed to be indestructible and the minimum W:H ratio is 10.0. The minimum

pillar width will then be 10m. This will only be incorporated as the mining depth increases. The barrier

pillars are designed to clamp major geological features and ensure stability around major water zones.

To a depth of 100m below surface the pillars are designed with a safety factor of 2.0 preventing any

surface subsidence. These pillars forms part of the crown pillar design and no mining will be allowed in

the crown pillar.

In-stope Pillar and Support

In-stope grid pillars were introduced following a series of back-breaks, large scale falls of ground in the

past. The hanging wall beam that fell measured up to 3.5m thick. Drilling information depicts that this

parting is situated between 2.4m and 3.8m into the hanging and has very little cohesion. The major

function of the in-stope grid pillars is to limit open spans. The grid pillar dimensions is 4m x 7m (dip x

strike) and a 2m ventilation holing. These pillars situated below the gully. The in-stope support design

is to prevent small localized falls of ground and to assist the stability of the hanging wall beam. The

support in the panels is placed at 1.5m x 1.5m (dip x strike) using190mm to 220mm diameter elongates

which can safely carry the hanging wall beam to a fallout height of 4,5m. Tendons are installed into

the hanging wall to assist in the lamination process to form a beam.

Mining Spans

The panels are limited to a maximum length of 22,5m and all new raise lines are planned at a panel

length of 20m grid pillar to grid pillar. This will assist in reducing any risk of hanging wall beam failure.

Vertical Spacing between Excavations

All footwall drives are located in excess of 10m below any stoping activity. A 3m wide pillar is left at

all reef intersections to protect the travelling ways. At Maroelabult a 5m wide bracket pillar is left on

both side of the on-reef developments and no mining is allowed in those pillars. Shaft pillars are

designed using computer modelling ensuring a safety factor of at least 2.0.

Ventilation

Ventilation Planning Parameters

Mine design and planning on Crocodile River mine was done taking into account the following

ventilation planning factors.

Table 23-2: Ventilation Planning Parameters


Air Quantities Required

The ventilation requirements and capacities are sufficient to supply the working with ventilation as

required. Maroelabult uses 2 vertical ventilation shafts and the two portals to circulate air through the

operation. Ventilation shaft 1 with a diameter of 2.5m uses 4 x 75kW fans to extract 114m3/s of air

from the workings. Air is supplied through the portals at 110m3/s and Ventilation shaft 2 at 15m3/s.

Ventilation shaft 2 is 3.5m in diameter with a capacity of 170m3/s. At Maroelabult the air utilisation is

measured at 55.9% with the measured downcast quantities at 110m3/s and the upcast at 114m3/s.

Zandfontein is divided in to 4 ventilation zones. Zone 1 uses Decline 1 as an intake at 90m3/s and a

3.6m diameter ventilation shaft with 2 x 75kW fans extracting a required quantity of 120m3/s. Decline

2 supplies Zone 2 with 40m3/s with some additional air from a raise bore hole supplying 15m3/s.

Ventilation Crocodile River Mine Mining Report shaft 2 is a 3.6m diameter shaft from which 49.2m3/s of

air is extracted by 2 x 90kW fans. The shaft capacity is 120m3/s. Decline 3 supplies 70.4m3/s to Zone 3

air extracted from Zone 3 is extracted from a 3.6m diameter Ventilation shaft 3 at a quantity of

47.5m3/s by 2 x 90kW fans. In Zone 4 air is supplied from Decline 4 at 70.8m3/s and from the vertical

shaft which has a capacity of 500m3/s. Ventilation shaft 4 extracts 158m3/s from a 3.6m diameter shaft

with 6 x 75kW fans. In total Zandfontein utilises air at 36.9%. The measured total upcast quantity is

307.5m3/s and the downcast is 304.4m3/s.

Crocette in not in production yet so there is no data available for that operation. On the current

planning the quantity required for Crocette will be 160m3/s.

Refrigeration

Currently Crocodile River does not need refrigeration and can remove heat sufficiently from the

working with the current ventilation system.


23.2 PRODUCTION SCHEDULE

The production schedule was based on parameters from the mine design criteria using the modifying

factors discussed earlier in this document. Production from Crocodile River mine is expected to

increase until 2011. A relative steady state will then be maintained as Crocette starts producing and

Maroelabult nears the end of its life. The life of mine is dependent on accessing Zandfontein below

level 9. In the following figure it can clearly be seen what Zandfontein 10 – 15 Level adds to the life of

mine.

Figure 23-1: Life of Mine (m2)

In the table below the actual square meters for 2008 and the planned square meters for the next five

years can be seen, the years following that can clearly be seen in the graph in Table 23-3

Table 23-3: Equivalent stoping square meters 2008 - 2013

The production profile of the run of mine tons to the mill is illustrated in the Figure 23-2 . The

contribution by each section on the mine can clearly be seen in the graph. Note the increase in

production from Zandfontein this is due to the planned production ramp up to reach steady state of

production.


Figure 23-2: Life of mine - Tons

23.3 INFRASTRUCTURE

All surface infrastructure is easily accessible on well maintained gravel or tar roads. The mine property

is fenced and all access controlled by effective security.

Shaft

Zandfontein Section

Zandfontein is accessed by four separate declines from surface. There is also a recently re-equipped

vertical shaft which is not in full production yet. The declines 1 to 4 provide access from separate

portals on surface to different areas in the Zandfontein mining area. Decline 4 splits up into two

different declines underground, the split known as Decline 5. All declines are designed at the same

parameters which is trackless footwall declines going down at 9° and at 35⁰ below strike.

Maroelabult Section

Maroelabult is accessed by two 6m x 2m declines that are spaced 10m apart going down skin to skin at

approximately 9°.

Change House

The industrial change houses used with our underground visit was in a good hygienic condition and

purposely built to supply the employees with a comfortable and easily accessible washing and changing

area. The lamp room facilities seen on the visit was well maintained and managed and able to supply

the necessary well maintained equipment needed by underground employees.

Offices


The offices are well constructed buildings and easily accessible. The layout of the offices is such that

the various disciplines are located close to each other to ensure ease of communication and interaction

between these disciplines. Some offices are portable offices which are spaced in constructed in such a

way that operations can be managed from them. Offices in the workshops are also maintained to these

standards. The offices provide an effective base for operations at CRM.

23.4 CURRENT OPERATIONS

Maroelabult Mine

On the Maroelabult mine plan it can be seen that the planning for the total mining area has been done

for the life of the shaft. Although this has only been done in detail for the next two years the future

planning is accounted for in all planned development. The location of the two declines from surface is

indicated on the plan (Figure 23-3). Underground a third decline is added to the system as they

advanced past the bord and pillar section. The three decline system is also indicated on the plan,

indicating the planned decline development to access the remaining reef. The major geological

structures are also displayed on the plan.

Figure 23-3: Maroelabult Mine Plan


Access

Maroelabult utilises two 6.0 x 4.0m on reef trackless declines from surface going down skin to skin, one

of them is equipped with a conveyor. The other decline used for men to access the workings and it is

also used by the utility vehicles and LHD‟s. Underground a third decline is added to the system. The

layout is then a decline on either side of the belt decline. These declines are used for men and vehicles

as travelling ways. The declines will be developed to extend to the end of the mining area ensuring

access and a means to remove ore from the mine via the conveyor belt in the centre decline.

Development

All development at Maroelabult is done on reef. Trackless reef drives are developed on the reef horizon

to create the faces. Trackless loading bays from where the conveyor belts are loaded are also

developed on the reef horizon together with a Raise/Winze on reef to connect to the top or bottom

level with conventional mining.

Ore transport:

Faces are cleaned using scrapers into the reef drives. The 10t Scooptram LHD‟s then loads out the ore

and waste from the reef drives and tips it in one of the various loading bays developed along the

conveyor belt. A 7km conveyor system consisting of a 900mm belt running at approximately 400t/hr

then transports the ore and waste to surface.

Mining Method:

The stoping method used at Maroelabult is conventional breast mining. Advance strike gullies spaced at

21.5m are 1.5m wide and 2.8m deep and will be developed on reef. The panel width will be

maintained below 22.5m for stability and safety reasons. The panels will have 6.0m long by 4.0m wide

strike pillars with 2m holings between pillars, with the ASG maintained approximately 3m ahead of the

advancing stope face.

Equipment

The single boom drill rigs used at Maroelabult works mainly on the development of the reef drives and

loading bays. The 10t LHD‟s are uses to load ore and waste on the conveyor belts. There are also 7 light

delivery vehicles in operation at Maroelabult. The table below summarizes the equipment on

Maroelabult:-

Table 23-4: Maroelabult equipment

Maroelabult

Type Make Model Capacity Total

Drill rig Boart Longyear CE/LPD-1007 1



LHD Atlas Copco Scooptram 10t 13


Zandfontein Mine

Figure 23-3: Zandfontein Mine

From the Zandfontein mine plan it can be noted that the planning has been done for the total mining

area. Although this has only been done in detail for two years, the future planning is accounted for in

all planned development. The four decline portals can be seen on the northern end of the plan with

their associated declines accessing the working areas with the vertical shaft that has recently been re-

equipped. The potholes, dykes and faults can be seen clearly be seen for this area.

Access:

Zandfontein utilizes four separate declines from surface and a recently re-equipped vertical shaft

which is not in full production yet but will remove some of the ore and waste. The declines numbered 1

to 4 provide access from separate portals on surface to different areas in the Zandfontein mining area.

Decline 4 splits up into two different declines underground, the split known as Decline 5. All the

declines are trackless footwall declines going down at 9° and at 35⁰ below strike.

Development:

From the declines trackless footwall haulages are developed 25m vertically below the reef horizon.

From the haulage trackless cross cuts are developed that advance towards the reef horizon and stop

8m vertically below the reef horizon. A travelling way will then be developed up at 34° with

conventional mining, until it intersects the reef position. Box holes are developed by mechanized

inverse blasting. Development on the reef horizon consists of a step over on reef with conventional

mining and a Raise/Winze on reef to connect to the top or bottom level with conventional mining.

Ore transport:


Cleaning from the face is done by conventional face and gully scrapers. The ore is then loaded from the

cross cut by the 10t Scooptram LHD‟s into 20t haul trucks. The haul trucks then transport the ore to 2

level were the ore is transferred to a 1050mm conveyor belt transporting the ore a further 1.5km to a

900mm surface conveyor taking the ore to the processing plant. Waste is transported by haul trucks

from the workings through the declines to surface were it is dumped on the waste rock dump.

Mining Method:

The stoping method used at Zandfontein is conventional breast mining. Advance strike gullies spaced at

21.5m are 1.5m wide and 2.8m deep and will be developed on reef. The panel width will be




Equipment

Zandfontein uses 20 ton haul trucks to transport ore and waste. These trucks remove waste to surface

using the various declines. Ore is tipped on 2 Level and removed with a conveyor belt. Zandfontein

uses 10t Scooptram LHD‟s in various applications on the mine. The total number of machines used on

Zandfontein is listed in the table below. Being a trackless operation there is a fleet of 53 light delivery

vehicles on site at Zandfontein.

Table 23-5: Zandfontein Equipment

Zandfontein

Type Make Model Capacity Total

Drill rig Atlas Copco RB281 1

Drill rig Atlas Copco RB282 5

Drill rig Rham Rham 1

Dump truck Sandvik EJC 522 20t 2

Dump truck Fermel Fermel 20t 2

Dump truck Atlas Copco MT2000 20t 2

Dump truck Atlas Copco MT2010 20t 16

LHD K9 K9-4 10t 1

LHD Atlas Copco Scooptram 10t 11

LHD Sandvik Toro 600 10t 2

Crocette Mine


Crocette is currently still in the development phase. The declines have been developed to reef

intersection, and development work will continue through 2011 before any ore can be produced.


Figure 23-4 Crocette Mine Plan

Access:

When in production Crocette will be utilizing two trackless footwall declines as the primary access with

the drives going down at 9°and 34⁰ below strike.

Development:

Development for Crocette is planned to be as follows. From the declines, trackless footwall haulages

will be developed. From the haulages, trackless cross cuts will be developed that will advance towards

the reef horizon. The trackless reef drives will be developed on the reef horizon this development

includes the trackless loading bays used by the LHD to load ore and waste on the conveyor belt. To

connect the levels a Raise/Winze is developed on reef conventional mining.

Ore transport:

Reef is loaded by LHD into a haul truck and tipped into a box hole at the footwall declines positions

and then transported out of the mine via dump trucks.

Mining Method:

The stoping method that will be used at Crocette is conventional breast mining. Advance strike gullies

spaced at 21.5m are 1.5m wide and 2.8m deep and will be developed on reef. The panel width will be





Equipment

Currently the only equipment used for Crocette is 2 light delivery vehicles as the shaft is not in

operation yet.

23.5 MINE SERVICES

Water

Crocodile River mine uses fresh water stored in 5 different dams on surface for all the water utilized on

the mine. This water from the dams is used underground in the various applications needed and in the

change houses where the employees use it to wash and it is also used as the drinking water on the

property. All used and dirty water is pumped to the settling dams where all the solids settle and the

resulting cleaner water is pumped through the water treatment plant.

The treated water, which is now again safe to be consumed by humans is pumped to the surface dams

and reused in the same application as mentioned above. Crocodile River mine uses approximately 0.5t

of water for 1.0t of rock. All excess clean water is pumped back into the Crocodile River which can also

be used should there be a shortage of water on the mine as a source of water.

Figure 23-5 : Schematic diagram of water reticulation on CRM

Pumping

Underground water management is done by utilising a number of Flight and RNE pumps. Currently the

system is under pressure and a project to construct a water dam on 3 Level with an improved pumping

arrangement should improve the situation at Zandfontein. Currently approximately 60 –80l/sec of

water is pumped from Zandfontein vertical shaft, this water comes mainly from the many water

fissures underground.

Compressed Air


Currently Zandfontein uses decentralized air compressors on surface to supply the necessary

compressed air for the conventional mining methods used. Diesel and electric compressors supply the

compressed air. The diesel compressors used mainly to safe on the consumption of electricity. At the

Zandfontein a total of 18 compressors are in use. These compressors are 11 x 1500cfm, 2 x 1000cfm

and 5 x 900cfm compressors running on diesel. On Maroelabult there are 6 compressors, 3 supplying

1500cfm and the other 3 supplying 900cfm of compressed air.

Electricity

Crocodile River mines Zandfontein section utilized 20MVA supplied by Eskom, since the 20% reduction

in power consumption the mine now only use 18MVA. This saving in power consumption was possible by

implementing diesel compressors to supply compressed air to the underground workings. Maroelabult

receives 6.6kV from the local Municipality. Ventilation

The mine is ventilated through a series of vertical raise bore holes with exhaust fans serving the east

and west ventilation districts of each of the two underground sections. Refrigeration is not warranted

due to the shallow depth of the Crocodile River mine.

23.6 PLANT DESIGN

The plant is designed to treat 170,000t per month of UG2 ore. The ore is conveyed to the plant from

underground and surface sources and is crushed in a three stage crushing circuit before feeding two

2,500 crushed ore silos. The milling and flotation circuit has two independent streams designed to treat

85,000t per month each. Each of the circuits is MF2 configuration. Th epromary mills are in closed

circuit with 650 micron screens and hydro-cyclones. Tails from the primary flotation circuit as well as

the tails from secondary cleaners are pumped to a spiral plant for the upgrading and the removal of

chrome concentrate. Secondary regrind mills which further reduce the product size to a grind to

liberate the sulphides containing PGMs and other valuable metals. The milled slurry then goes through

a bank of rougher cells which recovery the PGMs and base metals, the tails report to scavenger cells

and the concentrates report to cleaner cells. The final concentrate is settled and dispatched to

Impala‟s smelter for final smelting and refining. High grade is transported by tanker while low grade is

filtered and transported by road in concentrate bins. The concentrate contains 150 to 200 g/t of PGMs

and between 1.5 and 3.0% Cr2O3 and is treated under the terms and conditions set out in the

concentrate take-off agreement with IRSL. The overall plant design is robust. Maintenance and

housekeeping standards are good.

Plant Performance

The tonnage treated during 2010 has averaged 105,000 t/m. The year to date PGE recovery is 79% and

has varied between 78% and 82%. The PGE head grade has shown a steady downward trend from 4.1g/t

in January to 3.95 g/t in October 2010

The chrome spiral plant contributes to the lowering of Cr2O3 in the PGE final concentrate but also

generates revenue from the sale of metallurgical and chemical grade chrome concentrate.

A tailings retreatment plant was commissioned in 2007 and construction of a chrome recovery plant

was completed in 2008. The chrome tailings treatment plant was previously operated by a contractor,

Tailings Technologies. This has however changed and Eastplats personnel are now managing and

operating the plant. Retreatment of the old slimes dam will in future be carried out in the tailings


scavenging plant. Sampling and Evaluation Plant samples are taken by automatic sampler from the feed

belts by means of a go-belt sampler and automatic samplers are used of the pulp streams. There is a

comprehensive laboratory situated in the plant which is convenient for sample preparation and assay.

Equipment is of a high standard and should have a positive effect on the quality of analysis.

23.7 CROCODILE RIVER MINE CAPITAL

The capital budget on Crocodile River mine in 2009 was ZAR 223 M ($32 M). This is less than that

planned in 2010 which is ZAR 356 M ($51 M) due to the economic constraints imposed in 2009 due to

the financial meltdown. A breakdown of the budget capital for the next thee years is also given in the

table.

Table 23-6: CRM capital

CRM - Capital Summary 2008 2009 2010 2011 2012 2013 Total Crocette 80.33 Expansion: Crocette Phase 1 3.50 1.53 113.74 175.70 120.67 Total Concentrator 48.22 Expansion: Zandfontein Phase 1 to level 9 62.55 177.18 171.88 243.38 654.01 Total Drilling 45.09 Expansion: Development 93.70 105.06 158.31 132.25 132.25 Total Kareespruit 27.12 Expansion: Drilling 5.00 - - - - Total Maroelabult 34.52 Expansion: Concentrator 12.11 38.17 - - - Total Protection Services 7.44 Sustaining: Crocette Phase 1 - - - 21.09 14.88 Total IT 15.92 Sustaining 46.48 34.92 102.62 108.51 118.79 Total Shared Services 40.80 Total Enviromental 21.75 Total Projects Offices 13.49 Total Zandfontein 491.40 Total CRM 826.10 223.35 356.85 546.56 680.93 1,040.60 +

23.8 CROCODILE RIVER WORKING COST

The cash cost on Crocodile River mine is broken down in certain elements, namely Labour, Stores,

Sundries, Utilities and variable electricity. The figure below shows an increase in the total cash cost of

the Crocodile River mine‟s cost profile. This is attributed to various reasons including a ramp up in

production from Zandfontein and Crocette. The line indicates that the working cost over the life of

mine is constant, showing that the increased costs can be attributed to the increased production.


Figure 23-6 Cash Cost breakdown by element with working cost profile

Comparing the detailed cash cost breakdown of the budget and actual figures from 2008 and 2009 the

following can be noticed. The future planning and budgeting of costs is based on the historical cash

cost values of the previous year‟s production. The 2009 actual costs were taken from the mine cost

centres. In determining the cash costs in the budget some of the cost centres used are grouped

together and not broken into the same level of detail as the actual costs from the cost centres. The

cost estimations is deemed to be accurate as the costs is based on actual costs incurred by Crocodile

River mine over a production year. The following table shows a summary of the cash cost breakdown

indicating 2008 actual costs and costs for the next three years of the life of mine budget, the rest of

the costs is can be seen in the graph in Table 23-7.

Table 23-7: Actual and Budget cash cost summary to 2013 (Rm)

The previous figure showed the working cost (R/t) as a line and in Figure 23-7 and in the following

figure a breakdown of the working cost can be seen. The contribution from the various business units

towards the working cost can clearly be seen in the figure. Zandfontein contributing the biggest portion

later in the mine‟s life as the other operations are phased out and the bulk of the production is carried

by Zandfontein.


Figure 23-7: Working cost breakdown by area

The following table shows a total working cost for 2008 and then the breakdown of the working cost

per business unit from 2009 to 2013.

Table 23-8 – Working Cost

23.9 ENVIRONMENTAL CONSIDERATIONS

A full EIA has been produced for the site and accepted by the Ministry of Mines. An environmental

guarantee of ZAR 18 million ($2.7 million) has been set aside to cover potential close out liability,

reclamation and rehabilitation.


23.10 TAXES AND ROYALTIES

There has been a new mining royalty introduced in March 2010 which applies a 1% royalty against gross

revenues. Prior to tax being paid, the total value of the capital expenditures can be written off

against taxable income. After that, ordinary income taxes are assessed at a flat rate of 28% of taxable

income.

23.11 MARKETS

The concentrator produces a flotation concentrate containing 6E PGMs plus nickel and copper. The 6E

refers to the 5 PGM elements: platinum, palladium, rhodium, ruthenium and osmium, plus gold. This

concentrate is relatively refractory and must be smelted and refined in specialist facilities. Eastplats

currently sell the concentrates from the Crocodile River mine to Impala Refining Services (IRS).

PGMs apart from being precious metals also have industrial use, especially as catalysts. The largest

consumer of PGMs is the automotive industry and the price of the PGM group is dependent to a large

degree on the health of the automotive industry


CERTIFICATE of

QUALIFIED PERSON

I, Brian Alfred Montpellier, P. Eng. do hereby certify that:

1. I am Vice President, Projects for Eastern Platinum Ltd 250 – 1075 West Georgia Street,

Vancouver B.C., Canada V6E 3C9.

2. This certificate applies to the report titled “Technical Report Update on the Crocodile River

Mine” dated 6 December 2010.

3. I graduated with a BSc (honours) degree in Chemistry from the University of British Columbia in

1970. I am a member in good standing of the Association of Professional Engineers and

Geoscientists of the province of British Columbia and am also a member of the South African

Institute of Mining and Metals. I have worked in the mining industry for a total of 37 years since

my graduation from university. I am a Qualified Person for the purposes of the Instrument.

4. I was present on site initially in January 2007 and visited the site quarterly, in visits of one to

two days duration. The latest visit was 30 November 2010.

5. I am responsible for the preparation of all the sections of the technical report titled „‟Technical

Report Update on the Crocodile River Mine‟‟ and dated 6 of December 2010 (the “Technical

Report”).

6. I am not independent of the issuer applying all of the tests in section 1.4 of National Instrument

43-101.

7. I have had prior involvement with the property that is the subject of the Technical Report. The

nature of my prior involvement is planning and managing project development work on behalf

of Eastplats.

8. I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been

prepared in compliance with that instrument and form.

9. As of the date of the report, to the best of my knowledge, information and belief, the

Technical Report contains all scientific and technical information that is required to be

disclosed to make the Technical Report not misleading.

Dated at Vancouver, B.C. this 6th day of December, 2010.

Signed

Brian Montpellier, P. Eng.

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