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 April, 2006

                        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_____ 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):_____

        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):_____

        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_____ No__X__


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

- -------



        This report on Form 6-K is being furnished for the sole purpose of
providing a copy of the Paracatu Mine Technical Report dated March 30, 2006
prepared for the mine located in Paracatu, Minas Gerais State, Brazil in support
of the December 31, 2005, resource and reserve disclosure of Kinross Gold
Corporation with respect to the Paracatu mine.


                                  EXHIBIT INDEX

99.1    Paracatu Mine Technical Report dated March 30, 2006.




                                        2


                                   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



                                        By  /s/ Shelley M. Riley
                                           -------------------------------------
                                           Shelley M. Riley, Corporate Secretary


Date:  March 31, 2005




                                        3

EX-99.1

                                 [KINROSS LOGO}









                         PARACATU MINE TECHNICAL REPORT





        PARACATU, MINAS GERAIS STATE, BRAZIL




        Prepared by:

        W. Hanson P.Geo

        Vice President, Technical Services

        Kinross Gold Corporation

        March 30, 2006






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1.0      EXECUTIVE SUMMARY......................................................................................1-1

   1.1     INTRODUCTION.........................................................................................1-1
   1.2     KEY METHODOLOGY CHANGES..............................................................................1-2
      1.2.1      DRILL HOLE SPACING AND RESOURCE CLASSIFICATION.................................................1-2
      1.2.2      SAMPLE PREPARATION AND ANALYSIS................................................................1-3
      1.2.3      GEOLOGICAL INTERPRETATION......................................................................1-3
      1.2.4      ORE HARDNESS...................................................................................1-4
      1.2.5      METALLURGICAL RECOVERY.........................................................................1-5
      1.2.6      BENCH HEIGHT...................................................................................1-5
      1.2.7      RESOURCE MODEL OPTIMIZATION....................................................................1-6
   1.3     DESCRIPTION AND LOCATION.............................................................................1-6
   1.4     ACCESSIBILITY CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY..............................1-7
   1.5     PROJECT HISTORY......................................................................................1-8
   1.6     GEOLOGY..............................................................................................1-9
   1.7     DEPOSIT TYPE........................................................................................1-11
   1.8     MINERALIZATION......................................................................................1-12
   1.9     EXPLORATION.........................................................................................1-12
   1.10    DRILLING............................................................................................1-13
   1.11    SAMPLING METHOD AND APPROACH........................................................................1-14
   1.12    SAMPLE PREPARATION, ANALYSIS AND SECURITY...........................................................1-14
   1.13    DATA VERIFICATION...................................................................................1-16
   1.14    ADJACENT PROPERTIES.................................................................................1-16
   1.15    MINERAL PROCESSING AND METALLURGICAL TESTING........................................................1-17
   1.16    MINERAL RESOURCE AND RESERVE ESTIMATE...............................................................1-18
   1.17    CONCLUSIONS.........................................................................................1-23
   1.18    RECOMMENDATIONS.....................................................................................1-24

2.0      INTRODUCTION AND TERMS OF REFERENCE....................................................................2-1


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   2.1     INTRODUCTION.........................................................................................2-1
   2.2     TERMS OF REFERENCE...................................................................................2-1
   2.3     GLOSSARY.............................................................................................2-2
   2.4     SCOPE AND OBJECTIVES.................................................................................2-3
   2.5     REPORT BASIS.........................................................................................2-3
   2.6     INDEPENDENT THIRD PARTY PARTICIPANTS.................................................................2-3
   2.7     STUDY PARTICIPANTS...................................................................................2-4
   2.8     DISCLAIMER...........................................................................................2-5

3.0      PROPERTY DESCRIPTION AND LOCATION......................................................................3-1

   3.1     PROPERTY DESCRIPTION.................................................................................3-1
   3.2     LOCATION.............................................................................................3-2
   3.3     TITLE AND OWNERSHIP..................................................................................3-3
   3.4     PERMITTING...........................................................................................3-9
      3.4.1      BRAZILIAN FRAMEWORK FOR THE ENVIRONMENT........................................................3-9
      3.4.2      CURRENT OPERATIONS STATUS.....................................................................3-13
   3.5     ROYALTIES...........................................................................................3-15

4.0      ACCESS, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY......................................4-1

5.0      PROJECT HISTORY........................................................................................5-1

6.0      GEOLOGICAL SETTING.....................................................................................6-1

   6.1     REGIONAL GEOLOGY.....................................................................................6-1
   6.2     LOCAL GEOLOGY........................................................................................6-3
   6.3     DEPOSIT GEOLOGY......................................................................................6-7

7.0      DEPOSIT TYPE...........................................................................................7-1

8.0      MINERALIZATION.........................................................................................8-1


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   8.1     PETROGRAPHY..........................................................................................8-1
   8.2     SULPHIDES............................................................................................8-1
   8.3     GOLD.................................................................................................8-3

9.0      EXPLORATION............................................................................................9-1

10.0     DRILLING..............................................................................................10-1

   10.1    DRILL SPACING.......................................................................................10-5

11.0     SAMPLING METHOD AND APPROACH..........................................................................11-1

   11.1    BULK DENSITY AND CORE SPECIFIC GRAVITY..............................................................11-2
   11.2    BOND WORK INDEX.....................................................................................11-3

12.0     SAMPLE PREPARATION, ANALYSES AND SECURITY.............................................................12-1

   12.1    SAMPLE PREPARATION AND ANALYSES.....................................................................12-1
   12.2    SECURITY............................................................................................12-3

13.0     QUALITY CONTROL, QUALITY ASSURANCE....................................................................13-5

   13.1    RESULTS.............................................................................................13-7
   13.2    RERUNS.............................................................................................13-10
   13.3    ROUND ROBIN TESTS - COARSE AND PULD REJECT ANALYSES................................................13-14
   13.4    LAB BIAS...........................................................................................13-14

14.0     DATA VERIFICATION.....................................................................................14-1

15.0     ADJACENT PROPERTIES...................................................................................15-1

16.0     MINERAL PROCESSING AND METALLURGICAL TESTING..........................................................16-1

   16.1    EXISTING PROCESS PLANT..............................................................................16-1
   16.2    EXPANSION PLAN......................................................................................16-2
      16.2.1     IN PIT CRUSHING AND CONVEYING.................................................................16-3


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      16.2.2     NEW 32 MTPA MILL..............................................................................16-4
      16.2.3     TAILINGS......................................................................................16-5
      16.2.4     MODIFICATIONS TO THE EXISTING PLANT...........................................................16-6
   16.3    EXPANSION PLAN III METALLURGICAL TESTWORK...........................................................16-6

17.0     MINERAL RESOURCE AND RESERVE ESTIMATES................................................................17-1

   17.1    MINERAL RESERVE AND RESOURCE STATEMENT..............................................................17-2
   17.2    HISTORICAL ESTIMATES................................................................................17-4
   17.3    MODELING METHODOLOGY................................................................................17-6
      17.3.1     OVERVIEW......................................................................................17-6
      17.3.2     GEOLOGICAL INTERPRETATION.....................................................................17-6
   17.4    SAMPLE ANALYSIS....................................................................................17-11
      17.4.1     ARSENIC......................................................................................17-12
      17.4.2     BOND WORK INDEX..............................................................................17-12
      17.4.3     SPECIFIC GRAVITY.............................................................................17-13
   17.5    COMPOSITING........................................................................................17-13
   17.6    GRADE CAPPING AND RESTRICTING OF HIGH GRADE........................................................17-14
   17.7    GEOSTATISTICS......................................................................................17-14
   17.8    BLOCK MODEL........................................................................................17-15
      17.8.1     GRADE INTERPOLATION..........................................................................17-16
      17.8.2     SPECIFIC GRAVITY.............................................................................17-17
      17.8.3     ORE HARDNESS.................................................................................17-17
      17.8.4     RECOVERY.....................................................................................17-17
      17.8.5     MODEL CHECKING...............................................................................17-18
   17.9    RESOURCE CLASSIFICATION............................................................................17-19
   17.10      PIT OPTIMIZATION................................................................................17-21
      17.10.1      BASE CASE..................................................................................17-21
      17.10.2      CUT-OFF GRADES.............................................................................17-25
      17.10.3      PIT DESIGN.................................................................................17-26


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18.0     OTHER RELEVANT DATA AND INFORMATION...................................................................18-1

19.0     INTERPRETATION AND CONCLUSIONS........................................................................19-1

20.0     RECOMMENDATIONS.......................................................................................20-2

21.0     ADDITIONAL INFORMATION FOR OPERATING PROPERTIES.......................................................21-1

   21.1    PROCESS PLANT.......................................................................................21-1
      21.1.1     CRUSHING......................................................................................21-1
      21.1.2     GRINDING CIRCUIT..............................................................................21-2
      21.1.3     GRAVITY CIRCUIT...............................................................................21-2
      21.1.4     FLOTATION.....................................................................................21-2
      21.1.5     HYDROMETALLURGY PLANT.........................................................................21-3
      21.1.6     SMELTING......................................................................................21-3
   21.2    MARKETS AND CONTRACTS...............................................................................21-4
   21.3    RECLAMATION AND MINE CLOSURE........................................................................21-4
   21.4    TAXES...............................................................................................21-4
   21.5    CAPITAL AND OPERATING COST ESTIMATES................................................................21-5
      21.5.1     OPERATING COST ESTIMATE.......................................................................21-7
      21.5.2     ECONOMIC ANALYSIS.............................................................................21-7

22.0     REFERENCES............................................................................................22-1


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                                                   LIST OF TABLES
                                                   --------------


TABLE 1-1 PROVEN AND PROBABLE MINERAL RESERVES - DECEMBER 31, 2005, 2005........................................1-1

TABLE 1-2 MEASURED AND INDICATED MINERAL RESOURCES - DECEMBER 31, 2005..........................................1-2

TABLE 1-3 BOND WORK INDEX ORE HARDNESS ESTIMATES BY HORIZON....................................................1-10

TABLE 3-1 SUMMARY OF RPM MINING LICENSES AND EXPLORATION CONCESSIONS............................................3-7

TABLE 5-1 PARACATU LIFE OF MINE PRODUCTION SUMMARY..............................................................5-3

TABLE 5-2 HISTORICAL MINERAL RESOURCES AND RESERVE ESTIMATES....................................................5-4

TABLE 10-1 DRILL HOLES SUMMARY TABLE...........................................................................10-2

TABLE 10-2: CONFIDENCE LIMITS FOR GOLD.........................................................................10-7

TABLE 10-3: CONFIDENCE LIMITS FOR ARSENIC......................................................................10-7

TABLE 12-1 SUMMARY OF SIMPLE PREPARATION PROCEDURES BY LAB.....................................................12-3

TABLE 13-1: STANDARDS AND THEIR ACCEPTED LIMITS................................................................13-6

TABLE 13-2: SUMMARY OF QAQC BY LABORATORY......................................................................13-7

TABLE13-3: LABORATORY PERFORMANCE SUMMARY FOR 2005 EXPLORATION.................................................13-8

TABLE 13-4 SELECTED RERUN RESULTS.............................................................................13-11

TABLE 13-5 SUMMARY OF BATCH RERUNS............................................................................13-12

TABLE 14-1 PARACATU PRODUCTION RECONCILIATION..................................................................14-2

TABLE 16-1 PROCESS PLANT METALLURGICAL RECOVERY SUMMARY........................................................16-2

TABLE 17-1 PROVEN AND PROBABLE MINERAL RESERVES - DECEMBER 31, 2005............................................17-2

TABLE 17-2 MEASURED AND INDICATED MINERAL RESOURCES - DECEMBER 31, 2005........................................17-3

TABLE 17-3:  UPDATED DRILL HOLE DATABASE.......................................................................17-6


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TABLE 17-4 BASIC STATISTICS FOR GOLD, RAW SAMPLE DATA.........................................................17-12

TABLE 17-5: BASIC STATISTICS FOR ARSENIC ASSAYS...............................................................17-12

TABLE 17-6: BASIC STATISTICS FOR BOND WORK INDEX..............................................................17-13

TABLE 17-7: BASIC STATISTICS FOR SPECIFIC GRAVITY IN CORE SAMPLES.............................................17-13

TABLE 17-8: PARACATU CORRELOGRAM SUMMARY......................................................................17-15

TABLE 17-9 GRADE INTERPOLATION PARAMETERS.....................................................................17-17

TABLE 17-10 COMPARISON OF LM VS NLM WEST OF RICO CREEK........................................................17-19

TABLE 17-11 GRADE TONNAGE SUMMARY OF IMPORTED AND EXPORTED MODEL..............................................17-22

TABLE 17-12: BASE CASE OPTIMIZATION PARAMETERS................................................................17-23

TABLE 17-13: PIT DESIGN CRITERIA..............................................................................17-27

TABLE 17-14 COMPARISON OF PIT DESIGN RESULTS TO WHITTLE 4X(C) OPTIMIZATION RESULTS FOR THE BASE CASE ESTIMATE...17-29


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                                                   LIST OF FIGURES
                                                   ---------------


FIGURE 3-1 - PARACATU MINE LOCATION MAP.........................................................................3-2

FIGURE 3-2 PARACATU MINING AND EXPLORATION CLAIM MAP............................................................3-8

FIGURE 3-3 BRAZILIAN ENVIRONMENTAL LICENSING AND CONTROL PROCESS...............................................3-12

FIGURE 6-1 REGIONAL GEOLOGY PARACATU DISTRICT...................................................................6-3

FIGURE 6-2 TYPICAL SULPHIDE MINERALIZATION IN BOUDINAGE STRUCTURES..............................................6-4

FIGURE 6-3 SMALL SCALE THRUST FAULTING..........................................................................6-5

FIGURE 6-4: LOCAL GEOLOGY OF THE PARACATU DEPOSIT...............................................................6-6

FIGURE 6-5 CONCEPTUAL GEOLOGICAL CROSS SECTION OF THE PARACATU DEPOSIT..........................................6-7

FIGURE 6-6 CONCEPTUAL PRE-MINING WEATHERING PROFILE.............................................................6-8

FIGURE 8-1 PARACATU THIN SECTION GOLD ON ARSENOPYRITE GRAIN BOUNDARY............................................8-3

FIGURE 10-1 DRILL HOLE LOCATION MAP............................................................................10-3

FIGURE 10-1:GOLD ESTIMATION UNCERTAINTY BY DRILL HOLE SPACING..................................................10-8

FIGURE 10-2: ARSENIC ESTIMATION UNCERTAINTY BY DRILL HOLE SPACING..............................................10-8

FIGURE 13-1: STANDARD PERFORMANCE - RPM LAB....................................................................13-9

FIGURE 13-2: STANDARD PERFORMANCE - ALS CHEMEX................................................................13-10

FIGURE 13-3: STANDARD PERFORMANCE - LAKEFIELD.................................................................13-11

FIGURE 13-4 - K-508 SAMPLES 112 TO 135 INITIAL VS RERUN BY ALIQUOT............................................13-13

FIGURE 13-5 PLAN VIEW - DIAMOND DRILLING DISTRIBUTION BY ANALYTICAL LAB.......................................13-15

FIGURE 16. -1 SAG MILL PERFORMANCE CURVE (MORRELL REVISED CURVE)...............................................16-8

FIGURE 16. -2 TYPICAL GOLD ON ARSENOPYRITE GRAIN BOUNDARIES....................................................16-9


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FIGURE 17-1 TONNAGE MINED AND IN RESERVE AS OF DECEMBER 31, 2005...............................................17-5

FIGURE 17-2 OUNCES MINED AND IN RESERVE AS OF DECEMBER 31, 2005................................................17-5

FIGURE 17-3 GRADED BEDDING IN UNMINERALIZED PHYLLITE...........................................................17-7

FIGURE 17-4 PHYLLITE WITH VERGING ASYMETRIC FOLDS, SHEAR BANDS & BOUDINS.......................................17-8

FIGURE 17-5 LARGE ARSENOPYRITE PORPHYROBLAST IN CORE...........................................................17-9

FIGURE 17-6 DRILL SECTION 05N - LOOKING NORTH.................................................................17-11

FIGURE 17-7 BASE CASE WHITTLE 4X(C) RESULTS.....................................................................17-24

FIGURE 17-8 TYPICAL HAUL ROAD PROFILE.........................................................................17-28

FIGURE 21-1: SIMPLIFIED FLOW SHEET EXISTING PARACATU PROCESS PLANT.............................................21-1


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1.0     EXECUTIVE SUMMARY

1.1     INTRODUCTION

Rio Paracatu Mineracao (RPM), a 100% owned subsidiary of Kinross Gold
Corporation (Kinross) operates the Morro do Ouro (Paracatu) mine in Brazil.

The following Technical Report has been prepared in support of the December 31,
2005 resource and reserve disclosure. This report has been prepared to comply
with Canada's National Instrument 43-101.

Table 1-1 summarizes the Proven and Probable mineral reserve estimate for the
Paracatu mine as of December 31, 2005 at a gold price of US$ 400 per ounce, a
Foreign Exchange Rate (FEX) of 2.65 Reais per US $1.00 and a cut off grade of
0.21 g/t Au.

    TABLE 1-1 PROVEN AND PROBABLE MINERAL RESERVES - DECEMBER 31, 2005, 2005

       -------------------------------------------------------------------
               CLASSIFICATION       TONNES        GRADE          GOLD
                                  (X 1,000)      (AU G/T)       (OUNCES)
       -------------------------------------------------------------------
        Proven                    1,103,677        0.40       14,194,000
        Probable                     83,131        0.38        1,016,000
       -------------------------------------------------------------------
        PROVEN & PROBABLE         1,186,808        0.40       15,210,000
       -------------------------------------------------------------------

Table 1-2 summarizes the Measured and Indicated mineral resource estimates
(excluding mineral reserves) for the Paracatu mine as of December 31, 2005 at a
gold price of US $450 per ounce, a Foreign Exchange Rate (FEX) of 2.65 Reais per
US $1.00 and a cut off grade of 0.18 g/t Au.




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     TABLE 1-2 MEASURED AND INDICATED MINERAL RESOURCES - DECEMBER 31, 2005


       -------------------------------------------------------------------
               CLASSIFICATION       TONNES        GRADE          GOLD
                                  (X 1,000)      (AU G/T)       (OUNCES)
       -------------------------------------------------------------------
        Measured                    89,784         0.27         771,000
        Indicated                    5,540         0.38          68,000
       -------------------------------------------------------------------
        MEASURED AND INDICATED      95,324         0.27         839,000
       -------------------------------------------------------------------

 NB MEASURED AND INDICATED RESOURCES ARE REPORTED EXCLUSIVE OF MINERAL RESERVES


In addition to the Measured and Indicated mineral resources stated in Table 1-2,
Paracatu hosts an Inferred resource of 40.1 million tonnes averaging 0.37 g/t
Au. Inferred resources are estimated at a gold price of US $450 per ounce and a
FEX of 2.65 Reais per US $1.00.

The resource and reserve estimates described in this report are classified
according to the Canadian Institute on Mining, Metallurgy and Petroleum (CIM)
Standards on Mineral Resources and Reserves.

1.2     KEY METHODOLOGY CHANGES

The following section summarizes key changes in estimation methodology relative
to the historical estimation methods employed at RPM and previously reported by
Kinross.

1.2.1   DRILL HOLE SPACING AND RESOURCE CLASSIFICATION

Historically, RPM required a minimum drill hole spacing of 100 x 100 meters to
support a classification of Measured and Indicated Resources. In April 2005,
Kinross commissioned Dr. B. Davis (Davis 05), an independent geostatistical
consultant, to complete a study to determine the minimum drill hole spacing
necessary to support a Measured and Indicated resource classification at
Paracatu. Dr. Davis concluded that a 200 x 200 meter "five spot" pattern,
resulting in an average drill hole spacing of 140 meters, was sufficient to
support an Indicated classification at Paracatu. Additional discussion of Dr.
Davis' conclusions is included in Section 10.0 of this report.


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1.2.2   SAMPLE PREPARATION AND ANALYSIS

Historically, RPM assayed six (6), 50-gram sample aliquots for every 1.0 meter
sample submitted for analysis. The average grade reported for each sample was an
average of the six individual aliquots. In June 2005, Kinross received a report
from Agoratek International (Gy, Bongarcon 2005), an independent consulting firm
specializing in gold sampling and hetrogenity studies. Agoratek's principle
consultants are Dr. P. Gy and Dr. D. Francois Bongarcon, recognized industry
experts in sampling theory. Agoratek's scope was to review the historical
sampling methodology employed at Paracatu and recommend changes to maintain
sample integrity and precision. In their June report, Agoratek concluded that
three to four 50g sample aliquots would be adequate to ensure the precision of
the sample results is maintained. As a result of their recommendation, Kinross
abandoned the six aliquot practices and began using three (3) 50-gram aliquots
to determine the grade of each sample interval. Additional discussion regarding
Agoratek's conclusions is included in Section 12.0 of this report.

1.2.3   GEOLOGICAL INTERPRETATION

Historical resource models at Paracatu, estimated by RPM, limited gold grade
interpolation to a mineralized horizon determined largely by limits interpreted
by RPM geologists based on geology and assay data from drill holes. The
mineralized horizon was further sub-divided by weathering profiles and arsenic
content, establishing the C, T, B1 and B2 horizons and Calha, non-Claha and IDS
ore types. Metallurgical recovery was assigned to the Calha, n-n Calha, IDS
units and gold grade interpolation utilized conditional simulation of composite
data within the defined ore limits.

The resource model reported herein is based on a revised geological
interpretation that subdivides the mineralized horizon west of Rico Creek into
two distinct layers, developed largely from geological features logged in the
core and verified against gold assay results.

The hangingwall and footwall contacts of the mineralized zone correspond to the
first and last occurrences of arsenopyrite and/or deformation features such as


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boudins, shears and folds in the core. This defines a zone ranging from 120 to
160 meters in thickness that averages 0.35 to 0.45 g/t Au.

Within this zone is the Boudin Deformation Zone (BDZ) which can be visually
identified based on an increase in the intensity of deformation features and an
increase in arsenopyrite. The BDZ averages 60 to 80 meters in thickness with a
gold grade of 0.60 g/t Au.

East of Rico Creek, in the historical and current mine area, several holes
failed to test the entire thickness of the mineralized horizon, failing to
identify the footwall contact of the mineralization. As a result, grade
interpolation in this area relied on a geological solid that estimated the
footwall limits of the mineralized zone by projecting a limited distance beyond
the last data point available. The estimated contacts are considered by Kinross
to be conservative, rarely extending more than 8.0 meters below the available
drill data.

For the December 31, 2005 model, the mineralized zone limits were based on
several new holes that did identify the footwall contact. Drilling indicates the
BDZ is absent. Kinross interprets the absence to be the result of historic
mining with the BDZ mined out as C-T and B1 ore.

Additional information on the changes in the geological modeling is provided in
Section 17.0 of this report

1.2.4   ORE HARDNESS

Ore hardness has always been recognized as a critical success factor in modeling
the Paracatu deposit. Historically, hardness, measured according to Bond Work
Index (BWI), was assessed based on an 8.0 meter downhole composite sample equal
to the mine`s bench height. The December 31, 2005 model is based on a 12.0 meter
bench height which resulted in recompositing of the historical 8.0 meter data to
reflect the change in bench height.

The 8.0 meter composite samples were composed of a fraction of each meter after
initial sample crushing to 1.4 mm. The BWI test is carried out by the RPM


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process lab following the BWI standard test. BWI values were interpolated into
the model blocks using multi indicator kriging without lithology discretization.

The interpolated BWI values were then used to estimate a Process Cost Adjustment
Factor (PCAF) for each block. The PCAF was evaluated during optimization of the
resource model by Whittle 4X(C), an industry recognized software program.
Whittle 4X(C) determined the profitability of each block considering the PCAF,
recovery and gold content. More detail on the PCAF is provided in Section 17.0
of this report.

1.2.5   METALLURGICAL RECOVERY

Previous models at Paracatu estimated average recoveries for individual ore
types based on the arsenic content of the ore. Sectional interpretation, based
on arsenic analytical data, outlined polygons for Calha, non-Calha and IDS ore
types. Average metallurgical recoveries were then assigned to each unit.

In the resource model reported herein, metallurgical recovery is estimated for
each model block based on arsenic and sulphur data collected from the drill
core. The net result is a variable metallurgical recovery for each model block
based on the same data originally used to define Calha and IDS ores on section.
More detail on how recovery was estimated for the resource model reported herein
is provided in Sections 16.0 and 17.0 of this report.

1.2.6   BENCH HEIGHT

The historical models at Paracatu were based on 4.0 meter composite samples and
an 8.0 meter block height. With the planned increase in throughput capacity, a
12.0 meter bench height was more favourable from a design and operation
perspective.

The December 31, 2005 model is based on a 12.0 meter block height. Gold
composites are based on 6.0 meter composite samples derived from the raw sample
data collected on a 1.0 meter sample interval.


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1.2.7   RESOURCE MODEL OPTIMIZATION

Historically, RPM used MSO, a proprietary software program developed in Brazil,
for optimization of the resource model. For this report, pit optimization is
reported exclusively from Whittle 4X(C), a standard software program recognized
by the international mining community. Kinross has completed several comparisons
between Whittle 4X(C) and MSO and results indicate that MSO typically mines a
larger volume of rock than Whittle 4X(C). The MSO algorithm does not attempt to
identify the highest Net Present Value for the pit shells generated. MSO equates
marginal cost to marginal revenue at the outer boundary of the shells.

1.3     DESCRIPTION AND LOCATION

The Paracatu mine is located 2 km north of the city of Paracatu (population
75,000), in the north western portion of the state of Minas Gerais, Brazil, 230
km southeast of the national capital Brasilia and 480 km northwest of the state
capital Belo Horizonte.

The current mine includes an open cut mine, process plant, tailings impoundment
area and related surface infrastructure, with a throughput rate of 18 million
tonnes per annum (Mtpa). Historically, mining in the pit has not required
drilling or blasting prior to excavation. Ore is ripped using CAT D10 dozers,
pushed to CAT 992 front-end loaders and loaded to CAT 777 haul trucks for
transport to the crusher. In 2004, RPM began blasting harder portions of the
deposit exposed in certain areas of the mine.

The mineral resources and mineral reserves supported by this Technical Report
assume completion of Expansion Plan III, described in detail in Section 16.0 of
this report. Expansion Plan III will increase plant throughput to 50 Mtpa,
allowing more efficient treatment of harder ores at depth and improved recovery
of arsenopyrite rich ores.

RPM currently holds clear mineral rights title to two mining licenses (1,258
hectares) and twenty eight exploration concessions (21,250 hectares) in the


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immediate mine area. RPM has also applied for an additional nine exploration
concessions (16,974 hectares). By way of their application for these additional
concessions, RPM has guaranteed priority rights to the subsurface
mineralization.

The mine and most of the surface infrastructure, with the exception of the
tailings impoundment area, lie within the two mining licenses. The mining
licenses are confirmed by legal survey. An application to convert additional
exploration concessions to mining leases has been submitted to the DNPM for
review. RPM has expressed that there is reasonable certainty that DNPM will
approve the application within the next six months.

In many cases, third party landowners own the surface rights to the exploration
concessions. RPM is guaranteed access to the exploration concessions by
law,through a process known as Servidao. The legal process require RPM to
negotiate a fair price for the surface rights with the landowner. If negotiation
fails to reach an agreement, the matter is put before the Brazilian courts for
settlement.

Servidao was used to successfully secure the surface rights for the existing
operation.

1.4     ACCESSIBILITY CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

Access to the site is provided by paved federal highway or by charter aircraft.
A paved airstrip, suitable for small aircraft is maintained on the outskirts of
city of Paracatu.

The mine is the largest employer in Paracatu, directly employing 750 workers in
what is predominantly an agricultural town (dairy and beef cattle and soy bean
crops) located in Brazil's tropical savannah. Average annual rainfall varies
between 850 and 1800 mm, the average being 1300 mm, with the majority realized
during the rainy season between October and March. Temperatures range from
15(degree) to 35(degree) Celsius.


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The mine draws power from the Brazilian national power grid.

The mine is dependent on rainfall as the primary source of process water. During
the rainy season, the mine channels surface runoff water to temporary storage
ponds from where it is pumped to the beneficiation plant. Similarly, surface
runoff and rain water is stored in the tailings impoundment, which constitutes
the main water reservoir for the concentrator. The objective is to capture and
store as much water as possible from the rainy season to ensure adequate water
supply during the dry season. The mine is permitted to draw make up water from
three local rivers that also provide water for agricultural purposes.

1.5     PROJECT HISTORY

Gold mining has been associated with the Paracatu area since 1722 with the
discovery of placer gold in the creeks and rivers of the Paracatu region.
Alluvial mining peaked in the mid -1800's and until the 1980's; mining activity
was largely restricted to garimpiero (artisinal) miners.

In 1984, Rio Tinto began exploring the property using modern exploration methods
and by 1987, the RPM joint venture was formed between Rio Tinto and Autram
Mineracao e Participacoes (later TVX Gold Inc). The RPM joint venture
constructed the mine and processing facility for an initial capital cost of $65
million.

Production commenced in 1987 and the mine has operated continuously since then.
As of December 31, 2004, the mine has produced close to 3.0 million ounces of
gold from 237.0 M tonnes of ore. Average life of mine mill feed grade is 0.50
g/t Au. The average metallurgical recovery is 78.1%.

Production for the period January through October 2005 was 13.9 M tonnes
averaging 0.43 g/t Au. The resource model described herein has been adjusted to
reflect mine production.

In January 2003, TVX's 49% interest in RPM was acquired by Kinross as part of
the merger between Kinross, TVX and Echo Bay Mines Ltd (EBM).


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In December 2004, Kinross purchased Rio Tinto's 51% interest in RPM to obtain a
100% ownership position in the property.

In 2004, ECM, a Brazilian consulting engineering company completed a Feasibility
Study on Expansion Project III, proposing a throughput increase to 30 Mtpa.

In June 2005, RPM and Kinross staff prepared the Plant Capacity Scoping Study,
which is the primary supporting cost document for the resource and reserve
estimates disclosed herein. Various throughput rates and process options were
examined. The Study indicates that a staged expansion from 18 Mtpa to 32 Mtpa
with a second phase bringing total throughput to a 50 Mtpa returned the best Net
Present Value based on discounted cash flow analyses of the options.

In September 2005, Kinross awarded SNC-Lavalin Engineers and Constructors Ltd
and MinerConsult Engenharia, a Brazilian engineering firm, a contract for the
Basic Engineering of Expansion Plan lll.

1.6     GEOLOGY

Mineralization at Paracatu occurs within the Morro do Ouro sequence, a series of
phyllites that have been thrust from SW to NE producing extensive deformation.
Anamalous gold and sulphide mineralization is localized within a 120 - 140 meter
thick high strain zone that dips gently (20(Degree)) to the SW and is traceable
for over 6 km along a NE-SW trend, and more than 3 km in width. Grade variation
can be visually identified within the high strain zone based on readily
observable geologic features, the most important of which are the frequency of
boudins, intensity of shearing and arsenopyrite content,

Holcombe, Coughlin and Associates (Holcombe 2005), an independent structural
geology company, concluded that the timing of gold and sulphide mineralization
was syn-deformational. Gold and sulphides are scavenged from the Morro do Ouro
sedimentary sequence during deformation and localized within the high


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strain zone(s) that acted as chemical traps due to dissolution of silica and
carbonate and resulting increase in graphite.

Subsequent surface weathering produced four, distinct, weathering horizons. The
individual weathering horizons, known as the C, T, B1 and B2 are described in
detail in Section 6.0 of this report. Mining to date has exhausted the majority
of the softer C and T horizons. The remaining reserves for the project are
hosted in the B1 and B2 horizons with the majority (90%) hosted in the B2
horizon. Ore hardness, based on Bond Work Index (BWI) tests of the cores,
increases with depth from surface. Table 1-3 presents the average range of BWI
measurements by horizon.

           TABLE 1-3 BOND WORK INDEX ORE HARDNESS ESTIMATES BY HORIZON

                        ---------------------------------
                            HORIZON         BWI RANGE
                                             (kWh/t)
                        ---------------------------------
                               C             2 to 3
                               T             3 to 4
                               B1            5 to 7
                               B2            8 to 16
                        ---------------------------------

Historically, sulphide mineralization in mineralized horizon has been
sub-divided based on the arsenic content. The historically units, know as Calha
(arsenic greater than 2500 ppm), non-Calha (arsenic less than 2500 ppm) and
Intensely Deformed Sulphide (IDS) mineralization (the central portion of Calha
lenses with an arsenic content greater than 4000 ppm) were traditionally
interpreted and differentiated during resource modeling.

The percentage of arsenopyrite in the ore directly affects metallurgical
recovery. Ore with higher arsenic content typically has slightly lower
metallurgical recovery.

In the resource and reserve estimate summarized in this report, Kinross utilized
sulphur and arsenic assays collected during the drill programs to estimate the
metallurgical recovery for each block in the resource model. Complete details on


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how recovery has been estimated for this estimate are provided in Sections 16.0
and 17.0 of this report.

Mineralization is confined to the finely laminated phyllites of the Morro do
Ouro sequence immediately overlying the massive Serra da Landim metasiltstone
member that forms the base of the Paracatu formation. Gold and sulphide
mineralization is believed to be syngenetic with the deposition of the
phyllites.

In late Proterozoic times, the weaker phyllites responded more easily to
tectonic pressures than the enveloping siltstone units. Regional east-west
deformation and a later phase of north-south buckling (interpreted to be
responsible for formation of a high strain zone, occurred simultaneously with
remobilization of gold and sulphide mineralization.

Evidence supporting the two phase deformational history is provided by mapping
of the boudin axes. Outside of the high strain zone, boudin axes trend
north-south. Within the high strain zone, the axes are rotated to an east-west
orientation.

1.7     DEPOSIT TYPE

The Paracatu deposit is a metamorphic gold system with finely disseminated gold
mineralization hosted within an original bedded sedimentary host. Mineralization
is syn-deformational with the thrusting of the rocks of the Morro do Ouro
sequence from WSW to ENE. To the authors knowledge, Paracatu is a unique deposit
and therefore is termed a Morro do Ouro type deposit. The deposit has
extraordinary lateral continuity and exhibits very predictable grade
distribution and recovery characteristics. It is considered unlikely, given the
genesis of the deposit, that there would be significant deviation in the tenor
or physical properties of the gold mineralization at Paracatu.


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1.8     MINERALIZATION

The Paracatu phyllites have been metamorphically altered to lower greenschist
facies resulting in pervasive quartz-sericite alteration. Metamorphic grade
increases from east to west.

Sulphide mineralization is dominantly arsenopyrite and pyrite with pyrrhotite
and lesser amounts of chalcopyrite, sphalerite and galena.

Gold is closely associated with arsenopyrite and pyrite and occurs predominantly
as fine-grained free gold along the arsenopyrite and pyrite grain boundaries or
in fractures in the individual arsenopyrite and pyrite grains. Thin section
analyses indicate 92% of the gold is free. Gold grains typically average 50-150
microns in size. The size and amount of the gold grains does not correlate well
with the size or amount of the arsenopyrite grains. It is however essential that
arsenopyrite be available as a substrate on which gold can occur.

1.9     EXPLORATION

Rio Tinto was the first company to apply modern exploration methods at Paracatu.
Northeast of Rico Creek, the deposit had been drilled on a nominal 100 x 100
meter drill spacing.

Exploration at Paracatu evolved in lock step with knowledge gained through
production experience. Essentially, the success of mining in the C and T
horizons focused attention and exploration effort on the B1 horizon. Continued
production success in the B1 horizon led to increased interest in the B2
horizon.

Recent drilling by Kinross has indicated that portions of the deposit NE of Rico
Creek have not been drill tested for the entire thickness of the mineralized
horizon hosting gold. This largely reflects the historical mining theory at
Paracatu where softer C, T and B1 ores were targeted and harder B2 ores were
considered uneconomic due to limitations in the existing process plant
technology in operation at that particular moment in time.


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Expansion Plan III will allow processing of the harder ores of the B2 horizon.
Originally, Kinross focused on increasing reserves to the SW of Rico Creek,
exploiting the B2 mineralization that continues down dip of the surface exposure
being mined in the current pit.

1.10    DRILLING

The dominant sample collection method supporting the December 31, 2005 resource
and reserve model is diamond drilling. A total of 1,427 drill holes totalling
79,961 meters support the resource and reserve estimate stated in this report.
Table 10-1 summarizes the data used in this report.

During 2005, Kinross added 267 holes (48,660 meters) which represents the single
largest drill program in the history of the Paracatu mine. The resource model
described by this report incorporates gold results from 228 out of 267 drill
holes completed in 2005. Analytical results for the remaining 39 holes were
pending at the time of the estimate.

The nominal drill spacing ENE of Rico Creek is 100 x 100 meters. An Optimum
Drill Spacing Study (Davis 05) commissioned by Kinross established that a 200 x
200 meter five spot pattern (a 200 x 200 m grid plus one hole in the middle)
would satisfactorily define Indicated mineral resources. This pattern results in
a nominal 140 meter hole spacing and represents a departure from historical RPM
practices.

Diamond drilling has demonstrated that anomalous gold grades (greater than 0.20
g/t Au) occur within a 120-150 meter thick tabular zone that has been traced for
more than 4.0 km (NE-SW) by 3.0 km. (NW-SE). Anomalous gold grades remain open
down dip and laterally.

The portion of the deposit demonstrated to be economically viable is
approximately 3.0 km by 2.0 km in size.


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1.11    SAMPLING METHOD AND APPROACH

The diamond drill holes have been systematically sampled using a 1.0 meter
sample interval. In all, a total of 48,238 samples have been collected and
analyzed. Core recovery is typically greater than 95%. The core is logged and a
photographic record of each hole is collected prior to any sampling. The core is
systematically sampled on 1.0 meter intervals without adjustment for geological
boundaries. Sampling consumes 100% of the core except for the 8.0 cm pieces
selected from every two meter interval which are retained and stored for S.G and
Point Load Testing (PLT) analysis.

Specific gravity measurements are collected during the core logging process
using the water displacement method. These measurements are checked against
samples collected from the upper levels of each mining bench during mining of
the deposit.

Samples for BWI analysis are collected as composite samples during sample
preparation and are subjected to RPM's standard BWI analysis method.

1.12    SAMPLE PREPARATION, ANALYSIS AND SECURITY

Historical sample preparation and analysis was performed recognizing the low
average grade of the deposit. The historical method reduced each one meter core
sample to 95% passing 1.44mm. Crushed samples were homogenized and split with
approximately 7 kg stored as coarse reject. Approximately 200 grams of the
remaining sample were split off for ICP analysis and 1.35 kg of sample was split
out for Bond Work Index analysis. The remaining sample (4.5kg) was dried and
further reduced to 95% passing 65 mesh. This sample was homogenized and split
with 4.2 kg stores as pulp reject and the remain 300g was fully analyzed using
standard fire assay with AA finish in a series of six, individual 50 g aliquots.
Results from the six individual aliquots were weight averaged together to
determine the final grade for each sample.


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The process described above was time consuming, adversely impacted sample turn
around time and QAQC sample turnaround. In an effort to streamline the
preparation and analysis of the drill samples collected during Kinross'
exploration effort, Kinross completed several studies at the start of the
exploration program.

In April 2005, an audit of the RPM mine lab was undertaken by Kinross'
Laboratory Manager from the Fort Knox Mine (Oleson 05) to assess lab equipment
and procedures. The audit recommended changes in preparation and fluxing that
were implemented immediately resulting in markedly improved productivity and
QAQC performance. Variability between 50 g aliquots was reduced significantly.

In May 2005, Kinross commissioned Agoratek International (Gy, Bongarcon 05) to
review sample preparation and analysis procedures with a specific mandate to
assess the historical practice of assaying six individual 50 g aliquots per
sample and averaging the results. Agoratek, concluded that three (3) 50 g
analyses would be sufficient for determining the grade of any given sample.

Based on the lab audit and the Agoratek study, Kinross' standardized sample
preparation and analytical procedure for the remainder of the exploration
program was as follows:

Samples (typically 8.0 kg) are crushed to 95% passing 2.0 mm and homogenized at
the RPM sample preparation lab. Approximately 6 kg of sample is stored as coarse
reject; the remaining 2 kg of sample is split out and pulverized to 90% passing
150 mesh. This sample is homogenized and three (3) 50 g aliquots are selected
for fire assaying with an AA finish. The remaining pulverized sample is
maintained as a sample pulp reject.

Sample analyses were performed at three separate analytical labs during the
exploration program. Two independent labs: Lakefield, Brazil and ALS Chemex,
Vancouver was utilized due to the number of samples generated. RPM's lab at
Paracatu also analyzed samples during the exploration program.


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The use of three separate analytical facilities in compiling the results for the
89 additional drill holes added and used in completing this resource and reserve
estimate is beneficial in that it results in reduced potential of lab bias
influencing the accuracy of the estimate.

1.13    DATA VERIFICATION

RPM staff has indicated that Rio Tinto employed rigorous data verification
procedures. Kinross has not independently verified the data transcription
against original sources for historical data. Kinross has verified 10% of the
historical data collected between 1999 and 2004 against original source
documents. The verification did not identify any concerns regarding the quality
or accuracy of the historical data used in the December 31, 2005 resource model.

For the 2005 drill program, Kinross' exploration geologists managing the program
verified all data. Gold grades were all double entered and weight averaged per
sample, then the two databases were crosschecked with no significant errors or
differences detected. Arsenic and sulphur assays have undergone initial
cross-checking at the time of this report. Final checks were ongoing as are some
QAQC batch re-runs. Batch reruns were redone if the blind standards inserted in
the sample stream exceeded 2 standard deviations from the mean for any samples
within the mineralized horizon.

The summary database spreadsheet was compared to the individual digital files
sent by the different laboratories. Kinross is confident that the database is
sufficiently free of errors to support the present mineral resource and mineral
reserve estimates.

1.14    ADJACENT PROPERTIES

There are no other producing mines near the Paracatu mine. Fazenda Lavras is a
gold prospect located approximately 13 km from Paracatu. It has some
similarities with the Paracatu deposit but it is not in production. On a
regional scale there are additional anomalies being investigated by Kinross.


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1.15    MINERAL PROCESSING AND METALLURGICAL TESTING

The existing process plant at Paracatu has operated continuously since 1987 and
has had expansion upgrades in 1997 and 1999. In 2005, the plant processed 17.2
Mtpa and achieved an average gold recovery of 78.8%. The plant includes primary
and secondary crushing, grinding, gravity and flotation circuits. A
hydrometallurgical circuit leaches the concentrates and produces gold bullion.

Plant recoveries are estimated on the basis of sulphur and arsenic content in
each block. The maximum possible flotation plant recovery is 86% and this
decreases linearly with increasing sulphur and arsenic assays. Hydromet gold
recovery is modeled at a constant 96.5%.

RPM recognized that further plant improvements were necessary to maintain
current production levels in the face of increasing ore hardness. Exploration
drilling had successfully traced the Paracatu deposit to depth and west of Rico
Creek. Sampling indicated that ore hardness increased with increasing depth from
surface.

In response to the increasing ore hardness, RPM began evaluating options to
further increase plant throughput. In 2004, a Feasibility Study was completed by
ECM, a Brazilian engineering firm. Aker-Kvaerner contributed technical expertise
to ECM's study.

In June 2005, ECM completed a Plant Capacity Scope Study which considered
several alternatives to increase plant throughput. All options considered in the
Study assumed the installation of an in pit crushing and conveying system (IPCC)
and 38 foot diameter Semi-Autogenous Grinding (SAG) mill which were the
cornerstone assumptions in the original Feasibility Study.

Data on SAG mill performance was collected during a pilot plant program
completed by RPM staff. The pilot plant data was run on 1,500 tonnes of Paracatu
ore with WIs ranging from 5.5 to 12.0 kWh/t. In all, six different ore types
were processed through a Koppers 6x2 foot SAG mill that was leased from


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CETEM, Rio de Janeiro, Brazil. The pilot plant testwork and analysis of the
results were all completed under the supervision of a team of recognized expert
in the filed of SAG mill design and operation.

In Q4 2005, SNC Lavalin and Minerconsult were contracted to complete basic
engineering for the Expansion III Project. The scope of work included the IPCC,
covered stockpile, 32 Mtpa mill, hydromet expansion, power supply, tailings
delivery and water systems. The SAG mill and ball mills were purchased in
December 2005. The basic engineering design and supporting capital and operating
costs estimates will form the basis of the 2006 Feasibility Study.

The Expansion III Project will proceed in two stages over a four year period
commencing in 2006. The first stage will increase plant capacity from 18 to 32
Mtpa. The new 32 Mtpa SAG mill plant will be constructed and once commissioned,
the existing 18 Mtpa plant will be shut down and refurbished. Once refurbishment
of the 18 Mtpa plant is complete, it will be restarted and tasked with
processing the remaining B1 reserve. This will bring total plant throughput for
the two lines to 50 Mtpa. When the soft BI ore is depleted in 2017, the
throughput capacity will be limited to 41 Mtpa and then capacity will decrease
further as work index increases above a value of 11 in 2024.

1.16    MINERAL RESOURCE AND RESERVE ESTIMATE

The mineral resource model for Paracatu is interpreted and estimated using
Maptek Pty Ltd.s Vulcan(C) software.

The mineral resource model for Paracatu is developed from a series of NW-SE
oriented drill sections that include analytical data from the drill programs,
pre-mining topography and current mine development. The sections are used to
define the contacts between the various mineral horizons of interest.

The December 31, 2005 resource model is based on visual observations resulting
from Kinross' 2005 exploration drill campaign. Gold mineralization in the model
is strongly related to visual geological factors such as the frequency of


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boudins, folding, shearing and arsenopyrite content. Higher grade gold results
were found to correspond with a marked increase in boudins, folding, shearing
and arsenopyrite content. This correlation was used to refine gold grade
estimation in the resource model.

For the mineralization west of Rico Creek, the hanging and footwall contacts of
the mineralized zone were based on visual observation in the drill core of the
first and last occurrences of arsenopyrite and/or structural textures such as
boudins, folding and shearing. West of Rico Creek, the mineralized unit dips at
20(0) to the SW, averages 120 to 150 meters in thickness with a gold grade of
0.40 g/t. The mineralized horizon remains open down dip and along strike, a
result of limiting the 2005 exploration campaign within a $400 pit shell. The
mineralized horizon demonstrate remarkable grade and geological continuity.

Within the mineralized horizon, a zone of intense structural deformation can be
visually identified in drill core. The zone features increased occurrences of
boudins, folding and shearing and an increased concentration of arsenopyrite.
The Boudin Deformation Zone (BDZ) ranges in thickness from 60 to 80 meters,
averages 0.60 g/t Au and also demonstrate remarkable grade and geological
continuity.

East of Rico Creek, the mineralization is interpreted from the current mine
working to the footwall contact of the zone as defined by the last occurrence of
arsenopyrite. Kinross completed several holes in the NW quadrant of the deposit
to ensure that the footwall limit was properly identified. The footwall limit
had previously been interpreted by RPM geologists from drill data that had
stopped well short of the footwall contact.

The geological information is interpreted on the sections and the resulting
interpretation is imported into Vulcan(C) for further processing. Linear
features (faults, lithologic contacts, and mineralization polygons are modeled
as continuous three-dimensional surfaces and wireframes in Vulcan(C).


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The mineralized wireframes are used to extract sample data (gold, arsenic,
sulphur, BWI, specific gravity) and code model blocks in a 50 x 50 x 12 (x, y,
z) meter block model.

Raw assay data for gold (1.0 meter samples) is composited into 6.0 meter
intervals. The composite data is extracted using the wireframes produced from
sectional interpretation. Each composite is coded according to the geological
wire frame. Composites less than 2.0 meters in length are discarded and any
duplicate (twinned) composites are also discarded. Grade capping for original
1.0 m assays is considered on a zone by zone basis. High-grade results
occasionally occur in the 1.0 m sample results. Cumulative probability plots
were calculated for B1, B2 and BDZ. A capping grade of 1.4 g/t was selected for
both B1 and B2 based on the 99th percentile of the grade distribution. Within
the BDZ the capping level was set at 1.6 g/t.

The extracted composite data for gold, arsenic and sulphur for each zone is
analyzed using directional semi-variograms to determine the major, semi-major
and minor axes and the influence of individual composites. The variograms are
used to interpolate grades into the individual model blocks.

Gold grades are interpolated using Ordinary Kriging with each geological unit
(zone) estimated independently. The zone solids are used as hard boundaries and
the composites must have the identical domain code item as the solids to be used
in the interpolation process.

An octant search is used in all cases for grade interpolation. A minimum of 1
composite and a maximum of 12 composites are used within the search ellipsoid. A
maximum of four adjacent samples are used from the same drillhole.

The resource model described in this report estimates specific gravity for each
model block (50 x 50 x 12 meters). Specific gravity measurements for core
samples are collected and assessed based on 4.0 m composite samples comprised of
8.0 cm core intervals selected for every 2.0 meters of core. Statistical and
geostatistical analysis of the data is used to develop correlograms


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and fitted models for interpolation of specific gravity for the model blocks.
Interpolation extracts data for each geological zone and used the composite data
to estimate the grade for each block within the zone using ordinary kriging.

A comparison with overall tonnage estimates estimated by the previous regression
model compared to pit production statistics showed an increase in tonnage of
about 7%. The December 2005 density model for B1 ore was therefore factored down
by 7% to reflect operating experience. Only the B1 ore was factored downward.
The density for the B2 horizon is on field measurements taken during the drill
campaigns.

Most of the data supporting the 7% reduction originates form the B1 horizon
which is notably more weathered and possibly, more variable as far as specific
gravity is concerned. It is Kinross' opinion that in all likelihood, the more
competent rocks in the B2 horizon will likely not demonstrate the same trend as
the B1 rocks.

BWI data is also modeled from the composite data collected from the drill holes
during sample preparation. BWI is interpolated for each block in the model using
a nearest neighbour interpolation method.

Finally, each model block is assigned a metallurgical recovery based on results
for sulphur (S) and arsenic (As). The metallurgical recovery is based on the
following equation.

        Recovery = (a +(-2.36230 x S%) +(-0.0017 x As ppm)) x b) where

        a = theoretical maximum flotation recovery of 85.95352% and

        b = theoretical hydrometallurgical recovery or 96.5%

The resource model is classified according to the Canadian Institute on Mining,
Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves.

The resource model classification uses a combination of geostatistical methods
and manual verification. The primary classification is the result of drill
spacing analysis completed by Dr. B. Davis in April 2005, which is then manually
verified.


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The resource model is exported to Whittle 4X(C), an accepted industry standard
program used to estimate mineral reserves. Grade tonnage tables of the exported
model are compared to a grade tonnage table from Vulcan(C) to ensure the
accuracy of the transfer. Whittle 4X(C) optimizations are completed on the
Measured and Indicated mineral resources to develop a series of nested pit
shells. The shells are analyzed assuming a $US 400 per ounce gold price and a
FEX of 2.65 Reais per US$. An optimum shell is selected to guide the design of
the final pit.

The operating and capital costs used to complete the Whittle 4X(C) analysis are
those estimated in the Plant Capacity Scoping Study. Capital costs were updated
by ECM in 2005 to reflect current price increases and were based on recent
supplier quotations for the major plant equipment. Operating costs for the
recommended mine equipment fleet were estimated from first principles.

Geotechnical parameters are consistent with those provided by Golder and
Associates in their report dated June, 2005 (Golder 05).

Pit design is completed using Datamine(C) modeling software. The optimum pit
shell selected from Whittle 4X(C) is exported to Datamine(C) and used to guide
manual pit design. The pit design parameters are described in detail in Section
17.0 of this report.

The final pit design is modeled in Datamine(C) to generate a final surface. The
Vulcan(C) resource model is imported into Datamine(C) and the grade tonnage
curve is verified to ensure the model matches the model exported from Vulcan(C).
The pit design is used to extract the resource model blocks within the pit
design and the blocks are reported by class (Measured, Indicated and Inferred)
within the pit shell. Measured resources convert to Proven reserves, Indicated
resources convert to Probable reserves and Inferred resources are reported
separately.


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Mineral resources are estimated directly from Whittle 4X(C). The mineral
resources presented in this report assume a gold price of US$ 450 per ounce and
a FEX of 2.65 Reais per US$.

The mineral resources reported are the incremental difference between the
optimum pit shell at US$ 450 per ounce and the design pit at US$ 400 per ounce.
Total Proven Reserves at US$ 400 per ounce are subtracted from total Measured
Resources at US$450 per ounce and the difference is reported as the measured
resource at the US$ 450 per ounce price level. The same calculation is performed
on the Probable and Indicated component.

1.17    CONCLUSIONS

The Paracatu mine is a model mining operation. Gold production has consistently
met targeted levels in the 19 years the mine has been in operation. Over that
period of time, the predictive accuracy of the mineral resource model has been
verified by actual production experience.

RPM have completed a thorough pilot plant test confirming the amenability of the
Paracatu ores to SAG milling. A Feasibility Study was completed on an option to
increase throughput to 30 Mtpa with the addition of a SAG mill and in pit
crushing and conveying system was completed in 2004 and updated in 2005 to
reflect rising costs.

A Plant Capacity Scoping Study, examining the Net Present Values of various
plant throughput rates was completed in Q2-2005. This study considered four
throughput options:

        o       Maintain current plant throughput

        o       Expansion to 30 Mtpa

        o       Expansion to 50 Mtpa

        o       Expansion to 66 Mtpa


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The Plant Capacity Scoping Study was based on the updated costs from the
Feasibility Study. It is a detailed study and Kinross considers the costs to
have been estimated to a pre-feasibility level of accuracy. Most of the major
plant equipment is based on updated vendor quotations and site operating and
construction costs are well known from operating experience at Paracatu.

The Plant Capacity Scoping Study indicated the best Net Present Value of the
options considered was an expansion to a 50 Mtpa throughput rate. Discounted
cash flow analyses indicate that the project has a positive cash flow at gold
prices above US $ 400 per ounce.

1.18    RECOMMENDATIONS

Kinross considers the resource model to be very robust with minor risks
associated with the estimation of gold grade.

The remaining arsenic, sulphur, work index and density results should be
completed and added to the model. Kinross does not consider the missing data to
pose any significant risk to the resource and reserve estimates stated in this
report.


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2.0     INTRODUCTION AND TERMS OF REFERENCE

2.1     INTRODUCTION

The mineral resource and mineral reserve estimates summarized in this report are
classified according to the Canadian Institute on Mining, Metallurgy and
Petroleum (CIM) Standards on Mineral Resources and Reserves as required by
Canada's National Instrument 43-101. This report has been prepared under the
direct supervision of:

W. Hanson, P.Geo, Vice-President, Technical Services, Kinross Gold Corporation.

Mr. Hanson has personally visited the Paracatu mine on several occassions and
has been directly involved in the work supporting the estimate disclosed herein.

This report has been prepared in support Kinross Gold Corporations (Kinross')
December 31, 2005 resource and reserve disclosure. The resources and reserves
are based on an updated resource model prepared in December 2005, and assumes
that the existing operation would be expanded to increase plant throughput to 50
Mtpa.

ECM, a Brazilian consulting engineering company completed a Feasibility Study on
the planned expansion in 2004. In June 2005, RPM and KTS staff prepared the
Plant Capacity Scope Study that examined various plant throughputs and process
options to determine the highest Net Present Value for each option. On September
2005, Kinross awarded SNC-Lavalin Engineers and Constructors and MinerConsult
Engenharia a contract to complete Basic Engineering activities related to
Expansion Plan lll.

2.2     TERMS OF REFERENCE

All units of measure (distance, area, etc,) unless otherwise noted are in metric
units of measure.


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All monetary units are expressed in terms of October 2005 US dollars unless
otherwise specified.

An exchange rate of $2.65 Brazilian reais per $1.00 US has been assumed
throughout the Capacity Study and the 2006 Life of Mine Plan and, by extension,
this Technical Report. KTS prepared a sensitivity analysis of the mineral
reserves at an exchange rate of $2.45 Brazilian reais per $1.00 US.

2.3     GLOSSARY

        CIM      Canadian Institute of Mining Metallurgy and Petroleum

        CONAMA   National Environmental Council

        DNMP     Departamento Nacional da Producao Mineral (National Department
                 for Mineral Production)

        EIA      Environmental Impact Assessment

        g/t      grams per tonne

        IBAMA    Brazilian Institute for the Environment and Renewable Resources

        JORC     Joint Ore Reserves Committee

        KTS      Kinross Technical Services

        KWh/t    kilowatt-hours per tonne

        M        million

        Ha       hectares

        Mtpa     million tonnes per annum

        MW       megawatts

        oz(s)    troy ounce(s)

        PAE      Economic Development Plan

        ROM      run of mine


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        SAG      semi-autogenous grinding

        SGA      Environmental Management System

        SISNAMA  National Environmental System

        t        tonne(s)

        Tpa      tonnes per annum

        Tpd      tonnes per day

        Tph      tonnes per hour

2.4     SCOPE AND OBJECTIVES

This report is prepared in support of Kinross' December 31, 2005 resource and
reserve estimate for the Paracatu Mine.

2.5     REPORT BASIS

This Technical Report is based on costs and financial analyses completed as part
of the RPM Plant Capacity Scope Study completed in June 2005. The resource model
and reserve estimate have been prepared by RPM and Kinross staff. Reserve
estimates are based on a mine plan within design pit developed based on an
optimized pit shell estimated by Whittle 4X(C). Current operating costs were
adjusted to reflect increased throughput rates after completion of the proposed
plant expansion detailed in the Plant Capacity Scoping Study.

The underlying data supporting the reserve estimate has been verified for
accuracy by RPM staff and Kinross experts. No errors have been noted.

The lead author of this report has personally visited the project on several
occasions and has reviewed the estimation methodology.

2.6     INDEPENDENT THIRD PARTY PARTICIPANTS

The following independent consultants have contributed indirectly to this
report:


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Agoratek International                          Sampling Hetrogeneity Study

Dr. B. Davis, Independent Consultant            Optimum Drill Hole Spacing

ECM Engineering                                 Feasibility Study Expansion
                                                Project III

Holcombe, Couglin & Associates                  Structural Assessment


2.7     STUDY PARTICIPANTS

The following employees of Kinross have contributed to the report:

M. Belanger, P.Geo, Kinross Americas            Resource Estimation

C. Frizzo, Kinross Americas Exploration         Geology & QA/QC

B. Gillies, P.Geo, Kinross Gold Corporation     Geology & QA/QC

R. Henderson, P.Eng, Kinross Gold Corporation   Metallurgy & Process

K. Morris, P.Eng, Kinross Gold Corporation      Reserve Estimation

J. Oleson, Kinross, Fort Knox Operations        Laboratory Audit

Dr. R. Peroni, Rio Paracatu Mineracao           Resource Estimation

W. Phillips, Kinross Americas                   Metallurgy and Process

L. A. Tondo, Rio Paracatu Mineracao             Metallurgy and Process


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2.8     DISCLAIMER

This document has been prepared by Kinross Gold Corporation's Technical Services
Department (KTS). The document summarizes the professional opinion of the
author(s) and includes conclusions and estimates that have been based on
professional judgement and reasonable care. Said conclusions and estimates are
consistent with the level of detail of this study and 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. Written or verbal excerpts
from this report may not be used without the express written consent of the
author(s) or officers of Kinross Gold Corporation.


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3.0     PROPERTY DESCRIPTION AND LOCATION

3.1     PROPERTY DESCRIPTION

The Paracatu Mine (locally known as Morro do Ouro) is operated by Rio Paracatu
Mineracao (RPM), a wholly owned subsidiary of Kinross Gold Corporation
(Kinross). The mine has been in continuous operation since 1987.

The mine includes an open cast mine, process plant, tailings impoundment area
and related surface infrastructure and support buildings. Current plant
throughput averages 18 Mtpa.

Currently, mining does not require any waste removal (stripping) and just a
limited amount of explosive is necessary to blast the harder ores prior to
excavation. Ore is ripped and pushed into piles by CAT D10 dozers. CAT 992
front-end loaders load the ore from the piles into CAT 777 rigid frame haul
trucks that transport the ore to the existing crusher.

Ore hardness increases with the depth from surface and as a result, modeling the
hardness of the Paracatu ore is as important as modeling the grade. Ore hardness
is modeled based on Bond Work Index (BWI) analyses of diamond drill samples. RPM
currently estimates that blasting of the Paracatu ore will be necessary for
blocks with a BWI greater than 8.5 kWh/t

The mineral resources and mineral reserves supported by this Technical Report
assume implementation of Expansion Plan III.

The planned Expansion Plan III proposes to increase plant throughput to 30 Mtpa
then 50 Mtpa, allowing more efficient treatment of harder ores at depth and the
arsenopyrite rich ores. It is expected that with the Expansion Plan lll a fleet
of larger mining equipment comprising 218 tonne trucks and either hydraulic or
electric shovels will be purchased. KTS has prepared a trade-off study comparing
various combinations of trucks and loading units. Capital and operating costs
were also estimated.


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3.2     LOCATION

The mine is located less 3 km north of the city of Paracatu (population 75,000)
in the northwest part of the state of Minas Gerais, Brazil. Paracatu is located
approximately 230 km from Brazil's capital, Brasilia at latitude 17(degree)3'S
and longitude 46(degree)35'W. Figure 3-1 is a location map showing the location
of Paracatu (in red).

                     FIGURE 3-1 - PARACATU MINE LOCATION MAP






                                    [PICTURE]






The mine is located at an elevation of 780 m above sea level.


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3.3     TITLE AND OWNERSHIP

In Brazil, the Departamento Nacional da Producao Mineral (DNPM) issues all
mining leases and exploration concessions. Mining leases are renewable annually,
and have no set expiry date. Each year RPM is required to provide information to
DNPM summarizing mine production statistics.

RPM currently holds title to two contiguous mining claims totalling 1,258
hectares:

        o       DNPM Nos. 830.241/80 and 800.005/75 are outlined in blue in
                Figure 3-2 below. The mine and most of the surface
                infrastructure, with the exception of the tailings impoundment
                area, lie within the two mining licenses. The mining claims are
                confirmed by legal survey.

The current tailings impoundment is located on lands to which RPM has negotiated
surface rights with the former landowner(s).

RPM also holds title to 28 exploration concessions (21,250 hectares), shown in
red and magenta outlines in Figure 3-2. RPM also has applications before the
DNPM for an additional 9 concessions (16,974 hectares), shown in black in Figure
3-2, in and around the Paracatu area.

Exploration concessions are granted for a period of three (3) years. Once a
company has applied for an exploration concession, the applicant holds a
priority right to the concession area provided no previous ownership exists. The
owner of the concession can apply to have the exploration concession
successively renewed. Renewal is at the sole discretion of DNPM.

Granted exploration concessions are published in the Official Gazette of the
Republic (OGR), which lists individual concessions and their change in status.

The exploration concession grants the owner the sub-surface mineral rights.
Surface rights can be applied for if the land is not owned by a third party.

The owner of an exploration concession is guaranteed, by law, access to perform
exploration field work, provided adequate compensation is paid to third party


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landowners and the owner accepts all environmental liabilities resulting from
the exploration work.

In instances where third party landowners have denied surface access to an
exploration concession, the owner maintains full title to the concession until
such time as the issue of access is negotiated or legally enforced by the
courts. Access is guaranteed under law so eventually; the owner will gain access
to the exploration concession. Once access is obtained, the owner will have
three (3) years to submit an ER on the concession. This process is known as
Servidao and RPM has used it to obtain the surface rights from the landowners
during development of the current mine.

The owner of a mineral concession is obligated to explore the mineral potential
of the concession and submit an Exploration Report (ER) to DNPM summarizing the
results of the fieldwork and providing conclusions as to the economic viability
of the mineralization. The content and structure of the report is dictated by
DNPM and a qualified professional must prepare the report.

DNPM will review the ER for the concessions and either:

        o       approve the report, provided DNPM concurs with the report's
                conclusions regarding the potential to exploit the
                mineralization,

        o       dismiss the report should the report not address all
                requirements in which case the owner is given a term in which to
                address any identified deficiencies in the report or,

        o       postpone a decision on the report should it be decided that
                exploitation of the deposits are temporarily non-economic.

Approval, dismissal or postponement of the ER is at the discretion of the DNPM.
There is no set time limit for the DNPM to complete the review of the ER. The
owner is notified of the DNPM's decision on the ER and the decision id published
in the OGR.


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On DNPM approval of the ER, the owner of an exploration concession shall have
one year to apply for a mining lease. The application must include a detailed
Development Plan (DP) outlining how the deposit will be mined.

DNPM will review the DP and decide whether or not to grant the application. The
decision is at the discretion of DNPM but approval is virtually assured unless
development of the project is considered harmful to the public or the
development of the project compromises interests more relevant than industrial
exploitation. Should the application for a mining lease be denied for
exploration concessions for which the ER has been approved, the owner is
entitled to government compensation.

On approval of the DP, DNPM will grant the mining license, which will remain in
force until the depletion of the mineral resource. DNPM will publish the change
in the OGR.

RPM holds clear title to all the exploration concessions listed in Table 3-1. As
previously noted, access to said concessions is guaranteed under law. Given the
mines exemplary operations record for the past 18 years, there is no reason to
suspect that application to convert said exploration concessions to mining
leases would be denied.

RPM currently has applications before DNPM to convert four exploration
concessions to mining lease status. The four concessions are highlighted with
green shading in Figure 3-2. The current status of this application is
summarized below for each exploration concession.

        o       EXPLORATION PERMIT 831205/85

                The ER was submitted and approved on April 22, 2002. The mine
                claim request was submitted on April 17, 2005 and is dependent
                on the subsequent presentation of the DP that is planned for 25
                November, 2005. Once all necessary material is submitted to the
                DNMP, it is expected to take approximately six months to obtain
                the final mining claim.


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        o       EXPLORATION PERMIT 830907/99

                The ER was submitted and approved on April 22, 2002. The mine
                claim request was submitted on April 17, 2005. As per the claim
                above, its acceptance depends on the presentation of the DP to
                be submitted on November 25, 2005. The mining claim is expected
                after a period of six months following the presentation.

        o       EXPLORATION PERMIT 832228/93

                Title was effectively changed from GALESA (Rio Tinto) to RPM on
                November 22, 2005. RPM must present the ER and DP for this area
                to obtain the mining lease. It is expected to take approximately
                six months.

        o       EXPLORATION PERMIT 832225/93

                This exploration concession renewal is due January 1, 2006. RPM
                must present the ER and DP to obtain a mining lease. Once all
                reports are submitted, it is expected to take six months to go
                through the process established by the DNPM.

Table 3-1 summarizes RPM's current mining licenses and exploration concessions.


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      TABLE 3-1 SUMMARY OF RPM MINING LICENSES AND EXPLORATION CONCESSIONS



                                                                             
- -------------------------------------------------------------------------------------------------------
      DNPM                    TYPE                     DATE             MINING LEASE            AREA
      (#)                                            Acquired         Application Date       (Hectares)
- -------------------------------------------------------------------------------------------------------
830.241/80       Mining Lease                        03/11/80                                      828
800.005/75       Mining Lease                        01/02/75                                      430
- -------------------------------------------------------------------------------------------------------
                 SUBTOTAL                                                                        1,258
- -------------------------------------------------------------------------------------------------------
831.205/85       Exploration Concession              08/26/85                04/17/05               20
830.907/99       Exploration Concession              05/17/99                04/17/05               28
835.561/93       Exploration Concession              10/18/93            -                         131
832.228/93       Exploration Concession              06/21/93                11/22/05              990
832.225/93       Exploration Concession              06/21/93                01/01/06              938
832.227/93       Exploration Concession              06/21/93            -                          21
832.229/93       Exploration Concession              06/21/93            -                         950
805.862/75       Exploration Concession              07/02/75            -                         187
805.863/75       Exploration Concession              07/02/75            -                         130
831.848/93       Exploration Concession              06/07/93            -                         409
832.224/93       Exploration Concession              06/21/93            -                         171
831.823/99       Exploration Concession              09/24/99            -                         908
831.561/99       Exploration Concession              10/18/99            -                         976
830.253/00       Exploration Concession              02/10/00            -                       1,538
830.742/05       Exploration Concession              04/04/05            -                         381
830.743/05       Exploration Concession              04/04/05            -                       1,275
830.800/05       Exploration Concession              04/11/05            -                         461
830.801/05       Exploration Concession              04/11/05            -                         229
831358/05        Exploration Concession              06/13/05            -                         139
831537/05        Exploration Concession              07/04/05            -                         403
831892/05        Exploration Concession              08/17/05            -                           1
831893/05        Exploration Concession              08/17/05            -                         210
831894/05        Exploration Concession              08/17/05            -                       1,776
831895/05        Exploration Concession              08/17/05            -                       2,000
831896/05        Exploration Concession              08/17/05            -                       1,879
831897/05        Exploration Concession              08/17/05            -                       1,992
831898/05        Exploration Concession              08/17/05            -                       1,750
831899/05        Exploration Concession              08/17/05            -                       1,358
                 --------------------------------------------------------------------------------------
                 SUBTOTAL                                                                       21,250
                 --------------------------------------------------------------------------------------
831900/05        Exploration Concession              08/17/05            -                       1,881
832064/05        Exploration Concession              09/02/05            -                       2,000
832065/05        Exploration Concession              09/02/05            -                       2,000
832233/05        Exploration Concession              09/21/05            -                       2,000
831942/05        Exploration Concession              08/22/05            -                       1,967
831943/05        Exploration Concession              08/22/05            -                       1,316
831944/05        Exploration Concession              08/22/05            -                       1,986
831945/05        Exploration Concession              08/22/05            -                       1,841
832389/05        Exploration Concession              10/04/05            -                       1,984
                 --------------------------------------------------------------------------------------
                 SUBTOTAL                                                                       16,974
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              FIGURE 3-2 PARACATU MINING AND EXPLORATION CLAIM MAP











                                    [PICTURE]











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3.4     PERMITTING

3.4.1   BRAZILIAN FRAMEWORK FOR THE ENVIRONMENT

The Brazilian environmental policy is executed at three different levels of
public administration - federal, state and municipal. Coordinating and
formulating the Brazilian Environmental Policy is the responsibility of the
Ministry for the Environment. Directly linked to it is the National
Environmental Council (CONAMA), the deliberative and consultative board for
environmental policy. CONAMA's responsibility is to establish the rules,
standards and criteria guidelines so that environmental licensing can be granted
and controlled by the state and municipal local environmental agencies which are
part of the National Environmental System (SISNAMA), and by the Brazilian
Institute for the Environment and Renewable Resources (IBAMA). IBAMA is the
government agency under the jurisdiction of the Ministry for the Environment,
and is the agency responsible for executing the Brazilian Environmental Policy
at the federal level.

The basic environmental process is initiated with the collection of baseline
data, following the submission of a conceptual mine plan. Baseline data
collection is followed with an Environmental Impact Assessment (EIA), leading to
an Environmental Impact Report (RIMA), which is a summary of the EIA presented
in simple language adequate to public communication and consultation. The EIA
and RIMA are made available for public review and comment.

Once the EIA/RIMA process is complete, the Environmental License (LA) is
required to move the project forward. The LA is issued by the State Agency,
under guidelines developed by the CONAMA. There are a number of components to
the Environmental License:

        o       PRIOR LICENSE (LP) - this is relevant to the mining project's
                preliminary planning stage and contains the basic requirements
                to be met during the location, installing and operating stages,
                in accordance with the municipal, state or federal plans for
                soil use.


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                Requirements must meet regulations, criteria and standards set
                out in the general guidelines for environmental licensing issued
                by the CONAMA. In addition, the criteria and standards
                established by the state environmental agency must be met, in
                the scope of the agencies area of jurisdiction, providing there
                is no conflict with federal level requirements.

                o       The Mining Plan and the EIA/RIMA are technical documents
                required for obtaining the Prior License. This process is
                concurrent with the request for a mining concession.

        o       INSTALLATION LICENSE (LI) - authorizes the start of the mining
                project, including implementation and installation of the
                project, according to the specifications in the approved
                Environmental Control Plan. After the LP is granted, an Economic
                Development Plan (PAE) is prepared, to be approved by the
                National Department for Mineral Production (DNPM), as well as an
                Environmental Control Plan (PCA, based on the Environmental
                Management System (SGA), to be approved by local Environmental
                Agency in order for the Installation License and the land
                clearing (deforestation) license to be issued. At this stage, a
                closure plan is also required, to be submitted for the DNPM's
                approval.

        o       OPERATING LICENSE (LO) - authorizes, after necessary
                confirmation, the start of the licensed activity and functioning
                of its pollution control equipment, according to that set out in
                the Prior and Installation Licenses. During the operating phase
                of the Project, Annual Mining Reports (RAL) are submitted by the
                company for DNPM's approval. In the closure phase, the company
                applies for a Conformity Certificate from the environmental
                agency and DNPM, after the decommissioning, restoration and
                environmental monitoring operations are finished.


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Figure 3-3 is a simplified diagram of the environmental and mining rights,
licensing and control processes.

Kinross is confident that RPM holds clear mineral title to the resources and
reserves discussed in this report.





























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        FIGURE 3-3 BRAZILIAN ENVIRONMENTAL LICENSING AND CONTROL PROCESS















                                    [PICTURE]














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Environmental licenses related to Expansion Plan III have been obtained from the
Environmental Regulatory Authorities; these include the Preliminary License (PL)
allowing mining below the water table and the Installation License (IL) for
installing the major plant equipment for phase I of Expansion Plan III.

3.4.2   CURRENT OPERATIONS STATUS

One of the initial conditions satisfied by RPM in obtaining a mining license was
that an Environmental Impact Assessment (EIA) was successfully filed with the
State of Minas Gerais environmental agency. During the time that the mining
license is effective, the Operation License must be renewed every four years. In
the year 2000, RPM was the first Brazilian gold mining company to receive ISO
14001 certification. The mine has implemented excellent environmental care and
monitoring programs. They include complete acid rock drainage (ARD) prediction
and control program for mining the B2 sulphide ore and reclamation research and
studies carried out in partnership with Vicosa Federal University (UFV) to
define the final profile and vegetation for mined areas and the tailings dam.

RPM is currently licensed to draw a set amount of water from the Sao Domingos,
Santa Rita and Sao Pedro rivers. As previously discussed, any additional water
demands are likely to be a sensitive issue in the community. It is likely that
applications to increase water drawdown from the rivers will require public and
government consultation and possibly additional environmental study. RPM staff
has expressed confidence that Expansion Plan III can be completed under current
water drawdown rates.

Another permitting factor affecting Expansion Plan III is mining on the
exploration claims west of Rico Creek. Rico Creek is a historic placer mining
area and the soils in and around the creek are contaminated with mercury. The
creek plays an important role in the community however and any disruption of the
creek had to be carefully presented to the community.

A communication process about Expansion Plan III was initiated in November 2003.
A series of presentations outlining the benefits to the community and


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describing the environmental impact of the planned expansion were initially
presented to a representative group of RPM employees and selected opinion maker
groups in the town of Paracatu including the local press. From these initial
meetings, a strategy was developed to communicate Expansion Plan III to the
community. The planned diversion of Rico Creek to allow continued mining of the
deposit was one of the main focus areas for the community.

Public perception of this process has been very positive as evidenced by the
factthat no public hearings were requested after the EIA study was submitted to
FEAM (the environmental agency). Legally, any party could call for a public
hearing, at any time, within 45 days of submission of the EIA. This clearly
indicates that the communication process was successful in building public
support for the project within the local community.

The final potentially significant permitting issue is related to approvals to
mine below the water table. Currently the mine is not permitted to mine below
the water table. This would require a specific permit that is issued by the
State Water Authority. RPM personnel have indicated that it is reasonable to
assume that the necessary government approvals will be granted in the first
quarter of 2006. RPM has not actively pursued the necessary permits as there
were sufficient mineral reserves above the water table to support the long-range
mine plan. For the Expansion Project lll, however, these reserves become very
important. The impact of lowering the water table in the areas influenced by the
mine was studied in details by RPM. A Geohydrological and Geohydrochemical model
to identify the underground flows and water quality has been generated to
support the EIA study. The results were also communicated to the community. No
public hearing was requested after the EIA submission.

On September 29, 2005 RPM was granted the Previous Licence (LP) during a meeting
of the Environmental Executive Committee (COPAM) at the Minas Gerais State
Environmental Agency (FEAM).


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RPM is finalizing some documents related to the LP and the Environmental Control
Plan. This documentation should be submitted to the Environmental Agency by
November 15, 2005. The Installation Licence (LI), which will allow starting with
the expansion installation works, is expected to be granted in April 2006.

Kinross is confident that all necessary permits for the planned expansion and
the acquisition of all necessary surface rights is guaranteed under Brazilian
mining law. Kinross is not aware of any limitations that would dent successful
permitting of the project described herein.

3.5     ROYALTIES

RPM must pay to the DNMP a royalty equivalent to 1% of net sales. Another 0.5%
has to be paid to the holders of surface rights in the mine area if the rights
are not already owned by RPM.


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4.0     ACCESS, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

Access to the site is provided by the federal highway system, a network of
modern, paved roads that are maintained by the federal government. A small paved
airstrip also services the community. The airstrip can accommodate small,
charter aircraft.

The Paracatu mine is located 230 km southeast of the national capital, Brasilia
(pop. 2.1 million) and 480 km northwest of the state capital Belo Horizonte
(pop. 2.5 million). Both cities are modern cities with industrial and
manufacturing facilities. Belo Horizonte is considered the "mining capital" of
Brazil and several major mining suppliers and engineering companies are
headquartered there.

Paracatu is located in the Brazilian savannah, a region characterized by low
rolling hills that have been largely cleared of vegetation to support farming
along with cattle ranching. The elevation at the mine site is 780 meters above
sea level. The region is largely dependent on agriculture with soya beans being
the predominant crop.

The Paracatu mine is the largest industrial enterprise in the region, employing
750 people, most of who live in the city of Paracatu.

There are two distinct seasons, a rainy season from October to March and a dry
season from April through to September. Temperatures average 20(degree) Celsius,
ranging from a high of 35(degree) C to a low of 15(degree) C. Average annual
rainfall totals between 850-1800 mm.

Domestic water for the mine is obtained from the city of Paracatu, via pipelines
from the municipal water company provider. Process water is largely recycled
from the tailings pond. Make up water is drawn from the Sao Domingos and Sao
Pedro rivers during the rainy season to maintain the water level in the tailings
dam at a level sufficient to provide adequate water during the dry season. The
mine also has access to artesian wells as an emergency water supply.


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The Sao Domingos and Sao Pedro rivers provide all the water necessary to support
agricultural irrigation in the area. As such, the drawdown of additional water
is considered a sensitive issue in the community and was identified by RPM staff
as a potential limiting factor in the design of the SAG Mill Expansion Project.
RPM staff carefully monitored densities in the process circuit and concluded
that the SAG Mill Expansion could be operated without having to modify their
existing water drawdown permits.

The mine is connected to the national power grid, which relies mainly on
hydroelectric generation. Electricity is subject to a free market environment
with consumers able to select their supplier of choice. RPM obtains electricity
from Centrais Eletricas Minas Gerais (CEMIG). The mine has a small emergency
power capability, used for critical process equipment that cannot be suddenly
stopped such as thickeners and CIL tank agitators.

The mine has established surface areas for tailings disposal, and for its
mineral processing facilities. These are sufficient to meet the future needs as
defined by the Life of Mine Plan. In the case of the tailings storage, the
impoundment dam will be raised in a series of lifts to provide the necessary
storage volume.


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5.0     PROJECT HISTORY

The mining history of the Paracatu region is closely associated with the
activities of the Portuguese bandeirantes expeditions who prospected for gold in
Brazil's interior, arriving in the Paracatu region in 1722 after the discovery
of gold alluvial placers.

Alluvial mining peaked during the second half of the 18th century. The alluvial
mining was not limited to the placer deposits along Rico Creek, they also
exploited the oxidized ore outcrop on the top of Morro do Ouro hill or the "Hill
of Gold".

Gold production declined sharply in the region during the first decade of the
19th century. From this point forward, production was limited to "garimpeiros",
subsistence level mining practiced by local inhabitants. Various prospectors
explored the region but economically viable operations were limited as a result
of the low-grade nature of the deposits.

Beginning in 1970, Paracatu attracted some attention from mineral exploration
companies looking for lead and zinc deposits in the area. The interest in the
gold of Morro do Ouro was secondary as the majority of the companies were not
attracted by the gold grade, considered to be too low to be economically
extracted.

In 1980, Rio Tinto, operating in Brazil under the name of Riofinex do Brasil,
joined with Billiton in a partnership to explore land in Brazil. Billiton owned
the Morro do Ouro area but had no interest in investing in the area. In 1984
Billiton sold the balance of its shares in the Morro do Ouro area to Riofinex.
Riofinex embarked on a surface exploration program that focused on the oxidized
and weathered horizons of the Moro do Ouro area. At the end of 1984, based on
the data from hundreds of test pits (up to 25 m deep) and further supported by a
total of 44 drill holes, a reserve of 97.5 Mt at 0.587g/t Au was estimated at
what is currently known as the Paracatu Mine.


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This estimate only included the superficial oxidized ore, then categorized as
type C or T ore. Despite the low gold grade, Riofinex's geologists believed that
profitable extraction of the ore could be realized. In 1985 this was confirmed
by a feasibility study. Total investment up to that period was $7.3 million
including ground acquisition costs, exploration costs, and the cost of the
feasibility study.

Approval was granted by Rio Tinto to construct a mining project at a capital
cost of approximately US$ 65 million, on the condition that a Brazilian partner
could be secured for the venture. At the end of 1985, RTZ Mineracao, successor
to Riofinex, struck a joint venture agreement with Autram Mineracao e
Participacoes (Autram) to joint venture the project through a new company, Rio
Paracatu Mineracao (RPM), with Rio Tinto holding a 51% operating interest and
Autram the remaining 49%.

Autram's interest was ceded to TVX Participacoes who later became TVX Gold Inc.
(TVX). TVX entered into an agreement with Newmont that resulted in Newmont and
TVX holding a 24.5% interest in Paracatu. In early 2003, TVX acquired Newmont's
24.5% interest resulting in TVX having a 49% interest in Paracatu. Almost
immediately, Kinross acquired TVX's interest as part of the Kinross, TVX, Echo
Bay Mines Ltd (EBM), merger agreement.

Production at Paracatu commenced in October 1987 treating oxidized and highly
weathered ore from the C and T ore horizons described in Section 5.0 of this
report. The first gold bar was poured in December 1987. The following year, the
mine throughput reached the design capacity of 6.1 Mtpa.

After start up, the throughput rate was progressively increased to 13 Mtpa, as a
result of a number of improvement programs. In 1993, an $18.3M Optimization
Project was commissioned providing extra water and flotation capacity for the
circuit.

Throughput at Paracatu was increased again to 16 Mtpa in 1997 after completion
of Expansion Project I with a capital cost expenditure of $47.3 M.


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Expansion Project II (1999) increased the mill throughput to 20 Mtpa after a
capital investment of $6.2M. Due to an increase in ore hardness, throughput has
now fallen to the 18.0 Mtpa level.

Total capital investment to December 31, 2004 totalled $249.4 M dollars. This
includes the initial purchase costs of the land, all engineering, the initial
construction costs, later optimization and expansion capital costs, the purchase
of the mining fleet and other smaller capital investments to optimize the
existing project.

The plant currently produces approximately 200,000 ounces of gold annually at an
average cash cost of $220 per ounce

In December 2004, Kinross purchased Rio Tinto's 51% interest in the RPM mine
giving Kinross a 100% interest in RPM and the Paracatu mine.

Table 5-1 summarizes the historic life of mine production at Paracatu since it
began commercial production in 1987.

               TABLE 5-1 PARACATU LIFE OF MINE PRODUCTION SUMMARY



- ------------------------------------------------------------------------------------------------------------------------------------
            YEAR            1987      1988       1989       1990       1991       1992       1993      1994       1995        1996
- ------------------------------------------------------------------------------------------------------------------------------------
                                                                                                
 Tonnes milled (million)       0.5       6.2        8.2        9.3       10.1       10.5       13.0      13.4       13.6        13.5
 Feed grade (Au g/t)          0.78      0.77       0.67       0.64       0.61       0.58       0.50      0.50       0.49        0.50
 Gold Produced (oz)          3,884   113,257    145,844    160,258    166,053    167,000    174,699   169,003    162,844     165,646
- ------------------------------------------------------------------------------------------------------------------------------------
            YEAR              1997      1998       1999       2000       2001       2002       2003      2004       2005       TOTAL
- ------------------------------------------------------------------------------------------------------------------------------------
 Tonnes milled (million)      15.3      15.6       17.5       19.7       16.5       18.4       18.4      17.3       17.2       254.2
 Feed grade (Au g/t)          0.47      0.48       0.45       0.47       0.45       0.48       0.44      0.44       0.42        0.50
 Gold Produced (oz)        156,687   181,305    188,938    228,866    186,915    224,539    200,691   188,574    180,522   3,165,524
- ------------------------------------------------------------------------------------------------------------------------------------


Table 5-2 summarizes the mineral resource and reserve estimates for the Paracatu
mine since Kinross acquired an interest in the property in December 2002. In
2002 and 2003, Kinross held a 49% interest in the property with Rio Tinto, the
operator, holding the


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remaining 51%. RPM estimated and reported mineral resources and reserves in 2002
and 2003 according to the Australian Institute of Mining and Metallurgy (AusIMM)
Joint Ore Reserves Committee (JORC) Code. Kinross acquired Rio Tinto's 51%
interest in December 2004 and reported mineral resources and reserves according
to the Canadian Institute of Mining, Metallurgy and Petroleum's (CIM) Standards.
There are no material differences between JORC and CIM resource and reserve
classifications.

          TABLE 5-2 HISTORICAL MINERAL RESOURCES AND RESERVE ESTIMATES



- -----------------------------------------------------------------------------------------------------------------------------
                KINROSS      GOLD       REPORTING
     Date      Ownership    Price          Code                Classification          Tonnes        Grade           Gold
                  (%)      (US$/oz)                                                  (x 1,000)      (Au g/t)       (Au ozs)
- -----------------------------------------------------------------------------------------------------------------------------
                                                                                          
  31-Dec-02       49%        $300          JORC      Proven                            156,547         0.43        2,163,000
                                                                                 --------------------------------------------
                             $300                    Probable                           24,402         0.43          337,000
                                                     ------------------------------------------------------------------------
                             $300                    PROVEN & PROBABLE                 180,859         0.43        2,500,000
                                                     ------------------------------------------------------------------------
                             $325                    Measured                           14,700         0.46          217,000
                                                                                 --------------------------------------------
                             $325                    Indicated                          69,580         0.38          850,000
                                                     ------------------------------------------------------------------------
                             $325                    MEASURED AND INDICATED             84,280         0.39        1,067,000
                                                     ------------------------------------------------------------------------
                             $325                    INFERRED                           27,400         0.40
- -----------------------------------------------------------------------------------------------------------------------------
  31-Dec-03       49%        $325          JORC      Proven                            163,971         0.42        2,225,000
                                                                                 --------------------------------------------
                             $325                    Probable                           31,829         0.38          388,000
                                                     ------------------------------------------------------------------------
                             $325                    PROVEN & PROBABLE                 195,800         0.42        2,613,000
                                                     ------------------------------------------------------------------------
                             $350                    Measured                                -            -                -
                                                                                 --------------------------------------------
                             $350                    Indicated                          76,627         0.39          966,000
                                                     ------------------------------------------------------------------------
                             $350                    MEASURED AND INDICATED             76,627         0.39          966,000
                                                     ------------------------------------------------------------------------
                             $350                    INFERRED                           30,508         0.37
- -----------------------------------------------------------------------------------------------------------------------------
  31-Dec-04      100%        $350          CIM       Proven                            425,947         0.44        6,025,000
                                                                                 --------------------------------------------
                             $350                    Probable                          178,464         0.43        2,437,000
                                                     ------------------------------------------------------------------------
                             $350                    PROVEN & PROBABLE                 604,411         0.44        8,463,000
                                                     ------------------------------------------------------------------------
                             $400                    Measured                            1,645         0.30           16,000
                                                                                 --------------------------------------------
                             $400                    Indicated                             647         0.31            6,000
                                                     ------------------------------------------------------------------------
                             $400                    MEASURED AND INDICATED              2,292         0.30           22,000
                                                     ------------------------------------------------------------------------
                             $400                    INFERRED                           71,881         0.40
- -----------------------------------------------------------------------------------------------------------------------------
  22-Nov-05      100%        $400          CIM       Proven                            807,341         0.44       11,212,000
                                                                                 --------------------------------------------
                             $400                    Probable                          139,633         0.46        2,068,000
                                                     ------------------------------------------------------------------------
                             $400                    PROVEN & PROBABLE                 946,974         0.44       13,280,000
                                                     ------------------------------------------------------------------------
                             $450                    Measured                          110,837         0.43        1,530,000
                                                                                 --------------------------------------------
                             $450                    Indicated                          11,069         0.41          147,000
                                                     ------------------------------------------------------------------------
                             $450                    MEASURED AND INDICATED            121,906         0.43        1,677,000
                                                     ------------------------------------------------------------------------
                             $450                    INFERRED                          122,981         0.43
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6.0     GEOLOGICAL SETTING

In May 2005, R. Holcombe of Holcombe Coughlin and Associates, an independent
structural geology consulting firm, visited the site and conducted fieldwork to
isolate the structural controls on mineralization at Paracatu.

Holcombe hypothesizes that the mineralization at Paracatu is closely related to
the thrust faulting that emplaced the Paracatu Formation to the NW over top of
younger rocks of the Vazante Formation. Gold and sulphide mineralization was
emplaced syn-deformationally, localized from the surrounding sediments through
metamorphic alteration and concentrated into high stress areas where shearing
was greatest during thrusting. Silica and carbonate were stripped out of the
high strain zones resulting in an increase in graphite, providing an ideal
chemical trap to precipitate gold and sulphide minerals out metamorphic
remobilization fluids generated by pressure from the lithostatic pile.

6.1     REGIONAL GEOLOGY

The mineralization is hosted by a thick sequence of phyllites belonging to the
basal part of the Upper Proterozoic Paracatu Formation and known locally as the
Morro do Ouro Sequence. The sequence outcrops in a northerly trend in the
eastern Brasilia Fold Belt, which, in turn, forms the western edge of the San
Francisco Craton. The Brasilia Fold Belt predominantly consists of clastic
sediments, which have undergone lower greenschist grade metamorphism along with
significant tectonic deformation.

A series of east-northeast trending thrust faults are extensively developed
along the belt. Metamorphic grade increases towards the west as the thickness of
the fold belt increases. The timing of deformation is estimated at between
800-600 Ma during the Brasiliano orogenic cycle and the mineralization is
believed to originate syngenetically with this period of deformation.

A number of anomalous gold occurrences have been mapped in the area. Most are
hosted in rocks similar to those being mined at Paracatu. Stratigraphic


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correlation between the know occurrences is difficult, largely as a result of
fault offsets and lack of true marker units. It is not certain that these other
mineralized occurrences are within the same stratigraphic horizon as Paracatu.

Mineralization at Cabeca Seca and Luziania occurs along the same northwest
linear trend as Paracatu. This trend defines a significant regional gravity
anomaly.

Figure 6-1 is a regional geological map of the Paracatu district modified as per
Holcombe 2005.




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                  FIGURE 6-1 REGIONAL GEOLOGY PARACATU DISTRICT










                                    [PICTURE]









6.2     LOCAL GEOLOGY

The phyllites at Paracatu lie within a broader series of regional phyllites. The
Paracatu phyllites exhibit extensive deformation and feature well developed
quartz boudins and associated sulphide mineralization. Sericite is common,
likely as a result of extensive metamorphic alteration of the host rocks.

Primary sedimentary features and bedding planes are easily recognizable but are
intensively deformed with development of thrusting, bedding plane thrusting,
sygmoidal and boudinage structures as can be observed in Figures 6-2 and 6-3
below.


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       FIGURE 6-2 TYPICAL SULPHIDE MINERALIZATION IN BOUDINAGE STRUCTURES








                                    [PICTURE]








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                     FIGURE 6-3 SMALL SCALE THRUST FAULTING






                                    [PICTURE]






Mineralization at Paracatu is closely related to a period of ductile
deformation, associated shearing and thrust faulting. Overall, the Morro do Ouro
sequence has been thrust to the northeast. Intense, low angle isoclinal folds
are commonly observed. The mineralization plunges to the west-southwest at 15 to
20(degree) and there is secondary folding with axial planes striking to the
northwest resulting in kink bands and egg box folds in some areas.

The mineralization appears to be truncated to the north by a major normal fault
trending east-northeast as mapped in Figure 6-4. The displacement along this
fault is not currently understood but the fault is used as a hard boundary
during mineral resource estimation. The current interpretation is that the fault
has displaced the mineralization upwards and natural processes have eroded away
any mineralization in this area.


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                FIGURE 6-4: LOCAL GEOLOGY OF THE PARACATU DEPOSIT

























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Figure 6-5 presents a conceptualized geological cross section looking to the
Northwest through the Paracatu deposit. The section shows the high strain zone
in pink surrounded by the weakly mineralized phyllites of the Morro do Ouro
sequence. Kinross' exploration results and the resource and reserve estimate
summarized in this report are the results collected from following the high
strain zone to the southwest, down dip from Rico Creek.

     FIGURE 6-5 CONCEPTUAL GEOLOGICAL CROSS SECTION OF THE PARACATU DEPOSIT












                                    [PICTURE]













6.3     DEPOSIT GEOLOGY

The Paracatu mineralization is subdivided into 4 horizons defined by the degree
of oxidation and surface weathering and the associated sulphide mineralization.
These units are, from surface, the C, T, B1 and B2 horizons. Figure 6-6 presents
the conceptual pre-mining weathering surface and established the relative
relationship between the various zones. Mining to date has exhausted the C and T


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horizons. The remaining mineral reserves are exclusively hosted in the B1 and B2
horizons.

               FIGURE 6-6 CONCEPTUAL PRE-MINING WEATHERING PROFILE






                                    [PICTURE]







Type C mineralization occurs at surface and extends to 20 - 30 meters from
surface. Type C mineralization is completely altered with no remaining
sulphides. It also features localized laterite development.

The T horizon is generally only a couple of meters thick. It is varicoloured and
is essentially marks the transition from the C-horizon to the B1 horizon.

The B1 horizon is dark in colour and carbonaceous with less oxidation than the
C-horizon. Sulphides have been completely oxidized but some fresh sulphide
material is visible in the quartz boudins.

B2 mineralization was originally described as un-weathered or fresh
mineralization with primary sulphides.

The contact between un-mineralized host rock (Type A) and the various
mineralized horizons is gradational, occurring over a 10m wide zone that is
characterized by arsenic values of 200-500ppm and up to 0.2 g/t of gold.


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7.0     DEPOSIT TYPE

The Paracatu deposit is a metamorphic gold system with finely disseminated gold
mineralization hosted within an original bedded sedimentary host (phyllite).
Very fine, evenly distributed gold (associated with sulphides) is finely
disseminated throughout a thinly bedded phyllite (metamorphosed argillaceous
sedimentary rock) of Upper Proterozoic age.

The phyllites at Paracatu are highly deformed as a result of tectonic processes.

Gold mineralization at Paracatu was introduced syn-tectonically, the result of
metamorphic alteration during thrusting of the Morro do Ouro sequence over top
of the rocks of the younger Vazante Formation. Metamorphic grade increases from
east to west.. Structural interpretation suggests that mineralization was
precipitated within a high strain zone where silica and carbonate were scavenged
out of the host phyllites resulting in an increase in graphite content that may
have acted as a chemical trap, precipitating out gold and sulphide
mineralization remobilized during metamorphic alteration of the Morro do Ouro
Sequence.


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8.0     MINERALIZATION

8.1     PETROGRAPHY

The mineralization at Paracatu is indicative of metamorphic alteration of lower
greenschist facies intensity. Early petrographic studies of the B1
mineralization indicated that quartz and sericite make up 80% of the rock mass.
Carbon occurs in the form of a fine opaque dust disseminated within the
individual sericite bands. Carbon content varies from 5-20%. Minor amounts of
ilmenite, tourmaline, anatase, rutile and limonite are also commonly observed.

In 2000, a suite of 50 samples of typical Paracatu mineralization was submitted
for petrographic study. The samples were collected from different ore horizons,
at different locations and at different depths from surface and are considered
to be representative of the Paracatu mineralization.

West of Rico Creek a similar sized suite was collected from B2 rocks of the 2005
drilling campaign and confirmed that these rocks are mineralogically the primary
equivalent of slightly more weathered analogues to the east.

Results indicated that 60-90% of unoxidized phyllites were composed of quartz
and sericite producing the distinctive banding noted. Individual bands typically
are less than 2 cm in thickness.

The phyllites also contain carbonate (calcite and ankerite) locally up to 20%
and the same fine grained carbon noted in the previous petrographic work was
also observed in the latter samples. Accessory minerals included muscovite,
biotite, albite, tourmaline, ilmenite, chlorite, zircon and rutile.

8.2     SULPHIDES

The amount of sulphides present typically doesn't exceed 3-4%. The most common
sulphides observed are pyrite, arsenopyrite and pyrrhotite. Galena is relatively
common and may be accompanied by sphalerite. Chalcopyrite occurs locally in
fractures in the main sulphide minerals noted above.


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The sulphides typically occur as individual crystals or coarse crystalline
aggregates.

Arsenopyrite is the most common sulphide and occurs as a fine grained (less than
1mm) to coarsegrained (greater than 3mm) aggregates. Crystals up to 1 cm in size
are not uncommon. Arsenopyrite crystals increase in size to the southwest.

The mineralization at Paracatu exhibits distinct mineralogical zoning with the
arsenopyrite content increasing towards the center and west and in the zones of
intense deformation. Gold grades increase in lock step with the arsenopyrite so
that the highest gold grades occur where arsenopyrite content is greatest.

Pyrrhotite occurs in the western part of the deposit and gold grade are elevated
where pyrrhotite increases. There is evidence for the existence of a high-grade
pyrrhotite body at depth, which has been intersected in a number of drillholes.

The paragenetic model proposed for Paracatu suggests that gold and arsenopyrite
were introduced concurrently, syn-tectonically with deformation.

Holcombe suggests that the boudins typically observed in the higher grade
portions of the Paracatu deposit, represent original, attenuated quartz veins.
Holcombe notes that the quartz boudins crosscut bedding at a shallow angle. The
boudin thickness likely represents the original thickness of the quartz vein and
these have been considerably attenuated implying moderately high to very high
strain in the system.

Holcombe interprets a two-stage process related to the boudins, the first stage
emplaces the quartz veins early in the deformation event. As stress builds,
these veins are folded, boudinaged and separated. It is interesting to note the
apparent absence of continuous quartz veins in the Paracatu rocks. Mineralized
boudins are consistently foliation parallel, while a later barren quartz
boudinage phase is noted to cross cut folation. A final late barren quartz
stockwork phase also cross cuts foliation in the low grade hanging wall.


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8.3     GOLD

Gold occurs either as free gold or electrum. Microscopic analysis indicates that
92% of the gold at Paracatu is free milling with less than 8% encapsulated by
sulphide grains or silica.

RPM examined 50 polished sections of Paracatu ore and identified 79 gold grains
in 16 of the samples. 50 grains were associated with arsenopyrite either
occurring on the grain boundaries or as inclusions. The remaining 29 gold grains
were associated with pyrite.

No gold was observed with pyrrhotite and no gold was noted without sulphide.

The gold grains varied from sub-rounded to highly irregular (angular).
Typically, gold grains were less than 10 microns in size and occur on the
sulphide grain boundaries as seen in Figure 8-1.

      FIGURE 8-1 PARACATU THIN SECTION GOLD ON ARSENOPYRITE GRAIN BOUNDARY







                                    [PICTURE]








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The gold varies in color from pale to deep yellow reflecting variation in the
silver content.

Another mineralogical assessment made by Rio Tinto in Bristol has analysed ore
samples ground at a grinding size of 106 microns. 634 gold particles were
identified, 27 % being bigger than 53 microns and 16 % bigger than 75 microns.
These grains represented around 60 % of the total gold area of the samples. By
the same talk, only 7 % of the grains were bigger than 106 microns, but those
represented 40 % of the total gold area of the samples.

In summary, all mineralogical assessments conducted so far indicate that gold is
associated preferentially with arsenopyrite. Gold is predominantly free milling
and responds to cyanidation. The majority of grains are ultrafine (less than 20
microns) but the few coarse grains that occur are responsible for the highest
percentage of the contained gold in the ore.


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9.0     EXPLORATION

Rio Tinto was the first company to apply modern exploration methods at Paracatu.
The initial production decision was based on a mineral reserve estimate based on
44 drill holes and 458 surface pits (25 m maximum depth) testing the C and T
horizons at Paracatu.

The deposit, with the exception of the exploration permits west of Rico Creek,
is currently drilled off on nominal 100 x 100 meter drill spacing.

The exploration history at Paracatu has evolved in lock step with the mine
development. Initially, the exploration effort was focused only on defining
mineral reserves within the C and T horizons. As a result, the majority of the
sample support was limited to within 25-30 meters of surface.

As mining of the C and T horizons advanced and the initial capital investment
was recovered, the decision was made to evaluate the B1 horizon and exploration
drilling was focused on defining the deposit through drilling only to the bottom
of the B1 horizon.

As more knowledge was gained through mining of the B1 horizon, the potential of
the B2 horizon became increasingly important and exploration drilling was
extended to test the entire thickness of the C, T, B1 and B2 horizons.

As a result of the staged recognition of the mineral reserve potential at
Paracatu, several drill holes do not test the entire thickness of the B2
horizon.

After acquiring a 100% interest in RPM, Kinross reviewed the engineering support
prepared by RPM in support of a further mill expansion. At the same time,
Kinross evaluated the exploration potential at Paracatu and identified two
priority target areas:

        o       Deepening of holes in the northeast portion of the pit where the
                full extent of the B2 had not previously been defined and


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        o       Drilling to the west of Rico Creek where the B2 has been
                identified with similar characteristics as in the pit area but
                had been tested with a very limited number of drill holes.

In Q1, 2005, Kinross approved an exploration drill campaign totalling 30,000
meters and consisting of 154 diamond drill core holes. The purpose of this
program was to upgrade the Inferred mineral resources west off Rico Creek to
Measured and Indicated classification. A theoretical US$ 400 pit shell was used
to confine the drilling program.

Total costs for the program were estimated to be US$ 4.5 million. Drilling was
planned in Phases with subsequent phases contingent of results of the preceding
phase. All the planned drilling phases were completed prior to the November 2005
resource model however analytical results for 65 of the holes were pending when
the resource model was updated.

In addition to the drilling outlined above, in Q3, 2005, an additional drill
program was planned consisting of 50-75 diamond core holes (20,000 meters) that
were targeted to test the potential resources below the footwall contact defined
for the mineralized horizon below the existing mine pit in areas where
historical drilling was stopped short. Some holes were also targeted to test
lateral continuity of the mineralization beyond the limits that were in place
for the initial drill campaign. Total costs for this program were estimated to
be US $3.0 million.


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10.0    DRILLING

The 2005 exploration drill program was managed and supervised by B. Gillies, P.
Geo., Kinross Director of Exploration and C. Frizzo, Kinross Americas Project
Geologist.

The current database at Paracatu includes 458 test pits (5,070 meters) and 785
drill holes (42,489 meters). Table 10-1 summarizes the drill database as of July
29, 2005.




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               TABLE 10-1 DRILL HOLES SUMMARY TABLERAPHIC OMITTED]

- -------------------------------------------------------------------------------
     YEAR              CAMPAIGN       HOLE TYPE        NUMBER OF      TOTAL
                                      (diameter)         holes        meters
- -------------------------------------------------------------------------------
        1984     PMP                  6"                        44      2,462
   1983-1986     POCOS                PIT (1m)                 459      4,987
        1988     PAR                  6"                        26        708
        1989     PRF                  RC                        67      2,067
        1990     PRI                  6"                        15        465
        1992     PMP                  6"                        21        360
                 POCOS                PIT (1m)                  11         40
        1993     PMP                  6"                        33        686
                 PB2                  6"                         9        319
                 FPA                  6"                         8        240
                 POCOS                PIT (1m)                   9         29
        1994     PMP                  6"                        42      1,329
                 FPA                  6"                        35      1,261
        1995     PMP                  6"                        50      1,516
                 FPA                  6"                        22        802
        1996     PMP                  6"                        19        396
                 PB2                  6"                        10        753
                 FPA                  6"                        32      1,095
                 RAB                  6"                        21        592
                 ALB                  6"                        11        335
        1997     PMP                  6"                        52      1,650
                 PB2                  6"                        14        604
        1999     PMP                  6"                        29      1,320
        2000     PMP                  HX(3")                    20        600
                 PEC                  HX(3")                    38      3,597
        2004     PE                   HX(3")                    60      1,997
        2004     WCR                  HX(3")                     3      1,091
        2005     K                    HQ, HTW, NQ              267     48,660
- -------------------------------------------------------------------------------
TOTAL                                                     1,427     79,961
- -------------------------------------------------------------------------------

The database used in estimating mineral resources and reserves for this report
includes results from 89 drill holes completed in 2005.

Diamond drilling has demonstrated that anomalous gold grades (greater than 0.20
g/t Au) occur within a 125-150 meter thick tabular zone that has been traced


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for more than 4.0 km (NE-SW) by 3.0 km. (NW-SE). Anomalous gold grades remain
open down dip and laterally.

The portion of the deposit demonstrated to be economically viable is
approximately 3.0 km by 2.0 km in size.

Figure 10-1 is a plan map of the drill holes included in the resource model
documented in this report.

                       FIGURE 10-1 DRILL HOLE LOCATION MAP






                                    [PICTURE]







Included in the hole totals are 67 reverse circulation drill holes that were
drilled to test the mineralization. Assay results from the RC drill campaign
were 25 - 30 % lower than results from twinned diamond drill holes. The observed
bias is thought


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to be related to losses of gold in the dust that was produced during drilling,
some of it being retained inside the drill hole. RPM typically excludes RC data
where the data has been twinned by a diamond drill hole. Where holes have not
been twinned, RPM includes the RC results in the mineral resource modeling
process. Inclusion of the RC data in the mineral resource estimate does not have
any impact as the upper portions of the deposit tested with the RC holes have
been mined out.

All drill hole collars were established in field by RPM's mine surveyor using
standard Topcon GPS system. The drill hole is collared as close as possible to
the collar coordinates established by the surveyors with most holes collared
within 5 meters of plan.

All drill setups (-90 degrees) are checked by RPM geologists before beginning
drilling. RPM geologists controlled the hole shut down depths. A minimum of 20
meters of barren core (no arsenopyrite, no boudins), beyond the interpreted
footwall contact, was the criteria used to terminate drilling.

Several holes west of Rico Creek were surveyed using a downhole instrument. The
initial drill holes were surveyed using acid tube tests and a tropari. Deviation
was typically 2(Degree) per 100 meters. Azimuth readings from tropari were often
suspect.

Later in the program, an E-Z shot system was used. Results from the E-Z shot
instrument confirmed that some of the tropari readings were erroneous. Generally
pyrrhotite content was low enough that magnetic error is thought to be marginal.
Given the continuity and homogeneity of the mineralized zone and the wide
spacing of drilling, inclinometry variance is thought to have marginal effect.

Hole collars were surveyed again by the mine surveyor after drilling. 6 meter
PVC casing was placed downhole in as many collars as possible and collars were
cemented into a cairn, labelled, and photographed with landmark backgrounds. All
drill sites were cleaned up, drill cuttings removed and stored at the RPM waste
dump site and the water sumps were backfilled.


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Core was collected continuously from the collar. Wooden tags were placed in the
core trays and labelled according to the drill run. All core boxes were clearly
labelled with the hole number and drill interval. Lids were nailed on each core
box at the drill site to facilitate transport to the RPM logging facility.

Drill reports identified all zones of broken ground, fault zones and water gain
or loss. Water gain or loss was almost non-existent. Rusty water seams in the B2
horizon were almost non-existent, suggesting active hydrology occurs almost
exclusively in the weathered zone only.

10.1    DRILL SPACING

Until 1993, drilling and test pitting focused on the C and T horizons but since
that time, drilling has been extended into the B2 horizon. The nominal drill
spacing across the mineralized area east of Rico Creek roughly defines a 100 x
100 meter grid.

In 2005, the focus of Kinross' exploration efforts was the B2 horizon west of
Rico Creek. Kinross commissioned Dr. B. Davis, an independent consultant
specializing in geostatistical resource estimation, to complete a Drill Spacing
Study (Davis 05) to determine the optimal drill spacing required for defining
Measured and Indicated mineral resources at Paracatu.

The Drill Spacing Study is based on an estimation of confidence intervals for
various theoretical drill hole patterns. Spatial variation patterns are
incorporated in the variogram and the drill hole spacing can be used to help
predict the reliability of estimation for gold, arsenic, density and work index.
The measure of estimation reliability or uncertainty is expressed by the width
of a confidence interval or the confidence limits. By determining how reliably
gold, arsenic, density, and/or work index results must be estimated to meet
resource classification criteria, it is possible to calculate the drill hole
spacing necessary to achieve the target level of reliability


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Confidence intervals are intended to estimate the reliability of estimation for
different volumes and drill hole spacing. A narrower interval implies a more
reliable estimate. Using hypothetical regular drill grids and the variograms for
gold, arsenic, work index and specific gravity, confidence intervals or limits
can be estimated for different drill hole spacing and production periods or
equivalent volumes. The limits for 90% relative confidence intervals should be
interpreted as follows:

o       If the limit is given as 8%, then there is a 90 percent chance the
        actual value of production is within +/-8% of the estimated value for a
        volume equal to that required to produce enough ore tonnage in the
        specified period (e.g., quarter or full year). This means it is unlikely
        the true value will be more than 8 percent different relative to the
        estimated value (either high or low) over the given production period.

The method of estimating confidence intervals is an approximate method that has
been shown to perform well when the volume being predicted from samples is
sufficiently large. Dr Davis considered drill hole grids measuring 100 x 100
meters, 200 x 200 meters, and 300 x 300 meters in completing his study.

Further assumptions made for the confidence interval calculations are:

        o       The variograms are appropriate representations of the spatial
                variability for all variables

        o       Most of the uncertainty in metal production is due to
                fluctuations in the values of these variables

        o       Daily production rates range from about 17 - 50 Mtpa

Dr. Davis concluded that variability for density and work index at Paracatu was
marginal and not material to isolating optimum drill spacing. Confidence limits
for the gold and arsenic defined by different grids are shown in the Tables 10-1
and 10-2.


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               TABLE 10-2: CONFIDENCE LIMITS FOR GOLDPHIC OMITTED]

                    ---------------------------------------
                     DRILL GRID       17 MTPA     30 MTPA
                        (m)
                    ---------------------------------------
                     100 x 100           6.5%        4.9%
                     200 x 200           8.2%        7.5%
                     300 x 300          14.0%       13.0%
                    ---------------------------------------


               TABLE 10-3: CONFIDENCE LIMITS FOR ARSENICC OMITTED]

                    ---------------------------------------
                     DRILL GRID       17 MTPA     30 MTPA
                        (m)
                    ---------------------------------------
                     100 x 100           9.0%        7.9%
                     200 x 200          12.4%       10.8%
                     300 x 300          19.3%       18.0%
                    ---------------------------------------

Results for the 30 Mtpa production rate, the estimated production rate planned
for Expansion Project III at the time of Dr. Davis' work, are presented in
Figures 10-1 and 10-2.


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          FIGURE 10-1:GOLD ESTIMATION UNCERTAINTY BY DRILL HOLE SPACING





                                    [PICTURE]






        FIGURE 10-2: ARSENIC ESTIMATION UNCERTAINTY BY DRILL HOLE SPACING







                                    [PICTURE]








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Dr. Davis concluded that in order to support a classification of Indicated,
drill spacing should be maintained at a nominal 140 meter spacing. Drilling on a
200 x 200 meter grid pattern with a fifth hole in the center provides this drill
coverage. As a result, Kinross adopted the 200 x 200 meter five spot pattern for
their exploration work west of Rico Creek.

Sulphur content seems quite homogeneous and shows a very distinctive increase in
overall content at the bottom of the weathered-oxidized zone or top of water.
This is confirmed reasonably well with geological logging of first presence of
sulfides. Sulphur content can be used to define the contact of B1-B2, using the
geological log as backup.


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11.0    SAMPLING METHOD AND APPROACH

Core recovery from all diamond drill programs is reported to be excellent,
averaging greater than 95%. The greatest areas of core loss were from the collar
to 15.0 meters downhole in laterite zones. RPM employed a systematic sampling
approach where the drilling (and test pitting) were sampled using a standard 1.0
meter sample length from the collar to the end of the hole.

All samples were marked up and collected by geologists or technicians employed
by RPM.

It is standard practice at RPM to send the entire core for analysis after the
core had been logged and photographed. Reference pieces are 8 mm cores (1/ 4
meters) used for density and PLT testwork. These pieces are labelled and stored
at the core logging facility. This practice was continued for the duration of
sampling programs until Kinross acquired a 100% interest in RPM in 2004.

This practice of sampling large diameter core whole is not uncommon in deposit
with a low average grade and good grade continuity. Kinross does not consider
the sampling of whole core to be a concern especially when viewed in light of
the property's production history where typically, actual production is well
within 5% of estimated annual gold production.

It should be noted that only mineralized zones have been sampled. The remaining
non-mineralized core has been stored in metal tagged boxes both at the logging
facility and an enclosed secured storage building near the plant. Some core that
was assessed to be low grade was chip sampled in 2 x 5mm discs per 1 meter for
creating a single 8 meter composite (to fit with mining benches.) If the sample
returned close to 0.2 g/t au cut-off, the entire 8 meters was re-sampled in the
traditional 1 meter interval pattern. However, it is a very rare occurrence.


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11.1    BULK DENSITY AND CORE SPECIFIC GRAVITY

Bulk density analyses have been completed at various times throughout the
exploration and development of the project. The original values were based on
the results of samples collected from the surface test pits. Mining of the
deposit indicated that the bulk density values were low so efforts were made to
obtain a more representative number.

Changes were made to the calculation methodology and a linear regression method
was employed up to 1999. Reconciliation to actual production statistics
indicated problems with the density calculations and a study was commissioned to
examine the bulk density estimates.

Rio Tinto Technical Services Ltd (RTTSL) developed a new method that combined
statistical evaluation of near surface sampling for the C, T and B1 horizons
with a linear regression approach for the data within the B2 horizon in those
areas where deep drill coverage was limited. This new method has improved
reconciliation relative to the actual mill production to within 1.5% of
predicted tonnage figures.

At the mine, in situ density measurements are taken by extracting a 30cm cubic
block from the upper level of a bench. Generally two samples are taken and
averaged to give a value for the bench. The results from these samples will not
take into account any variations with depth and the density determination at the
top of the bench is applied through the entire depth of that bench (8.0 meters).

For the core samples, specific gravity is measured using the water displacement
method. This method is considered appropriate for the B2 horizon targeted in the
2005 exploration campaign.

A comparison between in situ density measurements and the recent specific
gravity measurements from the core samples shows the core being biased high with
an average difference of approx. 10%. The correlation in the B2 horizon improves
with increasing depth but in situ density information at depth in B2 is limited.
Since it will be the focus of future mining activity it was decided to use a


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reduction of 7% of the SG in the estimation process to produce a more
conservative estimate. It is recommended that more data be collected in the B2
horizon. The relationship between the in situ density on the bench and the core
specific gravity should be re-examined.

11.2    BOND WORK INDEX

Samples for Bond Work Index (BWI) testing are collected during sample
preparation of the 1.0 meter raw samples. Composite samples are based on an 8.0
meter downhole length representing the current mining bench height. Each
composite is composed of a fraction of each meter after initial sample crushing
to 2.0 mm. The BWI test is completed at the RPM process lab according to the
Bond Work Index standard test methodology.

KTS reviewed the lab's testing and quality control procedures and found them to
be within industry accepted industry standards.

The BWI composite data is used to interpolate the BWI for individual blocks in
the model using ordinary kriging.


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12.0    SAMPLE PREPARATION, ANALYSES AND SECURITY

12.1    SAMPLE PREPARATION AND ANALYSES

Prior to the start up of the mine, all samples were shipped to independent
analytical labs in Brazil for analysis. After construction of the mine, all
samples were processed at the on site lab by RPM employees. The RPM lab is not
an internationally certified analytical facility. Historically, gold assays were
completed on 50 g sample aliquots with a total of six (6) analyses done for each
sample. A sulphur assay value is also determined for each sample. Additional
elements assayed are arsenic, copper, lead, zinc, manganese, cadmium and silver.

In order to meet the demands of the 2005 drill program, Kinross contracted three
laboratories to perform analyses. They are listed below in decreasing order of
overall project workload.

        o       ALSChemex sample preparation facility in Luziania and ALSChemex
                analytical facility in Vancouver, Canada. 40% (ISO 9001
                Certified).

        o       Lakefield laboratories - Belo Horizonte, Brazil. 40% (ISO 17025
                Certified)

        o       RPM sample preparation and analytical facility, Paracatu. 20%
                (ISO 14001 Certified)

All facilities are ISO certified facilities.

The initial exploration program started with six (6) 50 g aliquots as per the
established procedure at RPM. A series of factors such as the number of samples
generated by the drill program, resulting requirements of the QAQC program,
workload and turnaround time at all commercial labs in Brazil forced Kinross to
re-evaluate different aspects of its exploration program.


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In May 2005 an audit of the RPM mine lab was undertaken by Kinross' Laboratory
Manager at the Fort Knox Mine to assess its equipment and procedures. Some
changes in preparation and fluxing were implemented resulting in markedly
improved productivity and QAQC performance. The variability between 50 g aliquot
was also reduced significantly.

In June 2005, Kinross commissioned a study by Agoratek International (Gy,
Bongarcon 05) to review exploration sampling procedures and assess the
requirements for six (6) 50 g aliquots assays per sample. Agoratek led by
Dominique Francois-Bongarcon, a recognized expert in sampling, reviewed the
sampling procedures and concluded that three (3) 50 g analyses would be
sufficient for the purposes of the exploration program.

Kinross standardized sample preparation and analytical procedures for all three
labs as closely as possible given equipment limitations and differences in
internal lab QA/QC protocols.

All three labs used fire assay with AA finish procedures on 3 x 50 g pulp
aliquots. Table 12-1 summarizes the sample preparation procedures employed by
the three laboratories in completing analyses for the exploration drill program
for the 89 holes added for this estimate.


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           TABLE 12-1 SUMMARY OF SIMPLE PREPARATION PROCEDURES BY LAB



                                                                             
- ------------------------------------------------------------------------------------------------------------------------------------
                    LAKEFIELD                                ALS CHEMEX                                          RPM
                    ---------                                ----------                                          ---
             Belo Horizonte, Brazil                       Luziania, Brazil                                Paracatu, Brazil
                                                          Vancouver, Canada
- ------------------------------------------------------------------------------------------------------------------------------------

- ------------------------------------------------------------------------------------------------------------------------------------
                     DRYING                                    DRYING                                       CRUSHING (1)
Total sample 100(degree) - 110(degree)C    Total sample 100(degree) - 110(degree)C         Total sample 100% < 1cm
                                                                                           Renard jaw crusher
                                                                                           Air cleaning every sample
                                                                                           LS cleaning every 20 samples

                                                                                                            CRUSHING (2)
                                                                                           Total sample, 95% <2.4mm
                                                                                           Renard roll crusher
                                                                                           Air cleaning every sample
                                                                                           LS cleaning every 20 samples
- ------------------------------------------------------------------------------------------------------------------------------------

- ------------------------------------------------------------------------------------------------------------------------------------
                    CRUSHING                                  CRUSHING                                         DRYING
                    --------                                  --------                                         ------
Total sample 90% < 2mm                     Total sample 90% < 2mm                          Drying, 2kg:110(degree) - 120(degree)C
Rhino jaw crusher                          Rhino jaw crusher
Air cleaning every sample                  Air cleaning every sample
Qtz cleaning every 40 samples              Qtz cleaning every 20 samples
Sieve test every 20 samples                Sieve test every 20 samples
- ------------------------------------------------------------------------------------------------------------------------------------

- ------------------------------------------------------------------------------------------------------------------------------------
                  PULVERIZATION                             PULVERIZATION                                   PULVERIZATION
                  -------------                             -------------                                   -------------
2kg: 95% < 150 mesh                        2kg: 95% < 150 mesh                             2kg: 90% < 100#
LM2 pulverizers                            LM2 pulverizers                                 Setamil pulverizer
Air cleaning every sample                  Air cleaning every sample                       Silica cleaning every sample
Qtz cleaning every 40 samples              Qtz cleaning every 20 samples
Sieve test every 20 samples                Sieve test every 20 samples
- ------------------------------------------------------------------------------------------------------------------------------------

- ------------------------------------------------------------------------------------------------------------------------------------
                  FINAL SAMPLES                             FINAL SAMPLES                                   FINAL SAMPLES
                  -------------                             -------------                                   -------------
3- 50g aliquots                            150g opacked for FA/AA analysis                 3-50g aliquots
FA/AA analysis                             ALS Chemex Vancouver, Canada                    FA/AA analysis
- ------------------------------------------------------------------------------------------------------------------------------------

- ------------------------------------------------------------------------------------------------------------------------------------
                 INTERNAL QA/QC                            INTERNAL QA/QC                                  INTERNAL QA/QC
                 --------------                            --------------                                  --------------
Batch size = 50 aliquots                   Batch size = 84 aliquots                        Batch size = 30 aliquots
1 standard                                 2 standard                                      1 standard
1 blank                                    1 blank                                         1 blank
2 duplicates                               3 duplicates
- ------------------------------------------------------------------------------------------------------------------------------------


12.2    SECURITY

All core boxes are shut with nailed wooden lids and transported by RPM personnel
from Geoserve or Geosol rigs to the logging facility located inside the fenced
mine gates. After photographing, logging and sample mark-up (1.0 meter standard
core interval), the whole core is placed in heavy gauge plastic bags with


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a unique sample tag. The sample tag number is also written in indelible marker
on the outside of each sample bag.

Samples to be analyzed at the RPM lab are loaded by RPM personnel onto prickup
trucks and transported to the RPM crushing facility. After crushing, samples are
again transported by pickup truck to the RPM preparation lab where samples are
riffle split. Approximately 6 kgs are stored as a coarse rejects and 2 kgs are
transported by pickup truck to the RPM assay lab for pulverization and analysis.

Samples that are to be analyzed by either Lakefield or ALS Chemex are loaded
onto transport trucks operated by the respective labs and delivered to the
respective sample preparation facilities in Belo Horizonte or Luziania.

Sample collection, preparation, transportation and analysis have all been
completed to industry standards. The samples used to estimate the mineral
resources and reserves described herein are, in the author's opinion, of
sufficient quantity and quality to support the resource classification.


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13.0    QUALITY CONTROL, QUALITY ASSURANCE

Quality Control and Assurance for the 2005 drilling was managed by B. Gillies
P.Geo, Kinross' Director of Exploration and R. Peroni, RPM's Director of
Technical Services.

Quality control and quality assurance programs were limited during early
exploration at Paracatu. The dominant quality control procedure involves the use
of inter-laboratory check assays comparing results from RPM's analytical lab to
Lakefield Research in Canada. Additional check assay work was carried out at the
Anglo Gold laboratories in Brazil (Crixas and Morro Velho).

Currently, inter-laboratory checks are run against all RPM's samples including
flotation rejects (low grade), geology samples (intermediate grade) and hydromet
plant samples (high grade). Results from the inter-laboratory check assaying
have not been reviewed by the author.

The RPM lab procedure includes insertion of certified analytical standards and
blanks. At least one blank and standard is inserted with each batch (30 samples)
analyzed. Results are statistically analysed and if they lie outside the
determined boundaries, all the samples within the batch are repeated. Other
checks are also conducted throughout the fire assay process, such as lead
recovery to the buttons and silver recovery for the prills. If recoveries are
below the criteria, the analyses are repeated.

For the 2005 exploration program, all procedures have been under direct control
of RPM KTS staff.

A QA/QC program was implemented for the three labs used during the 2005
exploration program. The program consists of inserted standards and blanks in
the sample streams. All three labs also reported using round robin checks. The
labs were visited on an infrequent and unannounced basis by RPM representatives.
No major sample preparation discrepancies were noted. The ALSC analytical
facility in Vancouver was not visited.


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Kinross purchased certified standard from Rocklabs (New Zealand) in two lots.
The standards were selected to meet typical Morro do Ouro grade ranges. These
standards were OXA26, OXC30, OXD27, SE19, SF12. Their certified values and
acceptable limits are listed in Table 13-1

                 TABLE 13-1: STANDARDS AND THEIR ACCEPTED LIMITS



- ------------------------------------------------------------------------------------------------------------------
                                                            CERTIFIED
   STANDARD            CERTIFIED        CERTIFIED           STANDARD            CERTIFIED           ACCEPTED
   (REF #)               VALUE         VARIABILITY          DEVIATION         QA/QC LIMITS        QA/QC LIMITS
                       (AU G/T)         (AU G/T)            (AU G/T)            (AU G/T)            (AU G/T)
- ------------------------------------------------------------------------------------------------------------------
                                                                                  
OxA26                   0.080           +/- 0.006                  -         0.068 to 0.092      0.065 to 0.095
OxC30                   0.200           +/- 0.014                  -         0.172 to 0.228      0.165 to 0.235
OxD27                   0.416           +/- 0.025                  -         0.366 to 0.466      0.354 to 0.478
SE19                    0.583           +/- 0.011          +/- 0.026         0.529 to 0.637      0.518 to 0.648
SF12                    0.819           +/- 0.012          +/- 0.026         0.763 to 0.875      0.751 to 0.887
- ------------------------------------------------------------------------------------------------------------------


For blanks, a local crushed (gravel 1-2 cm) calcareous metasiltstone was used
but was clearly identifiable by its white colour.

A model numbering code system was generated that could accommodate the 3
different batch sizes of the 3 labs. Table 13-2 presents a comparison between
internal QAQC for the labs and the QAQC system implemented by Kinross for the
2005 exploration-drilling program.


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               TABLE 13-2: SUMMARY OF QAQC BY LABORATORYC OMITTED]



- -----------------------------------------------------------------------------------------------------------------------------------
                                            INTERNAL LAB QA/QC                                   CLIENT QA/QC
- -----------------------------------------------------------------------------------------------------------------------------------
   Lab          Batch Size       Standards        Blanks     Duplicates            Standards       Blanks      Samples / batch
- -----------------------------------------------------------------------------------------------------------------------------------
                    (#)             (#)            (#)          (#)                   (#)           (#)              (#)
- -----------------------------------------------------------------------------------------------------------------------------------
                                                                                                
Chemex              84               2              1            3                     2             3               73
- -----------------------------------------------------------------------------------------------------------------------------------
Lakefield           50               1              1            2                     1             2               43
- -----------------------------------------------------------------------------------------------------------------------------------
RPM                 30               1              1            0                     1             1               26
- -----------------------------------------------------------------------------------------------------------------------------------


Each batch contained a minimum of one standard and one blank per analytical
furnace tray. Standards were numbered according to the number model and were
shipped in a separate bag to be inserted into the sample stream at the
preparation facilities. The standards were inserted in a manner that assured
that the analytical lab would not be able to identify the standards from the
submitted samples. But, as five different standards were used, it is a
reasonable to assume that they satisfy the requirement that they be blind.

13.1    RESULTS

Results available are from March 1, 2005 to August 11, 2005 and include data for
103 exploration holes analyzed by RPM, Lakefield and ALSChemex.

Results received to date for the certified standards indicate that ALS Chemex is
returning results that are 2% higher than the certified standard values,
Lakefield is 4% lower and RPM's lab is returning results that confirm the
certified standards. As sample lots were shipped to all three labs throughout
the program, no one lab significantly dominates a spatial area of the
mineralized resource.

Overall results returned from all labs were well within industry accepted
tolerences with failure rates of 1,6% to 2,7% for the analyses performed. A
failure on a standard is classified as +/- 2 standard deviations from the
certified mean for each standard.

All failures occurring within the identified mineralized horizon were requested
to be re-run Results for the failures noted during the exploration program are
pending and corrections, (if necessary) will be made to the database on receipt


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of re-run results. Given the low number of failures it is unlikely that the
changes (if warranted) will result in a material difference in the estimate.

A significant number of swaps between standards were noted possible due to
sample numbering mistakes by the geologists inserting the standards or
transcription errors at the receiving labs. Sample swaps were readily
identifiable when plotting standard performance.

Overall laboratory performance is summarized in Table 13-3


         TABLE13-3: LABORATORY PERFORMANCE SUMMARY FOR 2005 EXPLORATION

              ----------------------------------------------------
                  Lab          Standards     Failures     Swaps
                                  (#)           (#)        (#)
              ----------------------------------------------------
               RPM               1004           44         28
               Chemex            1233           20         11
               Lakefield         1470           81         33
              ----------------------------------------------------

          Figures 13-1 to 13-3 summarize QA/QC standard results by lab.




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                   FIGURE 13-1: STANDARD PERFORMANCE - RPM LAB










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                 FIGURE 13-2: STANDARD PERFORMANCE - ALS CHEMEX












                                    [PICTURE]













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                  FIGURE 13-3: STANDARD PERFORMANCE - LAKEFIELD












                                    [PICTURE]













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While, in general, standards performance of all three labs is considered
acceptable, QAQC analysis indicates a large number of standards sample "swaps"
have occurred. The source of these swaps has not been determined yet. RPM
logging staff onsite has been repeatedly reminded about labelling errors and
minor procedural adjustments have been made to reduce these occurrences.

The exploration geologists in charge of the 2005 program reviewed the results of
the standards analyses and filtered the data to isolate the reruns with the
biggest potential to reduce confidence in the resource estimate. After
identifying all outlier values, the outliers were examined to determine if there
was a failure or were the results related to a swap of standards. The outliers
identified as failures were then evaluated relative to their position within the
mineralized zone (HWZ vs FWZ), their position within the $400 pit limit and the
position relative to other sample data. All these factors were evaluated to
filter the outlier values with the greatest potential to affect the resource
model.

Based on these filters several intervals from different holes, analyzed by
different labs, were selected for rerun. Given the low number of failures it is
unlikely that the changes (if warranted) will result in a material difference in
the estimate

13.2    RERUNS

A total of 308 samples from 16 hole intervals were selected for reruns at the
respective labs:

        o       Lakefield: 198 samples / 8 intervals of 6 holes;

        o       RPM : 62 samples / 4 intervals of 3 holes;

        o       ALSChemex : 48 samples / 4 intervals of 4 holes.

The reruns confirmed the sample variance observed betweenthe individual aliquot
analyses. Typical results from a portion of the rerun analyses are provided in
Table 13-4.


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                        TABLE 13-4 SELECTED RERUN RESULTS


                                                                             
- ---------------------------------------------------------------------------------------
  HOLE   SAMPLE                              INITIAL ANALYSIS
- ---------------------------------------------------------------------------------------
   (#)     (#)   Aliquot 1   Aliquot 2   Aliquot 3   Aliquot 4   Aliquot 5   Aliquot 6
- ---------------------------------------------------------------------------------------
                 (Au g/t)     (Au g/t)   (Au g/t)    (Au g/t)     (Au g/t)   (Au g/t)
- ---------------------------------------------------------------------------------------
  K-508    170     1.16         0.48       5.80        1.24         0.67       0.86
- ---------------------------------------------------------------------------------------
  K-512    235     0.66         0.79       1.43
- ---------------------------------------------------------------------------------------
  K-601    112     0.56         0.71       0.27        0.95         0.48       0.55
- ---------------------------------------------------------------------------------------
  K-601    175     0.35         0.59       0.89        6.31         0.61       0.54
- ---------------------------------------------------------------------------------------
  K-1-5    128     0.65         1.27       0.19
- ---------------------------------------------------------------------------------------
  K-510    179     1.46         0.86       1.12
- ---------------------------------------------------------------------------------------
  K-207    26      0.94         0.53       0.58        0.57         0.43       0.76
- ---------------------------------------------------------------------------------------
  K-207    28      0.63         0.34       1.13        0.62         1.10       0.39
- ---------------------------------------------------------------------------------------
  K-613    171     0.13         0.16       0.09
- ---------------------------------------------------------------------------------------
  K-908    222     1.08         0.93       1.46
- ---------------------------------------------------------------------------------------
  K-908    226     1.11         0.89       1.88
- ---------------------------------------------------------------------------------------


- ------------------------------------------------------------------------
  HOLE   SAMPLE                         RERUN ANALYSES
- ------------------------------------------------------------------------
   (#)     (#)     Avg    Aliquot 1   Aliquot 2   Aliquot 3   Result
- ------------------------------------------------------------------------
                (Au g/t)   (Au g/t)   (Au g/t)    (Au g/t)   (Au g/t)
- ------------------------------------------------------------------------
  K-508    170    1.70       0.90       0.88        1.82       1.19
- ------------------------------------------------------------------------
  K-512    235    0.97       1.04       1.26        1.32       1.20
- ------------------------------------------------------------------------
  K-601    112    0.58       0.59       1.38        0.89       0.94
- ------------------------------------------------------------------------
  K-601    175    1.54       0.81       0.38        0.37       0.52
- ------------------------------------------------------------------------
  K-1-5    128    0.71       0.23       0.28        0.13       0.21
- ------------------------------------------------------------------------
  K-510    179    1.15       1.42       1.26        0.69       1.12
- ------------------------------------------------------------------------
  K-207    26     0.64       0.11       0.09        0.07       0.09
- ------------------------------------------------------------------------
  K-207    28     0.70       0.17       0.06        0.09       0.11
- ------------------------------------------------------------------------
  K-613    171    0.12       0.11       0.83        0.19       0.42
- ------------------------------------------------------------------------
  K-908    222    1.15       3.60       1.05        0.83       1.83
- ------------------------------------------------------------------------
  K-908    226    1.29       0.93       1.28        0.61       0.93
- ------------------------------------------------------------------------


Results also indicated that the grade variance is reduced when comparing the
averages of the individual aliquots. Of the 16 intervals rerun, 14 returned
average grades that were +/- 0.04 g/t Au. The remaining two intervals
demonstrated greater variability (0.11 g/t). The correlation coefficients
calculated for both the first analysis and rerun results, for each lab, were
0.72 to 0.80 respectively.

Table 13-5 summarizes the rerun results for the 16 batchs submitted for rerun
analysis.


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               TABLE 13-5 SUMMARY OF BATCH RERUNS[GRAPHIC OMITTED]

        ---------------------------------------------------------------
             LAB         HOLE        BATCH      INITIAL       RERUN
                          (#)       Sample       Result      Result
                                      (#)       (Au g/t)    (Au g/t)
        ---------------------------------------------------------------
        Lakefield        K-508      112-135      0.562        0.506
                         K-508      166-190      0.365        0.407
                         K-506      163-179      0.311        0.301
                         K-512      226-242      0.387        0.404
                         K-601      084-150      0.590        0.647
                         K-601      151-200      0.508        0.563
                         K-1-5      88-152       0.281        0.236
                         K-510      163-188      0.762        0.646
        ---------------------------------------------------------------
        CHEMEX           K-207       26-42       0.255        0.222
                         K-211      109-125      0.453        0.388
                         K-613      161-177      0.349        0.321
                         K-205       19-35       0.203        0.222
        ---------------------------------------------------------------
        RPM              K-407      198-214      0.458        0.424
                         K-908      98-114       0.427        0.319
                         K-908      206-227      0.729        0.741
                         K-116      172-204      0.324        0.335
        ---------------------------------------------------------------

Evaluation of the rerun data is difficult as the results mimic the results
observed in comparing individual sample aliquots. It is difficult to reproduce
grades due to the nugget effect albeit the effect is tempered by the low grade
nature of the deposit.

Figure 13-4 demonstrates the gold grade variance between individual aliquots of
the initial analysis and rerun analysis for reruns from hole K-508, sample
numbers 112-135.


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       FIGURE 13-4 - K-508 SAMPLES 112 TO 135 INITIAL VS RERUN BY ALIQUOT










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No reruns have been requested due to blanks failures. The number of blank
failures in ore zones to date is regarded as minimal.

13.3    ROUND ROBIN TESTS - COARSE AND PULD REJECT ANALYSES

Two round robin inter lab tests are currently in progress. Coarse and pulp
rejects (300 of each), selected from holes drilled in the mineralized zone west
of Rico Creek, were sent for round robin analysis at the three labs used during
the exploration program. Results of the round robin analyses are pending at this
time.

13.4    LAB BIAS

With three separate labs involved in analyzing the core collected from the drill
program the likelihood of lab bias materially affecting the estimate is
considered low. Figure 13-5 presents a drilling plan for the 2005 exploration
program showing the drill hole location and identifying the primary lab that
completed the analysis. The plan demonstrates the good distribution between the
three labs, highlighting the fact that no one lab is concentrated in one area of
the deposit. It is believed the distribution


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     FIGURE 13-5 PLAN VIEW - DIAMOND DRILLING DISTRIBUTION BY ANALYTICAL LAB











                                    [PICTURE]










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14.0    DATA VERIFICATION

Kinross has not completed any independent verification routines against original
data sources. Rio Tinto employed a rigorous data verification process at
Paracatu where the database was manually verified against original assay and
field certificates.

Rio Tinto Technical Services completed bi-annual reviews of RPM's procedures and
methodology. The review process was very detailed and generally involved 2-3
full days of detailed review and verification. Results of the reviews are
maintained in RPM's archives. The 1998, 2000 and 2002 reviews concluded that
RPM's procedures met Rio Tinto's corporate guidelines for resource modeling and
reserve estimation.

For the December 31, 2005 model, Kinross independently verified 10% of the data
collected between 1999 and 2004 against original source documents. The holes
were chosen at random and any errors against original sources were documented.
Results identified a single transcription error was made in the arsenic values
for an entire hole. No other errors were identified.

For the 2005 drill program, Kinross' exploration geologists managing the program
verified all data. Gold grades were all double entered and weight averaged per
sample, then the two databases were crosschecked with no significant errors or
differences detected. As and S assays have been cross checked at the time of
this report.

The summary database spreadsheet was compared to the individual digital files
sent by the different laboratories. Kinross is confident that the database is
sufficiently free of errors to support the present mineral resource and mineral
reserve estimates.

Paracatu's production history suggests that the accuracy of the data is beyond
reproach. Kinross has reviewed the production accounting records in detail and
have found these to be exceptionally detailed and thorough. Kinross is confident


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that the production reconciliation data is accurate and indicative of the
performance of the reserve estimate.

Table 14-1 summarizes the production reconciliation for the period 1990 to 2004.

                  TABLE 14-1 PARACATU PRODUCTION RECONCILIATION



- ---------------------------------------------------------------------------------------------------------
            YEAR               1990     1991      1992      1993      1994      1995     1996      1997
- ---------------------------------------------------------------------------------------------------------
                                                                          
Reserve Grade (Au g/t)        0.652    0.631     0.590     0.517     0.485     0.505    0.519     0.486
Actual Grade (Au g/t)         0.644    0.613     0.575     0.499     0.497     0.492    0.502     0.465
Mine Call Factor              0.988    0.971     0.975     0.965     1.025     0.974    0.967     0.957
- ---------------------------------------------------------------------------------------------------------
            YEAR               1998     1999      2000      2001      2002      2003     2004      2005
- ---------------------------------------------------------------------------------------------------------
Reserve Grade (Au g/t)        0.514    0.472     0.467     0.471     0.438     0.446    0.439    0.442
Actual Grade (Au g/t)         0.482    0.453     0.473     0.449     0.483     0.438    0.442    0.423
Mine Call Factor              0.938    0.960     1.013     0.953     1.103     0.982    1.007    0.956
- ---------------------------------------------------------------------------------------------------------


For the 2005 drill program, Kinross' exploration geologists managing the program
verified all data. Gold grades were all double entered and weight averaged per
sample, then the two databases were crosschecked with no significant errors or
differences detected. As and S assays have been cross checked at the time of
this report. QAQC procedures are on going. Batch reruns are in process of being
redone if standards exceeded 2 standard deviations from mean and if the
standards failure occurred within mineralized zones.

The summary database spreadsheet was compared to the individual digital files
sent by the different laboratories. Kinross is confident that the database is
sufficiently free of errors to support the present mineral resource and mineral
reserve estimates.

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15.0    ADJACENT PROPERTIES

There are no other producing mines near the Paracatu mine. . Fazenda Lavras is a
gold prospect located approximately 13 km from Paracatu. It shows some
similarities with the Paracatu deposit but it is not in production.




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16.0    MINERAL PROCESSING AND METALLURGICAL TESTING

The metallurgical and processing information presented herein was collected
under the supervision of L. A. Tondo, RPM's Manager of Projects, W. Phillips,
Kinross Americas Director of Technical Services and R. Henderson, P. Eng.,
Kinross' Director of Technical Services.

The resource and reserve estimates summarized by this report assume modification
of the existing plant according to Expansion Project III, which consists of the
installation of an in pit crushing and conveying system (IPCC), a 38 foot
diameter SAG mill, two 24 foot diameter ball mills operating in closed circuit
with cyclones, four new jigs, a new flotation plant and an upgrade of the
existing hydrometallurgical plant.

16.1    EXISTING PROCESS PLANT

The existing process plant at Paracatu has operated continuously since 1987 and
has had expansion upgrades in 1997 and 1999. In 2005, the plant processed 17.2
Mtpa and achieved an average gold recovery of 78.2%. A detailed discussion on
the existing process facilities is presented in Section 20.0 of this report. In
summary the plant consists of primary and secondary crushing, ball milling to
80% passing 75 micron, gravity recovery using jigs, rougher and cleaner
flotation, concentrate regrinding and cyanide leaching (Hydromet Plant). Final
gold bullion is produced from the carbon adsorption, desorption and
electrowinning circuit.

Table 16-1 summarizes the average annual metallurgical recoveries of the
flotation and hydrometallurgical process as well as the average global plant
recovery for the Paracatu plant since commercial production began.


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             TABLE 16-1 PROCESS PLANT METALLURGICAL RECOVERY SUMMARY



- -------------------------------------- ------ ------ ------ ------ ------ ------ ------ ------ ------ ------
YEAR                                    1987   1988   1989   1990   1991   1992   1993   1994   1995   1996
- -------------------------------------- ------ ------ ------ ------ ------ ------ ------ ------ ------ ------
                                                                     
Hydromet Recovery (%)                     NA   95.1   97.4   97.5   99.1   99.2   99.2   99.2   99.2  99.3
Flotation Recovery (%)                    NA   83.8   84.8   84.6   83.7   83.7   81.8   79.5   76.4  76.7
- -------------------------------------- ------ ------ ------ ------ ------ ------ ------ ------ ------ ------
GLOBAL METALLURGICAL RECOVERY (%)       59.0   75.7   82.4   82.7   83.3   83.2   81.4   78.8   75.8  76.0
- -------------------------------------- ------ ------ ------ ------ ------ ------ ------ ------ ------ ------
Year                                    1997   1998   1999   2000   2001   2002   2003   2004   2005  TOTAL
- -------------------------------------- ------ ------ ------ ------ ------ ------ ------ ------ ------ ------
Hydromet Recovery (%)                   97.5   92.2   94.3   96.2   96.7   97.1   96.8   96.3   96.3  97.2
Flotation Recovery (%)                  75.6   77.9   77.8   78.8   80.9   81.3   79.1   79.8   81.2  80.4
- -------------------------------------- ------ ------ ------ ------ ------ ------ ------ ------ ------ ------
Global Metallurgical Recovery (%)       73.7   71.8   73.4   75.8   78.3   79.0   76.6   76.8   78.2  77.9
- -------------------------------------- ------ ------ ------ ------ ------ ------ ------ ------ ------ ------


16.2    EXPANSION PLAN

The Paracatu Expansion III Project is the product of a number of years of
testing, development and planning. In 2002, RPM took action to counter the
gradually increasing work index of the deposit. The existing circuit was not
designed for hard ore and capacity and operating costs would be significantly
affected unless additional grinding capacity was installed.

A SAG mill pilot plant program was run in 2002/2003 and in 2004, a Feasibility
Study for Expansion Plan III was completed by ECM, a Brazilian engineering firm.
Aker-Kvaerner contributed technical expertise to ECM's study. This study
recommended expanding the current 18 Mtpa process facility to 30 Mtpa with the
addition of an in pit crushing and conveying (IPCC) system, a 38 foot diameter
semi-autogenous grinding (SAG) mill and expansion of the existing gravity
circuit.

In January 2005, Kinross and RPM commenced the exploration drill program west of
Rico Creek and became aware of the potential for a significant reserve increase.
A Plant Capacity Scope Study was completed in June 2005, which evaluated several
alternatives to increase plant throughput. All options considered in the Study
assumed the installation of an in pit crushing and conveying system (IPCC) and
38 foot diameter

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Semi-Autogenous Grinding (SAG) mill which were the cornerstone assumptions in
the original Feasibility Study.

The Plant Capacity Scope Study recommended that production be increased from 18
Mtpa to 50 Mtpa. The Expansion III Project will proceed in two stages over a
four year period commencing in 2006. The first stage will increase plant
capacity from 18 to 32 Mtpa. The new 32 Mtpa SAG mill plant will be constructed
and once commissioned, the existing 18 Mtpa plant will be shut down and
refurbished. Once refurbishment of the 18 Mtpa plant is complete, it will be
restarted and tasked with processing the remaining B1 reserve. This will bring
total plant throughput for the two lines to 50 Mtpa. When the soft BI ore is
depleted in 2017, the throughput capacity will be limited to 41 Mtpa and then
capacity will decrease further as work index increases above a value of 11 in
2024.

In Q4, 2005, the Basic Engineering for Expansion Plan III was awarded to
SNC-Lavalin Engineers and Constructors Ltd, an internationally recognized
consulting engineering and construction company and MinerConsult Engenharia, a
Brazilian engineering firm. The scope of work included the IPCC, covered
stockpile, new 32 Mtpa mill, hydromet expansion, electrical substation, tailings
delivery and water systems. Process design details were finalised and purchase
orders were awarded for the SAG mill and ball mills. The basic engineering
designs and supporting capital and operating costs estimates form the basis of
the 2006 Feasibility Study.

The following sections provide additional details on Expansion Project III.

16.2.1  IN PIT CRUSHING AND CONVEYING

The in pit crushing and conveying system (IPCC) is an integral component to the
Expansion Project III.

The IPCC is sized to treat 4870 tph of run of mine (ROM) ore. To meet this
specification, RPM estimates the system must average 75% availability. The
system will consist of a primary feed hopper that can be fed from three sides by
218 tonne trucks. Run of mine


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(ROM) ore will be fed by an apron feeder to a MMD-type, twin shaft sizer and ore
will be reduced to 80% passing 200 mm. The MMD sizer is the industry standard
for limestone quarry operations (WI=12.0) and is considered adequate to treat
the harder B2 ore type in the mineral reserve base. Samples of the hardest B2
ore were sent to the manufacturer for testing, results indicated that the ore
hardness was not an issue but the abrasive characteristics of the ore may
increase cost due to accelerated wear. Alternate primary gyratory crusher trade
off studies were completed during basic engineering and confirmed that the MMD
sizer was the most appropriate installation.

Ore will be transported by a 1.5 km long, 1.8 meter wide conveyor to a covered
stockpile. Material handling testwork and design by Jenike and Johansen has
shown that the ore will be prone to ratholing. Consequently, the stockpile
design incorporates eight transverse belt feeders to maintain a live capacity of
12 hours. The total stockpile capacity of 24 hours will provide enough buffer
capacity to prevent production disruption during IPC maintenance and during the
rainy season when rainfall reduces mining efficiency in the pit.

16.2.2  NEW 32 MTPA MILL

New feed to the SAG mill is rated at 3971 t/h. The grinding circuit will consist
of one 11.6 m diameter by 6.7 m long (38' diameter by 22' long) SAG mill
followed by two parallel 7.3 m diameter by 12.0 m long (24' diameter by 39.5'
long) ball mills. The SAG mill will operate in closed circuit with a trommel
screen and vibrating screen and the ball mills will operate in closed circuit
with hydrocyclones. Plant capacity has been selected to give a nominal flotation
feed grind of 80% passing 75 im.

The SAG mill will be driven by a 15,000 kW (20,000 HP) gearless wrap-around
motor. The wrap-around motor has inherent variable speed capability, which is
required to efficiently process the wide range of ore hardness anticipated. The
ring gear driven ball


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mill will be powered by twin clutch and pinion gears driven by two 6,500 kW
(8,720 HP) fixed speed motors.

A bank of four radial jigs will be installed in the new mill building to recover
coarse arsenopyrite particles from the ball mill circulating load. The jig
concentrate will flow by gravity to a vertical mill regrind circuit where it
will be reground, thickened and pumped to the CIL leach circuit.

The flotation circuit will consist of rougher flotation followed by a single
stage of cleaning. The cleaner flotation concentrate will be pumped to a
vertical mill regrind circuit where it will be reground, thickened and pumped to
the CIL leach circuit. Cleaner tails will be recirculated to the rougher
flotation cells and rougher tailings will be discarded to final tailings. A
total of 21 rougher cells (160 m(3) tanks cells) are included, arranged in three
rows of seven cells each.

16.2.3  TAILINGS

The increase in throughput will require an increase in tailing capacity. Golder
Associates Ltd are currently studying a second tailing dam with sufficient
capacity for containing 1.5 billion tonnes of tailings The new dam will also
serve as a water catchment area and reservoir, providing additional water that
will be required for the 50 Mtpa production level. Sulphide tailings from the
hydrometallurgical plant will continue to be stored in lined ponds.

A pre-feasibility study examining the installation of a sulphuric acid plant for
treating hydrometallurgical plant tailing is in progress. This is an
environmentally preferable alternative with the potential to reduce disposal
costs and potentially generate revenues from acid sales.


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16.2.4  MODIFICATIONS TO THE EXISTING PLANT

The Expansion Project III is being developed with a strategy designed to
minimize disruption to the current operation.

The new grinding plant will be a stand alone circuit that will feed its own
flotation cells..The only interaction between the existing circuit and the new
circuit will occur at the existing hydrometallurgical plant. This plant will be
upgraded to cope with the increase in concentration production. The
hydrometallurgical plant will be designed to maintain throughput at 100tph
(equivalent to 50 Mtpa mill feed) and additional equipment will be required for
the regrinding, CIL, elution, carbon regeneration and electro winning circuits t

The increase in flotation and hydrometallurgical capacities will ensure that
process residence times will not be reduced due to the increase in the ore
processing rates. Therefore it is expected that current gold recoveries will be
maintained after the proposed expansion.

16.3    EXPANSION PLAN III METALLURGICAL TESTWORK

The test work supporting the installation and operation of the SAG mill
originated from a series of 64 pilot plant tests conducted on the Paracatu ores.
The tests were run on 1,500 tonnes of Paracatu ore with WIs ranging from 5.5 to
12.0 kWh/t. In all, six different ore types were processed through a Koppers 6x2
foot SAG mill that was leased from CETEM, Rio de Janeiro, Brazil. The pilot
plant operated from April 2002 to February 2003. A staff of two process
engineers, 3 technicians and 10 laborers were permanently assigned to the pilot
plant operation.

The samples are considered to be representative of the variability in ore
hardness expected during the remainder of the mine life.


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The pilot plant testwork and analysis of the results were all completed under
the supervision of a team of recognized expert in the filed of SAG mill design
and operation. These experts were:

        o       Mr. Anthony Moon, Rio Tinto Technical Services;

        o       Dr. Steve Morrell, SMCC and

        o       Mr. George Grandy, Aker-Kvaerner.

        o       Dr Homero Delboni Jnr, University of Sao Paulo

The pilot plant test work evaluated ores independently as well as composite ores
formed by blending the available ore types together to produce a representative
blend of future mill feed.

Specific details on the pilot plant testwork are included in a 2004 Feasibility
Study. The results were reviewed by Dr. Morrell and Mr. Grandy who independently
concluded that a 38 foot diameter SAG mill with a 3,700 tph throughput rate
would be best suited to process the Paracatu ores.

This study was later updated to the 50 Mtpa level and used as the basis for the
Expansion III Project. The major modifications were the addition of two 24 x 40
foot ball mills that will permit the SAG mill to be run in open circuit. This
circuit design reduces the risk of high volumes of slurry going through the SAG
mill, by eliminating the circulating load (cyclone underflow) going back to SAG
mill.

Figure 16. -1 presents a graph showing mill throughput related to the ore
hardness. At the design ore hardness work index of 11 the SAG mill in open
circuit has a maximum throughput of 4000 tph.


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        FIGURE 16. -1 SAG MILL PERFORMANCE CURVE (MORRELL REVISED CURVE)





                                   [PICTURE]





Mineralogical studies carried out at the JKMRC-MLA laboratories in Australia
have shown that a large part of the Paracatu plant gold losses were associated
with mixed particles of arsenopyrite with gold. Figure 16. -2 illustrates a
typical occurrence.




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           FIGURE 16. -2 TYPICAL GOLD ON ARSENOPYRITE GRAIN BOUNDARIES





                                    [PICTURE]





The relatively large natural size of the arsenopyrite crystals in the deposit
makes them readily recoverable by gravity concentration. The JKMRC-MLA
mineralogical study showed that at 65 mesh, 90 % of the arsenopyrite crystals
are liberated. Since thin section analysis has demonstrated that arsenopyrite
crystals contain gold, increasing arsenopyrite recovery also results in
increased gold recovery. RPM has studied options to improve arsenopyrite
recovery from the ore. An obvious alternative for achieving this objective is to
improve gravity concentration efficiency. After the pilot plant testwork results
were analyzed, a number of optimization efforts were made in the current


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industrial jigging circuit, leading to an improvement in arsenopyrite (and gold)
recovery for some of the arsenic rich ores. The main change in operating
parameters was the removal of the steel shot previously being used as ragging to
create the jig dense media bed. It was found that the coarse arsenopyrite
crystals in the ore are sufficient to create an autogenously bed in the jigs.
The problem of bed compaction, resulting from the steel shot agglomerating after
operating for a number of hours, was thus eliminated. This resulted in a more
consistent production of jig concentrate, which in turn improved overall
recovery of the circuit. For the Expansion Project III, the use of jigs treating
part of the ball mill circuit-circulating load is being incorporated into the
process design. A modification of the existing system will be made: PAN AMERICAN
style jigs will be used instead of the current YUBA design. Testwork showed that
a PAN AMERICAN jigs achieve a more consistent concentrate production. This type
of jig is more robust and can fluidise the dense media bed more effectively,
thus resulting in better mass recovery to the concentrate, without prejudicing
concentrate quality.

In 2002, RPM joined the AMIRA Program P260D and as a project sponsor, RPM was
entitled to have an extensive program of fieldwork conducted in the plant at
Paracatu. Researchers from three institutions (University of Sao Paulo, CETEM in
Rio and IWRI from Australia) conducted a series of measurements in the
laboratory and industrial scales tests. They discovered that one of the major
factors limiting efficient arsenopyrite recovery in the RPM flotation circuit
was being caused by chemical oxidation of arsenopyrite surfaces during the
treatment in the plant. The conclusion was that the key for success in improving
flotation performance at RPM was to find a new suite of reagents that could cope
with this problem. In 2005, two new collectors developed by a large reagent
producing company were successfully tested in the process lab, and have resulted
in improved gold recovery.

The metallurgical recovery of gold decreases with increasing sulphur and arsenic
content. Laboratory testwork has been conducted on core samples to replicate the


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proposed flowsheet. The data has been factored to correspond with actual plant
operation and the following equation has been established:

Recovery = (a +(-2.36230 x S%) +(-0.0017 x As ppm)) x b) where

a = theoretical maximum flotation recovery of 85.95352% and

b = theoretical hydrometallurgical recovery or 96.5%












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17.0    MINERAL RESOURCE AND RESERVE ESTIMATES

Mineral resources and classification were estimated by M. Belanger, P.Geo,
Kinross Americas Director of Technical Services and Dr. R. Peroni, RPM's
Director of Technical Services.

Mineral reserves were estimated by K. Morris, P. Eng., Kinross' Manager of Open
Pit Mining.

W. Hanson, P.Geo., Kinross' Vice President of Technical Services supervised the
preparation or the resource and reserve estimates.

The mineral resource model for Paracatu is interpreted and estimated using
Vulcan(C) software. The modeL was updated December 31, 2005. The model
incorporates the results from 228 out of the 267 drill holes completed in 2005.
These holes were drilled to test the down dip extent of the deposit to the west
of Rico Creek and the extension of the B2 below the pit floor.

The estimate is based on a revised geological interpretation. The interpretation
is based on geological factors observed in the drill core where there is a
direct relationship between gold grade and the frequency of boudins,
asymmetrically folded quartz veins and arsenopyrite content. Boudin frequency
and arsenopyrite content are directly proportional with gold grade.

Ore hardness (BWI) and metallurgical recovery are estimated for each block in
the model.

Mineral reserves are estimated within design pits developed from optimized pit
shell using Whittle 4X(C) software, a program that has become a standard in the
mining industry.


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Kinross is not aware of any reason that would materially affect the resource and
reserve estimate. There is reasonable certainty that all necessary permits will
be obtained to allow continued exploitation of the resources and reserves at
Paracatu.

17.1    MINERAL RESERVE AND RESOURCE STATEMENT

The Proven and Probable mineral reserve estimate for the Paracatu mine as of
December 31, 2005 is summarized in Table 17-1. Proven and Probable mineral
reserves are estimated at a gold price of US$ 400 per ounce and a Foreign
Exchange Rate (FEX) of 2.65 Reais per US $1.00. The estimate is based on the
assumptions and costs documented in the Plant Capacity Scoping Study, June 2005.
The cut off grade used to report mineral reserves is 0.21 g/t Au.

       TABLE 17-1 PROVEN AND PROBABLE MINERAL RESERVES - DECEMBER 31, 2005


         -------------------------- ----------- ----------- -----------
                CLASSIFICATION         TONNES      GRADE        GOLD
                                     (X 1,000)   (AU G/T)     (OUNCES)
         -------------------------- ----------- -----------  -----------
         Proven                      1,103,677        0.40    14,194,000
         Probable                       83,131        0.38     1,016,000
         -------------------------- ----------- -----------  -----------
         PROVEN & PROBABLE           1,186,808        0.40    15,210,000
         -------------------------- ----------- -----------  -----------


Table 17-2 summarizes the Measured and Indicated mineral resource estimate
(excluding mineral reserves) for the Paracatu mine as of December 31, 2005 at a
gold price of US $450 per ounce and a Foreign Exchange Rate (FEX) of 2.65 Reais
per US $1.00. The cut off grade used to report mineral resources is 0.19 g/t Au.


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     TABLE 17-2 MEASURED AND INDICATED MINERAL RESOURCES - DECEMBER 31, 2005


         -------------------------- ----------- ----------- -----------
                CLASSIFICATION         TONNES      GRADE        GOLD
                                     (X 1,000)   (AU G/T)     (OUNCES)
         -------------------------- ----------- -----------  -----------
         Measured                       89,784        0.27      771,000
         Indicated                       5,540        0.38       68,000
         -------------------------- ----------- -----------  -----------
         MEASURED & INDICATED           95,324        0.27      839,000
         -------------------------- ----------- -----------  -----------

 NB MEASURED AND INDICATED RESOURCES ARE REPORTED EXCLUSIVE OF MINERAL RESERVES

In addition to the Measured and Indicated mineral resources stated in Table 1-2,
Paracatu hosts an Inferred resource of 40.1 million tonnes averaging 0.37 g/t
Au. Inferred resources are estimated at a gold price of US $450 per ounce and a
FEX of 2.65 Reais per US $1.00.

The resource and reserve estimates stated above are classified according to the
Canadian Institute on Mining, Metallurgy and Petroleum (CIM) Standards on
Mineral Resources and Reserves.

The mineral resources and mineral reserves estimates were completed by RPM's
staff and supervised of Wes Hanson, P.Geo, Kinross' Vice-President of Technical
Services.

Approximately 65% of the mineral resources and mineral reserves lie below the
current water table. The Preliminary License (PL) for extending the mine pit
below the water table has recently been approved by the Environmental Regulatory
Authorities.

The mineral resources and reserves for the project are hosted entirely on mining
leases and exploration concessions controlled by RPM. RPM is the sole owner of
the sub-surface mineral rights for all of the resource and reserve estimates
disclosed herein.

The mineral rights to these lands are controlled by RPM through the exploration
concessions. Permits to allow mining have, as yet, not been granted. RPM has
indicated


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that the necessary permits can be obtained once the decision to mine
the reserves on these exploration concessions has been confirmed and the proper
reports filed with DNPM. There is no reason to suggest that the necessary
permits will be denied.

17.2    HISTORICAL ESTIMATES

The reserve history at Paracatu indicates continuous growth of the reserve base
reflecting increased geological knowledge and improved process efficiencies.
Figures 17-1 and 17-2 are graphs that show the changes in mineral reserve
tonnages and contained ounces from the start of commercial production until
December 31, 2005.


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        FIGURE 17-1 TONNAGE MINED AND IN RESERVE AS OF DECEMBER 31, 2005





                                   [PICTURE]





         FIGURE 17-2 OUNCES MINED AND IN RESERVE AS OF DECEMBER 31, 2005





                                   [PICTURE]






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The historical resource and reserve estimates for Paracatu have been classified
according to the JORC Code. There are no significant differences between the
JORC resource and reserve estimates and the CIM classification described in this
report.

17.3    MODELING METHODOLOGY

17.3.1  OVERVIEW

The Paracatu resource model was updated as of December 31, 2005.

The resource and reserve estimate reported herein is based on the topographic
mining surface as of December 31, 2005.

A total of 267 diamond drill holes, have been added to the project database.
Table 17-3 summarizes the data added to the estimation database.

                     TABLE 17-3: UPDATED DRILL HOLE DATABASE

  ---------------- --------- --------- --------- --------- --------- ---------
                    GEOLOGY     GOLD    ARSENIC   SULPHUR     BWI       SG
                   --------- --------- --------- --------- --------- ---------
  # drill holes         267       228       141       110       111       234
  # DATA POINTS      48,660    30,334    19,681    14,883     1,699     9,080
  ---------------- --------- --------- --------- --------- --------- ---------


17.3.2  GEOLOGICAL INTERPRETATION

The mineral resource model for Paracatu is developed from a series of oriented
drill sections on which all exploration results have been plotted. Major fault
zones are interpreted from section to section, typically as a linear feature.
Observation of the drill core is used to define the A (waste)-C-T-B1 and B2
contacts, which are interpreted on individual sections as surfaces and later
converted to three-dimensional solids.

Previous models, estimated by RPM staff, interpreted the Calha, non-Calha and
IDS ore types on sections based on the arsenic content. The Calha, non-Calha,
IDS interpretation was used to assign global recovery in the model.


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Historically, grade interpolation for the Paracatu resource model interpolated
grades into a broad zone defining the entire thickness of the zone. This
modeling methodology produced a non-layered model (NLM) that failed to isolate
zonation within the hangingwall and footwall contacts of the zone.

Logging of the exploration core collected in 2005 has identified several
important geological clues that can be used to visually identify zonation within
the mineralized horizon. The observations are consistent with the strong
structural controls proposed by Holcombe.

Unmineralized phyllites exhibits well-developed lamination, largely due to
original bedding that dips at about 10(Degree) to the SW. Figure 17-3 shows
bedding structures typically observed in the host phyllites.

              FIGURE 17-3 GRADED BEDDING IN UNMINERALIZED PHYLLITE




                                   [PICTURE]




Anomalous gold grades correspond to the first and last occurrence of
arsenopyrite and mark the hangingwall and footwall contacts of the mineralized
zone which ranges from 120 to 150 meters in thickness and averages greater than
0.40 g/t. Pyrite ranges from 1-3% as fine laminae and arsenopyrite ranges from
trace to 1/2% as fine needles and grains typically less than 1 mm in size. Shear
cleavage begins to develop and, as strain
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increases, trends from a 20(Degree) SW dip to parallel to bedding. Interfolial,
isoclinal folds can be observed.

Gold grades increase steadily from the hangingwall and footwall contacts towards
the center of the zone where strain is highest. Gold grades increase in direct
proportion to the size and frequency of boudins (bedding and quartz), intensity
of shear banding, the presence of asymmetric folds where axial plane cleavage
begins to parallel bedding and the amount and size of arsenopyrite grains which
in the higher grade zones tends to occur as coarse porphyroblasts. Figures 17-4
and 17-5 typify structural textures and arsenopyrite mineralization within the
high strain zone.

    FIGURE 17-4 PHYLLITE WITH VERGING ASYMETRIC FOLDS, SHEAR BANDS & BOUDINS





                                   [PICUTRE]





                         Green - verging asymetric folds

                         White - shearing

                         Yellow - Foliation boudins



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              FIGURE 17-5 LARGE ARSENOPYRITE PORPHYROBLAST IN CORE




                                   [PICTURE]




The visual guides noted during core logging were used to create an updated
geological model for the mineralized phyllite to the west of Rico Creek and the
B2 identified below the actual pit bottom.

For the mineralization west of Rico Creek, the mineralized horizon has been
divided into two distinct zones producing a layered interpretation. First, a
global B2 zone is defined by the geologists based on the first and last
occurrences of arsenopyrite and/or deformation features. This step defines the
mineralized envelope from hangingwall to footwall. The overall thickness ranges
from 120 meters to 150 meters.

Within the B2, RPM geologists have identified the Boudin Deformation Zone (BDZ)
a zone of more intense deformation characterized by an increase in the presence
and size of boudins, and in the intensity of shear banding. The BDZ ranges in
thickness from 60 to 80 meters and averages 0.60 g/t Au.

East of Rico Creek, core logging has not identified the BDZ. The mineralization
is therefore interpolated within the B1 horizon and a broad B2 zone defining the
entire thickness of the mineralized horizon as modelled by RPM geologists.


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Figure 6-5 presents a conceptual model of the geology of the Paracatu deposit
outlining the layered interpretation. Figures 17-5 and 17-6 present typical
exploration drill results west of Rico Creek. The BDZ is represented by the
>0.40 g/t outline while the overall mineralized interval is represented by the
yellow outline which corresponds to the first and last occurrence of
arsenopyrite.

                  FIGURE 17-5 DRILL SECTION 07N - LOOKING NORTH






                                   [PICTURE]





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                  FIGURE 17-6 DRILL SECTION 05N - LOOKING NORTH




                                   [PICTURE]




The zone limits and, where applicable, the individual layers, are digitized and
imported into Vulcan(C) mine modeling software. Vulcan(C) is used to convert the
sectional polygons and lines to three-dimensional wireframes and surfaces
representing the mineralized units and features that have been interpreted.

17.4    SAMPLE ANALYSIS

The 1.0 meter raw sample data are extracted and grouped by using the wireframes
to clip out the sample data. For gold, the populations were separated for B1 and
B2 (east of Rico Creek); B2 and BDZ (west of Rico Creek).

Statistical analysis of the 1.0 meter samples indicates that within the defined
mineralized horizons, gold grades have excellent lateral and downdip continuity.
Table 17-4 summarizes the basic statistics of the 1.0 meter raw sample data for
gold.


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                                       TABLE 17-4 BASIC STATISTICS FOR GOLD, RAW SAMPLE DATA

- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------
                 DOMAIN                NUMBER OF      MEAN      MEDIAN     MINIMUM     MAXIMUM      COEFFICIENT OF     STANDARD
                                        SAMPLES                                                       VARIATION        DEVIATION
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------
                                                                                                       
B1                                         14,538     0.440       0.380       0.00         9.90            0.774            0.342
B2                                         39,015     0.360       0.310       0.00         7.01            0.830            0.301
B2 (Boudin-deformation zone)               12,102     0.440       0.370       0.00         5.43            0.730            0.320
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------


17.4.1  ARSENIC

Assay data for arsenic was used, in conjunction with sulphur analyses, to
estimate a metallurgical recovery for each model block as per the recovery
equation detailed in Section 17.8.4 of this report.



                                           TABLE 17-4 BASIC STATISTICS FOR ARSENIC ASSAYS

- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------
                 DOMAIN                NUMBER OF      MEAN      MEDIAN     MINIMUM     MAXIMUM      COEFFICIENT OF     STANDARD
                                        SAMPLES                                                       VARIATION        DEVIATION
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------
                                                                                                     
B1                                          1,905        759        612           0        6702             1.220          929.41
B2                                         24,243       1148        662           0      41,687             1.320         1513.96
B2 (Boudin-deformation zone)                1,238       1148       2675         250      10,988             0.430         1191.02
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------


17.4.2  BOND WORK INDEX

Hardness is assessed based on 8.0 m composite samples that represent the mine`s
bench height. Each sample is composed of a fraction of each meter after initial
sample crushing to 2.0 mm.

BWI composite data for the resource model was used to interpolate BWI estimates
into each model block. The composite data for the B1, B2 and BDZ were extracted
and interpolated separately. BWI interpolation used a nearest neighbour
interpolation to estimate the BWI of individual model blocks.

Basic statistics tabled below in Table 17-6 highlight the difference in ore
hardness between the B1 and B2 horizons.


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                                  TABLE 17-6: BASIC STATISTICS FOR BOND WORK INDEXRAPHIC OMITTED]

- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------
                                       NUMBER OF      MEAN      MEDIAN     MINIMUM     MAXIMUM      COEFFICIENT OF     STANDARD
                                        SAMPLES                                                       VARIATION        DEVIATION
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------
                                                                                                        
BWI (total)                                 1,990      10.94      12.17        0.59       20.30             16.31            4.04
BWI (B1)                                      213       4.65       4.39        0.76       14.97              3.46            1.86
BWI (B2)                                    1,445      11.83      12.65        2.38       18.60             10.99            3.32
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------  --------------


17.4.3  SPECIFIC GRAVITY

Specific gravity measurements for core samples are collected and assessed based
on 4.0 m composite samples comprised of 8.0 cm core intervals selected for every
2.0 meters of core. As shown in Table 17-7 the core specific gravity
measurements show minimal spread around the mean with a coefficient of variation
of 0.04. The higher specific gravity results are related to an increase in the
sulphide content.



                                 TABLE 17-7: BASIC STATISTICS FOR SPECIFIC GRAVITY IN CORE SAMPLES

- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------
                 DOMAIN                NUMBER OF      MEAN      MEDIAN     MINIMUM     MAXIMUM     COEFFICIENT OF
                                        SAMPLES                                                      VARIATION
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------
                                                                                           
B1                                          1,593       2.45       2.15        2.03        2.87              0.05
Total                                      10,674       2.76       2.81        1.89        4.42              0.05
- ------------------------------------ ------------- ---------- ---------- ----------- -----------  ----------------


17.5    COMPOSITING

After reviewing the statistics of the raw data, the 1.0 meter raw samples are
composited into 6.0 meter composite intervals. Compositing uses a bench
compositing routine with the 6.0 meter composite length is equivalent to half
the planned mining bench height.


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The composite data is then extracted using the same geological wireframes used
to evaluate the raw 1.0 meter sample results. Each composite is coded according
to the geological unit used for the extraction.

Any duplicate (twinned) composites are discarded. During the interpolation
process the composites are length-weighted to account for composites with a
length shorter than 3.0 meters.

Composite statistics are evaluated in exactly the same manner that the 1.0 meter
sample data was evaluated as a check against any introduced error resulting from
the compositing process. No errors were noted in comparing the composite sample
statistics against the raw sample data.

17.6    GRADE CAPPING AND RESTRICTING OF HIGH GRADE

Grade capping for original 1.0 m assays is considered on a zone by zone basis.
High-grade results occasionally occur in the 1.0 m sample results. Cumulative
probability plots were calculated for B1, B2 and BDZ. A capping grade of 1.4 g/t
was selected for both B1 and B2 based on the 99th percentile of the grade
distribution. Within the BDZ the capping level was set at 1.6 g/t.

17.7    GEOSTATISTICS

The 6.0 m composites for the different variables are then subjected to
geostatistical analysis. First a downhole correlogram is calculated to determine
the nugget to be used in a fitted model. Directional correlograms are then
computed to define the direction of best continuity. For gold, different
correlograms are used for the B1 and B2 ores. For the blocks west of Rico Creek
the B2 horizon is further divided into overall B2 and the BDZ domains with their
own variography and estimation parameters.


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Table 17-8 summarizes the correlogram models estimated for gold, arsenic,
sulphur, Bond Work Index and density.



                                                                           
                          TABLE 17-8: PARACATU CORRELOGRAM SUMMARY

- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
                                              ROT.   RANGE   ROT.            ROT.
                                      Sill     Z      Z'     X'               Y'    Range
   ZONE    ITEM  STR. TYPE   NUGGET                                RANGE X'           Y'
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
    B1      Au    2    Sph   0.320    0.288   -64    63.5    -15     72.0     21    76.9
                       Sph            0.392   -45    74.2     5     1229.8     3    769.8
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
    B2      Au    2    Sph   0.212    0.405   -62   277.6   -106     71.5     -6    78.0
                       Sph            0.383  -121   145.2    0.6    1615.2    -3   1706.4
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
   BDZ      Au    2    Sph   0.248    0.473   59     65.7    -10     65.7     60    142.9
                       Sph            0.279   -11   101.7    -1     2173.5     3    816.8
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
    B1      As    2    Sph   0.300    0.542   58     17.7     1     100.0     -4    257.4
                       Sph            0.158   50    121.5    -1     853.2      1   1847.2
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
    B2      As    2    Sph   0.234    0.376   -41    75.4     7      64.5     71    139.2
                       Sph            0.390  -131   100.0    -6     803.0      1   1231.0
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
   BDZ      As    2    Sph   0.220    0.513   -18    63.9    -4      41.2     -72   156.6
                       Sph            0.267   -21    93.0    -4     4368.8     0    826.5
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
    B1       S    2    Sph   0.130    0.721    2     10.9    30      64.9      2    155.0
                       Sph            0.149   35     65.7     1     517.5     48    209.3
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
    B2       S    2    Sph   0.050    0.468   -59    94.8    75      47.9     35    68.8
                       Sph            0.482    4    297.1    -3     5063.3    15   1318.9
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------
   BDZ       S    2    Sph   0.054    0.595   -6     34.8     2      27.7     90    115.0
                       Sph            0.351  -119   147.6    -7     919.1      3   1994.5
- --------- ------ ---- ----- -------- ------- ------ ------- ----- --------- ------ --------


17.8    BLOCK MODEL

The block model is created using a two-step process. First, a block model with a
50 x 50 x 12 meter (x,y,z) block dimension is coded using the same geological
wireframes used to evaluate the sample data.

The block model was initialized in Vulcan(C) using the following parameters:


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X MIN = 6,000

Y MIN = 8,000

Z MIN = 56

NUMBER OF BLOCKS X DIMENSION = 128

NUMBER OF BLOCKS Y DIMENSION = 84

NUMBER OF BLOCKS Z DIMENSION = 65

17.8.1  GRADE INTERPOLATION

Gold grades are interpolated using Ordinary Kriging with each geological unit
(zone) estimated independently. The zone solids are used as hard boundaries and
the composites must have the identical domain code item as the solids to be used
in the interpolation process. At Paracatu assay grade capping is set at the 99th
percentile of the gold grade for the zone being estimated. It results in capping
grades of 1.4 g/t Au for both B1 and B2 and 1.6 g/t Au for the BDZ.

An octant search is used in all cases for grade interpolation. A minimum of 1
composite and a maximum of 10 composites are used within the search ellipsoid. A
maximum of four adjacent samples are used from the same drillhole.
Discretization is as follows: 5 steps in the X direction, 5 steps in the Y
direction, and 2 steps in the Z direction for a total of 50 discretization
points.

Table 17-9 summarizes the search parameters used to control grade interpolation
in the resource model for all items in the different zones. It is assumed that
the regional NE trend represents the dominant control for all mineralization
types. First correlogram structures with much shorter ranges such as the NW
structure observed in B2 mineralization are accounted for in the kriging
algorithm.


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                               TABLE 17-9 GRADE INTERPOLATION PARAMETERS

- -----------------------------------------------------------------------------------------------------
  VARIABLE     ROCKTYPE     BEARING        DIP        PLUNGE      RADIUS 1    RADIUS 2     RADIUS 3
                           (degrees)    (degrees)    (degrees)    (meters)    (meters)     (meters)
- ------------ ----------- ------------ ------------ ------------ ------------ ----------- ------------
                                                                          
Au            B1                  46           -3            5      1230.00       770.00       75.00
             ----------- ------------ ------------ ------------ ------------ ----------- ------------
              B2                 239           -3            6      1710.00      1615.00      150.00
             ----------- ------------ ------------ ------------ ------------ ----------- ------------
              BDZ                 79           -1            3      1000.00       400.00      100.00
- ------------ ----------- ------------ ------------ ------------ ------------ ----------- ------------
AS            B1                  50           -1            1      1850.00       850.00      125.00
             ----------- ------------ ------------ ------------ ------------ ----------- ------------
              B2                 229           -6            1      1231.00       803.00      100.00
- ------------ ----------- ------------ ------------ ------------ ------------ ----------- ------------
              BDZ                 69            0           -4         2000          800         200
- ------------ ----------- ------------ ------------ ------------ ------------ ----------- ------------



17.8.2  SPECIFIC GRAVITY

Correlograms were calculated and models fitted. Block densities were estimated
using a nearest-neighbour interpolation method on a zone by zone basis.

17.8.3  ORE HARDNESS

Each model block is assigned an ore hardness based on the results of the BWI
analyses. BWI values are interpolated into the model blocks using a
nearest-neighbour assignment.

17.8.4  RECOVERY

Unique metallurgical recoveries are estimated for each model block (50 x 50 x 12
meters) based on the arsenic and sulphur analytical results of drill core
analysis. The interpolation method used is ordinary Kriging.

The metallurgical recovery is based on the following equation.

Recovery = (a +(-2.36230 x S%) +(-0.0017 x As ppm)) x b) where

a = theoretical maximum flotation recovery of 85.95352% and

b = theoretical hydrometallurgical recovery or 96.5%


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17.8.5  MODEL CHECKING

Grade tonnage tables at a range of cut off grades were used to determine the
impact the changes in geological interpretation have had on the model's
predictive capability. Historically, production at Paracatu, based on mill
production statistics, agrees well with the resource model. Paracatu's 18 years
of production history and the detailed reconciliation to the reserve estimates
confirms the predictive accuracy of the historic resource model grade
estimation. The data indicates that after processing more than 237.0 M tonnes of
ore, the estimated grade is within 2% of the actual grade as measured by the
process plant.

With this standard in mind, it was necessary to confirm that changes in the
modeling method have not materially affected the overall grade distribution
within the model limits. Based on the different modeling methodology between the
Layered Model (LM) and the Non-Layered Model (NLM) for the mineralization west
of Rico Creek where the LM was employed. It would be expected that the LM would
result in less tonnes at a higher grade, a result of confining higher grade
values to a higher grade zone, as opposed to diluting the value of this
mineralization with lower grade material on the periphery as is the case in the
NLM estimation methodology.

Grade tonnage distributions at various cutoff grades for the portion of the
model wet of Rico Creek is summarized in Table 17-10 for both the LM and NLM
models.


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                          TABLE 17-10 COMPARISON OF LM VS NLM WEST OF RICO CREEK

- -------------- -------------------------------------------- --------------------------------------------
                                    LM                                           NLM
- -------------- -------------------------------------------- --------------------------------------------
    CUTOFF         TONNES           AU             AU           TONNES         AU              AU
   (AU G/T)     (T X 1,000)      (AU G/T)       (OUNCES)      (T X 1,000)    (AU G/T)        (OUNCES)
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
                                                                         
        0.10         427,943           0.46      6,301,472        427,974           0.46      6,301,935
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
        0.20         423,217           0.46      6,272,701        427,974           0.46      6,301,935
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
        0.30         374,965           0.49      5,870,977        412,243           0.47      6,189,581
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
        0.40         274,512           0.54      4,730,614        319,219           0.50      5,100,776
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
        0.50         165,049           0.59      3,141,410        127,298           0.57      2,316,481
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
        0.60          61,975           0.67      1,335,003         20,551           0.71        471,104
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
        0.70          15,488           0.75        374,966          7,256           0.87        201,803
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
        0.80           2,276           0.84         61,468          4,016           0.96        123,947
============== ============== ============== ============== ============== ============== ==============


At a 0.20 g/t cut off grade, roughly equivalent to the economic cut off grade
estimated by Whittle 4X(C), the LM model contains virtually the same amount of
gold with the tonnage and grade being within 1% of each other. The data supports
the conclusion that within the mineralized horizon at the economic cut off grade
level, the two models have the same level of contained ounces.

17.9    RESOURCE CLASSIFICATION

Paracatu historically reported resources and reserves classified according to
the AusIMM JORC Code. JORC is essentially identical to the CIM Standards, which
are the required reporting format under Canada's National Instrument NI 43-101.

The resource and reserve estimates dated December 31, 2005 and described in this
report, are classified according to the Canadian Institute on Mining, Metallurgy
and Petroleum (CIM) Standards on Mineral Resources and Reserves.

Model classification is based on drill density and confidence limits. Resource
blocks are classified as Measured if; the grade within a grouping of blocks
equal to the average rate of annual production, is estimated to +/-5.0% accuracy
with a 90% confidence level. In other words, in 9 out of 10 years, the average
grade of all mill feed will agree within 5% of


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that predicted by the model. Blocks are classified as Indicated if; the grade
within a grouping of blocks equal to the average quarterly production, is
estimated to a +/-10% level of accuracy with a 90% confidence level.

The Drill Spacing Study completed at RPM suggests that indicated resources can
be delineated from a 140-meter grid and measured resources from a grid spacing
of less than 100 meters. It is important to note that the highest estimation
variability is associated with arsenic and not gold. The drill spacings
recommended in the Drill Spacing Study are shorter than optimal for gold due to
the fact that arsenic is more variable.

The Drill Spacing Study also indicated that reducing the grid spacing to less
than 100 meters will not significantly increase confidence limits. This suggests
that drilling on spacings of less than 100 meters will not increase the
predictive accuracy of the estimate.

The calculations of confidence intervals only consider the variability of grade
within the deposit. There may be other aspects of deposit geology and geometry
such as geological contacts or the presence of faults that would impact the
drill spacing. However, based on the overall knowledge of the deposit after 18
years on mining experience and the demonstrated continuity of the B1 and B2
horizons, KTS used the following classification scheme:

o       Measured resources require a minimum of three samples from three holes
        within a 100 meter distance of the block that is being estimated;

o       Indicated resources require a minimum of three samples and a minimum of
        one hole with a 140 meter distance of the block being estimated;

o       All remaining mineralized model blocks are classified as Inferred.


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o       Block classification checked manually to determine any blocks which may
        require re-classification if the geologist feels that grade and/or
        geological continuity warrants an increase or decrease in confidence of
        the block value.

Block classification checked manually to determine any blocks which may require
re-classification if the geologist feels that grade and/or geological continuity
warrants an increase or decrease in confidence of the block value.

17.10   PIT OPTIMIZATION

17.10.1 BASE CASE

The design process for the open pit mine at Paracatu began by completing a
series of pit optimizations in order to create a pit shell that would form the
basis, or template, for the pit design. Pit optimization was performed by Kevin
Morris, P.Eng, Kinross' Manager of Open Pit Mine Engineering. Mr Morris has more
than 20 years of industry experience in the optimization and design of open pit
mines.

Pit optimization for the Paracatu open pit was completed using proprietary
software known as Whittle 4X(C). This software uses the Lerchs-Grossman
algorithm. The optimization proceeds by mining blocks that add value to the pit
shell. In other words, an individual block is released to the optimum shell only
if the mining of that block, along with the cumulative values of all blocks
within the pit shell, produces an overall net positive cash flow.

Prior to optimization, the grade tonnage curve from the Vulcan(C) model is
compared to the grade tonnage curve for the model imported to Whittle 4X(C) to
ensure there are no transcription errors during the manipulation from one
software system to the other. Table 17-11 summarize the grade tonnage summaries
of the Vulcan(C) model as compared to the model imported to Whittle 4X(C).


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                     TABLE 17-11 GRADE TONNAGE SUMMARY OF IMPORTED AND EXPORTED MODEL

- ----------------------------------------------------------- --------------------------------------------
                           EXPORTED MODEL (VULCAN)               IMPORTED MODEL (DATAMINE-WHITTLE)
               -------------------------------------------- --------------------------------------------
    CUTOFF         TONNES           AU             AU           TONNES         AU              AU
   (AU G/T)     (T X 1,000)      (AU G/T)       (OUNCES)      (T X 1,000)    (AU G/T)        (OUNCES)
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
                                                                           
   0.1           2,979,995            0.34      32,575,057      2,885,865           0.34     31,546,099
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.2           2,489,940            0.38      30,420,281      2,396,884           0.38     29,283,390
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.3           1,766,342            0.43      24,419,344      1,680,937           0.43     23,238,636
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.4             989,167            0.50      15,901,217        934,357           0.50     15,020,126
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.5             428,106            0.58       7,969,308        408,128           0.58      7,610,534
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.6             121,552            0.66       2,594,902        117,173           0.66      2,486,350
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.7              25,226            0.75         610,709         24,193           0.75        583,367
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.8               3,320            0.84          89,555          3,125           0.84         84,396
- -------------- -------------- -------------- -------------- -------------- -------------- --------------
   0.9                 225            0.92           6,655            225           0.92          6,655
- -------------- -------------- -------------- -------------- -------------- -------------- --------------


The minor differences noted in the table are believed to be software related and
are not considered to be material.

Optimization parameters included the operating costs, process recovery, metal
price and pit slope angles. The optimization parameters used for this design
exercise are presented in Table 17-12 and represent the Base Case. Operating
cost assumptions are based on the operating costs as estimated in the June 2005
Plant Capacity Scoping Study.




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               TABLE 17-12: BASE CASE OPTIMIZATION PARAMETERSTTED]

     --------------------------------- -----------------------------------
                 PARAMETER                            VALUE
     --------------------------------- -----------------------------------
                Mining Cost                        $0.53/tonne
     --------------------------------- -----------------------------------
              Mining Recovery                         100%
     --------------------------------- -----------------------------------
              Mining Dilution                          0%
     --------------------------------- -----------------------------------
              Pit Slope Angles                         55o
     --------------------------------- -----------------------------------
          Process Cost (incl. G&A)          Contained in Model Blocks
     --------------------------------- -----------------------------------
            Process Recovery Au             Contained in Model Blocks
     --------------------------------- -----------------------------------
              F.E.X. (R$:US$)                        2.65:1
     --------------------------------- -----------------------------------
                 Gold Price                         $US400/oz
     --------------------------------- -----------------------------------
                Selling Cost                  $7.90/ounce (1.976%)
     --------------------------------- -----------------------------------
                    DCFR                               5%
     --------------------------------- -----------------------------------
              Throughput Rate                        41 Mtpa
     --------------------------------- -----------------------------------

Process recoveries were estimated during the modeling process with a unique
process recovery estimated for each 50 x 50 x 12 meter model block. The process
costs were calculated within the block model based on the bond work index (WI)
that was also estimated during resource modeling.

Process costs were estimated as a Process Cost Adjustment Factor (PCAF) in
Datamine(C) prior to exporting to Whittle 4X(C). In Whittle 4X(C) the base
process cost was set at $1.00 per tonne. The base cost was then adjusted during
optimization based on the PCAF formula presented below.


        PCAF = 23.7/(88.2-4.5(WI)) + 0.1158(WI)

A similar method was used to estimate the mining cost of each block. A Mining
Cost Adjustment Factor (MCAF) was established in the block model. The MCAF
increased


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costs, as the pit deepened by a quantity of $0.02 per 12-metre bench
starting at pit exit. For this exercise the pit entrance/exit was assumed at the
800-elevation.

Figure 17-7 graphically summarizes the base case optimization results for
Paracatu disclosed above.

                   FIGURE 17-7 BASE CASE WHITTLE 4X(C) RESULTS




                                   [PICTURE]



The optimum pit shell is that which produced the highest average cash flow
discounted at 7%. The average cash flow is based on coarse schedules produced
within Whittle 4X(C). One schedule the Best Case, mines the nested pit shells in
a series of "push backs". The other schedule, the Worst Case, mines each shell
"bench by bench" to


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exhaustion. Neither schedule is truly valid however the average is a better
approximation then either the best or worst case.

The pit shell selected for Paracatu was shell number 17. It was not the
"optimum" shell as it did not produce the highest discounted cash flow. KTS
elected to utilize pit shell 17 as the basis for final pit design because in the
opinion of Kinross, a 2% loss in cash flow for a gain of 16% in contained ounces
was an acceptable risk.

17.10.2 CUT-OFF GRADES

Resources and reserves are reported above a minimum cut off grade that
represents the incremental cut off. That is to say it does not mining costs.
Mining costs are considered during pit optimization to determine if a block in
the model will be mined or not mined by the optimum pit. The incremental cut-off
grade represents the cut off grade once the ore reaches the pit rim and the
decision must be made to process it or send it to the waste dump.

The incremental cut off grade formula used for the reserves at $400 is presented
below:


Cut -Off Grade =     (Processing Costs (G&A incl.))

           (Gold Price - Selling Cost) * % Au Recovery

Cut -Off Grade =                (2.12)

                (12.86 - (12.86*0.198%)) * 79.46%


Cut -Off Grade = 0.21 grams per tonne.


The cut-off grade formula used for the resources at $450 was as follows:


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Cut-Off Grade =           (Processing Costs (G&A incl.))

                   (Gold Price - Selling Cost) * % Au Recovery

Cut-Off Grade =                        (2.12)

                        (14.47 - (12.86*0.198%)) * 79.53%


Cut-Off Grade = 0.18 grams per tonne.

17.10.3 PIT DESIGN

To design a practical open pit for Paracatu, the selected pit shell (17)
developed in Whittle 4X(C) waS imported into Datamine(C), commercial mining
software. The chosen pit shell is contoured on A bench-by-bench basis in the
model and the resulting contour lines are used to guide the pit design process.
The pit design was completed by K. Morris, P.Eng., Kinross' Manager of Open Pit
Mine Engineering. Mr. Morris has over 20 years of open pit optimization and
design experience.

The design criteria are summarized as follows in Table 17-13.



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               TABLE 17-13: PIT DESIGN CRITERIAD][GRAPHIC OMITTED]

              ---------------------------- -----------------------
                        PARAMETER                    VALUE
              ---------------------------- -----------------------
                       Bench Height                  12m.
              ---------------------------- -----------------------
                     Bench Face Angle                 75o
              ---------------------------- -----------------------
                     Inter-ramp Angle                 55o
              ---------------------------- -----------------------
                   Catchment Berm Width             10.4m.
              ---------------------------- -----------------------
                      Berm Interval                  24m.
              ---------------------------- -----------------------
                     Haul Road Width                 30m.
              ---------------------------- -----------------------
                    Haul Road Gradient                10%
              ---------------------------- -----------------------


Haul roads and in-pit ramps were designed at 10% gradient and 30m width, based
on approximately four times the width of a CAT 793 haul truck (~7.41m). This
will provide sufficient room for 2-way road traffic and also included an
allowance for a drainage ditch and safety berm. A typical road cross-section is
presented in Figure 17-8


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                      FIGURE 17-8 TYPICAL HAUL ROAD PROFILE




                                   [PICTURE]




Mineral reserves are estimated by reporting the model blocks within the design
pit above the incremental cut off grade described in section 17-10.2. Resource
model blocks classified as Measured are reported as Proven reserves, model
blocks classified as Indicated are reported and Probable reserves.

The Proven and Probable reserves within the design pit have been adjusted to
reflect mine position as of December 31, 2005. This was based on the end of year
mine surveyed topographic surface.

The Proven and Probable reserves are then scheduled and entered into a
Discounted Cash Flow (DCF) spreadsheet to estimate the project NPV and rate of
return. Operating and capital costs used in the DCF analysis originate from the
Plant Capacity Scoping Study. Review of the DCF for the Proven and Probable
reserves disclosed in this report


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indicates that the Paracatu Expansion Plan III is profitable at gold prices
above US $400 per ounce.

There will be differences in tonnes, grades and contained ounces between the
design pit when compared to the optimized pit shell. They can vary by as much as
10%, but this variance is inversely proportional to the size of the pit (i.e. a
small pit will typically have greater variances when compared to a larger pit).
Table 17-14 provides a comparison of the tonnes, grades and ounces contained in
the Whittle 4X(C) pit shell (No. 17) with that in the design pit.



                 TABLE 17-14 COMPARISON OF PIT DESIGN RESULTS TO WHITTLE 4X(C) OPTIMIZATION
                                     RESULTS FOR THE BASE CASE ESTIMATE

- ------------------------ --------------- ---------- -------------- --------------- ---------- -------------
                              ORE          GRADE       WASTE            TOTAL        STRIP      CONTAINED
                             TONNES         G/T        TONNES           TONNES       RATIO        AU OZ
- ------------------------ --------------- ---------- -------------- --------------- ---------- -------------
                                                                              
   PIT DESIGN RESULTS     1,186,204,614       0.400   406,714,069   1,592,918,684        0.5    15,201,487
- ------------------------ --------------- ---------- -------------- --------------- ---------- -------------
    WHITTLE RESULTS       1,174,967,566       0.400   414,255,105   1,589,222,671       0.35    15,172,149
- ------------------------ --------------- ---------- -------------- --------------- ---------- -------------
       DIFFERENCE            11,237,048           0    -7,541,036       3,696,013        N/A        29,338
- ------------------------ --------------- ---------- -------------- --------------- ---------- -------------
      % DIFFERENCE                0.96%       0.00%        -1.82%           0.23%        N/A         0.19%
- ------------------------ --------------- ---------- -------------- --------------- ---------- -------------







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18.0    OTHER RELEVANT DATA AND INFORMATION

This section is not applicable to the Paracatu mine.













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19.0    INTERPRETATION AND CONCLUSIONS

The Paracatu mine is a well-managed operating gold mine with a long history of
meeting production schedules and quotas.

Kinross has been very successful in identifying the extension of the
mineralization west of Rico Creek. The deposit exhibits excellent geological and
gold grade continuity.

The resource and reserve estimate described herein is well supported with a
detailed Plant Capacity Scoping Study that includes firm supplier quotations for
the proposed plant and mine equipment necessary to increase plant throughput
from 18 to 50 Mtpa in stages over a four-year period.

The recent data collected during Kinross' exploration program and added to the
database is free of gross error and omission and has been collected using
reasonable care and supervision to ensure the data meets industry best
practices. While QAQC checks are still in process, ongoing monitoring of
standards results indicates no significant bias in any of the labs used.

The revised geological model is based on a structural geological interpretation
of the Paracatu deposit. The changes in modeling method have not imparted a bias
in the estimate and are a better reflection of the geology observed.

The data density is sufficient to support the resource model classification.

All work supporting the resource and reserve estimate described herein has been
performed by or supervised by individuals who meet the definition of a Qualified
Person as described in Canada's National Instrument 43-101.

The reserves as estimated demonstrate positive financial returns for the project
on a discounted cash flow basis and therefore, meet the definition of a reserve
as defined by the CIM's Standards and Guidelines.


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20.0    RECOMMENDATIONS

Based on the pilot plant test results and the Plant Capacity Scoping Study, RPM
has recommended construction of Expansion Plan III. Kinross has reviewed the
data and conclusions presented by RPM and are in agreement with their
recommendation to proceed with the planned expansion. In Q3, 2005, Kinross'
Board of Directors approved funding for Basic Engineering and financial
commitments to allow SAG and Ball mill fabrication to proceed. In Q4, 2005, the
Basic Engineering for Expansion Plan III was awarded to SNC-Lavalin Engineers
and Constructors Ltd, an internationally recognized consulting engineering and
construction company and MinerConsult Engenharia, a Brazilian engineering firm.

It is recommended that:

        o       the resource model is re-estimated to incorporate the remaining
                analytical results,

        o       update the pit optimization, mine design and reserve estimate
                based on the updated resource model,

        o       update capital and operating costs based on results developed
                during Basic Engineering and Feasibility Study work






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21.0    ADDITIONAL INFORMATION FOR OPERATING PROPERTIES

21.1    PROCESS PLANT

Figure 21-1 is a simplified flow sheet of the current process plant. The
production statistics in the flow sheet are budget estimates only.

       FIGURE 21-1: SIMPLIFIED FLOW SHEET EXISTING PARACATU PROCESS PLANT





                                   [PICTURE]




21.1.1  CRUSHING

Typical run of mine ore is about 80% passing 70mm. The current plant features
four separate crushing lines, three of which are operated at one time while the
fourth is on standby or down for scheduled maintenance. Three crushers provide a
crushing rate of 800 tonnes per hour (tph). Each circuit consists of a primary
impact crusher followed by


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a secondary cone crusher. The final crushed product has a specification of 80%
passing 10 mm.

The crushed product feeds to a 5000 tonne fine ore bin. Two feeders from the
fine ore bin transfer the ore to one of two blending bins that feed into the
grinding circuit.

21.1.2  GRINDING CIRCUIT

The existing grinding circuit features four separate process streams consisting
of a single stage ball mill (1800 kW) which are fed at a rate of 600 tph from
the two blending bins. The ball size is 60 mm diameter with consumption of 300
g/t. The ball mills operate in closed circuit with 500mm hydro cyclones.

A fifth ball mill is used for regrinding a portion of the circulating load.

The final product specification from the grinding circuit is 80% passing 75
microns (200 mesh).

21.1.3  GRAVITY CIRCUIT

The current mill circuit includes sixteen jigs that are set up as part of each
grinding line. The jigs are fed with a portion of the circulating load from the
grinding circuit.

21.1.4  FLOTATION

The flotation circuit features three stages, flash, scavenger and cleaner.
Product from each grinding line is fed to four flash flotation units for gold
and sulphide recovery. All flotation reagents are added in the flash flotation
stage and include: Mercapto, MIBC and Dow froth. Approximately 60% of the
flotation gold recovery occurs in flash flotation + jig circuit. The remaining
20-25 % occurs in the scavenger circuit.

The scavenger units are 120 m(3) WEMCO cells.


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40 % of the flotation tails are thickened to 42 % solids in two 70m -diameter
thickeners, which joins with the other 60 % at 30% solids and are sent to the
main tailings pond.

21.1.5  HYDROMETALLURGY PLANT

Two separate concentrate products are sent to the hydrometallurgical plant, a
jig concentrate and a cleaner concentrate, which cleans the flash flotation and
scavenger flotation concentrates. All three average roughly 20-30 g/t gold.

Concentrates are first reground in two parallel mills to a size of 90% passing
325 mesh. The concentrates are then processed in a Knelson concentrator in line
with the regrind mills, recovering approximately 20% of the contained gold. The
Knelson concentrate is directed to a bank of shaking tables and then on to the
smelting furnace.

The reground sulphide concentrate is thickened to 45% solids in two 15m-diameter
thickeners prior to leaching. The thickened concentrate product is leached in
eight, 300-m3 CIP tanks. Oxygen is injected in to the first tank to reduce
cyanide consumption. Activated carbon is added to the leaching tanks
(configuring a CIL circuit) to collect gold from the solution. Loaded carbon is
produced from the first CIL tank.

The loaded carbon is stripped in two, 3 tonne Zadra process elution columns at
130(0) C using a caustic soda solution. Gold is precipitated onto steel wool by
electro winning. The carbon is reactivated in a 200-kg/hr kiln.

21.1.6  SMELTING

The process plant produces gold bullion using two induction furnaces. Typically,
the bullion averages 70-80% gold content with 20-30% silver and minor copper and
iron content.


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21.2    MARKETS AND CONTRACTS

This section is not applicable as gold production from Paracatu is sold on the
open market at spot gold prices. There are currently no gold loans or gold
derivative products that influence the gold price.

21.3    RECLAMATION AND MINE CLOSURE

RPM has a comprehensive and up to date closure plan including a closure cost
estimate. The plan is based on the "Rio Tinto Health, Safety & Environment -
Closure Planning Guidelines". The current estimate of closure costs is US $32
million (excludes any credits for salvage value).

Currently in Brazil there are no laws requiring the posting of a reclamation
bond. RPM is making an annual financial provision for closure costs, but this is
an accrual only, not an actual expense.

The planned closure of the main tailings pond proposes to mine oxide ore only
during the last year of production. This will provide a cover for the pond,
which will then be drained.

The closure plan involves placing a 1-meter thickness of cover materials on the
final pit floor, the top 0.8m being soil material.

21.4    TAXES

The following three types of taxation apply to RPM's mining operation at
Paracatu.

        1.      A tax on profit equal to the greater of:

                o       (a) based on actual profit 34% of actual profit (25%
                        federal and 9% social contribution)


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                o       (b) based on a presumed profit 3% of net sales (same
                        distribution to federal and social contribution)

        2.      CPMF (Tax on Financial Movement) Every movement of funds between
                banks is taxed at 0.38%. This is a federal tax.

        3.      Property taxes. RPM must pay property tax on its mining land and
                property in Paracatu. It is distributed to rural (county) and
                municipal (city) governments.

21.5    CAPITAL AND OPERATING COST ESTIMATES

The capital and operating costs estimates for Expansion Plan III have been
prepared by RPM and KTS staff and are documented in the Plant Capacity Scoping
Study. The Plant Capacity Scoping Study considered four options:

        o       Base Case - current plant configuration, no expansion,

        o       30 Mtpa - updated Feasibility Study expansion plan,

        o       50 Mtpa and

        o       66 Mtpa.

The Plant Capacity Scoping Study was based on an unclassified resource model
that included mineralization west of Rico Creek. It did not include
mineralization below the current mining areas.

Unclassified resources were estimated for each of the four expansion options
considered in the Plant Capacity Scoping Study. Operating costs were estimated
for each of the four expansion options and these costs were used to optimize the
unclassified resource model and develop a life of mine schedule for input into
Discounted Cash Flow (DCF) models to evaluate the Net Present Value (NPV) of
each option.


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The Plant Capacity Scoping Study concluded that the highest NPV option was an
expansion to a 50 Mtpa throughput rate. The expansion would be completed in
stages over a four year period from 2006 through to 2009. The initial stage
would maintain production at the current 18 Mtpa rate while a 32 Mtpa SAG Mill
line was constructed and commissioned. Once completed, the existing 18 Mtpa line
would be upgraded and mine production would continue, uninterrupted, at a rate
of 32 Mtpa. After completion of the refurbishment of the 18 Mtpa line, it would
be brought back on stream to process the remaining B1 reserves bringing total
plant throughput to 50 Mtpa.

The life-of-mine capital cost for the 50 Mtpa expansion is estimated to be US $
700 million of which US $ 326 million is estimated to be the capital cost for
the plant expansion with a 32 Mtpa capacity. Closure cost is estimated to be US
$ 47 million.

Most major plant equipment has been estimated based on firm supplier quotations.
Mine equipment capital costs have been estimated based on internal Kinross
equipment cost databases. Kinross considers the level of estimation accuracy to
be sufficient to support a reserve classification.

Construction costs for the plant site, steel fabrication, concrete and
earthworks have been estimated based on RPM's operating experience by ECM, the
Brazilian Engineering firm the completed the original Feasibility Study.

All cost estimates assume an exchange rate of 2.65 Reais per US dollar.

All costs were estimated in terms of dollars of the day as of June 1, 2005.

SNC-Lavalin Engineers and Constructors Ltd, an internationally recognized
consulting engineering and construction company and MinerConsult Engenharia, a
Brazilian engineering firm, have been awarded the Basic Engineering contract for
Expansion Plan III.


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The current economic climate, especially with regards to the pricing of oil and
steel, combined with the devaluation of the American dollar represent the key
areas of risk with respect to the capital and operating costs estimates.

Kinross has reviewed the cost estimates and has concluded that they meet
generally accepted industry standards for evaluating the economic viability of
the project.

21.5.1  OPERATING COST ESTIMATE

Operating costs are based on the 2006 Life-of-Mine Budget and are estimated to
be US $2.94 per tonne. Total cash costs per ounce are estimated to be US $260
per ounce and the life of mine production cost is estimated to be US $ 340 per
ounce. The project has a 28 year mine life and average annual gold production is
estimated to be 370,000 ounces per annum.

21.5.2  ECONOMIC ANALYSIS

Discounted cash flow analyses for a 50 Mtpa throughput rate and the Proven and
Probable reserves disclosed herein have been completed demonstrating that the
project is viable and has a positive rate of return at gold prices greater than
US $400 per ounce. The cash flows are based on life of mine plans estimated by
Kinross from the resource and reserve model described in this Technical Report.
The life of mine plans have been reviewed by Kinross and meet generally accepted
industry standards.

Kinross considers the financial models to be confidential and have not
incorporated said models into the body of this report. Said financial models may
be made available with the execution of a confidentiality agreement with
Kinross.


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22.0    REFERENCES

Davis B., Paracatu Estimation Reliability by Drill Spacing, April, 2005

Gy P., Bongarcon D. Francois, Agoratek International, Study of Sampling
Protocols and Ore Heterogeneity, May 2005;

Holcombe R., Holcombe, Coughlin and Associates, Structural Assessment of the RPM
Mine, Paracatu, Minas Gerais, May 2005;

J.C. Moller, M. Batelochi, Y. Akiti, M. Sharratt, and A.L. Borges: 2001, The
Geology and Characterization of Mineral Resources of Morro do Ouro, Paracatu,
MG;

Oleson J., Preliminary Report of Sample Prep and Laboratory Audit, April 2005

Rio Paracatu Mineracao S.A., 2004: RPM Expansion Plan III Feasibility Study;

Rio Paracatu Mineracao and Kinross Technical Services, June 2005: Plant Capacity
Scoping Study;


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