EX-99.1 2 a10-15597_4ex99d1.htm EX-99.1

Exhibit 99.1

TECHNICAL REPORT

ON THE

TASIAST GOLD MINE

ISLAMIC REPUBLIC OF MAURITANIA

FOR

RED BACK MINING INC

Author:

Hugh Stuart, B.Sc., M.Sc, MAusIMM

Date: August 10, 2010

August 2010

TABLE OF CONTENTS

1.0	SUMMARY		7

1.1	INTRODUCTION		7
1.2	LOCATION		7
1.3	OWNERSHIP		7
1.4	GEOLOGY		8
1.5	MINERALISATION		8
1.6	EXPLORATION CONCEPT		8
1.7	STATUS OF EXPLORATION		8
1.8	MINERAL RESOURCES		9
1.9	MINERAL RESERVES		9
1.10	DEVELOPMENT AND OPERATIONS		10
1.11	ENVIRONMENTAL ISSUES		11
1.12	CONCLUSIONS		12
1.13	RECOMMENDATIONS		12

2.0	INTRODUCTION AND TERMS OF REFERENCE		13

2.1	TERMS OF REFERENCE		13
2.2	THE PURPOSE OF THIS REPORT		13
2.3	PRINCIPAL SOURCES OF INFORMATION		13
2.4	QUALIFICATIONS AND EXPERIENCE		13
2.5	ABBREVIATIONS		13

3.0	RELIANCE ON OTHER EXPERTS		15

4.0	PROPERTY DESCRIPTION AND LOCATION		16

4.1	THE TENEMENT AREA		16
4.2	LOCATION		18
4.3	OWNERSHIP		18
4.4	ROYALTIES, PAYMENTS, AGREEMENTS AND ENCUMBRANCES		18
4.5	ENVIRONMENTAL LIABILITIES		18
4.6	PERMITS		18
4.7	SURVEYING OF THE LICENCES		19
4.8	BACKGROUND INFORMATION ON MAURITANIA		19
4.8.1		Background	19
4.8.2		Economy	20
4.8.3		Currency and Foreign Exchange	20
4.8.4		Geography and Infrastructure	20
4.8.5		Administration	21
4.8.6		Political Situation	21
4.8.7		Mining Law in Mauritania	21

5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		24

5.1	SITE TOPOGRAPHY, ELEVATION AND VEGETATION		24
5.2	ACCESS		24
5.3	CLIMATE		24
5.4	SURFACE RIGHTS		24
5.5	SITE DEVELOPMENT		24
5.6	WATER SUPPLY		24
5.7	POWER SUPPLY		25
5.8	MINE PERSONNEL		25
5.9	TAILINGS STORAGE FACILITY		25
5.10	ADMINISTRATION AND PLANT SITE BUILDINGS		25
5.11	ACCOMMODATION		25
5.12	COMMUNICATIONS		26
5.13	MOBILE EQUIPMENT		26
5.14	SECURITY		26
5.15	GOODS AND CONSUMABLES		26

6.0	HISTORY		27

6.1	OWNERSHIP HISTORY		27
6.2	EXPLORATION HISTORY		27

6.2.1		Exploration Work Prior To 1993	27
6.2.2		1993 - 1996 European Development Fund - OMRG	27
6.2.3		1997 - 2001 NLSD	28
6.2.4		2003 Midas Gold plc	29
6.2.5		2003 GeomaqueE. - Defiance Mining Corporation	29
6.2.6		Rio Narcea	30
6.3	RESOURCE HISTORY		30
6.4	PRODUCTION HISTORY		30

7.0	GEOLOGICAL SETTING		31

7.1	REGIONAL GEOLOGICAL SETTING		31
7.2	PROPERTY GEOLOGY		32

8.0	DEPOSIT TYPES		36

9.0	EXPLORATION		38

9.1	INTRODUCTION		38
9.2	EXPLORATION METHODS USED		38
9.3	INTERPRETATION OF EXPLORATION INFORMATION		38
9.3.1		Introduction	38
9.3.2		Soil Sampling	38
9.3.3		Geological Mapping	38
9.3.4		Trenches	39
9.3.5		Drilling	39
9.3.6		Ground and Airborne Geophysics	39
9.3.7		Data Reliability	39

10.0	DRILLING		40

10.1	INTRODUCTION		40
10.2	1996 TO 1999 NLSD DRILL PROGRAMME		40
10.3	2003 DEFIANCE DRILL PROGRAMME		40
10.4	2004 DEFIANCE DRILL PROGRAMME		41
10.5	2007 RIO NARCEA DRILL PROGRAMME		41
10.6	RED BACK MINING DRILL PROGRAMMES		41

11.0	SAMPLING METHOD AND APPROACH		43

11.1	NLSD		43
11.2	2003-2004 DEFIANCE AND 2007 RIO NARCEA		43
11.3	2007 AND 2008 RED BACK		44
11.4	BULK DENSITY SAMPLING		44

12.0	SAMPLE PREPARATION, ANALYSES AND SECURITY		45

12.1	1996-1999 NLSD		45
12.2	2003-2004 DEFIANCE AND 2007 RIO NARCEA		45
12.3	2007 RED BACK		46
12.4	SUMMARY		46

13.0	DATA VERIFICATION		47

13.1	INTRODUCTION		47
13.2	HISTORICAL DATA VERIFICATION		47
13.2.1		NLSD Analytical Data	47
13.2.2		Defiance Analytical Data	48
13.2.3		Red Back Data Verification	54
13.2.4		2009 Red Back QAQC Data	55
13.2.5		SGS KAYES	55
13.2.6		SGS TASIAST	61
13.2.7		SGS MORILLA	66
13.2.8		Conclusion	71

14.0	ADJACENT PROPERTIES		72

15.0	MINERAL PROCESSING AND METALLURGICAL TEST WORK		73

15.1	MINERAL PROCESSING		73
15.1.1		Metallurgical Test Work	73
15.2	REFINING		75

16.0	MINERAL RESOURCES AND MINERAL RESERVES		76

16.1	REVISED MINERAL RESOURCE ESTIMATE		76
16.1.1		Resource Data Sets	76
16.1.2		Block Model	76
16.1.3		Geological Interpretation	76
16.2	DATA PREPARATION AND TREATMENT		77
16.2.1		Mineralisation Wireframes	77
16.2.2		Compositing	77
16.3	EXPLORATORY DATA ANALYSIS		78
16.4	SPATIAL CONTINUITY ANALYSIS		78
16.4.1		Measures of Spatial Continuity	78
16.4.2		Directional Controls on Gold Mineralisation	78
16.5	RESOURCE ESTIMATE		79
16.5.1		Indicator Kriging for Recoverable Resources	79
16.5.2		Indicator Kriging Parameters	80
16.5.3		Bulk density modeling	80
16.6	BLOCK SUPPORT ADJUSTMENT (VARIANCE ADJUSTMENT)		80
16.6.1		The Variance Adjustment	81
16.6.2		Shape of the Block grade Distribution	81
16.6.3		The Information Effect	81
16.6.4		Variance Adjustments Applied to the Tasiast Model	82
16.7	RESOURCE CLASSIFICATION		82
16.8	MINERAL RESOURCE STATEMENT		83
16.9	MINERAL RESERVE ESTIMATE		83
16.10	NON-GEOLOGICAL FACTORS RELEVANT TO RESOURCES AND RESERVES		85

17.0	OTHER RELEVANT DATA AND INFORMATION		86

17.1	RECONCILIATION		86

18.0	REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION PROPERTIES		87

18.1	MINING OPERATIONS		87
18.2	PROCESS RECOVERIES		88
18.2.1		CIL	88
18.2.2		Dump Leach	89
18.2.3		Recoveries	89
18.3	TAXATION		90
18.4	CAPITAL & OPERATING COST ESTIMATES		90
18.4.1		Capital Expenditure	90
18.4.2		Operating Costs	90
18.5	ECONOMIC ANALYSIS		91
18.5.1		Cash Flow Forecast	91
18.5.2		Sensitivity Analysis	94
18.6	PAYBACK		94
18.7	MINE LIFE		94

19.0	INTERPRETATION AND CONCLUSIONS		95

20.0	RECOMMENDATIONS		96

21.0	REFERENCES		97

22.0	DATE AND SIGNATURE		98

23.0	CERTIFICATES OF AUTHORS		99

LIST OF TABLES

Table 1-1: Resource Statement	9
Table 1-2: Reserve Statement	10
Table 1-3: Production History	10
Table 2-1: Listing of Abbreviations	14
Table 4-1: Tasiast Mine Licence Summary	16
Table 4-2: Licence Coordinates	17
Table 4-3: Operating Permits	19
Table 4.8.7-1: Summary of Exploration and Mining Licence Terms, Rules and Mechanisms	22
Table 4.8.7-2: Fiscal Provisions— Mining Code (“MC”) and Model Mining Convention (“MMC”)	23
Table 6.2.3: Summary of NLSD Exploration Activities at Tasiast	29
Table 7.2: Table of Formations, Aouéouat Greenstone Belt, Mauritania	33
Table 10.1: Tasiast Drill Summary	40
Table 13.2.2-.1: Summary of Basic Statistics for Duplicate Assays	48
Table 13.2.2-2: Comparison of Assay Data of RC Samples	52
Table 13.2.2-3: Comparison of Assay Data of Samples Collected by SNC	53
Table 13.2.4: Sample statistics by laboratory	55
Table 13.2.8-1: Tasiast Resource QAQC Summary	71
Table 16.2.1: Tasiast resource modeling domains	77
Table 16.8.1: Mineral Resource estimate for Tasiast	83
Table 16.9-1: Tasiast Open Pit Ore Reserve Statement	83
Table 16.9-2: Breakdown of Ore Reserve by Processing Route	84
Table 17.1-1: Reconciliation: Exploration Resource Model to Grade Control Model	86
Table 17.1-2: Milled Material to Block Model Reconciliation	86
Table 18.1: Summary of Pit Design Parameters	87
Table 18.2.3: Metallurgical Recoveries	89
Table 18.7.1: Life of Mine Capital Expenditure	90
Table 18.7.2: Life of Mine Operating Cost	90
Table 18.8.1: Life of Mine Financial Model	92
Table 18.8.2: Sensitivity Analysis	94

LIST OF FIGURES

Figure 4.2-1: Location	18
Figure 7.1: Geology of the West African Craton	32
Figure 7.2-1: Property Geology of the Piment Zone—Tasiast Permit Area	34
Figure 7.2-3: Tasiast - Main Exploration Trends	35
Figure 8: Tasiast Mineralisation	36
Figure 13.2.2-1: Scatter Plot of the Tasiast Original versus Duplicate Assays (n=1904)	49
Figure 13.2.2-2: Thompson & Howarth Precision Plot of All Duplicate Assay Data	49
Figure 13.2.2-3: Scatter Plot of the Tasiast Original vs Howe Duplicate Check Assay	50
Figure 13.2.2-4: Scatter Plot of the Tasiast Original vs Howe Duplicate Check Assays (cut to 20 g/t Au)	51
Figure 13.2.2-5: Scatter Plot of the Tasiast Original vs Howe Duplicate Check Assays (cut to 10 g/t Au)	51
Figure 13.2.2-6: Precision Plot of the Tasiast Original vs Howe Duplicate Check Assays	52

1.0 SUMMARY

1.1 Introduction

On August 2, 2007, Red Back Mining Inc (Red Back) completed the acquisition from Lundin Mining Corporation of a 100% interest in the Tasiast gold mine located in Mauritania, West Africa, through the purchase of Tasiast Mauritanie Limited (TML) and Tasiast Mauritanie Limited S.A. (TMLSA). The Acquisition was completed pursuant to a share purchase agreement, dated as of August 1, 2007, among Tasiast Holdings S.A., Rio Narcea Tasiast Luxembourg S.A., Lundin and Blokvink Holdings B.V. (a subsidiary of Red Back now known as Red Back Mining B.V.).

1.2 Location

The Tasiast Permit Area is located in north-western Mauritania, approximately 300 km north of the capital Nouakchott and 250 km southeast of the major city of Nouâdhibou. The Tasiast Permit Area falls within the administrative purview of the Inchiri and Dakhlet Nouâdhibou Districts and comprises one Permis d’Exploitation (PE) and five individual and contiguous Permis de Recherche Mineral (PRM) totalling 4,806 km2 in area. A well equipped camp consisting of concrete block construction and container-type buildings at the Tasiast site serves as the base from which all development and exploration activities are conducted.

1.3 Ownership

Upon the completion of regional reconnaissance exploration program carried out by the Office Mauritanien de Recherches Géologiques (“OMRG”) in 1996, the areas within and around the Tasiast area were made available to third parties. As a result, Normandy LaSource Development Ltd. (“NLSD”) (a subsidiary of Normandy Mining Ltd. of Australia) acquired the Tasiast area. In 2001, NLSD was acquired by Newmont Mining Corporation creating Newmont LaSource. Midas Gold plc (“Midas”) was incorporated in England and Wales in 2002 for the purpose of acquiring Newmont LaSource’s assets in Mauritania including the Tasiast Project as well as various other permit areas. Midas completed its acquisition of the Tasiast Project from Newmont LaSource on April 1, 2003 as per the terms of a 2002 sales agreement. In April 2003, Geomaque Explorations Inc. (“Geomaque”) announced the acquisition of Midas and the extension of a loan allowing Midas to complete its acquisition of the Tasiast Project. The merger of Geomaque and Midas ultimately created a new entity - Defiance Mining Corporation. In June 2004 Rio Narcea acquired Defiance and took ownership of the Tasiast Project.

Red Back acquired the Tasiast project from Lundin Mining Corporation in August, 2007 following Lundin’s acquisition of Rio Narcea Gold Mines, Ltd.

The Tasiast mine and the mining lease on which it is based are owned 100% by TMLSA. TMLSA is owned by the Corporation through Red Back Mining B.V., a Dutch company. The surrounding exploration licences are held 100% by a second subsidiary, TML.

Tasiast mine achieved commercial production in January 2008.

1.4 Geology

The Tasiast Permit Area encloses the 4 main Precambrian greenstone belts of the western compartment of the Reguibat shield, known as the Tasiast-Libzenia Domain. The Reguibat Shield consists of a core of west to east accreted, north-south trending Archean and Lower Proterozoic terranes dating from 2.6 Ga in the domal basement gneisses to 1.78Ga as the peak metamorphism in the metavolcanics. The package was cratonised at the end of the Eburnean Event and has been stable since 1700Ma. The Reguibat shield is bound on all sides by Pan African orogenic belts and covered in the south by the extensive intra-cratonic sediments of the Taoudeni Basin.

The Tasiast Mine Area is underlain by the Aouéouat greenstone belt, a 70 km long by 15 km wide N-S trending belt.

Gold mineralization in this area is similar to that found in Canadian Archaean terranes such as the Abitibi Greenstone Belt. Deposit types that occur within the Tasiast area consist primarily of gold-bearing shear hosted deposits.

Exploration work has identified north-south trending, intensely sheared, gold-mineralized zones extending over a strike length of 10km.

Mineralisation dips to the east between 30 and 60 degrees and is hosted by a folded sequence comprising Banded Iron Formation (BIF), Felsic Volcanics and intermediate to mafic volcaniclastics.

The structure is interpreted as a broad, regional antiform that is cored by a felsic volcanic unit approximately 200 m wide at the present surface elevation with structurally controlled mineralisation sitting close to either margin and within the core of the fold..

1.5 Mineralisation

Gold mineralisation is associated with structurally controlled late, discrete faults and shears, quartz-veining and silica-flooding, within all rock types.. There is ubiquitous bleaching due to prograde epidotization and retrograde chloritisation and later overprinting hydrothermal sericitization. Generally late, brittle, quartz-carbonate veining is overprinted by the remobilization and replacement of magnetite by secondary pyrrhotite. Coarse visible gold is common within the micro-fracturing and veinlets associated with the main shears.

1.6 Exploration Concept

The Tasiast deposit lies within an extensive gold system that is largely under-explored. The deposit is open along strike and at depth. Tasiast is the first mine in the highly prospective 70 kilometre long by 15 kilometre wide north-south trending Archaaen age Aoueouat greenstone belt, which is geologically similar to other Archaen greenstone belts in the world that host major gold deposits. The Tasiast properties cover a 70 kilometre strike length of the Aoueouat greenstone belt and exploration will therefore comprise of definition drilling combined with reconnaissance exploration to locate further areas for follow up.

1.7 Status of Exploration

Whilst the known open pit resources have been defined sufficiently by drilling for Mineral Resource estimation, further drilling is required to fully define the strike and depth extent of the mineralised zones at Tasiast.

1.8 Mineral Resources

In February 2010, based on additional drilling by the Company, Hellman and Schofield (H&S) re-calculated the Mineral Resource estimate as at December 2009.

Open pit resources were calculated from the 31st December 2009 mining surface and are reported based on a cut-off grade dependent on material type. Gold grades for the reported open pit resource have been determined using Multiple Indicator Kriging (MIK) based on block dimensions of 15m (east) x 25m (north) x 5m (elevation) and using a selective mining unit of 3m (east) by 5m (north) by 2.5m (elevation). Gold estimation and model blocks were constrained within geologically derived wireframes.

The summarised mineral resource statements are reported in accordance with Canadian National Instrument 43-101, Standards of Disclosure for Mineral Projects of December 2005 (National Instrument 43-101) and the classifications adopted by CIM Council in August 2000. The resource classification is also consistent with the Australasian Code for the Reporting of Mineral Resources and Ore Reserves of December 2004 as prepared by the Joint Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia (JORC). The resource statement is summarised Table 1.1.

	Cut-	Measured			Indicated			Measured + Indicated			Inferred
Zone	Off	Mt	Au g/t	Moz	Mt	Au g/t	Moz	Mt	Au g/t	Moz	Mt	Au g/t	Moz
Oxide	0.2	20.99	0.83	0.56	22.23	0.70	0.50	43.22	0.76	1.06	6.28	0.6	0.12
Fresh	0.5	41.23	1.55	2.05	70.41	1.50	3.40	111.64	1.52	5.45	26.52	1.4	1.18
Total		62.22	1.30	2.61	92.64	1.30	3.90	154.86	1.30	6.51	32.8	1.24	1.30
All ore types	1.0	27.55	2.17	1.93	41.87	2.13	2.86	69.42	2.15	4.79	13.38	2.1	0.91

Table 1-1: Resource Statement

1.9 Mineral Reserves

On the basis of the December 2009 resource AMC Consultants Pty Ltd (AMC) estimated the Ore Reserves for the Tasiast mine. The Ore Reserve estimate was based on the H&S December 2009 resource block model depleted to the 31st December 2009 pit surfaces. The statement includes ROM and low grade stockpiles.

The Ore Reserve Statement set out below has been determined and reported in accordance with National Instrument 43-101, and the classifications adopted by CIM Council in August 2000. Furthermore, the reserve classifications are also consistent with the ‘Australasian Code for Reporting of Mineral Resources and Ore Reserves’ of December 2004 as prepared by the Joint Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia (JORC).

Summary Ore Reserve as at December 2009

	Tonnes	Au	In situ Au
Classification	(Mt)	(g/t)	(Moz)
Total Proven	49.4	1.36	2.17
Total Probable	61.5	1.40	2.77
Total Stockpile	4.3	0.68	0.09
Total	115.2	1.36	5.03

Table 1-2: Reserve Statement

Breakeven cut off grades based on $US800/oz gold, metallurgical recoveries, process cost and royalty inputs to the open pit optimization are estimated as CIL oxide 0.75g/t, CIL fresh 0.81g/t, Dump leach oxide 0.10g/t.

1.10 Development and Operations

The Tasiast gold mine was officially opened by the President of Mauritania, His Excellency Sidi Mohamed Ould Cheikh Abdallahi, on July 18, 2007. Red Back completed the purchase of Tasiast on August 2, 2007. Commissioning of the Tasiast plant continued through 2007 with commercial production declared in January 2008.

Mining commenced in April 2007. Annual production is summarised in Table 1-3.

	Tonnes Milled	Grade	Gold Produced
Year	(Mt)	(g/t)	(‘000 ozs)
2009	1.69	3.10	161
2008	1.49	3.10	140
2007	0.22	4.77	21

Table 1-3: Production History

Based on positive results from the ongoing resource conversion drill program, an expansion of the Tasiast plant from 1 million tonnes per annum (mtpa) to approximately 2.5 mtpa was implemented with operation commencing in the fourth quarter of 2009.

All ore and waste is mined via conventional, open pit mining methods. The operation utilises selective mining techniques to separate ore and waste. The mining fleet is a combination of 120 tonne hydraulic excavators loading 90 tonne trucks. Provision has been made for drilling and blasting all primary materials.

The treatment plant flowsheet is based on three stage crushing, ball milling, pre-leach thickening, and a six stage CIL circuit. Gold is recovered by an elution circuit with electro-winning of the gold onto wire wool cathodes. The loaded wire wool is smelted to produce a final bullion product. A gravity circuit will be installed in the grinding circuit to prevent build up of coarse gold during processing of the higher grade primary ores.

Following a positive testwork program in 2008, low grade oxide material is now being processed by dump leaching.

At the end of 2009 TMLSA employed 454 people, of which 64 are expatriates.

The mine power is provided by three 2.7MW HFO generator sets with eight 1.0MW diesel generators in reserve have been commissioned.

The source of mine water supply is located 60 km west of the mine and is comprised of a semi-saline underground aquifer, which is exploited by twenty wells. Water is pumped to the mine site through two HDPE water pipelines to a 35,000m3 storage facility at the mine site. The final pumping capacity of the system is estimated to be 14,000m3 per day. Reverse osmosis water treatment plants provide drinking water and reagent mixing water. In the first quarter of 2010 pipe failure in the new 500mm pipeline reduced pumping capacity, adversely impacting the irrigation of the dump leach pads. The failing sections of the pipeline are being replaced with work expected to be completed in the fourth quarter 2010, at which time full dump leach irrigation rates will resume.

Gold produced at the mine site is shipped from site, under secured conditions, to a refining company. Under pre-established contractual instructions, the refiner delivers the refined gold directly to an account held by Tasiast with an international financial institution. Once received at the financial institution, the refined gold is sold with proceeds automatically credited to a Tasiast bank account. Gold is sold in the market at spot as Tasiast is not a party to any contract for the sale of its gold.

In the period 2008 to 2010, Tasiast’s profits may be exonerated from income taxes under a Mining convention signed in 2006 with the government of Mauritania. Tasiast’s future profits, once the exoneration from income taxes ceases, will be subject to tax based on a 25% income tax rate. Amortization and depreciation of Tasiast’s past and future capital projects can be applied using the established tax rates of amortization to reduce the income otherwise subject to tax.

The mine is estimated to have a life of a further 20 years.

1.11 Environmental Issues

An Environmental Impact Study (EIS) was undertaken by SNC Lavalin (SNC, 2004) as part of the full feasibility study for the project. The EIS report was submitted to the Ministry of Mines and Industry (MMI) now the Ministry of Petroleum and Mines (MPM) on 31st May 2004 and subsequently approved by the Director of Mines and Geology on 12th April 2005.

Following the publication of new legislation, namely Decrees No. 2004-094 and No. 2007-105, TMLSA was requested to retrospectively update the application to comply with new EIS legislation. The following updates were completed:

· Terms of Reference for the EIS;

· A conforming Environmental Management Plan (EMP);

· Formal public inquiry;

· Rehabilitation and Closure Plan; and

· Non-technical summary of the EIS aimed at the public and decision-makers.

TMLSA commissioned Scott Wilson, an international environmental and engineering mining consultancy, which was completed in the third quarter of 2009.

The structure of the Environmental Management Plan (EMP) has been agreed with the Ministry of the Environment. A key objective of the EMP is to achieve the management targets set against TMLSA’s environmental and social objectives.

The preliminary decommissioning, rehabilitation and closure plan has been presented to the Ministry. The plan focuses on decommissioning and demolition of the plant, followed by

rehabilitation.

In June 2009 TMLSA submitted an EIA to cover the following new developments:

· a new Tailings Storage Facility (TSF II) — The existing TSF I is nearing capacity and a second facility is required to cater for the increased reserve;

· a Dump Leach Facility (DLF) — A new process whereby gold is recovered from low grade ore through cyanide leaching; and

· an Expansion of Borefield — The increased milling rate and DLF have increased the water demand, which will be met by expanding the existing water borefield.

The associated construction and operating permits were received in August 2009.

1.12 Conclusions

The Tasiast mine has been developed and is now in operation. All necessary permits and licenses are in place.

Continued exploration is further expanding the resources and reserves and the Company is continuing to expand the scope of the Tasiast operation.

1.13 Recommendations

Exploration should continue throughout the Tasiast area in order to fully define the open pit potential of the permits.

2.0 INTRODUCTION AND TERMS OF REFERENCE

2.1 Terms of Reference

The author has been requested by Red Back to prepare a Technical Report on the Tasiast Gold Mine in the Islamic Republic of Mauritania to reflect updates in the Mineral Resources and Ore Reserves.

The report complies with National Instrument 43-101 and the Resource and Reserve classifications adopted by CIM Council in August 2000. The report is also consistent with the “Australasian Code for Reporting of Mineral Resources and Ore Reserves” of December 2004 (the Code) as prepared by the Joint Ore Reserves committee of Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia (JORC).

All monetary amounts expressed in this report are in United States of America dollars (US$) unless otherwise stated.

2.2 The Purpose of this report

The purpose of this report is to update the May 8, 2009 technical report in light of a revised Mineral Resource and Ore Reserve estimates.

2.3 Principal Sources of Information

This report has been updated using the following information:

· Resources:		Hellman & Schofield Pty Ltd.
· Reserves:		AMC Consultants Pty Ltd.

2.4 Qualifications and Experience

The primary author of this report is Mr. Hugh Stuart, who is a professional geologist with 22 years experience in the exploration and evaluation of mineral properties internationally. Mr. Stuart is a Member of the Australasian Institute of Mining and Metallurgy (AusIMM). Mr Stuart is Vice President Exploration of Red Back Mining Inc and is a Qualified Person as defined by National Instrument 43-101. The author has been involved with Tasiast since 2007 and is fully acquainted with all aspects of the Tasiast mine including exploration history, geology, metallurgy, mineral resource and ore reserve estimations and project development.

Mr Stuart has visited the Tasiast Gold Mine on numerous occasions, the most recent being in June 2010.

2.5 Abbreviations

A full listing of abbreviations used in this report is provided in table 2-1 below.

Abbreviations List

$	United States of America Dollars	km	Kilometres
“	Inches	kWhr/t	Kilowatt hours per tonne
µ	Microns	l/hr/m2	Litres per square metres
3d	Three dimensional	LM2	Labtechnics 2kg mill
AAS	Atomic absorption spectrometry	M	Million
AMC	Australian Mining Consultants	m	Metres
Aa	Gold	Ma	Million years
Bcm	Bank cubic metres	MIK	Multiple indicator kriging
BFP	BFP Consultants	ml	Millilitres
CC	Correlation coefficient	mm	Millimetres
Cfm	Cubic feet per minute	MMI	Mobile metal ion
kw	kilowatts	Moz	Million ounces
CIC	Carbon in column	Mtpa	Million tonnes per annum
CIL	Carbon in leach	N (Y)	Northing
cm	Centimetre	NaCN	Sodium cyanide
Cusum	Cumulative sum of the deviations	NPV	Net present value
CV	Coefficient of variation	NQ2	Size of diamond drill bit
DFS	Definitive bankable feasibility study	ºC	Degrees Celsius
DTM	Digital terrain model	OK	Ordinary kriging
E(X)	Easting	oz	Troy ounce
EDM	Electronic distance measuring	P80-75 µ	80% passing 75 microns
EV	Expected value	PAL	Pulverize and leach
G	Gram	ppb	Parts per billion
Ga	Billion Years	ppm	Parts per million
g/m3	Grams per cubic metre	psi	Pounds per square inch
g/t	Grams per tonne	PVC	Poly vinyl chloride
HARD	Half the absolute relative difference	QC	Quality control
HDPE	High density poly ethylene	Q-Q	Quantile — quantile
HQ	Size of diamond drill bit	RAB	Rotary air blast
hr	Hours	RC	Reverse circulation
HRD	Half relative difference	RL (Z)	Reduced
H&S	Hellman and Schofield	RQD	Rock quality designation
ICP-MS	Inductively coupled plasma mass spectroscopy	km2	Square kilometres
ID	Inverse distance weighting	SD	Standard deviation
ID2	Inverse distance squared	SG	Specific gravity
IPS	Integrated pressure stripping	SGS	Societe Generale de Surveillance
IRR	Internal rate of return	SMU	Selective mining unit
ISO	International standards organization	t	Tonnes
ITS	Inchcape testing services	t/m3	Tonnes per cubic metres
kg	Kilogram	tpa	Tonnes per annum
kg/t	Kilogram per tonne	w:o	Waste to ore ratio

Table 2-1: Listing of Abbreviations

3.0 RELIANCE ON OTHER EXPERTS

This report is based on information provided by TMLSA and various consultants which reflect various technical and economic conditions prevailing at the time of compilation of the report. These conditions can change significantly over relatively short periods of time and as such the information and opinions contained in this report may be subject to change.

This report includes technical information, which requires subsequent calculations to derive subtotals, totals and weighted averages. Such calculations may involve a degree of rounding and consequently introduce an error. Where such errors occur, the author does not consider them to be material.

The consultancy and engineering groups who have provided updated information are as follows:

· Resources: The revised resource estimates included in this report were prepared by Nicolas James Johnson of Hellman & Schofield Pty Ltd., an independent consulting firm. A separate certificate of responsibility in relation to Sections 16.1-16.8 is included in this report.

· Reserves/Mining: The revised Ore Reserve included in this report was prepared by Patrick Anthony Smith of AMC Consultants Pty Ltd., an independent consulting firm. A separate certificate of responsibility in relation to Section 16.9 is included in this report.

4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 The Tenement Area

The Tasiast Mine comprises the 312 km2 Tasiast Mining License of El Gaicha located centrally within a surrounding permit block of five contiguous exploration permits (PRM), totalling 4,494 km2 as listed in Table 4-1 below.

Name

District

Type

No.

Km2

Granted

Expiry

Tasiast (El Gaicha)

Wilaya de l’Inchiri

Mining Licence

PE 229

312

19 January 2004

19 January 2034

Tasiast South

Wilayas Dakhlet Noudhibou et Inchiri

Exploration licence

PRM 428

355

02 April 2008

02 April 2017

N’Daouas East

Wilaya de l’Inchiri

Exploration licence

PRM 437

1,478

02 April 2008

02 April 2017

Tasiast West

Wilayas Dakhlet Noudhibou et Inchiri

Exploration licence

PRM 157

1,376

22 January 2001

26 July 2010

Imkebdene

Wilayas Dakhlet Noudhibou et Inchiri

Exploration licence

PRM 237

539

19 January 2004

20 September 2014

Temeinchat

Wilaya de l’Inchiri

Exploration licence

PRM 238

746

19 January 2004

21 September 2014

Table 4-1: Tasiast Mine Licence Summary

The coordinates defining the boundaries of the licences are listed below in Table 4-2

	Licence
	Number
	for Group 2		Coordinates	Coordinates
Name	minerals	Point	UTM (E)	UTM (N)
El Gaicha	PE 229	A	441000	2287000
		B	454000	2287000
		C	454000	2263000
		D	441000	2263000
Imkebdene	PRN 237	A	435000	2311000
		B	446000	2311000
		C	446000	2287000
		D	441000	2287000
		E	441000	2263000
		F	445000	2263000
		G	445000	2258000
		H	432000	2258000
		I	432000	2285000
		J	435000	2285000
Temeinchat	PRM 238	A	446000	2330000
		B	460000	2330000
		C	460000	2263000
		D	454000	2263000
		E	454000	2287000
		F	446000	2287000
Tasiast South	PRM 428	A	460000	2263000
		B	460000	2248000
		C	432000	2248000
		D	432000	2258000
		E	445000	2258000
		F	445000	2263000

Continued

	Licence
	Number
	for Group 2
Name	minerals	Point	Coordinates	Coordinates
Tasiast West	PRM 157	A	420000	2322000
		B	420000	2299000
		C	435000	2299000
		D	435000	2285000
		E	432000	2285000
		F	432000	2272000
		G	405000	2272000
		H	405000	2270000
		I	400000	2270000
		J	400000	2322000
Ndaouas East	PRM 437	A	460000	2322000
		B	500000	2322000
		C	500000	2303000
		D	495000	2303000
		E	495000	2300000
		F	488000	2300000
		G	488000	2293000
		H	484000	2293000
		I	484000	2280000
		J	490000	2280000
		K	490000	2270000
		L	480000	2270000
		M	480000	2260000
		N	473000	2260000
		O	473000	2264000
		P	476000	2264000
		Q	476000	2268000
		R	477000	2268000
		S	477000	2271000
		T	479000	2271000
		U	479000	2283000
		V	460000	2283000

Table 4-2: Licence Coordinates

4.2 Location

Figure 4.2-1: Location

4.3 Ownership

The Company holds a 100% interest in the mine.

4.4 Royalties, Payments, Agreements and Encumbrances

A royalty is payable to the government. This is 3% of the gross revenue. In addition, Franco Nevada Corporation hold a 2% net royalty on production in excess of 600,000 ounces.

4.5 Environmental Liabilities

There are no environmental liabilities inherited by TMLSA from previous owners and explorers.

4.6 Permits

TMLSA holds a valid Mining lease ME 229 (EL Gaicha) covering 312km2 granted on 19 January 2004 and valid for a period of 30 years.

The Company also holds 5 contiguous Exploration permits which are in good standing.

The operating permits are listed in Table 4-3.

Brief Name		Issue Date
Original Operating Permit		#407 - 27 August 2009

New installation’s Permit (Dump Leach, TSF II)		# 408 - 27 August 2009

Authorization of water (12 drills from bore field)		#560 - 24 July 2008

Table 4-3: Operating Permits

4.7 Surveying of the Licences

The mining licence boundary is defined by a list of the coordinates of its corners (“pillar points” see Table 4-2). The boundaries are not physically marked on the ground and have not been surveyed, however, extensive surveying has been carried out within both the Mining Licence and adjoining exploration licenses. To date approximately 30,000 points have been located via formal surveying by qualified surveyors using EDM total station instruments, and many additional points have been picked up by DGPS and GPS methods. Field personnel working in the licence area are well aware of its boundaries. All the known gold deposits are well inside the boundaries, and the size and shape of the Mining Lease is adequate for the intended mining and processing activities.

4.8 Background Information on Mauritania

4.8.1 Background

The Islamic Republic of Mauritania (“Mauritania”), with an area of approximately 1,030,400 km2, is located in north-western Africa and covers the western portion of the Sahara Desert. Mauritania is bordered by the Atlantic Ocean to the west, Western Sahara and Algeria to the north, Mali to the east and Senegal to the south. Nouakchott is the capital and hosts one of two international airports of the country, the other being Nouâdhibou. As well, there are many other points of entry by road leading into Mauritania from Senegal, Mali, Algeria and Western Sahara. Other major urban centres in Mauritania are Nouâdhibou, Zouérate, Atar, Rosso, Bogué Kaédi, Kifa and Néma.

The capital Nouakchott is located along the Atlantic coast. Hassaniya Arabic is the official language in the country, and, other languages spoken are Pulaar, Soninke, Wolof and French. French is commonly used in most business transactions. There are three major ethnic groups in Mauritania: the mixed Maur/black (40%), Maur (30%) and black (30%) (2010 CIA World Fact Book).

Mauritania’s population, estimated at 3.2 million (2010 estimate) is concentrated in both the cities of Nouakchott and Nouâdhibou and along the Senegal River which is located in the south of the country. The remainder of the population inhabit small villages, living and working according to traditional customs. The vast majority population are Muslims.

While education is compulsory, only about 50% of the eligible children in Mauritania receive primary education, and secondary and higher education facilities are limited. Approximately

51% of the population is literate. Mauritania’s only university, the Université de Nouakchott, is located in the capital city and has an enrolment of about 10,000 students.

4.8.2 Economy

Mauritania is one of the most underdeveloped countries in Africa. According to the CIA World Fact Book, half of the population depends on agriculture and livestock for a livelihood, although most of the nomads and many of the farmers were forced into the cities during the severe droughts of the 1970’s and 1980’s. The basic agricultural products in the country include dates, millet, sorghum, rice, corn, cattle and sheep.

Mauritania has several large iron ore deposits which account for 40% of total exports. Mauritania’s coastal waters are amongst the richest fishing areas in the world. However, over-fishing by foreign fleets threatens this key source of revenue. The primary industries are fish processing, oil production, mining of iron ore, gold, and copper.

The government emphasizes reduction of poverty, improvement of health and education, and promoting privatization of the economy.

4.8.3 Currency and Foreign Exchange

The currency utilized in Mauritania is called the Ouguiya (“MRO”). Any foreign exchange and international capital movements must be made through the various “chartered” banks in the country.

4.8.4 Geography and Infrastructure

Mauritania is located along the north-western coast of Africa and is bordered by the Atlantic Ocean to the west. The country’s land mass covers the western portion of the Sahara Desert. Mauritania’s landmass consists mainly of flat and barren desert landscape surfaces which are cross cut by three large NE - SW trending longitudinal dune fields. In the central part of the country, near Adrar and Tagant, several hills and mountains rise up to 915 m a.s.l. In the desert regions, vegetation is sparse but consists of various species of trees (acacia, etc.) and grasses.

Mauritania’s climate is classified as an arid - desert climate (under the Köppen climate classification), with the average annual high temperature of above 44°C between May — August. Minimum temperatures may go below 10°C in December —January. From January to March, sandstorms frequently occur in the country which causes sand build-up and dune formation. Sandstorms do vary in intensity and visibility can be reduced to several metres. A rainy season, usually between July and September, does exist, however, the amount of rainfall and length of season varies spatially and temporally in the various regions of the country. Annual rainfall varies from a few millimetres in the desert regions to as high as 450 mm in the south along the Senegal River. During the last 15 years, the country has recorded two periods of drought: 1984-85 and 1991-92.

Routine access within the country is provided by a 11,000 km network of paved (~3,000 km) and unpaved (~8,000 km) highways as well as numerous desert tracks. A paved 470 km long two-lane highway runs between the cities of Nouakchott and Nouâdhibou.

A 717 km long rail line located along the border between Mauritania and Western Sahara is owned and operated by the Société Nationale Industrielle et Mines (“SNIM”). This rail line is primarily used to haul iron ore from SNIM’s iron ore mines in Zouérate to the port of Nouâdhibou.

Access to the major urban centres of Mauritania is also possible via air. Nouakchott is accessible via international flights operated by numerous West and North African carriers and Air France provide a direct connection to Paris.

Electricity for the capital is primarily supplied by fuel driven generator sets. Urban centres not connected to the national grid have autonomous generating units. The smaller towns and villages either have no electricity or depend on privately owned generators for their power.

Telecommunication services in the country consist of telephone and cellular phone networks as well as the Internet. Communications between small villages and camps is carried out by HF radios and satellite phones in some cases.

Supplies such as fuel, food stuffs and timber are generally obtained in Nouakchott and Nouâdhibou.

4.8.5 Administration

The Mauritanian government is a republic modeled under the French system of government. The President is democratically elected by popular vote to a six year term and the Prime Minister is appointed by the President. The legislature is a bicameral legislature which consists of the Senate and the National Assembly. The Senate consists of 56 seats and a part of the seats are up for election every two years whereas the members of the National Assembly are elected by popular vote to serve five year terms. The country is sub-divided into 12 regions and one capital district. The legal system of Mauritania is a combination of Shari’a (Islamic law) and French civil law.

4.8.6 Political Situation

Mauritania gained Independence from France in 1960. Maaouya Ould Sid Ahmed TAYA seized power in a coup in 1984 and ruled Mauritania for more than two decades. A bloodless coup in August 2005 deposed President TAYA and ushered in a military council that oversaw a transition to democratic rule. Independent candidate Sidi Ould Cheikh ABDALLAHI was inaugurated in April 2007 as Mauritania’s first freely and fairly elected president. His term ended prematurely in August 2008 when a military junta led by General Mohamed Ould Abdel AZIZ deposed him and ushered in a military council government. AZIZ was subsequently elected president in July 2009.

4.8.7 Mining Law in Mauritania

The conditions embodied in the new Mauritanian Mining Code (Law No. 2008-011), adopted April 2008, and the Model Mining Convention (No. 2002.02) adopted January 2002, are regarded as among the most attractive set of conditions in West Africa, and are designed to stimulate and encourage investment in both exploration and mining. Obtaining exploration licences / permits is not difficult and the granting of mining licences / permits is not expected to present difficulties in view of Mauritania’s stated aim to promote the mining industry. The mining industry is seen as one of the main growth industries for the improvement of the country’s economy.

The Decree on Mining Titles (Décret portant sur les Titres Miniers) of December 1999, most recently updated in 2008 by decrees 2008-158 and 2008-159 contains the rules and regulations for the granting, renewal and transfer of exploration licences. Mineral resources are classified into seven groups of minerals, with gold and precious metals belonging to Group 2. The Mining Code provides for five different types of licences / permits including Exploration Licence and Mining Licence as listed in Table 4.8.7 below:

Licence Type	Duration	Renewal Period	Number of Permissible Renewals	Rules and mechanisms
				· Surface Area: 1,000 km2 blocks;
Exploration Licence	3 years	3 years	Two	· Confers right to explore for resources to any depth within licence area;
			At renewal, licenced surface area can be reduced	· Number is limited to 20 exploration licences per legal entity or individual judged to have the technical and financial capability to carry out the work;
			At renewal, licenced surface area can be reduced	· Licences taken under a joint venture in which the title is not taken into consideration for the calculation of the above limit;
				· Transferable under conditions set out by the Decree on Mining Titles.

				· Necessary for operating a mine;
Mining Licence	30 years	10 years	Several	· Within an area covered by an exploration licence, for the same commodities and on the basis of a feasibility study;
				· Granted only to a legal entity incorporated under Mauritanian Law and set-up by the holder of the exploration licence;
				· Transferable under conditions set out by the Decree on Mining Titles;
				· Personnel health and safety reports to be lodged with Ministry every 6 months, environmental and activity reports every year;
				· Land needs to be rehabilitated after mining

Table 4.8.7-1: Summary of Exploration and Mining Licence Terms, Rules and Mechanisms

Exploration Licences (Permis de Recherche Minière or “PRM”) grant exclusive exploration rights over a specific block (maximum 1,000 km2) and are granted for a three (3) year period, renewable twice for additional of three (3) years maximum. At each permit renewal, the permit area can be reduced. PRM’s require expenditure and technical commitments by the licensee. Mining Licenses (“Permis d’Exploitation” or “PE”) are granted for a term of 30 years and are renewable thereafter for a period of 10 years each. A condition of each license is that the holder is required to hire Mauritanian tradespersons, services and to contract with national suppliers and businesses in preference to foreign service providers where the national suppliers and businesses can offer at least the same terms, quality and pricing.

As an incentive to investment in Mauritania, foreign companies are eligible for certain privileges as documented in the Mauritanian Mining Code and Model Mining Convention. These incentives are tabulated in Table 4.8.7-2 below:

Applicable fees, duties, taxes & levies	Exploration Licence	Mining Licence
Registration Fee (at granting, renewal or transfer)	UM 2,000,000	UM 10,000,000

Annual surface fee	1st period: UM 2,000-6,000/km2	UM 50,000/km2
	2nd period: UM 10,000-14,000/km2
	3rd period: UM 20,000-24,000/km2
Royalty		For gold, 3% of the sales value of the metal at the final stage of processing within Mauritania, deductible from taxable income.
Customs duties and other taxes	Complete exemption on all equipment and supplies including fuel	Complete exemption on all imported equipment and supplies, including fuel, for five years after the start of production. Customs duties of 5% thereafter on equipment and supplies imported, except fuel, lubricants, mine supplies and spares that will continue to be exempted from duty.
Corporate Income Tax		The corporate income tax rate of mining operations is set at 25%, tax holiday for the first three (3) financial years. A decree is to set-up precise application rules.
Withholding tax on dividends		16% on repatriated dividends and interest; 0% on dividends reinvested in national territory.
Taxation of expatriates		Expatriates and non-residents employed by the mining licence holder are taxable according to the rules set by decrees of application. In general, they are taxed at half the normal rates.

Table 4.8.7-2: Fiscal Provisions— Mining Code (“MC”) and Model Mining Convention (“MMC”)

5.0 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Site Topography, Elevation and Vegetation

The topography of the Tasiast Permit Area consists mainly of flat, barren plains which are primarily covered by regolith and locally by sand dunes, or eroded paleo-lateritic profiles. Locally, the drainage pattern within and outside of the Tasiast Permit Area consists of several intermittent dendritic first and second order streams that generally flow south-westerly.

The average elevation is approximately 500 feet above sea level.

Vegetation found on the Tasiast Permit Area is sparse and consists primarily of grasses and occasional acacia trees.

Current land use in the mine area consists of occasional nomadic livestock farmers. There are no villages, agricultural farms, nor artisanal mining activity within or around the mine area. The nearest permanent settlements are located some 100 km north of the mine area, on the Société Nationale Industrielle et Minière (“SNIM”) rail line at the railway maintenance station PK22.

5.2 Access

The Tasiast Permit Area is accessed from Nouakchott by using the paved Nouakchott to Nouâdhibou highway for 370km and then via 66km of graded mine access road which is maintained by TMLSA.

An airstrip has been constructed at the Mine Site and is used for light aircraft from Nouakchott.

5.3 Climate

The climate is hot most of the year and characterized by minimal rainfall and strong prevailing NE-SW winds. Maximum temperatures can exceed 45°C and reach lows of ~10°C).

5.4 Surface Rights

The Tasiast Gold Mine is based on granted Mining Lease number P229 held by TMLSA and granted on 19 January 2004.

There are no competing mining rights in the project area.

5.5 Site Development

The development of the Tasiast Gold Mine has required development at eight major locations:

· The treatment plant area which includes the mine administration building and power station.

· The mine services area, including workshops and refuelling area.

· Explosives storage area.

· The staff village.

· The airstrip.

· The borefield

· The dump leach pad

· The tailings storage facility.

5.6 Water Supply

The source of mine water supply is located 60 km west of the mine and is comprised of a semi-saline underground aquifer, which is exploited by twenty wells. Water is pumped to the mine

site through two HDPE pipelines to the raw water storage facility at the mine site. The final capacity of the system is estimated to be 14,000m3 per day.

Reverse osmosis (RO) water treatment plants and storage basins/tanks are located at the mine site. The saline water produced from the RO plant is used to water the haul roads. Potable water for human consumption is bought from the Boulanouar pumping fields. Water is also supplied for local nomadic people located within a radius of 20 km of the mine site.

Two lined raw water storage dams have been constructed adjacent to the process facility with a combined capacity of 35,000m3.

5.7 Power Supply

The mine is located in a remote area where there is no electrical utility grid. Three 2.7MW HFO generator sets have been commissioned. The mine power has back-up generation provided by eight 1.0MW diesel generators.

5.8 Mine Personnel

As at December 2009, TMLSA employed 454 people, of which 64 were expatriates.

5.9 Tailings Storage Facility

The original tailings storage facility, located to the west of the mine, has ceased to be used and is being de-commissioned.

A second tailings facility is situated to the south east of the concentrator. Tailings slurry is generated at the concentrator and pumped to a tailings facility where they are discharged with a high solids content of 50% in order to minimize additional water requirements. Runoff water is re-circulated to the concentration process.

5.10 Administration and Plant Site Buildings

Administration offices, infirmary and warehouse are made up of pre-fabricated units.

An assay laboratory has been constructed on site for the analysis of exploration samples and production samples from the mine and concentrator. This facility is run under contract by Societe Generale de Surveillance SA (SGS).

Mining equipment maintenance garage and mechanical shops have a pre-fabricated, roofed service area and a bay with an overhead crane and access pit for inspection and repair of vehicles.

Covered areas are provided for mechanical spares and cyanide. Open storage is used for lime.

5.11 Accommodation

A staff village has been constructed using pre-fabricated and containerised bedrooms and dormitories, 2km from the process plant.

A mess hall, fully equipped kitchen, food storage and laundry facilities serve all employees.

Religious facilities are provided.

Sewerage is disposed of through septic tanks fitted with soak away overflow systems. Currently there are septic tank systems at the mine camp and at the Mine Offices. Tanks are emptied on an ‘as required’ basis and the effluent is placed in a bunded area to dry.

5.12 Communications

The plant is provided with the following communication and radio facilities:

· Telephone system with battery back-up facility.

· VOIP Satellite telephone system suitable for phone, fax and data transmission.

· Base station radio system.

· Vehicle radios.

· Hand-held radios.

· Cell phone coverage through a dedicated mast located at the mine village.

5.13 Mobile Equipment

Sufficient mobile equipment for the efficient running of the operations is in place comprising light vehicles (including ambulance), light trucks, cranes, forklifts, buses and generators.

5.14 Security

Security is provided by a contractor.

Appropriate secure facilities are provided for the storage of fuel and explosives.

5.15 Goods and Consumables

The principal ports of entry for goods and consumables are either Nouakchott or Nouadibhou. Materials are transported by road to the mine site.

6.0 History

6.1 Ownership History

Upon the completion of a regional reconnaissance exploration program carried out by the Office Mauritanien de Recherches Géologiques (OMRG) in 1996, the areas within and around the Tasiast area were made available to third parties. As a result, Normandy LaSource Development Ltd. (NLSD) (a subsidiary of Normandy Mining Ltd. of Australia) acquired the Tasiast area. In 2001, NLSD was acquired by Newmont Mining Corporation creating Newmont LaSource.

Midas Gold plc (Midas) was incorporated in England and Wales in May, 2002 for the purpose of acquiring Newmont LaSource’s assets in Mauritania including the Tasiast Project as well as various other permit areas. Midas completed its acquisition of the Tasiast Project from Newmont LaSource on April 1, 2003.

In April 2003, Geomaque Explorations Inc. (Geomaque) announced the acquisition of Midas and the extension of a loan allowing Midas to complete its acquisition of the Tasiast Project. The merger of Geomaque and Midas ultimately created a new entity- Defiance Mining Corporation.

In June 2004 Rio Narcea Gold Mines Ltd (Rio Narcea) took ownership of the Tasiast Gold Project in addition to Defiance’s other exploration properties in Mauritania through the acquisition of Defiance.

In July 2007 Lundin Mining Corporation acquired the Tasiast Gold Mine through the acquisition of Rio Narcea.

On August 2, 2007, Red Back completed the acquisition of the Tasiast gold mine from Lundin Mining Corporation through the purchase of TML and TMLSA. The Acquisition was completed pursuant to a share purchase agreement, dated as of August 1, 2007, among Tasiast Holdings S.A., Rio Narcea Tasiast Luxembourg S.A., Lundin and Blokvink Holdings B.V. (a subsidiary of Red Back now known as Red Back Mining B.V.).

6.2 Exploration History

6.2.1 Exploration Work Prior To 1993

From 1962 to 1993, the Tasiast region has been the subject of three regional exploration programs which were carried out by the Bureau de Recherches Géologique et Minières (BRGM) as well as SNIM. Three regional exploration programs with different objectives and commodity searches were undertaken in the Tasiast region:

Date	Mission	Operator	Target
1960-1962	The Pegmatite Mission	BRGM	Beryllium and Lithium
1972	The Nickel Sulphide Mission	BRGM	Nickel
1973-1975	Iron Ore Mission	SNIM	Iron Ore

Table 6.2-1: Regional Exploration Programs

6.2.2 1993 - 1996 European Development Fund - OMRG

Three exploration programs were carried out in the Tasiast region between 1993 and 1996 as a European Development Fund project (EDF Project). The EDF Project was carried out by the OMRG and the BRGM. The first exploration program was focused on an area covered by the 1:200,000 scale Châmi topographical map (République Islamique de Mauritanie_Feuille NF-28-

III-IX) and consisted of regional-scale reconnaissance geological mapping and geochemical sampling. It was this initial exploration work that identified the Tasiast area as being anomalous in gold.. Traverse lines were oriented E-W with samples collected at 500 m centres. This was followed up with more detailed soil sampling of the Tasiast area (250m centres) and by a third exploration program which identified a series of geochemical anomalies at Tasiast. Several geochemical anomalies were tested by the OMRG by manual trenching. One of these soil geochemical anomalies, then called the C6-9 anomaly, eventually became the main Tasiast prospect. Since the EDF Project’s primary mission to locate new mineral indices of potential economic importance in the Tasiast area had been accomplished, and given the fact that the OMRG did not have the financial means to advance the Tasiast prospect to the drilling stage, the areas explored within and around the Tasiast area were made available to third parties. NLSD acquired the Tasiast area in 1996.

6.2.3 1997 - 2001 NLSD

Between 1996 and 2001, NLSD completed a major exploration program within the Tasiast Permit Area, including some 32,000m of RC drilling and 3,300m of diamond drilling. The program was managed and operated on a contractual basis by the BRGM who provided the supervising chief geologist and other technical expertise as required. An extensive amount of work (summarized in table 6.2.3) was completed during the program.

· Geological mapping and regolith mapping

· Satellite imagery interpretation;

· Airborne and ground magnetic surveys;

· Geochemical auger drilling carried out along strike of Tasiast;

· Drilling: drilled initially on 200m spaced E-W sections with a 50 m hole spacing along each section, to depths of 50-100m in three drill campaigns from 1996 to 1999. Drilling methods were predominantly reverse circulation (RC) with lesser diamond drilling (HQ diameter core) and including diamond tails to some RC holes (NQ diameter core);

· The strongly mineralized central area was renamed the Colonne Piment Zone (or Piment Zone). Drill spacing was reduced to 50m E-W sections along strike and ~ 30m down dip;

· During later stages, selected RC holes were twinned via diamond drilling and additional trenching and infill drilling completed in order to better understand the geometry and structural controls affecting high-grade gold mineralization;

· Hydrological drilling for ground water;

· Sampling and analysis of samples for gold and associated metals;

· Petrographic, mineralogic, and specialist geological studies; and,

· Preliminary metallurgical test work.

Table 6.2.3: Summary of NLSD Exploration Activities at Tasiast

Exploration Work	Area/Amount	Number of Samples
Aeromagnetic survey	2,000 km2
Satellite image interpretation	2,000 km2
1:1500,000 geological mapping	2,000 km2
1:10,000 geological mapping	72 km2
1:2,500 geological mapping	7 km2
Trenching (mechanised)	55 trenches, 26,593 m	11,747
Trenching (manual)	27 trenches, 1,309 m	380
Soil geochemistry		17,992
Litho geochemistry		130
RC drilling	338 holes, 29,606 m	18,979
Diamond drilling	35 holes, 3,313 m	3,751
Hydro geological drilling	4 holes, 280 m
Hydro geological studies	2
Geophysical study for water	75 lines, 35 line km, 40 electric holes
Topography/surveying	3,600 points
Metallurgical testing	BRGM	3 bottle roll tests
	CSMA	15 leach tests
		10 gravity / leach tests
		5 column tests
		3 bond ball mill index tests

6.2.4 2003 Midas Gold plc

Following Newmont Mining Corporation’s acquisition of NLSD in 2001 and divestiture of various NLSD assets including the Tasiast Project, Midas Gold plc (“Midas”) conducted an evaluation of the Tasiast data in 2003, completing a resource estimation for the West Branch and Piment areas.

6.2.5 2003 GeomaqueE. - Defiance Mining Corporation

In April 2003 Geomaque announced the acquisition of Midas and the extension of a loan allowing Midas to complete its acquisition of the Tasiast Project. The merger of Geomaque and Midas ultimately created a new entity - Defiance Mining Corporation.

Between March 2003 and October 2004 Defiance completed some 415 RC holes totalling 32,700m of drilling, and 33 diamond drill holes totalling 2,270m of drilling on the Piment Zone.

Defiance subsequently contracted ACA Howe International to provide geological services for the Tasiast Project and a NI 43-101 compliant technical report and resource estimate and ultimately SNC to prepare a bankable feasibility study based on the Howe resource.

Defiance also contracted Golder Associates (Golder) to undertake a preliminary pit slope design study; Prospections Hydrauliques, a Mauritanian consulting firm specializing in hydrogeology, to investigate the existence of aquifers some 60 kilometres from the project site; SGS Lakefield Research Limited (Lakefield) to undertake mineralogical and metallurgical test work; and Outokumpu Technology Canada (Outokumpu) to test the thickener properties of Tasiast ore with High Rate and Paste thickeners.

6.2.6 Rio Narcea

Between June 2004 and August 2007 Rio Narcea completed 4,160m of diamond drilling and 22,062m of reverse circulation drilling in the Piment and West Branch areas. Some of this drilling was directed at testing the depth extents of the Piment ore body in the central area whilst the bulk of the work was directed at tracing the mineralised zone to the north and infill drilling at the West Branch.

6.3 Resource History

Several Mineral Resource estimates have been made for the project and are listed in the table below:

		Measured and Indicated				Inferred
Date	Source	COG	Mt	Au g/t	Moz	Mt	Au g/t	Moz
2000	Ankobra Resource Services for NLSD	1.00	18.80	2.24	1.36	12.00	1.85	0.71
January 2003	Midas Gold	1.00	8.30	2.29	0.61	21.00	1.74	1.18
September 2003	ACA Howe for Defiance	1.00	10.20	3.10	1.01	14.40	2.31	1.07
October 2003	ACA Howe for Defiance	1.00	12.07	3.06	1.18	12.43	2.25	0.90
February 2008	H&S for Red Back	1.00	26.47	2.22	1.89	8.90	1.90	0.55
February 2009	H&S for Red Back	1.00	38.20	2.06	2.53	7.40	2.00	0.47

Table 6.3: Summary of Mineral Resource History

The resources and reserves associated with the Tasiast Gold Mine have been updated by Hellman and Schofield and AMC, respectively, using data available up to end December 2008.

The updated Mineral Resource and Ore Reserve estimates, which have been classified and reported in compliance with the requirements of both National Instrument 43-101 and the Australasian JORC Code, are respectively discussed in detail in Section 16 of this report.

6.4 Production History

There has been no historical gold production from the Tasiast area. Commercial production of gold at Tasiast was commenced by the Corporation in the first quarter of 2008.

	CIL				Total
	Tonnes Milled	Grade	Recovery	DL	Gold Produced
Year	(Mt)	(g/t)	(%)	(‘000 ozs)	(‘000 ozs)
2009	1.69	2.87	91.6%	16	159
2008	1.49	3.10	93.6%		140
2007	0.22	4.77	62.8%		21

Table 6.4: Summary of Production History

7.0 GEOLOGICAL SETTING

7.1 Regional Geological setting

The Tasiast Permit is located within the Archean-age Aouéouat greenstone belt, a 70 km long by 15 km wide N-S trending belt situated within the SW sector of the Reguibat Shield (Figure 7-1). The Reguibat Shield is a geological subdivision of Mauritania that comprises Precambrian basement granitoid and gneissic domes, mafic and acidic volcanic rocks, volcano-sedimentary rocks and mafic intrusions. The Reguibat Shield forms part of the West African Craton which is bound on all sides by younger age fold belts. Towards the South West of the Reguibat Shield and more specifically in the Tasiast district, four North-south trending greenstone belts include ferruginous (magnetite-rich) quartzites, which are often referred to as Banded Iron Formations or BIF and mafic dominated volcanic-sedimentary sequences, often referred as “Greenstones”.

The Tasiast Permit area overlies four north-south trending greenstone belts (LaSource-BRGM, 1997) which are, from east to west:

· N’daouas greenstone belt (20 km long x 5 km wide);

· Aouéouat greenstone belt (+100 km long x 15 km wide);

· Kneffissat greenstone belt; and (+ 50 km long x 15 km wide); and

· Hadeïbt Agheyâne greenstone belt (+25 km long x 5 km wide).

These four greenstone belts are geologically similar to other Archean greenstone belts in the world, which are known to host major gold deposits.

All of the above mentioned greenstone belts consist mainly of amphibolite, amphibolite-garnet schist and mica schist. However, the Aouéouat greenstone belt is the only one out of the four containing abundant ferruginous (magnetite) quartzites/banded iron formation.

The Aouéouat belt shows a typical volcano-sedimentary stratigraphy composed of metavolcanic and metasedimentary sequences of massive-to-pillowed mafic to ultramafic metavolcanic rocks, dacite, epiclastic rocks interbedded with BIF and mica schist (table 7-1). This stratigraphy has been intruded and is surrounded by a syn-to-post kinematic suite of granodiorite and large granite bodies. The belt is transected by younger age, discontinuous NNE-SSW and WNW-ESE or WSW-ENE gabbro/dolerite dykes and sills. The volcano-sedimentary succession uncomformably overlies an Archean age basement complex composed of granulite, granitic gneiss and migmatite. The Aouéouat greenstone belt has not been geo-chronologically dated. However, based on U-Pb model age dates obtained from gneiss/granodiorite and pegmatite intrusives that occur in the area, an age range from 3,070 Ma to 2,600 Ma is inferred (LaSource-BRGM 1997, Maurin et al, 1996). Maurin (1996) states that the 2,600 Ma age date most likely represents a thermal episode in the region.

Figure 7.1: Geology of the West African Craton

Structurally, the Aouéouat greenstone belt has undergone at least two major tectonic deformation events. The first, referred to as “D1”, involved an E-W compression that resulted in N-S trending isoclinal folds, shearing and low angle, ductile thrust (reverse) faults. The second event, referred to as “D2”, involved a strong NW-SE compression that resulted in tighter folding, shearing of folds, brittle N-S strike-slip faulting, low angle reverse faults and associated second-order faulting and shearing. The greenstone belts within the Tasiast Project Area were subsequently cut by younger cross faults and NNE trending gabbro - dolerite dykes.

During the Tertiary period, the then prevailing sub-tropical conditions formed lateritic profiles over the eroded basement, remnants of which occur at Tasiast. Elsewhere, the duricrust has been eroded and re-deposited as widespread sheets of gravel lag.

The geology and Archean age of the Reguibat Shield greenstone belts are similar to other well known gold bearing greenstone belts containing auriferous BIF, including the Yilgarn Craton of Western Australia and the Superior Craton of Canada.

7.2 Property Geology

Two principal mineralised zones have been identified at Tasiast, the Piment Zone which hosts the bulk of the mineral resources and ore reserves, and the West Branch prospect which lies 1km south west of the current southern limit of the Piment zone.

The Piment Zone is situated along the east limb of an interpreted broad, regional antiform that is cored by a felsic volcanics, approximately 200 m wide at the present surface elevation. The West Branch is located on the reverse limb of this antiform.

In the Tasiast area, the BIF units or horizons consist of an oxide-silicate facies ferrugineous magnetite quartzite, and is part of the following stratigraphic sequence (Table 7.2).

· Epiclastic sediments, greywacke;

· Oxide-silicate facies ferrugineous magnetite quartzite (BIF) ±grunerite-cummingtonite;

· Garnetiferous green schist;

· Biotite green schist; locally contains garnet crystals and grades into the garnetiferous green schist;

· Intercalated BIF/garnetiferous green schist, with varying proportions of the two components occurring in centimetre- or decimetre-size beds;

· Dacite; fine to medium grained; overlain by an aphanitic, pyrite/pyrrhotite-rich interval at the upper contact, interpreted as a chert by previous operators;

Late cross-cutting mafic (dolerite-gabbro) dykes.

PHANEROZOIC

CENOZOIC

RECENT

Fluvial gravel, sand, clay, silt, latosols, duricrust, sand dunes etc.

Formation of lateritic soils and saprolite (older??)

Gabbro — dolerite dyke swarms (120-65 Ma?)

UNCONFORMITY~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~

PRECAMBRIAN

PROTEROZOIC

LOWER TO MIDDLE (?)

Unmetamorphosed quartzites and mica-schists (with intercalated amphibolites and orthogneiss)

UNCONFORMITY~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~~

ARCHEAN

Late intrusives

Calk-alkaline granite-granodiorite, pegmatites, granitoids (~2,600 Ma)

Gabbro — Diorite, dolerite dykes (age?)

INTRUSIVE CONTACT-----------------------------------------------

Volcano-Sedimentary pile (> 2,600 Ma)

Greywackes (epiclastites), mafic volcanics (basalts?)

Banded iron formations inter-layered with alternating garniteferous schists, micaschists and tholeiites Amphibolites and dacites

Basement (>2,600 Ma)

Granite gneiss, orthogneiss, migmatite domes

Table 7.2: Table of Formations, Aouéouat Greenstone Belt, Mauritania

On a mine-scale, there are two primary N-S mineralised BIF “tram-lines” which straddle a core of felsic rhyo-dacitic to rhyolitic volcanic. The Piment ore bodies form on the eastern limb, the West Branch on the western limb. The recent discovery of the Greenschist zone lies within the interpreted axis of the fold.

All of the ore bodies defined to date dip moderate to steep to the east. All of the mineralisation discovered to date is demonstrably epigenetic and strongly structurally controlled.

Figure 7.2-1: Property Geology of the Piment Zone—Tasiast Permit Area

Several late NNE-SSW and WNW-ESE or WSW-ENE (fault related in some cases) mafic dykes intrude this package. A few narrow regional faults that appear to strike NE and NNW, with easterly dips, were encountered in some of the core holes drilled by Defiance. They are not volumetrically important.

Exploration work has now identified four prospective gold proximal to the Tasiast Mine (refer to figure 7.2-2):

1. Tasiast Main Trend, is the main Mine BIF corridor which trends N-S, axially through the El Gaicha mining license.

2. Pantaloon Trend, is a BIF trend, located sub-parallel, 3km to the East of Tasiast Main and comprises several sub-parallel NNW trending gold shears which converge onto the main trend north of El Gaicha.

3. Imkebdene Trend, is a N-S regional trending shear, located 4km to the west of the main trend.

4. Tasiast West Trend, is a NNE regional trending shear, located 30km west of the mine.

Figure 7.2-3: Tasiast - Main Exploration Trends

8.0 DEPOSIT TYPES

The gold deposits forming the basis for the Tasiast Gold Mine are structurally and lithologically controlled, epigenetic, BIF hosted deposits common in most Archean terranes. Exploration is therefore directed towards identifying structural and lithological settings favourable to the formation of gold mineralisation.

The traditional BIF-hosted gold deposits in Piment are thought to have formed through a reaction of auriferous and sulphur-bearing hydrothermal fluids with the iron oxide (or sulphide) present in country rocks, leading to the precipitation of gold and sulphides. Elevated gold concentrations associate with the more iron-rich BIF and greenschist lithologies. These lithologies would seem to be providing more preferential chemical and rheological “trap” characteristics within the host architecture.

MINERALISATION

Gold mineralisation has been defined over a strike length of over 7km and to a maximum vertical depth of 300m. To date two main zones of mineralisation have been delineated (see figure 4.2.2), the Piment Zone and the West Branch. Mineralisation is continuous within both zones.

Gold mineralisation is associated with structurally controlled late, discrete faults and shears, quartz-veining and silica-flooding, within all rock types. The mine was originally based primarily on BIM/BIF hosted gold mineralisation. Red Back have redirected their exploration model to expand the mine resource search into all potential structural hosts. There is ubiquitous bleaching due to prograde epidotization and retrograde chloritisation and later overprinting hydrothermal sericitization. Generally late, brittle, quartz-carbonate veining is overprinted by the remobilization and replacement of magnetite by secondary pyrhotite. Coarse visible gold is common within the micro-fracturing associated with the main shears.

Figure 8: Tasiast Mineralisation

Macroscopic gold grains have been observed in virtually all the exploration core holes in the following settings:

· isolated grains in quartz veins, commonly near the contacts;

· isolated grains near stoped fragments of host rock;

· closely associated with pyrrhotite: generally against or at the periphery, occasionally as inclusion;

· associated with pyrite (Piment North);

· in fractures within garnets crystals;

· seldom associated with arsenopyrite;

· rare occurrence within carbonate veins.

The vast majority of the quartz veins containing coarse visible gold cut the bedding (S0) at low angle and dip gently to the east. To a significantly lesser degree, visible gold is found in veins cross-cutting S0 at high angle as well as in the fabric of the rock. Sulphides in the gold mineralized zones, as determined by visual examination, are largely represented by pyrrhotite and, in decreasing order of occurrence, pyrite, arsenopyrite, and chalcopyrite. Pyrrhotite is commonly seen replacing magnetite, but sulphidation is generally limited to vein selvages. Pyrite rather than pyrrhotite occurs principally in the Piment North mineralization while subordinate amounts of pyrite are found in the Piment Central mineralization. Lesser arsenopyrite occurs as isolated, euhedral crystals. Chalcopyrite has been observed on a few occasions as small inclusions within pyrrhotite.

Carbonate and chlorite assemblages, with occasional tourmaline, are commonly found in the vein selvages. Amphiboles of the grunerite-cummingtonite series are the dominant silicate minerals found in the BIF unit.

Although the presence of graphite has been reported by NLSD in the descriptions of polished sections from four boreholes, no definite graphite occurrences were observed in the core drilled by Defiance. Metallurgical test work undertaken by Lakefield for Defiance has however confirmed the presence of minor graphite throughout the mineralized zones.

The oxide zone cut by core drilling is characterized by leaching of the quartz, carbonate veins and sulphides, calcrete precipitation and enrichment in iron hydroxides caused by meteoric water. No well-defined transition zone was identified as the strongly weathered upper portion of the deposits grades into fresh. The depth of oxidation is in the order of 50 m and no supergene enrichment of gold is apparent.

9.0 EXPLORATION

9.1 Introduction

Exploration at Tasiast is at an advanced stage with two significant zones of mineralisation delineated to date. Significant potential remains to increase the known Mineral Resources and Mineral Reserves and exploration will continue during the life of the mining operation.

9.2 Exploration Methods Used

The Tasiast area has been covered by a significant quantity of soil geochemical sampling. This has produced many anomalies the strongest of which have proved to be directly related to underlying gold mineralization.

Trenches across the core of the soil anomalies have been seen to yield widths and grades of gold comparable to those seen in later drilling beneath the trenches.

Geological mapping has been undertaken over the entire Tasiast area.

Both ground and airborne geophysics have been used in the area.

Both reverse circulation (RC) and diamond drilling have been employed for exploration and resource definition purposes.

9.3 Interpretation of Exploration Information

9.3.1 Introduction

Interpretation of the exploration information is continuous and ongoing.

9.3.2 Soil Sampling

Geochemical data derived from soil sampling undertaken in the early stages of the project has generally produced clearly delineated anomalies. The largest anomaly in the area is clearly related to the Tasiast mineralisation. Numerous other, lesser anomalies are distributed throughout the project area. Most have been tested to some degree by trenching and drilling and in most cases the anomalies are seen to be related to bedrock mineralisation.

Regolith development in most of the Tasiast area is generally favourable for soil sampling however in the south and east of the project area sand dunes inhibit the effectiveness of this method of exploration.

9.3.3 Geological Mapping

Tasiast represents the ideal environment for geological mapping with bedrock lying very close to the surface. Numerous phase of geological mapping have been undertaken during the life of the project leading to a solid understanding of the geological setting of the Tasiast mineralisation. This mapping working is being continued by Red Back.

9.3.4 Trenches

Excavation of trenches as an exploration technique has been very successful and was extensively used during the NLSD phase of exploration. Significant gold intersections in trenches typically overlie sub-surface zones of similar grade and width, as defined by subsequent drilling.

9.3.5 Drilling

Drilling indicates that the two main mineralised zones at Tasiast are generally quite simple in morphology, and that the continuity of mineralized width and grade are predictable. The mineralisation is not significantly interrupted by faulting or dyke development, nor complicated by folding.

Over ninety percent of the RC holes have been completed under dry drilling conditions.

9.3.6 Ground and Airborne Geophysics

Geophysical surveys at Tasiast have not materially influenced either the exploration approach or resulted in the discovery of mineralisation to date but has aided considerably in geological and structural analysis of the area. An airborne geophysical survey was flown by Red Back in 2008 and interpretation is ongoing.

9.3.7 Data Reliability

The Company has inherited a significant amount of data from previous operators of the Tasiast Project covering a period of over 10 years. All work appears to have been carried out by technically qualified personnel, and has been planned and supervised by trained and experienced geoscientists. The location of all exploration data is known with adequate accuracy and is the subject of a continuing check process by the Company. With the exception of early drilling by NLSD, the quality of geochemical analysis has been monitored by the use of blanks, standards, field duplicates, and check analysis via primary and umpire laboratories.

Overall the data acquired on the Tasiast Mine is considered to be reliable.

10.0 Drilling

10.1 Introduction

The Tasiast Property has been the subject of numerous drill campaigns since its discovery in the mid 1990’s as listed in Table 10.1. The table below shows a summary of the drilling and trenching activity on the resource area to date. This excludes drilling on outlying projects that has yet to be validated and entered into the Tasiast drill database.

		Reverse Circulation		Core
Company	Period	Holes	Metres	Holes	Metres
Normandy La Source	1996 - 1999	338	29,606	35	3,313
Defiance	2003	303	25,856	29	1,978
	2004	112	6,844	4	292
Rio Narcea	2007	179	22,062	18	4,160
Red Back	2007	159	16,459
Red Back	2008	1,029	113,331	21	2,069
Red Back	2009	582	116,439	29	3,784
Total		2,702	330,597	136	15,596

Table 10.1: Tasiast Drill Summary

10.2 1996 to 1999 NLSD Drill Programme

NLSD completed three drill campaigns between 1996 and 1999. This comprised 338 RC holes for 29,606m and 35 diamond core holes for 3,313m.

Drilling was initially undertaken on 200 m spaced E-W sections with 50 m hole spacing along each section, to depths of 50-100 m in three drill campaigns from 1996 to 1999. Drilling methods were predominantly reverse circulation (RC) with lesser diamond drilling (HQ diameter core) and including diamond tails to some RC holes (NQ diameter core).

10.3 2003 Defiance Drill Programme

Reverse Circulation Drilling

From March 1 to June 18, 2003, a total of 303 reverse circulation holes (“RC”) totalling 25,859 metres, or a cumulative 26,774 metres with RC pre-collars, were completed on the Piment zone by Defiance. RC hole diameter was 5 5/16”.

RC holes were drilled in between old NLSD RC holes along drill fences at 25 m spacing along east — west fences. The majority of the RC holes were drilled at an azimuth of 270° (grid orientation) and at an inclination of -60°.

Diamond Drilling

From March 1 to May 25, 2003, a total of 29 diamond drill holes totalling 1,976 metres of core were completed on Piment Central, Piment South, and Piment North (southern extension) by Defiance. DDH core diameter used was HQ3 for 25 of the 29 DDH holes, while one diamond drill hole (SC062) was drilled utilising NQ core diameter. Seven of the diamond drill holes were

drilled primarily for geotechnical purpose and three vertical PQ3 diamond drill holes (SC059, SC060 and SC061) were drilled to collect samples for metallurgical test work.

Of the 29 diamond drill holes completed, 18 holes were pre-collared by RC to pre-determined depths in order to avoid expensive drilling in non-mineralized rock.

For drill holes collared from surface, triple tube core barrel equipment was utilised for recovering core from the oxide material, and conventional core barrels were utilised in the primary rock. Orientation marks of the drill core were made every core run, either at three or six metres, using conventional down-hole spearing methods. Orientation of the vertical PQ3 drill holes was not carried out. All drill core was stored in clearly marked, one metre long plastic core boxes and subsequently transported and stored next to the core logging facility by Defiance staff.

All of the geological and geotechnical logging and photographic records were undertaken before the core was marked and cut for assaying.

10.4 2004 Defiance Drill Programme

Reverse Circulation Drilling

From March 2004 to October 2004, a total of 112 RC holes totalling 6,844 metres (including 4 RC pre-collars of four (4) deep diamond holes) were completed on the Piment Zone and to the west (as sterilisation of the waste dumps and tailings dam areas). RC hole diameter was 5 5/16”.

The procedures followed and the drilling equipment used was the same as that used in the January-July 2003 drilling program.

Diamond Drilling

From March 2004 to October 2004, a total of four (4) deep diamond drill holes totalling 292 m (1,122 m with RC pre-collars) were completed on the Piment Zoneto check the down-dip extension of the northern Piment Central shoot. The procedures followed and the equipment used was the same as that used in the January-July 2003 drilling program.

10.5 2007 Rio Narcea Drill Programme

During 2007 Rio Narcea re-commenced exploration using an in-house CS2000 drill rig and an RC rig supplied by Drillcorp Sahara.

During the period to August 10th 2007, when Red Back acquired the Tasiast project Rio Narcea drilled a total of 179 RC holes totalling 22,063m and 18 diamond core holes totalling 4,160m.

The RC drilling was specifically aimed at testing the northern extensions of the Piment Zone and on infill drilling at the West Branch.

Drill and sampling procedures for both RC and Core drilling where essentially the same as those employed by Defiance.

10.6 Red Back Mining Drill Programmes

Following the acquisition of the project Red Back commenced an aggressive programme of RC drilling to fully define the Mineral Resources in the Piment and West Branch mineralised zones.

During 2009 a further 116,439m of RC and 3,784m of core was completed.

Diamond drilling has continued, providing material for ongoing metallurgical testwork relating to the dump and heap leach potential of the project and for quality control purposes.

A programme of deeper, exploratory core drilling has also commenced.

11.0 SAMPLING METHOD AND APPROACH

11.1 NLSD

Little information has been preserved on the detailed procedures used in this drilling.

11.2 2003-2004 Defiance and 2007 Rio Narcea

RC Sampling

All of the RC holes were sampled at one-metre intervals and each sample was collected in a large plastic sample bag that was held below the cyclone spigot by a drill helper. All samples were sent off for assay except those that originated from the non-mineralized hanging wall at the start of each hole. To avoid sample contamination after a drill run was completed, blow-backs were carried out at the end of each 6.0m run by the driller whereby the percussion bit was lifted off the bottom of the hole and the hole was blown clean. When water was encountered in the hole, the driller would dry out the hole by increasing air pressure into the hole and lifting and lowering the rods prior to continuing the drilling.

Throughout the Defiance RC drill program, logging of all RC drill holes was conducted by the field geologist at the drill site. After each drilled 1.0m interval, the sample was weighed, sieved and split to give a 2-3 kg sample for analysis.

A representative sub-sample for geological logging was collected from the large sample bag by spearing a small diameter PVC pipe into the bag and emptying the contents of the PVC pipe into a hand sieve.

At the end of each day or at the completion of an RC hole, calico sample bags for RC drill holes completed that day were loaded onto a 4x4 pick-up truck by the field geologist and then delivered directly to the on site sample preparation laboratory. Once the samples were unloaded from the pick-up truck and both the field geologist and lab technician confirmed receipt of all calico sample bags, the field geologist then registered the sample number bag sequence.

Diamond Drill Core sampling

Upon the completion of the geological and geotechnical core logging of a diamond drill hole, Defiance’s core logging geologist identified the sections of core to be sampled and analysed for gold. Once identified, the core-logging geologist measured and marked out the sample intervals onto the uncut core’s down-hole right hand side of the orientation line and recorded the individual sample intervals onto a core-sampling sheet. The core was sampled according to lithological boundaries and vein widths, but the maximum sample interval did not exceed 1.50m in length.

At the core cutting facility the drill core boxes were stacked in ascending order so as to avoid sampling mix-ups. The core was cut on the line marked by the geologist and one half of the core placed in a numbered calico bag.

Once the core for a drill hole was cut and sampled, the core cutter and the core logging geologist then delivered the samples, with the core sampling sheet, to the preparation laboratory technician for sample preparation.

11.3 2007 and 2008 Red Back

On acquisition of the property Red Back halted the diamond drilling programme in order to concentrate on RC infill drilling.

RC Sampling Procedure

To minimise down-the-hole deviation, RC drilling is conducted with contract single and multi-purpose rigs using a standard 5½”face sampling hammer leading a 4½” 6m rod string.

The entire sample is collected in a large plastic bag tightly clamped onto the cyclone base. The entire length of each RC hole is sampled. A one-meter sample length is used in all holes. Dry samples, of nominal 20kg to 25kg weight, are reduced in size by riffle splitting using a two stage Jones riffle splitter to about three kilograms, and then placed in pre-numbered sample bags for dispatch to the assay laboratory. A record is made at the drill site of the sample identity numbers and corresponding intervals, and this is also recorded in the geological log.

Routine analytical sample ‘Duplicates’ are collected every 20th sample, and submitted in blind sequence every 20th and 21st within the sample stream. Further representative triplicate sample intervals are also routinely collected every 60th original sample in the sequence and retained for later submission to a 3rd party, independent referee laboratory. Analytical ‘RC Blanks’ are inserted every 20th original sample and are taken from barren dune sand collected form a source distant from the mine. GANNET, ROCKLABS and GEOSTATS certified reference materials (CRM) pulps are selected relative to important ‘resource thresholds’ and inserted as standards every 20th sample. All QAQC samples are inserted by the rig geologist at the rig. Grades of standards to be used are selected by the senior geologist and provided to the rig geologist in his rig box. TMLSA submit 16% routine QC samples within the sample stream. Holes are submitted by the rig geologist direct to the on-site lab as individual batch jobs.

In contrast to Defiance and Rio Narcea, Red Back routinely samples every metre drilled.

11.4 Bulk Density sampling

The results from 1,699 bulk density determinations completed by NLSD at Tasiast during previous drilling programs are available. The origin of the sample, its borehole number and sample depth was entered as an individual MS Access database file into NLSD’s project database. However, information on the sample size/length, lithology and oxidation state was not recorded in the NLSD database.

A total of 131 bulk density measurements were carried out on lengths of complete drill core by Defiance during their programmes. Density determinations were undertaken prior to core sawing on 131 samples of about 8 to 15 cm in length and of both HQ and HQ3 diameter.

During 2008 Red Back completed 495 determinations of bulk density (BD) using the Archimedes method have been undertaken on core samples. The samples were selected to provide a representative suite of densities covering all major lithology types and from all oxidation levels. This data was used in conjunction with 3-dimensional geological modelling and a remodelling of oxidation surfaces based on re-logging of all Red Back drill holes as input into the revised resource estimate.

12.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

12.1 1996-1999 NLSD

Little information has been preserved on the detailed procedures used in this drilling.

12.2 2003-2004 Defiance and 2007 Rio Narcea

Sample preparation was undertaken on site by Defiance and Rio Narcea personnel.

The entire RC calico sample bag was oven-dried for 24 hours, and then weighed prior to pulverization of the entire 2-3 kg sub sample using a Labtecnics LM5 mill

Each core sample was crushed to —10 mm in a jaw crusher and the entire sample was pulverised to P90 — 75 µm using a Labtecnics LM5 mill.

Barren dune sand was used to clean the bowls after every sample.

The pulverised material was sampled using a spatula and two 120g pulp sub-splits were then taken, one packet was prepared for shipment to the assay laboratory and one packet remained on site for future checks.

Blanks of dune sand and certified reference materials were added at this stage.

Sample pulp shipments were carried out on a weekly basis, with samples packed in wooden boxes. The sample preparation laboratory manager completed a Sample Submission Sheet, the original of which was placed inside the boxes, and then the boxes were secured and transported to Nouakchott, where Mauritanian Customs clearance was completed prior to shipment.

The samples were then shipped by airfreight to SGS Analabs (“Analabs”) in Kayes, Mali.

During Defiance’s RC and diamond drill program, the analytical work was carried out by Analabs in Kayes, Mali and by Abilabs located in Bamako, Mali. Analabs is an ISO accredited laboratory whereas Abilab is not ISO accredited.

A total of 21,686 RC sample pulps, including field duplicates, blanks and standards, and 904 DDH core sample pulps, including field duplicates, blanks and standards, were shipped in 16 batches, of which 14 went to Analabs and 2 went to Abilabs. Included within these sample batches were a total of 774 field duplicate samples, each one being a second split from a 1 metre interval field sample bag, and 1,136 preparation duplicates, each one being a second split from the pulverised RC and core sample at the preparation laboratory. All of the sample pulps were analysed for gold using a 50 g fire assay with an AAS finish at both laboratories. The Analabs 50 g fire assay/AAS method (FA50) has a lower detection limit of 0.005 g/t Au compared with Abilabs lower detection limit is 0.010 g/t Au.

Analabs routinely ran random check assays in all batches. However, when the laboratory was notified of possible samples containing high values of gold for the core samples, Analabs carried out a fire assay/AAS method, with repeats in some case, as well as fire assay/gravimetric analysis for samples grading greater than 5.00 g/t Au. Analabs also provided Defiance with its internal QA-QC data during the analysis period.

12.3 2007 Red Back

Closely following Red Back’s acquisition of the project in August 2007, the on site SGS Analabs assay facility became operational. Prior to that time samples had been prepared on site by staff of TMLSA under supervision of senior geological staff. Since that time samples have been prepared and analysed under contract by SGS on site and by SGS Kayes, Mali and SGS Ouagadougou, Burkina Faso.

Samples, including duplicates, blanks and certified reference materials are delivered daily from the drill rig to a secure storage area within the Tasiast office complex.

SGS Tasiast Procedures

The entire RC sample bag is oven-dried for 24 hours, and then weighed prior to pulverization. A 1.5kg sub sample is split using a Jones riffle splitter and pulverised in a Labtechnics LM2 mill.

Sample pulps are analysed for gold using a 50 g fire assay with an AAS finish with a detection limit of 0.01g/t.

SGS Kayes and Ouagadougou procedures

RC samples are stockpiled in a secure area within the Tasiast office complex and are picked up by a truck contracted by Analabs for shipment to Kayes.

The entire RC sample bag is oven-dried for 24 hours, and then weighed prior to pulverization. Two 1.5kg sub sample are split using a Jones riffle splitter and pulverised in a Labtechnics LM2 mill before the two pulps are re-combined.

Sample pulps are analysed for gold using a 50 g fire assay with an AAS finish with a detection limit of 0.01g/t.

12.4 Summary

The sampling methods, chain of custody procedures, sample preparation procedures and analytical techniques are all considered appropriate and are compatible with accepted industry standards.

13.0 DATA VERIFICATION

13.1 Introduction

The revised Mineral Resources have been estimated using over 2,300 drill holes combining to provide in excess of 300,000 metres of drilling.

Red Back has installed a Century systems database management system under the supervision of an experienced on site database manager. All drill data has been imported into this system and has been re-validated. The drill data for resource estimation purposes was exported as comma delimited ASCII files.

13.2 Historical Data Verification

The following section describes the data verification procedures and quality control measures relating to the historical data that has been incorporated into the current resource calculations.

13.2.1 NLSD Analytical Data

Prior to the commencement of Defiance’s RC and diamond drill program, Midas and SRK (2003) carried out an independent check sampling of the RC and core samples from 1996 to 2001. Samples were submitted to OMAC Labs in Ireland and to SGS France Laboratories (Montpellier) for check assaying.

Most of the gold assays in the NLSD database were initially analysed by Aqua Regia digestion using a 30 g aliquot. All assays >1.0 g/t Au were then re-assayed by a 30 g fire assay with AAS finish.

The earlier sampling programs by NLSD did not incorporate QA-QC (i.e. blanks, duplicates, or standards), although NLSD did carry out a check sampling and analytical program during their 3rd phase of drilling. SRK’s review (2003) of the available data detected no significant problems with the NLSD data. As mentioned by both NLSD’s consultants (BRGM) and by Midas’s consultants RSG Global, more QA-QC work was required to determine sampling error, sample preparation error and assay accuracy.

In early 2003, on the recommendations of RSG Global, Midas collected a total of 429 pulp samples from selected NLSD-drilled mineralized zones. Midas inserted blanks and standards and submitted their sample batch to Genalysis, an ISO Guide 25 accredited laboratory in Perth, Australia. The QA-QC program indicated that some 25% of the total check assay pair data have a precision higher than ± 20%. The Genalysis results also compared well with the NLSD assays and the standards and blanks inserted Midas assayed within acceptable limits.

13.2.2 Defiance Analytical Data

Prior to the commencement of Defiance’s infill drill program, Midas retained RSG Global to design and implement a QA-QC program for the infill drill program. RSG Global’s QA-QC program incorporated the use of standards, blanks and duplicate samples. RSG Global carried out the QA-QC audit of the sampling procedures and assay data until May 3, 2003, after which Howe personnel assumed supervision of the QA-QC monitoring program in conjunction with Defiance personnel.

Samples collected during Defiance’s infill drill program were submitted to SGS Analabs (“Analabs”), in Kayes, Mali, and Abilabs Afrique de l’Ouest SARL, in Bamako, Mali. The primary laboratory, Analabs, is an ISO accredited laboratory and employs a procedure of internal submission of standards and blanks as well as carrying out repeat assays on approximately 10% of the client submitted samples. Defiance was provided with Analabs internal QA-QC data for comparison with Defiance’s QA-QC standards and blanks. Standards were inserted at every 20th sample and internally prepared coarse blank sand was inserted at every 10th sample within the RC and core sample stream.

RSG Global audited both laboratories in April 2003 and reported that the quality of operations at Analabs were satisfactory and up to recommended standards. With regard to the Abilab laboratory, RSG Global reported that the laboratory was functioning well but that there were a number of internal laboratory issues in need of improvement.

Repeat assays from Defiance’s RC and diamond drill program included 774 RC field duplicate samples and 1,136 RC and core prep duplicate samples for a total of 1,904 assay pairs. Analysis of the assay pairs show an acceptable repeatability (Table 13.2.2-1). The high correlation coefficient of 0.879 reflects the degree of similarity between the repeat values. However, the high coefficient of variation values of the assay pair data indicates the variability of the grades within the sample data.

	Original Au (g/t)	Repeat Au (g/t)	% Difference
Number of Assay Pairs	1904	1904	0
Maximum	42.600	114.400	-71.800
Minimum	0	0	0
Mean	1.491	1.544	-0.052
Median	0.218	0.240	-0.023
Standard Deviation	3.360	4.131	-0.775
Variance	11.260	17.063	-5.799
Coefficient of Variation	2.250	2.676	-0.426

Table 13.2.2-.1: Summary of Basic Statistics for Duplicate Assays

A bivariate scatter plot of original and repeat RC field duplicate and RC and core pulp duplicate assays (Figure 13.2.2-1) confirms no major bias across all grade ranges.

Figure 13.2.2-1: Scatter Plot of the Tasiast Original versus Duplicate Assays (n=1904)

The Thompson & Howarth precision plot constructed for both the RC field duplicate and RC pulp duplicates and illustrated below in Figure 13.2.2-2 indicate that over 30% of the total duplicate assay dataset between the 1.0 g/t Au and 10 g/t Au range have a precision higher than ± 20%.

Figure 13.2.2-2: Thompson & Howarth Precision Plot of All Duplicate Assay Data

Further examination of the RC field duplicate and RC pulp duplicate assay pair data also indicates that over 50% of the original assays versus RC pulp duplicate assay pair values have a precision higher than ± 50% than that of the original assay versus RC field duplicate assay pairs. This higher precision is attributed to better sample homogeneity of the original RC sample after sample preparation.

At the completion of Defiance’s RC and diamond drill program, ACA Howe re-split and collected a total of 134 one-metre interval RC samples from six RC holes, and 27 core pulp samples from two diamond drill holes for check assay purposes. The samples were submitted

to ALS Chemex Laboratories in Ontario, Canada; an ISO 9002 and ISO 9001:2000 accredited laboratory.

ACA Howe carried out an assay comparison of its RC field duplicate and diamond drill data with the results Defiance received from Analabs for the same assay pair. Figures 13.2.2-3, 13.2.2-4, 13.2.2-5 show bivariate scatter plots of the RC field duplicate data, Figure 14-6 shows a Thompson & Howarth precision plot of the same data.

The scatter plots show a relatively large amount of scatter is the assay pairs across the entire grade range. However, > 85% of the RC field duplicate samples occurs within the ± 50% precision envelopes. The precision plot shows very noisy data with >50% of the RC and diamond drill samples returning relative precision of >±20%.

Figure 13.2.2-3: Scatter Plot of the Tasiast Original vs Howe Duplicate Check Assay

Figure 13.2.2-4: Scatter Plot of the Tasiast Original vs Howe Duplicate Check Assays (cut to 20 g/t Au)

Figure 13.2.2-5: Scatter Plot of the Tasiast Original vs Howe Duplicate Check Assays (cut to 10 g/t Au)

Figure 13.2.2-6: Precision Plot of the Tasiast Original vs Howe Duplicate Check Assays

Defiance also selected mineralized intersections from 30 RC holes covering the four mineralized areas of the Piment Zone, which were sent to SNC Lavalin in Canada for metallurgical test work. SNC reviewed the drill hole information on the geological sections prepared by ACA Howe and combined the sampled intersections of several drill holes to obtain nine samples considered to be more or less representative for the various ore zones and their high and low gold grades. These samples were sent to SGS Lakefield (“Lakefield”) in Ontario, Canada; an ISO/IEC 17025 accredited laboratory for assay. The comparison of the assay results of the initial samples and those from Lakefield was acceptable and showed a reasonable correlation. Table 13.2.2-2 shows the comparison of the calculated assays from the drill hole data with those from Lakefield.

	Assay (g Au/t)		Difference
Sample	Calculated	Lakefield	(%)
1	3.32	2.94	-11
2	6.53	5.00	-15
3	4.12	3.84	-7
4	1.25	1.48	+18
5	3.05	3.30	+8
6	4.14	4.77	+15
7	3.64	3.92	+8
8	5.34	5.56	+4
9	6.57	5.75	-11
Average	4.22	4.06	-4

Table 13.2.2-2: Comparison of Assay Data of RC Samples

During the course of the Feasibility study SNC representatives also collected eight samples of RC drilling chips that had previously been assayed by Analabs. These samples were sent to Lakefield for assay. These independently collected samples showed that gold was present in the indicated mineralized zones even though the correlation was rather erratic due to the statistically low number of samples. Table 13.2.2-3 shows a comparison of the results of these sample tests.

	Between	Assay (g Au/t)		Difference
Hole	Metres	Lakefield	Analab	(%)
SR353	83 – 84	1.80	7.82	+334
SR362	95 – 96	7.06	1.97	-72
SR421	13 – 14	3.89	4.38	+13
SR423	55 - 56	6.05	2.22	-63
SR434	40 – 41	3.19	3.25	+2
SR466	54 – 55	1.77	3.64	+106
SR518	62 - 63	13.40	5.45	-59
SR572	36 – 37	7.44	10.50	+41
Average		5.58	4.90	-12

Table 13.2.2-3: Comparison of Assay Data of Samples Collected by SNC

13.2.3 Red Back Data Verification

Red Back conducted a further analysis of the available, historical QAQC data (Defiance and Rio Narcea) as part of the February 2008 resource update comparing all historical data with data generated by Red Back up to February 2008. The following conclusions were noted:

Globally, all of the laboratories used to compile the Tasiast resource have reported the ore grade standards well. On average 85% of the six +1.5g/t internationally accredited CRM standards submitted, reported to within an accuracy of +/-10%. The historical Pre-Red Back database reports 86%. The Red Back database reports 84%.

A minor negative bias is repeated in each of the standards tested at each of the labs. The negative direction of the standards bias, however, results in a degree of conservatism in the assays reported.

The <1.0g/t standards perform less well with a range of 67-75% of the standards submitted reporting to within +/-10%. The poorer precision and accuracy of the QAQC data below 1.0 g/t is evident across the whole of the resource timeframe.

Both routine Red Back and Pre-Red Back blank submissions performed well, exhibiting only a minor low level <50ppb Au cross-contamination. Evidence suggests a component of poor blank selection may have contributed to the Red Back higher bias and was noted for further attention.

The Total Operational Precision (TOP) achieved by Red Back, demonstrated by a %MAHD = +/- 14% of resource grade assays > 0.2 g/t, is generally within acceptable limits of a coarse gold deposit such Tasiast. The coarse gold nature of the deposit is apparent within the range of errors expressed by the 90th Percentile AHD = +/- 58%.

Analysing the historical pre-Red Back duplicate data it was observed that the historical resource data reported similar “nuggetty” duplicate assaying, closely comparative to the RBK data with %MAHD = +/- 14% and a P90 AHD = +/- 50% > 0.2 g/t. The datasets have equivalent coarse gold features and equivalent assaying precision.

The imprecision consequent of coarse gold is evident across the entire Tasiast grade profile from 0.2 g/t to 10 g/t. Clustering of “nuggetty assaying” is often observed in mesothermal greenschist facies, epigenetic, structurally controlled deposits at the high grade end of the profile due to the coarse gold being hosted dominantly in the quartz vein materials, the gold being finer, closer to sulphide lattice within the disseminated selvages.

Red Back considered the extant of the assay data included in the Tasiast resource data to be accurate and precise to within the inherent, natural coarse grade variation observed in the grade profile of this structurally controlled, sub-amphibolite, BIF hosted style of gold mineralisation.

13.2.4 2009 Red Back QAQC Data

The data presented below covers all QAQC data for 2009.

Red Back maintained 10% total QC through the sample stream.

A total of 16,907 QAQC samples including standards (STD), blank (BLK) and duplicates (Dups) were blindly inserted as part of the routine sample preparation and were submitted for analysis. Due to the volume of samples, three different SGS laboratories, namely Tasiast, Kayes and Morilla in Mali were used for the sample analysis (shown in table 13.2.4 below).

	std	blanks	dups
SGS KAYE	4972	5359	572
SGS TASIAST	566	577	108
SGS MORILLA	2007	2279	467
Total	7545	8215	1147

Table 13.2.4: Sample statistics by laboratory.

13.2.5 SGS KAYES

Kaye SGS is the principal laboratory used for Tasiast Exploration sample analysis and is located in Mali. 10,903 QAQC samples were submitted blindly for analysis.

The results are shown in the graphs below:

BLIND FIELD STANDARDS

303 samples of the 0.208g/t CRM standards were submitted but 296 out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 77.4%

604 samples of the 0.410g/t CRM standard were submitted but 596 out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 97.8%.

The 0.416g/t CRM standard with a total of 60 were submitted and all the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 100%.

The 0.72g/t CRM standard with a total of 82 were submitted and all the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 100%.

A total of 794 samples of the 0.73g/t CRM standard were submitted but 788 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 92.8%. High bias treatment was evident in the early part of the year before the performance became regularised. There was an evidence of batch drift.

A total of 723 samples of the 1.34g/t CRM standard were submitted but 712 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 99.5%.

A total of 417 samples of the 1.76g/t CRM standard were submitted but 412 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 96.9%.

The 3.33g/t CRM standard with a total number of 673 were submitted but 662 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 96%. Even though most of the values were within the limits, the distribution during the later part of the year was better than the beginning of the year. There was an evidence of high bias from the beginning to the middle part of the year (Batch drift).

The 4.5g/t CRM standard with a total number of 567 were submitted but 562 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 96%.

The 6.76g/t CRM standard with a total number of 598 were submitted but 588 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 96%. This is an excellent performance as almost all the values reported are close to the mean value.

A total of 97 samples of the 6.78g/t CRM standard were submitted but 96 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 100%. The distribution of the values was very close to the mean value.

The 7.635g/t CRM standard with a total of 54 were submitted and all the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 69.6%. There was is a clear evidence of low bias in the treatment of this standard.

FIELD BLANKS

The general performance by Kaye SGS on the field blanks that were submitted blindly a good one as 99.5% of the samples reported had values less than or equal to 0.02g/t. Blanks of about 5359 submitted, 5328 were reported with only 0.5% of the samples reporting values greater than 0.02g/t.

DUPLICATES

The Total Operation Precision (TOP) of the duplicates as expressed by a 50th percentile (MAHD) is ±18.15% of resource grade assays >0.1g/t and is generally within the acceptable limits of a coarse gold deposit such as Tasiast. The coarse nature of the deposit is expressed within the range of errors expressed by the 90th Percentile AHD = ±51%. TOP sampling precision of P90 +/-50% is equivalent for mesothermal, coarse, quartz-veined, gold deposits globally.

13.2.6 SGS TASIAST

TML SGS is one of the labs used for the sample analysis and is located at the mine site in Mauritania. 1251 QAQC samples including standards, blank and duplicates were submitted blindly for analysis.

The results are shown in the graphs below:

BLIND FIELD STANDARDS

The 0.208g/t CRM standard was submitted 20 times and all the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 84.6%.

The 0.410g/t CRM standard was submitted 20 times and all the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 80%.

A total of 94 samples of the 0.73g/t CRM standard were submitted but 87 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 72.5%.

A total of 43 samples of the 1.34g/t CRM standard were submitted but 42 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 88.5%. The point above the upper limit appears to be a gross error.

The percentage of the number of samples that reported within the ±10% boundaries out of the 13 CRM standard 1.76g/t submitted is 69.2%.

The 3.33g/t CRM standard with a total number of 202 were submitted but 194 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 78.9%. Just as a low bias in the treatment of this standard is prominent, the values observed to be too erratic. Some of the points outside the limits could be gross errors. Supervision at TML lab should be strengthened.

The 4.18g/t CRM standard with a total of 18 were submitted and all the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 98.8%. The chart illustrates an evidence of a low bias in the treatment of these samples. The point below the lower limit appears to be a gross error.

The 4.5g/t CRM standard with a total number of 59 were submitted but 50 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 84%. Some of the points outside the limits could be gross errors.

The 7.76g/t CRM standard with a total number of 94 were submitted but 89 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 85.4%.

FIELD BLANKS

TML SGS performance saw a tremendous improvement in the treatment of the field blanks that were submitted blindly. Blanks of about 577 submitted, 568 were reported with only 1.8% of the samples reporting values greater than 0.02g/t.

DUPLICATES

The Total Operation Precision (TOP) of the duplicates as expressed by a 50th percentile (MAHD) is ±16.67% of resource grade assays >0.1g/t and is generally within the acceptable limits of a coarse gold deposit such as Tasiast. The coarse nature of the deposit is expressed within the range of errors expressed by the 90th Percentile AHD = ±57.23% (See table 4) and is high since the acceptable limit the industry should not exceed ±50% for P90.

13.2.7 SGS MORILLA

Morila SGS laboratory is also used for sample analysis and is located in Mali. 4,755 QAQC samples including standards, blank and duplicates were submitted blindly for analysis.

Results are shown in the graphs below:

STANDARDS

A total of 102 samples of the 0.208g/t CRM standard were submitted but 101 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 87.2%. The point below the lower limit appears to be a gross error.

A total of 339 samples of the 0.410g/t CRM standard were submitted but 336 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 98.5%.

The 0.72g/t CRM standard with a total number of 152 were submitted but 148 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 97.9%.

The 0.73g/t CRM standard with a total number of 250 were submitted but 248 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 92.3%.

A total of 384 samples of the 1.34g/t CRM standard were submitted but 379 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 97.2%. The two points below the lower limit appears to be a gross error.

The 1.76g/t CRM standard with a total of 63 were submitted and all the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 100%.

A total of 203 samples of the 0.41g/t CRM standard were submitted but 201 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 99.5%. The point below the lower limit appears to be a gross error.

A total of 161 samples of the 4.5g/t CRM standard were submitted but 159 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 99.4%.

A total of 244 samples of the 6.76g/t CRM standard were submitted but 242 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 74.4%. The chart illustrates the presence of high biasness in the treatment of almost all the batches.

A total of 102 samples of the 6.78g/t CRM standard were submitted but 282 samples out of the values reported were used for plotting the chart. The percentage of the number of samples that reported within the ±10% boundaries is 99.1%. The chart illustrates the presence of high biasness in the treatment of the batches.

FIELD BLANKS

The performance by Morila SGS on the field blanks that were submitted blindly a good one as 98.4% of the samples reported had values less than or equal to 0.02g/t. Blanks of about 2279 submitted, 2266 were reported with only 1.6% of the samples reporting values greater than 0.02g/t.

DUPLICATES

13.2.8 Conclusion

2009 TASIAST RESOURCE QAQC

	TOTAL	STD within		BLANKS		DUPS		DUPS
	QAQC	+/-10%		<20 ppb		MAHD P50		MAHD P90
KAYE	11189	94	%	100	%	18	%	51	%
TASIAST	1305	84	%	98	%	17	%	57	%
MORILLA	4987	96	%	98	%	17	%	56	%
	17481	94	%	99	%	+/-18	%	+/-53	%

Table 13.2.8-1: Tasiast Resource QAQC Summary

In summary, the QAQC data reported is of industry accepted standards and the assay data is considered reliable for inclusion in the December 2009 resource estimation.

14.0 ADJACENT PROPERTIES

There are no mineral exploration properties owned by mining/exploration companies adjacent to the Tasiast Permit area.

15.0 MINERAL PROCESSING AND METALLURGICAL TEST WORK

The metallurgical characteristics as developed through metallurgical studies have largely been confirmed by operating experience.

Several characteristics are routinely evaluated:

· Screen analysis is performed daily on a sub-sample of Ball Mill Feed and CIL Feed as an indicator of mill performance.

· CIL Recovery vs Leach time is tested using standard bottle roll tests on weekly composite samples of CIL Feed and CIL Tails.

Samples are sent as required to external laboratories for both Bond Work Index (BWI) and SMC testing. The latter test gives a Mia value to fully evaluate ore hardness characteristics.

The testwork is performed by an experienced metallurgical technician and supervised by a Metallurgical Engineer. Both grinding and CIL results are reported in a standard format.

15.1 Mineral Processing

The process flow route comprises three stage crushing, wet ball mill grinding, thickening, carbon in leach (CIL), gold recovery by the Zadra elution/electro-winning system and water recovery from tailings.

15.1.1 Metallurgical Test Work

The CIL metallurgical testing was performed on the 9 composite samples, which were subsequently combined into 6 composites representing the various ore zones. To obtain sufficient sample weight for cyanide destruction, thickening and filtration tests, two super-composites representing the main geological features i.e. the oxide and primary ores were also prepared.

The main findings are summarised below.

· Oxide ores are amenable to grinding in a SAG/ball mill circuit while the introduction of a crusher to give an SABC (SAG/ball mill/ pebble crusher) would be recommended to handle primary ores;

· Magnetic separation showed poor selectivity with respect to gold;

· Both oxide and primary ores contain coarse gold and accordingly, are amenable to gravity concentration, an average recovery of 22.6% (range 4-42%) being obtained when the samples were treated in a laboratory Knelson concentrator with the gravity concentrate being further upgraded on a Mozley mineral separator;

· Initial work indicated pre-aeration before cyanidation to be beneficial but further work demonstrated that the combined effect of gravity concentration followed by CIL was equal or superior to the effect of pre-leaching followed by carbon in pulp (CIP). The use of the former allows a reduction in the number of tanks compared to pre-aeration/CIP, thus has economic attractions;

· All samples when ground to an 80% passing size of 90 µm gave an overall gold recovery of 95% or higher when subjected to gravity concentration followed by cyanidation of the gravity tailings;

· A cyanide contact time of 36 hours gave significantly higher recoveries than one of 24 hours: compared to the 36 hour results, extending the contact time to 48 hours gave improved recovery ranging from 0.1-1.4% in half of the six tests and poorer results

ranging from 0.7- 3.9% in the other half. On the basis of this work a contact time of 36 hours was selected;

· Generally, the results of cyanidation tests conducted in saline water from site showed little difference when compared to those obtained when the same samples were tested under identical conditions using tap water. However, as might be expected, the use of saline water resulted in a slight increased consumption of reagents, the average cyanide consumption rising from 0.28 kg/t to 0.39 kg/t while that of lime rose from 0.69 kg/t to 0.97 kg/t;

· Elution with a strip solution of sodium hydroxide and sodium cyanide dissolved in saline water unexpectedly proved more efficient than the same solution prepared with deionised water but it was proposed that deionised water be used since a white precipitate, which could result in serious scaling problems in the plant, was formed on addition of the reagents to the saline water;

· The cyanide content of tailings liquor can be reduced to 1 mg/I CNWAD by use of sodium metabisulphite and air with copper as catalyst. The general guideline from the Cyanide Management Code specifies that a concentration of less than 50 mg/I CNWAD is acceptable;

· The cyclone overflow obtained by grinding oxide ore will require a large surface area for thickening and the underflow density will be relatively low. The primary ore, on the other hand, can be thickened to about 64% density with the same flocculant dosage. The thickening rates were not greatly dependent on water quality or pH;

· Filtration of the thickener underflow slurries from both slurries will be impractical due to the large filter area required to treat the oxide material;

· The poor thickening characteristics of oxide material after cyanide destruction in both high rate and paste thickeners, precludes water recovery from tailings. Further investigations with respect to tailings disposal are underway in order to maximize water recovery from the tailings.

In November 2008 Kappes, Cassiday & Associates’ (KCA) visited site and carried out a review of applicable metallurgical testwork that had been carried out at site to establish the gold recovery that can be achieved from oxide ore using a conventional Run-of Mine (ROM) dump leaching method.

The testwork focussed on lower grade material, i. e. material which is below the economic cut-off grade for the CIL mill circuit, and which has to be mined. The oxide portion of this lower grade material would be moved directly to a conventionally designed dump leach pad and to leach this low grade material to recover the cyanide soluble gold.

The testwork samples originated from the Piment South South (PSS), Piment South North (PSN) and the Piment Central (PC) mine areas. The dump leach material constitutes a low grade “halo” that typically is located adjacent to or in association with the higher grade ore zones that are being mined for feed to the CIL mill. The ore zones are hosted predominantly in BIM-BIF (Banded Iron Formation) or BIMBIF/ GST (Banded Iron Formation/Greenstone) lithologies.

A number of column tests and two larger scale simulated ROM dump leach field trials or pilot tests have been carried out by TMLSA at site. The metallurgical tests confirm that the low grade mineralized portions of the BIM-BIF and BIM-BIF/GST lithology group will respond well to ROM dump type leaching.

In KCA’s opinion the ROM “Pilot Test” or DUMP LEACH Test 1 that was carried out at site on 1,648 tonnes of ROM type material mined from the PSS area provides the confirmation that this type of material is amenable to ROM dump leaching. Based on the available metallurgical testwork using this methodology KCA recommended that TMLSA uses a 75% gold recovery value to forecast the ultimate gold production from their ROM dump leach.

15.2 Refining

Gold production from the Tasiast Mine is shipped and refined under contract by MKS Finance of Zurich, Switzerland at their PAMP refinery in Switzerland.

16.0 Mineral Resources and Mineral Reserves

16.1 Revised Mineral Resource Estimate

Hellman & Schofield Pty Ltd (H&S) was retained by Red Back to estimate recoverable gold resources at the Tasiast Gold Mine. The new resource estimates are required for open pit mine optimisation studies and mine planning purposes.

Estimates were prepared with reference to the Canadian Institute of Mining Metallurgy and Petroleum (CIM) Definition Standards (2005) and CIM Best Practice Guidelines (2003) for preparing Mineral Resources and Mineral Reserves.

Under JORC reporting requirements and guidelines Nicolas James Johnson, a member of the Australian Institute of Geoscientists, with more than five years experience in the use of geostatistics for estimation of recoverable resources in gold deposits, is the Competent Person for the purposes of this work. Mr Johnson visited the mne in February 2008.

The recoverable resource models have been built using GS3©, the MIK software developed by Hellman and Schofield Pty Ltd, and are suitable for use in open pit optimisation studies.

16.1.1 Resource Data Sets

The Tasiast drillhole database is maintained through a Century Systems Database managed on site by the exploration team at Tasiast.

The data set represents all drill hole and assay information available at 14th February 2010.

16.1.2 Block Model

The block model was constructed using dimensions of 15m (east) by 25m (north) by 5m (RL).

16.1.3 Geological Interpretation

Geological wireframes were generated using Micromine software by site geologists. Sectional interpretations based on a 25m section spacing were wireframed to provide a seamless geological model for the resource area. This model was then used to code resource blocks by lithology, bulk density and metallurgical recovery.

Logging available in the geological data base has been used to interpret the oxidation profiles on cross-sections and these were then joined to form surfaces. Red Back has provided H&S with a surface that separates the transition material into upper and lower transition and which is based on metallurgical criteria. Metallurgical test work has demonstrated that the upper transition material is amenable to Dump Leaching and is therefore included with oxide material for modelling purposes.

Red Back has interpreted wireframes to represent barren dykes that intrude the Tasiast mineralisation. H&S have used these wireframes to remove the dyke volume from the resource model.

16.2 Data Preparation and Treatment

16.2.1 Mineralisation Wireframes

The interpretation of mineralised zones was based on the resource composites and based primarily on separating broad mineralised areas from un-mineralised areas in the resource data. A series of mineralised lenses were interpreted over the seven kilometres established strike length of gold mineralisation at Tasiast, delimiting mineralised zones based on a combination of logged geology (footwall contact), identifying mineralisation of similar tenor and directional trends seen in the assay data and, where required, a nominal cut-off grade of about 0.2g/t Au. Outlines were digitized on E-W cross-sections, with points snapped to drill traces in three-dimensions and those outlines then joined to form three-dimensional wireframes.

Table 16.2.1 lists the domain identifier used in the resource model with the respective deposit names and domain extents in northings.

Domain		Description
0		Peripheral weakly/barren mineralised zones
1		West Branch South
2		West Branch (Southern Footwall Zone)
3		West Branch (Southern Hangingwall Zone)
4		West Branch (Central Zone)
5		West Branch (Northern Footwall Zone)
6		West Branch (Northern Hangingwall Zone)
7		West Branch (East Zone)
8		Piment South South (Footwall Zone)
9		Piment South South (Main Zone)
10		Piment South North
11		Piment Central (Footwall Zone)
12		Piment Central (Main Zone)
13		Piment North

Table 16.2.1: Tasiast resource modeling domains

16.2.2 Compositing

H&S composited the resource data into two metre down hole intervals before commencing the model building process. Un-sampled or un-assayed intervals in the dataset are the result of un-sampled intervals which are peripheral to the study area so have no bearing on the resource model. These intervals were assigned “-9999” grade before compositing and resulting composites with negative grades were removed from the resource data set.

For each resource composite a gold composite grade is recorded and located by the mid-point de-surveyed coordinate. In addition to the gold grade the composites are identified as belonging to either one of 14 modelling primary domains and also to either the oxide+upper transition, lower transition or fresh horizon, 1, 2 or 3, respectively (secondary domains). The primary domaining were assigned via 3D solid models (wireframes) and the secondary domains have been coded from a DTM surface, both described in an earlier section.

16.3 Exploratory Data Analysis

Exploratory data analysis (EDA) consisted of histograms, frequency plots, box plots and summary statistics.

The mean grade of resource composites flagged as Domain 0 is very low, having an average grade less than 0.08g/t Au indicating that the domaining has effectively separated out the bulk of the gold mineralisation at Tasiast.

The mean sample grades of the mineralised domains range from 0.4 to 2.0g/t Au and show coefficients of variation (CV) generally in the range of about 1.5 to 4.0, which are high and are typical for gold deposits with gold mineralisation similar to that seen at Tasiast. CV’s at these levels indicate that reliable estimation of recoverable gold grades using a linear estimator would be difficult.

The grade populations of all domain and secondary domain combinations show the positive skewness typical of gold deposits and moderate to high maximum composite grades. Due to the nature of the gold mineralisation, coupled with the high CV’s for the mineralisation, H&S have chosen to exclude some high composite grades from the data sets used to calculate the indicator statistics used in the MIK model and It has also been chosen to accept the median as the average grade of the highest indicator class for input into the MIK model.

16.4 Spatial Continuity Analysis

16.4.1 Measures of Spatial Continuity

Most resource estimation methods use a measure of spatial continuity to estimate the grade of blocks in a resource model. In some methods the measure is implicit; for example, a polygonal method assumes that the grade is perfectly continuous from the sample to its surrounding polygon boundary. Geostatistical methods like Ordinary Kriging and Indicator Kriging are among those methods for which the continuity measure is explicit and is customised to the data set being studied. This measure in its many forms is usually called the variogram.

16.4.2 Directional Controls on Gold Mineralisation

Gold and indicator variograms were calculated and modelled for the various data subsets within the mineralised trend. Results are summarised below:

West Branch mineralisation: The spatial continuity model indicates that the dominant control on the gold mineralisation is within a moderate east dipping plane. Variogram ranges in the along strike and down dip direction appear isotropic.

Piment South North: The spatial continuity model indicates that the dominant control on the gold mineralisation is within a moderate east dipping plane. Variogram ranges in the along strike and down dip direction appear isotropic.

Piment Central: The spatial continuity model indicates that the dominant control on the gold mineralisation is within a plane dipping moderately towards approximately grid north.

Piment North: The spatial continuity model indicates that the dominant control on the gold mineralisation is the east west direction with continuity in cross section tending to be approximately isotropic.

16.5 Resource Estimate

16.5.1 Indicator Kriging for Recoverable Resources

The MIK method was developed in the early 1980’s with a view toward addressing some of the problems associated with estimation of resources in mineral deposits. These problems arise where sample grades show the property of extreme variation and consequently where estimates of grade show extreme sensitivity to a small number of very high grades. These characteristics are typical of many lode gold deposits, where the coefficient of variation in samples normally exceeds 2. MIK is one of a number of methods that can be used to provide better estimates than the more traditional methods such as ordinary kriging and inverse distance weighting.

It is fundamental to the estimation of resources that the estimation error is inversely related to the size of the volume being estimated. To take the extreme case, the estimate of the average grade of a deposit generated from a weighted average grade of the entire sample data set is much more reliable than the estimate of the average grade of a small block of material within the deposit generated from a local neighbourhood of data.

Another fundamental notion relevant to the optimisation of resources to develop an open pit mine and schedule is that the optimisation algorithm does not require the resource be defined on extremely small blocks relative to data spacing. Small blocks cannot provide the basis for reliable estimates of recoverable resources.

The basic unit of an MIK block model is a panel that normally has the dimensions of the average drill hole spacing in the horizontal plane. The panel should be large enough to contain a reasonable number of blocks, or Selective Mining Units (SMUs; about 15). The SMU is the smallest volume of rock that can be mined separately as ore or waste and is usually defined by a minimum mining width. At Tasiast, the dimensions of this block are assumed to be in the order of 3mE x 5mN x 2.5mRL.

The goal of MIK is to estimate the tonnage and grade of ore that would be recovered from each panel if the panel were mined using the SMU as the minimum selection criteria to distinguish between ore and waste. To achieve this goal, the following steps are performed:

1. Estimate the proportion of each domain within each panel. This estimation can be achieved by kriging of indicators of domain classifications of sample data points. In the Tasiast model proportions of each domain in each panel were calculated by passing the panels through the domain wireframes.

2. Estimate the histogram of grades of sample-sized units within each domain within each panel using MIK. MIK actually estimates the probability of the grade within each panel being less than a series of indicator threshold grades. These probabilities are interpreted as panel proportions.

3. For each domain, and for each panel that receives an estimated grade greater than 0.0g/t Au, implement a block support correction (variance adjustment) on the estimated histogram of sample grades in order to achieve a histogram of grades for SMU-sized blocks. This step incorporates an explicit adjustment for Information Effect.

4. Calculate the proportion of each panel estimated to exceed a set of selected cut-off grades, and the grades of those proportions.

5. Apply to each panel, or portion of a panel below surface, a bulk density to achieve estimates of recoverable tonnages and grades for each panel.

Apart from considerations of resource confidence classification (Section 16.7), Step 5 completes construction of the resource model. The estimates of recoverable resources for

each panel may be combined to provide an estimate of global recoverable resources for the deposit.

16.5.2 Indicator Kriging Parameters

The input parameters to the Tasiast Indicator Kriging model of the gold mineralisation include;

· resource data file containing metal grades located by their mid-point coordinates together with a primary domain code and secondary oxidation code.

· appropriate panel size, search and data constraint selection given the density of the available sample information.

· the indicator variogram models for the gold grade distributions within each domain,

· the class means for the various indicator classes on which the indicator variograms are based with orientations customised for each modelling domain.

· Bulk Densities for use in converting estimated volumes to tonnes.

16.5.3 Bulk density modeling

The BD’s in the current study were assigned to the resource panels using the following procedure;

· A lookup table of average BD generated based on 495 determinations with BD’s averaged across each major lithology type and also separated into oxidation level.

· Wireframes of lithology types interpreted from the logged lithology of the drill holes and provided the estimates of proportions of each lithology type within each resource panel.

· The interpretation of oxidation surfaces provides the proportion of oxide, upper and lower transition and fresh material within each resource panel.

16.6 Block Support Adjustment (Variance Adjustment)

The block support adjustment is one of the most important properties of a recoverable resource model based on non-linear estimation methods like MIK. It is an essential part of the model and involves important assumptions about the nature of the block grade distribution within each panel of the model.

Indicator Kriging provides a direct and reliable estimate of the histogram of grades of sample-sized units within each panel of the model provided the panel dimensions are of an appropriate size. However, ore is not selected on sample-sized units during mining; it is selected by shovels that have a minimum mining width and loaded into trucks that are despatched to either ore or waste. The flexibility of digging equipment and the size of the trucking equipment provide an indication of the size of the smallest block of rock that will be mined as ore or waste. To estimate with some accuracy the resources in a deposit that will be recovered with a certain set of mining equipment, the histogram of grades of sample-sized units in a panel provided by MIK must be adjusted to account for the size of the mining block.

There are a number of adjustment methods that can be used and most of these are described well in Journel & Huijbregts (1978) or Isaaks & Srivastava (1989). These methods make three reasonable assumptions:

· The average grade of sample-sized units and blocks within the panel is the same and is equal to the estimated average grade of the panel.

· The variance, or spread, of the block grades within the panel is less than the variance of grades of sample-sized units within the panel and the change of variance from sample-sized units to blocks can be calculated from the variogram of gold grades.

· The approximate shape of the histogram of block grades can be reasonably predicted by some appropriate assumptions.

16.6.1 The Variance Adjustment

The size of the variance adjustment needed to obtain the variance of the block grade distribution within the panel can be calculated using the rule of additivity of variances, which in the case of block support adjustment is often called Krige’s Relationship:

Var(samples in a panel) = Var(samples in a block) + Var (blocks in a panel)

The variance of sample grades in a panel and the variance of samples within a block can be directly calculated from the variogram of gold grades for the particular domain. The ratio of Var(blocks in panel) to Var(samples in panel) is that required to implement the block support adjustment.

16.6.2 Shape of the Block grade Distribution

There are a number of rules of thumb that are useful when making judgements about the shape of the block grade distribution within each panel and they relate to the size of the variance adjustment ratio:

1. If the variance adjustment ratio is greater than 0.7, it may be useful to assume that the shape of the histogram of block grades is similar to that of the histogram of grades of sample-sized units. This is known as the Affine Correction method. Its application to gold deposits is usually inappropriate.

2. If the variance adjustment ratio is between 0.3 and 0.7 and the information adjustment is negligible, then the Indirect Lognormal Correction method of Isaaks & Srivastava (1989) can be useful. This is a rule of thumb based on the experience of the authors.

3. If the variance adjustment ratio is less than 0.3, a high degree of symmetrisation in the block grade histogram will occur and a lognormal assumption (Journel & Huijbregts, 1978, page 481) for the shape of the block histogram is an appropriate choice. This model is well supported by reconciliation studies of resource and grade control models.

16.6.3 The Information Effect

The variance adjustment described above is only part of the adjustment required in many gold deposits because the short scale variation in gold grades is extreme, as is the case at Westonia. This variance adjustment provides an estimate of the variance of true block grades under the assumption that grade control selection will operate with knowledge of the true block grades. While this assumption is never absolutely true, it can be a reasonable assumption in some deposits where the short scale variability is small and the grade control sampling density is high. In many deposits, however, an additional variance adjustment must be undertaken to account for the “Information Effect”.

In the absence of production information or grade control sampling, the Information Effect ratio is based on the variograms of gold grade and on the grade control sample spacing expected to be used during mining.

16.6.4 Variance Adjustments Applied to the Tasiast Model

On this basis Variance adjustment ratios have been applied in estimating the Tasiast recoverable gold resources. These ratios have been applied using the Direct Lognormal Correction method (Journel & Huijbregts, 1978, page 481). Selective mining (SMU) dimensions of 3mE x 5mN x 2.5mRL and grade control sample spacing of 6mE x 10mN x 1mRL have been assumed. The variance adjustments applied to the models represent large reductions of variance and in H&S experience are typical of shear hosted gold deposits.

16.7 Resource Classification

Panels in the resource models have been allocated confidence categories based on the number and location of samples used to estimate proportions and grade of each panel. The approach is based on the principle that larger numbers of samples, which are more evenly distributed throughout the search neighbourhood, will provide a more reliable estimate. The number of samples and the particular geographic configurations that may qualify the panel as Measured rather than Indicated or Inferred are essentially the domain of the Qualified Person. The search parameters used to decide the classification of a panel resource in this study are:

· Minimum number of samples found in the search neighbourhood.

For Measured and Indicated resources, this parameter is set to sixteen. For Inferred category, a minimum of eight samples is required. This parameter ensures that the panel estimate is generated from a reasonable number of sample data.

· Minimum number of spatial octants informed.

The space around the centre of a panel being estimated is divided into eight octants by the axial planes of the data search ellipsoid. This parameter ensures that the samples informing an estimate are relatively evenly spread around the panel and do not all come from one drill hole. For Measured and Indicated resources, at least four octants must contain at least one sample. For Inferred panels, at least two octants must contain data.

· The distance to informing data.

The search radii define how far the kriging program may look in any direction to find samples to include in the estimation of resources in a panel. Panel dimensions and the sampling density in various directions usually influence the length of these radii. It is essential that the search radii be kept as short as possible while still achieving the degree of resolution required in the model. For Measured resources, the easting, northing and elevation search radii were set to 20, 35 and 10 metres respectively. For Indicated and Inferred resources the radii were expanded by 50 per cent to 30mE x 52.5mN x 15mRL.

At Tasiast, the majority of the main corridors of gold mineralisation have been drilled at 25m x 25m spacing or closer, and in these areas the MIK model blocks report to Measured category, most panels in areas consistently drilled at 50m x 50m spacing or less report to Indicated category and panels in peripheral areas and at depth with less consistent drill coverage report to Inferred category.

16.8 Mineral Resource Statement

Table 16.8.1 lists estimated recoverable resources at Tasiast below the December 2009 mining surface. Estimated panel proportions and grades above cut-off are considered recoverable by mining and application of ore loss and dilution factors is not recommended.

Table 16.8.1: Mineral Resource estimate for Tasiast

Cut-

Measured

Indicated

Measured + Indicated

Inferred

Zone

Off

Au g/t

Moz

Au g/t

Moz

Au g/t

Moz

Au g/t

Moz

Oxide

0.2

20.99

0.83

0.56

22.23

0.70

0.50

43.22

0.76

1.06

6.28

0.6

0.12

Fresh

0.5

41.23

1.55

2.05

70.41

1.50

3.40

111.64

1.52

5.45

26.52

1.4

1.18

Total

62.22

1.30

2.61

92.64

1.30

3.90

154.86

1.30

6.51

32.8

1.24

1.30

All ore types

1.0

27.55

2.17

1.93

41.87

2.13

2.86

69.42

2.15

4.79

13.38

2.1

0.91

· The Company reports resources on the basis of mining cut-off grades to be applied to the various ore types and, for comparison purposes, at a 1.0 g/t cut-off grade

· Minerals Resources are reported below the December 31, 2009 mined surface.

· Figures may not add correctly due to rounding.

· Oxide is referred to as material amenable to Dump Leaching and CIL. Fresh is referred to as material amenable to Heap Leaching and CIL.

· The resources are estimates of recoverable tonnes and grades using Multiple Indicator Kriging with block support correction into 15 metres (East) by 25 metres (North) by 5 metres (Elevation) model blocks and assuming smallest mining unit for ore selection in mine grade control of 3 metres (East) by 5 metres (North) by 2.5 metres (Elevation).

· Measured resources lie in areas where drilling is available at a nominal 25 x 25 metre spacing, Indicated resources occur in areas drilled at approximately 25 x 50 metre spacing and Inferred resources exist in areas of broader spaced drilling.

· Gold estimation and model blocks were constrained within geologically derived wireframes.

16.9 Mineral Reserve Estimate

The Tasiast open pit resource model was supplied by H&S in CSV format and converted to Datamine format. AMC conducted a high level review of the resource models in order to assess its suitability for pit optimisation and design, and ore reserve estimation. It should be noted that this review was not a Technical Review, Technical Audit or Technical Due Diligence as defined by AMC but a high level desktop review aimed at ensuring that there are no gross errors in the block model itself.

The Tasiast Open Pit Ore Reserve Estimate (as of 31 December 2009 and estimated in February 2010), reported in accordance with the Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves, prepared by the Joint Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia December 2004 (The JORC Code), is summarised in Table 16.9-1.

Table 16.9-1: Tasiast Open Pit Ore Reserve Statement

	Tonnes	Au	In situ Au
Classification	(Mt)	(g/t)	(koz)
Proved	49.4	1.36	2.17
Probable	61.5	1.40	2.77
Stockpiles	4.3	0.68	0.09
Total	115.2	1.36	5.03

· Numbers may not add correctly due to rounding

· The Ore Reserve estimate used a gold price of US$800

· Cut-off grades: CIL: oxide 0.75 g/t, fresh 0.81 g/t.

· Cut-off grades: Dump leach oxide cut-off grade varies depending on lithology with the lowest being 0.1 g/t.

A breakdown of the Ore Reserve by processing route is shown in Table 16.9-2.

Table 16.9-2: Breakdown of Ore Reserve by Processing Route

	Tonnes	Au	In situ Au
Classification	(Mt)	(g/t)	(koz)
CIL Circuit
Proven	30.2	1.98	1.92
Probable	39.8	1.96	2.51
Stockpile	0.6	1.48	0.03
Total CIL Circuit	70.7	1.96	4.46

Dump Leach
Proven	19.2	0.40	0.24
Probable	21.7	0.37	0.26
Stockpile	3.7	0.54	0.06
Total Dump Leach	44.5	0.40	0.57

· Numbers may not add correctly due to rounding

· Oxide ore between the cut-off grade, which is dependent on lithology and the lowest being 0.1 g/t and 0.9 g/t was allocated to the Dump Leach Process.

The metal price used in the determination of the ore reserve estimate was US$800/oz for gold.

The ore reserve estimate is based on the mineral resource contained within the ultimate open pit mine design classified as ‘Measured’ and ‘Indicated’ after consideration of all mining, metallurgical, social, environmental and financial aspects of the property. All Proven Ore Reserves have been derived from the Measured Mineral Resource and all Probable Ore Reserves have been derived from the Indicated Mineral Resource.

The sections in this report that relate to ore reserves are based on information reviewed by Patrick Smith. Patrick Smith is a Member of the Australasian Institute of Mining and Metallurgy and is employed by AMC. He has sufficient experience which is relevant to the style of mineralisation and type of deposit under consideration in open pit and underground mining activities (respectively) to qualify as the Competent Person as defined in the 2004 Edition of the “Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves”. He visited the mine site during March 2008.

Patrick Smith consents to the inclusion of the Ore Reserve Statement in this report in the form and context in which it appears in this report.

All optimization parameters were supplied by TMLSA or derived in consultation with TMLSA and are consistent with a nominal 3.0 Mtpa on site mineral processing operation. Pits were re-optimized, with detailed staged and ultimate pit designs developed. Using the inventories from these detailed designs, a mining schedule was produced by TMLSA.

All reserves mentioned are completely included within the quoted resources.

16.10 Non-geological Factors Relevant to Resources and Reserves

The ability to exploit mineral resources and reserves can be affected by many “external” factors. The necessary mining lease has been granted, a final environmental impact statement has been accepted by the Mauritanian authorities. Mining operations commenced in 2007.

Royalties, agreements, encumbrances, government policies and related matters are dealt with in Section 4.4. Taxation is covered in Section 18.6.

17.0 OTHER RELEVANT DATA AND INFORMATION

17.1 Reconciliation

Reconciliation of block model estimates to production is used to evaluate the predictive nature of the block model for future production. Two sets of data are available to compare with the block model estimates: grade control data and milled production data.

Grade control drilling is carried out independent of blast hole drilling, and employs the sampling of material from reverse circulation drill rigs. Grade control drilling using face-sampling hammers is carried out on an 8m x 6m grid. A simulation process is implemented with MP grade control modelling software. MP uses conditional simulation to determine the optimum ore outlines for mining. Optimization is done on the basis of maximizing the mine profitability given the uncertainty in knowing the true grade of an ore block.

A summary of the reconciliation to date between grade control modelling and the Mineral Reserve Model at 1.0g/t cut-off is shown in Table 17.1-1 below.

Table 17.1-1: Reconciliation: Exploration Resource Model to Grade Control Model

Block Model

Grade Control Model

Percentage Difference

All Pits

Tonnage

Grade

Ounces

Tonnage

Grade

Ounces

(1.0g/t cut off)

(‘000 t)

(g/t)

(‘000)

(‘000 t)

(g/t)

(‘000)

Tonnage

Grade

Ounces

2007 - 2009

3,534

2.81

319

3,497

2.72

306

-1%

-3%

-4%

The data in Table 17.1-2 comprises mill statistics, tonnages as mined in the field and block model gold estimates. Milled material versus block model tonnage, gold grade and gold metal are shown. The ratio of milled material to block model is a measure of how well the milled ore tonnage and grade reconciles to predictions made by the block model.

Table 17.1-2: Milled Material to Block Model Reconciliation

	Tonnage	Au	Au Metal
Life of Mine to end 2009	(‘000 t)	(g/t)	(‘000ozs)
Reconciled Milled Production	3,940	3.09	391
Block Model	4,112	2.94	388
Mill / Model Ratio	0.96	1.02	0.98

Results of the reconciliation review show good predictability of tonnage and gold grade by the block model. These results demonstrate the level of confidence in block model estimates, which support the designations of Mineral Reserves and Resources.

18.0 REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION PROPERTIES

The Tasiast Mine began commercial operation in January 2008. Pre-production stripping and mining commenced in February 2007 and first gold was poured on 13 October 2007.

18.1 Mining Operations

The mining method utilised is conventional truck and excavator open pit mining. The operation is selective in terms of separately mining ore and waste. The degree of selectivity upon which the dilution and ore loss allowances are based reflects the scale of mining equipment, the proposed grade control method and the nature of the mineralization. A selective mining unit (SMU) of 3m wide by 2.5m high and 5m long is utilised. This SMU size is commensurate with the ore body model as described in the geological section of this report.

The excavation fleet on site is made up of three Komatsu PC-1250, 110t hydraulic excavators with bucket size of 8m3 loading ten 90t Komatsu 785 trucks The mining fleet on site includes the requisite ancillary equipment (wheel loader, support excavator, dozers, graders and water trucks) for haul and pit access road construction and maintenance, waste dump and ROM pad ore loading to the crusher.

Provision has been made for drilling and blasting all primary materials and 50% of the oxide material. Three drill rigs: two Tamrock Pantera 1100 and one Tamrock Pantera 1500 are available for this task.

Ore is hauled to the ROM pad adjacent to the primary crusher. The majority of ore is tipped to the ROM pad stockpile for reclaim by a front-end-loader (CAT 988) operated as part of the mining operation but costed as a processing cost. Low-grade oxide ore is placed on the dump leach pads and fresh sub-grade ore is stockpiled adjacent to the ROM pad for later treatment.

Waste is used for haul road construction and tailings dam construction as needed or hauled to waste dumps. The road network currently in place is well developed however continued road maintenance as well as additional roads will be required throughout the life of the mine.

The Pit Design Parameters are summarized in Table 18.1

Table 18.1: Summary of Pit Design Parameters

	Bench	Bench	Berm	Inter-Ramp
	Height	Face Angle	Width	Angle
Design Sector	(m)	(°)	(m)	(°)
Oxide (Highly Weathered)	10	60	4.5	45
Oxide (Moderately Weathered)	10	65	5.5	46
Fresh Bedrock North Wall	20	75	8.5	55
Fresh Bedrock East Wall	20	75	8.5	55
Fresh Bedrock South Wall	20	75	8.5	55
Fresh Bedrock West Wall	20	65	10.0	46

Slopes were applied to the pit optimization process through a profile file. This methodology provides the flexibility required by a number of pits and local slope design.

The overall slope angles for optimisation purposes were adjusted to reflect allowances for ramps. Ramp allowances were made considering the following:

· 11m wide one way roads at a gradient of 1 in 8 (7.1 degrees or 12.5%).

· 20m wide two way roads at a gradient of 1 in 10 (5.7 degrees or 10%).

· Minimize waste movement by having ramps predominately on the lowest pit wall, incorporating switchbacks where necessary.

· Combining of ramps and exits for pits in close proximity.

All dollars are in US$. All tonnes quoted are dry tonnes unless otherwise specified. A base gold price of $800/oz was used and a government royalty of 3.0% was applied. A further royalty of 2.0% for gold production above 600 koz is payable to Franco-Nevada Corporation. A combined royalty of 5.0% was assumed for optimization analysis.

All work was carried out using Datamine Studio software and Whittle Four-X optimization software.

18.2 Process Recoveries

18.2.1 CIL

Ore is transported from the open pits to the plant by truck and deposited onto the Run-Of-Mine pad (ROM). To aid blending, ore is transferred to the crushing plant feed bin by front end loader. Crushing of the ore takes place in three stages; a primary jaw crusher that reduces ore to less than 150mm; a secondary cone crusher and two tertiary cone crushers. Screens located before the secondary crusher remove ore that is at the final product size, nominally 10mm. Material that is greater than 10mm passes through the secondary crusher. Secondary crushed ore is conveyed to a screening section before two tertiary stage crushers. Oversized ore is subjected to further crushing and returned to the screens, the tertiary screens and crushers forming a closed circuit. Material passing through the tertiary screens joins the secondary screen undersize as final product which is then transferred to a Fine Ore Bin (FOB).

Grinding

Crushed ore is transferred from the fine ore bin at a controlled rate to the ball mill by means of a conveyor belt. Water is added to the mill feed to permit wet grinding and slurry pumping. Ore passes through the mill, is reduced in size and is pumped to hydrocyclones for classification. Ore that is less than 90 microns in size exits the grinding circuit together with some process solution and passes to the leaching circuit. Ore that has not been reduced to 90 microns is returned to the mill for a further pass through the mill for further grinding.

Carbon-in-Leach

The leaching circuit at Tasiast is a carbon-in-leach circuit. Ore that exits the milling circuit has an approximate pulp density of 42% solids, by weight. The slurry gravitates across a trash screen into the first of six agitated leaching tanks. Dilute sodium cyanide solution and lime are added to start the chemical dissolution of gold from the ore. Compressed air is pumped into each tank to accelerate the dissolution. Activated carbon granules adsorb the dissolved gold from solution. Carbon that has a high gold content is termed “loaded” and is pumped from time to time to the elution circuit for recovery of the gold.

Carbon Elution and Electrolysis

To recover gold from the carbon, batches of carbon are subject to a high pressure and temperature process called elution. A hot caustic solution is used to remove the gold from the carbon and into solution. The gold is recovered from the caustic solution by electrolysis onto mild steel wire wool cathodes in an electro-winning tank. The loaded wool is removed regularly,

mixed with fluxing chemicals and smelted on site to produce bullion bars of 90% purity or higher. The bullion is shipped to a refinery for further refining and sale. After gold recovery the “barren” carbon is heated to reactivate it, and returned to the circuit.

Tailings Waste

Tailings from the CIL process are thickened, to 50% solids, in a thickener to recover process solution for re-use and then pumped to tailings storage facility (TSF). The TSF is a specifically engineered facility, currently comprising two imperviously lined paddock dams located one km south west of the processing plant. After settling a further quantity of process solution is recovered and pumped to the plant for re-use. Solids are retained in the facility.

18.2.2 Dump Leach

The dump leach operation has been designed to process up to 4.5 mtpa of low grade oxide ore utilising five separate pads. The design of each pad allows for three ten metre lifts for a final stack height of 30m. All ponds are plastic-lined with installed leak detection systems.

The initial construction involves earthworks to ensure correct drainage and following compaction a 1.5mm HDPE plastic geomembrane liner is laid. The liner is covered with 0.5m layer of overliner (suitable waste or low grade ore) which acts as a cushion layer to protect the plastic when dumping. Slotted plastic drainage pipes are laid across the pad within the overliner to allow the solution percolating through the pad to be collected and passed to the main drainage pipe at the side of the pad and hence to the required process pond. In addition, leak detection systems are installed below the geomembrane liner.

Trucks deliver low grade run of mine ore to the pads following the addition of lime at the rate of 3.8 kg/t. Each truck places the ore directly onto the pad which is pushed with a bulldozer as required. Once sufficient ore has been placed to an initial height of 10m, the surface is ripped and is then available for irrigation.

Solution required for irrigation is obtained from the tailings dam return water pond. Fresh make-up water is added to the tailings dam allowing for precipitation of scale within the dam. Solution from the return water pond is pumped to the Barren Pond where cyanide is added as required to maintain an initial concentration of 200ppm. The lime added during stacking maintains a pH of ten. Irrigation is targeted at ten litres per hour per square meter and is applied using a system of plastic piping and mini-wobblers. Anti-scalant is added to prevent scale formation. Pregnant solution (~ 0.5 g/t) is then pumped to the carbon recovery plant at the CIL plant.

Water make-up requirement is approximately 4,000 m3 per day at maximum production to replace losses due evaporation and moisture retention within the pads.

18.2.3 Recoveries

The CIL recoveries for the Tasiast ore deposits based on the selected treatment route, metallurgical testwork and concentrator operation, are presented in Table 18.2.3 below.

Table 18.2.3: Metallurgical Recoveries

	Oxide	Primary
CIL Recovery	94%	95%
DL Recovery*	60% - 75%

· Dependent on lithology

18.3 Taxation

In the period 2008 - 2010, Tasiast’s profits have been exonerated from income taxes under a Mining convention signed in 2006 with the government of Mauritania. Tasiast’s future profits, once the exoneration from income taxes ceases, will be subject to tax based on a 25% income tax rate. Amortization and depreciation of Tasiast’s past and future capital projects can be applied using the established tax rates of amortization to reduce the income otherwise subject to tax.

18.4 Capital & Operating Cost Estimates

18.4.1 Capital Expenditure

The estimated life of mine capital expenditure is forecast to be $146.5 million. Table 18.7.1 below details the forecast capital expenditure for the life of mine:

Table 18.7.1: Life of Mine Capital Expenditure

		LOM	2010	2011	2012	2013	2014	2015	2016	2017	2018	2019
Capital
Plant Expansion	$ million	3.8	3.8
Dump Leach	$ million	13.9	13.9
Other Projects	$ million	58.5	8.5	3.0	3.0	3.0	3.0	3.0	3.0	3.0	3.0	3.0
Fleet	$ million	56.8	16.8		2.0	2.0	1.0			20.0	15.0
TSF	$ million	13.5	2.0	0.6		1.2		1.5		1.6		1.0
Total Capital	$ million	146.5	45.0	3.6	5.0	6.2	4.0	4.5	3.0	24.6	18.0	4.0

		2020	2021	2022	2023	2024	2025	2026	2027	2028	2029	2030
Capital
Plant Expansion	$ million
Dump Leach	$ million
Other Projects	$ million	3.0	3.0	3.0	3.0	3.0	3.0	3.0	1.0	1.0
Fleet	$ million
TSF	$ million		1.1		1.3		1.5		1.7
Total Capital	$ million	3.0	4.1	3.0	4.3	3.0	4.5	3.0	2.7	1.0

18.4.2 Operating Costs

The life of mine unit operating costs are forecast as detailed in table 18.7.2 below:

Table 18.7.2: Life of Mine Operating Cost

		LOM
OP Mining Cost	$/t mined	1.91
CIL Process Cost	$/t milled	10.83
HL Process Cost	$/t milled	—
DL Process Cost	$/t milled	1.13
Power Cost	$/t milled	2.72
G&A Costs	$/t milled	4.06
Total Cost *	$/t milled	32.05

*(Excluding Royalties)

18.5 Economic Analysis

18.5.1 Cash Flow Forecast

A financial model based on the Life of Mine Plan is presented in Table 18.8.1. The model is based on the December 2009 reserve model.

This model has been prepared on the basis of the following assumptions:

· Pre-tax state royalty of 3.0%, Franco Nevada royalty 2.0% on production greater than 600k ozs

· No escalation has been applied to operating costs.

· The mine life is 20 years.

The model is based on the mining and treatment of proven and probable reserves (5.03 million ounces) at 3.0 mtpa and total life of mine capital expenditures of $146.5 million.. The principal results of the evaluation are:

· Net Present value at 5% $1,247 million

· Net cash Flow $1,577 million

· Total Royalty Payments $196 million

Table 18.8.1: Life of Mine Financial Model

Aug-10	Units	LOM	2010	2011	2012	2013	2014	2015	2016	2017	2018	2019
CIL
Milled Tonnes	million t	61.2	2.5	3.0	3.0	3.0	3.0	3.0	3.0	3.0	3.0	3.0
Grade	g/t	2.05	1.94	2.43	2.94	2.92	2.60	2.16	1.85	2.16	1.96	2.28
DL
Placed Tonnes	million t	54.5	6.0	7.6	6.6	8.6	9.3	5.4	1.8	4.2	1.0	3.9
Grade	g/t	0.58	0.77	0.66	0.75	0.59	0.52	0.48	0.56	0.31	0.16	0.54
Recovered Ounces
CIL	‘000 oz	3,628	143	214	259	256	228	190	162	189	172	200
HL	‘000 oz	—	—	—	—	—	—	—	—	—	—	—
DL	‘000 oz	826	165	221	113	106	92	69	31	10	10	5
Total	‘000 oz	4,454	307	435	372	362	320	259	193	199	182	206
Gold Price	$/oz	871	1,000	1,000	1,000	900	900	900	800	800	800	800
Revenue	$ million	3,880	307	435	372	326	288	233	155	159	146	165
Total Operating
Costs	$ million	1,961	102	130	122	128	128	122	120	120	119	76
Operating Cash flow	$ million	1,920	206	305	250	198	160	112	35	39	27	88
Royalty	$ million	196	17	22	19	16	14	12	8	8	7	8
Total Capital	$ million	147	45	4	5	6	4	4	3	25	18	4
Cash Flow	$ million	1,577	143	279	227	175	141	95	24	6	2	76
NPV5%	$ million	1,247
Cash Cost		440	331	299	327	353	401	470	620	604	652	371

Aug-10	Units	2020	2021	2022	2023	2024	2025	2026	2027		2028	2029	2030
CIL
Milled Tonnes	million t	3.0	3.0	3.0	3.0	3.0	3.0	3.0	2.8		3.0	3.0	1.8
Grade	g/t	2.25	2.13	2.27	2.16	1.67	2.33	2.22	1.49		0.92	0.92	1.11
DL
Placed Tonnes	million t	0.0	—	0.0	—	0.0	—	0.0	0.2		—	—	—
Grade	g/t	0.54	—	0.54	—	0.44	—	0.37	0.46		—	—	—
Recovered Ounces
CIL	‘000 oz	197	187	199	190	146	204	195	101		81	81	33
HL	‘000 oz	—	—	—	—	—	—	—	—		—	—	—
DL	‘000 oz	1	1	0	0	0	0	0	1		1	1	—
Total	‘000 oz	199	188	199	190	147	204	195	102		82	81	33
Gold Price	$/oz	800	800	800	800	800	800	800	800		800	800	800
Revenue	$ million	159	150	159	152	117	164	156	81		65	65	26
Total Operating
Costs	$ million	86	106	74	70	74	69	81	94		54	54	31
Operating Cash flow	$ million	72	44	86	82	43	94	75	(13	)	11	11	(4	)
Royalty	$ million	8	8	8	8	6	8	8	4		3	3	1
Total Capital	$ million	3	4	3	4	3	4	3	3		1	—	—
Cash Flow	$ million	61	32	75	70	34	82	64	(20	)	7	7	(6	)
NPV5%	$ million
Cash Cost		435	565	370	369	507	339	417	926		663	668	936

18.5.2 Sensitivity Analysis

Sensitivity Studies have been undertaken on the financial model for gold price variations and changes in operating costs. The effect of these scenarios on net present value (NPV5%) are shown in table 18.8.2.

	%	NPV5%
	Change	($ million)
	120%	1,789
Revenue	Base	1,247
	80%	705

	120%	977
Operating Cost	Base	1,247
	80%	1,517

Table 18.8.2: Sensitivity Analysis

The sensitivity analysis demonstrates that the mine is sensitive to changes in revenue which is equally influenced by gold price, grade and recovery.

18.6 Payback

Following the completion of the mill expansion, the mine will be able to fund all capital and restoration requirements through cash flow.

18.7 Mine Life

The current life of the mine is 20 years based on current reserves. However, the exploration potential of the mine is considered extremely positive.

Further Ore Reserves will potentially be defined from the upgrade and inclusion of the West Branch prospect and in addition numerous prospects exist both within the 312km2 El Gaicha Mining lease and in the surrounding Prospecting licences that are within haulage distance of the Process Plant. Many of these prospects have been the subject of a minimal amount of drilling and have yielded ore grade intercepts that justify follow up.

19.0 INTERPRETATION AND CONCLUSIONS

Geological, mining and metallurgical data from Tasiast Mine is sufficient to obtain a good level of understanding to assess the Tasiast end December 2009 Mineral Reserve and Resource statement. The following is a list of general conclusions.

1. The geology is well understood for the Piment and West Branch deposits. The gold mineralization types and extents are well defined and that knowledge has been integrated into the geologic block models, mining practice and metallurgy.

2. The quality of the assay data used for block model grade estimates is supported by good reconciliation of material milled to block model grades and tonnages.

3. The block models have been developed using industry-accepted methods.

4. The cutoff grade strategy employed is based on industry-accepted parameters.

5. Metallurgical expectations are reasonable and reflect the metallurgical results during the first year of production from the oxide ore.

6. Operating cost estimates are reasonable and have been calculated using sound industry-accepted practices.

7. The assumptions used for the economic forecast are within market parameters and are valid assumptions for an economic forecast.

8. There is considerable potential to further increase the Tasiast Mineral Resources and Ore Reserves.

20.0 RECOMMENDATIONS

Exploration at Tasiast has been successful in converting Inferred Mineral Resources to Measured and Indicated Resources and hence to Ore Reserves. There is excellent potential for this to continue and work should continue in this regard to fully define the Tasiast Mineral Reserves.

21.0 REFERENCES

Anonymous, 2002. CIA The World Factbook 2002 - Mauritania. Internet (http://www.odci.gov.cia.publications/factbook/ct.html), 9p.

Demers, P., Gauthier, D., Kroon, A.S., Lafleur, P-J., 2004 Tasiast Gold Project, Islamic Republic of Mauritania, Feasibility Study Report. SNC Lavalin report for Defiance Mining Limited.

King, P.A., 1999. Laboratory Test work on Samples from the Tasiast Project for LaSource SAS. CSMA Minerals Limited Report, Ref: 64-0075, 36 pages.

Leroux, D.C., Roy, W.D., Orava, D., 2007. Technical Report on the Tasiast Gold Project, Islamic Republic of Mauritania for Red Back Mining Inc. A.C.A. Howe International Limited Report #910, 97 pages.

Scott Wilson (February 2008a) Tasiast Gold Project, Environmental Impact Study, Addendum II of IV, Environmental Impact Review of Tailings Storage Facility, Report prepared by Scott Wilson Limited for Tasiast Mauritania Limited S.A. for submission to Mauritanian Government.

Scott Wilson (February 2008b) Tasiast Gold Project, Environmental Impact Study, Addendum III of IV, Environmental Management Plan, Report prepared by Scott Wilson Limited for Tasiast Mauritania Limited S.A. for submission to Mauritanian Government.

Scott Wilson (February 2008c) Tasiast Gold Project, Environmental Impact Study, Addendum IV of IV Preliminary Rehabilitation and Closure Plan, Report prepared by Scott Wilson Limited for Tasiast Mauritania Limited S.A. for submission to Mauritanian Government.

22.0 DATE AND SIGNATURE

/s/ Hugh Stuart

Hugh Stuart, B.Sc., M.Sc, MAusIMM		Date: 10 August 2010
VP Exploration
Qualified Person

23.0 CERTIFICATES OF AUTHORS

CERTIFICATE OF QUALIFIED PERSON

RE: Technical Report in the Tasiast Gold Mine, Islamic Republic of Mauritania dated 10 August, 2010

I, Hugh David Stuart do hereby certify that:

1. I am a consulting Geologist and Britsih citizen residing in Coventry, United Kingdom. I am contracted full time to Red Back Mining Inc and fill the role of Vice President Exploration.

2. I hold a Bachelor of Science (Honours) degree in Geology, graduating from Manchester University University in 1985 and an Master of Science degree in Mineral Exploration and Mining Geology from the University of Leicester in 1988.

3. I am a member, in good standing, of the Australasian Institute of Mining and Metallurgy.

4. I have been practicing my profession relating to mining and mineral exploration for 22 years.

5. I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Properties (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I am a “qualified person” for the purposes of NI 43-101.

6. I am responsible for the overall preparation of the Technical report entitled “Technical Report in the Tasiast Gold Mine, Islamic Republic of Mauritania” and dated 10 August, 2010.

7. I am Vice President Exploration for Red Back Mining Inc and have been involved with the Tasiast Gold Mine since its acquisition by Red Back Mining Inc in 2007. I have visited the mine on numerous occasions.

8. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

9. As at the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated 10 August, 2010

/s/ Hugh David Stuart
Hugh David Stuart

100

CERTIFICATE OF QUALIFIED PERSON

RE: Technical Report in the Tasiast Gold Mine, Islamic Republic of Mauritania dated 10 August, 2010

I, Nicolas James Johnson do hereby certify that:

1. I am an independent consulting Geologist and Australian citizen residing in Perth, Western Australia. I am a full time employee of Hellman and Schofield Pty. Ltd. of Suite 6, 3 Trelawney Street, Eastwood, NSW, Australia.

2. I hold a Bachelor of Science (Honours) degree, graduating from La Trobe University in 1988.

3. I am a member, in good standing, of the Australian Institute of Geoscientists.

4. I have been practicing my profession relating to mining and mineral exploration for 20 years.

6. I am one of the authors of the report entitled “Technical Report in the Tasiast Gold Mine, Islamic Republic of Mauritania” and dated 10 August, 2010. I am responsible for section 16 (excluding the reserves listed in section 16.9) of the Technical Report.

7. I visited the mine from 17th to 22nd February 2008 for a period of 5 days.

8. I have been working on the resource estimation of the Tasiast Gold Mine since 2008.

9. I am independent of Red Back Mining Inc. applying the test set out in section 1.4 of NI 43-101.(i)

10. I have read NI 43-101 and section 16 of the Technical Report has been prepared in compliance with NI 43-101.

11. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

12. As at the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated 10 August, 2010

/s/ Nicolas James Johnson
Nicolas James Johnson

101

(i) Section 1.4 of NI 43-101 states a qualified person is independent of an issuer if there is no circumstance that could, in the opinion of a reasonable person aware of all relevant facts, interfere with the qualified person’s judgment regarding the preparation of the technical report.

102