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                         Paracatu Mine Technical Report

Paracatu, Minas Gerais State, Brazil

Prepared by:

R. D. Henderson, P. Eng

Acting Vice President, Technical Services

Kinross Gold Corporation

July 31, 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..............................................    1-22
        1.18   RECOMMENDATIONS..........................................    1-23

 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

 3.0   RELIANCE ON OTHER EXPERTS........................................     3-4
         3.1   INDEPENDENT THIRD PARTY PARTICIPANTS.....................     3-4
         3.2   STUDY PARTICIPANTS.......................................     3-4
         3.3   DISCLAIMER...............................................     3-5

 4.0   PROPERTY DESCRIPTION AND LOCATION................................     4-1
         4.1   PROPERTY DESCRIPTION.....................................     4-1
         4.2   LOCATION.................................................     4-2
         4.3   TITLE AND OWNERSHIP......................................     4-2
         4.4   PERMITTING...............................................     4-9

            4.4.1   Brazilian Framework for the Environment.............     4-9
            4.4.2   Current Operations Status...........................    4-13

         4.5   ROYALTIES................................................    4-16

 5.0   ACCESS, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND
          PHYSIOGRAPHY..................................................     5-1

 6.0   PROJECT HISTORY..................................................     6-1

 7.0   GEOLOGICAL SETTING...............................................     7-1
         7.1   REGIONAL GEOLOGY.........................................     7-1
         7.2   LOCAL GEOLOGY............................................     7-3
         7.3   DEPOSIT GEOLOGY..........................................     7-7

 8.0      DEPOSIT TYPE..................................................     8-1

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 9.0   MINERALIZATION...................................................     9-1
         9.1   PETROGRAPHY..............................................     9-1
         9.2   SULPHIDES................................................     9-1
         9.3   GOLD.....................................................     9-3

10.0   EXPLORATION......................................................    10-1

11.0   DRILLING.........................................................    11-1
        11.1   DRILL SPACING............................................    11-5

12.0   SAMPLING METHOD AND APPROACH.....................................    12-1
        12.1   BULK DENSITY AND CORE SPECIFIC GRAVITY...................    12-2
        12.2   BOND WORK INDEX..........................................    12-3

13.0   SAMPLE PREPARATION, ANALYSES AND SECURITY........................    13-1
        13.1   SAMPLE PREPARATION AND ANALYSES..........................    13-1
        13.2   SECURITY.................................................    13-3

14.0   QUALITY CONTROL, QUALITY ASSURANCE...............................    14-5
        14.1   RESULTS..................................................    14-7
        14.2   RERUNS...................................................   14-10
        14.3   ROUND ROBIN TESTS - COARSE AND PULD REJECT ANALYSES......   14-14
        14.4   LAB BIAS.................................................   14-14

15.0   DATA VERIFICATION................................................    15-1

16.0   ADJACENT PROPERTIES..............................................    16-1

17.0   MINERAL PROCESSING AND METALLURGICAL TESTING.....................    17-1
        17.1   EXISTING PROCESS PLANT...................................    17-1
        17.2   EXPANSION PLAN...........................................    17-2

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        17.3   EXPANSION PLAN III METALLURGICAL TESTWORK................    17-4

            17.3.1   Crushing...........................................    17-4
            17.3.2   Grinding Work Index................................    17-4
            17.3.3   Mill Sizing........................................    17-6
            17.3.4   Mineralogy.........................................    17-7
            17.3.5   Cyanide Destruction................................   17-10
            17.3.6   Gold Recovery......................................   17-10

18.0   MINERAL RESOURCE AND RESERVE ESTIMATES...........................    18-1

        18.1   MINERAL RESERVE AND RESOURCE STATEMENT...................    18-2
        18.2   HISTORICAL ESTIMATES.....................................    18-4
        18.3   MODELING METHODOLOGY.....................................    18-6

            18.3.1   Overview...........................................    18-6
            18.3.2   Geological Interpretation..........................    18-6

        18.4   SAMPLE ANALYSIS..........................................   18-11

            18.4.1   Arsenic............................................   18-12
            18.4.2   Bond Work Index....................................   18-12
            18.4.3   Specific Gravity...................................   18-13

        18.5   COMPOSITING..............................................   18-13
        18.6   GRADE CAPPING AND RESTRICTING OF HIGH GRADE..............   18-14
        18.7   GEOSTATISTICS............................................   18-14
        18.8   BLOCK MODEL..............................................   18-15

            18.8.1   Grade Interpolation................................   18-16
            18.8.2   Specific Gravity...................................   18-17
            18.8.3   Ore Hardness.......................................   18-17
            18.8.4   Recovery...........................................   18-17
            18.8.5   Model Checking.....................................   18-18

        18.9   RESOURCE CLASSIFICATION..................................   18-19
       18.10   PIT OPTIMIZATION.........................................   18-21

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            18.10.1    Base Case........................................   18-21
            18.10.2    Cut-Off Grades...................................   18-25
            18.10.3    Pit Design.......................................   18-26

19.0   OTHER RELEVANT DATA AND INFORMATION..............................   19-29

20.0   INTERPRETATION AND CONCLUSIONS...................................    20-1

21.0   RECOMMENDATIONS..................................................    21-2

22.0   ADDITIONAL INFORMATION FOR OPERATING PROPERTIES..................    22-1
        22.1   PROJECT IMPLEMENTATION PLAN..............................    22-1
        22.2   MINING...................................................    22-4
        22.3   PROCESS PLANT............................................    22-8

            22.3.1   Existing Circuit...................................    22-8
            22.3.2   New Circuit........................................    22-9

        22.4   TAILINGS DISPOSAL AND RECLAIM WATER......................   22-15
        22.5   INFRASTRUCTURE...........................................   22-16
        22.6   MARKETS AND CONTRACTS....................................   22-17
        22.7   OCCUPATIONAL HEALTH, SAFETY AND ENVIRONMENTAL ASPECTS....   22-17

            22.7.1   Occupational Health and Safety Aspects.............   22-17
            22.7.2   Environmental Aspects..............................   22-18
            22.7.3   Closure Related Aspects............................   22-23

        22.8   TAXES....................................................   22-26
        22.9   CAPITAL AND OPERATING COST ESTIMATES.....................   22-27
       22.10   ECONOMIC ANALYSIS........................................   22-30

23.0   REFERENCES.......................................................    23-1

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                                 LIST OF TABLES

Table 1-1 Proven and Probable Mineral Reserves..........................     1-1
Table 1-2 Measured and Indicated Mineral Resources......................     1-2
Table 1-3 Bond Work Index Ore Hardness Estimates by Horizon.............    1-10
Table 4-1 Summary of RPM Mining Licenses and Exploration Concessions....     4-7
Table 6-1 Paracatu Life of Mine Production Summary......................     6-3
Table 6-2 Historical Mineral Resources and Reserve Estimates............     6-4
Table 11-1 Drill Holes Summary Table....................................    11-2
Table 11-2: Confidence Limits for Gold..................................    11-7
Table 11-3: Confidence Limits for Arsenic...............................    11-7
Table 13-1 Summary of Simple Preparation Procedures by Lab..............    13-3
Table 14-1: Standards and their Accepted Limits.........................    14-6
Table 14-2: Summary of QAQC by Laboratory...............................    14-7
Table 14-3: Laboratory Performance Summary for 2005 Exploration.........    14-8
Table 14-4 Selected Rerun Results.......................................   14-11
Table 14-5 Summary of Batch Reruns......................................   14-12
Table 15-1 Paracatu Production Reconciliation...........................    15-2
Table 16-1 Process Plant Metallurgical Recovery Summary.................    17-2
Table 18-1 Proven and Probable Mineral Reserves.........................    18-2
Table 18-2 Measured and Indicated Mineral Resources.....................    18-3
Table 18-3: Updated Drill Hole Database................................     18-6

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Table 18-4 Basic Statistics for Gold, Raw Sample Data...................   18-12
Table 18-5: Basic Statistics for Arsenic Assays.........................   18-12
Table 18-6: Basic Statistics for Bond Work Index........................   18-13
Table 18-7: Basic Statistics for Specific Gravity in Core Samples.......   18-13
Table 18-8: Paracatu Correlogram Summary................................   18-15
Table 18-9 Grade Interpolation Parameters...............................   18-17
Table 18-10 Comparison of LM vs NLM West of Rico Creek..................   18-19
Table 18-11 Grade Tonnage Summary of Imported and Exported Model........   18-22
Table 18-12: Base Case Optimization Parameters..........................   18-23
Table 18-13: Pit Design Criteria........................................   18-27
Table 22-1: Mine Fleet Requirements 2006-2009...........................    22-6
Table 22-2 Paracatu Life of Mine Schedule...............................    22-7
Table 22-3 Paracatu Taxation............................................   22-27
Table 22-4  Paracatu Production and Cost Summary........................   22-28
Table 22-5  CAPEX Breakdown for the 61 Mtpa Case @ R$ 2.3/US$...........   22-29
Table 22-6  Paracatu Expansion III Project Sensitivity to Gold Price....   22-30

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                                 LIST OF FIGURES

Figure 4-1 - Paracatu Mine Location Map.................................     4-2
Figure 4-2 Paracatu Mining and Exploration Claim Map....................     4-8
Figure 4-3 Brazilian Environmental Licensing and Control Process........    4-12
Figure 7-1 Regional Geology Paracatu District...........................     7-3
Figure 7-2 Typical sulphide mineralization in boudinage structures......     7-4
Figure 7-3 Small scale thrust faulting..................................     7-5
Figure 7-4: Local Geology of the Paracatu Deposit.......................     7-6
Figure 7-5 Conceptual Geological Cross Section of the Paracatu Deposit..     7-7
Figure 7-6 Conceptual Pre-Mining Weathering Profile.....................     7-8
Figure 9-1 Paracatu Thin Section Gold on Arsenopyrite Grain Boundary....     9-3
Figure 11-1 Drill Hole Location Map.....................................    11-3
Figure 11-2: Gold Estimation Uncertainty by Drill Hole Spacing..........    11-8
Figure 11-3: Arsenic Estimation Uncertainty by Drill Hole Spacing.......    11-8
Figure 14-1: Standard Performance - RPM Lab.............................    14-9
Figure 14-2: Standard Performance - ALS Chemex..........................   14-10
Figure 14-3: Standard Performance - Lakefield...........................   14-11
Figure 14-4  K-508 Samples 112 to 135 Initial vs Rerun by Aliquot.......   14-13
Figure 14-5 Plan View - Diamond Drilling Distribution by Analytical
   Lab..................................................................   14-15
Figure 17 -2 Typical Gold on Arsenopyrite Grain Boundaries..............    17-8
Figure 18-1 Tonnage Mined and in Reserve as of December 31, 2005........    18-5

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Figure 18-2 Ounces Mined and in Reserve as of December 31, 2005.........    18-5
Figure 18-3 Graded bedding in Unmineralized Phyllite....................    18-7
Figure 18-4 Phyllite with Verging Asymetric Folds, Shear Bands &
   Boudins..............................................................    18-8
Figure 18-5 Large Arsenopyrite Porphyroblast in Core....................    18-9
Figure 18-6 Drill Section 05N - Looking North...........................   18-11
Figure 18-7 Base Case Whittle 4X(C) Results.............................   18-24
Figure 18-8 Typical Haul Road Profile...................................   18-28
Figure 22-1 Paracatu Expansion III Implementation Schedule..............    22-3
Figure 22-2: Simplified Flow Sheet Existing Paracatu Process Plant......    22-8

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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 2006
Feasibility Study. This report has been prepared to comply with Canada's
National Instrument 43-101.

The scope of the Feasibility Study is to increase the present ore production
from approximately 18 Mtpa to approximately 61 Mtpa via the installation of a
new 41 Mtpa treatment plant, designed to treat the harder B2 sulphide ore being
encountered as the mine goes deeper. The existing plant will treat the softer
near-surface B1 ore at a throughput rate of 20 Mtpa until the soft ore is
depleted.

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 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

                      tonnes      Grade       Gold
Classification      (x 1,000)   (Au g/t)    (ounces)
-----------------   ---------   --------   ----------
Proven              1,106,420      0.40    14,277,000
Probable               79,864      0.38       979,000
                    ---------      ----    ----------
Proven & Probable   1,186,284      0.40    15,256,000
                    ---------      ----    ----------

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 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

                           tonnes       Grade      Gold
Classification            (x 1,000)   (Au g/t)   (ounces)
----------------------   ----------   --------   --------
Measured                 60,894,841     0.38      735,072
Indicated                 6,944,356     0.37       81,546
                         ----------     ----      -------
Measured and Indicated   67,839,197     0.37      816,617
                         ----------     ----      -------

 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 38.8 Mt averaging 0.37 g/t Au. Inferred
Resources were estimated at a gold price of US$450 per ounce and a FEX of 2.65
Reais per US$1.00.

An updated resource model was prepared in April 2006 and the Proven and Probable
Reserves within the design pit reflect the mine position as of December 31, 2005
based on the end of year mine surveyed topographic surface.

The resource and reserve estimates described in this report are classified
according to the Canadian Institute of 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,

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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.

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 principal
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-Calha and IDS
ore types. Metallurgical recovery was assigned to the Calha, non-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

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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
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 content. The BDZ averages 60 to 80 meters in thickness
with a gold grade of approximately 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 July 2006 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 current model is based

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on a 12.0 meter bench height which required 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
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. A 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 is provided in
Sections 17.0 and 18.0 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.

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The current 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.

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 21.0 of
this report. Expansion Plan III will increase plant throughput up to a maximum
of

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61 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
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 requires 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.

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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 DEG. to 35 DEG. Celsius.

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

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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).

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 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 Project III. The engineering drawings and cost
estimates were completed in July 2006 and form the basis of the 2006 Feasibility
Study.

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 DEG.) 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

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was syn-deformational. Gold and sulphides are scavenged from the Morro do Ouro
sedimentary sequence during deformation and localized within the high 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 on the core samples,
generally increases with depth. Table 1-3 presents the average range of BWI
measurements by horizon.

           Table 1-3 Bond Work Index Ore Hardness Estimates by Horizon

          BWI Range
Horizon    (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.

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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
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 at Paracatu is diamond drilling. A
database of 1,427 drill holes and test pits totalling 79,961 meters supports the
mineral reserve estimate for the 2006 Feasibility Study.

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
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 in the database collected prior to
1999. 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

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similarities with the Paracatu deposit but it is not in production. On a
regional scale there are additional anomalies being investigated by Kinross.

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%.

The RPM plant was initially designed for ore with a work Index of 3.0 and has
since 2003 treated ore in excess of 6.1kWh/t. Without substantial investment in
increased crushing and grinding capacity the throughput level at a work index of
6.9kWh/t (2006 estimate) is estimated at around 17Mtpa. As work index increases
with depth, the projection is that existing mill throughput will decline
steadily to 11Mtpa in 2026, gold output being then only 120.114 ounces/year.

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. 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
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 experts
in the field of SAG mill design and operation.

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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 and in January 2005 initiated an
aggressive 48,000m drill campaign at Paracatu significantly increasing project
reserves from 8.5M oz to 15.2M oz of gold.

A Plant Capacity Scoping Study was completed in July 2005, based on the
successful drill campaign and the SAG pilot plant data and RPM recommended
construction of Expansion Plan III.

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

The scope of the Feasibility Study is to increase the present ore production
from approximately 18Mtpa to approximately 61Mtpa by the installation of a new
41Mtpa treatment plant, designed to treat the harder B2 sulphide ore being
encountered as the mine goes deeper. The existing plant will treat the softer
near-surface B1 ore at a throughput rate of 20Mtpa until these reserves are
depleted.

1.16 Mineral Resource and Reserve Estimate

The resource model for the Feasibility Study relies on the same procedures and
methodology described in the Paracatu Mine Technical Report issued on March 30,
2006. The Technical Report was issued supporting the December 31, 2005 mineral
resource and reserve estimate.

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

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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 resource model is based on 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 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 DEG. 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 demonstrates 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 demonstrates 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.

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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).

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 meter
intervals. The composite data is extracted using the wire frames produced from
sectional interpretation. Each composite is coded according to the geological
wire frame. Any duplicate (twinned) composites are also discarded. Grade capping
for original 1.0m 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 10 composites are used within the search ellipsoid. A
maximum of four adjacent samples are used from the same drill hole.

The resource model estimates specific gravity for each model block (50 x 50 x 12
meters). Specific gravity measurements for core samples are collected and

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assessed based on 4.0 m composite samples comprised of 8.0 cm core intervals
selected for every 2.0 meters of core. Interpolation extracts data for each
geological zone and used the composite data to estimate the grade for each block
within the zone using a nearest neighbour methodology.

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 neighbor interpolation method.

Finally, each model block is assigned a metallurgical recovery based on sulphur
(S) and arsenic (As) content of the block. 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.

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.

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

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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.

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 2004 Feasibility Study was completed on an

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option to increase throughput with the addition of a SAG mill and in pit
crushing and conveying system.

Basic engineering that started in 2005 has culminated in a control cost estimate
with an accuracy level of +/-15%. As part of the Engineering study, SNC-Lavalin
confirmed that the 38' SAG mill would be adequate to attain the 61Mtpa
production level and the study has quantified the capital and operating costs to
support this Expansion III project.

The Feasibility Study has shown that the project is economically viable at a
gold price of $400/oz.

1.18 Recommendations

On the basis of the 2006 Feasibility Study, RPM have requested full release of
funds for continuing the implementation of the Expansion III project. Kinross
has reviewed the data and conclusions presented by RPM and are in agreement with
their recommendation to proceed with the planned expansion.

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 data from the 2005 drill campaign 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:

R. D. Henderson, P. Eng, Acting Vice-President Technical Services, Kinross Gold
Corporation.

Co-authors of this report include:

M. Belanger, P Geo, Director Technical Services, Kinross Americas, and

K. Morris, P. Eng, Manager Mining, Kinross Gold Corporation

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

The resource and reserve estimates are based on an updated resource model
prepared in July 2006, and assumes that the existing operation will be expanded
to increase plant throughput to 61 Mtpa.

2.2  Terms of Reference

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

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

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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
SAG       semi-autogenous grinding
SGA       Environmental Management System
SISNAMA   National Environmental System
t         tonne(s)
Tpa       tonnes per annum

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Tpd       tonnes per day
Tph       tonnes per hour

2.4  Scope and Objectives

This report is prepared in support of the 2006 Rio Paracatu Mineracao
Feasibility Study for the Paracatu Expansion III Project.

2.5  Report Basis

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.

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3.0  RELIANCE ON OTHER EXPERTS

3.1  Independent Third Party Participants

The following independent consultants have contributed indirectly to this
report:

Agoratek International                 Sampling Heterogeneity Study
Dr. B. Davis, Independent Consultant   Optimum Drill Hole Spacing
Holcombe, Couglin & Associates         Structural Geology Assessment
Minerconsult / SNC Lavalin             Process Plant and Infrastructure
Golder Associates                      Pit Design, Waste Rock and Tailings

3.2  Study Participants

The following employees of Kinross have contributed to the report:

C. Frizzo, Kinross Americas Exploration         Geology and QA/QC
B. Gillies, P.Geo, Kinross Gold Corporation     Geology and QA/QC
M. Belanger, P.Geo, Kinross Americas            Resource Estimation
W. Hanson, P. Geo, Kinross Gold Corporation     Resource Estimation
Dr. R. Peroni, Rio Paracatu Mineracao           Resource Estimation
K. Morris, P.Eng, Kinross Gold Corporation      Reserve Estimation
J. Oleson, Kinross, Fort Knox Operations        Laboratory Audit
R. Henderson, P.Eng, Kinross Gold Corporation   Metallurgy and Process
W. Phillips, Kinross Americas                   Metallurgy and Process
L. A. Tondo, Rio Paracatu Mineracao             Metallurgy and Process

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3.3  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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4.0  PROPERTY DESCRIPTION AND LOCATION

4.1  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 61 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 will be purchased and a new mill will be installed to
supplement the existing mill.

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4.2  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 DEG. 3'S and
longitude 46 DEG. 35'W. Figure 4-1 is a location map showing the location of
Paracatu (in red).

                     Figure 4-1 - Paracatu Mine Location Map

                                      [MAP]

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

4.3  Title and Ownership

In Brazil, the Departamento Nacional da Producao Mineral (DNPM) issues all
mining leases and exploration concessions. Mining leases are renewable

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

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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.

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.

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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 4-2. 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.

     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

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          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 4-1 summarizes RPM's current mining licenses and exploration concessions.

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      Table 4-1 Summary of RPM Mining Licenses and Exploration Concessions

                                                  Mining Lease
   DNPM                                 Date       Application       Area
    (#)               Type            Acquired        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 4-2 Paracatu Mining and Exploration Claim Map

                                      [MAP]

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4.4  Permitting

4.4.1 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 4-3 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 4-3 Brazilian Environmental Licensing and Control Process

                             [GRAPHIC APPEARS HERE]

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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.

4.4.2 Current Operations Status

One of the initial conditions satisfied by RPM in obtaining a mining licence 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.
RPM is the first Brazilian gold mining company to receive ISO 14001
certification. The mine practices very good environmental care and monitoring
programs.

RPM is currently licensed to draw a set amount of water from the Sao Domingos,
Santa Rita and Sao Pedro rivers. As 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 have 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.

Public perception of this process has been very positive as evidenced by the
fact that 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.

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A potentially significant permitting issue was the request to mine west and
below the Rico Creek and this permit has been approved in 2006.

As outlined in the section above there are a number of licences to be granted by
governmental bodies associated with the implementation of RPM Expansion Project
III. The current status of these licenses is as follows:

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                             Estimated
   Type of                  completion
   Licence        Status        by                       Comments
-------------   ---------   ----------   ---------------------------------------
Previous         granted     July 2007   Previous licence is a pre-requisite of
Licence (LP)                             the Installation License (LI) required
Expansion                                for lowering the water table and Rico
Project III                              Creek deviation. This licence 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. A draft version of the
                                         EIA-RIMA has been concluded and the
                                         final document should be forwarded to
                                         the agency in August 2006.

Installation     ongoing     February    Installation License (LI) authorises
Licence (LI)                   2008      the start of the mining project,
Expansion                                including implementation and
Project                                  installation of the project.
                                         Installation License (LI) estimated to
                                         February 2008.

Deforestation    ongoing     February    All these permits should be issued
Permits                        2008      accordingly to plan agreed with IEF -
                                         Forestry Department before the
                                         installation works begin.

Rico Creek      completed    July 2005   This permit was granted on July 2005
Deviation                                and published in the official
Permit                                   government newspaper.

Water Table     completed    May 2006    Permit granted in May 2006 during
Lowering                                 meeting of the Water Committee - IGAM
Permit                                   in Belo Horizonte.

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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.

4.5  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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5.0  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 DEG. Celsius,
ranging from a high of 35 DEG. C to a low of 15 DEG.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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6.0  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 17.0 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.

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 6-1 summarizes the historic life of mine production at Paracatu since it
began commercial production in 1987.

               Table 6-1 Paracatu Life of Mine Production Summary

<TABLE>
<CAPTION>
          Year             1987     1988      1989      1990      1991      1992      1993      1994      1995      1996
-----------------------   -----   -------   -------   -------   -------   -------   -------   -------   -------   -------
<S>                       <C>     <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>
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
</TABLE>

<TABLE>
<CAPTION>
          Year              1997      1998      1999      2000      2001      2002      2003      2004      2005      TOTAL
-----------------------   -------   -------   -------   -------   -------   -------   -------   -------   -------   ---------
<S>                       <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>
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>

Table 6-2 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 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

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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 6-2 Historical Mineral Resources and Reserve Estimates

<TABLE>
<CAPTION>
             Kinross      Gold
            Ownership     Price    Reporting                              Tonnes      Grade       Gold
   Date        (%)      (US$/oz)      Code         Classification       (x 1,000)   (Au g/t)    (Au ozs)
---------   ---------   --------   ---------   ----------------------   ---------   --------   ----------
<S>              <C>        <C>       <C>      <C>                      <C>             <C>    <C>
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
31-Dec-05        100%       $400      CIM      Proven                   1,103,677       0.40   14,194,000
                            $400               Probable                    83,131       0.38    1,016,000
                            $400               Proven & Probable        1,186,808       0.40   15,210,000
                            $450               Measured                    89,784       0.27      771,000
                            $450               Indicated                    5,540       0.38       68,000
                            $450               Measured and Indicated      95,324       0.27      839,000
                            $450               Inferred                    40,100       0.37
</TABLE>

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7.0  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.

7.1  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 7-1 is a regional geological map of the Paracatu district modified as per
Holcombe 2005.

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                  Figure 7-1 Regional Geology Paracatu District

                                      [MAP]

7.2  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 7-2 and 7-3
below.

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       Figure 7-2 Typical sulphide mineralization in boudinage structures

                                    [GRAPHIC]

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                     Figure 7-3 Small scale thrust faulting

                                    [GRAPHIC]

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 DEG. 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 7-4. 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 7-4: Local Geology of the Paracatu Deposit

                                      [MAP]

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Figure 7-5 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 7-5 Conceptual Geological Cross Section of the Paracatu Deposit

                                    [GRAPHIC]

7.3  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 7-6 presents
the conceptual pre-mining weathering surface and established the relative
relationship between the various zones. Mining to date has exhausted the C and

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

               Figure 7-6 Conceptual Pre-Mining Weathering Profile

                                    [GRAPHIC]

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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8.0  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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9.0  MINERALIZATION

9.1  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.

9.2  Sulphides

The amount of sulphides present typically does not 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 (<1mm) to
coarsegrained (>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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9.3  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 9-1.

      Figure 9-1 Paracatu Thin Section Gold on Arsenopyrite Grain Boundary

                                    [GRAPHIC]

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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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10.0 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 two phases with the subsequent phase 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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11.0 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 drilling and sampling at Paracatu includes 479 test pits (5,056 meters) and
948 drill holes (74,905 meters). Table 11-1 summarizes the drill database as of
July, 2006.

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                      Table 11-1 Drill Holes Summary Table

                        Hole type    Number of    Total
   Year     Campaign    (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 228 of the 267 drill holes completed in 2005.

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

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

The portion of the deposit demonstrated to be economically viable comprises an
area approximately 3.0 km long by 2.0 km wide.

Figure 11-1 is a plan map of the drill holes included in the resource model
documented in this report.

                       Figure 11-1 Drill Hole Location Map

                                      [MAP]

Included in the hole totals are 67 reverse circulation (RC) 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

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bias is thought 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 the 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 2Deg. 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.

11.1 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 11-2
and 11-3.

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                     Table 11-2: Confidence Limits for Gold

Drill grid
    (m)      17 Mtpa   30 Mtpa
----------   -------   -------
 100 x 100      6.5%      4.9%
 200 x 200      8.2%      7.5%
 300 x 300     14.0%     13.0%

                    Table 11-3: Confidence Limits for Arsenic

Drill grid
    (m)      17 Mtpa   30 Mtpa
----------   -------   -------
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 11-2 and 11-3.

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         Figure 11-2: Gold Estimation Uncertainty by Drill Hole Spacing

                                     [CHART]

        Figure 11-3: Arsenic Estimation Uncertainty by Drill Hole Spacing

                                     [CHART]

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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.

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12.0 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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12.1 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.

For the core samples in B1 the recent specific gravity measurements were
factored down 5% based on the average moisture content measured by the process
plant in the last 6 months. In B2, the dry and wet density measurements on the
core samples showed no significant differences. It was therefore decided not to
modify the B2 values.

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12.2 Bond Work Index

Samples for Bond Work Index (BWI) testing are collected during sample
preparation of the 1.0 meter raw samples. Composite samples were originally
based based on an 8.0 meter down hole length representing the current mining
bench height. The current model is based on a 12.0 meter bench height which
required re-compositing of the historical 8.0 meter data to reflect the change
in bench height.

Each composite is composed of a fraction of each meter after initial sample
crushing to 2.0mm. 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. In April 2006, a calculation
error was detected in the BWI laboratory spreadsheet for the 2005 drill data and
was corrected.

The BWI composite data is used to interpolate the BWI for individual blocks in
the model using a nearest neighbour interpolation method.

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13.0 SAMPLE PREPARATION, ANALYSES AND SECURITY

13.1 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 13-1 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 13-1 Summary of Simple Preparation Procedures by Lab

<TABLE>

<S>                             <C>                               <C>
          Lakefield                        ALS Chemex                         RPM
          ---------                        ----------                         ---
   Belo Horizonte, Brazil               Luziania, Brazil                Paracatu, Brazil
                                       Vancouver, Canada

           Drying                           Drying                        Crushing (1)
           ------                           ------                        ------------
Total sample                    Total sample                      Total sample 100% < 1cm
100 DEG. - 110 DEG. C           100 DEG. - 110 DEG. C             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 DEG. - 120 DEG. 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
</TABLE>

13.2 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 pickup
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 kg are stored as a coarse rejects and 2 kg 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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14.0 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 Head of Mining
Department.

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 and 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 14-1

                 Table 14-1: Standards and their Accepted Limits

                                     Certified
           Certified    Certified     Standard      Certified        Accepted
Standard     Value     Variability   Deviation    QA/QC Limits     QA/QC limits
 (Ref #)    (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 color. In order to make the blanks
less obvious to lab employees, samples of barren hangingwall phyllite with
similar characteristics as regular samples were used in the latter part of the
drilling program

A model numbering code system was generated that could accommodate the 3
different batch sizes of the 3 labs. Table 14-2 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 14-2: Summary of QAQC by Laboratory

<TABLE>
<CAPTION>
                                Internal Lab QA/QC            Client QA/QC
                         -------------------------------   ------------------
            Batch Size   Standards   Blanks   Duplicates   Standards   Blanks   Samples / batch
   Lab         (#)          (#)        (#)        (#)         (#)        (#)          (#)
---------   ----------   ---------   ------   ----------   ---------   ------   ---------------
<S>                 <C>          <C>      <C>          <C>         <C>      <C>              <C>
Chemex              84           2        1            3           2        3                73
Lakefield           50           1        1            2           1        2                43
RPM                 30           1        1            0           1        1                26
</TABLE>

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 not inserted in a manner that assured
that the analytical lab would be unable to identify the standards from the
submitted samples. But, as five different standards were used, it is reasonable
to assume that the standards satisfy the requirement that they be blind.

14.1 Results

Results available are from March 1, 2005 to January 11, 2006 and include data
for 228 exploration holes analyzed by RPM, Lakefield and ALSChemex.

Results received to date for the certified standards indicate that both
ALSChemex and RPM have returned results mostly within the +1 / -1 standard
deviation limits and Lakefield has returned results within + 0.5 / -1.5 standard
deviation, showing a consistent bias of - 0.5 standard deviation. 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
tolerances with failure rates of 0.9% to 2.7% for the analyses performed. A
failure on a standard is classified as +/- 2.5 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

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pending and corrections, (if necessary) will be made to the database on receipt
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 14-3

         Table 14-3: Laboratory Performance Summary for 2005 Exploration

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

Figures 14-1 to 14-3 summarize QA/QC standard results by lab.

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                   Figure 14-1: Standard Performance - RPM Lab

                                     [CHART]

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                 Figure 14-2: Standard Performance - ALS Chemex

                                     [CHART]

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                  Figure 14-3: Standard Performance - Lakefield

                                     [CHART]

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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

14.2 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 14-4.

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                        Table 14-4 Selected Rerun Results

<TABLE>
<CAPTION>
                                            Initial Analysis                                             Rerun Analyses
               --------------------------------------------------------------------------  -----------------------------------------
 Hole  Sample  Aliquot 1  Aliquot 2  Aliquot 3  Aliquot 4  Aliquot 5  Aliquot 6     Avg    Aliquot 1  Aliquot 2  Aliquot 3  Result
 (#)     (#)    (Au g/t)   (Au g/t)   (Au g/t)   (Au g/t)   (Au g/t)   (Au g/t)  (Au g/t)   (Au g/t)   (Au g/t)   (Au g/t)  (Au g/t)
-----  ------  ---------  ---------  ---------  ---------  ---------  ---------  --------  ---------  ---------  ---------  --------
<S>      <C>      <C>        <C>        <C>        <C>        <C>        <C>       <C>        <C>        <C>        <C>      <C>
K-508    170      1.16       0.48       5.80       1.24       0.67       0.86      1.70       0.90       0.88       1.82     1.19
K-512    235      0.66       0.79       1.43                                       0.97       1.04       1.26       1.32     1.20
K-601    112      0.56       0.71       0.27       0.95       0.48       0.55      0.58       0.59       1.38       0.89     0.94
K-601    175      0.35       0.59       0.89       6.31       0.61       0.54      1.54       0.81       0.38       0.37     0.52
K-1-5    128      0.65       1.27       0.19                                       0.71       0.23       0.28       0.13     0.21
K-510    179      1.46       0.86       1.12                                       1.15       1.42       1.26       0.69     1.12
K-207     26      0.94       0.53       0.58       0.57       0.43       0.76      0.64       0.11       0.09       0.07     0.09
K-207     28      0.63       0.34       1.13       0.62       1.10       0.39      0.70       0.17       0.06       0.09     0.11
K-613    171      0.13       0.16       0.09                                       0.12       0.11       0.83       0.19     0.42
K-908    222      1.08       0.93       1.46                                       1.15       3.60       1.05       0.83     1.83
K-908    226      1.11       0.89       1.88                                       1.29       0.93       1.28       0.61     0.93
</TABLE>

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 14-5 summarizes the rerun results for the 16 batches submitted for rerun
analysis.

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                       Table 14-5 Summary of Batch Reruns

                     Batch     Initial     Rerun
             Hole    Sample    Result     Result
   Lab       (#)      (#)     (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 14-4 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 14-4 K-508 Samples 112 to 135 Initial vs Rerun by Aliquot

                                     [CHART]

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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.

14.3 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.

14.4 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 14-5 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.

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     Figure 14-5 Plan View - Diamond Drilling Distribution by Analytical Lab

                                     [MAP]

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15.0 DATA VERIFICATION

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

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Table 15-1 summarizes the production reconciliation for the period 1990 to 2005.

                  Table 15-1 Paracatu Production Reconciliation

<TABLE>
<CAPTION>
           Year              1990    1991    1992    1993    1994    1995    1996    1997
-------------------------   -----   -----   -----   -----   -----   -----   -----   -----
<S>                         <C>     <C>     <C>     <C>     <C>     <C>     <C>     <C>
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
</TABLE>

<TABLE>
<CAPTION>
           Year              1998    1999    2000    2001    2002    2003    2004    2005
-------------------------   -----   -----   -----   -----   -----   -----   -----   -----
<S>                         <C>     <C>     <C>     <C>     <C>     <C>     <C>     <C>
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
</TABLE>

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16.0 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 is not significant in the context of
this Technical Report.

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17.0 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.

17.1 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

<TABLE>
<CAPTION>
Year                                1987   1988   1989   1990   1991   1992   1993   1994   1995    1996
---------------------------------   ----   ----   ----   ----   ----   ----   ----   ----   ----   -----
<S>                                 <C>    <C>    <C>    <C>    <C>    <C>    <C>    <C>    <C>     <C>
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
                                    ----   ----   ----   ----   ----   ----   ----   ----   ----    ----
</TABLE>

<TABLE>
<CAPTION>
Year                                1997   1998   1999   2000   2001   2002   2003   2004   2005   TOTAL
---------------------------------   ----   ----   ----   ----   ----   ----   ----   ----   ----   -----
<S>                                 <C>    <C>    <C>    <C>    <C>    <C>    <C>    <C>    <C>     <C>
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
                                    ----   ----   ----   ----   ----   ----   ----   ----   ----    ----
</TABLE>

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

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The Plant Capacity Scope Study recommended that production be increased from 18
Mtpa to 50 Mtpa via the installation of a new circuit with a capacity of 32
Mtpa.

In Q4, 2005, Basic Engineering commenced including the in pit crusher, covered
stockpile, new 32 Mtpa mill, hydromet expansion, electrical substation, tailings
delivery and water systems. As part of the Engineering Study SNC/Minerconsult
investigated the possibility of upgrading the 50 Mtpa design to 61 Mtpa. The
study demonstrated that with the upgrade of some equipment, the bottlenecks
restricting production to 50 Mtpa could be removed. The additional capital
requirements for the upgrade formed part of this study. The study also confirmed
that the 38' SAG mill would be adequate to attain the 61 Mtpa production level.

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 100 tph (equivalent to 50 Mtpa mill feed) and additional
equipment will be required for the regrinding, CIL, elution, carbon regeneration
and electro winning circuits.

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.

In the Expansion Plan, the existing plant will be operated on soft B-1 ore at a
treatment rate of up to 20 Mtpa. In mid 2008, the new plant consisting of 1
crusher (1,300 mm toothed roll type), one 20 MW SAG mill (38' dia. x 22' long
EGL), one 13 MW ball mill (24' dia. X 39.5' long EGL), gravity plant, flotation
plant and the hydrometallurgical plant

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expansion will be commissioned. This plant will be operated at approximately 20
Mt/a to process a blend of soft B-1 and harder B-2 ore.

In late 2008, two additional 13 MW ball mills will be added to the new plant and
commissioned. The plant will then treat 41 Mtpa of harder B-2 ore. The existing
plant will continue to operated at up to 20 Mtpa of soft B-1 ore for a total of
61 Mtpa. A new tailings dam and water reclaim system are required for operation
at 61 Mtpa.

17.3 Expansion Plan III Metallurgical Testwork

17.3.1 Crushing

Crushing test work on the hardest RPM ore types (WI=>12.0) has been conducted by
MMD using an MMD type sizer and the equipment has proved adequate for treating
the harder B2 ore in the reserve inventory. A trade off study comparing MMD type
sizers with gyratory crushers was conducted by SNC-Lavalin/Minerconsult and
concluded that the MMD sizer was the most suitable equipment for the task. On
this basis, an MMD type sizer has been selected as the most suitable equipment
for the IPCC.

17.3.2 Grinding Work Index

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

     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 single stage 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 41 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.

The historical work index measured in the existing Paracatu grinding circuit has
been significantly lower than that predicted by the laboratory Bond Ball WI. A
correction factor called the RPM factor has traditionally been applied to the
Bond Ball WI to estimate ball mill power requirements. The accepted factor at
Paracatu has been 1.6, that is the laboratory Bond Ball WI is divided by 1.6 to
estimate the ball mill power requirements. The pilot plant test work
investigated this factor, and independent consultants, (SMCC

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and OMC) confirmed that the measured WI was lower than the laboratory Bond Ball
WI and that the RPM factor of 1.6 was appropriate for the design of the new
grinding circuit. The factor is attributed to the large amount of fines that are
in the feed to the SAG mill. There are some indications that the factor could
increase with increasing Bond Ball WI, however this was not considered in the
sizing of the mills and may represent an opportunity for increased throughput
with the harder ores.

17.3.3 Mill Sizing

SAG mill power and sizing is based on the pilot plant test work which consisted
of 65 runs, including 58 closed circuit tests and 7 open circuit tests. The test
results have been reviewed by others and their reports are included in the
Feasibility Study.

     o    Mill Sizing for the RPM Expansion Options A and B - SMCC Pty Ltd,
          Queensland Australia (SMCC).

     o    Mill Sizing for RPM Grinding Circuit Expansion Project Progress Report
          November 2005 - Homero Delboni and Associates Services, Queensland,
          Australia (HAD).

     o    Mill Sizing for RPM Grinding Circuit Expansion Project Progress Report
          July 2005 - HAD.

     o    Kinross Gold Mines Rio Paracatu Mineracao (RPM) Review of Comminution
          Circuit - Orway Mineral Consultants, Perth WA (OMC).

The initial study was completed by ECM S.A. Projetos Industriais (ECM) for the
RPM Expansion Project was for a single stage SAG mill producing a P80 of 212
(mu)m. This product was then transferred to the existing ball mills to produce a
final product P80 of 75 (mu)m. The reviews by SMCC and HDA expressed concerns
about the potential of

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slurry pooling in the SAG mill due to the anticipated high volumetric flows
through the single stage SAG mill.

The grinding circuit for the 2006 Feasibility Study was modified to consist of
one 20 MW SAG mill (38' dia. x 22' long EGL) and two 13 MW ball mills (24' dia.
X 39.5' long EGL) processing up to 5,100 t/h at a P80 of 75 (mu)m. As the early
capacity of the grinding circuit will be restricted by the availability of power
to the project, the grinding circuit will initially operate with a single SAG
mill followed by a single ball mill. The second ball mill will be started when
full power supply is on line. The ore hardness is variable, increasing with time
as the mine progresses deeper into the pit. As the installed grinding power is
fixed, the tonnage processed will decrease with increasing ore hardness. There
may be potential to install a third ball mill to sustain mill throughput at 41
Mtpa and space has been reserved for a third ball mill.

The project criteria established by RPM for sizing and selecting the grinding
mills, was to use sizes of mills and drives that were operating successfully
elsewhere. As the drives for 40' SAG mills have not been without problems, the
SAG mill maximum size was limited to 38' where there is a large population of
successful installations. There are six similar mill installed worldwide
including the Brazilian mill at Sossego (CVRD). The maximum tonnage for a single
38-foot SAG mill is anticipated to be limited to 42Mtpa due to volumetric
constraints in the mill chamber. Larger capacity 40-foot SAG mills may have
higher tonnage, however the technical risk will increase, as there are currently
only two 40-foot SAG mills in operation.

17.3.4 Mineralogy

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 17-2 illustrates a
typical occurrence.

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           Figure 17 -2 Typical Gold on Arsenopyrite Grain Boundaries

                                    [GRAPHIC]

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.

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17.3.5 Cyanide Destruction

Studies have been undertaken by RPM to evaluate the substitution of the existing
AVR cyanide recovery process with the more modern and widely used SO2 /Air
Cyanide Destruction Process. Laboratory testing of the SO2 /Air Cyanide
Destruction Process on RPM cyanidation tailings has demonstrated that the
process has a series of advantages over the AVR process currently being used. A
new SO2 /Air cyanide destruction plant will be installed as part of the
Expansion III Project.

17.3.6 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
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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18.0 MINERAL RESOURCE AND RESERVE ESTIMATES

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

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 of the resource and reserve estimates.

The mineral resource model for Paracatu was interpreted and estimated using
Vulcan software. 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 resource model of December 31, 2005 was updated in April 2006 for the
Feasibility Study. The update was necessary to correct the ore hardness (BWI)
model as a calculation error was identified originating from the reporting
laboratory. As a result of the error, Kinross reviewed the entire BWI results
used in resource estimation, verified that the correct calculation method was
employed, entered the correct BWI values into the resource model blocks,
re-optimized the revised model and re-designed the ultimate pit. Kinross notes
that no material difference resulted from the BWI error relative to the December
31, 2005 resource and reserve estimate.

With the exception of the changes noted above, the procedures and methodology
described in the Paracatu Mine Technical Report, March 30, 2006 were used to
complete the resource model for the Feasibility Study.

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The resource model 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. Increased boudin
frequency and arsenopyrite content result in higher gold grades.

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

Mineral resources were estimated within optimized pit shells based on Whittle
4X[_] software, a program that has become a standard in the mining industry.
Mineral reserves were estimated within design pits developed from the optimized
pit shells.

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.

18.1 Mineral Reserve and Resource Statement

The Proven and Probable mineral reserve estimate as of December 31, 2005, for
the Feasibility Study was estimated from the April 2006 resource model update
and is summarized in Table 18-1. Proven and Probable mineral reserves were
estimated at a gold price of US$400 per ounce and 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 18-1 Proven and Probable Mineral Reserves

                     tonnes       Grade       Gold
  Classification    (x 1,000)   (Au g/t)    (ounces)
-----------------   ---------   --------   ----------
Proven              1,106,420       0.40   14,277,000
Probable               79,864       0.38      979,000
                    ---------   --------   ----------
Proven & Probable   1,186,284       0.40   15,256,000
                    =========   ========   ==========

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Table 18-2 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 of 2.65 Reais per US
$1.00 and a cut off grade of 0.18 g/t Au.

               Table 18-2 Measured and Indicated Mineral Resources

                            tonnes      Grade     Gold
    Classification        (x 1,000)   (Au g/t)  (ounces)
----------------------   ----------   -------   -------
Measured                 60,894,841      0.38   735,072
Indicated                 6,944,356      0.37    81,546
                         ----------   -------   -------
Measured and Indicated   67,839,197      0.37   816,617
                         ==========   =======   =======

 NB Measured and Indicated resources are reported exclusive of mineral reserves

In addition to the Measured and Indicated Mineral Resources stated in Table
18-2, Paracatu hosts an Inferred Resource of 38.8 Mt averaging 0.37 g/t Au.
Inferred Resources were 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 were classified according to the
Canadian Institute on Mining, Metallurgy and Petroleum (CIM) Standards on
Mineral Resources and Reserves.

The mineral resources 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.

18.2 Historical Estimates

The reserve history at Paracatu indicates continuous growth of the reserve base
reflecting increased geological knowledge and improved process efficiencies.
Figures 18-1 and 18-2 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 18-1 Tonnage Mined and in Reserve as of December 31, 2005

                                     [CHART]

         Figure 18-2 Ounces Mined and in Reserve as of December 31, 2005

                                     [CHART]

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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.

18.3 Modeling Methodology

18.3.1 Overview

The December 31, 2005 resource model was updated by KTS in April 2006 to correct
an error in the ore hardness (BWI) data. No other data or edits were made to the
December 31, 2005 resource model. The resource model reported herein was based
on the topographic mining surface as of December 31, 2005.

The model incorporates the results from 228 out of the 267 drill holes completed
in 2005. Table 18-3 summarizes the data added to the estimation database.

                     Table 18-3: Updated Drill Hole Database

                Geology    Gold    Arsenic   Sulfur    BWI      SG
                -------   ------   -------   ------   -----   -----
# drill holes       267      228       141      110     142     234
# data points    48,660   30,334    19,681   14,883   2,035   9,080

18.3.2 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 DEG. to the SW. Figure 18-3 shows bedding
structures typically observed in the host phyllites.

              Figure 18-3 Graded bedding in Unmineralized Phyllite

                                    [GRAPHIC]

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 DEG. 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 18-4
and 18-5 typify structural textures and arsenopyrite mineralization within the
high strain zone.

    Figure 18-4 Phyllite with Verging Asymetric Folds, Shear Bands & Boudins

                                    [GRAPHIC]

          Green - verging asymetric folds

          White - shearing

          Yellow - Foliation boudins

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              Figure 18-5 Large Arsenopyrite Porphyroblast in Core

                                    [GRAPHIC]

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 0.6 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 7-5 presents a conceptual model of the geology of the Paracatu deposit
outlining the layered interpretation. Figures 18-5 and 18-6 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 18-5 Drill Section 07N - Looking North

                                    [CHART]

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                  Figure 18-6 Drill Section 05N - Looking North

                                    [CHART]

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.

18.4 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 18-4 summarizes the basic statistics of the 1.0 meter raw sample data for
gold.

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              Table 18-4 Basic Statistics for Gold, Raw Sample Data

<TABLE>
<CAPTION>
                               Number of                                        Coefficient of    Standard
           Domain               samples     Mean   Median   Minimum   Maximum      Variation     Deviation
----------------------------   ---------   -----   ------   -------   -------   --------------   ---------
<S>                               <C>      <C>      <C>        <C>       <C>             <C>         <C>
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
</TABLE>

18.4.1 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 17.3 of this report.

                 Table 18-5: Basic Statistics for Arsenic Assays

<TABLE>
<CAPTION>
                               Number of                                       Coefficient of    Standard
           Domain               samples    Mean   Median   Minimum   Maximum      Variation     Deviation
----------------------------   ---------   ----   ------   -------   -------   --------------   ---------
<S>                               <C>      <C>      <C>        <C>     <C>              <C>       <C>
B1                                 1,905    759      612         0      6702            1.220      929.41
B2                                24,243   1148      662         0     41687            1.320     1513.96
B2 (Boudin-deformation zone)       1,238   1148     2675       250     10988            0.430     1191.02
</TABLE>

18.4.2 Bond Work Index

Hardness was assessed based on 8.0 m original composite samples that are
adjusted for the new 12.0 m bench height considered in this study. Each sample
was 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 neighbor
interpolation to estimate the BWI of individual model blocks.

Basic statistics tabled below in Table 18-6 highlight the difference in ore
hardness between the B1 and B2 horizons.

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                Table 18-6: Basic Statistics for Bond Work Index

<TABLE>
<CAPTION>
              Number of                                        Coefficient of    Standard
               samples     Mean   Median   Minimum   Maximum      Variation     Deviation
              ---------   -----   ------   -------   -------   --------------   ---------
<S>             <C>       <C>      <C>        <C>      <C>          <C>            <C>
BWI (total)     3,422     12.13    14.25      2.15     21.56         0.35          4.25
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
</TABLE>

18.4.3 Specific Gravity

Specific gravity measurements for core samples were 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 18-7 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 18-7: Basic Statistics for Specific Gravity in Core Samples

         Number of                                       Coefficient of
Domain    samples    Mean   Median   Minimum   Maximum      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

18.5 Compositing

After reviewing the statistics of the raw data, the 1.0-meter raw samples were
composited into 6.0 meter composite intervals. Compositing used a bench
compositing routine with the 6.0 meter composite length is equivalent to half
the planned mining bench height. The composite data was then extracted using the
same geological wire

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frames used to evaluate the raw 1.0 meter sample results. Each composite was
coded according to the geological unit used for the extraction.

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

Composite statistics were 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.

18.6 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.

18.7 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.

Table 18-8 summarizes the correlogram models estimated for gold, arsenic,
sulphur, Bond Work Index and density.

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                    Table 18-8: Paracatu Correlogram Summary

<TABLE>
<CAPTION>


                                             Rot.   Range   Rot.               Rot.    Range
Zone   Item   Str.   Type   Nugget    Sill     Z      Z'     X'    Range X'     Y'      Y'
----   ----   ----   ----   ------   -----   ----   -----   ----   --------   ----    ------
<S>     <C>     <C>   <C>    <C>     <C>     <C>    <C>     <C>     <C>        <C>    <C>
 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
</TABLE>

18.8 Block Model

The block model was created using a two-step process. First, a block model with
a 50 x 50 x 12 meter (x,y,z) block dimension was 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

18.8.1 Grade Interpolation

Gold grades were interpolated using Ordinary Kriging with each geological unit
(zone) estimated independently. In general, the zone solids were used as hard
boundaries and the composites must have the identical domain code item as the
solids to be used in the interpolation process. Because of the limited number of
composites, the B2 material that was outside the BDZ zone was interpolated using
both the B2 and BDZ composites. At Paracatu assay grade capping was set at the
99th percentile of the gold grade for the zone being estimated. It resulted 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 was used in all cases for grade interpolation. A minimum of 1
composite and a maximum of 10 composites were used within the search ellipsoid.
A maximum of four adjacent samples were used from the same drill hole.
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 18-9 summarizes the search parameters used to control grade interpolation
in the resource model for all items in the different zones.

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                    Table 18-9 Grade Interpolation Parameters

<TABLE>
<CAPTION>
                       Bearing       Dip        Plunge    Radius 1   Radius 2   Radius 3
Variable   Rocktype   (Degrees)   (degrees)   (degrees)   (meters)   (meters)   (meters)
--------   --------   ---------   ---------   ---------   --------   --------   --------
<S>        <C>              <C>          <C>         <C>   <C>        <C>         <C>
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
</TABLE>

18.8.2 Specific Gravity

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

18.8.3 Ore Hardness

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

18.8.4 Recovery

Unique metallurgical recoveries were estimated for each model block (50 x 50 x
12 meters) based on the arsenic and sulphur block grades. estimated by Ordinary
Kriging.

The metallurgical recovery was 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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18.8.5 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 18-10 for both the LM and NLM
models.

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             Table 18-10 Comparison of LM vs NLM West of Rico Creek

<TABLE>
<CAPTION>
                           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)
--------   -----------   --------   ---------   -----------   --------   ---------
<S>            <C>           <C>    <C>             <C>           <C>    <C>
    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
</TABLE>

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.

18.9 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 July 2006 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%

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of 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.

18.10 Pit Optimization

18.10.1 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 was
compared to the grade tonnage curve for the model imported to Whittle 4X(C) to
ensure there were no transcription errors during the manipulation from one
software system to the other. Table 18-11 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 18-11 Grade Tonnage Summary of Imported and Exported Model

                Exported Model (Vulcan)        Imported Model (Datamine-Whittle)
           ---------------------------------   ---------------------------------
 Cutoff      tonnes     Avg Au        Au         tonnes     Avg Au        Au
(Au g/t)    (000's)    (Au g/t)    (ounces)     (000's)    (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 18-12 and represent the Base Case.

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                 Table 18-12: Base Case Optimization Parameters

Parameter                            Value
------------------------   -------------------------
Mining Cost                          $0.75
Mining Recovery                       100%
Mining Dilution                        0%
Pit Slope Angles                    55(DEG.)
Process Cost (incl. G&A)   Contained in Model Blocks
Process Recovery Au        Contained in Model Blocks
F.E.X. (R$:US$)                      2.65:1
Reserve Gold Price                 $US400/oz
Resource Gold Price                $US450/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.

Whittle(R) generates a series of pit shells during the optimization by applying
a revenue factor that varies the price of gold. These factors were input as a
range within the Whittle(R) optimization program. For the Paracatu pit
optimizations, the revenue factors ranged from a low of 0.2, which represented a
gold price of US$80 per ounce (i.e. 0.2 x US$400), to a high of 2.5, which
represented a gold price of US$1000 per ounce (i.e. 2.5 x US$400), in increments
of 0.05. Given this range, Whittle(R) produced a series of 47 nested pit shells
(i.e.(((2.5-0.2)/0.05)+1)=47). Optimization results are also presented
graphically in Figure 18-7 summarizing the base case optimization results for
Paracatu.

                   Figure 18-7 Base Case Whittle 4X(C) Results

                                    [CHART]

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The lines on the graph represent the optimum pit shell (15) and the pit shell
selected for design (18). In Kinross' opinion, it was considered that a 3%
decrease in Net Revenue in order to achieve a 20% increase in contained gold
ounces would be an acceptable risk when selecting the pit shell. It was further
reasoned that the optimization process was an approximation based on preliminary
data and that for the purpose of this study it would be appropriate to use the
shell number 18 as the basis for completing the pit design.

18.10.2 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 include 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.

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

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.

18.10.3 Pit Design

To design a practical open pit for Paracatu, the selected pit shell (18)
developed in Whittle 4X(C) was imported into Datamine(C), commercial mining
software. The chosen pit shell was 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 18-13.

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                        Table 18-13: Pit Design Criteria

      Parameter          Value
--------------------   --------
Bench Height             12m.
Bench Face Angle        75 DEG.
Inter-ramp Angle        55 DEG.
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 18-8

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                      Figure 18-8 Typical Haul Road Profile

                                    [CHART]

Mineral reserves were estimated by reporting the model blocks within the design
pit above the incremental cut off grade described in section 18-10.2. Resource
model blocks classified as were reported as Proven Reserves, model blocks
classified as Indicated were reported as 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.

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19.0 OTHER RELEVANT DATA AND INFORMATION

This section is not applicable to the Paracatu mine.

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20.0 INTERPRETATION AND 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.

The 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.

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

Basic engineering that started in 2005 has culminated in a control cost estimate
with an accuracy level of +/-15%. As part of the Engineering study, SNC-Lavalin
confirmed that the 38' SAG mill would be adequate to attain the 61 Mtpa
production level and the study has quantified the capital and operating costs to
support this Expansion III project.

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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21.0 RECOMMENDATIONS

On the basis of the 2006 Feasibility Study, RPM have requested full release of
funds for continuing the implementation of the Expansion III project. Kinross
has reviewed the data and conclusions presented by RPM and are in agreement with
their recommendation to proceed with the planned expansion.

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 data from the 2005 drill campaign 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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22.0 ADDITIONAL INFORMATION FOR OPERATING PROPERTIES

22.1 Project Implementation Plan

The scope of the Expansion III project is to increase the present ore production
from approximately 18 Mtpa to approximately 61 Mtpa by the installation of a new
41 Mtpa treatment plant, designed to treat the harder B2 sulphide ore being
encountered as the mine goes deeper. The existing plant will treat the softer
near-surface B1 ore at a throughput rate of 20 Mtpa until these reserves are
depleted.

Project Management of the Feasibility Study and Implementation Phases is vested
in an Owners' Team, drawn largely from Rio Paracatu Mineracao (RPM). This team
incorporated local Consultants, mainly ex-employees of RPM who had participated
in previous expansions, and other international Consultants of repute. This RPM
team was placed under the direction of an internationally experienced Project
Director, provided by the Kinross Group.

An Engineering Procurement and Construction Management contract for the
Expansion of the process plants and associated infrastructure was awarded to a
joint venture of the local Brazilian company Minerconsult and the Canadian
company SNC Lavalin. The procurement element of this contract is limited to the
technical recommendation together with placing of orders, expediting, QA/QC and
transport; the commercial aspects will be handled by the Owners' Team. SNC
Lavalin is an extremely well qualified Canadian company with a track record of
successfully completing major mining projects throughout the world. Minerconsult
is a Brazilian company with some years in the mining scene in Brazil. Many of
its senior managers and directors have worked for older companies in the
Brazilian market place, and have great experience locally completing projects
for the iron ore and fertilizer industries. They are also currently carrying out
expansion projects

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for Anglo Gold at Morro Velho and again for Anglo in the Barro Alto nickel
project. Together these companies provide strong local and international
experience.

Supply of electrical power is a constraint and RPM are negotiating with current
provider CEMIG, to supply the full 90 MW electrical demand that will be required
in 2009. This demand necessitates the upgrading of the 43-SE-01 sub-station and
a separate contract has been awarded to CEMIG to design and execute the upgrade.
However due to limitations in CEMIG capacity this demand has to be satisfied by
means of a regional development project which requires the installation of a 500
kV line from Emborcacao to Luziania.. This line is currently being constructed
and is scheduled for completion by mid 2008.

Golder Associates has been contracted to carry out a Feasibility Study of the
Tailings Disposal requirements. These will necessitate a completely new Tailings
Dam probably constructed by a similar centerline method as the existing dam.
Again special attention will be given to arsenic and sulfur control as in the
existing facility, and the new dam will also serve as a water storage reservoir.
While the conventional earth dam construction has been considered for this
study, further options are under development by Golder Associates through other
studies.

A simplified implementation schedule is shown in Figure 22-1 below:

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          Simplified Implementation Plan for the Expansion III Project

                             [GRAPHIC APPEARS HERE]

           Figure 22-1 Paracatu Expansion III Implementation Schedule

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22.2 Mining

The mine design prepared for this study was completed for the purpose of
preparing a feasibility level cost estimate. All mine engineering and reserve
estimation work was completed by professional mining engineers who meet the
requirements of "qualified persons" as required by Canada's NI 43-101
legislation.

Pit and waste dump design recommendations were prepared by Golder Associates of
Belo Horizonte, Brazil (Golder).

The open pit design criteria used for the pit design are summarized below:

o    Bench Height        12m

o    Bench Face Angle    70 DEG.

o    Berm Width          8.4m

o    Berm Interval       24m

o    Inter-ramp Angles   (Weathered Rock) 38 DEG.

o    Inter-ramp Angles   (Fresh Rock) 55 DEG.

o    Haul Roads: Main haul roads and in-pit ramps were designed at 10% gradient
     and 30m width, based on approximately four times the width of a 220t truck
     (~7.4m). This provided room for 2-way traffic and included a drainage ditch
     and safety berm.

In addition, the following hard boundary limits were used to prevent the
feasibility study pit from impacting existing critical infrastructure,
specifically, this included:

o    Protect the plant site with a 100 meter buffer zone;

o    Protect the townsite with a 200 meters buffer zone; and\

o    Protect the existing highway with a 200 meter buffer zone.

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Pit design work was completed with Datamine(R) software based on optimization
results from Whittle(R) software. Pit optimizations for the mine design assumed
the following economic parameters:

o    $400/oz gold price;

o    A Foreign Exchange rate of R2.65 per $US;

o    Operating costs based on the 2006 Feasibility Study data;

o    Mine design parameters as recommended by Golder Associates.

Golder recommended the following design criteria for waste dump storage:

o    Lift height = 12 meters;

o    Angle of Repose = 37 degrees;

o    Catch Berm Width = 15 meters; and

o    Overall Slope Angle = 26.5 degrees (1:2 - V:H)

On completion of the designs, bench-by-bench reserves were estimated using
Datamine(R) software. The total bench-by-bench reserve totals constitute the
entire mineral reserve estimate. A detailed life-of-mine schedule was prepared
tracking the B1 ore, B2 ore and waste mining schedule. The LOM plan segregated
waste mining by overburden, saprolite, oxide and sulphide to allow segregation
for purposes of Acid Rock Drainage (ARD) control.

It was assumed for this study that the sulphide waste rock would be backhauled
into the NE portion of the ultimate pit, which in this plan will be mined early
in the project life. A separate stockpile will be constructed for overburden and
saprolite storage which will

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enable the use of these materials for reclamation to further reduce acid rock
drainage (ARD) potential once mining has been completed.

The LOM plan, interim and ultimate pit designs, plant infrastructure and waste
dump locations were used to estimate the mining fleet requirements, manpower
loading and "first principles" mine operating cost estimate for the Feasibility
Study. The mining fleet requirements for the next four years of production at
Paracatu are as follows:

                  Table 22-1: Mine Fleet Requirements 2006-2009

                                              No. of
    Equipment Type        Size       Model     Units
---------------------   --------   --------   ------
Electric Cable Shovel     35 m3    P&H 2800      1
Loaders 992               12 m3     CAT 992      2
Loaders 994               18 m3     CAT 994      1
Haul Trucks 777            91t      CAT 777     13
Haul Trucks 793           220t      CAT 793      9
Drills                   200 mm.    IR DML       2
Dozers                     kW      CAT D10R      3
Graders                    kW       CAT 16H      1
Wheel Dozers               kW       CAT 834      2
Water Trucks            100,000l    CAT 777      1

Production rates vary by material type. The B1 ore was scheduled at a throughput
rate of 20Mtpa. Nominally, the B2 ore was scheduled at a rate of 41Mtpa. As the
bond work index for the B2 ore increased the throughput rate decreased.

It was estimated that at the given plant throughput rates that the reserves at
Paracatu would be depleted in year 2036. Total mine life would be 31 years as of
the end of 2005. The table below presents the life of mine plan for the Project.

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                        Table 22-2 Paracatu Life of Mine

                                    Schedule

<TABLE>
<CAPTION>
                         1         2         3         4         5         6         7         8
    PRODUCTION          2006      2007      2008      2009      2010      2011      2012      2013
-------------------   -------   -------   -------   -------   -------   -------   -------   -------
<S>                   <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>
B1-Ore Ktonnes          7,788     6,874    15,464    19,593    19,997    18,505    19,093    19,469
B1-Ore g/t               0.34      0.34      0.34      0.34      0.36      0.34      0.27      0.32
B2-Ore Ktonnes          9,518    10,311    24,749    41,001    41,003    40,468    39,288    33,532
B2-Ore g/t               0.45      0.42      0.43      0.42      0.41      0.42      0.33      0.36
Total Ore Ktonnes      17,306    17,185    40,213    60,594    61,000    58,973    58,381    53,002
Total Ore g/t            0.40      0.39      0.40      0.40      0.39      0.40      0.31      0.34
Total Waste Ktonnes         0         0     1,640     2,305     1,476     4,702     8,262    22,306
Total Ktonnes          17,306    17,185    41,853    62,900    62,475    63,675    66,643    75,308
Total Work Index         6.33      7.00      6.25      5.59      5.46      6.05      7.32      8.75
Avg. Recovery            78.8%     78.2%     80.1%     80.3%     79.9%     79.8%     81.1%     80.2%
Recovered Au Oz       175,439   167,570   412,564   622,541   615,858   599,720   475,661   469,799
Contained Au Oz       222,479   214,354   515,106   775,468   770,606   751,462   586,686   585,611

<CAPTION>
                         9        10         11        12        13        14        15        16
    PRODUCTION          2014     2015       2016      2017      2018      2019      2020      2021
-------------------   -------   -------   -------   -------   -------   -------   -------   -------
<S>                   <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>
B1-Ore Ktonnes         15,794    10,985         0         0         0         0         0         0
B1-Ore g/t               0.34      0.33      0.00      0.00      0.00      0.00      0.00      0.00
B2-Ore Ktonnes         32,783    37,887    41,001    40,999    41,001    40,139    39,744    40,994
B2-Ore g/t               0.36      0.36      0.43      0.38      0.36      0.36      0.35      0.34
Total Ore Ktonnes      48,578    48,872    41,001    40,999    41,001    40,139    39,744    40,994
Total Ore g/t            0.36      0.36      0.43      0.38      0.36      0.36      0.35      0.34
Total Waste Ktonnes    26,288    19,344    17,923    21,789    17,611    15,271    17,450    18,156
Total Ktonnes          74,866    68,216    58,925    62,788    58,612    55,410    57,194    59,151
Total Work Index         9.73      9.62     11.16     10.51     10.87     11.60     11.60     11.13
Avg. Recovery            80.2%     79.9%     79.0%     79.7%     79.7%     80.0%     79.9%     80.2%
Recovered Au Oz       448,101   446,924   444,751   396,957   378,160   366,967   353,483   359,349
Contained Au Oz       558,379   559,301   562,782   498,117   474,431   458,455   442,091   448,099
</TABLE>

<TABLE>
<CAPTION>
                         17        18        19        20        21        22        23        24
    PRODUCTION          2022      2023      2024      2025      2026      2027      2028      2029
-------------------   -------   -------   -------   -------   -------   -------   -------   -------
<S>                   <C>       <C>       <C>       <C>       <C>       <C>       <C>       <C>
B1-Ore Ktonnes              0         0         0         0         0         0         0         0
B1-Ore g/t               0.00      0.00      0.00      0.00      0.00      0.00      0.00      0.00
B2-Ore Ktonnes         39,973    40,299    37,855    35,924    34,385    33,841    32,899    33,051
B2-Ore g/t               0.34      0.40      0.40      0.41      0.42      0.42      0.44      0.47
Total Ore Ktonnes      39,973    40,299    37,855    35,924    34,385    33,841    32,899    33,051
Total Ore g/t            0.34      0.40      0.40      0.41      0.42      0.42      0.44      0.47
Total Waste Ktonnes    23,691    19,336    26,000    26,425    27,094    25,827    25,481    25,371
% PAG                    65.3%     81.7%     88.5%     92.7%     92.8%     92.6%     92.2%     90.5%
Total Ktonnes          63,665    59,636    63,855    62,348    61,479    59,668    58,380    58,422
Total Work Index        11.48     11.16     12.49     13.16     13.75     13.97     14.37     14.31
Avg. Recovery            79.7%     78.9%     79.0%     79.0%     79.0%     79.1%     79.2%     79.3%
Recovered Au Oz       348,267   404,106   385,431   377,007   367,339   362,952   370,059   397,780
Contained Au Oz       436,867   511,895   487,857   477,261   464,747   459,018   467,133   501,380

<CAPTION>
                         25        26        27        28        29        30        31
    PRODUCTION          2030      2031      2032      2033      2034      2035      2036      Total
-------------------   -------   -------   -------   -------   -------   -------   -------  ----------
<S>                   <C>       <C>       <C>       <C>       <C>       <C>       <C>      <C>
B1-Ore Ktonnes              0         0         0         0         0         0      714      154,278
B1-Ore g/t               0.00      0.00      0.00      0.00      0.00      0.00     0.45         0.34
B2-Ore Ktonnes         31,713    31,998    31,341    30,787    30,198    29,825    7,226    1,035,737
B2-Ore g/t               0.50      0.48      0.47      0.50      0.48      0.54     0.29         0.41
Total Ore Ktonnes      31,713    31,998    31,341    30,787    30,198    29,825    7,941    1,190,015
Total Ore g/t            0.50      0.48      0.47      0.50      0.48      0.54     0.31         0.40
Total Waste Ktonnes    25,398    24,323    22,666     9,985     1,643       642    2,928      481,335
% PAG                    88.4%     88.7%     94.4%     93.6%     58.4%     26.5%    29.5%        64.2%
Total Ktonnes          57,111    56,321    54,007    40,772    31,841    30,467   10,869    1,671,350
Total Work Index        14.91     14.78     14.52     14.79     15.09     15.27     9.66        10.66
Avg. Recovery            79.0%     79.1%     79.3%     78.9%     79.6%     79.5%    83.0%        79.6%
Recovered Au Oz       400,020   388,213   373,255   393,053   373,299   408,515   62,851   12,145,993
Contained Au Oz       506,291   490,521   470,458   497,845   469,048   514,068   75,727   15,259,441
</TABLE>

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22.3 Process Plant

22.3.1 Existing Circuit

The existing RPM process plant has a nominal capacity of 18 Mtpa when processing
ore with a work index of up to 8 kWh/t. The facility consists of crushing to
minus 25 mm, closed circuit ball mill grinding to 80% passing 75 micron and
flotation to produce a sulphide concentrate. A gravity concentrate is also
produced from jigs located in the ball mill circuit. The concentrate products
are reground in a ball mill and leached with cyanide in a carbon in pulp
circuit. Gold is recovered from the carbon and smelted to produce a dore bar for
export. Figure 22-2 is a simplified flow sheet of the current process plant. The
production statistics in the flow sheet are budget estimates only.

       Figure 22-2: Simplified Flow Sheet Existing Paracatu Process Plant

                                  [FLOW CHART]

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22.3.2 New Circuit

The new process plant has a nominal capacity of 41 Mtpa when processing ore with
a work index below 8.7 kWh/t. Tonnage throughput will decrease as work index
increases.

Primary Crushing

The primary crusher is located within the open pit and is scheduled to operate
for 18 hours per day. Run-of-mine ore is delivered by 240 tonne capacity,
rear-dump, haulage trucks to the 480 tonne capacity crusher dump hopper. An
apron feeder mounted at a 15 DEG. inclination withdraws run-of-mine ore from the
dump hopper and feeds it at a controlled rate of approximately 6,240 t/hr to an
MMD 1300 Series Twin Shaft Sizer. The MMD Sizer crushes the rock from a maximum
size of 1300 mm to a nominal size of 350 mm and discharges directly onto a
"sacrificial" conveyor, which in turn discharges onto the overland conveyor.

Three truck-dump positions are provided to allow the MMD Sizer to operate at its
maximum instantaneous capacity. The middle position is smaller than the two side
positions and can accommodate up to 150 tonne trucks.

A stationary hydraulic rock breaker located at the MMD Sizer feed chamber is
used to break oversize rock that may be delivered. Large rocks can be ejected
from the Sizer itself, through a side-door in the hopper by the action of the
Sizer. The Sizer can be removed from its operating position to a maintenance
position by a winch, slide rails are provided. A tower crane is used for
maintenance of heavy components.

Stockpile

The crushed ore stockpile is a rectangular "A" frame with a dust and rain-cover.
Ore is delivered into the stockpile by a tripper conveyor. The reclaim tunnel
and a small part of

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the 45,000 t live volume of the stockpile is situated below ground. The total
storage volume is 282,000 t, which can be accessed by using a dozer.

The reclaim tunnel has six variable speed belt feeders. During normal operation
five of these feeders feed the SAG mill feed conveyor. The reclaim tunnel has a
baghouse, escape tunnel, access stairway and a sump pump.

Grinding

The grinding circuit is designed to operate at 5,100 t/hr at an availability of
92%. The grinding circuit consists of one 11.6 m diameter by 6.7 m long (38'
diameter by 22' long EGL) 20 MW SAG mill followed by two parallel 7.3 m diameter
by 12.0 m long EGL (24' diameter by 39.5' long) 13 MW ball mills. The SAG mill
operates in closed circuit with a trommel screen and vibrating screen and the
ball mills operate in closed circuit with hydrocyclones. The ball mills are
equipped with a single trommel magnet to remove tramp steel from the mill
discharge. Plant capacity has been selected to give a nominal flotation feed
grind of 80% passing 75 um.

The SAG mill is driven by a 20,000 kW (26,800 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 scheduled in the mine
plan. Each ring gear driven ball mill is powered by twin pinion gears driven by
two 6,500 kW (8,720 hp) fixed-speed, wound rotor motors.

Slurry discharges from the SAG mill through a trommel screen onto a double-deck
vibrating screen. The top deck aperture is 25 mm and the bottom deck aperture 12
mm. A standby screen is included adjacent to the operating screen, to allow
rapid change-out. This change-out arrangement is based on an existing design,
used at the Kinross Fort Knox mine in Fairbanks, Alaska.

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Screen oversize is transferred to the SAG mill feed conveyor by three pebble
conveyors in series. A weigh scale is mounted on the first conveyor to monitor
the recycle rate to ensure that the conveyors are not overloaded due to unusual
ore conditions. A pebble crusher is not provided for crushing SAG mill oversize,
although provision is made for a possible future installation.

SAG mill discharge screen undersize and trommel screen undersize flow by gravity
to a pump box. SAG mill discharge is pumped from the SAG mill discharge pump box
to the ball mill cyclone feed pump box. A distribution box is used to split the
slurry from the SAG mill to the two ball mill pump boxes. Water is added to the
cyclone feed pump box at a controlled rate to produce a cyclone overflow that
contains 39% solids by weight. Ball mill product also discharges to the cyclone
feed pump box. The combined slurry is pumped by a variable speed pump to three
cyclone clusters, one for each mill, each containing up to fourteen 660 mm
diameter (26" diameter) cyclones. Cyclone underflow, at 73% solids, is
recirculated to the ball mill for additional grinding. Cyclone overflow flows by
gravity to the rougher flotation feed distribution box.

The capacity of the cyclone feed pump and cyclone cluster are adequate for a
circulating load of 250% (solids mass flow to cyclone underflow is two and one
half times solids mass flow to cyclone overflow). Flotation collector and
frother can be added to the cyclone feed pump box to provide additional
conditioning time.

Liner handlers are provided for the SAG mills and the ball mills. One liner
handler services the SAG mill and a second one is provided to service the two
ball mills. Jib cranes are located at each mill feed end to transfer new liners
and scrap liners between the floor area, accessible by the overhead cranes, and
the liner handlers.

A 100 tonne capacity overhead crane with a 25 tonne auxiliary hoist is provided
for the grinding bay.

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Gravity Separation / Gravity Concentration

There are six jigs using the radial "Pan-American" design successfully used on
the existing plant. Each pair of jigs are fed by a slurry distributor which in
turn are fed from variable speed pumps located on the cyclone feed box. The jigs
are arranged so that the concentrate flows by gravity to a sieve bend. The
coarse products from the sieve bend flow by gravity to the regrind mill and the
fine product to the solution removal thickener. The jig tails gravitate back to
the cyclone feed pump box.

Flotation and Regrinding

The flotation circuit consists of rougher flotation followed by a single stage
of cleaning of the rougher concentrate. Cleaner tails are recirculated to the
rougher flotation cells and the cleaner concentrate is ground using a vertical
stirred regrind ball mill operating in closed circuit with hydrocyclones.
Rougher tailings are discarded to the tailings mix tank by gravity.

Flotation collector and frother are added to the slurry as it enters the
four-way distribution box. A total of 24 rougher flotation cells are included,
arranged in four rows of six cells each. The cells are 160 m3 tanks cells fitted
with self-aspirating mechanisms. Rougher flotation tailings streams are combined
and are transported by gravity pipeline to the tailings mix tanks to mix with
oxide (B-1) tailings from the existing plant. From the tailings mix tank, the
tailings can gravitate to either tailings dam. Rougher flotation concentrate is
collected in a single pump box and pumped by horizontal pumps to the cleaner
cells.

The cleaner cells consist of two rows of five self-aspirated 60 m3 tank cells.
Cleaner tailings flow to a pump box and are pumped by a horizontal slurry pump
to the rougher flotation distribution box. Cleaner concentrate is collected in a
single pump box and pumped by horizontal slurry pumps to the solution removal
thickener.

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Both the gravity concentrate and the cleaner concentrate report to the Solution
Removal thickener. Sieve bends are used to remove the coarse heavy particles in
the gravity concentrate and direct them to the regrind mill. The purpose of the
thickener is to remove solution containing flotation reagents and also to
prepare the slurry to the optimum density for regrinding.

The concentrate regrind mill grinds the combined cleaner and gravity concentrate
to approximately 80% passing 40 im. Concentrates, combined with regrind mill
discharge, are pumped by variable speed pumps to a single cyclone cluster,
containing ten 254 mm diameter (10" diameter) cyclones. The 13.5 m high (44')
vertical stirred regrind ball mill is driven by a 931 kW (1250 hp) fixed speed
motor.

Carbon In Leach Circuit

The existing circuit will be upgraded from 35 t/hr to 100 t/hr via installation
of a new pre-aeration tank and 4 new CIL tanks. Pre-aeration residence time will
be 4 hours and CIL residence time will be 36 hours.

Concentrate thickener underflow is pumped to a new trash screen located over the
new pre-aeration tank at the CIL plant. The cleaned slurry passes through the
screen to the 750 m3 agitated pre-aeration tank where milk of lime slurry is
added. This tank overflows to four new 750 m3 CIL tanks. The flow from the
fourth tank is split to flow into two lines of 4 existing 300 m3 CIL tanks.
Carbon is retained in each tank by swept "NKM"-type pumping screens. The screens
in the existing tanks will be upgraded. Cyanide and lime from existing make-up
systems are staged added to each tank train. On an intermittent basis, loaded
carbon is pumped counter current to the slurry flow in order to increase the
gold loading. Loaded carbon is removed from the first CIL tank and is
transferred to the new loaded carbon bin after draining and washing on two new
vibrating screens. A carbon safety screen is located over the CIL tails pump
box.

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Carbon Elution and Regeneration

The new carbon elution and regeneration plant is located in a building
previously occupied by a redundant filtration plant. The plant is designed to
handle a carbon batch size of 14 tonnes at 4,500 g Au/t.

The pregnant and barren eluate tanks are located in an area previously occupied
by two small regeneration kilns. The new 600 kg/h kiln will be constructed and
commissioned prior to demolishing the smaller existing kilns.

Loaded carbon is pumped to the new acid wash vessel. After acid washing with 3%
HCl is complete, the spent acid is neutralized with sodium hydroxide from the
existing make-up system before discarding it to the tails pump box.

The elution cycle operates with a 0.2% sodium cyanide and 1% sodium hydroxide
solution at a temperature of 145 DEG. C and a pressure of 450 kPa for
approximately 7 1/2 hours.

Stripped carbon is evacuated from the bottom of the elution vessel and the
activity of the stripped carbon is restored in a new regeneration kiln.

Electrowinning and Refining

Pregnant solution is pumped to new four electro-winning cells located adjacent
to the existing single cell on the upper floor of the existing refinery. Gold
metal is electro-won on the stainless steel wool cathodes in the electro-winning
cells. At the end of the run, the cathodes are removed from the cells and the
gold bearing sludge is recovered in a small new filter press. The filter cake is
mixed with fluxes, usually borax, soda ash and occasionally sodium nitrate and
fed to an existing electric induction furnace. The dore

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metal and slag separate in the furnace, and the slag is poured off to slag pots
then the dore metal is poured into bars for shipment.

22.4 Tailings Disposal and Reclaim Water

Previous studies completed in 2005 identified the need to build a new tailings
disposal facility rather than continue to raise the existing dam.

At that time, because of a 20 year record of successful tailings disposal
management it was decided to maintain the dam construction methodology and
disposal system in order to facilitate the licensing of the new facility. This
is the technical solution that has been maintained for the purposes of cost
estimating for this Feasibility Study.

Golder Associates has confirmed the site selection chosen by RPM and prepared a
Basic Engineering design for the new starter dam and centerline deposition
system. Costs of construction and operation of the new facility have been
applied to the Golder Basic Engineering quantities by RPM, based on actual
contracted rates and operating costs currently experienced by RPM.

Licensing of the new facility will be based on exactly the same parameters on
which the current facility is licensed viz. maintaining PH, arsenic and sulfur
levels within strictly controlled limits. RPM's success in maintaining control
within these limits is evidenced by the presence of fish in the pond. This is
achieved under current mining and processing condition by blending of the B1 and
B2 tailings before disposing them together in the dam.

This will continue until exhaustion of the B1 ore in 2015. The Expansion III
project will provide more sophisticated arsenic recovery equipment, and when B1
ore is exhausted a desulfurization plant will be installed. Meanwhile tests are
now in progress to utilize much of the sulfur in the tailings for the production
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a valuable by product in Brazilian terms it will be an important environmental
protection project. This will be the subject of a separate Feasibility Study.

The new tailings disposal dam is important for water capture; together with the
existing dam basin it will provide sufficient drainage area, an additional 49
square Kms, for the extra water required for the expansion project. Water from
the new dam will be pumped or eventually gravitated into the existing dam and
pumped back with the existing pumping installations duly expanded.

22.5 Infrastructure

For the existing mine installations RPM is fully equipped in respect of mine and
process plant workshops, power supply, water supply, communications, fire
protection, sewage, site drainage road access etc. All this will continue to
serve and service the existing installations.

New service stations, maintenance workshops, fuel storage and fuelling stations
will be required for the large new higher capacity mining fleet. In addition,
the new Process Plant installations will be removed somewhat from the old, and
consequently, new infrastructure such as power supply, water supply,
communications, site drainage, sewage and road access will be required.

The increased power demand cannot at present be supplied from the CEMIG
distribution network until a new 500kV transmission line is built by the
National Grid; this is in progress and current predictions are that it will be
completed by mid 2008, in advance of the contract date of January 2009.
Negotiations are in hand to furnish RPM with a 500/230kV supply directly from
the National Grid instead of a 138kV supply from the Minas Gerais State power
Company CEMIG. This would require RPM to build a 500/230kV sub-station and a 30
km long power line. This brings the advantage of a much reduced tariff,
US$47/MWh as against the current 138kV tariff, US$66MWh. RPM will also
investigate participation in a privately sponsored hydroelectric power plant.

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Accommodation blocks, office buildings, storage facilities and temporary
infrastructure will be provided to service the Construction phase and will be
utilized after construction by mine operating personnel who will eventually be
displaced from their current accommodation by mining activities.

22.6 Markets and Contracts

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.

22.7 Occupational Health, Safety and Environmental Aspects

22.7.1 Occupational Health and Safety Aspects

Occupational health and safety aspects are not expected to significantly change
with the RPM expansion project. These aspects are mainly related to noise, dust
and vibration emissions from new equipment and facilities.

RPM health and safety system includes a series of tools to effectively identify
health and safety risks and specific standards and procedures to control them.
The same approach will be adopted to cover the expansion project occupational
health and safety related risks.

The health program initiates with regular monitoring of dust, noise, vibration,
thermal stress, etc in the workplace for SEG - Similar Exposure Groups. These
monitoring campaigns provide RPM with the required information to propose
hierarchy criteria of controls in this sequence:

     1.   elimination of the risks at source in the planning stage of the
          project: purchase of low noise equipment and vehicles and or supplied
          with devices to reduce dust and noise emissions;

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     2.   reduction of the risks through encapsulation or enclosure of equipment
          and systems or

     3.   adoption of appropriate PPE - Personal Protective Equipment for
          specific tasks and exposure areas.

The RPM safety programme includes a risk inventory for activities routinely
performed and preliminary risk assessment for new tasks. Safety operational
procedures, permits for critical works and adoption of behaviour safety
observations are the main tools used to control safety related risks at all RPM
premises including contractor facilities and works. Fire fighting and Emergency
Programmes are also in place for small incidents as well as a Contingency Plan
for addressing serious incidents.

22.7.2 Environmental Aspects

Diversion of Rico Creek

In order to allow mining of the west portion of the pit, RPM will intercept a
small branch of a local water course, Rico Creek. This process required RPM to
obtain pertinent permits with the State Water Agency - IGAM. The Rico Creek has
been intensively impacted in the past by small prospector activities
("garimpeiros") and, hence, siltation and mercury contamination occur along its
course. The Rico Creek has also received domestic sewage from the city of
Paracatu for a long time. Recently, the local water and sewage company has
installed a collecting sewage system in order to treat it in a new sewage plant
that was put in operation in October 2005.

As part of its mining plan, RPM has proposed to the environmental agency to
divert a portion of the Rico Creek to a small canal while mining in this portion
of the mine. Pertinent hydrogeological studies were performed by RPM and a
minimum flow will be maintained in accordance with what was recommended in these
studies (30 m3/h)

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In the end of the mine life, a lake will be formed and integrated into the
landscape and the original path flow of the creek re-established. The overall
area will be reclaimed and vegetated accordingly to a specific Closure Plan. RPM
has also agreed with the local authorities, the Paracatu mayor and the local
based Public Attorney, a rehabilitation project to recontour and re-establish
the creek nearby woods in order to create leisure areas in the downstream
portions of the creek that cross the city, as part of the compensation measures
of the project.

Tailings Disposal

The main tailings facility is subject to a rigorous auditing and reviewing
process. In 2004, a workshop was held with external consultants and reviewers
which consolidated all previous expert findings and obtained a consensual
evaluation of the project and construction up to that date.

The key finding of this workshop is that dam can be raised by a modified
centre-line method. The implications of accelerated tailings placement is taken
into account for this strategy and is incorporated into the scope of work for
the next stages of the dam raise. There are no direct environmental impacts
associated with a faster rate of tailings deposition.

The 61Mt expansion scenario will require a new tailings dam facility to allow
storage of 1.2 billion tones of mineral waste for the life of the mine. The
requirement for this new facility is justified by the increasing amount of
borrowing soils that RPM needs for further raises of the current dam embankment,
which has significantly impacted on construction costs and on new areas where
removal of vegetation and top soil is required.

The new tailings facility should cover approximately 2,500 hectares and should
be constructed on a nearby valley next to the current dam. In the region that
includes the area of the new dam, the disappearance of primitive vegetal
formations, the degradation

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of quality of water courses by anthropic actions in the past and predatory
hunting have been the three major factors responsible for the disappearance of
typical species of the site. Some regions had most part of their natural
habitats converted into pastures and crops (Biodiversitas, 1998).

A few endangered and endemic species have been registered in the area of the new
tailings dam. However, these species have also been found in the operation
nearby environments that will be protected by RPM, as part of the minimization
and compensation measures established in the Environmental Impact Assessement-
EIA-RIMA (Brandt Meio Ambiente, 2006).

CIL Tailings Management

In order to clean the tailings that are stored in the tailings dam from sulphide
and metals, RPM has collected and segregated a sulfur concentrate in specific
sumps, which are located in the mine pit area. This has ensured that the water
quality from the present tailings pond and discharged into the Santo Antonio
Creek complies with the local regulatory standards and World Bank guidelines
(Limits for Process Wastewater, Domestic Sewage and Contaminated Stormwater
Discharged to Surface Waters for General Application, July 1998).

The amount of concentrated sulphide tailings generated will not increase,
although the rate of deposition will be accelerated. The current management of
these tailings is considered to provide appropriate containment with the
installation of composite liners at the floor of the tailings pond and subdrains
to intercept and collect any potential leakage. Multi-layer cover on closure has
been assessed.

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Cyanide Management

RPM has used cyanide as part of its CIL circuit, where it is used to leach the
gold and passing it into solution. In the current process, the CIL tailings
containing a sulphur concentrate and cyanide is send to sumps located in the
mine area, where the sulphur concentrate tails are settled and the cyanide
solution returned to an AVR (Acid, Volatization and Recovery) plant which
recovers around 60% of the cyanide and recycle it into the process. Finally, the
residual cyanide is discharged together with the flotation tailings in the
tailings dam, where it is naturally degraded by volatilization and the UV sum
light. RPM keeps a rigorous monitoring of residual cyanide discharge and
degradation on the tailings dam pond, where concentrations have been below the
World Bank General Environmental Guidelines: < 0.1 mg/L for free cyanide and <
1.0 mg/L for total cyanide. This information is forwarded monthly to the
environmental agency as part of the RPM operational licence (LO) requirements.

In the last years, RPM has also implemented a series of improvements to reduce
cyanide consumption such as the implementation of automatic feed systems, which
has allowed to reduce the specific use (per unit) of this reagent in more than
50%. A further improvement that has been introduced in the expansion design is a
pre-aeration step prior to leaching, which will enable the current cyanide
consumption levels to be maintained (and not increased) despite the higher
amount of concentrate that will have to be processed.

The 61Mt Expansion Project should slightly improve the current strategy for
cyanide management. The current AVR circuit should be discontinued and a SO2 -
air (similar to INCO Process) circuit installed to treat cyanide effluent from
the CIL process. This will allow RPM to eliminate almost completely the residual
cyanide in the final effluent that nowadays goes to the tailings pond where it
is destroyed by the sum light (photo-degradation).

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Dust, Vibration and Noise Management

Replacing the existing Crushing Plant with a mineral sizer and a SAG mill is
expected to have a positive impact on occupational health management, due to a
decrease in dust generation, particularly due to reduction of truck traffic in
the mine area. The conveying system transporting the crushed ore from the mining
area to the SAG mill will be semi-enclosed and will be running at a low
velocity. Dust collection systems will be installed at transfer points. An air
quality prediction model used by RPM to investigate the future impact from new
equipment on the local environment demonstrated that no additional impact is
expected with the installation of those at the mine and beneficiation areas.
Vibration levels will also be kept below the standard. 5 HI-VOL (high volume)
samplers are installed in strategic points around the mine site since the
beginning of RPM operation in Paracatu to monitor the air quality. Information
provided by these samplers will be used as reference for controlling the dust
emissions to local communities.

Preliminary analysis performed by RPM during blasting studies and drills showed
that noise generated by the future mine activities and by the in-pit mineral
crusher is within guidelines set by the World Bank, or below 70 dB(A). Given the
location of the conveying system and its low operating velocity no significant
noise is expected. The replacement of the current Crushing Plant with the SAG
Mill is considered to be an improvement of current noise emissions conforming to
the principle of noise exposure control by elimination and substitution of risk
at source, which in line with the RPM Occupational Health and Environmental
Standards.

Monitoring of noise levels has been conducted to establish a baseline for the
current mining operation. These monitoring results will be compared to noise
levels subsequently to completion of the expansion project in order to
demonstrate that noise levels and exposure remained similar.

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22.7.3 Closure Related Aspects

Since 1999, RPM has in place a Closure Plan that covers all areas affected by
its activities including the mine pit area, tailings dam and specific ponds that
are used to dispose of the sulphide concentrate. Accordingly to the plan
outlined above, RPM should implement specific measures to safely close down its
facilities as following:

a)   Mine Pit Area

Oxidized or non sulphide areas will be reclaimed by direct treatment of the
surface (grading) and growing of selected native species. Sulphide exposed areas
will receive a cover of overburden material and soil to prevent acid mine
generation. A drainage system for controlling run off water will be also be
implemented and maintained until the reclaimed areas become stable. A pit lake
will be formed in the west part of the mine. RPM performed specific
geohydrological studies (by Golder Consultants) to identify the potential risk
of impacting the groundwater flow and quality. These studies proved that the low
permeability of the local rocks and the adoption of specific control measures
such as the installation of a water treatment plant and the pumping of the
stored acid water back to the beneficiation plant can appropriately manage the
risk outlined above.

b)   Tailings Dam Areas

The closure of the tailings dam areas will initiate in the last 2 years of
operation. A low grade CTB1 ore will be mined with the purpose of installing a 1
m cover over the tailings disposed in the dam. This cover will considerably
reduce the risk of exposing any residual sulphide contained in the tailings
materials to atmospheric oxygen and hence of generating acid drainage. Actually,
the closure of the tailings dam pond will occur in a segmented manner with the
installation of dikes in specific branches of the tailings pond from the upper
elevations to the lower elevations. In order to ensure a higher safe

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coefficient, the current pond lake will be firstly moved upstream during the
operational stage and will be eventually eliminated in the final closure of the
dam. Peripheral canals will be constructed in both sides of the tailings pond in
order to allow the drainage of any stored water and drying of stored tailings.
An evaluation of the final water table is required to properly select the
species to grow in dry and wetlands. The tailings embankment slopes and
borrowing areas downstream the dam will be resloped and adequately rehabilitated
with the installation of drainage system over the benches and with the use of
native grass selected to prevent erosion.

c)   Specific Ponds (tanks)

The sulphur concentrate disposal facilities will require a special strategy for
closure in order to prevent oxygen income and hence acid mine drainage. Firstly,
a layer of oxide overburden material will be directly dumped and compacted over
the tailings. Following, a thin layer of coarse material (grinded limestone, for
instance) will be installed to break the capillary movement of salts to surface.
Finally, a local clay (red clay), that has proved to have a high adsorption
capacity of potential contaminants, will be installed on the top and selected
native grass planted. In order to reduce the seepage of water to lower levels,
which could accelerate the transportation of contaminants, and to prevent
erosion, an adequate drainage system to collect and divert run off water will
also be implemented. Periodical maintenance of surface drain systems and of
revegetation will be required for at least 5 years after closure of the specific
ponds to ensure a sustainable reclamation.

RPM has been performing reclamation surveys with UFV - Vicosa Federal University
to identify the potential species to be used in the revegetation works. An
onsite plant nursery has been maintained by RPM to generate seeds and saplings
to be used in the final revegetation of the site.

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d)   Buildings and Ancillary Facilities

Buildings and ancillary facilities such as roads, power lines and pumps should
be maintained and passed to local Paracatu authorities as long as they have
interest in using them for any beneficial purpose.

e)   Future Land Use and Sustainability of Closure

RPM wants to integrate its closure strategy with the local expectations,
although, taking also account of the restricted use of some areas such as the
sulphide mined areas, tailings dam and specific sumps. In this context, some
future potential land uses include the development of a park in the areas with
restricted use and use of the installations for educational and recreational
purposes. Buildings and ancillary facilities can also be used by the local
authorities for installing public services facilities (town hall, workshops,
ware stores, etc).

RPM has already conducted some surveys and discussions with the local community
to identify their perception and expectations in relation to the site closure.
Additionally, RPM has initiated a series of programs to address the social
impacts arising from the site closure. For instance, a local agency for
developing alternative sources of income has been created with support of the
company.

f)   Funding Closure

Kinross Gold Corporation has funded cash expenditures for reclamation and
closure costs out of cashflow from operations. Reclamation activities that are
performed concurrent during a mines operating life are funded from the mines
cashflow. Reclamation activities that are performed after a mine has completed
operations is funded from Kinross's portfolio of operating mines.

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Kinross accounts for reclamation obligations according to Canadian and US GAAP
under HB 3110 / FASB 143 Asset Retirement Obligations (ARO) reporting guidance.
For Paracatu, for example, this means that Kinross is reporting the net present
value (the discount rate is based on the cost of capital for Kinross) of the
estimated full cost of reclamation for all existing disturbance as a long-term
liability.

22.8 Taxes

Taxes on income include income tax (25% of EBITDA less depreciation of capital,
amortisation of goodwill and interest charges) and the social tax (9% of EBITDA
less depreciation of capital, amortisation of goodwill and interest charges).
Both income tax and social tax are assumed to be applied identically, with the
sole exception of the historic tax losses to be carried forward.

Note that tax is calculated in BRL terms, then converted to US$ based on the
assumed exchange rate.

In years where income before taxes (i.e., EBITDA less depreciation of capital,
amortisation of goodwill and interest charges) is negative, no taxes are charged
and the full amount of the loss is added to historic losses, which are assumed
to be carried forward indefinitely.

In years where income before taxes is positive, a maximum of 30% of historic
losses can be first applied to reduce the taxable amount. Any remaining positive
income is then taxed at the rates of 25% and 9% (aggregate = 34%).

There is no distinction between accounting taxes and cash taxes in Brazil - the
taxes calculated using the depreciation schedule represent the cash taxes paid.
There is no deferred tax liability on the Brazilian balance sheet.

Landowner's royalties are 0.5% of gold revenue for the portion of leases that
landowners have rights to (containing approximately 67% of recoverable gold).

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RPM pays ICMS taxes of 18% on the electricity price, but is able to sell ICMS
credits to third parties for typically 60% - 80% of the tax cost (i.e., 12% -
15%). The mid point of this range has been selected as the spread between taxes
paid and recoveries (4.5% = 18.0% - 13.5%).

CPMF (Tax on Financial Movement) applies to all movements from bank accounts,
which have been approximated by combining capital and operating costs.

Taxes that have been applied to revenues and costs are as tabulated in
Table 22.3 below.

                          Table 22-3 Paracatu Taxation

Taxes                      Description                                Value
------------------------   ---------------------------------------   ------
PIS + COFINS               % of silver revenue                         9.25%
ICMS                       % of silver revenue                        12.00%
ICMS                       % of transport costs                       10.62%
Landowners Royalties       % of gold revenue                           0.33%
State Royalties - gold     % of gold revenue                           1.00%
State Royalties - silver   % of silver revenue                         0.20%
Social Tax                 % of Earnings Before Tax                    9.00%
Income Tax                 % of Earnings Before Tax                   25.00%
Spread on ICMS for power   pay 18%, recover 13.5% from 3rd parties      4.5%
CPMF                       % of combined opex and capex                0.38%
Ad valorem with taxes      % of total revenue                        0.0401%

22.9 Capital and Operating Cost Estimates

The capital and operating costs estimates for the Paracatu Expansion Plan III
are documented in detail in the 2006 Feasibility Study.

The mine plan indicates a mine life of more than 30 years based on current
reserves. The project is anticipated to begin production in 2008 with expected
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557,000 oz of gold per year from 2009 through 2013 at an average cost of
$230/oz. The table below presents a summary of the Paracatu production data and
operating costs.

                 Table 22-4 Paracatu Production and Cost Summary
<TABLE>
<CAPTION>
                                                                        2009-2013        2009-2018       2009-2036
                                                                       --------------------------------------------
<S>                                     <C>                              <C>               <C>               <C>
Average throughput                      MM tpa                              58.4              51.2              40.9
Average Grade                           g/t                                 0.37              0.37              0.40
Average Recovery                        %                                  80.2%             80.0%             79.6%
Average Annual Gold                     oz pa                              556.7             489.8             418.1
Total Gold                              oz total                         2,783.6           4,898.5          11,392.4

Average Mining Cost                     $/tonne rock                      $ 0.47            $ 0.50            $ 0.69
Average Mining cost                     $/tonne ore                       $ 0.53            $ 0.64            $ 0.99
Average Milling cost                    $/tonne ore                       $ 1.50            $ 1.67            $ 1.96
Average G&A Cost                        $/tonne ore                       $ 0.18            $ 0.19            $ 0.22

Total Cash Costs                        $/oz                            $ 229.17          $ 257.97          $ 305.82
Total Production Costs                  $/oz                            $ 368.01          $ 356.61          $ 389.37

Initial Capital (2006 - 2009)           $ MM                            $ 469.53          $ 469.53          $ 469.53
Sustaining Capital for period           $ MM                            $  82.07          $ 141.33          $ 424.71 until 2037
Working Capital for period              $ MM                            $  (7.71)         $  (9.22)         $ (23.62)until 2037
Closure for period                      $ MM                            $      -          $      -          $  58.55 until 2037
Total capital for period                $ MM                            $ 543.89          $ 601.64          $ 929.17

sustaining 2006 - 2007/2008             $ MM                            $   40.7          $   40.7          $   40.7
working cap 2006 - 2007                 $ MM                            $   18.1          $   18.1          $   18.1
                                                                       ----------------------------------------------
total capital                                                            $602.68           $660.43           $987.96
</TABLE>
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                          Paracatu Mine Technical Report

                                      22-28

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The capital cost of the Expansion III Project is estimated to be US$470 million
as summarized in the table below.

          Table 22-5 CAPEX Breakdown for the 61 Mtpa Case @ R$ 2.3/US$

                              SUMMARY BY DISCIPLINE

<TABLE>
<CAPTION>
                                                                                                   TOTAL
                                                                                            -------------------
                                  Total        %     ALLOWANCE     GROWTH     contingency      Cost         %
                               -----------   -----   ---------   ----------   -----------   -----------   -----
<S>                            <C>           <C>     <C>         <C>           <C>          <C>           <C>
 DIRECTS
1  CIVIL                        10,289,165             424,635      633,900            --    11,347,700     3.9%
2  CONCRETE                     14,236,666             350,002      823,692                  15,410,360     5.3%
3  STRUCTURES                   22,774,269             454,357    1,297,082                  24,525,708     8.5%
4  ARQUITECTURE                  4,815,757              52,699      252,049                   5,120,505     1.8%
5  MECHANICAL                  153,645,429           2,243,197    9,007,015                 164,895,641    56.8%
6  PIPING                       20,453,060           2,558,210    1,176,598                  24,187,868     8.3%
7  ELECTRICAL                   31,068,962             678,958    2,161,014                  33,908,934    11.7%
8  INSTRUMENTATION               9,453,690             402,238      955,315                  10,811,243     3.7%
                                                     ---------   ----------                 -----------   -----
                                                     7,164,296   16,306,665                 290,207,959   100.0%
                                                     ---------   ----------                 -----------   -----
   10 Total                                                                                       100.0%
                                                                                            ===========
INDIRECTS
5  MECHANICAL                    7,447,982                          414,011                   7,861,993    10.5%
7  ELECTRICAL                      235,610                           14,066                     249,676     0.3%
8  INSTRUMENTATION               1,026,342                           61,273                   1,087,615     1.5%
9  INDIRECTS                    21,109,140           5,259,988      498,552    38,674,556    65,542,236    87.7%
                                                     ---------   ----------    ----------   -----------   -----
                                                     5,259,988      987,902    38,674,556    74,741,520   100.0%
                                                     ---------   ----------    ----------   -----------   -----
   20 Total                                                                          13.3%        100.0%
                                                                               ==========   ===========
                               Allowance+Growth+Contingency over
                                  (Directs + Indirects):                             18.7%
OWNER COST
   Owner's cost                106,578,428             528,532      510,703       439,600   108,057,263    65.6%
   Mine Fleet                                                                                41,293,953    25.1%
   Existing Dam                                                                               3,962,037     2.4%
   New Dam                                                                                   11,231,605     6.8%
   Waste Dump                                                                                   275,332     0.2%
                               106,578,428             528,532      510,703       439,600   164,820,190   100.0%
                               -----------           ---------   ----------    ----------   -----------
   30 Total                           64.7%                0.3%         0.3%          0.3%        100.0%
                               ===========           =========   ==========    ==========   ===========

   REDUCTION ON NEGOTIATIONS                                                                (15,297,915)
   Tax Reductions on
      Equipment                                                                             (44,941,580)
                                                                                            -----------
                                                                                    TOTAL   469,530,174
                                                                                            ===========
</TABLE>

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                          Paracatu Mine Technical Report


                                      22-29

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22.10 Economic Analysis

Discounted cash flow analyses have been completed for a 61 Mtpa throughput rate
using the Proven and Probable Reserves and the results demonstrate that the
project is viable and has a positive rate of return at gold prices greater than
US $400 per ounce, the price at which the reserves were estimated. The cash
flows are based on life of mine plans estimated by Kinross from the resource and
reserve model described in this Technical Report.

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

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                          Paracatu Mine Technical Report


                                      22-30

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23.0 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;

Rio Paracatu Mineracao S.A., 2006: Expansion Plan III Feasibility Study

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