6-K

SECURITIES AND EXCHANGE COMMISSION

Washington, DC  20549

FORM 6-K

REPORT OF FOREIGN PRIVATE ISSUER

PURSUANT TO RULE 13a-16 OR 15d-16

UNDER THE SECURITIES EXCHANGE ACT OF 1934

For the month of March, 2008

Commission File Number:  001-13382

KINROSS GOLD CORPORATION

(Translation of registrant’s name into English)

52nd Floor, Scotia Plaza, 40 King Street West

Toronto, Ontario  M5H 3Y2

(Address of principal executive offices)

Indicate by check mark whether the registrant files or will file annual reports under cover of Form 20-F or Form 40-F:

Form 20-F  o Form 40-F  x

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(1):o

Note:  Regulation S-T Rule 101(b)(1) only permits the submission in paper of a Form 6-K if submitted solely to provide an attached annual report to security holders.

Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(7):o

Note:  Regulation S-T Rule 101(b)(7) only permits the submission in paper of a Form 6-K if submitted to furnish a report or other document that the registrant foreign private issuer must furnish and make public under the laws of the jurisdiction in which the registrant is incorporated, domiciled or legally organized (the registrant’s “home country”), or under the rules of the home country exchange on which the registrant’s securities are traded, as long as the report or other document is not a press release, is not required to be and has not been distributed to the registrant’s security holders, and, if discussing a material event, has already been the subject of a Form 6-K submission or other Commission filing on EDGAR.

Indicate by check mark whether by furnishing the information contained in this Form, the registrant is also thereby furnishing the information to the Commission pursuant to Rule 12g3-2(b) under the Securities Exchange Act of 1934.

Yes o        No x

If “Yes” is marked, indicate below the file number assigned to the registrant in connection with Rule 12g3-2b: o

This Current Report on Form 6-K, dated March 31, 2008, is specifically incorporated by reference into Kinross Gold Corporation’s Registration Statement on Form S-8 (Registration No. 333-141896), filed on April 4, 2007.

This report on Form 6-K is being filed for the sole purpose of providing the technical reports with respect to the Maricunga Gold Mine effective December 31, 2007 and with respect to the Fort Knox Mine dated March 31, 2008.

INDEX

Table of Contents

SIGNATURES

EXHIBIT INDEX

99.1Maricunga Gold Mine Technical Report effective December 31, 2007

99.2Fort Knox Mine Technical Report dated March 31, 2008

99.3         Consent of Expert - Robert Henderson, P. Eng.

99.4         Consent of Expert - Maryse Belanger, P. Geo

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SIGNATURES

Pursuant to the requirements of Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.

	KINROSS GOLD CORPORATION
	Signed:	/Shelley M. Riley
	V.P. Administration and Corporate Secretary
March 31, 2008.

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EX-99.1

Exhibit 99.1

Technical Report

for the

Maricunga Gold Mine

(Located in the Maricunga District of Region III, Chile)

Prepared for Compañia Minera Maricunga

and

Kinross Gold Corporation

Prepared by:

M. Bélanger, P. Geo.

Director of Technical Services

Kinross Gold Corporation

Effective Date: December 31, 2007

	Kinross Gold Corporation
	Maricunga Mine Technical Report

ACKNOWLEDGEMENTS

The author would like to thank the following individuals for their contributions to and their assistance in the preparation of this report.

Juan Carlos Buguena	Planning Superintendent, CMM
Rene Albornoz	Geology Superintendent, CMM
Jorge Rojas	Reserves Technician, CMM
Barry Gillies	Director of Exploration, Kinross
Mark Sedore	Director, Corporate Reserves, Kinross
Kevin Morris	Director Mine Engineering Services, Kinross
Jeremy Niemi	Senior Geologist, Kinross
Rob Henderson	Vice President Technical Services, Kinross
Kevan Ford	Manager, Metallurgy, Kinross
Sergio Neira Martinez	Exploration Geologist, CMM
Ivo Ahumada Cortes	Exploration Geologist, CMM

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Certificate of Author

I, Maryse Bélanger, P. Geo., do hereby certify that:

·I am Director of Technical Services, Kinross Americas for Kinross Gold Corporation, Av. Afonso Pena, 4001-0 andar Ala - Ed. Presidente Tancredo Neves - Bairro Serra - Belo Horizonte, MG, Brazil, CEP-30130.

·I graduated with a Bachelor of Science degree (BSc) in Earth Sciences from the Université du Québec à Chicoutimi in 1985. I studied Geostatistics at the Centre de Géostatistique in Fontainebleau, France in 1986.

·I am a member of the Association of Professional Geoscientists of Ontario (Registration Number # 0125).

·I have worked as a geologist for a total of 23 years since my graduation from University.  I have been involved in gold exploration and mining in Canada, United States of America, Russia, Niger, Burkina Faso, Ivory Coast, Ethiopia, Gabon and Chile.

·I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purposes of NI 43 -101.

·I have read NI 43-101 and certify that the Technical Report has been prepared in compliance with NI-43-101 and Form 43-101F1.

·I am responsible for supervising the writing of the technical report titled “Technical Report on the Maricunga Gold Mine” dated December 31, 2007.

·I visited the property every quarter in 2004 and 2005 and spent more than 75 days at site during the course of active exploration in 2006. In 2007 I visited the mine on a regular basis.

·I have not had prior involvement with the property that is the subject of the technical report.

·I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

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

·As of the date of this certificate, to the best of my knowledge and belief, the Technical Report contains all the scientific and technical information that is required to omission to make the technical report not misleading.

·I am not independent of the issuer.  Per section 5.3.2 of National Instrument 43-101 an independent qualified person was not required to write the technical report on the Maricunga Mine.

Dated this 31^st day of March, 2008 at Brasilia, Brazil.

“Signed and Sealed”

Maryse Bélanger, P. Geo.

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TABLE OF CONTENTS

1.0	SUMMARY		1-11
2.0	INTRODUCTION AND TERMS OF REFERENCE		2-17
2.1	INTRODUCTION		2-17
2.2	TERMS OF REFERENCE		2-17
2.3	SCOPE OF WORK		2-17
2.4	SOURCES OF INFORMATION AND REPORT BASIS		2-18
2.5	FIELD INVOLVEMENT OF THE QUALIFIED PERSON		2-18
3.0	RELIANCE ON OTHER EXPERTS		3-19
4.0	PROPERTY DESCRIPTION AND LOCATION		4-20
4.1	CLAIM STATUS		4-20
4.1.1		Surface Rights	4-21
4.1.2		Water Rights	4-21
4.1.3		Royalty Payments	4-21
4.1.4		Environmental Liabilities	4-21
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		5-27
5.1	ACCESSIBILITY		5-27
5.2	CLIMATE AND PHYSIOGRAPHY		5-27
5.3	LOCAL RESOURCES AND INFRASTRUCTURE		5-27
6.0	HISTORY		6-29
7.0	GEOLOGICAL SETTING		7-32
7.1	ALTERATION		7-32
7.1.1		Verde	7-32
7.1.2		Pancho	7-33
7.2	STRUCTURE		7-34
8.0	DEPOSIT TYPES		8-35
8.1	VERDE		8-35
8.2	PANCHO		8-36

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9.0	MINERALIZATION		9-38
9.1	VERDE		9-38
9.2	PANCHO		9-38
10.0	EXPLORATION		10-40
11.0	DRILLING		11-42
11.1	RIG SETUP AND SURVEY		11-43
11.2	TOPOGRAPHIC BASE AND SURVEY AUDIT		11-43
11.3	CORE HANDLING, STORAGE AND SECURITY		11-43
12.0	SAMPLE METHOD AND APPROACH		12-45
12.1	RC DRILLING		12-45
12.2	CORE DRILLING		12-46
12.3	LOGGING		12-46
12.4	GEOTECHNICAL CORE LOGGING		12-47
12.5	COMPOSITE SAMPLE SUMMARY		12-47
13.0	SAMPLE PREPARATION, ANALYSES AND SECURITY		13-48
13.1	SAMPLE PREPARATION		13-48
13.2	SAMPLE PREPARATION 2006		13-49
13.3	ANALYSIS		13-49
13.4	QUALITY CONTROL / QUALITY ASSURANCE		13-50
13.4.1		Drilling 2002-2003	13-50
13.4.2		Drilling 2006	13-51
13.4.3		Table of Failures	13-51
13.4.4		Analysis of QA/QC Data	13-52
13.4.5		Referee Laboratory	13-52
13.4.6		Statement of Author’s Opinion	13-52
14.0	DATA VERIFICATION		14-53
14.1	GENERAL 2002-2003		14-53
14.2	PRE-2002 DATA VERIFICATION		14-54
14.2.1		Collar Data	14-54
14.2.2		Survey Data	14-54
14.2.3		Lithology, Alteration, Oxidation, Vein and Mineralization Data:	14-54

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14.2.4		Analytical Data	14-54
14.3	2002-2003 DATA VERIFICATION		14-54
14.3.1		Collar Data	14-54
14.3.2		Survey Data	14-55
14.3.3		Lithology, Alteration, Oxidation, Vein and Mineralization Data	14-55
14.3.4		Analytical Data	14-55
14.3.5		Geotechnical Data	14-55
14.4	DATA ENTRY, STORAGE, MANAGEMENT AND SECURITY		14-56
14.4.1		Data Entry	14-56
14.4.2		Data Storage	14-57
14.4.3		Data Management and Security	14-57
15.0	ADJACENT PROPERTIES		15-58
16.0	MINERAL PROCESSING AND METALLURGICAL TESTING		16-59
16.1	METALLURGICAL TEST WORK 2003		16-60
16.1.1		Cyanide Soluble Assays	16-60
16.1.2		Direct Agitated Cyanidation (Bottle Roll) Tests	16-60
16.1.3		Column Leach Tests	16-62
16.1.4		Bond Work Index	16-64
16.1.5		Specific Gravity	16-65
16.1.6		Acid Rock Drainage Study	16-65
16.2	METALLURGICAL TEST WORK 2006		16-66
16.2.1		Introduction	16-66
16.2.2		Comments on Historical Metallurgical Composite Selection Criteria	16-66
16.2.3		Metallurgical Composite Selection Criteria (2006)	16-67
16.2.4		McClelland Test Work Summary	16-68
16.2.5		SART Process Feasibility Study	16-73
17.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE		17-78
17.1	INTRODUCTION		17-78
17.2	INTERPRETATION AND GRADE DOMAINS		17-78
17.2.1		Verde	17-78
17.2.2		Pancho	17-80
17.3	OXIDATION MODEL AND RECOVERY		17-81
17.4	GRADE MODEL		17-81

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17.4.1		Statistical and Geostatistical Analyses	17-81
17.4.2		Grade capping	17-83
17.4.3		Sample Compositing	17-83
17.4.4		Contact Profiles	17-84
17.4.5		Variography	17-84
17.4.6		Model Framework	17-85
17.4.7		Gold Grade Interpolation Verde	17-85
17.4.8		Gold Grade Interpolation Pancho	17-86
17.4.9		Densities	17-87
17.4.10		Model Checking and Verification	17-88
17.5	RESOURCE CLASSIFICATION		17-88
17.5.1		Resource Classification Verde	17-89
17.5.2		Resource Classification Pancho	17-89
17.6	MINERAL RESOURCE ESTIMATE		17-89
17.7	CONSIDERATIONS FOR RESERVES		17-90
17.8	MINERAL RESERVE ESTIMATE		17-90
17.9	PIT OPTIMIZATIONS		17-91
17.9.1		Verde	17-91
17.9.2		Pancho	17-91
17.10	PIT DESIGNS		17-93
17.10.1		Verde	17-93
17.10.2		Pancho	17-95
18.0	OTHER RELEVANT DATA AND INFORMATION		18-97
19.0	INTERPRETATION AND CONCLUSIONS		19-98
20.0	RECOMMENDATIONS		20-99
21.0	REFERENCES		21-100
22.0	DATE AND SIGNATURE PAGE		22-104
23.0	ADDITIONAL REQUIREMENTS FOR OPERATING PROPERTIES		23-105
23.1	MINING OPERATIONS		23-105
23.1.1		Mobile Equipment	23-106
23.1.2		Leach Pad and Plant Infrastructure	23-106

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23.2	RECOVERABILITY	23-107
23.3	MARKETS	23-107
23.4	CONTRACTS	23-107
23.5	ENVIRONMENTAL CONSIDERATIONS	23-107
23.6	CAPITAL COSTS	23-108
23.7	OPERATING COSTS	23-109
23.8	ECONOMIC ANALYSES	23-110
23.9	PAYBACK	23-110
23.10	MINE LIFE	23-110

LIST OF TABLES

Table 1-1: Maricunga Proven and Probable Mineral Reserves	1-15
Table 1-2: Maricunga Measured and Indicated Mineral Resources	1-15
Table 4-1: Mineral Rights and Claims	4-24
Table 6-1: Annual Gold Production of the Maricunga Mine	6-30
Table 11-1: Exploration Drill Summary	11-42
Table 16-1: Summary of Bottle Roll Sample Selection	16-61
Table 16-2: Summary of Bottle Roll Results	16-61
Table 16-3: Summary of Column Leach Sample Selection	16-62
Table 16-4: Summary of Column Leach Test Results	16-63
Table 16-5: Specific Gravity Results	16-65
Table 16-6: Pancho Metallurgical Composites 2006	16-68
Table 16-7: Pancho and Verde Gold Recovery Estimates	16-71
Table 16-8: Pancho Bottle Roll Summary 2006	16-72
Table 16-9: Pancho Column Test Summary 2007 (Part 1)	16-72
Table 16-10: Pancho Column Test Summary 2007 (Part 2)	16-73
Table 17-1: Oxidation Codes	17-81
Table 17-2: Statistics for Uncapped 2 m Gold Assays - Verde	17-82

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Table 17-3: Statistics for Uncapped 2 m Gold Assays - Pancho	17-82
Table 17-4: Statistics for 5 m Gold Composites - Verde	17-83
Table 17-5: Statistics for 5 m Gold Composites – Pancho	17-84
Table 17-6: Statistics for 5 m Copper Composites – Pancho	17-84
Table 17-7: Statistics for 5 m Cyanide Soluble Copper Composites – Pancho	17-84
Table 17-8: Verde Variogram Table	17-85
Table 17-9: Pancho Variogram Table	17-85
Table 17-10: Density Measurements and Oxidation Level for Verde	17-88
Table 17-11: Density Measurements and Oxidation Level for Pancho	17-88
Table 17-12: Measured and Indicated Mineral Resources	17-90
Table 17-13: Proven and Probable Mineral Reserves	17-91
Table 17-14: Optimization Parameters for Verde	17-92
Table 17-15: Optimization Parameters for Pancho	17-93
Table 23-1: Capital Costs – Maricunga	23-108
Table 23-2: 2007 Actual Operating Costs – Maricunga	23-109
Table 23-3: 2007 Life of Mine Plan	23-110

LIST OF FIGURES

Figure 4-1: Maricunga Location Map	4-20
Figure 4-2: Land Claims and Infrastructure Location	4-22
Figure 4-3: Details of Mineral Rights and Claims	4-23
Figure 7-1: Geology of the Maricunga Mine Area	7-34
Figure 8-1: Simplified Geology for Pancho	8-37
Figure 11-1: Verde Drilling Coverage as December 31, 2007	11-44
Figure 11-2: Pancho Drill Coverage as at December 31, 2007	11-44
Figure 16-1: Metallurgical Sample Locations – Verde Pit – 4300 Level	16-63
Figure 16-2: Metallurgical Sample Locations – Pancho Pit – 4300 Level	16-64
Figure 16-3: SART Plant Flowsheet	16-74

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Figure 16-4: Modified Maricunga Plant Flowsheet with SART	16-77
Figure 17-1: Typical Plan View of Verde Domaining (Bench 4280)	17-79
Figure 17-2: Interramp angles for Verde Pit Design	17-94
Figure 17-3: Verde Pit Design Phase 4 with Topography	17-95
Figure 17-4: Pancho Pit Design with Topography	17-96

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1.0          SUMMARY

This Technical Report for the Maricunga Mine has been prepared to support the 2007 Kinross Gold Corporation (Kinross) Annual Mineral Resources and Reserves Statement and has been prepared under the direct supervision of the qualified person, Maryse Bélanger, P. Geo., Director Technical Services, Kinross Americas with contributions from CMM technical staff.

The Maricunga Gold Mine is located in the Maricunga Gold Belt in Region III of northern Chile. The property is located 120 km due east of the city of Copiapó at elevations between 4,200 and 4,500 meters above mean sea level.  The mine operated from 1996 to 2001, producing more than 920,000 ounces of gold from 46.0 million tonnes of ore. The mine was placed on care and maintenance in 2001, due to a downturn in gold prices. The Maricunga Mine complex comprises two open-pits, Verde East and Verde West and a near-development project called Pancho.

The Maricunga heap leach mine is owned and operated by Compañía Minera Maricunga (CMM), a Chilean company that is wholly-owned by Kinross since its acquisition of Bema Gold Corporation effective at the end of February 2007.

The Verde and Pancho gold deposits at Maricunga occur in the Maricunga Gold Belt of the high Andes in northern Chile.  Since 1980, a total of 40 million ounces of gold have been defined in the belt, (Muntean and Einaudi, 2000).  Gold mineralization at Maricunga has been interpreted as porphyry style gold systems and is hosted in the Maricunga volcanic-intrusive complex of Early Miocene age. The porphyries occur within a sequence of intermediate tuffs, porphyries and breccias that are the host rocks to the gold mineralization.

Gold mineralization at Verde is interpreted to be the result of the fracturing and concentration of fluids in the carapace of an intrusive plug or stock. Gold is closely associated with quartz, magnetite, calcite, and garnet stockworks. Gold mineralization at Pancho is characterized as porphyry hosted stockwork and sheeted veins. The veins are subvertical and have a strong, preferred north-westerly strike. The northwest structural control is evident not only at outcrop scale but is also reflected in the northwest alignment of intrusives and the three centers of mineralization in the district, Verde, Pancho and Guanaco.

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In September 2002, in response to rising gold prices, Compañia Minera Maricunga (CMM) approved an Exploration Program designed to increase the reserve base of the Maricunga Project to a level sufficient to support resumption of active mining. The drilling completed in 2002-2003 focused largely on definition drilling for a $350 gold pit design.

The exploration program ran from September 2002 to June 2003.  During this period, a total of 262 drill holes (51,478 meters) of drilling were completed. The drilling focused on increasing the confidence level of the known mineralization below the current Verde pits as well as increasing the confidence level in the mineralization at the nearby Pancho deposit, located approximately 2.0 km to the northeast.  Much of the 2002 – 2003 drilling was diamond drill core, allowing geologists an opportunity to clearly delineate geological and alteration features affecting gold mineralization and recovery.

The information was compiled and incorporated in new mineral resource and mineral reserve estimates for Verde and Pancho. The reserves formed the basis of a detailed engineering study examining the economics of the project assuming a capital investment of $101.1 million to upgrade the existing infrastructure, allowing the plant to process 40,000 tonnes per day of Verde ore and 35,000 tonnes per day of Pancho ore. The capital investment was based on preliminary design and first-principle engineering.

The Maricunga Mine resumed full production in October 2005 at a mining rate of 40,000 tpd. Total production of 233,000 ounces from 14,333,000 tonnes at 0.74 g/t Au is reported for 2007. Only data below the year-end 2007 topographic surface was used to complete the new resource model and reserve estimate.

Both the 2002-2003 and 2006 drill programs were carefully supervised, employing industry best practices and rigid quality management procedures in the collection and management of the field data. In addition, all historic data for the project was subjected to detailed verification programs where the data in the relational database was verified against original logs, survey calculation sheets and original assay certificates.  The result is a duly verified database containing data of high quality and free of gross errors and omissions. This database was then used to estimate mineral resource and reserves for the project.

In early 2006, CMM made the decision to drill at Pancho to better define the mineral resource both laterally and at depth.  During the course of 2007, a full review of the geologic

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interpretation and the spatial controls of mineralization was undertaken for both Verde and Pancho resulting in new resource models to support updated Mineral Resource and Mineral Reserve estimates.

In addition to the analytical data, the 2002-2003 drill program was also used to collect carefully controlled metallurgical samples from Verde and Pancho. A total of 58 direct agitation (bottle roll) and 18 column leach test samples were collected and analyzed. The metallurgical samples were carefully selected to ensure adequate sample coverage of various grade bins, recovery classes and lithologies. More metallurgical test work was completed in 2006 and 2007 with the addition of 19 column tests for Pancho. The data compiled to date indicates that gold recovery from the Pancho samples tested is approximately 68% for the sulphide ore and 85% for the oxide ore.  This recent test work plus the work completed in 2003 confirms that the gold recovery values used in the 2007 reserve estimate are reasonable.

In 2007, the SART process was tested for the Pancho ore given the higher concentrations of copper present in the ore. Some of the copper minerals are soluble in cyanide solutions and these minerals consume cyanide reagent and contaminate the leach solution with copper metal.  A feasibility study and pilot testing program were completed by Idesol and SGS Lakefield of Santiago. The study considers plant design, performance metrics, operating and capital costs and implementation strategy. The reserves defined at Pancho consider the application of SART.

Ore and waste hardness was also tested in 2002-2003. Specific gravity of the various rock types was confirmed through extensive field sampling. In 2006, specific gravity measurements were also performed on more than 1,500 core samples from Pancho.

The resource block models were first updated in 2003 to reflect the lithological models developed from observations taken from drill core. At the same time, a new recovery model was developed based on the visual logging of oxidation. The visual estimates were completed using a well-established scale previously used on the project. In 2006, with the additional core samples, the surfaces defining oxidation levels were updated (oxide, mixed and sulphide).

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Statistical and geostatistical analysis of the drill results was performed on the raw, 2.0 meter sample intervals to identify appropriate grade capping factors. Grade capping does not have a significant effect on the resource estimates, resulting in a less than 1% decrease in contained gold. The 2007 work also included compositing of the data using 5 m length. Contact profiles were prepared to determine the search strategies for copper and gold grade interpolation.  Directional variograms were modeled to identify anisotropy and search ranges that would guide grade interpolation. The variograms demonstrated excellent continuity, generally exceeding 100 meters in all directions.

Grade interpolation was limited to gold for the Verde deposit while gold; copper and cyanide soluble copper were modeled for Pancho. Ordinary kriging was selected as the primary interpolation method for gold. Both inverse distance and nearest neighbour grade interpolations were also performed to verify the kriged gold grades.

The resultant resource block models were manually verified to ensure that grade interpolation was reasonable. The model was classified as per CIM Guidelines as Measured, Indicated and Inferred based primarily on the search ranges indicated by the variograms and experience gained at Verde.

The block models were used as the basis for pit optimization that was completed with the Whittle algorithm. Optimization work was based on operating costs, adjustment factors and parameters updated in November 2007.  The optimum pit shells were imported into Vulcan for final pit design using design parameters approved by a third party geotechnical engineering company.

The final mine designs were used to estimate the mineral resources and reserves was and were prepared by a Qualified Person as required by National Instrument 43-101.

Mineral reserves listed below in Table 1-1 are based on a $550 gold price and constitute the official reserves as of 31, December 2007.

It is Kinross Policy that for all open pit mine projects, the mineral resource must fall within optimized or final design pits that are based on reasonable, long-term, gold price assumptions.

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In Table 1-2, mineral resources are based on a $625 gold price and reported exclusive of mineral reserves as of 31, December 2007.

Table 1-1: Maricunga Proven and Probable Mineral Reserves

		Ore Tonnes	Grade	Gold Ounces
Deposit	Class	(x 1,000)	(Au g/t)	(x 1,000)
Verde	Proven	71,071	0.85	1,947
	Probable	32,409	0.74	772
	2P	103,480	0.82	2,719
Pancho	Proven	106,627	0.71	2,436
	Probable	69,395	0.58	1,290
	2P	176,022	0.66	3,726
Total		279,502	0.72	6,445

Table 1-2: Maricunga Measured and Indicated Mineral Resources

		Ore Tonnes	Grade	Gold Ounces
Deposit	Class	(x 1,000)	(Au g/t)	(x 1,000)
Verde	Measured	19,335	0.73	456
	Indicated	26,987	0.71	619
	M&I	46,322	0.72	1,075
Pancho	Measured	7,585	0.52	128
	Indicated	57,549	0.58	1,071
	M&I	65,134	0.57	1,198
Total		111,456	0.63	2,274

The Maricunga deposits also host an Inferred Mineral Resource of 134.71 million tonnes averaging 0.56 g/tonne Au, given the same $625 US per ounce gold price.  Pancho hosts 132.90 million tonnes averaging 0.56 g/tonne Au and Verde hosts 1.81 million tonnes averaging 0.66 g/tonne Au.

The Maricunga Mine is an operating mine with significant infrastructure in place. The author is of the opinion that the mineral reserve estimates presented in this document are sufficient to support the operation for 16 years based on daily production of 44,000 tpd increasing to 49,000 tpd in 2012 as per the Life-of-Mine Plan approved for 2008. The larger reserve base defined in 2007 at Maricunga will require an increase in leach pad facilities.  A study for establishing an additional permanent leach pad was completed in 2006.  In 2007, CMM undertook some engineering/design work to increase leach pad capacity.  Capital costs associated with pad development, the 16 years mine life and the envisaged SART plant

15

have all been considered for this mineral resource and mineral reserve update. This document supports the December 2007 Maricunga Mineral Resource and Reserve Statement.

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2.0          INTRODUCTION AND TERMS OF REFERENCE

2.1          Introduction

This Technical Report for the Maricunga Mine has been prepared to support the 2007 Kinross Gold Corporation (Kinross) Annual Mineral Resources and Reserves Statement and has been prepared under the direct supervision of the qualified person, Maryse Bélanger, P. Geo., Director Technical Services, Kinross Americas with contributions from CMM technical staff and Kinross Technical Services.

2.2          Terms of Reference

Unless otherwise noted:

·all units of measurement in the following report are in metric measure;

·all costs are expressed in terms of United States dollars;

·all metal prices are expressed in terms of United States dollars;

·a foreign exchange rate of $1.0 US = 530 Chilean pesos was used.

2.3          Scope of Work

The following Technical Report considered the following:

·regional and local geology, structure, alteration and mineralization;

·sample collection, preparation, security and analysis;

·quality assurance and quality control procedures;

·data entry, verification, management, security and storage;

·block modelling, grade interpolation and resource estimation;

·metallurgical recovery, plant design and performance;

·mine planning, scheduling and reserve estimation;

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·leach pad expansion;

·environmental and operational permitting;

·operating and capital cost estimates and

·financial models.

2.4          Sources of Information and Report Basis

This Technical Report is based on verified historical data as well as data collected during a recent field program at Pancho in 2006.  All data was collected under the direct supervision of experienced field geologists.  Operational data since production resumed in 2005 was also considered in the mineral reserves estimates presented herewith. The updated metallurgical test work, new block models for Verde and Pancho and updated operating and capital costs form the basis of the mineral reserve and mineral resource estimates.

This report has been prepared under the direct supervision of the qualified person with direct contributions from CMM technical staff.  A number of other sources of information have been used in the compilation of this report and a complete list of references is provided in Section 21 of this report.

2.5          Field Involvement of the Qualified Person

M. Bélanger has been associated with the Maricunga Mine since November 2003 and has visited the site on numerous occasions during that time frame. In 2006, a total of 75 days were spent at the Maricunga mine site. Regular visits were also made during the course of 2007.

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3.0          RELIANCE ON OTHER EXPERTS

This document has been prepared by Kinross Technical Services with input from Compañia Minera Maricunga. The document summarizes the professional opinion of the author and it includes conclusions and estimates that have been based on professional judgement and reasonable care and are based on the information available at the time this report was completed. All conclusions and estimates presented are based on the assumptions and conditions outlined in this report. This report is to be issued and read in its entirety.

The author has relied upon, and believes there is a reasonable basis to rely upon the contribution of other parties as defined below.

Services Provided by Other Parties

During the course of the fieldwork it was necessary to rely on expertise supplied by third party professionals with technical expertise beyond the experience of the field personnel managing the work. Areas impacted by the work of third party consultants are clearly noted. Kinross relied inherently on the conclusions and recommendations of the following third party consultants:

Golder Associates (2003)	Geotechnical Design Parameters
A. Karzulovic and Assoc. Ltda. (2007)	Geotechnical Design Parameters
McClelland Laboratories	Metallurgical Testing Program
Guillermo Contreras and Sons Limitada	Land Survey

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4.0PROPERTY DESCRIPTION AND LOCATION

The Maricunga Property is located in the Maricunga District of Region III in Chile. The property is located 120 km due east of the city of Copiapó at elevations between 4,200 m and 4,500 m above mean sea level. The Maricunga Mine (formerly known as Refugio) is located at 27 degrees 33 minutes south latitude and 69 degrees 18 minutes west longitude.

Figure 4-1 is a generalized location map of the project.

Figure 4-1: Maricunga Location Map

4.1Claim Status

All surface and mineral claims, surface rights and water rights are maintained in good standing. Mining claims total 8,380 hectares while the exploration properties held by CMM include 5,900 hectares.  Chilean attorneys monitor claim status on behalf of CMM annually. No ownership issues with respect to the mineral claims hosting the project’s mineral resource and reserve estimates have been identified. The project mineral rights and claims

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are listed in Table 4-1 below. The position of the CMM claims relative to the main infrastructure and open pit location is shown in Figure 4-2.  Figure 4-3 shows the detailed mineral rights and claims.

In Chile, mining claims are obtained through a judicial process that requires all survey information to be filed with the appropriate government agency. All CMM claims were surveyed by Guillermo Contreras and Sons Limitada of Santiago.

4.1.1Surface Rights

In addition to the mineral claim rights, CMM also holds title to surface rights at Maricunga, providing the land required for the leach pads, waste dumps, Verde pit, Pancho and Guanaco targets.  All surface rights affecting the ongoing mining activity have been reviewed. No fatal flaws have been identified that would negatively affect the development of new areas for mining activities.

4.1.2Water Rights

Water extraction rights, totalling 258 litres per second have been secured by CMM. Permits are in place and maintained. No issues regarding water rights have been identified that would adversely affect the project.

4.1.3Royalty Payments

The mine production at Maricunga is subject to royalty payments in US dollars based on the realized gold price per ounce. The underlying royalty payments for Maricunga consist of an NSR combined with payment based on Net Operating Margin (NOM).  It is effectively a sliding scale.  At a gold price of $400 the payments are equivalent to $9.471 per ounce.

4.1.4Environmental Liabilities

There is no underlying environmental liability affecting CMM.

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Figure 4-2: Land Claims and Infrastructure Location

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Figure 4-3: Details of Mineral Rights and Claims

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Table 4-1: Mineral Rights and Claims

Maricunga Mine Site  Properties
Claim Name	Assigned Blocks	Hectares
Maricunga 1-112	1 through 112	560
Maricunga 1-40	17 through 20	20
Maricunga 1-40	27 through 30	20
Maricunga 81-120	97 through 100	20
Maricunga 121-160	121 through 124	20
Maricunga 241-280	261 through 262	10
	TOTAL	650

Demasias on Maricunga 1 through 112 Western
Boundary 5,6 Aprox.

Buffer Zone Properties
Claim Name	Assigned Blocks	Hectares
Maricunga 1-40	1 through 16	80
Maricunga 1-40	21 through 26	30
Maricunga 1-40	31 through 40	50
Maricunga 81-120	93 through 96	20
Maricunga 81-120	113 through 120	40
Maricunga 121-160	125 through 160	180
Maricunga 161-200	161 through 167	35
Maricunga 161-200	171 through 177	35
Maricunga 161-200	181 through 187	35
Maricunga 161-200	191 through 197	35
Maricunga 201-240	201	5
Maricunga 201-240	221 through 223	15
Maricunga 241-280	241 through 260	100
Maricunga 241-280	263 through 280	90
Anillo Seis 1-30	3 through 15	130
Anillo Seis 1-30	18 through 30	130
Anillo Siete 1-20	1 through 5	50
Anillo Siete 1-20	11 through 14	40
Anillo Ocho 1-20	1 through 2	20
Anillo Ocho 1-20	6 through 7	20
Anillo Ocho 1-20	11 through 12	20
Anillo Ocho 1-20	16	10
Anillo Trece 1-20	8 through 10	30
Anillo Trece 1-20	20	10
Anillo Catorce 1-20	1 through 17	170
Anillo Quince 1-20	1 through 2	20
Anillo Dieciseis 1-30	10 through 11	20
Amparo Uno 1-60	45 through 60	80
Amparo Dos 1-26	25 through 26	10
Amparo Seis 1-60	31 through 32	10
Amparo Once 1-60	5 through 30	130
Amparo Once 1-60	35 through 60	130
Amparo Doce 1-36	1 through 36	180
	TOTAL	1960

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Outside Zone Properties
Claim Name	Assigned Blocks	Hectares
Maricunga 81-92	81 through 92	60
HIELO 1 1 AL 20		200
HIELO 2 1 AL 10		100
HIELO 3 1 AL 10		100
HIELO 4 1 AL 10		100
HIELO 5 1 AL 10		100
HIELO 6 1 AL 10		100
HIELO 7 1 AL 10		100
HIELO 8 1 AL 10		100
HIELO 9 1 AL 5		10
FRANCISCO 1 AL 10		100
LAGUNA 1 1 AL 10		100
LAGUNA 2 1 AL 10		100
LAGUNA 3 1 AL 20		200
LAGUNA 4 1 AL 20		200
LAGUNA 5 1 AL 30		300
LAGUNA 6 1 AL 30		300
LAGUNA 7 1 AL 30		300
LAGUNA 8 1 AL 20		200
LAGUNA 9 1 AL 18		180
LAGUNA 10 1 AL 14		130
LAGUNA 11 1 AL 8		80
CAMI 48 1 AL 20		100
CAMI 49 1 AL 40		200
RANCHO 1 1 AL 80		400
RANCHO 2 1 AL 60		300
	TOTAL	4160

Exploitation Concessions Located over the Buffer Zone and the Maricunga Mine Site
Claim Name	Assigned Blocks	Hectares
TADEO 1 AL 70		350
JUAN 1 AL 70		350
TOMAS 1 AL 70		350
PEDRO 1 AL 18		90
JOSE 1 AL 80		400
SANTIAGO 1 AL 78		390
JUDAS 1 AL 44		220
PABLO 1 AL 42		210
MATEO 1 AL 4		20
MARCOS 1 AL 46		230
	TOTAL	2610

Compañía Minera San Damian Exploitation Concessions
Claim Name	Assigned Blocks	Hectares
AMPARO 1 1/60		220
AMPARO 2 1/26		120
AMPARO 3 1/40		200
AMPARO 4 1/24		120
AMPARO 5 1/60		290

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AMPARO 6 1/60		290
AMPARO 8 1/36		180
AMPARO 9 1/40		200
AMPARO 11 1/60		40
ANILLO 1 1/30		300
ANILLO 2 1/30		300
ANILLO 3 1/20		200
ANILLO 4 1/30		300
ANILLO 5 1/30		300
ANILLO 6 1/30		40
ANILLO 7 1/20		110
ANILLO 8 1/20		130
ANILLO 9 1/20		200
ANILLO 10 1/30		156
ANILLO 11 1/20		200
ANILLO 12 1/20		200
ANILLO 13 1/20		160
ANILLO 14 1/20		30
ANILLO 15 1/20		180
ANILLO 16 1/30		280
ANILLO 17 1/30		300
ANILLO 18 1/5		50
ANILLO 19 1/30		300
ANILLO 20 1/30		300
ANILLO 21 1/20		200
ANILLO 22 1/20		200
ANILLO 23 1/20		200
ANILLO 24 1/20		200
ANILLO 25 1/20		200
CORINA 1/15		150
LAGUNA I 1/30		300
LAGUNA II 1/30		300
LAGUNA III 1/30		300
LAGUNA IV 1/30		300
LIZ 1/30		300
LOIDA 1/20		200
MARICUNGA 41/80		200
MARICUNGA 81/120		60
MARICUNGA 161/200		60
MARICUNGA 201/240		5
MARICUNGA 201/240		5
MARICUNGA 201/240		170
REBECA 1/11		29
TRANQUE 1 1/60		300
TRANQUE 2 1/60		300
TRANQUE 3 1/24		120
TRANQUE 4 1/24		120
YEZENIA 1/8		76
	TOTAL	9991

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5.0ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1Accessibility

Access to the property is via 156 km of a two-lane dirt road connecting with the paved highway C-35 approximately 10 km south of Copiapó. The first 96 km of the dirt road constitute a section of a national highway crossing into Argentina. The road then branches out to the northeast to the mine site. The final 60 km is a private CMM road and it is generally in good to very good condition.

5.2Climate and Physiography

The property is located 120 km due east of the city of Copiapó at elevations between 4,200 m and 4,500 m above mean sea level. The Maricunga Mine is located in steep, mountainous terrain with slopes up to 35%. The site is largely devoid of vegetation with the exception of the spring-fed marshes found along the valley floors. The climate is arid with an average annual precipitation of 87 mm, most of which is realized as snowfall during the winter months (March through August).  Generally, very little precipitation occurs during the summer months (September through February). Local wildlife is sparse.

5.3Local Resources and Infrastructure

The Maricunga Mine site includes leaching and ADR facilities, power generation, maintenance shops, office facilities and camp. The existing infrastructure also includes an established fresh water supply system, process water supply system, truck shop, warehousing and lay down areas, an in pit crushing and conveying system and a secondary/ tertiary crushing and screening plant. Significant upgrades designed to increase production throughput were completed in 2005. They included the in-pit crushing and conveying system as well as the secondary/tertiary crushing and screening infrastructure.  All facilities are in good operating conditions.  Surface and water rights are previously discussed in Section 4.0.

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Chile features a strong mining culture with well-established support centers in Santiago, Antofagasta and Copiapó. The town of Copiapó is the primary staging and support area for the mine with various well-established contractors.

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6.0HISTORY

David Thomson and Mario Hernandez discovered gold mineralization at Maricunga in 1984. Shortly after their discovery, Hernandez, Thomson, and three other partners acquired the existing claims at Maricunga for Compañia Minera Maricunga (CMR).  CMR completed geologic mapping and geochemical sampling, identifying anomalous gold values in three areas:  1) Cerro Verde, 2) Cerro Pancho, and 3) Guanaco.  A detailed discussion of exploration activities by previous owners is discussed in Section 10.

In 1985, Anglo American Chile Limitada (Anglo) optioned the property from CMR.  Anglo explored the property for three years, returning the claims to CMR in 1988.

In 1989, CMR signed a letter of intent to explore the Maricunga property with Bema Gold Corporation (Bema). Bema commenced exploration fieldwork in October 1989 and from 1989 to 1991, completed 51,765 meters of drilling at Verde with an additional 5,088 meters at Pancho. Bema also commissioned Mineral Resources Development Inc (MRDI) to complete a feasibility study on the project, which indicated positive project economics. In January 1993, Bema exercised its option rights, obtaining a 50% interest in the Maricunga properties.  At the same time, CMR sold its remaining 50% interest to Amax Gold Inc. (Amax).  Amax (operator) and Bema formed Compañia Minera Maricunga (CMM), a 50/50 joint venture to develop and operate the Maricunga project. From 1993 through 1997, CMM continued developing the project, beginning commercial production in 1996.

The mine first operated from 1996 to 2001, producing more than 920,000 ounces of gold from 46.0 million tonnes of ore. In 1998, Kinross Gold Corporation (Kinross) acquired Amax’s 50% interest through a merger agreement.

Mining operations at Maricunga were suspended during the second quarter of 2001, a result of depressed gold prices. Improving gold prices during 2002 prompted CMM to evaluate options for re-opening the mine. Scoping studies quantifying the conditions necessary to re-open the mine were completed. The studies indicated that the project economics required an increase in the existing reserve base. As a result, an exploration program was developed to evaluate the reserve potential at depth at the Verde deposits and the inferred resource at the nearby Pancho deposit, located approximately 2.0 km northwest of the Verde pit.

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A 35,000 meter drill program was approved in September 2002 with the objective of adding 20 million tonnes of Measured and Indicated resource at both Pancho and Verde. This mineralization had to be recoverable through open pit mining methods. Follow-up work to further delineate the mineralization, provide representative metallurgical samples and detailed geotechnical information would be contingent on the initial results of the drilling.

The information was collated and incorporated in new mineral resource and mineral reserve estimates for Verde and Pancho. The reserves were based on a detailed engineering study examining the economics of the project assuming a capital investment of $101.1 million to upgrade the existing infrastructure, allowing the plant to process 40,000 tonnes per day of Verde ore and 35,000 tonnes per day of Pancho ore. The capital investment was based on preliminary design and first-principle engineering. Capital cost estimates were prepared and verified by two, independent, North American engineering consultants with significant experience in Latin America.

The Maricunga Mine resumed full production in October 2005 at a rate of 40,000 tpd. Total production of 233,000 ounces from 14,333,000 tonnes at 0.74 g/t Au is reported for 2007.   Table 6-1 summarizes the production history of Maricunga.

Table 6-1: Annual Gold Production of the Maricunga Mine

	Mined	Grade	Gold
	tonnes	mined	production
Year	(x 1,000)	(Au g/t)	(ounces)
1996	7,617	1.03	101,276
1997	7,789	1.02	147,085
1998	8,207	0.93	161,046
1999	8,936	0.94	179,465
2000	8,801	0.94	169,832
2001	4,643	0.95	133,947
2002	—	—	26,094
2003	—	—	5,000
2005	6,214	0.87	121,670
2006	14,872	0.74	233,836
2007	14,333	0.74	233,000
TOTAL	81,412	0.88	1,512,251

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In early 2006, CMM made the decision to drill at the Pancho deposit to better define the mineral resource both laterally and at depth since the drilling completed in 2003 focused largely on definition drilling for a $350 gold pit design. During 2007, both Verde and Pancho models were refined to better reflect a new geologic interpretation and the mining experience gained over the years. The SART process was tested for the Pancho ore as uneconomic concentrations of copper are present in ore. Some of the copper minerals are soluble in cyanide solutions and these minerals consume cyanide reagent and contaminate the leach solution with copper metal.  A feasibility study and pilot testing program were completed by Idesol of Santiago. The study considers plant design, performance metrics, operating and capital costs and implementation strategy. The mineral reserves defined at Pancho consider the application of SART.

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7.0GEOLOGICAL SETTING

The Verde and Pancho gold deposits at Maricunga occur in the Maricunga Gold Belt of the high Andes in northern Chile.  Since 1980, a total of 40 million ounces of gold have been defined in the belt, (Muntean and Einaudi, 2000).

Basement rocks in the Maricunga area are Palaeozoic to early Tertiary age.  The oldest rocks are the Late Pennsylvanian to Triassic aged rhyolite ignimbrites and breccias of the Pantanosa Formation. This unit has been uplifted along a northerly trending, westerly dipping reverse fault.  The Pantanosa Formation is faulted over a package of interlayered redbeds and greenstones of the Monardes and Agua Helada Formations.  These formations are Late Jurassic to Early Cretaceous in age.  They comprise an east dipping (55°) sequence of strata with a thickness of 900 m. A 200 meter thick sequence of andesitic volcaniclastic sedimentary rocks overlies the redbeds and greenstones.  These strata are correlated with Late Cretaceous to Early Tertiary rocks of the Quebrada Seca, Quebrada Paipote, and Las Pircas Formations.

Gold mineralization at Maricunga is hosted in the Maricunga volcanic-intrusive complex of Early Miocene age. These rocks are largely of intermediate composition. Radiometric age dating indicates these rocks are between 22 to 24 Ma in age. (Muntean 1998).

The Maricunga volcanic-intrusive complex is exposed over an area of 12 km² and consists of andesitic to dacitic domes, flows, and breccias that are intruded by subvolcanic porphyries and breccias (Muntean 1998).  Distinguishing between volcanic and intrusive rocks in the complex is difficult as there are only minor differences between the various units.

7.1Alteration

7.1.1Verde

Alteration assemblages observed at Verde and Pancho are generally supportive of porphyry style mineralization but the intensity of the alteration fabric tends to be weak. Potassic alteration has been observed at Verde but generally tends to be rare. Silicification is local and patchy.  Propylitic alteration, although variable on small scale, appears ubiquitous on a

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mine scale and in global terms does not change laterally or vertically.  For that reason, primary alteration (including silicification) zonation has not been mapped.

Supergene alteration, which directly affects gold recovery, occurs deeper within fracture and fault zones.  Sericite and chlorite are replaced by clay minerals, magnetite by hematite and pyrite by jarosite.  The oxidation and the leaching by meteoric waters has penetrated to variable depths within the deposits depending on the fracture intensity and faulting. The supergene alteration is accompanied by the deposition of limonite, manganese oxides, clay, sericite, jarosite along with gypsum in the argillic altered zones.

7.1.2Pancho

Gold mineralization at Pancho is associated with a central zone of potassic alteration, which is manifested by replacement of mafic minerals, by fine grained, secondary biotite and magnetite. Partial replacement of plagioclase by k-feldspar can also be observed.  Muntean (1998) has documented a more restricted core to this potassic zone consisting of magnetite-k feldspar-plagioclase replacement which grades outward into the more widespread secondary biotite zone. Similar to Verde, much of the potassic alteration is obliterated or obscured by a later chlorite overprint.

The intermediate to upper parts of the system are dominated by pyrite-albite-clay-sericite alteration, primarily in the volcanics but also overprinting the upper parts of the potassic zone. This alteration is late and probably has a large supergene component.

In the uppermost parts of the system analysis indicates the presence of hypogene alunite, dickite and pyrophyllite, characteristic of epithermal, high sulfidation alteration. This hypogene assemblage occurs together with strong supergene alunite, kaolinite and other clays.

The presence of this high level alteration within 100 - 200 m of the potassic zone suggests strong telescoping of the system.

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7.2Structure

Most of the structural trends affecting the Verde and Pancho deposits are related to fracture systems rather than fault zones. The dominant trends are north-northeast and north-northwest fracture systems that can be readily identified at Verde. At Pancho, the dominant structural trend is northwesterly, reflected not only in mineralized structures but also in late, post mineral structures.

One of the main structural features influencing the Pancho deposit is Falla Guatita fault zone.  Field mapping suggests that there may be significant vertical displacement on this structure.  The distinctly higher copper grades and presence of potassic alteration in the main zone suggests that the central portion of the Pancho deposit may have been down dropped. Figure 7-1 is a geological map of the Maricunga Area (after Muntean, 2000).

Figure 7-1: Geology of the Maricunga Mine Area

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8.0DEPOSIT TYPES

8.1Verde

Gold mineralization at Maricunga has been interpreted to be porphyry style gold systems.  At Verde West, gold mineralization is centered about an elliptical porphyry plug measuring 175 by 100 m and oriented at N30°W. At Verde East, the porphyry plug measures 130 by 80 m and is oriented at N35°E. The porphyries occur within a sequence of intermediate tuffs, porphyries and breccias that are the host rocks to the gold mineralization. Lithological interpretation at Verde has identified six major lithologic units. These are:

·Post mineral intrusives (barren)

·Mineralized post mineral intrusives

·Verde breccia

·Dacite porphyry

·Dacite tuffs

·Laguna tuff (barren)

The most favourable ore hosts are the Verde Breccia and Dacite Porphyry units. The dacite porphyry is a volcanic to hypabyssal intrusive rock with 20% to 40% plagioclase phenocrysts in a fine-grained matrix.  It contains phenocrysts of biotite, hornblende, and sparse quartz.  The unit is the best host for stockwork veining, with some portions containing up to 20% quartz-magnetite ± pyrite veinlets.

The Verde Breccia consists of intrusive breccia and/or volcanic tuff breccia.  Its geometry suggests an intrusive origin. The unit consists of breccia with angular to rounded clasts, generally matrix supported.  The breccia is generally green to greenish gray, which is one of the diagnostic features of the unit.  The color of the rock is largely due to chlorite and occasionally epidote. The fragments range from 2 mm to greater than 2 meters in size.  Locally the unit is mostly matrix-sized material with only sparse clasts.   Clasts in the unit are generally monolithic volcanic fragments, which are often porphyritic with white plagioclase laths up to 5 mm in length.  The unoxidized rock typically contains 0.5 to 1% pyrite.  Quartz-magnetite veinlets are common in the mineralized portions of the unit.  In places, the breccia

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is cut by fine-grained matrix material that forms clastic-like dikes.  These are generally > 2 cm in width.  They may have been formed as “fluidized” material injected into fractures at the time of the formation of the unit.

8.2Pancho

The gold mineralization at Pancho is also described as porphyry style mineralization. It occurs within a sequence of intermediate tuffs, porphyries and breccias that are the host rocks to the gold mineralization. Lithological interpretation based on the recent 2006 drill program has identified six major lithologic units. These are from older to younger rocks:

·Hornfels – They represent less than 5% of the total volume. They were identified at depth in the Pancho porphyry system. They are present in a series of elongated bodies, sub-horizontal and intercalated with intrusive breccias and dioritic porphyry. The hornfels are also characterized by widespread intense silicification.

·Diorite Porphyry - Rocks hypabyssal that represent 60% of the volume of the Pancho intrusive complex.   It is the most frequent host of mineralized A, B and T veinlets.  The diorite porphyry also shows signs of being intruded by a smaller intrusive phase. The porphyry displays obvious porphyry texture with 20 to 40 % phenocrysts. Phenocrysts alteration pattern include biotite or hornblende being completed altered and replaced by assemblages of chlorite-quartz-sericite-magnetite-hematite.  At depth the porphyry clearly shows potassic alteration.

·Intrusive Breccias – The intrusive breccias represent approximately 15% of the volume of material in the Pancho complex. They are generally elongated and sub-horizontal.  These breccias are characterized by their fragments of dioritic porphyry. The fragments vary in size for mm to several cm within a matrix which is normally fine-grained. The intrusive breccias are a significant host of gold mineralization with A, B, T veinlets.

·Diorite Porphyry (2) - This diorite porphyry is defined as a small body identified between two drill sections and found at depths ranging from 300-400 meters. 400 m. In theory, it intrudes the main porphyry but in turn is intruded by the PMI.  It is very similar to the main porphyry body in terms of mineralogy, alteration and mineralization. The main difference observed is its primary biotite that is partially altered into chlorite. It also has less veinlets than the main diorite porphyry.

·Dioritic Post Mineral Intrusive – The Post Mineral Intrusive (PMI) identified at Pancho is not a PMI in a strict sense as there is some very low-grade mineralization associated with it.

·Volcanic Breccias – Found in the upper portion of the deposit they represent approximately 15 % of the volume of material explored to date at Pancho. The volcanic breccias are normally sub-horizontal and discordant with the other units described.

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Figure 8-1 shows the simplified geology model for Pancho.

Figure 8-1: Simplified Geology for Pancho

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9.0MINERALIZATION

9.1Verde

Gold mineralization at Verde is interpreted to be the result of the fracturing and concentration of fluids in the carapace of an intrusive plug or stock. Gold is closely associated with quartz, magnetite, calcite, and garnet stockworks. Approximately 80% of the stockwork veins are generally dark grey in color, finely banded, with magnetite. The remaining 20% are principally white quartz veins. Gold mineralization is postulated to have resulted from at least 2 phases of mineralization, the first is a lower grade phase associated with copper and probably the porphyry emplacement event.  The second phase is a higher-grade gold only event possibly associated with the structurally emplaced veinlet swarms and northwest trending sheeted veinlet zones more evident in Verde East. Verde East and Verde West cover an area of 1.4 km in length by 700 km in width. Drilling has identified gold mineralization to a depth of a 600 m.

9.2Pancho

The porphyritic diorite intrusives and intrusive breccias are the main hosts for mineralized veins. Mineralization is also hosted by the volcanic rocks. The mineralization is commonly related to the various geologic features listed below:

·directly associated with the presence of B, A and T veinlets

·in silicified zones, normally micro-granular and a dark grey color, with or without the presence of magnetite

·associated with intrusive breccias

·in contact zones such as the contact between the intrusive breccia with the volcanic breccia or the contact between the intrusive breccia and the diorite porphyry

In the porphyry the veins are present in stockwork or sheeted veins. They are generally subvertical and have a strong, preferred north-westerly strike. The northwest structural control is evident not only at outcrop scale but is also reflected in the northwest alignment of intrusives and the three centers of mineralization in the district, Verde, Pancho and

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Guanaco.  The mineralization outlined to date at Pancho comprises an area of 800 m by 700 m.  Gold porphyry mineralization was identified in drilling at depths exceeding 600 m.

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10.0EXPLORATION

Exploration of the Verde and Pancho deposits has been ongoing since 1984 when the Maricunga Property was identified as a gold prospect in 1984 by David Thomson and Mario Hernandez during a field visit to areas of alteration.  The existing claims on the property were acquired by the partnership Companio Minera Maricunga (CMR) consisting of Hernandez, Thomson, and three other partners.

CMR subsequently completed a program of geological mapping, rock chip and geochemical sampling and identified three large areas of alteration with anomalous gold values.   These areas were named Cerro Verde, Cerro Pancho and Guanaco.

In December 1985, CMR optioned the Maricunga Property to Anglo American for 3 years.  During that period Anglo American accomplished the following:

·Geological mapped the property at a scale of 1:5,000

·Collected 2,161 surface samples over an 8 km² area and assayed for gold, silver, copper, molybdenum and zinc.

·Constructed 7.5 km of roads

·Geologically mapped Pancho and Verde at 1:1,000 scale

·Cut 6,000 m trenches on Pancho; collected and assayed 1,682 samples

·Cut 7,500 m of trenches on Verde; collected and assayed 2,350 rock samples

·Drilled 45 percussion drill holes on Pancho to average depths of 50 m (total of 2,234 m), along with 6 diamond drill holes (total of 366 m).

·Drilled 35 percussion holes, also to a depth of 50 m on Verde (total of 1,744 m).

Anglo American’s exploration program outlined broad areas of gold mineralization on Pancho, Verde and Guanaco, targeted by geochemical anomalies of +100 ppb gold in the soil.

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In December 1988 the property reverted to the owners.  In early 1989, CMR solicited bids from twelve mining companies for Maricunga; in July, Bema was chosen and in September a letter of intent was signed.  Field work commenced the following month.

Phase I of the Bema program was designed to test the gold mineralized zones on Verde, which showed much greater depths of oxidation than Pancho.  Since 1989, drilling has been the primary means of exploration.  Drilling on the property is further summarized in Section 11.

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11.0DRILLING

Historically, most of the drilling at Maricunga consisted of reverse circulation drilling. The destructive nature of this drill method made identification of lithology, structure and alteration difficult. The 2002 - 2003 drilling consisted primarily of diamond drill core, providing site geologists with an opportunity to refine the geology model of the deposit.

The preferred orientation of drilling for the Verde deposits is 060 - 240° azimuths.  At Pancho, the preferred orientation for the drilling grid of 038 – 218° was established based on the northwest strike of veinlet swarms observed in drill roads. The surface expression of the Pancho system is approximately 700 m by 800 m. The depth for the most recent drill holes was generally determined using a $600 gold pit and the limits of the porphyry mineralization at depth have not been clearly defined as of yet. The mineralization being disseminated in nature also implies that its geometry and size are not limited to a vein. Therefore, the concept of true thickness in drill holes does not apply to Pancho and Verde.

The 2006 exploration campaign followed the same field procedures implemented in 2002-2003. A total of 667 holes (103,392 m) of drilling has been completed on the Verde deposit with an additional 210 holes (56,748 m) completed at Pancho. The drilling has resulted in a drill spacing of approximately 50 x 50 meters at Verde and 75 x 75 meters at Pancho.  Table 11-1 summarizes the various drill campaigns at Verde and Pancho.

Table 11-1: Exploration Drill Summary

		Verde				Pancho
Company	Year	RC	RC	Core	Core	RC	RC	Core	Core
		(#)	(m)	(#)	(m)	(#)	(m)	(#)	(m)
Bema	1989	45	5,060	—	—	—	—	—	—
Bema	1990	231	46,705	31	4,083	—	—	—	—
Bema	1991	—	—	6	1,090	24	5,088	—	—
Subtotal		276	51,765	37	5,173	24	5,088	—	—
CMM	1993	176	5,060	—	—	—	—	—	—
CMM	1994	—	—	6	4,083	—	—	—	—
CMM	1997	—	—	—	—	15	4,296	—	—
CMM	1998	—	—	18	6,689	—	—	—	—
Subtotal		176	5,060	24	10,772	15	4,296	—	—
CMM	2003	20	3,154	134	27,468	39	6,710	69	14,146
CMM	2006	—	—	—	—	18	4,012	45	22,496
TOTAL		472	59,979	195	43,413	96	20,106	114	36,642

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It should be noted that all the RC holes drilled in 2006 were effectively pre-collar holes down to a depth of approximately 200 m. The holes were completed with a diamond drill rig to depths averaging 400 m.  Figures 11-1 and 11-2 are plan maps of Verde and Pancho showing the completed drill holes as at December 31, 2007.

11.1Rig Setup and Survey

All proposed drill locations were laid out in plan and on section. The collar coordinates were provided to the mine survey crews who laid out the drill pad locations in the field using total station theodolites. Pad locations were verified prior to construction to ensure access and safety. After pad construction, the mine survey crews established the collar location and marked it in the field. They also established the front sight and back sights necessary to provide the drill direction. After the drill was moved onto the setup and prior to the start of drill, geological staff verified the drill alignment and inclination using a compass. The survey crew later verified alignment and inclination when the drill hole was in progress.

Downhole inclinometry was completed at the end of each hole. Gyroscopic azimuth and inclination readings were taken every 10 meters down the hole to within ten meters of hole bottom and every 50 meters back up the hole as a double check.

11.2Topographic Base and Survey Audit

In 2002-2003, all field surveys were tied into the established mine grid. Survey data was incorporated into current as-built plans that were updated and maintained in AutoCAD by the survey crews. Guillermo Contreras and Sons Limitada (Santiago), licensed Chilean surveyors, completed a survey audit that verified an approximate 10% of the drill collars using a differential GPS survey system. No significant errors were noted.

11.3Core Handling, Storage and Security

Drill core was placed in labelled boxes, fitted with lids and transported to the site logging facility by pick-up. All core logging was completed at a facility located at the Maricunga mine site.  Only CMM personnel worked with the drill core.

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Figure 11-1: Verde Drilling Coverage as December 31, 2007

Figure 11-2: Pancho Drill Coverage as at December 31, 2007

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12.0SAMPLE METHOD AND APPROACH

12.1RC Drilling

CMM provided all of the technical support for the sampling, geologic logging, and drill supervision.  Rig geologists and samplers were responsible for the quality/accuracy of each sample. Geologists and samplers typically had up to 15 years experience sampling.

The 2003 drill program adopted a 2.0 meter standard sample length for all samples. The same procedure was followed for the samples in 2006.

When drilling dry, reverse circulation drill cuttings were directed from the cyclone to a Gilson splitter where the sample was reduced to a 6-8 kg sample. Typically, this required a ¹/₈ split of the drill cuttings. The hole and cyclone were blown clean after each rod change. Two samples, an original and a duplicate, were produced for each 2.0 meter interval. The samples were weighed, and the weight, number of splits and hole diameter were recorded. The samples were bagged and tagged and prepared for transport to the on-site sample preparation facility. The splitter was air cleaned after each sample.

When sampling wet, a rotating wet sampler replaced the Gilson splitter below the cyclone.  Usually ground water flow was not sufficient and make up water was necessary to provide uniform flow to the splitter.  As with the dry sample collection, two samples
(6 – 8 kg.) were collected in perforated sample bags at the exit port of the wet splitter.  These samples were not weighed immediately but were left to drain for a 24 hour period prior to transport to the sample preparation facility.

It should be noted that the drill holes completed in the winter of 2006 did not encounter any water.

Typical chip trays were produced for logging and future reference. In addition, a 100 g, unwashed sample (reference sample) was collected for each 2.0 meter interval. All chip trays and reference samples were placed in permanent storage at site.

The samples are considered representative by the author who has visited the drill sites on many occasions.

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12.2Core Drilling

Drill core was received and laid out for logging and sampling at the core logging facility. The first sample of each core hole was marked from the start of core recovery to the end of the first even meter. From that point, samples were laid out every 2.0 meters downhole.  Sample numbers were marked on the core box edges, accounting for number gaps for QA/QC sample insertion. Logged and marked boxes of whole core were transferred to the ALS Chemex sample preparation facility where chain of custody was transferred from CMM to ALS Chemex.

CMM received the sawn core, in the original labelled boxes, from ALS Chemex after splitting. The half-cores were stored shipping containers at the Maricunga site. In all, ten containers were used to store the half-core library.

Core sample rejects are bagged in plastic, labelled and received from the ALS Chemex sample preparation facility.  These bags are placed in cubic meter wooden bins.  The bins are labelled according to hole number and sample range.  The bins are stored beside the core containers and are locked in a covered storage compound.

The samples are representative of the mineralization intercepted at Pancho. The author visited the drill sites at a minimum once a month during the 2006 exploration. The author is satisfied that all established procedures were properly followed by the field staff.

12.3Logging

Experienced geologists logged the reverse circulation chips and drill core at Maricunga. Reverse circulation chips were typically logged at the drill while core was typically logged at the core logging facility. All logging utilized standardized logging forms and a geological legend developed for the Verde and Pancho deposits. The legend has evolved from historic observation and is consistent with both the regional and local geology. The legend and logging records lithology, alteration, structure, geology, mineralization and oxidation (of pyrite) for each two meter sample interval, reducing the geological descriptions to numeric or alphanumeric codes. Unique features not accounted for in the legend are noted in written comments.  The legend stresses oxidation conditions of pyrite, veinlet mineralization, location and amount of magnetite and occurrence of typical porphyry alteration minerals.

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In 2002-2003, Golder Associates from Santiago (Golder) identified all holes requiring detailed geotechnical logging that was completed by Golder trained staff. CMM geologists typically recorded both Rock Quality Designation (RQD) and recovery data for each hole. Recoveries averaged better than 95% during the 2003 program. In 2006, average core recovery was 96%.

12.4Geotechnical Core Logging

Golder managed the geotechnical program.  Francisco Carrasco, (Project Manager) and Joaquin Cabello (Assistant Manager) supervised the filed programs and provided the initial training. The program included detailed geotechnical logging of selected drill holes, uniaxial compressive strength (UCS) testing of selected cores, completion of a seven (7) hole oriented core program and cell mapping of existing pit walls and road cuts. Golder also managed a hydrology study including installation of piezometer wells for monitoring water levels.

Golder provided all initial training and completed frequent site visits to maintain quality control of the work and upgrade training of the field staff performing the work.

CMM adopted identical geotechnical logging procedures in 2006 under the supervision of the same geotechnical engineer employed by CMM in 2003.

12.5Composite Sample Summary

The full composite summary for both Verde and Pancho are shown in Tables 17-4 and 17-5 respectively.

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13.0SAMPLE PREPARATION, ANALYSES AND SECURITY

13.1Sample Preparation

In 2002-2003, all sample preparation, including core splitting (sawing) was performed and supervised by ALS Chemex personnel. ALS Chemex established, equipped and staffed a portable sample preparation facility at the Maricunga mine for the duration of the program. The sample preparation facility included two rock saws, three drying ovens, two Rhino jaw crushers and a Jones riffle splitter.

After logging and sample mark up, the core boxes were sent to the sample preparation facility. ALS Chemex personnel physically sawed the core. The core splitter randomly selected the split line. Core splitters were trained to maintain the same cut line on contiguous pieces of core wherever possible. After splitting, one half of the core was placed in sample trays along with the sample tag. These were placed in the drying ovens for drying. The other half of the core was fitted back into the core box and returned to CMM for placement in permanent storage.

All samples were dried for 4 – 6 hours at 70° Celsius. For wet reverse circulation samples, the dry time was increased to 24 hours. After drying, sample weight was recorded and the sample was single pass crushed to 90% passing 2 mm using the Rhino Jaw crushers. The crushers were cleaned with air after each sample and cleaned with barren quartz after every five (5) samples. Crush size was monitored after every 20 samples to ensure that the final product met the crush size specification.

After crushing, the entire sample was passed through a Jones riffle splitter, reducing the sample size down to a nominal 1.0 kg size. A preparation duplicate was collected every forty (40) samples. The duplicate was distinguished by an “A and B” designation and treated as per the standard sample procedure.  The coarse rejects were returned to the original sample bag, labelled and returned to CMM for storage.

The 1.0 kg samples were sealed in plastic bags, labelled, packed in crates and transported to the primary analytical facility (ALS Chemex – La Serena) for analysis. The laboratory has ISO 9001-2000 accreditation.

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13.2Sample Preparation 2006

In 2006, only core splitting (sawing) was performed at the Maricunga site. All samples were shipped to Actlabs in La Serena where sample preparation was performed using exactly the same procedure as described above. Actlabs also has an ISO 9001-2000 accreditation.

13.3Analysis

The prepared samples were received at the ALS Chemex’s facility in La Serena (the primary analytical lab for the duration of the program) where analyses for Au, Ag, Cu and cyanide soluble Au and Cu analyses were performed.

On receipt of the samples the work order numbers and sample numbers were recorded and the samples were arranged for analysis based on the individual sample numbers. The entire 1.0 kg sample was pulverized to 95% passing -150 mesh using an LM-2 pulverizer. The pulverizer was cleaned with compressed air after each sample. Sieve analyses were performed after every 20^th sample to ensure the final product met the stated specification of 95% passing –150 mesh.

After pulverization, the sample was reduced to a 300 gram sample and 700 gram reject. The reject placed in labelled plastic bags and returned to CMM for storage.

The 300 gram sample is homogenized. After homogenization:

·a 50 gram aliquot is selected for fire assay with atomic absorption finish for gold,

·a second aliquot is selected for total digestion atomic absorption analyses for silver and copper,

·a 20 gram aliquot is selected for cold cyanide soluble gold analysis, leached for 4 hours with 60 ml of cyanide solution (0.1% NaOH & 0.5% NaCN).

·A 1 gram aliquot is selected for cold cyanide soluble copper analysis, leached for 2 hours with 20 ml of cyanide solution.

Dr. B. Smee (Smee and Associates Consulting Ltd.) completed operational audits of the La Serena lab for the 2002-2003 program. The operational audits were performed measuring compliance with analytical best practices as well as to NI-43-101 requirements with respect

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to quality control and quality assurance. Dr Smee did not note any significant problems with this facility, concluding that ALS Chemex’s lab and procedures met the highest industry standards.

Initially, ALS Chemex assigned work order numbers at the mine site but in March 2003 a Lab Information Management System (LIMS) was adopted and work order numbers were generated from a centralized ALS Chemex location.

Results containing Au, Ag, Cu or CN Au, CN Cu were sent via digital mail to the Project and Office Managers.  Attachments included 1 file per hole, identified by work order number and hole number.  The file included internal ALS Chemex blanks and duplicates, while results of ALS Chemex standards were reported monthly in a separate file.

ALS Chemex was also requested to send from a list chosen by CMM 3 batches of pulps (1/40 samples) to Assayers Canada in Vancouver.  Dr. Smee audited the Assayers laboratory in February 2003 and found them acceptable to qualify as a referee lab for the project.

13.4Quality Control / Quality Assurance

13.4.1Drilling 2002-2003

The original QA/QC program was designed and monitored by Dr. B. Smee. The program included the insertion of one blank, one standard, and one duplicate (randomly selected) for each batch of 40 samples. Standards were not available for insertion for the first 3 weeks of the program. The standards were designed by Dr. Smee and provided by Rocklabs, New Zealand. Dr. Smee selected the standards to best match Maricunga rock types and grades. Standards were designed to monitor gold grade only. The standards were not blind to the analytical lab.

Duplicate samples were randomly selected within the sample batches.  Reverse circulation duplicates were taken as an opposite split post cyclone and given a sequential sample number.  RC duplicates were not taken below the water table. As the duplicate was generated in the field, no ALS Chemex personnel could identify these samples as duplicates.  For core, two duplicates of one interval were taken.  The first (d1) was a second

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split of post crusher product.  The second (d2) was the actual other half of sawn core.  The intent was to see the variance according to crushing procedure (d1) and also the variance according to natural geological variance + crushing procedure.   This would also allude to any bias in core sawing activity.  In the core duplicate samples, obviously ALS Chemex sample preparation personnel knew which samples were being duplicated, while ALS Chemex analytical personnel did not.

ALS Chemex also inserted their own blanks, standards and duplicate samples for each sample batch as part of the labs own internal quality management program.

13.4.2Drilling 2006

The same overall QA/QC procedures were implemented during the 2006 exploration campaign. The results were compiled as soon as received by the site staff.  Results for standards were plotted using two standard deviations as lower and upper limits.

The charts were kept updated and any anomalous results were discussed with Actlabs.

13.4.3Table of Failures

The Senior Geologist extracted the QA/QC sample results from the digital result files based on the sample numbers. If a standard assay fell outside 3 standard deviations of the standard value, the entire batch of 40 samples was rerun.  If 2 consecutive standard assays fell within 2 standard deviations of the standard value, a batch rerun was requested. If a blank exceeded 0.05 g/t au, a batch rerun was requested.   No reruns were determined from duplicate data.

Monitoring of the QA/QC data required a total of 17 batch reruns (649 samples) based on standards failures. This represented approximately 2% of the total analyses run for the project. Inspection of batch failure results indicates minimal bias with both gold and copper relative to the repeat batch assays.  At the end of the project, an additional nine QA/QC batch assays (347 samples) were selected based on standards failures.  Three additional batches plus 11 individual samples (127 total) were found to have CNAu at least 0.10 g/t au greater than the total gold assay.  These were also rerun for both total and cyanide.

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13.4.4     Analysis of QA/QC Data

For the 2002-2003 drilling program all standard, blank and duplicate results were sent to Dr. Smee for analysis.  Duplicate results were analyzed for precision and plotted on Thompson Howarth graphs. Dr. Smee found the error rate and duplicate variance normal for porphyry gold deposits.  In his final report, (July 20,2003) he concluded the gold assay database to be of commercial quality but raised concerns about eight of the 650 ALS Chemex digital files.  All eight files, containing failures, had already been reanalyzed, were determined to have had CMM field labelling errors, or were duplicated assay sets.

The 2006 QA/QC data was compiled as it was received by the geologists at the site. Any anomalous results were flagged and the lab contacted directly to discuss the variations observed. A series of samples representing approximately 10% of the total were also sent to ALS Chemex as an independent check.

13.4.5     Referee Laboratory

Assayers Canada (Assayers), Vancouver, was selected as the referee laboratory for the program. Assayers Canada has achieved Certificates of Laboratory Proficiency from the Standards Council of Canada for precious and base metal analysis.  One pulp per ALS Chemex batch of 40 was selected at random, including the CMM standards, blanks and duplicates.  Additionally one of the 3 CMM standards was inserted in this sample stream to check Assayers at a frequency of 1 per 40. Three batches of checks (631 samples in total) were sent to Assayers over the duration of the program.  Results were forwarded to Dr. Smee who concluded that the two labs essentially agreed on gold and cyanide soluble gold analysis within normal precision limits.

13.4.6     Statement of Author’s Opinion

The author concurs with Dr. Smee’s opinion that the slight variations observed were within normal precision limits.

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14.0        DATA VERIFICATION

14.1        General 2002-2003

Kinross Technical Services verified the Initial Database (IDB) to copies provided by Bema Gold Corporation and MRDI.

Original drill logs, assay certificates, survey worksheets and other original data were located and used to verify the data contained in the IDB.  Verification of the historical data included:

·Collar location

·Collar azimuth and dip

·Downhole azimuth and dip

·Total gold and copper analyses

·Cyanide soluble gold and copper analyses

·Primary lithology code

·Primary alteration code

·Primary oxidation code

Assay certificates, and original logs were found for approximately 80% of the historical data.

For the 2006 Pancho drill program a system of double entry was established. It included all of the items listed above plus the geology logs as prepared by the field geologists. The final grade database was compiled by Kinross Technical Services (KTS). KTS and CMM assumed that the historical data used in the past was sufficiently free of errors or omissions. Details of the data verification process and database management for 2002-2003 are provided in the following paragraphs.

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14.2        Pre-2002 Data Verification

14.2.1     Collar Data

Verification of the historic collar data identified one error of note. Collar coordinates of drill holes completed between 1986 and 1987 did not match the coordinates recorded on original survey sheets.  This error was also corrected.

14.2.2     Survey Data

Much of the historical downhole survey data could not be located. The limited downhole survey data that was located was used to verify the data in the IDB. No significant errors were noted.

14.2.3     Lithology, Alteration, Oxidation, Vein and Mineralization Data:

Ten percent of all lithology, alteration, oxidation, vein description and mineralization data for the 1997 – 1998 drill campaign was verified as this data was collected using field practices and logging formats consistent with the 2002-2003 effort.

14.2.4     Analytical Data

Ten percent of the historical analytical results in the IDB were verified against original assay certificates.  Rare errors were detected during the verification process but these were almost exclusively related to the recording of results below the detection limit for successive programs.

14.3        2002-2003 Data Verification

14.3.1     Collar Data

Prior to entering collar data into the database, collar locations were visually checked for obvious errors such as transposed northings and eastings. Once entered, collar coordinates and hole lengths were extracted and printed for direct comparison to the original survey certificates.  The collar coordinates were compared against the signed surveyors’ certificate and the hole length were checked against the geologists’ logs.  Three independent checks of

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this information yielded rare errors. Collar information was also digitally checked against the surveyors’ final compilation file; this check also yielded rare decimal place errors. Erroneous data was corrected and reloaded.  The corrected data was subjected to two additional independent checks.

14.3.2     Survey Data

Prior to entering the downhole survey data into the database it was visually scanned for obvious errors. After entry, the data was extracted and printed for manual verification relative to the original downhole survey data contained in the files and provided by Comprobe. Data verification of 10% of the downhole survey data was completed for the survey data.  Two independent checks of this data detected no errors.  Visual inspections of the plotted hole traces also failed to identify any obvious errors or miss-plots.

14.3.3     Lithology, Alteration, Oxidation, Vein and Mineralization Data

Lithology, Alteration, Oxidation, Vein Description and Mineralization data collected during the 2002-2003 field season was verified against the original logs.  A 10% check of the database yielded an unacceptable amount of errors and led to a 100% check.  The 100% check yielded an error rate of 3.5%, with most of the errors related to the recording of alteration. These errors were corrected.

14.3.4     Analytical Data

A random 10% of the analytical data was verified against the original assay certificates.  Only rare errors were detected, usually pertaining to how results below detection were entered in the database.   Erroneous data was corrected and re-entered.

14.3.5     Geotechnical Data

Ten percent of the geotechnical data collected during the 2002-2003 field program was verified against the original geotechnical logs yielding a significant error rate. As a result, a 100% verification process was initiated. Significant error rates were noted in the fractures-per-meter field.   Error rates in the structural measurement and RQD fields were related to format errors rather than data entry errors. These errors were corrected.

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The QA/QC Database Manager has certified that the database as transferred to Kinross Technical Services on CD-R media is free of gross errors and omissions and is suitable for estimating resource and reserve estimates.

14.4        Data Entry, Storage, Management and Security

The QA/QC Database Manager was responsible for updating and managing the Project Database (PDB). Only the QA/QC Database Manger could access, update or otherwise modify the PDB. Regular backups were made to CD-R disks over the duration of the program. The final PDB was transferred to Kinross Technical Services to be used in resource modeling.

14.4.1     Data Entry

Collar Coordinates

The mine survey department entered collar coordinates, drill azimuth and inclination into spreadsheets that were provided to the QA/QC Database Manager. The QA/QC Database Manager checked the spreadsheet data against original survey field notes prior to importing the data into the PDB.   Cross sections and drill hole composite plans were plotted using Gemcom software and were inspected by the Senior Geologist for collar and directional errors. Final hole depths were verified against the drill log hole depth.

Downhole Survey Data

Comprobe provided the downhole survey data in spreadsheet format to CMM’s Office Manager who checked the data for obvious errors. Once satisfied that the downhole survey data was correct, the Office Manager provided the spreadsheet files to the QA/QC Database Manager for import into the PDB.

Lithology, Alteration, Mineralization and Oxidation

The Office Manager and Data Entry Clerks entered data from the field lithology logs into spreadsheets, one for each hole.  A double-blind entry system was not used in recording this data. A 100% manual check of this data was performed at the end of the drill program.

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Assay Data

Assay results were received as digital files, identified by work order number and hole number. The assay files included individual sample numbers allowing the samples to be matched to the sample intervals. The digital files were imported into spreadsheet files that were verified back to the original assay files prior to being imported into the PDB.

QA/QC Data

The Senior Geologist matched assays to sample numbers on the original log and extracted the QA/QC data from the sample stream into an excel spreadsheet which was used to monitor QA/QC data independently from the sample stream. The QA/QC data was sorted into worksheets for blanks, standards and duplicates.  The entire file was cross-checked between the Project Manager’s copy and the Senior Geologist’s copy. Once satisfied that the data was free of error it was sent via email to Dr. Smee for final analysis.

14.4.2     Data Storage

The PDB was established and maintained using Gemcom modeling software. The PDB was backed up on CD-R media on a regular basis for the duration of the field program. Once the PDB was fully updated with all the results collected during the 2002-2003 programs, the data was copied and forwarded to Kinross Technical Services in CD-R format. This database is referred to as the Final Database (FDB).

In addition to the digital files, all available hard copy data was organized and inventoried on a per hole basis. This information was copied in triplicate with one copy forwarded to Kinross Technical Services, one copy forwarded to Bema and a final copy stored at the Maricunga Mine. Original hard copy data was packaged and stored at the CMM office in Copiapó.

14.4.3             Data Management and Security

Only the QA/QC Database Manager had the necessary file access permissions to update the PDB. Only Kinross’ Manager of Technical Services could access and modify the FDB.

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15.0        ADJACENT PROPERTIES

The Maricunga Project is located within the Maricunga Gold Belt of northern Chile.  Since 1980, a total of 40 million ounces of gold have been defined in the gold belt, (Muntean and Einaudi, 2000).

The Maricunga Gold Belt hosts numerous mineral deposits of economic interest including:

·La Coipa, an epithermal gold-silver mine owned and operated by Kinross

·Marte-Lobo, a gold porphyry deposit that was partially developed by Teck Corporation and,

·Cerro Casale, a large (1.035 billion tonnes), undeveloped, porphyry copper-gold deposit that is a joint venture between Barrick Gold (51%) and Kinross (49%).

No information from adjacent properties has been used in the exploration program or in the estimation of the mineral resource.

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16.0        MINERAL PROCESSING AND METALLURGICAL TESTING

The 2002-2003 Exploration Program included the following metallurgical and physical property test work for both Verde and Pancho:

·cyanide soluble gold analyses

·cyanide soluble copper analyses

·direct agitated cyanidation tests (bottle roll)

·column leach tests

·Bond work index tests

·specific gravity determinations

·Acid Based Accounting (ABA) tests

The 2006 drill program on Pancho also provided representative material to perform additional metallurgical testing. Cyanide soluble gold and copper analyses were systematically completed for all samples.  Recent test work included bottle roll and 14 column leach tests. The results based on the 2006 data collection are summarized separately in this section.

Finally, during 2007 the SART process was considered for the Pancho ore as uneconomic concentrations of copper are present in ore. Some of the copper minerals are soluble in cyanide solutions and these minerals consume cyanide reagent and contaminate the leach solution with copper metal.  A feasibility study and pilot testing program were completed in 2007. The study considers plant design, performance metrics, operating and capital costs and implementation strategy. The reserves defined at Pancho consider the application of the SART process.

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16.1        Metallurgical Test Work 2003

16.1.1     Cyanide Soluble Assays

Both gold and copper cyanide soluble analyses were routinely analyzed for Verde and Pancho. Results were used to guide selection of samples for bottle roll and column leach tests. Cyanide copper analyses were also used to adjust process operating costs during pit optimization of the resource model for Pancho. The mechanics of the Process Cost Adjustment Factor (PCAF) are discussed elsewhere in this report.

16.1.2     Direct Agitated Cyanidation (Bottle Roll) Tests

Bottle roll composite samples were selected from the half core library based on lithology, oxidation state, grade and spatial location. A total of 43 bottle roll samples were selected from the Verde with an additional 15 samples collected from Pancho. The samples were shipped to McClelland Laboratories, Nevada for testing.

The gold grade selection criteria used for selection of the bottle roll tests were:

·	Low grade	0.50 to 0.75 g/t
·	Mid-grade	0.75 to 1.00 g/t and
·	High grade	>1.0 g/t

Oxidation state was determined based on the visual logging of oxidation noted in the geological logs. Three distinct oxidation levels were defined.

·	Oxide	>90% oxidation of sulphide
·	Mixed	between 10% and 90% oxidation of sulphides and
·	Sulphide	<10% oxidation of sulphides.

Samples were selected from the four major ore bearing lithologies, post mineral intrusives and Laguna tuffs were not sampled.

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Sample selection was restricted within an optimized pit shell generated at a US $400 gold price.  Table 16-1 summarizes the Bottle Roll sample selection.

Results of the bottle roll tests show distinct differences in recovery based on the level of oxidation. Oxide ores exhibit higher recoveries than mixed ores which, in turn, demonstrate higher recoveries than the sulphide ores. These findings agreed with the historical metallurgical test work and experience gained from mining in the Verde area.  A summary of the bottle roll test results is presented in Table 16-2.

Table 16-1: Summary of Bottle Roll Sample Selection

		Gold Grade
		Low	Medium	High
Deposit		(0.50 - 0.75)	(0.75-1.00)	(>1.00 g/t)
Pancho	# of samples	5	5	5
Verde W	# of samples	5	5	11
Verde E	# of samples	9	5	10

		Oxidation
		Oxide	Mixed	Sulphide
Deposit		(>90%)	(10-90%)	(<10%)
Pancho	# of samples	6	3	6
Verde W	# of samples	8	5	8
Verde E	# of samples	9	6	9

		Lithology
				Mineralized
				Post Mineral	Dacite
Deposit		Diorite	Breccia	Intrusive	Porphyry
Pancho	# of samples	6	9	—	—
Verde W	# of samples	—	7	5	9
Verde E	# of samples	—	9	6	9

Table 16-2: Summary of Bottle Roll Results

	Oxidation	Au Recovery
Deposit	Class	(%)
Verde	Oxide	72.3
	Mixed	63.8
	Sulphide	53.0
Pancho	Oxide	83.0
	Mixed	72.8
	Sulphide	68.6

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16.1.3Column Leach Tests

Results of the bottle roll tests were used to target specific areas of the deposit with the HQ diameter drill holes that were used to collect the sample material necessary for column leach analysis. Composite samples were collected from whole HQ diameter core that was drilled specifically for the metallurgical test program. A total of 10 HQ diameter holes (1,500 meters) of core were drilled at Verde for the purpose of column leach tests.  An additional seven (7) samples were collected at Pancho but unlike the Verde samples, the Pancho samples were taken from remaining half cores available at the time of sample selection.

For the Verde samples, column tests were established for three different product sizes (6.3, 9.5 and 15.8 mm). The same sample selection matrix established for the bottle roll tests guided column leach sample selection. Grade of the HQ diameter core was confirmed by cutting a sliver sample for each two-meter interval.  Table 16-3 summarizes the Column Leach Sample program.

Table 16-3: Summary of Column Leach Sample Selection

		Gold Grade
		Low	Medium	High
Deposit		(0.50 - 0.75 g/t)	(0.75 – 1.00 g/t)	(>1.00 g/t)
Pancho	# of samples	2	2	3
Verde West	# of samples	2	2	1
Verde East	# of samples	2	2	2

		Oxidation
		Oxide	Mixed	Sulphide
Deposit		(>90%)	(10-90%)	(<10%)
Pancho	# of samples	3	1	3
Verde West	# of samples	—	—	5
Verde East	# of samples	3	3	—

Results of the column leach tests matched the bottle roll results. Table 16-4 summarizes the results of the column leach tests completed as part of this study.

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Figures 16-1 and 16-2 are plan views of the Verde and Pancho pits showing the original metallurgical sample distribution within the 2002-2003 design pit shells.

Table 16-4: Summary of Column Leach Test Results

Deposit	Ore Type	Crush Size	Au Recovery
		(k80 mm)	(%)
West		15.8	44.0
Verde	Sulphide	9.5	53.7
		6.3	57.2
East		15.8	70.5
Verde	Mixed	9.5	73.6
		6.3	79.3
East
Verde	Oxide	9.5	77.0
Pancho	Sulphide	9.5	71.6
Pancho	Mixed	9.5	79.8
Pancho	Oxide	9.5	89.0

Figure 16-1: Metallurgical Sample Locations – Verde Pit – 4300 Level

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Figure 16-2: Metallurgical Sample Locations – Pancho Pit – 4300 Level

16.1.4Bond Work Index

Bond Work Index (BWI) samples were collected from the 10, HQ diameter core holes drilled at Verde. Bond work index tests required selection of whole HQ core pieces of approximate length of 0.15 – 0.20 meters.  For Verde, such pieces were selected every 4 meters down-hole.  All ten Verde HQ diameter holes were sampled in this manner, a total of 296 samples. At Pancho, BWI samples were collected from the standard diameter drill holes. Whole core pieces were taken approximately every 20 meters down hole.  A total of 67 samples were selected. All samples were shipped to Hazen Research, Colorado, for analysis.

Results of the BWI tests indicated limited variability in the rock mass therefore; the BWI data was not modeled when completing the Verde and Pancho resource models.

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16.1.5Specific Gravity

As part of the BWI tests above, Hazen completed specific gravity analyses on all the BWI samples submitted. In addition to this data, 1171 samples of half core were selected from the half core library. Samples, roughly 15 cm in length, were selected approximately every 30 meters downhole throughout the entire campaign. The samples were shipped to ACME laboratories, Santiago where they were analysed using the wax dipped immersion method.  A summary of the specific gravity results is presented in Table 16-5.

Table 16-5: Specific Gravity Results

Deposit	Unit	SG
Verde	Oxide	2.45
	Mixed	2.52
	Sulphide	2.57
	Laguna Tuff	2.00
	Post Mineral Intrusive	2.60
	Waste Dump Material	2.00
Pancho	Oxide	2.27
	Mixed	2.38
	Sulphide	2.49

Both Pancho and Verde exhibit trends of increasing Specific Gravity with depth, leading to the decision to model the deposit Specific Gravity by the oxidation surface rather than by rock type.

16.1.6Acid Rock Drainage Study

At the request of the CMM environmental department (Jorge Herrera), ALS Chemex La Serena sent 71 pulp samples to ALS environmental laboratories Santiago for acid based accounting testwork. A subset of 40 of these was proposed for testwork.   The forty samples selected were from widely spaced locations in the anticipated pit designs for both Verdes and Pancho.  Samples included both ore and waste in major rock types in all redox zones.

In 2007, follow up work included a sample collection program comprising 150 samples. To date partial results are available for 140 samples. Representative samples will be selected for further wet cell analysis work once the dataset is complete.

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16.2Metallurgical Test Work 2006

16.2.1Introduction

Gold recovery rates for the Verde and Pancho deposits have been estimated from test work conducted at McClelland Laboratories in Reno, Nevada.  In 2003, a total of 9 column leach tests were done on 7 Pancho composite samples and 11 column leach tests were done on 11 Verde composite samples.  In 2006, an exploration drill program on the Pancho property provided samples for an additional 14 column leach tests at McClelland.  Final test data was available by August 2007 and these gold recoveries have been integrated in the Pancho reserve estimate.

16.2.2Comments on Historical Metallurgical Composite Selection Criteria

The early work on Maricunga metallurgical type classification established that the most predominant metallurgical criteria were grade, oxide state and crush size.  The 1998 metallurgical column testing relied on the classification of four groups of sample types largely predicated by the amount of sheeted and stockwork veining as well as silicification. The four groups are outlined below.

·Group 1 – 81.4% recovery – Strong sheeted veining, moderate strong stockwork, and weak matrix silicification.

·Group 2 – 72.9% recovery – Moderate-strong sheeted veining, moderate stockwork, weak matrix silicification.

·Group 3 – 66.1% recovery – Weak-moderate sheeted veining, weak stockwork, and weak to moderate matrix silicification.

·Group 4 – 50.4% recovery – Weak-nil sheeted veining, weak-nil stockwork, and moderate strong silicification.

These groupings were specific to Verde, where East Verde was predominantly represented by Group 2, while West Verde was mostly comprised of Group 3.

In 2003, the methodology for metallurgical sample selection was changed.  Of concern was the issue that the head grades of the group 1 to 4 samples were 20 to 40% above average grade of the deposit and it was suspected that higher grade column samples would result in higher gold recoveries.  In addition, field geologists had concerns with the classification of

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silicification and the ability to differentiate between stockwork versus sheeted veining in chips from the predominantly RC based database.

As a result of the assessment of the Maricunga database and regression results and historic results, it was proposed to classify metallurgical column test samples, firstly by oxidation intensity, then by grade, and by elevation where appropriate. Due to questions of database suitability and analysis, a lower priority was given to veining, silicification, argillic alteration and lithology.

In 2003, before the final composites were selected for column tests, McClelland Laboratories in Reno completed bottle roll tests on 14 Pancho “point samples” selected to represent the likely metallurgical domains. Samples were differentiated by grade, lithology and oxidation state and subjected to a 96 hour bottle roll leach at a crush size of 80% passing 1.7 mm.  The results showed that the oxidation state is the only reliable predictor of heap leach gold recovery.  Head grade and lithology do not significantly affect gold recovery.  This correlation was borne out in the column tests and the current Pancho mine plan uses gold recovery values based only on oxide, mixed and sulphide ore designations.

16.2.3Metallurgical Composite Selection Criteria (2006)

In 2006, an extensive drill campaign was completed on the Pancho deposit and drill core composite samples were shipped to McClelland Laboratories in Reno for metallurgical test work. The first stage of test work was to complete seven column tests on Pancho drill core composite samples crushed to 80% passing 6.3 mm.  The objective was to test the variability of gold recoveries from ore in zones of the Pancho pit that had not been tested previously.  Composites were selected from single holes where possible to maintain the spatial integrity of the sample. The McClelland columns were 6” diameter x 10’ tall and held 68 kg of rock. The column tests were run at a crush size of 80% passing 6.3 mm in order to be compatible with the previous test work completed on the Verde deposit.

The second stage of test work was to select a typical sulphide sample and a typical oxide sample from a multi-hole composite and conduct a crush size variability study to determine the effect of particle size on gold recovery.  Crush sizes of 6.3 mm, 10mm and 16 mm were selected. The oxide column leach tests have only just begun and no data is available yet. In

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addition to the column test work, bottle roll cyanide leach tests were conducted on multiple point samples to determine the variability of gold recovery and copper recovery.

The metallurgical composite selection criteria was to obtain lithotypes with representative oxidation and hypogene alteration combined with representative Au and Cu grades in spatially well distributed areas within the original Pancho $450/Au oz pit.  The sample summary table below present the details for 2006 Pancho metallurgical composites (see Table 16-7).

Table 16-6: Pancho Metallurgical Composites 2006

Column	Sample		From	To	Au	Cu	%
Test No.	Ref. No.	Hole	m	m	g/t	ppm	Oxidized	Lithology	Hypo Alt str-wk	% py	% veins
P1	1	P2-1930	350	474	0.96	239	0	Porphyry	chl ser pot	1-5	0.5 - 1
P2	2	P2-2240	180	288	0.5	548	5	Breccia	pot,chl.ser	0.5-2	0.5 - 1
P3	3	P9.5-1690	40	104	0.71	698	5	Porphyry	chl ser pot	1-2	0.5
P4	4	P7-2140	80	160	0.62	1366	10	Porph./Breccia	clay,ser,chl	1-3	0.5 - 1
P5	5	P9.5-1840	80	150	0.58	755	2	Porphyry	pot, chl	0.5	0.5 - 1
P7	6a	P7-2140	10	50	0.57	103	100	Breccia	arg,jar	0.5	1-3
P7	6b	P10.5-2090	10	44	0.64	50	100	Porphyry	arg,jar,pot	0.5	0.5
P6	6	P9.5-1840	3	70	0.82	1166	50	Porphyry	chl pot arg	1-1.5	0.5-1
P8	8	P8-2175	140	248	0.63	1322	3	Fine Gr. Porph.	ser chl	0.5-5	0.5 - 2
P9	9	P4.5-1730	380	490	0.72	1041	0	Porphyry	chl sil	0.5	0.5-1
P10	10	P5.5-2000	290	400	0.63	504	0	Porphyry	pot chl ser	0.5	0.5
P11	11	P7-2020	300	390	0.59	1264	100	Porphyry	chl ser pot	0.2	1-2
P12	12	P5.5-2100	400	500	0.57	738	100	Porphyry	chl sil	0.5	0.5-1.5
P13	13	P3.2-1770	350	450	0.52	256	100	Porphyry	chl ser pot	0.2-1	0.2
P14	14	P4.5-1990	300	400	0.79	455	0	Porphyry	qtz-ser-chl	0.2-1	0.2

16.2.4McClelland Test Work Summary

A total of 13 drill core composites and 43 drill cuttings composites were prepared for heap leach cyanidation testing. Another composite was collected at a later date and was selected for its high copper content. Detailed head analyses and direct agitated cyanidation (bottle roll) tests were conducted on all of the composites. The cuttings composites were evaluated at a nominal 1.7 mm (as-received) feed size. The drill core composites were evaluated at 80% - -1.7mm and 80% -75mm feed sizes. Select (3) drill core composites were also evaluated at an 80%-6.3mm feed size. A column percolation leach test was conducted on 10 of the 13 drill core composites, at an 80% -6.3mm feed size. Select (2) drill core composites

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were also subjected to column leach testing at 80% -15.9mm and 9.5mm feed sizes. Metallurgical results from this testing program are summarized below.

The composites evaluated were representative of the various ore types from the Pancho deposit. Head analysis results showed that the cuttings composites contained between 0.35 and 2.54 gAu/mt ore. The drill core composites contained between 0.44 and 0.95 gAu/mt ore. Total copper content of all of the composites ranged from <0.01 to 0.2% Cu. On average, acid soluble copper and cyanide soluble copper represented 17% and 63%, respectively, of the total contained copper.

The Pancho drill cuttings and core composites generally were amenable to direct agitated cyanidation treatment at an 80%-1.7mm feed size. Gold recoveries from the cuttings composites ranged from 43.5% to 97.4%, and averaged 73.0%. Gold recoveries of less than 60% were obtained from only 4 of the 43 cuttings composites. Gold recoveries obtained from the drill core composites at an 80%-1.7mm feed size ranged from 52.9% to 88.6%, and averaged 74.2%.

The 2140 composite was the only drill core composite exhibiting a gold recovery of less than 60%. Grinding of the drill core composites to 80% -75mm improved the average gold recovery to 83.2%. Bottle roll test cyanide consumptions varied substantially, and tended to increase with increasing copper head grade. Cyanide consumption for the 1.7mm feeds ranged from 0.11 to 3.67 kg NaCN/mt ore, and averaged 1.00 kg NaCN/mt ore. Lime requirements varied substantially, and were very high (>10 kg/mt ore) for seven of the composites tested. Overall, 1.7mm feed bottle test lime requirements ranged from 2.2 to 61.9 kg/mt ore, and averaged 7.3 kg/mt ore.

The 10 drill core composites subjected to column leach testing were amenable to simulated heap leach cyanidation treatment, at the feed sizes evaluated. All 10 composites were evaluated at an 80% -6.3mm feed size. Gold recoveries obtained from the 6.3mm feeds ranged from 55.1% to 89.1%, and averaged 72.0%, in from 71 to 174 days of leaching. Gold recoveries of lower than 70% were obtained from the 1840 composite (62.3%), the 2140 composite (56.5%) and the Copper Column Test (CCT) composite (55.1%).

The Maricunga-High Elevation Variable Crush (RHEVC) and Oxide core composites were both also subjected to column testing at 80% -15.9mm and 80% -9.5mm feed sizes, to

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evaluate crush size sensitivity of the Pancho ore. Gold recovery increased incrementally (3% - 6%), by crushing from 15.9mm to 9.5mm in size. Gold recoveries obtained from both composites were essentially the same at the 9.5mm and 6.3mm feed sizes.

Column test gold recovery rates for the 2240, Pancho 6, CCT and Oxide composites were rapid, and gold extraction was substantially complete in about 25 days of leaching. Gold recovery rates were slower, and gold extraction progressed at a slow, but significant rate after 75 days of leaching, for the 1690, 1840, 1930 and 2140 composites. The slowest gold recovery rates observed were for the 1840 Shallow and RHEVC composites.

Column test cyanide consumptions generally were high, but were not as strongly correlated to copper head grade, as seen during bottle roll testing. Column test cyanide consumptions tended to increase with increasing leach solution application, as normally expected for column leach tests. Because some of the column test leach cycles were relatively long in duration (>140 days), the resulting cyanide consumptions were relatively high. Average column test cyanide consumption (6.3mm feeds) was 2.39 kg NaCN/mt ore.

Column test lime additions were made based on corresponding bottle roll test lime requirement data. Lime requirements for the drill core composites varied substantially. Initial column test lime additions ranged from 1.8 to 13.0 kg/mt ore. Those lime additions were sufficient for maintaining protective alkalinity during leaching, for most of the composites. Minor pH control problems were encountered during leaching of the 2140 and 2240 composites, and relatively small lime additions were made to the column test barren solutions during leaching, to help control leaching pH. Severe pH control problems were encountered during leaching of the Oxide and CCT composites, and the lime added during leaching for pH control was equivalent to more than the lime added to the ore initially. The highest observed combined lime requirement was 41.1 kg/mt ore, for the Oxide composite 9.5mm column feed. Higher initial lime additions should be evaluated for these ore types to determine whether or not pH control will be a problem during commercial heap leaching.

Overall, reagent consumption data indicate potential concerns during commercial heap leaching of the Pancho ore, with respect to high cyanide consumptions related to copper dissolution, and high lime consumptions related possibly to interaction with sulphates naturally occurring in the ore. Follow-up testing was recommended to better establish base requirements for pH control during heap leaching of the higher lime consuming Pancho ores.

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Table 16-7: Pancho and Verde Gold Recovery Estimates

Gold Recovery (%)
Deposit	Zone	2003 Column Test Data	SNC Study	2004  $350/oz Pit Design	2005 $400/oz Pit Design	2006 $475/oz Pit Design	2006-2007 Final Column Tests
	Oxide	77	73	67	73	73
Verde	Mixed	74	70	67	70	70
	Sulphide	54	53	48	53	53
	Oxide	89	80	80	85	85	86
Pancho	Mixed	80	75	75	75	75	77
	Sulphide	72	70	70	70	70	70

Pancho Column Test Work Results

Prior to conducting the column leach tests, bottle roll cyanide leach tests were conducted on the samples to confirm metallurgical response and estimate reagent consumptions.  The data is summarized in Table 16-8.  Gold recoveries and reagent consumption rates were similar to those established in the 2003 testwork. It is clear that cyanide consumption increases with copper content and lime consumption increases with oxidation state.  Gold recovery also increases with oxidation state.  It is also clear that gold recovery increases with a finer grind confirming that recovery is sensitive to crush size.

The column leach test data is summarized in Table 16-9 and charts of gold and copper extraction versus time are appended.  The data compiled to date indicates that gold recovery from the Pancho samples tested in 2006-2007 is in the region of 69% for the sulphide ore and 86% for the oxide ore.  This test work (Table 16-9) plus the work completed in 2003 confirms that the gold recovery values used in the 2007 reserve estimate are reasonable (68% for sulphide, 75% for mixed and 85% for oxide).

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Table 16-8: Pancho Bottle Roll Summary 2006

			1.7 mm crush, 96 hr bottle roll
			Calculated	Gold	Calculated	Copper	Cyanide	Lime
Sample	%		Head	Recovery	Head	Recovery	Consumption	Consumption
Number	Oxidized	Lithology	Au g/t	%	Cu ppm	%	kg/mt ore	kg/mt ore
P1	0	Porph.	1.04	70	248	19	0.45	2.2
P2	5	Bx.	0.47	79	429	53	1.13	6.1
P3	5	Porph.	0.70	70	631	37	0.98	3.7
P4	10	Porph./Bx.	0.70	53	1,225	51	2.20	2.6
P5	2	Porph.	0.55	62	769	22	0.92	2.4
P6	50	Porph.	0.80	73	1,076	54	1.81	7.1
P7	100	Porph./Bx.	0.64	88	111	10	0.18	7.0
P8	0	Porph./Bx.	0.73	64	765	22	0.83	3.2

			75 micron grind, 72 hr bottle roll
			Calculated	Gold	Calculated	Copper	Cyanide	Lime
Sample	%		Head	Recovery	Head	Recovery	Consumption	Consumption
Number	Oxidized	Lithology	Au g/t	%	Cu ppm	%	kg/mt ore	kg/mt ore
P1	0	Porph.	1.04	87	364	18	0.37	2.0
P2	5	Bx.	0.54	80	438	54	1.13	4.9
P3	5	Porph.	0.67	84	675	41	1.46	2.9
P4	10	Porph./Bx.	0.67	64	1,102	46	1.96	2.3
P5	2	Porph.	0.59	83	752	20	0.67	2.4
P6	50	Porph.	0.81	83	1,064	62	2.09	6.2
P7	100	Porph./Bx.	0.65	89	116	14	0.37	5.8
P8	0	Porph./Bx.	0.74	84	742	19	0.61	2.7

Table 16-9: Pancho Column Test Summary 2007 (Part 1)

Column Test No.	Composite	Oxidized	Lithological Type	Feed Size P80 (mm)	Leach/Rinse Time, days	Au Rec. %
P1	1930	0	Porphyry Sulphide	6.3	169	78.8
P2	2240	5	Breccia Sulphide	6.3	108	71.1
P3	1690	5	Porphyry Sulphide	6.3	223	77.6
P4	2140	10	Porphyry Sulphide	6.3	193	56.5
P5	P95-1840	2	Porphyry Sulphide	6.3	175	62.3
P6	Shallow Comp P95-1840	50	Mixed Sulphide	6.3	210	76.6
P7	Pancho 6a+6b Comp	100	Oxide	6.3	84	89.1
P8	RHEVC	0	Porphyry Sulphide	15.9	167	61.3
P9	RHEVC	0	Porphyry Sulphide	9.5	167	67.2
P10	RHEVC	0	Porphyry Sulphide	6.3	167	68.8
P11	Oxide	100	Oxide	15.9	81	82.1
P12	Oxide	100	Oxide	9.5	81	84.8
P13	Oxide	100	Oxide	6.3	81	84.1
P14	Copper Column Test	5	Porphyry Sulphide	6.3	81	55.1

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Table 16-10: Pancho Column Test Summary 2007 (Part 2)

Column							NaCN
Test		gAu/mt ore			gCu/mt ore		Consumed,
No.	Composite	Extracted	Tail Screen	Avg. Head	Extracted	Head Assay	kg/mt ore
P1	1930	0.78	0.21	0.99	60	300	2.03
P2	2240	0.32	0.13	0.47	173	600	2.49
P3	1690	0.52	0.15	0.7	173	800	2.82
P4	2140	0.39	0.3	0.69	539	1400	3.71
P5	P95-1840	0.33	0.2	0.5	104	1200	2.12
P6	Shallow Comp P95-1840	0.59	0.18	0.79	600	770	3.46
P7	Pancho 6a+6b Comp	0.57	0.07	0.62	27	100	1.18
P8	RHEVC	0.38	0.24	0.68	162	800	2.19
P9	RHEVC	0.43	0.21	0.68	178	800	2.55
P10	RHEVC	0.44	0.2	0.68	144	800	2.37
P11	Oxide	0.55	0.12	0.67	22	200	2.17
P12	Oxide	0.56	0.1	0.67	38	200	2.23
P13	Oxide	0.53	0.1	0.67	16	200	2.16
P14	Copper Column Test	0.38	0.31	0.75	133	1300	1.52

16.2.5     SART Process Feasibility Study

Uneconomic concentrations of copper are present in the heap leach ore at the Maricunga mine. Some of the copper minerals are soluble in cyanide solutions and these minerals consume cyanide reagent and contaminate the leach solution with copper metal.  Unless removed, this copper will decrease the purity of the dore bars and incur penalty refining costs. The copper grade in the leach solution has risen significantly from initial levels of below 50 ppm in 2005 to over 300 ppm in 2007. This copper concentration is expected to increase further as ores from the Pancho deposit are processed. The rising copper level in the Maricunga heap leach solution will need to be addressed to avoid increased lock-up of cyanide by the copper resulting in increased operating costs.

Over the last two years, Kinross has undertaken significant metallurgical testwork and engineering studies to establish a robust process for removing copper from the circuit and recycling the cyanide.  In 2005, SGS Lakefield completed preliminary tests confirming that SART technology had the best potential.  The SART process was developed in 1997 by SGS Lakefield Research and Teck Corporation and is currently in full-scale commercial application at Newcrest’s Telfer Mine in Australia.

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The SART name is derived from an acronym describing the unit operations in the process, as follows:

·Sulphidization

·Acidification

·Recycling of precipitate/s

·Thickening of precipitate/s

The SART process flowsheet is shown schematically below in Figure 16-3.

Figure 16-3: SART Plant Flowsheet

SART, as originally developed, involves adding chemical sulphide ions, such as sodium hydro-sulphide (NaSH), to an acidified CN solution (pH 10 to pH 5 or below) to precipitate copper, silver and zinc (if present) as metal sulphides and convert cyanide to HCN, under weakly acidic conditions. Other CN solubilized metals like Ni and Co would also be precipitated. HCN is a very soluble gas, and remains in solution until the Cu₂S, Ag₂S or ZnS

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is removed by thickening and filtration. The HCN solution is then neutralized with lime or caustic soda, prior to recycling to leach.

The overall chemical reactions applied to Cu, Zn and Ag cyanide complexes are:

2Cu(CN)₃^2-  +  S^2-  +  6H⁺   à   Cu₂S  +  6HCN

2Ag(CN)₂^-   +  S^2-  +  4H⁺  à   Ag₂S  +  4HCN

Zn(CN)₄^2-   +  S^2-  +  4H⁺  à   ZnS  +  4HCN

The above sulphidization reactions occur rapidly in a first stage (primary) precipitation reactor from which sulphide precipitates are fed to a thickener. Recycle of thickened slurry of sulphide precipitates back to the reactor is done until a suitable precipitate particle size develops and sufficient bed depth in the thickener is achieved. A precipitate slurry of 10 – 15% solids may then be drawn off the thickener underflow and washed, filtered and dried. The Cu₂S precipitate (with Zn, Ni and other metal sulphides, if present) is then sold to a smelter as a valuable co-product.

A second reactor is used to neutralize the aqueous HCN in the thickener overflow from pH 5 back to pH 10. Lime or caustic soda may be used to produce CaCN₂ or NaCN for recycle to the heap or tank leaching circuit. The neutralization reactions are:

2HCN  +  Ca(OH) ₂  à  CaCN₂  +  H₂O

2HCN  +  NaOH      à  NaCN  +  H₂O

All three process units (2 reactors and a thickener) are sealed and have vents to a lime scrubber system to capture any evolved HCN or H₂S gases. By acidifying with sulphuric acid in the primary reactor, the SART process leads to formation of a gypsum type precipitate on neutralization of HCN in the neutralization reactor. This gypsum precipitate needs to be removed from the neutralized CN solution prior to recovery of gold on carbon and recycle of barren CN solution to the leach circuit. To achieve this, a gypsum thickener is included in the circuit with the option of recycling the gypsum thickener underflow to enhance precipitation rates and settling / consolidation in the thickener.

The SART process may be applied to pregnant gold solutions prior to carbon adsorption or on barren solutions after carbon adsorption. For a heap leach operation the inherent leach

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solution / ore solids separation suits the SART process as clear solutions are fed directly to the primary precipitation reactor. For mill plants the leach slurry may be washed via CCD with the overflow pregnant solution feeding a SART plant prior to carbon adsorption. When the leach slurry feeds a CIP plant, the CIP tails (barren) may be washed in a CCD circuit and then fed to a SART plant.

In summary, the process involves lowering the solution pH to release the cyanide associated with the copper cyanide complex, allowing it to be recycled back to the leach process as free cyanide.  Sulphide reagent is added to the acidic solution to precipitate the copper as a solid high-grade copper sulphide precipitate which can be filtered, dried, bagged and shipped to a smelter to generate copper revenue.  Lime has to be added back into the solution to neutralize the acid and gypsum (calcium sulphate) will be formed as by-product.  This gypsum will be stored in lined ponds.

The location of the SART plant was chosen to be near the existing ADR plant at the base of the heaps in order to maximize the use of existing roads and services.  A SART plant of 750 m3/hr capacity was determined as optimal to treat a bleed stream from the leach solution (2,150 m3/hr) from the heaps.  In this way, the capital outlay for the plant is minimized and the SART plant can be taken off line without affecting gold production. A copper sulphide precipitate (Cu₂S at about 65% Cu) will be produced at a rate of approximately 5 to 10 tonnes per day and is planned to be sold to the Enami smelter near Copiapo. Figure 16-4 shows the Marigunga modified flowsheet incorporating the SART process.

A process to apply for environmental permitting has begun with preparation of DIA documents leading to a construction permit. This feasibility study has mapped out the DIA application process steps and contains the process technical information that will be required for the application documentation. The installation of a SART plant at Maricunga will involve some changes to the existing operation and a DIA permit will be required. Main changes to the Maricunga operation are: supply of sulphuric acid (about 15 t/day) and NaSH chemicals to the mine, dispatch of copper sulphide precipitate (about 5 t/day) from the mine and disposal of a gypsum product on site in a lined facility at the mine (about 20 t/day). A new EIS for Maricunga will not be required for the plant. The estimated timing for the permitting process following submission of the DIA is expected to be about 8 months and will not be on the project critical path.

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Figure 16-4: Modified Maricunga Plant Flowsheet with SART

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17.0        MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE

17.1        Introduction

The Mineral Resource and Reserve Estimates for the Maricunga Project have been prepared under the direct supervision of Maryse Bélanger, P. Geo., Director Technical Services, Kinross Americas.

During the course of 2007, a full review of the geologic interpretation and the spatial controls of mineralization was undertaken for both Verde and Pancho resulting in new resource models to support updated Mineral Resource and Mineral Reserve estimates.

17.2        Interpretation and Grade Domains

17.2.1     Verde

The gold grade model for Verde used to rely on a lithological interpretation completed by the site geologists to discriminate between mineralized and barren lithologic units. Mining experience at Verde has shown that the mineralization tends to cross over all lithologies.

The Verde model update prepared in 2007 incorporates a new interpretation based on a better understanding of the geology and the mineralization and its spatial distribution. Essentially, mining has shown that the gold content of the material mined is often related to its position relative to the known post-mineral intrusive bodies at Verde East and Verde West.

The model update is based on the new interpretation for Verde East and Verde West.  The concept and resulting geometry are the same for both open-pit areas.  The mineralization and its inner limits are defined within a zone of high-grade material centred about a post-mineral intrusive. The proximity to the post-mineral intrusive is what affects most the overall grade distribution.

 The halo surrounding the higher-grade core is slightly more diffuse and incorporates mineralization of lesser tenor.  The overall concept also concurs with observations in the mining areas where the material closer to the PMI displays stronger brecciation and more

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veinlets often resulting in grades exceeding 1 g/t Au.  This compares to grades of 0.5 g/t to 0.6 g/t for the rocks located at greater distances and then again decreasing as distances from the PMI become greater.  Finally, an external zone of low grade material was modelled broadly based on a cut-off of 0.15 g/t Au.  A typical plan view is presented below in Figure 17 -1.

Figure 17-1: Typical Plan View of Verde Domaining (Bench 4280)

The domains model developed from cross sectional and level plan interpretation, identified distinct zone boundaries that are used as soft boundaries during grade interpolation as per the conclusions drawn for the contact profiles (see Section 17.4.2.1).

The Figure above shows the peripheral zone of approx. 0.15 g/t Au, the low-grade mineralization (in green) defined as the material with a gold grade between 0.3 g/t and 0.6 g/t. The high-grade (in red) encompasses the material greater than 0.6 g/t Au. Throughout this report they are referred to as:

·Domain 1 = gold grade greater than 0.6 g/t

·Domain 2 = gold grade between 0.3 g/t and 0.6 g/t

·Domain 3 = low grade at the periphery based on 0.15 g/t

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The new interpretation incorporates outlines digitized first from sections and then transposed on plans in order to get the best possible representation of the mineralization in three-dimensions. The outlines were than linked to form solids that were validated in Vulcan.  Blast hole plans were also used to guide the interpretation as they provide close-spaced information.

17.2.2     Pancho

The model for Pancho is based on the core samples collected in 2006. Diamond drilling demonstrated that the mineralization crosses over all lithologies and appears centred on a Post-Mineral Intrusive that is in rare instances mineralised. A model similar to Verde was created where three domains were defined.  First, 9 sections oriented N40°W were plotted and interpreted with the outlines digitized in Vulcan. Then, 47 plans at 10 m intervals encompassing elevations from 4440 m to 3700 m elevation were interpreted. The outlines obtained in plan and section views were connected to form the solids defining the grade domains. They are also referred to as Domain 1, Domain 2 and Domain 3 corresponding to grade ranges similar to those used at Verde.

·Domain 1 = gold grade greater than 0.6 g/t

·Domain 2 = gold grade between 0.3 g/t and 0.6 g/t

·Domain 3 = low grade at the periphery based on 0.15 g/t

The northwest quadrant of the Pancho deposit shows significant copper content related to the presence of a small chalcocite blanket close to the surface and hypogene porphyry copper mineralization in the form of chalcopyrite at greater depths. A copper domain was interpreted on sections and plans as per the methodology used in defining the gold grade domains. The copper domain of interest is generally located between 4320m and 3840m elevation. The copper grades shown on plans and sections demonstrate that a 200 ppm cyanide soluble copper grade form a continuous domain in three-dimensions. It was therefore the criteria chosen for defining two domains and coding the block accordingly:

·Domain 0 = outside the copper envelope defined by 200 ppm cyanide soluble copper

·Domain 1 = inside the copper envelope

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17.3Oxidation Model and Recovery

Metallurgical recovery is a key aspect in modelling the Verde and Pancho deposits. Significant effort was directed towards establishing a relationship between metallurgical recovery and the level of oxidation visually estimated by the site geologists. Table 17-1 shows the criteria used in defining the material types encountered in the pit.

Table 17-1: Oxidation Codes

Material	Condition	Code
Oxide	greater than 50% oxidation	Block Code Redox = 3
Mixed	between 10% and 50% oxidation	Block Code Redox = 2
Sulphide	less than10% oxidation	Block Code Redox = 1

The original recovery model was developed by interpretation of cross section, longitudinal section and level plan data by the site geologists. The more recent recovery model for Verde integrates all blast hole and mine geology information collected to date. The interpreted surfaces were used as hard boundaries for assigning recoveries in the blocks during pit optimization. The surfaces are highly irregular and similar to the undulations observed in the pit walls at Verde.

The additional core samples available for Pancho allowed for an update of the surfaces marking the different levels of oxidation, namely oxide, mixed and sulphide. One noticeable difference is that the new interpretation includes less mixed material than indicated by the original interpretation completed in 2003.

17.4Grade Model

17.4.1Statistical and Geostatistical Analyses

Excluded Data

The following data was not used in completing the original resource block model for Verde:

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·blast hole data

·trench data and;

·drill data within the previously mined portion of the Verde pits.

For Pancho some historical data was also excluded. They included trench data and original Anglo American drill holes for which location could not be confirmed. These were part of hole series SHR and SDR.

Unexplained missing data was assigned a zero value for all analyses. Explained missing data was treated as missing data and ignored during processing.

Statistics

Basics statistics were calculated for the original 2 m gold assays. The results for Verde and Pancho gold assays are presented in Table 17-2 and Table 17-3 below.

Table 17-2: Statistics for Uncapped 2 m Gold Assays - Verde

		Mean Au		Coef. of
Zone	Number	g/t	Std. Dev.	Variation	Minimum	Maximum
All	49,789	0.67	0.69	1.03	0	15.94
1	19,153	1.07	0.80	0.75	0	15.94
2	15,274	0.56	0.47	0.84	0	10.2
3	6,908	0.25	0.35	1.39	0	10.63

Table 17-3: Statistics for Uncapped 2 m Gold Assays - Pancho

		Mean Au		Coef. of
Zone	Number	g/t	Std. Dev.	Variation	Minimum	Maximum
All	31,164	0.54	0.66	1.22	0	21.9
1	8,560	0.97	0.86	0.89	0	21.32
2	10,796	0.51	0.37	0.72	0	11.95
3	9,856	0.28	0.53	1.90	0	21.9

It should be noted that the coefficients of variation for gold are low reflecting the disseminated widespread nature of the mineralization.

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17.4.2Grade capping

Capping of gold values affects 0.2% of the total number of assays at Verde. Grade capping was determined based on analysis of the statistics of the three populations created by dividing into the separate grade domains. The levels chosen were 4 g/t Au in the high-grade zone (domain 1) and 2.0 g/t Au in the low-grade zone (domain 2). These capping levels correspond respectively to the 99^th percentile and the 98^th percentile. In the material to the periphery (domain 3) a cap of 1 g/t was selected effectively reducing the grade of approx. 2.0% of the samples

Similar capping levels were selected for the Pancho domains reflecting the similarity in the grade distributions. The levels chosen were also 4 g/t Au in the high-grade zone (domain 1) and 2.0 g/t Au in the low-grade zone (domain 2). These capping levels correspond to the 99^th percentile in both domains. In the material to the periphery (domain 3) a cap of 1 g/t was selected effectively reducing the grade of approx. 2.5% of the samples.

The grade capping was used in combination with restricting the influence of the higher-grade composites. (Please see Sections 17.4.4 and 17.4.5 on Gold Grade Interpolation Verde and Gold Grade Interpolation Pancho for more details).

At Pancho the copper and cyanide soluble copper grades were not capped.

17.4.3Sample Compositing

Capped assays were composited to build the resource models for Verde and Pancho. All grade interpolation was based on 5 m length composites to account for drill hole deviations.

Statistics for the 5 m composites using gold capped assays are presented in Tables 17-4 to 17-8. Statistics for copper and cyanide soluble copper are also presented for Pancho.

Table 17-4: Statistics for 5 m Gold Composites - Verde

		Mean Au		Coef. of
Domain	Number	g/t	Std. Dev.	Variation	Minimum	Maximum
All	20,577	0.66	0.55	0.84	0	4.00
1	7,901	1.05	0.58	0.55	0	4.00
2	15,274	0.56	0.31	0.55	0	10.20
3	6,908	0.25	0.35	1.39	0	10.63

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Table 17-5: Statistics for 5 m Gold Composites – Pancho

		Mean Au		Coef. of
Domain	Number	g/t	Std. Dev.	Variation	Minimum	Maximum
All	12,578	0.52	0.45	0.87	0	4.00
1	3,477	0.94	0.53	0.56	0	4.00
2	4,330	0.51	0.27	0.52	0	2.00
3	4,018	0.27	0.26	0.97	0	1.78

Table 17-6: Statistics for 5 m Copper Composites – Pancho

		Mean Cu		Coef. of
Domain	Number	ppm	Std. Dev.	Variation	Minimum	Maximum
Outside	9,314	303	303	1	0	4,254
Inside	2,592	1,048	1048	1	3	7,820

Table 17-7: Statistics for 5 m Cyanide Soluble Copper Composites – Pancho

		Mean Cu		Coef. of
Domain	Number	ppm	Std. Dev.	Variation	Minimum	Maximum
Outside	7,047	92	184	2	0	2,410
Inside	2,088	469	469	1	3	4650

17.4.4Contact Profiles

For the updated models, contact profiles were generated to determine cases where hard or soft boundaries should be used during grade interpolation. The graphs were examined and search strategies as described in the grade interpolation sections were adopted.

17.4.5Variography

Directional semi-variograms, (15 degree increments horizontal, 30 degree increments vertical) were modelled for both Verde and Pancho, examining only the composited drill data below the current topographic surface

The variography focused on the composites for the three zones defined at Verde and Pancho. The models generally indicated isotropic, two-structure models with well defined nugget values that were less than 35% of the modelled sill value.

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The models showed good ranges in all directions, typically in excess of 100 meters.

Correlograms also showed good continuity with ranges in excess of 200 m along the vertical.

Table 17-8: Verde Variogram Table

	Variogram Axis	Axis Angles
Domain	Direction	(Azimuth ; Dip)	Nugget	C1	C2	Range	Range 2
Domain 1,2	X (strike)	44;-0 , 127;22	0.15	0.571	0.279	13.4	107.8
	Y (dip)	317;89 , 32;13				23.5	148.3
	Z (across dip)	134;1 , 273;65				9.5	220.8
Domain 3	X (strike)	174;76 , 37;1	0.21	0.545	0.245	32	91.7
	Y (dip)	274;2 , 131;81				32	711.2
	Z (across dip)	185;-14 , 126;-9				23.6	111

Table 17-9: Pancho Variogram Table

	Variogram Axis	Axis Angles
Domain	Direction	(Azimuth ; Dip)	Nugget	C1	C2	Range	Range 2
Domain 1,2	X (strike)	64;18 , 60;-21	0.15	0.702	0.148	50.7	1022.3
	Y (dip)	335;- 3 , 329;-3				45.8	550.4
	Z (across dip)	254;72 , 52;69				33	177.8
Domain 3	X (strike)	72;17 , 37;12	0.194	0.639	0.167	87.2	406.4
	Y (dip)	346;-15 , 284;62				35.4	2170.7
	Z (across dip)	295;66 , 133;25				29.7	341

17.4.6Model Framework

The Verde model is not rotated whereas Pancho is rotated 38° NW around the Z-axis. Both models have sufficient extent to encompass all known mineralization. The model cell size at Verde was set at 10 x 10 x 10 meters (X, Y and Z direction) while at Pancho, the model cell size was increased to 25 x 25 x 10 meters to take into account the wider spaced drilling at depth.

17.4.7Gold Grade Interpolation Verde

The primary interpolation method for gold was ordinary kriging which used the anisotropy and ranges supported by the best variogram model for each domain. Grade interpolation was verified by inverse distance to the 2^nd power and nearest neighbour interpolation. Grade interpolation in Domain 1 did not utilize octant search restrictions but used a soft boundary

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with Domain 2 composites. A minimum number of 2 samples and maximum of 8 samples were required to estimate the grade of any block. The number of samples from one drill hole that could influence a block grade estimate was limited to four (4).

For Domain 2, soft boundaries were used with both Domain 1 and 3. A minimum number of 2 samples and maximum of 8 samples were required to estimate the grade of any block. The number of samples from one drill hole that could influence a block grade estimate was limited to four (4). The sphere of influence of the composites greater than 0.6 g/t Au was limited to 50 m.

Grade interpolation in Domain 3 utilized octant search restrictions, requiring a minimum of two octants to estimate the grade of a block. A soft boundary approach was applied using samples from both Domain 2 and Domain3. The minimum number of samples per octant was 1 with a minimum of two octants required to estimate the block grade. A minimum of 2 samples and maximum of 10 samples were required to estimate the grade of any block. The number of samples from one drill hole that could influence a block grade estimate was limited to four (4).

Copper grades were not modelled at Verde, as the results of the metallurgical test work originally did not indicate that copper was impacting grade recovery or operating costs. Experience has shown a build up of copper in solution going to the recovery plant.

17.4.8Gold Grade Interpolation Pancho

The gold grade interpolation at Pancho also included kriging and inverse distance to the 2^nd power but did not use octant search restrictions. Domain1 gold grade were interpolated with a soft boundary with Domain 2 samples. A minimum number of 1 samples and maximum of 8 samples were required to estimate the grade of any block. The number of samples from one drill hole was limited to four (4).

For Domain 2, soft boundaries were used with both Domain 1 and 3. A minimum number of 1 samples and maximum of 8 samples were required to estimate the grade of any block. The number of samples from one drill hole that could influence a block grade estimate was limited to four (4).

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A soft boundary approach was applied using samples from both Domain 2 and Domain3. The minimum number of samples was 1 and the maximum limited to 8. The number of samples from one drill hole was a maximum of four (4).

Copper Interpolation Pancho

Analysis of the drill data indicated that cyanide soluble copper grades related to the presence of a chalcocite blanket are significant. Therefore, copper grades were interpolated into the Pancho model blocks independently of the gold grade. Kriging was also the primary estimation method. Search ellipsoid anisotropy and ranges were based on the copper variograms for the two copper domains.

The copper grade interpolation included kriging and inverse distance to the 2^nd power but did not use octant search restrictions. Domain 0 copper grades were interpolated with a soft boundary incorporating some samples from Domain 1. A minimum number of 1 samples and maximum of 8 samples were required to estimate the grade of any block. The number of samples from one drill hole was limited to four (4).

The interpolation for Domain 1 blocks used a similar approach with the same minimum and maximum number of composites with the effect of a single hole being limited to 4 samples.

Cyanide Soluble Copper Interpolation

The cyanide soluble copper grade interpolation followed exactly the same methodology used for copper.

17.4.9Densities

The densities that are in the current Verde model were originally based on average values by redox state that were determined at the time of the pre-feasibility study. Since that time, the redox surfaces have been updated with the blasthole data, so a check was completed to ensure that the average values were still reasonable.

The density measurements were coded with the updated redox interpretation and the statistics were compared to the values used in the previous model. Table 17-10 shows that

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there is no significant difference and therefore the density assignment will remain the same as in previous models. In addition, a review of the data shows that there is insufficient data at this time to produce a spatial density model. Because the variation in the density values by redox code is not significant within the redox type, a mean density value is considered reasonable.

At Pancho the density values were coded using the newly interpreted oxidation surfaces. The statistics were calculated and compared with the previous model. It is, as in the case of Verde, considered reasonable to use a mean density value in the blocks. Table 17-11 shows the comparison between the original model and the model updated in 2007.

Table 17-10: Density Measurements and Oxidation Level for Verde

	Original Values		2007 Checks
Code	Mean	Mean	No. Data	Min.	Max.	C.V.
Oxide	2.45	2.45	55	2.12	2.63	0.06
Mixed	2.52	2.53	219	2.16	2.69	0.04
Sulphide	2.57	2.57	467	2.10	2.72	0.03

Table 17-11: Density Measurements and Oxidation Level for Pancho

	2003 Values		2007 Checks
Code	Mean	Mean	No. Data	Min.	Max.	C.V.
Oxide	2.27	2.20	642	1.38	2.73	0.09
Mixed	2.38	2.40	196	2.10	2.63	0.06
Sulphide	2.49	2.50	2,506	1.83	2.90	0.05

17.4.10Model Checking and Verification

Block model verification included visual inspection of the model after grade interpolation was completed to examine if the grades were consistent with the available raw data. The model was manually reviewed in plan and section using Vulcan visualization tools. Visual inspection did not identify any errors in the model.

17.5Resource Classification

The mineral resource estimates for Verde and Pancho have bee prepared and classified in accordance with the Canadian Institute of Mining, Metallurgical and Petroleum’s (“CIM”)

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Standards on Mineral Resources and Reserves, Definition and Guidelines Resources classification was based on the variogram ranges for each zone, number of holes found in the search ellipsoid and proximity of the composites to the center of the block being estimated.

17.5.1Resource Classification Verde

The model update resulted in a better understanding of the mineralization and its controls at Verde. The following conditions had to be met for classifying the blocks in the model.

•Measured Resource – Information from three drill holes found in a 50 m search ellipsoid with the closest composite at a distance of 35 m or less.

•Indicated Resource – Data from two drill holes found within a search of 70 m with the closest composite at a maximum distance of 50 m.

•Inferred Resource – Remainder of the blocks including all the blocks located in Domain 3.

17.5.2Resource Classification Pancho

The model is now including an epithermal upper portion that is drilled on approximately 35 m and a porphyry system drilled at depth with average drill spacing greater than 75 m in some areas. These considerations and the experience gained while mining at Verde also guided the new classification scheme used for Pancho. The following conditions had to be met for classifying the blocks in the model.

•Measured Resource – Information from three drill holes found in a 75 m search ellipsoid with the closest composite at a distance of 35 m or less.

•Indicated Resource – Data from two drill holes found within a search of 100 m with the closest composite at a maximum distance of 50 m.

•Inferred Resource – Remainder of the blocks.

17.6Mineral Resource Estimate

The mineral resources estimated using a $625 gold price for Verde and Pancho as of December 31, 2007 are presented in Table 17-5. Mineral resources are reported exclusively of reserves.

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Table 17-12: Measured and Indicated Mineral Resources

		Ore Tonnes	Grade	Gold Ounces
Deposit	Class	(x 1,000)	(Au g/t)	(x 1,000)
Verde	Measured	19,335	0.73	456
	Indicated	26,987	0.71	619
	M&I	46,322	0.72	1,075
Pancho	Measured	7,585	0.52	128
	Indicated	57,549	0.58	1,071
	M&I	65,134	0.57	1,198
Total		111,456	0.63	2,274

The Maricunga deposits also host an Inferred Mineral Resource of 134.71 million tonnes averaging 0.56 g/tonne Au at the same $525 US per ounce gold price. Pancho hosts 132.90 million tonnes averaging 0.56 g/tonne Au and Verde hosts 1.81 million tonnes averaging 0.66 g/tonne Au.

17.7Considerations for Reserves

It is Kinross policy that resource estimates for open pit mine projects are reported within optimized or designed pit shells at assumed gold prices that represent reasonable, long-term price projections. Final pit designs have been completed for both Verde and Pancho and as a result, only Proven and Probable Reserves are reported.

17.8Mineral Reserve Estimate

Kinross Technical Services (KTS) completed the pit optimization for defining the 2007 yearend Mineral Resource and Mineral Reserve estimates.  Mineral reserves are reported within actual pit designs while the resources are reported inside optimized pit shells. The results as of December 31, 2007 are summarized and presented in Table 17-13.

The Verde reserve change reflects a higher gold price in conjunction with a new resource model. The increase at Pancho is the result of modifications to the resource model, entirely revised costs as of September 30, 2007 and the application of the SART process. More details on the costs and parameters used for defining the reserves are described in the section below (Section 17.9 Pit Optimizations).

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Table 17-13: Proven and Probable Mineral Reserves

		Ore Tonnes	Grade	Gold Ounces
Deposit	Class	(x 1,000)	(Au g/t)	(x 1,000)
Verde	Proven	71,071	0.85	1,947
	Probable	32,409	0.74	772
	2P	103,480	0.82	2,719
Pancho	Proven	106,627	0.71	2,436
	Probable	69,395	0.58	1,290
	2P	176,022	0.66	3,726
Total		279,502	0.72	6,445

17.9Pit Optimizations

17.9.1Verde

The Verde block model was updated to reflect the change in topographic surface for 2007.  A Mining Cost Adjustment Factor (MCAF) was also included to take into account the increased haul distance as the pit will get deeper. The base mining cost used was $1.00 per tonne with an added $0.016 for longer ore haul to crusher and waste dump. Sustaining capital for leach pad development was also included ($0.26 per tonne).

Optimization parameters used in Whittle optimization software are presented in Table 17-14.

17.9.2Pancho

Pancho ore has variable operating costs, gold recoveries and copper grade.  The geological block model contains attributes for copper grade, cyanide soluble grade and ore type.  A Process Cost Adjustment Factor based on cyanide soluble copper grade but considering the application of the SART process was defined to optimize the block model.  Using 2007 costs and the test work completed in the SART Feasibility Study the following formulas were derived for the different ore types.

Oxide Process +G&A costs = $3.752

Mixed and Sulphide Process +G&A costs = 3.116 + 0.0037*Cu Soluble (ppm) + 1.116 * Cu Soluble (ppm) ^ - -0.112

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Table 17-14: Optimization Parameters for Verde

Mining Costs

Unit Mining Costs	$	1.000
Mining Recovery	97.50		%
Mining Dilution	2.50		%
MCAF based on elev	$	0.016

Pit Slopes

Profile 1 (East Verde)	Azimuth	Slope
	10	44.5
Below 4350 m	65	49
	155	52.5
	275	49

Profile 2 (East Verde)	Azimuth	Slope
From 4360 to 4430 m	70	44.5
	153	49

Profile 3 (East Verde)	Azimuth	Slope
Above 4440	0	38
	0	180

Profile 4	Azimuth	Slope
West Verde	0	51
	90	51
	180	49
	270	49

Process Recoveries

Oxide	Au	73	%
Mixed	Au	70	%
Sulphide	Au	53	%

Processing Costs

Process costs $US/t	$3.76
Gold Prices	$550 and $625
Selling costs $US/oz	Scaled to gold price

A Mining Cost Adjustment Factor (MCAF) was also included to take into account the increased haul distance as the pit will get deeper. The base mining cost used was $1.00 per tonne with an added $0.084 for longer ore haul to crusher and waste dump. Another $0.016 was also added per 10 m bench from 4300 m elevation to consider extra haul above or below the pit access road. Sustaining capital for leach pad development was also included ($0.26 per tonne).

The Pancho ore will be processed at a rate of approx. 49,000 tpd over the life of mine.  Average heap leach gold recoveries are based on test work results: 85% for oxide ore, 75%

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for mixed ore and 69% for sulphide ore.  The assumptions for the copper recoveries incorporate the dissolution rates for the different material types. They are 71%, 92% and 60% respectively for oxide, mixed and sulphide material. These rates are used in conjunction with a 95% plant recovery resulting in final recoveries of 67% for oxides, 87% for mixed and 57% for sulphides.  Table 17-15 summarizes the optimization parameters for Pancho.

Table 17-15: Optimization Parameters for Pancho

Mining Costs

Unit Mining Costs		$		1.000
Mining Recovery		97.50%
Mining Dilution		2.50%
Pit Slopes
Profile 1	Azimuth		Slope
	70		44.5
Below 4300 m	153		49
	210		44.5
	308		52.5
Profile 2	Azimuth		Slope
Above 4300 m	90		39
	90		39
Process Recoveries
Oxide	Au		85		%
Mixed	Au		75		%
Sulphide	Au		68		%
Processing Costs
Process costs $us/t	Block Model PCAF
Gold Prices	$550 and $625
Selling costs $US/oz	Scale

17.10Pit Designs

For the 2007 reserve estimate final pit designs were prepared for the Verde pits and Pancho.  Waste dump designs and overall site layout were also updated.

17.10.1Verde

Verde pit design work was completed by CMM’s mine planning engineers using Vulcan mining software.  The software allows the user to input slope design criteria that varies both

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with azimuth and elevation.  The software enables the user to specify the bench face angle and berm width for a given sector of the pit.  The slope sectors specified by Golder had two azimuths, a “from” and a “to” (i.e. from 70(o) ENE to 120(o) ESE).  In Vulcan, these were specified transition azimuths for each slope sector.  Vulcan also inserts access ramp segments based on the design parameters and the location specified by the user.

The following sections provide design details for each of the pit designs completed for this study.

East Verde Ultimate Pit Design

Due to the complex nature of the geology in the East Verde pit area a number of different slope sectors had to be used for completing the original pit design.  The slope recommendations provided by Golder in 2003 formed the basis of that design.

Since then, mining experience and geotechnical reviews allowed to modify slightly the angles. The most recent work was completed by A. Karzulovic and Ass. Ltda of Santiago in April 2007. Figure 17-3 shows the pit geometry adopted in the yearend 2007 design.

Figure 17-2: Interramp angles for Verde Pit Design

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West Verde Ultimate Pit Design

Slight adjustments to the interramp angles were made following the most recent geotechnical review completed in April 2007. An angle of 53 degrees is now used in the Verde Breccia compared to 51 degrees in the original design. In the porphyry material an angle of 53 degrees is also used in the design which corresponds to an increase of 3 degrees relative to the original design.

Figure 17-3: Verde Pit Design Phase 4 with Topography

17.10.2Pancho

The Pancho pit was designed using the same criteria as that used during the 2003 feasibility study.  Golder had recommended two slope design criterion and the appropriate rosettes were established in Vulcan.  The first rosette controlled the slope design from the pit bottom up to the 4300 elevation where the rock mass was generally considered more competent.   Inter-ramp angles varied from 44.5 to 49 degrees in the region southeast to the southwest. 

95

All other inter-ramp slopes were 52 degrees.  Above the 4300 elevation, controlled by the second rosette, the rock mass was considered to be weathered and therefore slope angles were shallower.  Golder recommended an inter-ramp angle of 38 degrees.  More recent geotechnical work by A. Karzulovic and Associates Ltda. of Santiago (April 2007) indicate that the angle in the volcanic material above 4300 m elevation could be increased to 43 degrees.  Figure 17-3 shows the pit geometry adopted.

Figure 17-4: Pancho Pit Design with Topography

The author, Maryse Bélanger, has no knowledge of any environmental, permitting, legal, ownership, taxation, political or other relevant issue that would materially affect the current Mineral Resource and Mineral Reserves estimates

96

18.0OTHER RELEVANT DATA AND INFORMATION

There is no other data or information relevant to the project that is not covered in other sections of this report.

97

19.0INTERPRETATION AND CONCLUSIONS

The Mineral Resource and Reserve Estimates for the Maricunga Project have been prepared under the direct supervision of Maryse Bélanger, P. Geo., Director Technical Services, Kinross Americas.

Ms. Bélanger is satisfied the data used in the estimation of Mineral Resources and Reserves is free of gross errors and omissions and is of suitable quality and sufficient quantity to use in estimating resources and reserves for the Verde and Pancho deposits.

98

20.0RECOMMENDATIONS

The Maricunga mine is a producing gold mine for Kinross Gold Corporation.  The recommendation for successive work phases is not required.

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21.0        REFERENCES

Anonymous, 1991, Final feasibility study, Verde Gold Deposit Maricunga Property, Northern Chile:  Mineral Resources Development Inc., unpublished report for Bema Gold Corp. and Compañia Minera Maricunga, several volumes.

Anonymous, 1992, Revised feasibility study, Verde Gold Deposit Maricunga Property, Northern Chile:  Mineral Resources Development Inc., unpublished report for Compañia Minera Maricunga, Bema Gold Corp. and Amax Gold Inc., one volume.

Anonymous, 2007, Application of the SART Process at Maricunga, Project Feasibility Study, Idesol Ingenieros S.A, unpublished report for Compañia Minera Maricunga, one volume.

Ardilla, R.J., 1997. Summary Of The Geology of The Pancho Gold Deposit, Maricunga. Internal Report, Minera Cyprus Chile Ltda.

CMM Staff Geologists, Carrasco, M, Crawford, K, Duran, M. Rodriquez, B. Romero, I, Strusvieci, M, Urrutia, H. Cortes, 2001, Geologia del Yacimento Verde (Geology of the Deposit Verde, English Translation):  Compania Minera Maricunga, unpublished geologic map (1:2,000).  NOTE:  The mapping was conducted in 2000 and 2001 after the mining was completed; therefore, it is a “final pit map”.  The map is of good quality and the mapping is generally confirmed by the Kinross mapping and compilation conducted in 2003 as part of the feasibility study.  AutoCad File:  geologia2001.dwg.

Compania Minera Maricunga, Procedimiento Mapeo Pozos Tronaduras.  Procedures Files, in CMM files.

Cortes, C., Omar 2000, Informe de Pruebas Geolmetalurgicas de Mineralizacion Mixta. Noviembre 1999 - Febrero 2000, In CMM files.

Emmons, D, Verde East and West Geologic Mapping, 2003,  Kinross internal report, May 10,2003.

100

Flores V., Roman, undated, Precious Metal Deposits of the Maricunga Area, Northern Chile, In Kinross Files.

Khin, B.G. 2003. Petrographic Descriptions, Pancho Project, Chile. Consultants Report Prepared For Kinross Gold Corp.

Kinross Technical Services, 1998, Maricunga Property Review, June 1998, report in Kinross Files.

Magri, Eduardo and NCL Consulting, Drillhole Database, Audit, SD001 a SD177., , July 1995.  in files Maricunga Mine Site

McDermott, Ryan, 2000, Maricunga Ore Type Classification, Memorandum to Brian Scott October 12 2000, in Kinross files.

McDermott, Ryan, 2001, Maricunga Model Summary and Explanation, Memorandum to Rick Dye, January 30 2001, in Kinross files.

Miller, Vic, and McDermott, Ryan, 1999, Maricunga Trip, Memorandum to Rick Dye Dated September 14, 1999, in Kinross files.

Meister, Susan N., 1998, Verde Block Model and Resource, Estimate Report, dated December 16, 1998.  In CMM files.

Miller, Victor, 2002, Maricunga Resources and Reserves, Memorandum to Rick Dye dated February 20 2002,  in Kinross files.

Muntean, J. L., 1998, Magmatic-hydrothermal gold deposits of the Maricunga Belt, Northern Chile:  Stanford University, Ph. D. dissertation, 407 p.

Muntean, J. L. and Einaudi, M. T., 2000, Porphyry gold deposits of the Maricunga District, Maricunga Belt, Northern Chile:  Economic Geology, v. 95, pp. 1445-1472.

101

Muntean, J.L. and Einaudi, M.T. 2001. Porphyry Epithermal Transition: Maricunga Belt, Northern Chile. Economic Geology, v 96, pp 743-772.

Pritting, J, Geologic Summary of the Pancho Project, Maricunga District, Chile, May 2003.

Romero, Ivan, 2000, Informe de Cierre Temporal de Actividades Cia Minera Maricunga, Noviembre de 2000, Modificacado enero 9 2001. In CMM files

Romero, Ivan, 2001, Programa de Termino de Geologia y Planificación Mina. In CMM files

Scott, Brian, 1998, Trip to Maricunga-Jan 12-22, 1998, Bema Gold Corporation Memorandum to Tom Garagan,  in CMM files.

Scott, Brian, 1999, Deep Verde Drilling, Consolidated Westview Resources Memorandum to Tom Garagan, dated April 23, 1999, in CMM files.

Scott, Brian, 1999b, Maricunga Metallurgical Info 1991 - 1998, Bema Gold Corporation Memorandum to Tom Garagan, dated August 31. in CMM files.

Scott, Brian, 2000, Maricunga - Met Program 2000 - Geologic Observations, Bema Gold Corp. Memorandum to Tom Garagan, in CMM files.

Sillitoe, R.S., 1991. Gold Metallogeny of Chile, An Introduction. Economic Geology v.86,  pp 1187-1205.

Sillitoe, R. H. and Camus, F., 1991, A special issues devoted to gold deposits in the Chilean Andes:  Economic Geology, v. 86, no. 6.  NOTE: contains papers on the Maricunga gold deposits and the Maricunga Gold Belt.

Spiteri, J. G. and Cohoon, G. A., 2002, Report on resources Verde deposits Maricunga, Chile (draft):  Spiteri Geological and Mining Consultants Inc., unpublished report for Kinross Gold Corp., 101 p.

SNC-Lavalin, Pre-Feasibility Study, Maricunga Project, September, 2003

102

SNC-Lavalin, Maricunga Mine Metallurgical Summary, December, 2003

Strusievici, Mihai 1998,  A Review of the Geology of Verde Gold Deposit, Preliminary Conclusions - Implication in the Resource Model, in CMM files

Strusievici, M., 1999, Geological report, mineralization model of Verde – targets for mine site exploration:  Compania Minera Maricunga, unpublished report with geologic map, 24 p.  NOTE:  Describes the staff geologic mapping for the 1999 Verde pit maps.

Strusievici, M., et al., 1999, Geologic maps Verde deposit:  Compania Minera Maricunga, unpublished geologic map (1:1,000).  NOTE:  The map is the product of a 1998 pit mapping project by CMM staff including geologists Crawford, K., Duran, Mauricio, and Urrutia, Hegel.  The map was compiled and interpreted by Strusievici and others.  AutoCAD File:  Verde Geo.dwg

Vila, T. and Sillitoe, 1991, Gold-rich porphyry systems in the Maricunga Belt, Northern Chile:  Economic Geology, v. 86, pp. 1238-1260.  NOTE:  The paper has a good summary of the exploration history of the Maricunga Belt.

Williams, S.A., 1997. Petrographic Descriptions, Pancho Project, Maricunga District. Consultant’s Report Prepared For Minera Cyprus Chile, Ltda.

Wolverson, N.J. 1997. Pancho Summary Report, 1997. Internal Report, Compania Minera Maricunga.

103

22.0        DATE AND SIGNATURE PAGE

The undersigned prepared this Technical report, titled NI-43-101 Technical Report for the Maricunga Gold Mine, dated 31 December 2007.  The format and content of the report are intended to conform to Form 43-101F1 of the National Instrument (NI 43-101) of the Canadian Securities Administrators.

“Signed and Sealed”

Maryse Bélanger, P. Geo. 31 March 2008

104

23.0        ADDITIONAL REQUIREMENTS FOR OPERATING PROPERTIES

23.1        Mining Operations

The Maricunga Mine is an operating mine with significant infrastructure in place. The reserve estimates presented in this document are sufficient to support the operation for 16 years based on daily production of 45,000 tpd.

Maricunga production re-opened in October 2005 and achieved its targeted production rate of 14 million tonnes per year (40,000 tonnes per day) in late 2005.  The mine operates two 12-hour shifts per day for 352 days annually allowing for inclement weather interruptions.  Final pit design for Verde and Pancho assumed 10 meter bench heights, bench face angles of 65° to 70°, berm widths of 8 to 11 meters, berm interval of 20 meters, inter-ramp angles of 38° to 52.5° and haul road gradient at 10% with a 25 meter road width.

The Maricunga gold recovery process consists of a single line primary crushing, fines crushing (secondary and tertiary), heap leach and adsorption and regeneration (“ADR”) plant operation. The process treats 45,000 tonnes per day of dry Maricunga ore using primary crushing followed by a secondary and tertiary crushing plant.  The crushing plant product is approximately 80% passing 10.5 millimetres.  A pad type heap leach and an ADR plant are used for gold recovery.

A comprehensive program of metallurgical testing incorporating bottle roll tests and column leach tests of samples obtained from drilling was established to determine gold recovery and reagent consumption data for the remaining Verde resources and the Pancho resource.  Based on the recovery estimates by ore type, process recovery over the mine life averaged 69% of contained gold in the plant feed.  Life of mine annual gold production is expected to range from 230,000 to 305,000 ounces on a 100% basis.

The increase in reserve as of December 31, 2007 will require an expansion of the heap leach facility. This report considered the findings of a pre-feasibility study level completed by Dave Prins of PrinZ Mining Services based in Santiago, Chile. The increased reserves will

105

result in the need to permit additional leach pad capacity but this is not considered to be a risk, as the existing permitted space is sufficient for the majority of the remaining reserves.

23.1.1     Mobile Equipment

The mine operates as a traditional open pit operation. A partial list of the major mine equipment necessary to achieve the planned production rate includes:

·3 CAT 994D Front End Loaders

·12 CAT 785C Haul Trucks

·1 CAT 773E Haul Truck

·2 Ingersol Rand 250 mm Rotary Drills

·1 ECM 720 Rotary Drill

·2 CAT D9R Bulldozers

·2 CAT D8R Bulldozers

·1 CAT D11R Bulldozer

·2 CAT 14H and 16H Motor Graders

·1 CAT 834 Wheel Dozer

·1 50t Water Truck

·1 CAT 330CL Backhoe

Ancillary equipment include fuel and lube trucks, flatbed trucks, portable light towers, and service and welding trucks.

23.1.2     Leach Pad and Plant Infrastructure

The plant facilities were upgraded with all the work completed by October 2005. Key aspects of the upgrade included establishing a 110 km power line to replace the generated power currently at site. Additional capital expenditures went towards replacing certain components to improve availability and throughput of the crushing and conveying system. All equipment and systems are in place and allow for meeting a daily throughput target of 45,000 tpd.

106

In 2007 the SART process was tested for the Pancho ore as uneconomic concentrations of copper are present in ore. Some of the copper minerals are soluble in cyanide solutions and these minerals consume cyanide reagent and contaminate the leach solution with copper metal.  A feasibility study and pilot testing program were completed by Idesol of Santiago. The study considers plant design, performance metrics, operating and capital costs and implementation strategy. The reserves defined at Pancho consider the application of SART.

23.2        Recoverability

Metallurgical recovery information is discussed in detail in Section 16 of this report.

23.3        Markets

Kinross will continue with current marketing arrangements at Maricunga and does not envision any concerns related to marketing doré.

23.4        Contracts

Smelting, refining, handling and sales charges are within industry standards.

23.5        Environmental Considerations

A reclamation plan for the current mine disturbance was approved in 2002, based on the commitments made in the original environmental impact assessment for the site (1994).  The plan addressed physical activities, such as earthworks, but did not address chemical closure of the heap.  A closure plan for chemical stabilization of the heap has been prepared and has been submitted to the regulatory authorities in the form of a Declaration of Impact to the Environment (“DIA”).  The regulatory agencies are currently considering the Company’s proposed chemical stabilization approach and further discussions with agencies are expected prior to a decision regarding the chemical stabilization plan.  The agencies consider submittal of the chemical stabilization DIA as meeting the commitments in the original environmental impact

Future cash outflows for site restoration costs at Maricunga under CICA Handbook Section 3110, as at December 31, 2007, are estimated at approximately US $4.4 million.  There is

107

no requirement to post financial assurance to secure site restoration costs in Chile at present.

The increased reserves will result in the need to permit additional leach pad capacity but this is not considered to be a risk, as the existing permitted space is sufficient for the majority of the remaining reserves.

23.6        Capital Costs

Capital requirements were defined based on a mine life of 16 years utilizing the SART process in future years. The capital costs critical items include construction costs for the SART process facility, sustaining capital for continued expansion of the existing leach pad and partial replacement of the mining fleet in year 2015. The breakdown of costs was established by activity area. They are reported in three main categories: mining, processing and others.  The costs associated with upgrading some existing equipment in 2008 report to the Other category.  The schedule of capital costs is presented in Table 23-1.

Table 23-1: Capital Costs – Maricunga

	2008	2009	2010	2011	2012	2013	2014	2015	2016	2017
Mining
Stripping
Project construction
Equipment	1,900	7,200	13,290	10,650	2,050		13,530	28,900	1,500	4,300
Sustaining	10,233	3,227	10,965	841	755	473	325	583	341	855
Other
Subtotal	12,133	10,427	24,255	11,491	2,805	473	13,855	29,483	1,841	5,155
Processing
Project Construction	8,000	2,000	29,387	24,926	5,000	3,000	3,500	3,500	3,500	3,500
Equipment
Sustaining		1,000				500				500
SART	539	19,704
Subtotal	8,539	22,704	29,387	24,926	5,000	3,500	3,500	3,500	3,500	4,000
Other
Project Construction
Equipment
Sustaining	14,879	-13,916
Other
Subtotal	14,879	-13,916
Capital Costs	35,551	19,215	53,642	36,417	7,805	3,973	17,355	32,983	5,341	9,155

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	2018	2019	2020	2021	2022	2023	2024
Mining
Stripping
Project construction
Equipment	980
Sustaining	325	325	225
Other
Subtotal	1,305	325	225
Processing
Project Construction	3,500	3,500	3,500	3,500	3,500	3,500
Equipment
Sustaining
SART
Subtotal	3,500	3,500	3,500	3,500	3,500	3,500
Other
Project Construction
Equipment
Sustaining
Other
Subtotal
Capital Costs	4,805	3,825	3,725	3,500	3,500	3,500	0

Note: All amounts expressed in Table 23-1 are in thousands of dollars.

23.7        Operating Costs

Operating cost estimates were originally prepared based on first-principles and verified to historical mine operating costs available for the project. Operating costs used in this report are consistent with present operating experience. They are updated on a regular basis with a full cost review completed in November 2007.

A detailed 2008 and LOM budget has been developed for the mine based on an actual operating cost history and projected future performance.  Current operating costs are similar to most mines of this size and equipment age and are presented in Table 23-2.

Table 23-2: 2007 Actual Operating Costs – Maricunga

Item	2007 Actual Cost
Ktonnes Mined	25,800
Ktonnes to Leach	13,691
Mining cost ($/tonne mined)	$	1.09
Mining cost ($/tonne processed)	$	2.05
Leaching cost ($/tonne processed)	$	3.37
G&A ($/tonne processed)	$	1.26

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23.8        Economic Analyses

The discounted cash flow analysis indicates that the mine generates a positive cash flow.  The details of the economic and sensitivity analyses are considered by Kinross to be confidential information. The economic models are considered as complete, reasonable, and meeting generally accepted industry standards. This information is available and individuals seeking to review these models must request and sign a confidentiality agreement with Kinross Gold Corporation.

23.9        Payback

This is not applicable to the producing operation.

23.10      Mine Life

A complete Life-of-Mine plan was prepared based on the estimated mineral reserves. The mineral reserves are sufficient to support gold production until 2024 (see Table 23-3).

Table 23-3: 2007 Life of Mine Plan

Year	2008		2009		2010		2011		2012		2013		2014		2015
Ore (tonnes)	15,400,000		16,051,200		17,234,805		17,259,980		17,285,155		17,285,155		17,285,155		17,285,155
Strip ratio	0.68		0.97		0.97		0.97		0.93		0.93		0.93		0.93
Waste (tonnes)	10,471,820		15,610,696		16,717,761		16,742,181		16,075,194		16,075,194		16,075,194		16,075,194
Grade (g/t)	0.77		0.82		0.80		0.78		0.80		0.78		0.69		0.64
Recovery (%)	58.2	%	56.2	%	58.2	%	63.9	%	83.1	%	77.8	%	75.7	%	74.0	%
Recov. ounces	222,753		239,048		257,383		278,076		371,321		335,893		290,088		264,647
Total tonnes	25,871,820		31,661,896		33,952,566		34,002,161		33,360,349		33,360,349		33,360,349		33,360,349

Year	2016		2017		2018		2019		2020		2021		2022		2023		2024
Ore (tonnes)	17,285,155		17,285,155		17,285,155		17,285,155		17,285,155		17,285,155		17,285,155		17,285,155		6,603,551
Strip ratio	0.93		0.93		0.93		0.93		0.93		0.38		0.16		0.23		0.21
Waste (tonnes)	16,075,194		16,075,194		16,075,194		16,075,194		16,075,194		6,646,585		2,767,186		4,023,861		1,372,398
Grade (g/t)	0.58		0.78		0.73		0.65		0.63		0.73		0.63		0.66		0.62
Recovery (%)	71.4	%	65.6	%	57.9	%	58.7	%	75.3	%	69.8	%	72.4	%	68.2	%	68.0	%
Recov. ounces	231,371		285,514		234,528		213,115		262,729		283,548		253,468		251,302		89,768
Total tonnes	33,360,349		33,360,349		33,360,349		33,360,349		33,360,349		23,931,740		20,052,341		21,309,016		7,975,949

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EX-99.2

Exhibit 99.2

Technical Report

for the

Fort Knox Mine

Prepared for Kinross Gold Corporation

and

Fairbanks Gold Mining Incorporated

Fairbanks North Star Borough, Alaska, USA

Prepared by:

David Quandt, Chief Mine Engineer

Chris Ekstrom, Chief Mine Geologist

Klaus Triebel, P. Geo., Senior Mine Geologist

This report has been prepared under the supervision of the qualified person:

R. D. Henderson, P. Eng., Vice-President Technical Services

Kinross Gold Corporation.

Effective Date: March 31, 2008

 Fort Knox Mine Technical Report

Certificate of Author

I, Robert Duncan Henderson, P. Eng., residing at 3295 Spruce Avenue, Burlington, ON, L7N 1J5 do hereby certify that:

·I am employed by Kinross Gold Corporation, 40 King Street West, Toronto, Ontario, Canada, M5H 3Y2, in the capacity of Vice President, Technical Services.

·I am a graduate of the University of Cape Town (1984) with a B.Sc. Chemical Engineering degree and have practiced my profession continuously since 1984.

·I am a member in good standing of the Association of Professional Engineers of Ontario, licence number 100107661.

·I have been involved with the mining industry continuously since my graduation from University and have operating and engineering experience in gold mines located in South Africa, Canada, United States of America, Russia, Brazil and Chile.

·I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43 -101.

·I am responsible for the supervision of the preparation of the “Technical Report for the Fort Knox Mine” dated March 31, 2008. I have read National Instrument 43-101 and Form 43-101F1, and the technical report has been prepared in accordance with this instrument.

·I have personally visited the mine site on several occasions from 2004 to 2007.  The most recent visit was in November, 2007.

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

·I have not had prior involvement with the property that is the subject of the technical report.

·I am not independent of the issuer.  Per section 5.3.2 of National Instrument 43-101 an independent qualified person was not required to write the technical report on the Fort Knox Mine.

2

Dated this 31^st day of March, 2008 at Toronto, Canada.

“Signed and Sealed”

Robert Henderson, P.Eng.,

Vice President, Technical Services,

Kinross Gold Corporation

3

TABLE OF CONTENTS

1.0	Summary				8
2.0	Introduction and Terms of Reference				12
	2.1		Terms of Reference		12
	2.2		Sources of Data and Information Contained in Report		13
	2.3		Field Involvement of the Qualified Person		13
3.0	Reliance on Other Experts				14
	3.1		Other Experts		14
	3.2		Independent Audit		14
4.0	Property Description and Location				15
	4.1		Fort Knox Open Pit		15
	4.2		Gil Property		16
	4.3		True North Open Pit		16
5.0	Accessibility, Climate, Local Resources, Infrastructure and Physiography				20
6.0	History				22
7.0	Geological Setting				25
8.0	Deposit Type				26
9.0	Mineralization				28
	9.1		Fort Knox Deposit		28
10.0	Exploration				32
11.0	Drilling				34
12.0	Sampling Method and Approach				38
	12.1		Reverse Circulation		38
		12.1.1		Reverse Circulation - Dry Samples	38
		12.1.2		Reverse Circulation - Wet Samples	38
	12.2		Core Samples		39
	12.3		Sample Composite Summary		39
13.0	Sample Preparation, Analyses and Security				41
	13.1		Quality Control, Quality Assurance		41
	13.2		Check Assays		41
	13.3		Blanks Program		42
	13.4		Field Duplicates Program		43
	13.5		Standards		43
	13.6		Sample Security		44
14.0	Data Verification				45
	14.1		Drill Hole Data Rejection		45
15.0	Adjacent Properties				47
16.0	Mineral Processing and Metallurgical Testing				48
	16.1		Fort Knox Mill Facility		48
		16.1.1		Milling Methods and Capacity	48
		16.1.2		Mill Recovery – Fort Knox Deposit	49

4

		16.1.3		Mill Recovery – True North Deposit	49
	16.2		Fort Knox Heap Leach Facility		50
		16.2.1		Heap Leach Methods and Capacity	50
		16.2.2		Heap Leach Recovery – Fort Knox Deposit	51
	16.3		Past Considerations		51
		16.3.1		True North	51
		16.3.2		Fort Knox	51
17.0	Mineral Resource and Reserve Estimates				52
	17.1		Mineral Resource and Reserve Statement		52
	17.2		Fort Knox Deposit - Modeling Methodology		53
		17.2.1		Overview	53
		17.2.2		Survey Data	54
		17.2.3		Assay Grade Capping	54
		17.2.4		Drill hole Compositing	54
		17.2.5		Block Model	56
		17.2.6		Lithology	57
		17.2.7		Domains	58
		17.2.8		Variography	58
		17.2.9		Interpolation Methodology	59
		17.2.10		Search Parameters	59
		17.2.11		Specific Gravity - Tonnage Factors	60
		17.2.12		Dilution	60
		17.2.13		Metallurgical Recovery	61
		17.2.14		Model Checks	61
		17.2.15		Resource and Reserve Classification	61
	17.3		Pit Optimization		63
	17.4		Production Reconciliation		64
18.0	Other Relevant Data and Information				66
19.0	Interpretations and Conclusions				67
20.0	Recommendations				68
21.0	References				69
22.0	Date and Signature Page				72
23.0	Additional Requirements for Technical Reports on Production Properties				73
	23.1		Fort Knox Mining Operations		73
		23.1.1		Mine Equipment	73
		23.1.2		Mine Life Plan	74
		23.1.3		Other Considerations	75
	23.2		Recoverability		76
	23.3		Markets		76
	23.4		Contracts		77

5

23.5	Environmental Considerations	77
23.6	Taxes	77
23.7	Capital and Operating Cost Estimates	78
23.8	Economic Analyses	78
23.9	Payback	79
23.10	Mine Life and Production Schedule	79

6

LIST OF TABLES

Table 1-1: Proven and Probable Mineral Reserve Summary	10
Table 1-2: Measured and Indicated Mineral Resource Summary	10
Table 4-1: Mineral Title Summary	17
Table 12-1: Composite Summary Table	40
Table 16-1: Grade Recovery Relationship - Fort Knox Mine	48
Table 17-1: Mineral Reserve Summary as of December 31, 2007	53
Table 17-2: Mineral Resource Summary as of December 31, 2007	53
Table 17- 3: Variogram Parameters	61
Table 17-4: Classification Parameter Criteria	63
Table 17-5: Ore Reconciliation for 2007	65
Table 23-1: Estimated Operating Costs – Fort Knox	73
Table 23-2: Fort Knox Life of Mine Capital Spending	76
Table 23-3: Fort Knox Life of Mine Plan Production Schedule	76

LIST OF FIGURES

Figure 4-1: Fort Knox Project - General Location and Land Position Plan	18
Figure 4-2: Fort Knox Project - General Site Compilation Plan	19
Figure 7-3: Regional Geology	26
Figure 9-1: Fort Knox Deposit – General Geology	31
Figure 11-1: Drill Hole Location Plan – Fort Knox Deposit Area	34
Figure 17-1: Histogram – 5ft Uncapped Exploration Data used in 2007 Modeling	55
Figure 17-2: Histogram – 5ft Capped Exploration Data used in 2007 Modeling	55
Figure 17-3: Fort Knox Mine – Composite Sample Statistics	56

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	Kinross Gold Corporation
	Fort Knox Technical Report

1.0SUMMARY

This report has been prepared for Kinross Gold Corporation and Fairbanks Gold Mining Inc. to fulfill the requirements of National Instrument 43-101 as they relate to the estimation and reporting of mineral reserves and mineral resources situated on properties owned or controlled by Kinross Gold Corporation (Kinross) and its subsidiary Fairbanks Gold Mining, Incorporated (FGMI) in the Fairbanks mining district, Fairbanks North Star Borough, Alaska USA.  The report covers the period from January 1, 2007 through December 31, 2007.  Information disclosed prior to this date is covered in previous Technical reports that are located on SEDAR (sedar.com).

Kinross is the owner of the Fort Knox mine located in Fairbanks North Star Borough, Alaska.  The Fort Knox mine includes the main Fort Knox open pit mine, mill, heap leach and tailings storage facility, and an 80% ownership interest in the Gil property that is subject to a joint venture agreement with Teryl Resources Corp (“Teryl”), and the True North open pit mine (mining is currently suspended).

Fort Knox is located 40 kilometres (25 miles) by road, northeast of the city of Fairbanks, Alaska.  Kinross’ mining and exploration properties are located within the Fairbanks mining district, a northeast trending belt of lode and placer gold deposits that comprise one of the largest gold producing areas in the state of Alaska.

The Fairbanks district is situated in the north-western part of a geologic formation called the Yukon – Tanana Terrane (“YTT”).  The YTT consists of a thick sequence of poly-metamorphic rocks that range from Precambrian to upper Palaeozoic. The dominant rock types in the district are gray to brown, fine-grained micaceous schist and micaceous quartzite known as the Fairbanks Schist.  The Cleary Sequence, consisting of bimodal metarhyolite and meta-basalt with actinolite schist, chlorite schist, graphite schist, and impure marbles, is intercalated with the Fairbanks Schist.  Higher grade metamorphic rocks of the Chatanika Terrane are thought to be middle Palaeozoic to Ordovician and they outcrop in the northern part of the district.  Granodiorite to granite igneous bodies intrude YTT rocks.

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The mineral deposits are generally situated in a northeast trending, structurally complex zone characterized by a series of folds, shear zones, high angle faults, and occasional low angle faults.  Northeast striking high angle faults influence the location of gold deposits.

The Fort Knox gold deposit is hosted by a granitic body that intruded the Fairbanks Schist.  The surface exposure of the intrusive body is approximately 1,100 meters in the east-west direction and 600 meters north-south.  Gold occurs in and along the margins of pegmatite veins, quartz stockwork veins and veinlets, quartz-veined shear zones, and fractures within the granite.  The stockwork veins strike predominantly east and dip randomly.  Stockwork vein density decreases with depth.  Shear zones generally strike northwest and dip moderately to the southwest.  Gold mineralization in the quartz-filled shears is distributed relatively evenly, and individual gold grains are generally less than 100 microns in size. The gold occurrences have a markedly low (less than 0.10%) sulphide content.

The True North gold deposit is located in the Chatanika Terrane.  Gold is hosted in mafic to felsic schists and is frequently accompanied by carbon and carbonate alteration in sheared or otherwise structurally prepared zones.  The gold is very fine grained, and is closely associated with pyrite, arsenopyrite, and stibnite in the unoxidized zones.  It occurs in quartz veins, and in altered and brecciated rocks. There appears to be a direct relationship between veining and gold content, as weakly veined rocks generally carry lower gold values.

The Fort Knox deposit provides feed for the mill, a modern carbon-in-pulp gold extraction plant with a 32,658 to 45,359 (36,000 to 50,000 short tons) tonne per day capacity.  A heap leach facility is currently under construction.  The leach pad is designed to receive 32,658 to 45,359 tonnes per day (36,000 to 50,000 short tons) from the pit and stockpiles starting in the 3^rd quarter of 2009.

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FGMI controls an extensive property position, which covers the Fort Knox gold deposit and extensions of the mineralized zones. The Company has exploration and development drilling programs on the property. Kinross’ regional exploration within the Fairbanks district totalled $1.4 million in 2006 and $4.4 million in 2007.

At the end of 2007, FGMI had identified mineral reserves and mineral resources for the deposit as shown in Tables 1-1 and 1-2.   Table 1-1 summarizes the Proven and Probable Mineral Reserves for Fort Knox as of December 31, 2007 using an assumed gold price of $550 US per ounce.

Table 1-1: Proven and Probable Mineral Reserve Summary

	Proven		Probable		Proven and Probable
	Tonnes	Grade	Tonnes	Grade	Tonnes	Grade
Source	(,000)	(g/t)	(,000)	(g/t)	(,000)	(g/t)
Mine	31,498	0.69	104,098	0.59	135,596	0.61
Stockpile	105,319	0.35			105,319	0.35
TOTAL	136,817	0.43	104,098	0.59	240,915	0.50

In addition to the above noted mineral reserves, Measured and Indicated Mineral Resources as of December 31, 2007 at an assumed gold price of $625 US are summarized in Table 1-2.  The Gil and True North projects contain resources included in this table, but contain no mineral reserves.

Table 1-2: Measured and Indicated Mineral Resource Summary

	Measured		Indicated		Measured & Indicated
	Tonnes	Grade	Tonnes	Grade	Tonnes	Grade
Deposit	(,000)	(g/t)	(,000)	(g/t)	(,000)	(g/t)
Gil (80%)			2,838	1.03	2,838	1.03
True North			3,646	1.49	3,646	1.49
Fort Knox	2,664	0.67	26,643	0.59	29,307	0.60
Fort Knox Area	2,664	0.67	33,127	0.73	35,791	0.72

The Fort Knox area also hosts an Inferred Mineral Resource of 2.128 million tonnes averaging 0.55 g/tonne Au at the same $625 US per ounce gold price.  The majority (2.108 million tonnes) of this material is contained at the Fort Knox deposit.

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Fort Knox is mined as a conventional truck and shovel open pit mine. The deposit has yielded slightly less than 163 million tonnes (179.6 M tons) of ore containing 4.61 million ounces of gold since 1996.

Production experience indicates the current mineral resource and reserve models are within a marginally acceptable tolerance of variance for both tonnage and grade of the deposit when compared to surveyed mine production. The reconciliation showed that the mineral model estimated total ore tones, grade and ounces to be less (2.4%, 9.4% and 12.1% respectively) than the ore control model revealed. Over the two-year period consisting of 2006 and 2007 (not depicted), the mineral model underestimated total ounces by 4.7%.

FGMI’s Life of Mine Plan (LMP) was based on the design pit and Proven and Probable Mineral Reserves at a $550.00 per ounce US gold price assumption.  The LMP indicates the Fort Knox deposit host sufficient mineral reserves to support a seven year mine life. Pit production will continue from 2008 through 2014 at the Fort Knox deposit. Thereafter, re-handling of low-grade stockpiles to the leach pad will occur until early 2019. The LMP specifies an annual production rate from the pit of 39.5 million tonnes, including 135.6 million tonnes of mill and leach ore averaging 0.61 g/tonne Au.  Mill recoveries are estimated at 85% during the life of mine, and long-term leach recoveries are estimated to be 65%.  FGMI estimated both operating and capital cost requirements in support of the LMP. The financial analysis indicates a positive cash flow. The LMP, cost estimates, budgets and financial analysis have been prepared using industry standard procedures and assumptions. The authors of this report have reviewed these documents and found them to be complete and free of gross errors or omissions.  FGMI considers these documents to be confidential in nature and therefore, are not incorporated into the body of this report. Individuals requiring access to said documents are required to sign a confidentiality agreement with Kinross Gold Corporation.

The qualified person is of the opinion that this document supports the December 2007 Fort Knox and Area Mineral Resource and Reserve Statement.

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2.0INTRODUCTION AND TERMS OF REFERENCE

2.1Terms of Reference

This report has been prepared for Kinross Gold Corporation and Fairbanks Gold Mining Inc. to fulfill the requirements of National Instrument 43-101 as they relate to the estimation and reporting of mineral reserves and mineral resources situated on properties owned or controlled by Kinross Gold Corporation (Kinross) and its subsidiary Fairbanks Gold Mining, Incorporated (FGMI) in the Fairbanks mining district, Fairbanks North Star Borough, Alaska USA.  The report covers the period from January 1, 2007 through December 31, 2007.  Information disclosed prior to this date is covered in previous Technical reports that are located on SEDAR (sedar.com).

The FGMI Technical Services staff involved in the preparation of the mineral reserve and mineral resource estimates cited in the report are, David Quandt, Chief Mine Engineer, Chris Ekstrom, Chief Mine Geologist, and Klaus Triebel, P. Geo., Senior Mine Geologist. These individuals have been regularly involved in the day-to-day activities related to exploration, development, mineral reserve and resource estimation and production of FGMI’s properties in the Fairbanks mining district during 2007.

The mineral reserve and mineral resource estimates cited in this report were classified in conformance within the meanings ascribed by the Canadian Institute of Mining, Metallurgy, and Petroleum CIM Standards on Mineral Resources and Reserves Definitions and Guidelines.

Conversion of mine data to metric equivalents used the following conversion factors:

Tons to tonnes	multiply by 0.9072
Oz / ton to g/tonne	multiply by 34.2857

Remaining units of measure (distance, area, etc,) unless otherwise noted are presented in both metric and English units of measure.

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2.2Sources of Data and Information Contained in Report

The mineral resource and mineral reserve estimates reported in this document were prepared using technical information from the files of FGMI.  Previous owners of the Fort Knox property developed assay and geologic data, which FGMI has reviewed. The majority of information used in the estimates was collected during exploration and development drilling programs performed by FGMI and from data developed by contractors who conducted studies under FGMI’s supervision.  A number of other sources of information have been used in the compilation of this report and a complete list of references is provided in Section 21 of this report.

2.3Field Involvement of the Qualified Person

This report has been prepared under the supervision of R. Henderson, P. Eng, Vice-President Technical Services, Kinross Gold Corporation.   R. Henderson has been associated with the Fort Knox Mine since 2004 and has visited the site on several occasions during that time frame, including most recently in November 2007.

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3.0RELIANCE ON OTHER EXPERTS

3.1Other Experts

Mineral Resource Development, Inc. (MRDI) of San Mateo, California assisted in the resource modelling through 2001.  Fairbanks Gold Mining Inc. and Kinross Technical Services staff developed the 2007 model reserve and resource methodology.

Diligent care has been taken in the preparation of this report and the authors are confident about the accuracy of supporting studies not prepared under their direct supervision.  While the authors did not supervise the preparation of reports relating to reserve estimation techniques prior to 2004, geotechnical studies, or any drilling and assaying prior to 2004, they have examined the reports and supporting data and have concluded that the information, conclusions and recommendations are reasonable. Accordingly, the authors are confident that report recommendations and conclusions are suitable for use in the preparation of the appropriate sections of this report.

3.2Independent Audit

Roscoe Postle Associates Inc. (RPA) was retained by Kinross Gold Corporation to independently audit the December 31, 2004 Mineral Resource and Mineral Reserve estimates of the Fort Knox operation.  RPA did not identify any significant “fatal flaws” in the Fort Knox resource and given the knowledge and understanding that FGMI has accumulated with the operation, RPA expects that the Fort Knox Proven and Probable Mineral Reserves will likely represent a reasonable global estimate of the total in-situ gold reserves.

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4.0PROPERTY DESCRIPTION AND LOCATION

The Fort Knox mine is situated in the Fairbanks North Star Borough in the State of Alaska, United States of America. Fort Knox is located 40 kilometres (25 miles) by road northeast of the city of Fairbanks, Alaska.  Figure 4-1 is a general location map of the Fort Knox property.

The Company controls a large and diverse group of properties that comprise its mineral holdings in the Fairbanks mining district. These properties include patented and unpatented Federal lode and placer mining claims and State of Alaska mining claims. Some of the claims are owned outright, while others are controlled through leases and/or joint venture agreements. Figure 4-1 presents FGMI’s land position within the Fairbanks Gold Mining District. The blue shaded areas represent the company’s holdings.  The Fort Knox Uplands Mining Lease has been legally surveyed.

4.1Fort Knox Open Pit

At the Fort Knox project, the Company controls 1,316 State of Alaska mining claims covering an area of approximately 19,180 hectares (47,396 acres), an additional 502 hectares (1,239.5 acres) of mineral rights comprised of an Upland Mineral Lease issued by the State of Alaska, a Millsite Lease (issued by the State of Alaska), and one unpatented (federal) lode mine claim. Cumulatively, these properties cover an area of approximately 19,682 hectares (48,636 acres).  Mineral reserves at the Fort Knox mine are situated on 505 hectares (1,248 acres) of land that are covered by a State of Alaska Millsite Lease.  Although this lease expires in 2014, it may be renewed for a period not to exceed 55 years. The property package is located within the State of Alaska within UTM

NAD 27 zone 6 coordinates of 474,950mE, 7,195,700mN and 495,700mE, 7,215,300mN as illustrated in Figure 4-1.

The State of Alaska Millsite Lease carries a 3% production royalty, based on net income and recovery of the initial capital investment.  Mineral production from State mining claims is subject to a Mine License Tax, following a three-year grace period after

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production commences.  The license tax ranges from 3% to 7% of taxable income.  There has been no production from State claims situated outside the boundaries of the Millsite Lease at the Fort Knox mine.

A brief summary of the mineral title is shown below in Table 4-1.  Kinross maintains a current detailed listing of all Fort Knox Title and Ownership documents [Refer to Hansen, 2008 “Mother of All Lists” 11^th edition, 1/2008].  This document is available from Kinross upon request. The unpatented federal lode claim is owned by Kinross and is not currently subject to any royalty provisions. As a result of high metal prices, Kinross royalties and production taxes are estimated at $3.3 million for 2007 compared to $2.2 million in 2006.

All requisite permits have been obtained for mining and continued development of the existing Fort Knox open pit mine and are in good standing.    All federal and state permits were received for the Heap Leach on or before October 31, 2007 at that time construction commenced on the project.

4.2Gil Property

The Gil property mineral claims cover approximately 2,521 hectares located contiguous to the Fort Knox claim block.  The claim block consists of 182 State of Alaska mining claims and is subject to a joint venture agreement between Kinross and Teryl.  Kinross’ ownership interest in the Gil claim block is 80%.  All production from the State of Alaska mining claims is subject to the State of Alaska Mine License Tax following a three-year tax grace period after production commences.  The State of Alaska Mine License tax is graduated from 3% to 7% of taxable income.  Kinross continues to actively explore the Gil claims.  The property package is located within the State of Alaska within UTM NAD 27 zone 6 coordinates of 485,800mE, 7,204,800mN and 498,600mE, 7,212,400mN as illustrated in Figure 4-1.

4.3True North Open Pit

The True North open pit mine mineral claims and leases cover approximately 1,993 hectares, located 43 kilometres northeast of the City of Fairbanks, Alaska.  Kinross owns

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77 State of Alaska mining claims, covering 712 hectares which are subject to a State production royalty tax of 3%.  Mineral reserves are situated on two groups of State claims that Kinross has leased from private individuals.  Mineral production to date has been from one of the leased claim blocks.  Mineral leases have been executed with third parties for an additional 82 State mining claims that cover approximately 935 hectares.  Leased claims are subject to net smelter return royalties ranging from 3.5% to 5%.   Mining at the True North open pit has been suspended. The property package is located within the State of Alaska within UTM NAD 27 zone 6 coordinates of 468,850mE, 7,212,450mN and 477,050mE, 7,217,700mN as illustrated in Figure 4-1.

Table 4-1: Mineral Title Summary

FORT KNOX

Private Land	Claims	Hectares	Acres
Mental Health Lease		1,802	4,452
Stepovich patent (the worm)		49	121
Lower Fish Creek patent		321	793
Upper Fish Creek patent		267	660
Sub-total		2,439	6,026
State Claims
Clark Creek Lease	34	194	480
Eagan’s Deep Creek	10	186	460
F.G.M.I.	1272	16,345	40,390
Sub-total	1,316	16,726	41,330
Federal Claims
Clark Creek Lease	1	8	20
F.G.M.I.	1	8	20
Sub-total	2	16	40
Upland Leases
Fort Knox Lease		461	1,140
DIT lease		40	100
Sub-total		502	1,240
FORT KNOX TOTAL	1,318	19,682	48,636

Gil Venture

Joint Venture	Claims	Hectares	Acres
Gil Venture	182	2,521	6,230
Gil Venture Total	182	2,521	6,230

TRUE NORTH

Private Land	Claims	Hectares	Acres
Mental Health Lease		291	720
USMS 1924		30	73
USMS 2174		25	61
Sub-total		346	855
State Claims
Raines/Cope Lease	42	340	840
Shepard (Archimedes) Lease	20	312	770
Shepard (Murray) Lease	6	97	240
Woodruff Lease	8	113	280
Towse/Davies Lease	6	73	180
F.G.M.I.	77	712	1,760
Sub-total	159	1,647	4,070
TRUE NORTH TOTAL	159	1,993	4,925
GRAND TOTAL	1,659	24,196	59,791

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Figure 4-1: Fort Knox Project - General Location and Land Position Plan

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Figure 4-2: Fort Knox Project - General Site Compilation Plan

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5.0ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

The Company’s mineral prospects and mining operations are situated in close proximity to the City of Fairbanks, which is a major center for the Interior region of Alaska. Fairbanks is the second largest city in Alaska, and has an estimated population of more than 35,000. The surrounding areas of the Fairbanks North Star Borough have a further 40,000 to 50,000 residents. Fairbanks is served by major airlines and the Alaska Railroad, and is connected to Anchorage and Canada by a series of well-maintained paved highways. Services, supplies, and energy (fuel and electricity) are available in Fairbanks in ample quantities to support the local and regional needs, along with the mining and processing operations of FGMI.

The Fairbanks mining district is an arcuate belt of placer and lode gold deposits that is more than 64 kilometres (40 miles) long and about 13 kilometres (8 miles) wide.  The district is situated within the northern foothills of the Tanana Valley of central Alaska. The south-western part of the mining district is located about 18 kilometres (11 miles) west of the city of Fairbanks and the north-eastern extremity of the district is approximately 48 kilometres (30 miles) to the northeast. The area has generally good access, with numerous all-weather paved highways and gravel roads maintained by the Fairbanks North Star Borough and the State of Alaska.

The mining district is a region of low hills and broad valleys occupied by meandering streams.  The hills are generally rounded with gentle slopes and irregular ridge patterns. The meandering ridgelines are the result of numerous gulches and streams that cut the flanks of hills. The most prominent topographic features include Ester Dome, located in the western part of the district at 720 meters (2,364 feet) above sea level, and Pedro Dome, situated in the north-eastern part of the area at 609 meters (2,000 feet) above sea level.

The Fairbanks area has a sub-arctic climate, with long cold winters and short summers. Winter low temperatures drop to the range of

–40 F to –55 F, while in the summer, highs may occasionally exceed +90 F. The annual precipitation in Fairbanks is approximately

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30.5 centimetres (12 inches). Typically, inclement weather conditions, such as temperature inversions or slippery road surfaces, will only negatively impact production in the open pit for portions of a few shifts in any given year.

The Fort Knox milling operation obtains its process makeup water from a fresh water reservoir located within the permitted property area.  The tailings storage area on site and is under review for storage capacity for the remaining mine life of the Fort Knox.  Power is provided to the mine by Golden Valley Electric Association’s power grid serving the area over a distribution line paid for by Kinross.

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6.0HISTORY

The Fort Knox area was actively explored for gold placer deposits since 1902 when Felix Pedro discovered gold in Fish Creek located downstream of the Fort Knox deposit. Since that initial discovery, the surrounding Fairbanks Mining district has produced in excess of 8 million ounces of gold, predominantly from placer deposits. Placer mining operations continue to this day.

Exploration for lode gold deposits was very limited during the early history of the region and focused on tracing the source of the placer deposits up headwaters and tributaries to Fish Creek. Alfred Brooks mapped the area while working for the United States Geological Survey (USGS) in the early 1900’s. Brooks’ mapping described a large granitic intrusive within the project area.  In the 1968 Final Report for Mineral Resources of Northern Alaska, L.A. Heiner and E.N. Wolff noted that the great amount of placer gold in the Fairbanks District led many workers to conclude that only the roots of the gold veins were left and that there was not enough economic incentive to test the theory by exploration.

In 1913, H. A. Currier staked lode-mining claims covering auriferous quartz veins on the Melba Creek-Monte Cristo Creek divide (covering part of what is now the Fort Knox gold deposit). A three-stamp mill was constructed on the property, but there is little evidence of any production from the claims. USGS geologists who examined the prospect noted the association of bismuthinite and gold in quartz veins and suggested a relationship between the observed mineralization and the large granitic intrusive located nearby.

The claims were dormant until 1980 when two local prospectors, Joe Taylor and George Johnson staked 19 state mining claims to work placer deposits.  Taylor and Johnson worked the placers of Monte Cristo Creek from 1980 – 1982 and recovered bismuthinite nuggets containing abundant gold. The demonstrated correlation between the gold and bismuth led Taylor and Johnson to prospect the slopes and divide between Melba and Monte Cristo Creeks. The prospecting operations involved panning and trenching, which suggested that gold mineralization was widespread and resulted in the prospectors staking an additional 34 mining claims.

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In 1984, Rob Blakestead, a consulting geologist, noted the presence of visible gold in quartz veins hosted by granite in the Fort Knox area. This discovery led to increasing levels of exploration to locate the source of the gold. In 1986, the claims were leased to Nye Minerals who entered into a joint venture agreement with Electrum Resources in 1987. During 1987 to 1991 numerous small mining companies actively explored the claims.

The Denver, Colorado based Amax Gold, Inc (AMAX) purchased the Fort Knox project in 1992 and established Fairbanks Gold Mining (FGMI) as a wholly owned subsidiary to operate the project. FGMI initiated extensive exploration programs on the property including surface geochemical sampling, drilling and geophysics.

Soil sampling proved the most useful exploration tool in delineating the ore body during the initial exploration of the deposit. Later surface trenching and mapping of the anomalies developed by the soil geochemistry identified the favorable exploration targets.

Ground magnetometer surveys performed in 1987, 1991 and 1992 were employed with limited success.  427 drill holes totaling over 262,000 feet were completed on the property by late 1992. This work was followed by the completion of environmental and engineering studies examining the feasibility of beginning commercial production from the deposit.

In 1991, Amax Gold Inc. (now Kinam Gold Inc. (“Kinam”), a subsidiary of Kinross) entered into a joint venture agreement with Teryl to explore the Gil property.  In 1992, Kinam acquired ownership of the Fort Knox property.  Kinross Gold Corporation acquired the property in 1998 as part of the merger with Amax Gold Inc., which involved stock exchange with Cyprus.

Construction of the Fort Knox mine and mill operations began in 1995 and were completed in 1997.  Commercial production at Fort Knox was achieved in March 1997,

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and the property has operated continuously since start up.  Construction of the mine was completed at a capital cost of approximately $373 million, which included approximately $28 million of capitalized interest.

Fort Knox is mined as a conventional truck and shovel open pit mine. The deposit has yielded slightly less than 163 million tonnes (179.6 M tons) of ore containing 4.61 million ounces of gold since 1996.

After acquiring ownership of the True North property in 1999, Kinross completed pre-production capital expenditures, primarily permitting and the building of a haulage road to the Fort Knox mill.  Commercial production at True North was achieved on April 1, 2001, but is currently suspended.

In 2006, Kinross Gold began negotiations with the National Oceanographic and Atmospheric Administration (NOAA), the Bureau of Land Management (BLM), and the State of Alaska’s Trust Land Office (TLO) to acquire a section of land bordering the existing claim group on the western boundary of the existing Fort Knox pit.  Acquisition of the land package was completed on December 31, 2007, allowing for a Phase 7 pit expansion.

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7.0GEOLOGICAL SETTING

The Fort Knox property is located in the Fairbanks mining district, a southwest–northeast trending belt of lode and placer gold deposits that comprise one of the largest gold producing areas in the state of Alaska.  The Fairbanks district is situated in the northwestern part of the Yukon–Tanana - - Uplands.  The Yukon – Tanana terrane consists of a thick sequence of poly-metamorphic rocks that range from Precambrian to upper Paleozoic. The polymetamorphic protoliths were primarily sedimentary, volcanic, and volcanoclastic units, with only minor plutonic rocks. The region has undergone at least two periods of dynamo - thermal metamorphism, which included an early prograde amphibolite event, followed by a retrograde, greenschist facies event (Bundtzen, 1981). A more complex deformational history identifying four phases of penetrative tectonism has been suggested by Hall, 1985.

The Fairbanks Schist, which is Proterozoic to lower Paleozoic, is the dominant lithology in the district. It is composed of quartz-muscovite schist, muscovite-feldspar-quartz schist, micaceous quartzites, metaconglomerate, garnet - hornblende amphibolite, and marble, indicative of an emergent shelf environment.  The Cleary Sequence consisting of bimodal meta-rhyolite and meta-basalt with actinolite schist, chlorite schist, graphite schist, and impure marbles is intercalated with the Fairbanks Schist, indicating immature rift basins in the shelf environment. Geological work performed in 1996 has led to reassigning the meta-rhyolite in the Cleary Sequence to the Devonian Muskox Sequence and placing it in fault contact with the Fairbanks Schist. In the northern part of the district, metamorphosed rocks of the Chatanika terrane have been identified.  The sequence includes type C eclogites, impure marbles, amphibolites, calc-muscovite schist, garnet-muscovite schist, and muscovite schist, containing garnet, biotite, chlorite and graphite. The Chatanika unit in fault contact with the Fairbanks Schist is thought to be middle Paleozoic to Ordovician and may represent a telescoped, mature rift basin within the shelf environment. Amphibole from the Chatanika unit yielded an Ordovician K-Ar date (Robinson, 1990).  Figure 7-1 illustrates the location of the Kinross projects within this geologic setting.

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Figure 7-3: Regional Geology

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8.0DEPOSIT TYPE

The Fort Knox gold deposit is hosted by a granitic body that intruded the Fairbanks Schist.  The surface exposure of the intrusive body is approximately 1,100 meters in the east-west direction and 600 meters north-south.  Several late Cretaceous to early Tertiary intrusive bodies penetrate the Yukon–Tanana terrane (Blum, 1982). They range from ultramafic to felsic composition and are distinguished from older intrusive rocks by their lack of metamorphic textures.  The Fort Knox deposit is hosted by one of the west-northwest trending, late-Cretaceous granitic complexes.  The pluton is offset by two northeast structures, which display left-lateral strike slip movement. These structures, the Monte-Cristo Fault and Melba Fault, are regional in extent and offset the Gilmore Dome Pluton south of Fort Knox. The Fort Knox pluton is composed of a light gray, fine grained granodiorite, medium grained biotite granite, and coarse grained, biotite granite porphyry, which form blocky bedrock exposures. The diverse textural and chemical granitic varieties and sharp to gradational intrusive contacts suggest the Fort Knox pluton is a multi-phase intrusive. The local occurrences of orthoclase megacrysts, resorbed quartz phenocrysts, and quartz glomero-phenocrysts support that observation. Crenulated quartz layers (brain rock) and dendritic growths of quartz and potassium feldspar present in the Fort Knox pluton contacts help to evaluate intrusive paragenesis (Bakke, 1994).

The mineral deposits are generally situated in a northeast trending, structurally complex zone characterized by a series of folds, shear zones, high angle faults, and occasional low angle faults.  Northeast striking high angle faults influence the location of gold deposits.   The dominant structural zones in the Yukon-Tanana terrane are the north bounding Tintina Fault system and the south bounding Denali Fault system. Northeast trending faults and shear zones within the Yukon-Tanana terrane are related to movement along these major structures. The northeast structures, which display a strike-slip displacement, were important to the localization of gold mineralization.

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9.0MINERALIZATION

The Fairbanks mining district is a celebrated placer gold camp with production in excess of 8.0 million ounces of gold since 1902. Although a significant mining district in terms of total production, it had only limited lode production until the discovery and development of the Fort Knox deposit in the late 1990’s. The discovery and development of the True North deposit (Figure 4-1) produced ore for the Fort Knox mill from 2001 to 2004, and further enhanced the lode production from the district.

Although the regional distribution of gold mineralization in the Fairbanks district is beyond the scope of this report, the area hosts gold in a variety of geologic settings, as follows:

·The Fort Knox deposit (the largest lode deposit in the district) is characterized by mineralization in quartz, quartz-sericite and quartz pegmatite veins, stock work zones, and mineralized shear zones;

·True North mineralization is hosted by complexly deformed, carbonaceous meta-sediments associated with quartz veins;

·The Ryan Lode deposit, gold occurs in and adjacent to large-scale shear zones;

·The Gil project, gold is hosted in skarns and quartz veins within the skarns;

·The Cleary Hill mine, gold occurs in quartz veins within the Fairbanks Schist.

9.1Fort Knox Deposit

The Fort Knox deposit is hosted entirely within the Fort Knox Pluton. The pluton is elongate and measures approximately 1100m east-west and 600m north-south.  Drilling has identified mineralization to the 400 level which marks the lowest extent of detailed information.   The contact with the Fairbanks schist is abrupt, and drilling indicates the pluton contacts plunge steeply to the east and moderately to the north, south, and west. The Fort Knox pluton has been subdivided into three mappable phases based on textural differences. Intrusion of the biotite-rich fine-grained granite was followed by a medium-grained porphyritic granite. The youngest intrusive phase is a coarse-grained,

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seriate porphyritic granite. A volumetrically minor, biotite-hornblende rich phase (mapped as “mafic”), that commonly displays a medium-grained texture is locally present as pendants near the schist-granite contact. Bakke (1995) provides whole-rock, trace element and a classification profile for the three main phases of granite.

Gold-bismuth-tellurium mineralization is restricted to the Fort Knox pluton and is strongly structurally controlled. Gold occurs within, and along margins of pegmatite veins, quartz veins and veinlets, within shear zones, and along fractures within the granite. The overall sulphide content is very low (< 0.1%) and the orebody is oxidized to the depths of the drilling.

The following is a description of vein types and associated alteration styles found at Fort Knox (Bakke and others, 1998).

1)Pegmatite veins and veinlets:  Range in thickness from micro-scale to 8cm. Composed of clear to gray quartz, large K-spar megacrysts, and micaceous clots. Potassic alteration halos, rarely exceeding 1cm thick, consist of an assemblage of variable amounts of secondary biotite and K-spar overgrowths on primary K-spar within the granite matrix.

2)Pegmatite veins similar to (1) above with alteration envelopes consisting of a variably developed phyllic (sericite + pyrite) assemblage.

3)Quartz veins and veinlets (stockwork): Range in thickness from micro-scale to 15cm. These veins possess thin albitic alteration halos

4)Quartz veins and veinlets similar to (3) above with phyllic alteration envelopes that range in thickness from 0.5-3cm.

5)Low-temperature fracture coatings and chalcedonic veins and breccia:  Low temperature assemblage of zeolite + calcite + clay + chalcedony. Pervasive throughout the deposit in the form of fracture coatings and breccia zones. Argillic alteration halos as much as 7m in width are developed adjacent to the larger chalcedonic breccia zones.

Bakke (1995) and McCoy and others (1997) note the strong geochemical correlation of gold with bismuth and tellurium. Bismuth and tellurium mineral species that have been identified include: native bismuth, Bi; maldonite, AuBi; bismuthinite, Bi₂S₃; tellurobismutite, Bi₂Te₃; bismite, Bi₂O₃; tetradymite, Bi₂Te₂S; and eulytite, Bi₄(SiO₄)₃. Other ore minerals that are found include trace to minor amounts of molybdenite and scheelite.

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The major structural trend controlling vein orientation and mineralization is defined by southeast-northwest trending, moderately to shallowly southwest dipping shear zones. The shear zones are typically filled with granulated white quartz, and range in thickness from 0.3-1.5m. They possess mixed phyllic and argillic alteration assemblages and contain abundant iron oxide clay gouge along the margins. In the vicinity of the shear zones, vein density increases and vein orientations are predominantly parallel to the shear direction. Abundant, Thin, subsidiary shears, are abundant between, and especially adjacent to, major shears.  Figure 9-1 displays the local geology of the Fort Knox deposit. The plan shows the extent of the granite intrusive with distribution and orientation of the major shear zones.

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Figure 9-1: Fort Knox Deposit – General Geology

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10.0EXPLORATION

Although the Fairbanks district has produced over 8.0 million ounces since its discovery in 1902, most of the production has come from auriferous gravels, rather than lode deposits. Exploration for the lode sources of the placer deposits has been performed periodically since the discovery of the district with little success from these efforts, until the 1990’s.

Standard exploration procedures have been utilized at the Fort Knox project as follows:

·Reconnaissance and detailed geologic mapping on topographic maps or aerial photographs at scales that are suitable to display the details of observed geologic features. Geologic mapping on topographic base maps is normally done at a scale of 1 inch = 500 feet, and mapping that is done on aerial photographs is at a scale of 1 inch = 2,000 feet. This work is undertaken by employees of the Company or contract/consulting geologists;

·Soil and rock chip sampling to determine the presence of gold mineralization, or associated trace elements. These samples are regularly collected during the course of geologic mapping programs by either Company employees or contract geologists, under Company supervision;

·Soil anomalies were trenched to create exposures of bedrock. The trenches were cut with bulldozers owned and operated by contracting firms. The trenches were mapped and sampled in detail by either contract geologists or Company personnel;

·Drilling has been the principal exploration tool used by the Company. The two drilling methods used were diamond core and reverse circulation (RC).  Company personnel directed the work performed by the independent drilling contractors. Sampling of the drill holes was completed by drill contractor employees, under close supervision of Company or contract geologists;

·Geochemical and assay determinations for gold and associated elements are undertaken by independent commercial laboratories. The Company has utilized the services of three firms – ALS Chemex Laboratories, Bondar-Clegg (now owned by the ALS Chemex group), and Alaska Assay Laboratories. Prior to 2002, the assays for Fort Knox and True North were completed by Bondar-Clegg at their Vancouver, B.C. laboratory. The samples were dried, weighed, crushed, and pulped at the Bondar-Clegg preparation facility in Fairbanks. The duplicate assay and quality control procedures were performed by Chemex in Vancouver. From 2002 to 2005, all assaying was done by ALS Chemex at their Vancouver laboratory, although sample preparation was done at their facility in Fairbanks. In 2006 and 2007, primary assays were performed by Alaska Assay Laboratories.

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Check assay work was performed by American Assay lab of Reno, Nevada from 2003 to 2004; at Alaska Assay Laboratories from late 2004 to 2005; and at the Vancouver laboratory of ALS Chemex in 2006 and 2007.

·Certain aspects of the process of mineral exploration and evaluation have varying degrees of precision. The Company employs methods that will determine the precision of the most critical elements, which are assaying and drill sample reliability. The sample QA/QC (quality assurance/quality control) program is both rigorous and effective. The quality control program is more fully described in Section 11.0 of this report. The Company also utilizes visual, logging, and weight measurement methods to determine if reverse circulation samples are representative of rock penetrated by drill holes.

Kinross’ regional exploration within the Fairbanks district totalled $1.4 million in 2006 and $4.4 million in 2007.

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11.0    DRILLING

Drilling has been used by FGMI to define mineral deposits in the Fairbanks mining district. Diamond drilling and reverse circulation drilling were the two methods used at the Fort Knox project. The deposit has been defined by 751 drill holes (307 core holes and 444 reverse circulation holes totaling 492,663 feet), which have provided 98,651 nominal 1.52-meter (5 foot) long samples, 97,420 of which were assayed for gold.

Figure 11-1 displays the drill hole distribution at the Fort Knox deposit.

Figure 11-1: Drill Hole Location Plan – Fort Knox Deposit Area

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Reverse circulation (RC) drilling is a specialized method of “rotary” drilling.  Drilling medium consisting of water, foam, drilling mud with additives, or air is circulated to the drill bit face on the outside of the drill rods.  The drill cuttings, which consist of pulverized rock mechanically ground by the drill bit, are collected through the inside of the rods. The reverse circulation method is an accepted technique commonly used in mineral exploration and development drilling programs. The reverse circulation holes completed by FGMI are normally 139.70 millimeters (5.50 inches) in diameter, but may range as high as 146.05 millimeters (5.75 inches) in diameter.

Diamond drilling or core drilling is the process of obtaining continuous cylindrical samples of rock from drill holes by means of annular shaped rock cutting bits rotated by a borehole-drilling machine. Core drilling is commonly used to collect undisturbed and continuous samples. It may also be used for pre-selected intervals of holes that are of particular interest. Core hole intervals can provide detailed, comprehensive samples for improved geologic description, geotechnical and rock strength tests, metallurgical tests, and density determination.

The Company has commonly drilled PQ3 diameter core holes (83.1 mm, or 3.270 inches) at the Fort Knox deposit since 1998.  Prior to 1998, core holes were PQ diameter (85.0 mm, or 3.345 inches in diameter).  Both PQ3 and PQ diameter core are used for exploration and evaluation of mineral deposits where a larger sample is more representative of coarse grained gold distribution. The two PQ core varieties are the preferred sample diameter at Fort Knox because of the local, coarse-grained nature of gold within the deposit.

Core and RC samples are collected from each drill hole and are geologically logged with detailed descriptions of rock types, alteration, mineral identification, and examination of structural features. Reverse circulation holes are sampled on regular 1.52-meter (5 foot) intervals for the entire length of the hole. Each core interval and RC sample is submitted to an independent assay laboratory for geochemical analysis. The resultant geochemical data along with the lithologic data is entered into the project database. This drill data is used to construct subsurface geologic maps, cross sections and solids, as well as

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depicting geologic contacts, interpretations of structural settings and the extent of alteration zones and intervals of mineralization. The “raw” data, as well as the geologic interpretations, are used to prepare estimates of mineral resources and mineral reserves that may be present.

Historically, the core from each diamond drill hole is photographed and the photographs are retained for future use. A representative “split” of all reverse circulation drill hole samples, collected at 1.52 meter (5 foot) intervals, are preserved in plastic “chip trays”. The trays are designed to hold 20 five-foot samples, and each sample has a weight of approximately 200 grams. The total core, from diamond drill holes, completed in the Fort Knox deposit was pulverized and submitted for assay.

Starting in 1997, angled core holes have been routinely surveyed down the hole with a Reflex EZ-Shot instrument. Although early core drill holes were not surveyed down the hole, readings from more recent drilling indicate that holes usually deviate 3^o or less over 1,000 ft. of core length.  The FLEXIT SmartTool survey system was used during the 2005 drill season.  The system downloads multi-shot surveys from a down-hole tool to a data pad on the drill deck. The data pad containing the digital down-hole survey information is then loaded to a computer where the survey results are calculated.  In 2006, drilling reverted to using the Reflex EZ-Shot for downhole surveys and continued to utilize this system in 2007.

The Company employs a detailed program of weighing the RC and core samples to determine if the specimen is under weight, which might indicate loss of material in the sample interval. If individual 1.52-meter (5 foot) intervals have unusually high or low weights they could indicate sample contamination in a drill hole. The formula used in the Fort Knox deposit is as follows:

Percent sample recovery = (sample weight) (split fraction) (4.3633*(bit diameter) ^2)*100.

Mineralized intervals with a calculated recovery greater than 100 percent are evaluated. The anomalous hole is flagged and examined in cross-section. The drill hole is compared to adjacent holes, historical production and a decision is made to accept or

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reject the assay interval. Rejected samples are coded in the database and composite values are assigned a “-1” (no sample) in estimating mineral resources.

The mineralization and host rock at the Fort Knox deposit require that attention be given to RC samples, which penetrate the water table within the deposit. The method of drilling this part of the deposit has been validated by the results of mining.

The sample widths are generally significantly less than the true width of the ore zones.  The orientations of the ore-bearing structures are discussed in Section 9.1. Although the orientations of the ore zones and the drillholes are somewhat irregular, in the vast majority of cases the drillholes pierce the ore zones at a significant angle.

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12.0    SAMPLING METHOD AND APPROACH

12.1    Reverse Circulation

The reverse circulation drill samples are collected every 1.52-meter (5 foot) by a geologist or helper at each drill site.  The data for each sample is recorded on a log sheet and entered into an Excel file.  A small portion of each sample is collected and preserved in plastic “chip” trays, (twenty 1.52 meter samples per tray) and retained for future use. Once the sample has been collected, it is placed in a pre-numbered, 45.72 by 60.96 centimeter (18 by 24 inches) 8-mil polyethylene bag. A sample tag is placed in each bag, which is then sealed with a wire tie. The samples are placed in bulk bags, with approximately 50 samples per bag and are picked up by staff from a commercial laboratory.

12.1.1    Reverse Circulation - Dry Samples

Drill cuttings are passed through a collection hose into a cyclone-type dust collector and are then manually split through a hopper-feed Gilson splitter. The split fraction of each sample is recorded on the log sheet.  Dry RC samples are not considered to be prone to contamination.

12.1.2    Reverse Circulation - Wet Samples

Cuttings are fed into a cyclone that deposits a stream of sample and drilling fluid into a splitter with a variable speed hydraulic motor that rotates a set of vanes controlling the volume of split sample. Prior to 2006, this split sample was fed into four 5-gallon buckets set in cascading series to collect and settle out the cuttings. A flocculent was added to the first bucket to aid in the settling of the sample. The samples were then permitted to settle.  Beginning in 2006, the split sample was fed into one, perforated and lined, 5-gallon bucket placed in a washtub to collect all of the overflow.  A flocculent was added to the bucket to aid in the settling of the sample.  All of the overflow material was washed back into the bucket to avoid any loss of material.  The single bucket procedure was

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utilized for the 2007 drill sample collection as well.  Wet RC samples are reviewed carefully for the possibility of down-hole contamination (see Section 14.1).

12.2    Core Samples

PQ (85.0 millimeter, or 3.345 inch) and PQ3 (83.1 millimeter, or 3.270 inch) diameter core holes were drilled at the Fort Knox deposit. Detailed log sheets prepared for core record lithology, mineral type, alteration, oxidation, color, vein percent, structure and vein orientations. Geotechnical data including RQD, core recovery, and fracture logs are also recorded. Hardness values ranging from R0 to R6 were documented beginning in 2005 and is ongoing.  The core is then photographed, logged, and sampled at 1.52-meter (5 foot) intervals.  MRDI (1991) indicated that diamond drill core has the potential to slightly understate the grade at Fort Knox.

12.3    Sample Composite Summary

Sample composites reflect the weighted average grade per hole over the bench height. Composite length can vary. Vertical holes show composite lengths of 9.1 m (30 feet = bench thickness) while inclined holes display longer composites, depending on their dip. Table 12-1 depicts the most important statistical parameters of the composite sample population, segregated by domains.

There are 15,721 composite samples of which 2,181 are in the schist and the remaining 13,540 in the combined domains of the intrusive (1677 in Domain 10, 5750 in Domain 20 and 6041 in Domain 30).

The reliability of the data is expressed by the coefficient of variance (CV = Standard deviation divided by the mean). The closer CV is to 1 the more dependable is the population of samples. CV is furthest away from 1 in the schist, indicating the highest variability of samples in this unit. Within the intrusive, Domain 20 has the lowest value, but the difference to the other intrusive domains (Domain 10 and 30) appears not to be of significance.

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Composites are used to assign grade values to the block model.  Because of statistical and field observations the grade of the schist is modeled as missing and is therefore not included in reserve calculations.

Table 12-1: Composite Summary Table

Statistic	Schist	Domain 10	Domain 10	Domain 10	All Samples
No. Samples	2,181	1,677	5,750	6,041	15,721
No. Missing Samples	119	28	101	66	314
Min. (opt)	0.000	0.000	0.000	0.000	0.000
Max. (opt)	0.205	0.215	0.248	0.239	0.248
Mean. (opt)	0.003	0.015	0.020	0.016	0.016
Std. Deviation	0.010	0.020	0.025	0.022	0.022
C.V.	3.278	1.288	1.249	1.313	1.415

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13.0    SAMPLE PREPARATION, ANALYSES AND SECURITY

Core and reverse circulation drill samples, which are the basis for all analytical determinations, are collected from the drill hole under the direct supervision of Company staff. The samples are labeled and placed in bags at the Company facility and prepared for transport to commercial laboratories for preparation and assay. Employees of the laboratory pick up drill samples at the Company facility. The RC cuttings are weighed, dried and reweighed. The sample is then crushed to minus 25.4 millimeters (1 inch) and a 1250-gram split is retained for air shipment to the analytical facility for assay. Once the 1250-gram split has been delivered to the laboratory it is pulverized to 80-mesh and passed through a riffle splitter to produce a 200 to 300 gram sample. This sample is then ring pulverized to minus 150-mesh, rolled, and a 50 gram sample is taken for gold determination by fire assay with an atomic absorption “finish”. For the majority of the drilling at Fort Knox, any assay exceeding 10.29 g/t Au would be re-assayed by fire assay with a gravimetric finish. The detection limit of these analytical methods is 0.001 oz Au/short ton from 1987 to 2002. The detection limit is 0.0003 oz Au/short ton from 2002 to present.

13.1    Quality Control, Quality Assurance

The objective of FGMI’s QA/QC Program is to ensure the accuracy and integrity of the data. FGMI submits samples to check assays as part of the company’s standard Quality Control, Quality Assurance Program (QA/QC). The QA/QC Program includes the submission of duplicates, blanks and standards at rates described in the following section.

13.2    Check Assays

From 1991 to 2005, to monitor the precision of the analytical process, FGMI had collected separate 1250-gram samples from every tenth sample collected. The even numbered samples, 20th, 40th, and 60th, were air freighted along with the regular samples to the primary lab. Every odd-numbered 10th, 30th, and 50th, sample was picked-up on a weekly basis by a secondary lab, pulverized, further split to 250 to 300 grams and assayed.  In addition, every 40th sample was re-assayed by the primary lab.

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In 2006, the even numbered samples, 20^th, 40^th, 60^th, etc., were pulverized by Alaska Assay Laboratories and then shipped to ALS Chemex for “secondary lab” check assaying.  Since a new primary lab was being used in the 2006 program, a more rigorous program of primary lab duplicate checks was instituted to monitor their performance.  Primary duplicates were selected randomly by the primary lab, and increased in frequency to an average of 1 duplicate per 7 samples.  The current verification practice for the primary lab is to re-assay the first and then every 20^th sample in each job.

Assay results are received electronically from Chemex, American, and Bondar-Clegg (prior to 2002) and Alaska Assay labs. Final certified “hard” copies of the assay certificates are delivered later and retained in the Company’s files.

FGMI’s technical staff monitors the results on a regular basis. Any skewed results are investigated to determine the cause of the difference. The authors consider the results of the past operating year to be within acceptable error levels.

13.3    Blanks Program

FGMI also inserts blank or unmineralized samples into each sample shipment as part of the operation’s standard procedures. Returned sample rejects that assay below the detection limit (<0.0003) are submitted with the regular RC samples.  The frequency of blank insertion has varied from a minimum of one per hole; to as many as one blank every 100 feet or 20th sample, depending on the drill season.  The old and the new sample numbers are recorded on the log sheet for reference purposes.  Inclusion of blank samples in core drilling occurred from 1997 to present, and used either reject material from RC holes, or material from Browns Hill Quarry in North Pole, Alaska.

FGMI’s technical staff carefully monitors the results of the submitted blanks to check for possible contamination during the analytical process. Any skewed results are investigated to determine the cause and the set of samples will be re-run to clarify the

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assays. The authors consider the results of the past operating year to be within acceptable error levels.

13.4    Field Duplicates Program

Historically, FGMI has collected duplicate samples at random from the RC drill sample population.  These duplicates were collected from the reject portion of the sample splitter and are used to monitor sample analysis precision. The samples were bagged and tagged consistent with FGMI’s normal sample submission practices so that the duplicates are indistinguishable from the normal sample population. The FGMI technical staff carefully reviewed these duplicate samples. Any samples falling outside expected tolerance levels are reviewed to determine the cause of the discrepancy. The authors consider the results of the duplicate analyses for the previous operating years to be within acceptable error levels.

13.5    Standards

FGMI also submits known standards with the unknown drill samples. Pulp samples of known grade are submitted to the laboratory to monitor analytical precision.

From 2001 to 2005, standards were submitted at a rate of two standards per core hole, and every 100 feet for reverse circulation holes. These standards, representing different ranges of gold grades, were prepared from Bondar-Clegg samples that were collected from True North material.  The samples and their values are as follows:

TNA    0.000 opt Au

TNB    0.006 opt Au

TNC    0.019 opt Au

TND    0.058 opt Au

TNE    0.081 opt Au

TNG    0.190 opt Au

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In 2006, five commercial standards were purchased for use with the ongoing drill program.  These standards were created and certified by Rocklabs Ltd., of Auckland, New Zealand.

The samples and their values are as follows:

RA    0.0302 opt Au

RB    0.0057 opt Au

RC    0.0538 opt Au

RD    0.0117 opt Au

RE    0.0699 opt Au

In 2007 additional Standards were purchased from GeoStats Pty Ltd, Australia with the following grades:

GA    0.0061 opt Au

GB    0.0120 opt Au

GC    0.0303 opt Au

GD    0.0513 opt Au

GE    0.0691 opt Au

Reverse circulation drilling inserts a standard at the first sample number ending in 11 and then every 20th sample afterwards (11, 31,51,71).

Core drilling inserts standards at samples ending in 1 and then every 20^th sample afterwards (1, 21,41,61,81)

FGMI’s technical staff monitors the performance of the standard samples submitted for analysis to ensure that the results lie within acceptable tolerance levels. Recent sample performance indicates no significant areas of concern. The authors of this report have reviewed the preparation of and results of the standards used since 2001 and have found them to be within acceptable error levels.

13.6    Sample Security

Core and reverse circulation drill samples, which are the basis for all analytical determinations, are collected from the drill hole under the direct supervision of Company staff. The samples are labeled and placed in bags at the Company facility and prepared for transport to commercial laboratories for preparation and assay. Employees of the laboratory pick up drill samples at the Company facility. The qualified person is satisfied that the security of the samples is adequate.

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14.0DATA VERIFICATION

All drill data, including collar locations, down-hole surveys, and data intervals are collated in Excel spreadsheets. Survey and log data are hand-entered into the spreadsheets, while assay data are downloaded in digital format from the assay lab then subsequently merged into the spreadsheets. Collar locations are initially recorded in the local grid by the GPS surveying system.  Collar locations and downhole surveys are also rigorously reviewed via manual verification methods for erroneous data by the technical staff of FGMI.  Following input of data into Excel spreadsheets, the data are output to a MineSight-ready ASCII file and subsequently imported into the MineSight software package. The resulting assay and survey files in the MineSight directory include:

Assay file is identified as knx011.07d and,

Survey file is named knx012.07d.

Gold grade and lithology data are entered into the “au1” and “lith” fields within the assay file.  The majority of the drillhole data has had a minimum of 10% double-entry checks performed for data validation purposes.

The processes of data verification as described above are acceptable due diligence routines to assure the reliability of the Fort Knox data set. The authors feel comfortable that the presented data reflect actual conditions of the in-situ situation.

14.1Drill Hole Data Rejection

Twelve RC holes displayed down-hole contamination when compared to adjacent holes, blast hole drilling, and their respective drill logs, and select intervals were rejected from the assay data set. Gold values for the rejected data were converted to -1 (as if missing). This process resulted in the rejection of 806 intervals. Prior to 2004, computer programs recording down hole “cyclicity” and  “decay”, as well as excessive sample weight, would be used as guides for geologists to reject suspect intervals as aberrant gold values.  As mining production has progressed and production blastholes have allowed for verification of the presence of gold, these rejected intervals have been reviewed and

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updated on an ongoing basis, resulting in some previously rejected intervals being restored for use in modeling.

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15.0 ADJACENT PROPERTIES

This region of Alaska has seen significant gold production but almost exclusively this production was derived from placer deposits. The Fort Knox mine is the first lode gold deposit to be developed in this region.  Therefore, adjacent properties are not deemed relevant for the purposes of this Technical Report.

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16.0MINERAL PROCESSING AND METALLURGICAL TESTING

16.1Fort Knox Mill Facility

16.1.1Milling Methods and Capacity

High grade ore from the Kinross Fort Knox deposit is processed at the Company’s CIP (carbon-in-pulp) mill located at the Fort Knox mine. True North (Figure 4-1) ore was processed at the mill from 2001 to 2004. The Fort Knox mill facility was constructed in 1995, the first doré produced in December 1996 and commercial production achieved in March 1997.

The Fort Knox mill has a daily capacity of between 33,658 to 45,359 tonnes (36,000 to 50,000 short tons) per day. The mill schedule for 2008 will process an average of 38,010 tonnes (41,900 short tons) per day.  Mill feed is first crushed to minus 20 centimetres (8 inches) in the primary gyratory crusher located near the Fort Knox pit and conveyed 800 metres (2,625 feet) to a coarse-ore stockpile located near the mill. The crushed material is conveyed to a semi-autogenous (SAG) mill, which operates in closed circuit with two ball mills and a bank of cyclones for sizing.  A portion of the cyclone underflow is screened and then directed to a gravity recovery circuit.

Correctly sized material flows into a high rate thickener and then into leach tanks where cyanide is used to dissolve the gold. Activated carbon is used in the CIP (carbon-in-pulp) circuit to adsorb the gold from the cyanide solution. Carbon particles loaded with gold are removed from the slurry by screening and are transferred to the gold recovery circuit where the gold is stripped from the carbon by a high temperature/pressure cyanide solution. Gold ions in the solution are then plated onto a cathode by electrowinning. The gold is then melted into doré bars for shipment to a refiner. Mill tailings are detoxified and sent into the tailings impoundment below the mill.

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16.1.2Mill Recovery – Fort Knox Deposit

Gold recoveries at the Fort Knox mill have historically ranged from 87 percent to more than 90 percent since production began in 1996. During 2001 to 2004 it was necessary to add lead nitrate and slightly increase the cyanide and lime concentrations to maintain mill recovery rates with some of the feed coming from True North.

The historical production at Fort Knox allows a gold grade to recovery relationship to be developed. Table 16-1 summarizes this relationship for the Fort Knox mine based on several years of production experience.  At the $550 per ounce A-ore, cutoff grade (0.0150 oz Au/short ton) the predicted recovery is 88.25 percent, while for the average grade for reserves scheduled to be milled (0.0225 oz Au/short ton) the predicted recovery is 89.36 percent.

Table 16-1: Grade Recovery Relationship - Fort Knox Mine

Grade oz/ton	0.010		0.015		0.020		0.025		0.030		0.035
Recovery	86.82	%	88.25	%	89.31	%	90.16	%	90.85	%	91.5	%
Grade oz/ton	0.040		0.045		0.050		0.055		0.060
Recovery	91.97	%	92.43	%	92.85	%	93.23	%	93.57	%

16.1.3Mill Recovery – True North Deposit

The True North mineral model has both a fire assay grade for in situ reserves and resources and a cyanide soluble (CN) grade for the estimation of recoverable gold. The cutoff grade and daily grade control are based on the cyanide soluble grade. A monthly comparison between the predicted CN grades in the mineralization model and the back-calculated head grade from the mill indicated that a 96 percent recovery factor is required to estimate the True North gold produced from 2001 to 2004.  True North ore is not currently being processed at Fort Knox.

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16.2Fort Knox Heap Leach Facility

16.2.1Heap Leach Methods and Capacity

The heap leach facility currently being constructed at Fort Knox includes a valley fill leach pad, carbon adsorption plant, piping, haul road construction, relocation of access roads, power lines, tailings, and water lines.  Run of mine ore is scheduled to be hauled from the pit and from existing stockpiles to the leach pad.  The pad design anticipates ore loading in 40-foot lifts, of which the top five feet will be minus 1 1/4 inch crushed ore, intended to insulate the drip emitters from freezing in colder months. The loading rate is expected to vary from around 30,000 tons per day of run of mine ore in the early years of the project, to roughly 50,000 tpd later in the project.  The heap leach pad is designed to be built in five stages, covering approximately 310 acres with a total capacity of 161 million tons.  The location of the heap leach pad is in the upper end of the Walter Creek drainage immediately upstream of the tailings storage facility (TSF).  The liner system currently designed for the Fort Knox heap leach pad consists of a triple liner in the pond area of the pad and a double liner for the remainder of the heap.

Leach solution is designed to flow through the run-of-mine ore into a 110 million-gallon in-heap storage reservoir.   The pregnant solution would be pumped at a rate of 8,000gal/minute (gpm) directly to a new carbon-in-column plant (CIC) located adjacent to the existing mill.  After the pregnant solution has been processed through the CIC plant, barren solution would be returned to the heap via a bank of five barren solution pumps.

Modifications to the existing crushing circuit have been carried out to produce the large quantities of overliner material and crushed material needed to protect the leach solution application system.  This project required some modifications to the inside of the mill building as well as the installation of two conveyors to transport crushed reject material to a truck load-out facility located adjacent to the main leach pad haulage road. Construction was completed in 2007.

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16.2.2Heap Leach Recovery – Fort Knox Deposit

Projected heap leach recoveries were based on laboratory testing of eight bulk samples which were representative of the material to be leached.  Gold recoveries from the ROM feeds ranged from 28% (samples 103A and 200A) to 93% (sample 205B).  Average gold recovery was 61.2% with an average of 224 days of leaching and rinsing.  Gold extraction was progressing from all eight feeds at a slow rate when leaching was terminated, though a longer leaching cycle would increase gold recoveries somewhat.  From this testwork, the following recovery relation was derived and used in mine planning:

Gold recovery at time t (days) = 0.65 - 1/ (0.03t + 1/0.65)

16.3Past Considerations

16.3.1True North

The True North deposit is hosted in altered metamorphic rocks, and generally crushes and grinds at a more rapid rate than material from Fort Knox. Blending of True North and Fort Knox mineralization increased the throughput of the mill over feeding Fort Knox mineralization alone.

The True North mineralization locally contains concentrations of stibnite, which has the potential to decrease the pH, lower the dissolved oxygen content, and increase the consumption of cyanide. To offset the detrimental attributes of True North ore the feed rates were limited and additional mill reagents were added.

16.3.2Fort Knox

Based on the varying hardness of the ore, crushing and grinding rates can change significantly within the deposit having a direct impact on mill throughput and ultimately unit costs. The mill tonnage rates are adjusted to reflect crushing properties based on where mining has been scheduled in the pit.

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17.0MINERAL RESOURCE AND RESERVE ESTIMATES

The 2007 Mineral Resource and Reserve Estimates have been estimated using MineSight, a mine modeling software program developed by Mintec Inc.  Klaus Triebel, P.Geo., David Quandt and Chris Ekstrom, who are employees of FGMI, completed the estimates.  Mineral reserve estimates were reviewed by Kevin Morris, P.Eng., Director of Mining Services, Kinross Gold Corporation.

17.1Mineral Resource and Reserve Statement

The mineral reserve and resource estimates prepared by FGMI personnel have been completed with consideration given to title, taxation, marketing, political, mining, permitting, environmental and socio-economic issues, as well as the amenability of the mineralization to be processed by the mill at the Fort Knox mine. There are no adverse environmental or social issues that would affect the Company’s ability to develop the reserves on its properties, nor the resources above a mill cutoff grade.  Conversion of leach-grade resources into reserves was completed, as all the permits for heap leach were received on or before October 31, 2007.

Table 17-1 summarizes the Proven and Probable Mineral Reserves for the Fort Knox mine December 31, 2007. Mineral reserves are estimated using an assumed gold price of $550 US per ounce and are classified in accordance with the CIM Standards on Mineral Resources and Reserves Definitions and Guidelines.

Table 17-2 summarizes the Measured and Indicated Mineral Resources for the Fort Knox mine and area as of December 31, 2007. Mineral resources are estimated using a gold price of $625 US per ounce and are also classified according to the CIM Standards. Mineral resources as stated are exclusive of mineral reserves.

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Table 17-1: Mineral Reserve Summary as of December 31, 2007

	Proven		Probable		Proven and Probable
	Tonnes	Grade	Tonnes	Grade	Tonnes	Grade
Source	(,000)	(g/t)	(,000)	(g/t)	(,000)	(g/t)
Mine	31,498	0.69	104,098	0.59	135,596	0.61
Stockpile	105,319	0.35			105,319	0.35
TOTAL	136,817	0.43	104,098	0.59	240,915	0.50

Table 17-2: Mineral Resource Summary as of December 31, 2007

	Measured		Indicated		Measured & Indicated
	Tonnes	Grade	Tonnes	Grade	Tonnes	Grade
Deposit	(,000)	(g/t)	(,000)	(g/t)	(,000)	(g/t)
Gil (80%)			2,838	1.03	2,838	1.03
True North			3,646	1.49	3,646	1.49
Fort Knox	2,664	0.67	26,643	0.59	29,307	0.60
Fort Knox Area	2,664	0.67	33,127	0.73	35,791	0.72

The Fort Knox area also hosts an Inferred Mineral Resource of 2.128 million tonnes averaging 0.55 g/tonne Au at the same $625 US per ounce gold price.  The majority (2.108 million tonnes) of this material is contained at the Fort Knox deposit.

17.2Fort Knox Deposit - Modeling Methodology

17.2.1Overview

A 2007 resource model for the Fort Knox project was built with Mintec’s MineSight modeling and mine planning software. Grades are projected from composited drill hole data (grade and other parameters are averaged over a certain set length from the original drill hole assays) into a 3-dimensional matrix of blocks that are sized appropriately for the mining methods in current use. Additional drilling from 2007 improved confidence levels in zones of formerly low drill density. Reconciliation incorporating blast hole data from nine years of mining with the current exploration drill program established the rejection of contaminated drill holes from the current model.

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17.2.2Survey Data

Pre-mine topography is based on airborne photogrammetric elevation data contoured at 1.52 meter (5 foot) intervals. The coordinate system at the mine is a local grid modified from State Plane coordinates. Mining progress is constantly updated by survey traverses of bench toes using a backpack GPS (global positioning system) system and periodically updated using airborne photogrammetric methods. The airborne topography was completed in August 1998 by AeroMap U.S, of Anchorage, Alaska, and updated in April 2003, 2005 and May of 2007.

17.2.3Assay Grade Capping

Original gold assays are capped at 0.50 oz Au/short ton (17.14 g/tonne), which displays a shift in the lognormal distribution of the database.  Two hundred eighty-three intervals were capped at 0.50 ounce per ton, representing 1.06% of all assay data above the 0.0125 ounce gold per short ton (.43 g/tonne) cutoff grade. These data are capped directly in the MineSight database file knx011.07d.

Figures 17-1 and 17-2 display the univariate statistics of the uncapped and capped assay data respectively for the Fort Knox deposit, excluding rejected data.

17.2.4Drill hole Compositing

Grade-capped assay intervals were combined into 30-foot bench composites stored at the mid-point as “AU1” items in the knx009.07d composite file.  Rejected assay intervals were excluded from the compositing.

Figure 17-3 presents the 2006 histogram frequency plot and statistics for the bench composite data for the Fort Knox deposit.

Composites were assigned a code corresponding to one of the four defined lithologic domains (see section 16.2.6).  A composite was given the code corresponding to a domain if more than half of its length fell within that domain.

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Figure 17-1: Histogram – 5ft Uncapped Exploration Data used in 2007 Modeling

Figure 17-2: Histogram – 5ft Capped Exploration Data used in 2007 Modeling

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Figure 17-3: Fort Knox Mine – Composite Sample Statistics

17.2.5Block Model

The 2007 Fort Knox mineral block model is based on a block size of 50 feet x 50 feet x 30 feet (x, y, z), which is appropriately sized for the mining equipment and methods utilized at the Fort Knox mine. The dimensions of the block matrix are: 5,000 to 13,400 feet East (168 columns), 6,400 to 13,200 feet North (136 rows), and 400 to 2,470 feet in elevation (69 levels).

The December 2007 pit topography is stored in the model as a “topo%” item. It is assigned to the model from a MineSight file type-13 gridded surface model stored as a “topog” item. The original pre-mine topography is stored as a “topo” item. The MineSight gridded surface file 13 is created from a triangulated surface.

The following is a list of the important field items in the 2007 Reserve model file named knx015.07d:

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·	AUID	ID2 interpolated gold grade using AU1 composite item
·	AUPP	AUID with inferred blocks assigned missing values
·	DIST	Distance to nearest composite
·	DOMI	Domain identifier (10=west pit, 20=mid pit, 30=east pit, 4=granite/schist contact)
·	MCOST	Mining cost of block in $/ton
·	NCMPS	Number of composites used to estimate block grade
·	NHOLE	Number of drill holes used to estimate block
·	MII	Classification (1=Measured, 2=Indicated, 3=Inferred)
·	RECOV	Mill metallurgical recovery as a percentage
·	TOPO%	Percentage of block below Dec 2006 mining topography
·	VAL2	Net block value

17.2.6Lithology

Lithology plays an important role in the distribution of mineralization at Fort Knox.  Most of the mineralization above the cutoff grade is confined to the granitic Fort Knox stock that has been subdivided into fine-grained granodiorite, medium granite and coarse-grained granite porphyry. The granodiorite, granite, granite porphyry subdivisions were not used in the current, knx015.07d model.

The gold mineralization is hosted by quartz-sericite filled fractures and veins within northwest striking and southwest dipping shear packages. The quartz-filled structures form three gradational northwest striking zones in the Fort Knox deposit. The mineralized structures have variable trends which average 131 degrees, -63 degree dip; 135 degrees, -32 degree dip; and 147 degrees, -62 degree dip, in the west pit, middle pit and east pit, respectively.  These three structural domains were digitized; triangulated and subsequent solid geometry features were developed. The west, middle, and east structural domains were assigned “material types” 10, 20, and 30, respectively. The solid geometry features were then loaded to the knx009.07d composite file using the “spear”

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tab in the MineSight drill hole view properties.  The solids were loaded to the “DOMI” item in the composite file and coded as 10, 20, and 30.

In the schist, gold mineralization appears to consist of narrow extensions of mineralized structures that transect the granite. In order to constrain grade interpolation between granite and schist, a solid was constructed from pre-existing sectional interpretations of the schist lithology. The subsequent schist solid was loaded to the knx009.07d composite and coded as 4 in the “DOMI” item.

17.2.7Domains

The west, middle, east, and schist domains described in the Lithology section above were developed to reflect orientations of gold-bearing veins observed in exploration and blast-hole data across the deposit.  These domain solids were loaded to the composite file as described above, then assigned to the “DOMI” item in the knx015.07d block model as a block limiting item in the interpolation. The block search parameters, ellipsoidal search distances and orientations are adjusted for each of the four domains. The domain boundaries were “soft” between the various granite domains, meaning that composites from one domain were allowed to interpolate blocks in other granite domains.  However, the domain boundaries between the schist and the various granite domains were “hard” boundaries, so composites coded as schist were not used in interpolating blocks coded as granite, and vice-versa.

17.2.8Variography

Variograms were developed, using the variography program (SAGE2001), for the west, middle and east domains to develop kriging and search parameters.  The variography consists of directional correlograms created using nested spherical models for each of the domains, calculated from blast hole data. The variogram parameters, applied to the krige models, were derived from structural orientations and helped determine the ellipsoidal search parameters used in the final ID(2) model that define the current mineral reserve and resource estimates contained in this report.  Tables 17-3 presents the variogram parameters.

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17.2.9Interpolation Methodology

The “07d” block model was interpolated from MineSight Compass Multi-Run “06b IDW” developed in 2006. Gold grade is projected into model blocks from 30-foot bench composites with a series of interpolation runs using an ID(2) algorithm. In the first set of interpolations, the granitic host is modeled in the west, middle and east domains. Gold is interpolated with 250x150x75-foot search radii for the granitic host in the west, middle, and east domains and 200x200x200 for the schist domain. In the second pass, which overwrites the first pass, gold is interpolated with 150x100x75-foot search radii for the granitic host in the west, middle, and east domains and 110x110x110-foot search radii for the schist domain. The domains were assigned “soft boundaries” between the granite domains, allowing interpolation across the domain borders. However, the schist-granite domain boundaries were “hard”, and so composites in the granite were not used in interpolation of schist blocks, and vice-versa. The number of composites for a block is a minimum of 3 and maximum of 8. The maximum number of composites per hole is 4.  The maximum 3-D search distance for composites greater than or equal to 0.300 ounce per ton gold was set at 25 feet.

Schist blocks had their gold grades assigned a zero value.  Operational experience has shown that the very small zones of mineralized schist are difficult to identify during ore control.  Therefore, it was considered appropriate to discard from the estimate any ore contained in the schist, which represented approximately 1% of the total ore.

Drilling in 2007 added 41 holes to the 2007 “07d” model bringing the total number of holes to 751 versus the 710 utilized in the 2006 “06b” model. The search parameters and variography generated for the 06b were utilized for the 07d model.  Figure 17-4 presents the 2006 histogram frequency plot and statistical summary of the interpolated block model grades for item “AUID” in the “07d” model.

17.2.10Search Parameters

FGMI geologists and resource modelers established the ellipsoidal search directions based on blast hole and structural data in December 2004. The ellipsoidal search

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distances used in composite selection for the interpolation are based on blast hole variography established in April 2004.  The search parameter radii restrict the extrapolation of mineralization in areas of wide spaced drilling. As described in the Lithology, Domains, and Interpolation sections above, gold is interpolated with 250x150x75-foot search radii for the granitic host in west, middle, and east structural domains. Gold is interpolated with 200x200x200-foot search radii for the schist lithology domain. In the second pass, which overwrites the first pass interpolation, gold is interpolated with 150x100x75-foot search radii for the granitic host in west, middle, and east structural domains and with 110x110x110-foot search radii for the schist lithology domain.  Despite the differences in variography ranges in the three granitic-structural domains, identical major, semi-major, and minor axis search distances are being used; only their orientations differ from domain to domain.  However, local comparisons of the resource and reserve model with the blast hole model using variable search radii shows that the current search radii display a good fit to the data.

17.2.11Specific Gravity - Tonnage Factors

The dry tonnage factor used at Fort Knox is a global 12.5 ft³/ton and is based on analyses of samples using 25 immersion tests (FGMI), 33 tests without immersion (Golder Associates Inc. 1995), 99 pycnometer tests (Bondar-Clegg), and one volumetric determination of a bulk sample supplied by FGMI.

17.2.12Dilution

Although dilution was not added directly to the modeling as a percentage, it was indirectly incorporated into the modeling by use of appropriate interpolation parameters.  These parameters were selected so that the modeled ore was a reasonable fit to the actual A-ore and B-ore production to date (see Table 17-5).

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Table 17- 3: Variogram Parameters

First-Pass Search Parameters

Domain	West	Middle	East	Schist
Search Ellipse
Ellipse major	250	250	250	200
Ellipse minor	150	150	150	200
Ellipse vertical	75	75	75	200
Major Azimuth	131	135	147	0
Plunge at Major Azimuth	0	0	0	0
Dip at Major Azimuth	-63	-32	-62	0
Max. dist to nearest hole	250	250	250	200
Min. # composites	3	3	3	3
Max. # composites	8	8	8	8
Max. # comps/hole	4	4	4	4
Variogram
Range: 1st Var major	16.0	24.5	39.8	20.6
Range: 1st Var minor	32.5	26	63	21.9
Range: 1st Var vertical	56.4	31	24.9	24
Direction Major Axis	46	-22	-21	40
Plunge Major Axis	-15	16	19	24
Dip	-2	26	58	12
Range: 2nd Var major	506	583.3	394.5	150.7
Range: 2nd Var minor	226.4	662.4	280.7	408.9
Range: 2nd Var vertical	117.3	138.9	893.3	200.3
Direction Major Axis	-29	-21	18	-38
Plunge Major Axis	26	19	46	0
Dip	55	58	-65	-24
Nugget	0.417	0.350	0.540	0.500
Sill 1st structure	0.452	0.503	0.364	0.429
Sill 2nd structure	0.131	0.147	0.096	0.071

2nd-Pass Search Parameters

Domain	West	Middle	East	Schist
Search Ellipse
Ellipse major	150	150	150	110
Ellipse minor	100	100	100	110
Ellipse vertical	75	75	75	110
Major Azimuth	131	135	147	0
Plunge at Major Azimuth	0	0	0	0
Dip at Major Azimuth	-63	-32	-62	0
Max. dist to nearest hole	150	150	150	110
Min. # composites	3	3	3	3
Max. # composites	8	8	8	8
Max. # comps/hole	4	4	4	4
Variogram
Range: 1st Var major	16.0	24.5	39.8	20.6
Range: 1st Var minor	32.5	26	63	21.9
Range: 1st Var vertical	56.4	31	24.9	24
Direction Major Axis	46	-22	-21	40
Plunge Major Axis	-15	16	19	24
Dip	-2	26	58	12
Range: 2nd Var major	506	583.3	394.5	150.7
Range: 2nd Var minor	226.4	662.4	280.7	408.9
Range: 2nd Var vertical	117.3	138.9	893.3	200.3
Direction Major Axis	-29	-21	18	-38
Plunge Major Axis	26	19	46	0
Dip	55	58	-65	-24
Nugget	0.417	0.350	0.540	0.500
Sill 1st structure	0.452	0.503	0.364	0.429
Sill 2nd structure	0.131	0.147	0.096	0.071

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17.2.13Metallurgical Recovery

Metallurgical recovery is calculated for each model block and is based on a grade-recovery curve discussed in Section 16, Table 16-1. The following formula is applied to determine the recovery for ore sent to the mill:

Recovery % = (0.0411 * LN (Grade + 0.002) + 1.046) * 100

The following formula is applied to determine the recovery for ore sent to the heap leach pad:

Gold recovery at time t (days) = 0.65 - 1/ (0.03t + 1/0.65)

17.2.14Model Checks

The interpolated gold grades in the block model are checked in detail against both exploration and blasthole production drilling on both plan and sectional views using the MineSight software visualization tools. This checking has led to refinements of interpolation parameters to obtain a better fit with grade patterns in drilling data. The model and production reconciliation is also discussed in Section 17.4.

17.2.15Resource and Reserve Classification

Measured, Indicated, Inferred Resources

All model blocks with interpolated grade are assigned a confidence category identifier. The “anisotropically adjusted” distance to the nearest hole, number of composites and number of holes is stored during the interpolation using anisotropic ratios corresponding to the search ellipsoids.  Only one drill hole was required to interpolate a block, and therefore, the variables used to distinguish potential “Measured” blocks from “Indicated” block were the anisotropically adjusted distance to the nearest drill hole, and the number of holes in the interpolation. The variable used to distinguish potential “Indicated” blocks from “Inferred” blocks was the anisotropically adjusted distance to the nearest drill hole. Resources are also limited by gold price and are exclusive of reserves.  Distance requirements to the nearest sample for each domain are listed in Table 17-4.

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Table 17-4: Classification Parameter Criteria

Category	West	Middle	East	# Holes
Measured	100 ft.	100 ft.	100 ft.	3
Indicated	200 ft.	200 ft.	200 ft.	1
Inferred	250 ft.	250 ft.	250 ft.	1

Proven and Probable Reserves

All “Measured” and “Indicated” blocks, using an economic cutoff based on $550 Au, and contained within the $ 525 ultimate pit design, are converted to “Proven” and “Probable” reserves, respectively.

17.3Pit Optimization

Because each block in the model has a unique recovery, pit optimizations are based upon a net dollar value rather than simply block grade. The steps to calculate the net value in the block model are as follows:

·Mining cost is variable by depth and is applied to each block in the field MCOST;

·Revenue is calculated for all “Measured” and “Indicated” blocks and is set in the block model item REVT. REVT = (gold grade * tons * (recov/100) * gold price) - (ore costs * tons);

·If REVT is positive, it is assumed the block will be processed and MCOST is subtracted from REVT. VAL2 = REVT – MCOST;

·If REVT is negative, the block is assumed to be waste and is assigned just the negative mining cost. VAL2 = MCOST * -1;

·All other blocks including “Inferred” and waste blocks are assigned the negative mining cost.  VAL2 = MCOST * - -1.

Mining costs (MCOST) are variable between $1.10 and $1.44 based on increased pit depth and haul length from the pit rim. The “ore” costs include: milling - $4.118/ton, leach operating costs - $0.98/ton, leach incremental capital - $0.20/ton, G&A taxes etc - $57.31 /oz for mill ore and $20/oz for leach ore.

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Pit optimization slope angles are based on inter-ramp angles from studies by two firms Golder Associates and Call & Nicholas, Inc.  Due to the “flat bottom” and central geometry of the Fort Knox orebody allowing for unique ramp access opportunities, overall slopes in the pit optimization were not flattened beyond inter-ramp angles. A complex slope model was developed utilizing the location of the granite-schist contact. Slopes used in the schist were 37º on the south wall, 42º in the southwest corner, 40º on the north and northwest walls, and 45º on the east wall.  Slopes in the underlying granite were set at 45º. The 37º and 45º angles on the south wall schist and granite, respectively, are not an exact match to the design strategy suggested by Call & Nicholas; however, this would not influence the pit optimization, since the south wall was constrained by practical mining pushback limitations.

Previous years’ optimizations were limited by NOAA’s land holdings on the west side of the pit.  Negotiations with NOAA to obtain a westward extension of FGMI’s land holdings were completed in 2007 allowing the inclusion of additional material as mineral reserve and mineral resource.

Using the MineSight software package, a Lerchs-Grossmann pit optimization was run at $550 gold price using no discount rate.  A check was performed by Kinross Technical Services using Whittle software and methodology, and the pit shell with the optimal NPV was found to vary in total tonnage from the $550 MineSight pit by less than 1.7%.   A final pit was then designed using the optimized pit outline as a guide, and closely follows the $525 per ounce pit optimization on the north and east sides.  A very large variance exists between the optimization and the pit design, due to size limitations of practical pushbacks on the southeast and east sides, and land position restrictions on the west side.

17.4Production Reconciliation

The current mineral model for Fort Knox was reconciled to actual mining to provide an indication as to how well the current resource and reserve model predicts the tonnage and grade of the mineralization. Table 17-5 displays the reconciliation of the A-ore model

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(predicted) to A-ore mined (actual) at the Fort Knox Mine for the 2007 operating year. The reconciliation is based on grade control data using 0.0150 opt Au and 0.0125 opt Au cutoffs for A-ore and B-ore, respectively.

Table 17-5: Ore Reconciliation for 2007

	A-Ore						B-Ore						Total Mill Ore
	Tonnes		Grade		Gold		Tonnes		Grade		Gold		Tonnes		Grade		Gold
Source Model	(‘000’s)		(g/t)		(oz)		(‘000’s)		(g/t)		(oz)		(‘000’s)		(g/t)		(oz)
Resource	9,607		0.98		303,087		2,224		0.46		32,680		11,831		0.88		335,767
Ore Control	10,169		1.06		345,770		1,951		0.49		30,591		12,119		0.97		376,361
% Difference	5.9	%	7.8	%	14.1	%	-12.3	%	6.7	%	-6.4	%	2.4	%	9.4	%	12.1	%

The reconciliation showed that the mineral model estimated total ore tones, grade and ounces to be less (2.4%, 9.4% and 12.1% respectively) than the ore control model revealed. Over the two-year period consisting of 2006 and 2007 (not depicted), the mineral model underestimated total ounces by 4.7%.

The qualified person has no knowledge of any environmental, permitting, legal, ownership, taxation, political or other relevant issue that would materially affect the current mineral resource and mineral reserve estimates.

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18.0OTHER RELEVANT DATA AND INFORMATION

There is no other data or information relevant to the project that is not covered in other sections of this report.

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19.0INTERPRETATIONS AND CONCLUSIONS

The Mineral Resource and Reserve Estimates for the Fort Knox mine have been prepared under the supervision of R. Henderson, Vice President of Technical Services.

The authors are satisfied the data used in the estimation of Mineral Resources and Reserves is free of gross errors and omissions and is of suitable quality and sufficient quantity to use in estimating resources and reserves for the Fort Knox deposits.

The economic analysis and cash flow forecast is positive, demonstrating that the mineral reserves are economic.  The authors are not aware of any issues that have not been otherwise disclosed in this Report, which would materially affect the current estimate of the mineral reserves for the Fort Knox mine.

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20.0RECOMMENDATIONS

The Fort Knox mine is a producing gold mine for Kinross Gold Corporation.  The recommendation for successive work phases is not required.

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21.0REFERENCES

Bakke, A., 1994, The Fort Knox ‘Porphyry Gold Deposit- Structurally controlled stockwork and shear quartz vein, sulphide-poor mineralization hosted by a Late Cretaceous Pluton, East Central Alaska, Canadian Institute of Mining Special Volume 46.

Bakke, A., Morrell, B., and Odden J. (1998), The Fort Knox porphyry gold deposit, east-central Alaska: An overview and update. Porphyry and Hydrothermal Copper and Gold Deposits. Australian Mineral Foundation Conference Proceedings November, 1998. p. 89-98.Blum, J.D., 1982: The Petrology, Geochemistry, and Isotope Geochronology of the Gilmore Dome and San Pedro Dome Plutons, Fairbanks District, Alaska, Unpublished M.S. Thesis, University of Alaska Fairbanks pp1-107.

Bundtzen, T.K., 1981: The Geology and Mineral Deposits of the Kantishna Hills, Mt. McKinley Quadrangle, Alaska, Unpublished M.S. Thesis, University of Alaska Fairbanks, pp1-237.

Call & Nicholas, Inc, 2005, Fort Knox and Gil Mines – Site Visit and Slope Stability Review.

Ekstrom, C., Jan. 2007: 2006 Fort Knox Dump Evaluation Summary, Internal Memo, pp1-3.

Fairbanks Gold Mining, Inc, and Kinross Gold USA, 2007, Fort Knox Mine Walter Creek Heap Leach Project: Project Feasibility Evaluation, pp1-82.

Golder Associates, Inc., 1995 (c): Pit Slope Stability Investigation, Fort Knox Project;

Golder Associates, Inc., 2002 (a): Pit Slope Stability Study, Proposed 715 Pushback, Fort Knox Project;

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Golder Associates, Inc., 2002 (b): Review of North Wall Failures, January, 2002 site visit to Fort Knox Mine;

Golder Associates, Inc., 2002 (c): Slope Design and Controlled Blasting Criteria, Granite Slopes in the Fort Knox Pit;

Golder Associates, Inc., 2000: True North Waste Dump Site Investigation and slope Stability Evaluations.

Golder Associates, Inc., Scoping -Level Pit Slope Study, True North Project Hall, M.H., 1985: Structural Geology of the Fairbanks Mining District, Central Alaska, Unpublished M.S. Thesis, University of Alaska Fairbanks pp1-68.

Hansen, Erik, 2008: Mother of All Lists, 10th edition pp. 1-99.

Heiner, L.A. and Wolff, E.N., 1968, Final Report Mineral Resources of Northern Alaska: Mineral Industry Research Laboratory No. 16, University of Alaska.

Henderson, Rob, 2006, Fort Knox Heap Leach Recovery, internal memo.

Metz, Paul A., 1991: Metallogeny of the Fairbanks Mining District, Alaska and Adjacent Areas.

Mineral Industry Research Laboratory, School of Mineral Engineering, University of Alaska Fairbanks, MIRL Report No. 90;

Mineral Resources Development, Inc., 1991 (a): Fort Knox Gold Project Fairbanks, Alaska; Reserve Study, Volume 1;

Mineral Resources Development, Inc., 1991 (b): Fort Knox Gold Project Fairbanks, Alaska; Reserve Update;

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Mineral Resources Development, Inc., 1992: Fort Knox Gold Project, Alaska 1992 Reserve Update;

Mineral Resources Development, Inc., 1998: Reserve Update, Fort Knox Project;

Mineral Resource Development, Inc., 2001: 2001 Resource Model Review, True North Project.

Morrison, D. and Cicchini, P, 2005: Fort Knox and Gil Mines Site Visit and Slope Stability Review, Call & Nicholas, Inc.

Murphy, John M. and Bakke, Arne, 1993, Low-temperature thermal history of the Gilmore Dome area, Fairbanks Mining District, Alaska, Canadian Journal of Earth Sciences, pp.764-768.

Robinson, M.S., Smith, T.E., and Metz, P.A. 1990: Bedrock Geology of the Fairbanks Mining District, Alaska Division of Geological and Geophysical Surveys, Professional Report 106.

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22.0DATE AND SIGNATURE PAGE

The undersigned prepared this Technical report, titled NI-43-101 Technical Report for the Fort Knox Mine, dated 31 March 2008.  The format and content of the report are intended to conform to Form 43-101F1 of the National Instrument (NI 43-101) of the Canadian Securities Administrators.

Signed and Sealed

Robert D. Henderson 31 March 2008

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23.0ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION PROPERTIES

23.1Fort Knox Mining Operations

The Fort Knox mine includes the main Fort Knox open pit mine, mill, heap leach and tailings storage facility. The mine production rate in 2007 varied between 110,000 and 146,000 tonnes (121,000 to 161,000 short tons) per day of total material, averaging 126,400 tonnes per day in 2007 (139,400 short tons).  Open pit mining is carried out on a year round basis, seven days a week. Standard drilling and blasting techniques are used, and the blast holes are sampled and assayed for production grade control purposes. Broken rock is loaded with a shovel or a wheel loader into haul trucks. Depending on the grade control results, the mined material is delivered to either the primary crusher, low-grade stockpiles, or to waste rock dumps.

23.1.1Mine Equipment

The following major mining equipment is budgeted for use at Fort Knox in 2008:

·6 Blast hole drills;

·9 Caterpillar 785 haul trucks (150 short ton capacity)

·8 Caterpillar 789 haul trucks (190 short ton capacity)

·6 Caterpillar 793 haul truck (240 short ton capacity)

·2 Hitachi EX 3600 shovels

·1 Hitachi EX 5500 shovel

·1 Caterpillar 994 loader

·1 Caterpillar 992 loader

·3 Caterpillar D10 dozers

·2 Caterpillar 834B rubber-tired dozers

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·3 Caterpillar graders (1 x 16G, 2 x 16H)

·2 Water trucks

Other equipment includes: assorted support equipment and facilities, truck shop, wash bay, fueling facilities and site office.

23.1.2Mine Life Plan

FGMI’s Technical Department updates a detailed Life of Mine Plan (LMP) based on the available reserves and utilizing the available mining fleet. Truck productivity predictions (TPH and Cycle times) are adjusted to reflect production experience. The proposed haulage profiles are determined for each bench and pit phase, plus the four possible destinations (crusher, leach pad, low grade stockpile, and waste dumps). Using this data, the truck hours required to move the scheduled tonnage are limited to the actual producing truck hours from the mine fleet (after availability and utilization).

Production scheduling is driven by:

·Mill feed requirements and leach feed capacities;

·Stripping requirements for Phase-6 and Phase 7;

·Loading rates constraints, where only one shovel or loader can operate;

·Haulage limits determined by the size and number of trucks and their productivity.

The LMP uses the following pit design criteria:

·30 foot benches;

·Single benching;

·Variable Slope angles based on studies by Golder Associates, and Call & Nicholas Inc.;

·120-foot wide haul roads typically 8 percent grade, with some 10 percent segments near the bottom of the pit design;

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·All designs utilized toe and crest contours.

The ultimate pit was designed from a US $525 per ounce pit optimization, which was limited by practical pushback size limitations on the south side.

On the south wall, the zone of alteration 2-3 benches above and below the schist-granite contact has been identified as a zone that requires more conservative slopes.  In these areas, design includes slope angles as low as 32°, as well as 90 ft stepout berms up to 130 ft above the contact.  General design guidelines were documented by geotechnical consultants Call & Nicholas (“CNI”) in October 2006, and FGMI technical staff have worked closely with CNI in areas that require local design exceptions.

23.1.3Other Considerations

In 1996, a 1.3 million short ton slope failure developed in the central south wall above the granite-schist contact. The slide was stabilized with mining set backs at the toe of the failure. Golder Associates Inc. was not able to develop a predictive slope stability model, because of the complex structure in the area, so a range of potential slopes was recommended.  At a design angle of 28º Golder predicted little chance of failure and, at a design angle 35º, a high chance of failure was predicted.  In 2005 Call and Nicholas Inc. was retained to develop a recommendation for the slope failure area, with continuing work resulting in design parameters described briefly in the section above. Ground water was believed to be a contributing factor to the failure, and a dewatering program is in place.

The dewatering department drilled two dewatering wells in 2007.  The wells target water in advancing areas of mining.  There are a total of twenty-one groundwater dewatering wells, with a potential pumping capacity of 1,470 gallons per minute (gpm). Eighteen of the wells were active in and around the pit, as well as one well north of the pit, near the causeway. The average annual pumping rate for wells in the pit area was 355 gpm, with an additional 650 gpm from the well north of the pit, for a total average dewatering rate of approximately 1005 gpm for 2007.  The total pumping rate shows a 35% decrease in

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production when compared to 2006’s dewatering rate of 546 gpm. Mining activities in 2007 were concentrated in the Phase 6 portion of the pit resulting in a loss of access to the bottom of the pit for extended periods of time.  Low water levels in the Fish Creek causeway wells north of the pit resulted in shutting down one of the wells and operating the second at the static influx rate.

Haul road access to Phase-6 will be from the northern end. With the expansion of the Phase 6 design to the north, the mine plan requires the subdivision of Phase 6 into north and south mining areas to balance the strip ratio.  The first significant mill-grade ore will be reached on north Phase 6 on the 1360 bench in the second half of 2008.  To reach the 1360 bench on the north side, roughly 13.2Mt of waste and 1Mt of leach ore will be stripped at an average rate of 44,000 tonnes per day.  Fully sustainable mill feed from the north side of Phase 6 is expected by the 1240 bench.

Due to the lowering of the economic cutoff grade by heap leach economics, certain material that had been hauled to waste dumps in the past has been reclassified as Stockpile reserves.  Blasthole modeling indicates that the granite waste in the Barnes Creek and Fish Creek waste dumps grades 0.309g/t.  In 2006, 65 rotary drillholes were completed on the Barnes Creek and 40 in Fish Creek waste dump, and confirmed the blasthole grade estimates for this material.  As a result, 30.085 ktonnes of material in Fish Creek grading 0.309g/t have been added to the “Stockpile” category of the Proven and Probable Reserves for year-end 2007.

23.2Recoverability

Metallurgical recovery information is discussed in detail in Section 16 of this report.

23.3Markets

Kinross will continue with current marketing arrangements at Fort Knox and does not envision any concerns related to marketing doré.

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23.4Contracts

Smelting, refining, handling and sales charges are within industry standards.

23.5Environmental Considerations

Fort Knox has a letter of credit in place that includes all of the following:

Fort Knox Mine – Reclamation/Closure	$	34,314,418
Fort Knox Post Reclamation Obligations	$	734,537
True North – Reclamation Plan	$	1,155,774
True North Access Road Reclamation	$	80,000
True North Plan of Operations	$	1,301,100

Very little concurrent reclamation can be initiated at Fort Knox since most all areas of disturbance are subject to continuing disturbance.  Borrow areas for the construction of the tailing dam have been reclaimed.

At True North, all acreage that would not be subject to further disturbance if mining commences has been reclaimed.

23.6Taxes

The State of Alaska Millsite Lease carries a 3% production royalty, based on net income and recovery of the initial capital investment.  Mineral production from State mining claims is subject to a Mine License Tax, following a three-year grace period after production commences.  The license tax ranges from 3% to 7% of taxable income.  There has been no production from State claims situated outside the boundaries of the Millsite Lease at the Fort Knox mine.  The unpatented federal lode claim is owned by Kinross and is not currently subject to any royalty provisions. Refer to Hansen, 2008 “Mother of All Lists” 11^th edition, 1/2008 for a detailed listing of all Fort Knox Title and Ownership documents, CD1. As a result of high metal prices, Kinross royalties and production taxes are estimated at $3.3 million for 2007 compared to $2.2 million in 2006.

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23.7Capital and Operating Cost Estimates

A detailed 2008 and LOM budget has been developed for the mine based on an actual operating cost history and projected future performance.  Mining costs are similar to most mines of this size and equipment age and are presented in Table 23-1.  The capital spending is shown for the LOM in Table 23-2.

Table 23-1: Estimated Operating Costs – Fort Knox

Time Period	2008 Budget		LOM Plan
Ktons Mined	49,723		305,171
Ktons Milled	15,291		103,632
Ktons to Leach Pad			162,037
Mining ($/ton) (incl. Maintenance)	$	1.41	$	1.26
Milling ($/ton)	$	4.13	$	4.08
Leaching ($/ton)	$		1.08
General and Administrative ($ per recoverable oz.)	$	75.57	$	69.05

Table 23-2: Fort Knox Life of Mine Capital Spending

Capital	Total		2008		2009		2010		2011		2012		2013		2014		2015
Mine/MEM/Mill/Sustain	$	193,447	$	65,424	$	41,341	$	28,109	$	34,158	$	24,181	$	235	$	0	$	0
Heap Leach	$	101,116	$	23,080	$	42,979	$	0	$	11,477	$	0	$	10,476	$	0	$	13,104
Total	$	294,563	$	88,504	$	84,320	$	28,109	$	45,635	$	24,181	$	10,711	$	0	$	13,104

Budgeted costs for 2008 reflect mining rates = associated costs that assume no True North material being used to supplement the Fort Knox mill feed. In 2014, pit mining will have ceased but stockpile material is expected to continue to be hauled to the leach pad.

23.8Economic Analyses

A profitability estimate was calculated for each mining area that was included in the Fort Knox reserves. Revenue was determined by using the mill grade-recovery relationship and a gold price of US $550 per ounce. Operating costs are based on the expenses discussed above plus added expenditures for capital items, such as tailings capacity, reclamation and equipment needed to sustain operations.  The discounted cash flow analysis indicates that the mine generates a positive cash flow.  The details of the

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economic and sensitivity analyses are considered by Kinross to be confidential information. The economic models are considered as complete, reasonable, and meeting generally accepted industry standards. Individuals seeking to review these models must request and sign a confidentiality agreement with Kinross Gold Corporation.

23.9Payback

The payback on the LMP was evaluated using both the capital for heap leach and Phase 7 expansion capital and indicates 5.8 years to return profits on the project.  The heap leach project accomplishes a positive return in 4 years.

23.10Mine Life and Production Schedule

FGMI, as part of the LMP, has completed a production schedule that depletes the currently known reserves at Fort Knox. The current FGMI LMP production schedule includes no material from True North. Production is derived entirely from the Fort Knox deposit from 2007 until 2014, at which time continuing gold production is expected to occur from the shipping of stockpile material to the leach pad.  Table 23-3 presents the estimated production schedule for the Fort Knox deposit.

Table 23-3: Fort Knox Life of Mine Plan Production Schedule

Source	Total	2008	2009	2010	2011	2012
Leach Feed
Tons (‘000’s)	162,037	0	3,970	14,600	14,600	18,250
Contained Grade (opt)	0.0095	0	0.0107	0.0109	0.0113	0.0092
Contained Ounces	1,544,738	0	42,365	159,390	165,358	168,345
Mill Feed
Tons (‘000’s)	103,632	15,291	14,918	14,821	14,850	14,850
Contained Grade	0.0223	0.0238	0.0221	0.0221	0.0191	0.0178
Contained Ounces	2,310,707	364,031	329,983	328,260	283,716	264,968

Source	2013	2014	2015	2016	2017	2108	2019
Leach Feed
Tons (‘000’s)	18,300	18,250	18,250	18,250	18,300	18,251	1,016
Contained Grade (opt)	0.0092	0.0092	0.0091	0.0091	0.0091	0.0091	0.0091
Contained Ounces	167,884	167,314	166,095	166,095	166,550	166,095	9,248
Mill Feed
Tons (‘000’s)	14,850	14,052	0	0	0	0	0
Contained Grade	0.0245	0.0268	0	0	0	0	0
Contained Ounces	363,833	375,917	0	0	0	0	0

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EX-99.3

EXHIBIT 99.3

CONSENT OF EXPERT

March 31, 2008

I hereby consent to the inclusion of the Technical Report for Fort Knox Gold Mine dated March 31, 2008 in the report on Form 6-K dated March 31, 2008 to be filed by Kinross Gold Corporation.

I also hereby consent to the incorporation by reference of the Fort Knox Gold Mine technical report included with the report on Form 6-K, into the Registration Statements on Form S-8 (Registration Statement Nos. 333-141896, 333-110208, 333-05776, 033-93926, 033-82450, 333-08936, 333-09004, 333-12662, 333-13744 and 333-13742) of Kinross Gold Corporation.

Sincerely,

/s/ Robert Henderson, P. Eng

Robert Henderson, P. Eng.

EX-99.4

EXHIBIT 99.4

CONSENT OF EXPERT

March 31, 2008

I hereby consent to the inclusion of the Technical Report for Maricunga Gold Mine, effective date December 31, 2007 in the report on Form 6-K dated March 31, 2008 to be filed by Kinross Gold Corporation.

I also hereby consent to the incorporation by reference of the Maricunga Gold Mine technical report included with the report on Form 6-K, into the Registration Statements on Form S-8 (Registration Statement Nos. 333-141896, 333-110208, 333-05776, 033-93926, 033-82450, 333-08936, 333-09004, 333-12662, 333-13744 and 333-13742) of Kinross Gold Corporation.

Sincerely,

/s/Maryse Belanger, P. Geo

Maryse Belanger, P. Geo