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

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

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

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

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

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



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

This Current Report on Form 6-K, dated October 31, 2019 is specifically incorporated by reference into Kinross Gold Corporation’s Registration Statements on Form S-8 [Registration No. 333-217099, filed on April 3, 2017 and Registration Nos. 333-180824, 333-180823 and 333-180822, filed on April 19, 2012].

This report on Form 6-K is being furnished for the purpose of providing a copy of the press release dated October 31, 2019 in which the company announced the filing of a technical report for its Tasiast Project, Mauritania, and for providing a copy of the Technical Report along with the required certificate and consents as filed on SEDAR, dated October 31, 2019.

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

Toronto, Ontario – October 31, 2019 – Kinross Gold Corporation (TSX:K; NYSE: KGC) (“Kinross”) today filed an updated technical report for its Tasiast mine in Mauritania. The technical report incorporates updates related to the current Tasiast operation and provides comprehensive details regarding the recently announced Tasiast 24k project that is expected to incrementally increase throughput capacity to 24,000 tonnes per day, as well as the previously updated estimated mineral resource and reserve estimates at December 31, 2018, which were disclosed by news release on September 15, 2019.

The Tasiast 24k project takes a continuous improvement approach to increase throughput through minor upgrades and de-bottlenecking initiatives in the plant. The project includes modifications to the existing grinding circuit, adding new leaching and thickening capacity, as well as incremental additions to onsite power generation and water supply.

The technical report has been prepared pursuant to Canadian Securities Administrator's National Instrument 43-101, and may be found at www.kinross.com or under the Company's profile at www.sedar.com.

Kinross is a Canadian-based senior gold mining company with mines and projects in the United States, Brazil, Russia, Mauritania, Chile and Ghana. Kinross’ focus is on delivering value based on the core principles of operational excellence, balance sheet strength, disciplined growth and responsible mining. Kinross maintains listings on the Toronto Stock Exchange (symbol:K) and the New York Stock Exchange (symbol: KGC).

All statements, other than statements of historical fact, contained or incorporated by reference in this news release, including any information as to the future performance of Kinross, constitute "forward looking statements" within the meaning of applicable securities laws, including the provisions of the Securities Act (Ontario) and the provisions for "safe harbor" under the United States Private Securities Litigation Reform Act of 1995 and are based on expectations, estimates and projections as of the date of this news release. Forward looking statements include, without limitation, future events and opportunities including but not limited to the 24k project at Tasiast described in this news release. The words "estimate”, “expected” and “project”, and variations of such words identify forward-looking statements. Forward-looking statements are necessarily based upon a number of estimates and assumptions that, while considered reasonable by Kinross as of the date of such statements, are inherently subject to significant business, economic and competitive uncertainties and contingencies. The estimates and assumptions referenced, contained or incorporated by reference in this news release, which may prove to be incorrect, include, but are not limited to, the various assumptions set forth herein and in our Annual Information Form dated March 31, 2019 ("2018 AIF") and full year 2018, and Q1, Q2 and Q3 2019, Management's Discussion and Analysis . Known and unknown factors could cause actual results to differ materially from those projected in the forward-looking statements. In addition, there are risks and hazards associated with the business of mining and related project development, including financial, environmental hazards, government and labour relations, unusual or unexpected formations, pressures, cave-ins and flooding (and the risk of inadequate insurance, or the inability to obtain insurance, to cover such risks). Many of these uncertainties and contingencies can directly or indirectly affect, and could cause, Kinross' actual results to differ materially from those expressed or implied in any forward-looking statements made by, or on behalf of, Kinross. There can be no assurance that forward-looking statements will prove to be accurate, as actual results and future events could differ materially from those anticipated in such statements. Forward-looking statements are provided for the purpose of providing information about management's expectations and plans relating to the future. All of the forward-looking statements made in this news release are qualified by these cautionary statements and those made in our other news release dated September 15, 2019 regarding the 24k study results as well as other filings with the securities regulators of Canada and the United States including, but not limited to, the cautionary statements made in the "Risk Factors" section of our 2018 AIF. These factors are not intended to represent a complete list of the factors that could affect Kinross. Kinross disclaims any intention or obligation to update or revise any forward-looking statements or to explain any material difference between subsequent actual events and such forward-looking statements, except to the extent required by applicable law.

Kinross has prepared a Technical Report for the wholly-owned Tasiast mine (the Project) located in the Islamic Republic of Mauritania (Mauritania), Africa. Kinross is using this Technical Report to support disclosure of mineral resources and mineral reserves at the Project. The Technical Report conforms to National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) and has an effective date of October 31, 2019. Kinross will be using this Technical Report to support disclosure of mineral resources and mineral reserves at the Project.

In September 2019, Kinross completed a feasibility study to incrementally increase throughput capacity at Tasiast from approximately 15,000 t/d throughput to 24,000 t/d. The project is expected to ramp up to 21,000 t/d by the end of 2021, and then to 24,000 t/d by mid-2023. Throughput increases are expected to be achieved through minor upgrades and de-bottlenecking initiatives in the plant. The project includes modifications to the existing grinding circuit, adding new leaching and thickening capacity, as well as incremental additions to onsite power generation and water supply.

The Tasiast mine and the mining permit are owned by Tasiast Mauritanie Limited S.A. (TMLSA). SENISA (Société d’Extraction du Nord de l’Inchiri S.A., a sister company of TMLSA) holds two mining permits (for the Tmeimichat and Imkebdene areas). The two mining permits are contiguous with the Tasiast mining permit land (collectively, the Tasiast Lands). As part of the conversion process, Kinross has undertaken to transfer to the Government of Mauritania a 10% carried interest in SENISA. Kinross acquired TMLSA, including the Tasiast Lands, through its acquisition of Red Back in September 2010. There are exploration prospects in the 312 km2 Guelb El Ghaîcha Mining permit and in the surrounding permits.

Commercial production of gold at Tasiast began in January 2008, and a total of 2,288 koz. was produced by the end of 2018.

The Project is located in northwestern Mauritania, approximately 300 km north of the capital Nouakchott and 250 km southeast of the major city of Nouadhibou. The Tasiast Lands are accessed from Nouakchott by using the paved Nouakchott to Nouadhibou highway for 370 km and then via 66 km of graded mine access road, which is maintained by TMLSA. There is an airstrip at the mine site that is used for light aircraft travelling to and from Nouakchott.

Mining operations commenced in 2007, with commercial production reached in January 2008. Infrastructure on site supports an open pit mining operation and associated processing facilities consisting of a carbon-in-leach (CIL) mill and a run of mine (ROM) dump leach.

TMLSA holds a valid mining permit, 229C2 (Guelb El Ghaîcha), covering 312 km2 granted in January 2004 and valid for a period of 30 years. The mining operations and infrastructure lie entirely within the lands subject to the mining permit. There are also two additional contiguous permits (1,285 km2), each of which is in good standing. The Tasiast Lands fall within the Inchiri and Dakhlet Nouadhibou Districts, with 229C2 within the Inchiri District only.

Surface rights are granted along with permit 229C2 and are paid annually as determined by decree under the Mauritanian Mining Code. Surface rights for the permit are in good standing.

The operation’s water supply is located 64 km west of the mine and consists of a bore field of 43 wells in a semi-saline aquifer. Water is pumped from the bore field to the mine. The Tasiast permit, issued May 7, 2017 by the Ministry of Hydraulics and Sanitation, allows abstraction at a maximum rate of 30,000 m3/d through to December 31st, 2034.

A royalty equal to 3% of the gross revenue of TMLSA is payable to the government. In addition, Franco-Nevada Corporation (Franco-Nevada) holds a 2% net smelter return royalty on gold production in excess of 600,000 ounces. Production at Tasiast reached 600,000 ounces in July 2011 and the first royalty payment to Franco-Nevada was made in October 2011.

Exploration, development and mining activities to date have been performed under the appropriate permits, laws and regulations.

Current mine operations and the expansion project are based on the formal approval of a number of Environmental Impact Assessment (EIA) studies completed before and since mine commissioning in 2007.

A review of waste rock geochemistry to determine the potential for acid rock drainage concluded that the rock has excess neutralizing capacity. Given the excess neutralizing capacity and the very low precipitation at Tasiast, acid rock drainage is not anticipated.

The Tasiast facilities operate under an environmental management system (EMS) that specifies activities to be planned and implemented by the mine’s environmental management team. The EMS incorporates the project design and management, mitigation strategies and performance monitoring commitments outlined in the environmental assessments, applicable legislation and specific permit requirements.

An element of each EIA prepared for the Tasiast mine site is a preliminary reclamation and closure plan and associated cost estimate. The preliminary reclamation and closure plan outlines the measures that will be taken to reclaim and close the proposed activities assessed in each EIA. The preliminary reclamation and closure cost estimate forms the basis of the financial assurance. The current financial assurance for the existing operation is approximately $6.2 million. In 2016, Tasiast submitted an updated closure estimate of $37 million which included the 12 kt/d expansion. With the proposed 24 kt/d expansion the estimated closure costs increase to $45.4 million. The Government of Mauritania has requested an independent review of the closure estimate. Upon completion of the independent review, Tasiast will put additional financial assurance in place. At least two years before entering closure, a detailed reclamation and closure plan must be submitted to the appropriate ministries for approval.

Current environmental liabilities are those that would be expected from a mining operation, and include the mine, crushing and CIL processing plant, dump leach facilities, power plant, tailings and waste rock facilities, power grids, roads, accommodation camp, ancillary facilities and drill pads established to support mining and exploration activities.

The Tasiast district is situated in the south-western corner of the Reguibat Shield, which is a northeast-trending crustal block of the West African Craton. The Reguibat Shield contains the oldest rocks in Mauritania and consists of two major subdivisions separated by a crustal-scale shear zone representing a major accretionary boundary. The south-western part (which hosts the Tasiast deposits) consists of Mesoarchean to Paleoproterozoic rocks that include high-grade granite-gneiss and greenstone belt assemblages. The north-eastern part of the shield consists of younger Paleoproterozoic to Neoproterozoic successions, which hosts many orogenic gold occurrences in the West African Craton. This region is characterized by a series of volcanosedimentary belts and associated batholithic-scale granitic intrusive suites of different ages cut by major shear zones.

The district scale geology is characterized by basement rocks, largely composed of orthogneiss, overlain by deformed north-striking metavolcanic and metasedimentary successions intruded by stocks and plutons of mafic to intermediate composition (granite-greenstone belts). All of the rock units are cut by unfoliated and post-mineral mafic (gabbroic) dikes.

The Tasiast Mine consists of two deposits hosted within distinctly different rock types, both situated within the hanging wall of the Tasiast thrust. The Piment deposits are hosted within metasedimentary rocks including metaturbidites and banded iron formation. The West Branch geology succession comprises mafic to felsic volcanic sequences, iron-rich formations and clastic units that have undergone mid greenschist to lower amphibolite facies metamorphism and multiple deformation events.

The Tasiast gold deposits fall into the broad category of orogenic gold deposits. The regional geological setting and deposit features at Tasiast are similar to other well-known Archaean lode gold deposits hosted along greenstone belts in granitoid-greenstone terranes.

Exploration programs have included geological and regolith mapping, satellite image interpretation, airborne and ground magnetic geophysical surveys, soil, rock chip, and grab geochemical sampling, trenching, reverse circulation (RC) and core drilling, engineering studies, metallurgical test work, and specialist geological studies such as ore and alteration petrography. Work was completed by the Office Mauritanien de Recherches Géologiques (OMRG), Normandy LaSource Development Ltd. (NLSD), Midas Gold plc. (Midas), Geomaque Explorations Inc. (Geomaque), Defiance Mining Corporation (Defiance), Rio Narcea Gold Mines Ltd. (Rio Narcea), Red Back and Kinross.

The Tasiast project area has significant exploration potential to delineate additional resources both around the Tasiast mine (near mine exploration) and within the wider district (generative exploration).

Exploration efforts to date have discovered additional prospects, gold deposits and mineral resources along strike to the North and to the South of the main Tasiast mine area (West Branch and Piment-Prolongation), and generally along the Aouéouat (Tasiast) belt. The deformed greenstone rocks to the west (Imkebdene-Kneiffissat) of the Aouéouat belt are notable in that they host quartz-carbonate veins with anomalous gold values, however to date no significant deposits or mineral resource have been defined. To the immediate north of the Tasiast operation (5-12 km) and within the Guelb El Ghaîcha license, a cluster of deposits referred to as “North Mine Satellites” have been outlined, these are Fennec, C67 and C68. These gold deposits currently host approximately 0.5 Moz Au and are part of the near-mine resource growth strategy. Further north of the Tasiast operation (12-25 km) and within the Imkebdene and Tmeimichat licenses are another cluster of gold deposits referred to simply as “Morris”, these are Tef, Askaf, Central, NE, N1 and N2. This large area saw extensive exploration drilling from 2012 to 2014, which resulted in the discovery of several small deposits best described as narrow, high grade vein systems. Most of these deposits are open to depth down plunge.

Beyond 25 km from the Tasiast operation, within the northern extents of the Imkebdene and Tmeimichat licenses, are several gold exploration prospects that are pending follow-up exploration and drilling, of note are; C23, Kneiffissat and Grindstone. These prospects have significant surface geochemical footprints and are considered highly prospective.

To date, 15,360 drill holes (14,286 RC, 840 diamond core and 234 RC-DD) for an aggregate total of 1,683,635 meters have been completed within the three mining licenses that constitute the Tasiast project area; Guelb El Ghaîcha, Imkebdene and Tmeimichat. Drilling activities were conducted by various drilling contractors and supervised by geological staff of the Project operator.

The Tasiast mineral resource statement, as of year-end 2018, comprises Measured, Indicated and Inferred resources (Table 1-1). Mineral resources were classified in accordance with the 2014 CIM Definition Standards for Mineral Resources and Mineral Reserves, incorporated by reference into NI 43-101. Mineral resources have an effective date of December 31, 2018.

Mineral Resources are stated at variable cut-off grades, dependent on the metallurgical type, mining operating cost and variable process recoveries. The cut off grades were determined using a gold price of $1,400/oz.

The mineral resources were reported below the projected December 31, 2018 mined surface and are constrained using the Lerchs-Grossman (LG) 24 kt/d pit shell estimated by Kinross Technical Services. Kinross cautions that mineral resources that are not mineral reserves do not have demonstrated economic viability.

Mineral reserves for the Project incorporate appropriate allowances for mining dilution and mining recovery for the selected mining method. Mineral reserves have an effective date of December 31, 2018 and are summarized in Table 1-2.

Ore and waste rock is mined in 10 m benches by conventional open pit methods primarily from the West Branch pit. Tasiast currently operates a load and haul fleet of 46 Cat 793D (220 t) trucks, five Komatsu 785 (92 t) and five Cat 6060 shovels plus two RH340B excavators. Blasting techniques, including presplit and buffer hole blasting, are employed to protect the pit walls. The grinding circuit produces a product size of 80% passing 90 microns which is processed in a conventional CIL circuit to produce gold bullion. Gold recovery averages 93%. Tailings slurry from the CIL process is currently pumped to the tailings storage facility 4 (TSF4).

The existing Tasiast CIL plant has proven capable of processing approximately 15 kt/d. Kinross has decided to increase the existing CIL plant capacity in stages through a de-bottlenecking exercise. The modifications to achieve and initial target of 21 kt/d and later 24 kt/d form the basis of this project. The project includes modifications to the existing grinding circuit, adding new leaching and thickening capacity, as well as incremental additions to onsite power generation and water supply.

The project schedule is designed to incrementally alleviate each bottleneck in order of priority to first achieve 21 kt/d, and then advance to 24 kt/d. Commissioning will begin during the final stages of construction in each area. As the construction of a section of the plant is completed, the section is handed over to commissioning. Having both construction and commissioning underway at the same time minimizes the delay between final construction completion and the start-up of the plant. Each project area will have a detailed commissioning and tie-in plan prepared to minimize operational downtime.

Kinross typically establishes refining agreements with third-parties for refining of doré. Kinross’s bullion is sold on the spot market, by marketing experts retained in-house by Kinross. The terms contained within the sales contracts are typical and consistent with standard industry practice, and are similar to contracts for the supply of doré elsewhere in the world.

The total going forward capital cost estimate is $150 million. The Tasiast debottlenecking project assumes the total capital costs will be distributed over four years, approximately $85M of which will be spent through 2020, decreasing to $30M, $20M and $15M per year in 2021 to 2023, respectively.

Operating cost estimates are shown in Table 1-3. The operating costs for each area include allocations for power plant operating costs.

3. No additional tonnes placed on dump leach 2020+. Dump leach decommission planned completion by 2022.

The economics of the Tasiast Debottlenecking Project were evaluated using a real (non-escalated), after-tax discounted cash flow (DCF) model on a 100% project equity (unlevered) basis. Unless otherwise stated, all economic parameters are shown on an absolute basis (not incremental to existing operations). Production, revenues, operating costs, capital costs and taxes were considered in the financial model. The main economic assumptions are a US$1,200/oz gold price and a 5% discount rate.

The valuation date for the financial analysis was set for January 1, 2020. All cash flows assumed for the purposes of this study are from this date onward.

The results of the financial analysis, with sensitivities to gold price and discount rate assumptions, are shown in Table 1-4, based on $1,200/oz gold, a real discount rate of 5%, and an oil price of $55/bbl.

Tasiast is viewed as a long-term strategic asset for Kinross, located in a district that is believed to have significant future potential. The phased expansion project is believed to provide an opportunity to capitalize on the full potential of the operation and to solidify Tasiast as a low cost, long life asset within the company’s portfolio.

The project economics, as stated at a 5% discount rate and a $1,200 base case gold price, are robust and offer significant potential.

It is recommended that Kinross proceed with the project to incrementally increase throughput capacity at Tasiast from approximately 15,000 t/d throughput to 24,000 t/d.

Kinross Gold Corporation (Kinross) has prepared a Technical Report for the wholly-owned Tasiast gold deposit (the Project) located in the Islamic Republic of Mauritania (Mauritania), Africa (Figure 2-1). Kinross is using this Technical Report to support disclosure of mineral resources and mineral reserves at the Project. The Technical Report conforms to National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101) and has an effective date of October 31, 2019.

In September 2019, Kinross completed a feasibility study to incrementally increase throughput capacity at Tasiast from approximately 15,000 t/d throughput to 24,000 t/d. The project is expected to ramp up to 21,000 t/d by the end of 2021, and then to 24,000 t/d by mid-2023.

Throughput increases are expected to be achieved through minor upgrades and de-bottlenecking initiatives in the plant. The project includes modifications to the existing grinding circuit, adding new leaching and thickening capacity, as well as incremental additions to onsite power generation and water supply.

All measurement units used in this Technical Report are metric, and currency is expressed in US dollars unless stated otherwise. Mauritania uses the Ouguiya (MRU) as its currency.

Information used to support this Technical Report has been derived from the reports and documents listed in the References section of this Technical Report.

The use of the terms “we”, “us”, “our” or “Kinross” in this Technical Report refer to Kinross Gold Corporation.

The Qualified Person (QP) for this Technical Report is John Sims, AIPG Certified Professional Geologist and Vice President, Technical Services for Kinross.

Mr. Sims visited the site most recently in June 2018. During the site visit, Mr. Sims inspected core and surface outcrops, drill platforms and sample cutting and logging areas; discussed geology and mineralization with Project staff; reviewed geological interpretations with staff; and inspected the major infrastructure and current mining operations. There have been no material changes in site conditions since Mr. Sims’ most recent site visit. All sections in this Technical Report have been prepared under the supervision of Mr. Sims.

Mineral Resources: The mineral resource estimates included in this report were prepared under the supervision of Racquel Kolkert, Director, Resource Geology, Kinross Technical Services. Ms. Kolkert is a Chartered Professional Member of the Australasian Institute of Mining and Metallurgy. Ms. Kolkert visited the site most recently in July 2018.

Mineral Reserves / Mining: The mineral reserve estimate and economic analysis included in this report was prepared under the supervision of Scott Hicks, Vice President, Mine Planning and Technical Evaluations, Kinross Technical Services. Mr. Hicks is a Registered Professional Engineer in the Province of Ontario, and a Chartered Financial Analyst. Mr. Hicks visited the site most recently in July 2019.

Mineral Processing: Mineral processing aspects of this report were prepared under the supervision of Yves Breau, Vice President, Metallurgy and Engineering, Kinross Technical Services. Mr. Breau is a Registered Professional Engineer in the Province of Ontario. Mr. Breau visited the site most recently in October 2019.

Information used to support this Technical Report was derived from previous technical reports on the property, and from the reports and documents listed in the References section of this Technical Report.

The effective date of this Technical Report is October 31, 2019, and for the Tasiast expansion mineral resources and mineral reserves the effective date is December 31, 2018.

There were no material changes to the information on the Project between the effective date and the signature date of the Technical Report.

In the preparation of the Technical Report, the Qualified Person relied on information provided by internal Kinross legal counsel for the discussion of legal matters in Sections 4, 19, and 20.

Except for the purposes legislated under provincial securities law, any other use of this report by any third parties is at this party’s sole risk.

The Tasiast Lands are located in northwestern Mauritania, approximately 300 km north of the capital Nouakchott and 250 km southeast of the major city of Nouadhibou. The Tasiast Lands fall within the Inchiri and Dakhlet Nouadhibou Districts. The Tasiast mine is located at 446600E, 2275600N (UTM, WGS84, Zone 28N).

The Tasiast mine is owned and operated by Tasiast Mauritanie Ltd. S.A. (TMLSA), a wholly owned subsidiary of Kinross, under exploitation Permit No. 229C2.

The Tasiast mine is located within the 312 km2 Tasiast exploitation permit of Guelb El Ghaîcha. The mining operations and infrastructure are located entirely within the lands subject to the mining/exploitation permit (permis d’exploitation or PE). Exploitation permit No. 229C2 is located centrally within a surrounding permit block of two contiguous exploitation permits, totalling 1,597 km2, as listed in Table 4-1 and shown in Figure 4-1. All these permits are in good standing. The table also indicates tenure expiry dates. The Tasiast mine and the exploitation permit are owned by TMLSA.

The adjacent two permits, the underlying lands of which are contiguous to the Tasiast exploitation permit lands, are held by a sister company of TMLSA, Société d’Extraction du Nord de l’Inchiri S.A. (SENISA). SENISA holds two mining permits (for the Tmeimichat and Imkebdene areas) As part of the conversion process of two exploration permits, Kinross has undertaken to transfer to the Government of Mauritania a 10% carried interest in SENISA. Kinross acquired TMLSA, including the Tasiast operation and exploration permits and lands, through its acquisition of Red Back Mining Inc. (Red Back) in September 2010.

Tenure coordinates are shown in Table 4-2. A permit boundary is defined by a list of the coordinates of its corners or pillar points. The boundaries are not physically marked on the ground, and have not been surveyed. However, extensive surveying has been conducted within both the exploitation permit No. 229C2 and adjoining permits. All the known gold deposits are well inside the boundaries, and the size and shape of the exploitation permit are adequate for the intended exploration, mining and processing activities.

Surface rights are granted along with Permit No. 229C2, and applicable fees are paid annually, as determined by decree under the Mining Code. Surface rights for the permit are in good standing, and there are no competing mining rights in the area.

Exploration permits (Permis de Recherche Minière or PRM) grant exclusive exploration rights over a specific block (maximum 1,000 km2) and are granted for a three-year period, renewable twice for up to three years at each renewal. Exploitation permits are granted for 30 years, and are renewable for periods of 10 years each. A condition of each permit is that the holder is required to hire Mauritanian tradespersons to provide services, and to contract with national suppliers and businesses in preference to foreign service providers, where the national suppliers and businesses can offer at least the same terms, quality and pricing. Table 4-3 summarizes the durations of exploration and mining permits in Mauritania. Operating permits are discussed in Section 20.2.

Current environmental liabilities are those that would be expected from a mining operation, and include the mine, crushing and CIL processing plant, dump leach facilities, power plant, tailings and waste rock facilities, power grids, roads, accommodation camp, ancillary facilities and drill pads established to support mining and exploration activities.

After the two renewal periods lapse, the permit expires unless it is converted (in whole or in part) into an exploitation permit.

Mining activities in Mauritania are mainly governed by the Mining Code and its regulations, and by the Model Mining Convention Law, which provides the legal and tax framework for all mineral exploration and extraction activities. TMLSA is governed by the 1999 Mining Code.

The Mining Code establishes conditions and rules governing all phases of mining activity. The Model Mining Convention Law provides that each exploration permit is subject to a mining convention with the State of Mauritania, which outlines the framework of customs, economic, financial, legal and tax terms and conditions under which the permit holder proceeds with its exploration or mining activities inside the perimeter of its permit. A Mining Convention is attached to a given permit. Table 4-4 summarizes provisions of the TMLSA Mining Convention relating to fees, royalties, duties and taxes.

The Mining Code is also complemented by the Decree on Mining Titles, which provides more details on the process governing the grant, renewal, expansion or reduction, division or merger, transfer, termination, suspension and cancellation of a permit for exploration or exploitation. It also governs the conversion of an exploration permit into an exploitation permit.

The conditions embodied in the Model Mining Convention (Law No. 2002/02) subsequently replaced by Law 2012-012 are designed to stimulate and encourage investment in both exploration and mining. The mining industry is seen as one of the main growth industries for the improvement of the country’s economy.

In addition to the 3% royalty payable to the government, Franco-Nevada Corporation holds a 2% net smelter return royalty on gold production at the Tasiast mine in excess of 600,000 cumulative ounces produced. Production at the Tasiast mine reached 600,000 ounces in July 2011, and the first royalty payment to Franco-Nevada was made in October 2011.

1 Kinross has been in discussions with the Government of Mauritania regarding the applicability of the fuel exemption to contractors and since January 2018 the Government has not approved the exemption.  For the purposes of this Technical Report only, it is assumed that the exemption will apply from 2021 through the end of mine life.

The Tasiast Lands are accessed from Nouakchott by using the paved Nouakchott to Nouadhibou highway for 370 km and then via 66 km of graded mine access road which is maintained by TMLSA. An airstrip at the Mine Site is used for light aircraft primarily travelling to and from Nouakchott.

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

Routine access within the country is provided by an 11,000 km long road network, comprising approximately 3,000 km of paved highways and approximately 8,000 km of unpaved highways as well as numerous desert tracks. A paved 470 km long, two-lane highway runs between the cities of Nouakchott and Nouadhibou.

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

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

Mauritania has an arid desert climate, with an average annual high temperature of above 45°C between May and August. Minimum temperatures may go below 10°C in December and January. From January to March, sandstorms frequently occur in the country; this causes sand build up and dune formation. Sandstorms vary in intensity, and visibility can be reduced to several metres. There is a rainy season, usually between July and September; however, the amount of rainfall and length of season varies spatially and temporally in the various regions of the country. Annual rainfall varies from a few millimetres in the desert regions to as high as 450 mm in the south along the Senegal River.

Average annual precipitation at the mine site is approximately 90 mm, and usually occurs from July to September. The average recorded monthly evaporation is approximately 320 mm/month (3840 mm/a).

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

The mine is located in a remote area where there is no electric power grid. On-site power generation is discussed in Section 18.

The source of mine water supply is located 64 km west of the mine and consists of a semi-saline underground aquifer, with 44 wells for water production. Water is pumped from the bore field to the mine (see Section 18).

In 2019, the Tasiast mine employs approximately 1,285 employees, of whom approximately 1,220 are Mauritanian nationals. Staff accommodation is provided at the mine site (see Section 18).

The terrain surrounding the Tasiast deposit is flat, and is adequate for construction and operation of the camp, mine, plant, tailings, and waste rock disposal facilities.

The topography of the Tasiast Lands consists mainly of flat, barren plains which are primarily covered by regolith and locally by sand dunes, or eroded paleo-lateritic profiles. Elevation ranges from approximately 130 masl to 150 masl.

The drainage pattern around the Tasiast Lands consists of several intermittent dendritic first- and second-order streams that generally flow in a southwesterly direction. There are no permanent watercourses in the area. However, there are numerous, intermittent watercourses, known as “wadis”, which flow for only a few days per year. The largest wadi is the Khatt Ataoui wadi, which is located approximately 6 km from the mine site.

The Tasiast mine is located in the arid Saharan zone, where plant life is very scarce, consisting mainly of the low shrubs Zygophyllum album, the small tree Maerua crassifolia (atil) and the grass Aristida pungens (sbot). Acacias are also present along many of the wadis.

Hares, hamsters and gerbils are the most common mammals at the mine site, and jackals, fennec fox and polecat can also be found in the well area. There are no protected species in the Project area. The eastern boundary of the Banc D’Arguin National Park is located about 2 km west of the bore field area and 60 km from the mine site.

In 1996, the Office Mauritanien de Recherches Géologiques (OMRG) completed a regional reconnaissance exploration program within and around the lands hosting the Tasiast deposit and made this information available to third parties. As a result, NLSD (a subsidiary of Normandy Mining Ltd. of Australia) acquired the exploration rights to the Tasiast deposit.

In 2001, NLSD was acquired by Newmont Mining Corporation creating Newmont LaSource. Midas Gold PLC (Midas) was incorporated in England and Wales in 2002 for the purpose of acquiring Newmont LaSource’s assets in Mauritania including exploration permits over lands hosting the Tasiast deposit, as well as various other permit areas. Midas completed its acquisition of the Tasiast deposit from Newmont LaSource on April 1, 2003 and, in April 2003, Geomaque Explorations Inc. (Geomaque) announced the acquisition of Midas. The merger of Geomaque and Midas ultimately created a new entity - Defiance Mining Corporation (Defiance). In June 2004, Rio Narcea Gold Mines, Ltd. (Rio Narcea) acquired Defiance and took ownership of the Tasiast deposit.

Red Back acquired the Tasiast deposit from Lundin Mining Corporation (Lundin) in August, 2007 following Lundin’s acquisition of Rio Narcea. In September 2010, Kinross completed the acquisition of Red Back. Kinross, through TMLSA, holds 100% of the Project.

From 1962 to 1993, the Tasiast region was the subject of three regional exploration programs for pegmatites, iron ore, and nickel sulphides which were carried out by the BRGM (Bureau de Recherches Géologiques et Minières) and SNIM.

Three exploration programs were carried out in the Tasiast region between 1993 and 1996 as a European Development Fund project. Work completed included regional-scale reconnaissance geological mapping and geochemical sampling. Traverse lines for the mapping and geochemical sampling programs were oriented east-west with samples collected at 500 m centres; this work identified the Tasiast area as being anomalous in gold. More detailed soil sampling of the Tasiast area on 250 m spaced centres, and trenching was conducted.

NLSD, in the period 1996-2001 completed geological and regolith mapping, interpretation of satellite imagery, airborne and ground magnetic geophysical surveys, specialist petrographical, mineralogical, and geological studies, metallurgical test work, and auger, reverse circulation (RC) and core drilling.

Midas undertook a full review of all existing information in 2003, and prepared mineral resource estimates for the West Branch and Piment areas. From 2003 to 2004, Defiance completed mineralogical and metallurgical test work, hydrogeological studies, a preliminary pit slope design study, RC and core drilling, a mineral resource estimate, and a feasibility study.

Rio Narcea completed additional RC and core drilling from 2005-2006. Red Back also undertook RC and core drilling, re-estimated mineral resources, and updated engineering studies. Mine construction was completed in 2007, with the mine formally opened in early 2008.

Between August 2007 and September 2010 Red Back completed several large exploration campaigns in the Piment and West Branch areas, as well as at several district targets. Early drilling campaigns were directed at testing the lateral and vertical extents of the mineralization at Piment and drilling oxide resources at West Branch. In October 2009, Red Back discovered the Greenschist Zone at West Branch and commenced drilling the deposit.

From September 2010 to date, TMLSA has aggressively ramped up exploration with the majority of activities directed towards delineating the extents of the Greenschist Zone.

Mining at Tasiast commenced in April 2007 and the mine was officially opened by the President of Mauritania, His Excellency Sidi Mohamed Ould Cheikh Abdallahi, on July 18, 2007. A summary of gold production at Tasiast is included in Table 6-1. There has been no historical gold production from other deposits in the Tasiast area.

The Tasiast district is situated in the south-western corner of the Reguibat Shield, which is a NE-trending crustal block of the West African Craton (Figure 7-1). The Reguibat Shield contains the oldest rocks in Mauritania and consists of two major subdivisions separated by a crustal-scale shear zone representing a major accretionary boundary (Lahondère et al., 2003; Pitfield et al., 2004; Schofield et al., 2006). The south-western part (which hosts the Tasiast deposits) consists of Mesoarchean to Paleoproterozoic rocks that include high-grade granite-gneiss and greenstone belt assemblages. The north-eastern part of the shield consists of younger Paleoproterozoic to Neoproterozoic successions, which hosts many orogenic gold occurrences in the West African Craton. This region is characterized by a series of volcanosedimentary belts and associated batholithic-scale granitic intrusive suites of different ages cut by major shear zones.

The Reguibat Shield is bound on all sides by Pan African orogenic belts and covered in the south by the extensive intra-cratonic sediments of the Taoudeni Basin. The Taoudeni basin represents one of the largest Mesoproterozoic to Paleozoic cratonic sedimentary basins in Africa. It consists of many thousands of metres of continental sandstones, platform carbonate rocks and lesser shales.

The district scale geology is characterized by basement rocks, largely composed of orthogneiss, overlain by deformed north-striking metavolcanic and metasedimentary successions intruded by stocks and plutons of mafic to intermediate composition (granite-greenstone belts). All of the rock units are cut by unfoliated and post-mineral mafic (gabbroic) dikes. Two significant Archaean greenstone belts are exposed within the Tasiast District (Figure 7-2):

The greenstone belts are wholly enclosed by granitic to gabbroic intrusive rocks and gneissic domes that comprise the bulk of the rocks within the district and the Reguibat Shield overall. The greenstone belts comprise ultramafic to felsic volcanic and volcanosedimentary packages with variably preserved ferruginous quartzite, locally termed banded magnetite. Rock units within the belts have undergone mid greenschist to lower amphibolite facies metamorphism and multiple deformation events. Swarms of non-foliated mafic (basaltic) dykes striking north-northeast/south-southwest and more or less east-west crosscut all other rocks in the district, including undeformed pegmatite units.

A Precambrian lithostratigraphy was established by Kinross for the Aouéouat belt (Figure 7-3) including several U-Pb dates for rocks of the Aouéouat and Tasiast assemblages and for granodiorite intrusions. The mafic to felsic volcanic and intrusive units that host the West Branch deposit belong to the Aouéouat assemblage that crystallized between 2,990 Ma and 3,000 Ma. Metasedimentary rocks of the Tasiast assemblage that overlay the mineralized West Branch units contain detrital zircons of similar ages and older populations derived from approximate 3,200 Ma orthogneiss basement. Granodiorites that crosscut the metavolcanic rocks are dated 2,960 Ma to 2,970 Ma. An age of 2839 ± 36 Ma obtained from the hydrothermal overgrowth on zircons from a quartz vein at the Fennec deposit is interpreted to represent the age of mineralization at Tasiast (Heron, et al., 2016).

The principal north-south structural fabric in the Tasiast granite-greenstone belts is evident in satellite images (Worldview-2), airborne geophysics and regional geological maps. Steeply dipping foliations and isoclinals folds with north-south to northwest-southeast axial surface traces are common across the Aouéouat belt. Those structures formed through east-west transpressive shortening that occurred as a result of basin inversion. Strain was partitioned between tightly folded domains and north-south striking shear zones. Several families of late-stage cross-cutting faults with northeast and southwest strikes are occupied by fresh mafic dyke material.

All of the significant mineralized bodies defined to date dip moderately to steeply (45° to 70°) to the east and have a south–southeasterly plunge. Gold deposits on the Tasiast trend are associated with second order shear zones and splays cutting the hanging wall block of in inferred thrust. The volcano-sedimentary stratigraphy has been tightly to isoclinally folded, and is cut longitudinally by sub-parallel shears that are sub-parallel to the predominant foliation.

The main Tasiast gold trend includes the West Branch and Piment-Prolongation deposits. Gold mineralization is spatially associated with the west vergent Tasiast shear system that places mafic to felsic volcanic and intrusive rocks of the Aouéouat assemblage, including the host rocks of the West Branch deposit, on top of the younger metasedimentary rocks of the Tasiast assemblage. The Tasiast trend passes north-south through the Guelb El Ghaîcha mining permit and extends to the north and south onto adjacent licences.

The Tasiast Mine consists of two deposits hosted within distinctly different rock types, both situated within the hanging wall of the west-vergent Tasiast thrust (Figure 7-4 and Figure 7-5).

The Piment deposits are hosted within metasedimentary rocks including metaturbidites and banded iron formation where the main mineral association consists of magnetite-quartz pyrrhotite ± actinolite ± garnet ± biotite. Gold is associated with silica flooding and sulphide replacement of magnetite in the turbidites and in the banded iron formation units.

The West Branch geology succession comprises mafic to felsic volcanic sequences, iron-rich formations and clastic units that have undergone mid greenschist to lower amphibolite facies metamorphism and multiple deformation events.

a) Biotite rich, 2-5% sulphides (pyrrhotite>>pyrite) intense quartz- carbonate veining (5-10%) folded and boudinaged => BST zone; b) actinolite+ garnet >biotite => GST 1 zone; c) biotite rich => BST

Most of the gold mineralization at West Branch is hosted by quartz–carbonate veins within the sheared and hydrothermally altered meta-diorites that constitute the Greenschist Zone. All of the significant mineralized bodies defined to date dip moderately to steeply (45° to 70°) to the east and have a south–southeasterly plunge (Figure 7-6 through Figure 7-9).

The majority of the economic mineralization at West Branch is hosted by a diorite to quartz diorite intrusion (termed the GDI) which has intruded into FVC or felsite (clastic sediments). The GDI is light to medium grey, medium to fine-grained and composed of plagioclase, quartz and biotite with some potassium feldspar (up to half of the plagioclase content). It typically shows a zonation from a barren garnet-amphibole assemblage at its margins to an auriferous quartz-biotite-ankerite-pyrite-pyrrhotite assemblage and back into the barren garnet-amphibole assemblage. This unit is also characterized by a distinctive penetrative foliation that is most strongly expressed by the alignment of biotite crystals. It was previously referred to as plagioclase-biotite schist and was logged as schist (‘SHT’) and biotite schist (‘BST’). At West Branch, the GDI consistently averages 50 m to 100 m in thickness over a strike length that exceeds 2 km.

The FVC unit is a clastic sedimentary sequence of predominantly quartzite with minor locally polymict felsic conglomerate layers. Previously this unit had been interpreted to be of volcanic origin and hence its confusing name (FVC = Felsic Volcanic). At West Branch, the FVC unit is present both structurally above and below the GDI. The unit is intensely sheared and preserves a well-developed phyllosilicate foliation. Commonly this unit is albitized and is called ‘felsite’, which is not a rock type. Within the FVC, near to its contacts with other units, a cream coloured rock locally occurs that hosts fuchsitic (chromium rich) mica.

The rocks on the eastern and western sides of the project area are primarily sediments. The majority of these units are greywacke, siltstone, arenite and iron formation (locally termed BIM or banded-iron-magnetite).

The SVC or SGW (volcaniclastic – clastic sedimentary rock) is mainly a greywacke-turbiditic clastic sedimentary unit. The beds are typically 1-10 cm thick, well sorted and often graded. The logging code ‘SVC’ has been inherited from early days of exploration when the rocks were interpreted to be volcaniclastic sediments.

The BIM is composed of alternating layers of dark greenish magnetite-grunerite and light gray quartzofeldspathic compositions, typical of Algoma style iron formation. The aluminum component is detrital and has a composition-mixing trend to the clastic sedimentary rocks (SVC, SGW). The iron formation units can vary from cm to decimeter scale in thickness, with mm to cm beds common. Although the units are locally tightly folded, attenuated or boundinaged, individual units can in some cases be traced for hundreds to thousands of metres along strike. Three BIM units have been logged and mapped in the West Branch project area; two in the hanging wall, and one in the footwall of the Tasiast thrust system. The hanging wall BIM units are generally unmineralized and the footwall BIM is variably mineralized (dependent upon proximity to structure). The contact between the hanging wall BIM and the FVC units is locally defined by the presence of a discontinuous conglomerate that contains abundant clasts derived from the FVC unit and a subordinate proportion of clasts derived from mafic metavolcanic units.

Mafic dykes that are post schistosity and post mineralization are dark olive green, fine to medium-grained and are locally plagioclase phyric. Dykes are typically less than 5 m wide, weakly magnetic and have locally developed hornfelsed and brecciated margins with a carbonate-chlorite assemblage. The dykes are dominantly barren and crosscut mineralized units.

The Tasiast deposits are hosted within a package of strongly foliated and folded rocks in the hanging wall block of an assumed thrust fault or thrust fault system referred to as the Tasiast thrust system (Figure 7-10). Modelling and interpretation of high-resolution gravity data (Figure 9-1) shows deep geometry suggestive of a thrust system underlying the Aouéouat belt. The Tasiast thrust system displays zones of strong deformation typically 0.5 m to 10 m wide and characterized by laminated foliation with locally preserved mylonitic textures. Hydrothermal alteration assemblages, sulphides and quartz veins are commonly spatially associated with the zones of intense deformation.

All of the Tasiast deposits host an intense, generally N-S striking, variably dipping, penetrative foliation, S1, which is axial planar to tight isoclinal folds in the host sequence (Davis, 2018). The foliation fabric within the main mine sequence, at West Branch, dips moderately to the East at 40-50°, steepens to the North to 55-65° at Piment and becomes sub-vertical, at the north end of the mine sequence, near Prolongation.

Pit mapping at Tasiast includes the collection of structural fabric measurements for structural geology and geotechnical application. Numerous consultants have assisted Kinross in developing a structural model for the Tasiast deposits along with developing pit mapping procedures. Most recently, in 2018, Dominique Chardon worked with the Mine Geology department to review and analyse the significant database of structural data collected from pit mapping activities. A synthesis of this review are presented in Figure 7-11 and Figure 7-12 as detailed pit-scale maps.

Quartz-carbonate veins sets occur sub-parallel and oblique to foliation and range in style from boudinaged, buckled, folded to planar. The veins clearly formed in extensional and/or Riedel shear orientations and were progressively folded, rotated, locally boudinaged and partially or wholly transposed parallel to the foliation. In the core of the West Branch Greenschist Zone vein, densities are typically higher in the meta-intrusive dioritic unit (averaging between 2% to 5%) than in the meta-basalt (<2%). This higher density suggests the coarser-grained feldspar-rich dioritic facies focused stresses and readily developed brittle-ductile shears, as expected for quartzofeldspathic rocks under retrograde Greenschist metamorphic conditions. Along the margins of the West Branch deposit, both the dioritic and meta-mafic volcanic units have a low vein density (<1%). Quartz-carbonate veins also developed locally within FVC that envelops the Greenschist Zone and within the footwall meta-sedimentary units.

a) Fabric measurements (in red and blue) and fabric trajectories (in black); b) Fabric trajectories and interpretative shear zone pattern (in red). Pits are shown in grey. Source of fabric measurements: Geology team, Kinross technical service

a) Fabric measurements (in red and blue) and fabric trajectories (in black); b) Fabric trajectories and interpretative shear zone pattern (in red). Pits are shown in grey. Source of fabric measurements: Geology team, Kinross technical service.

All of the rocks in the project area have undergone lower amphibolite facies metamorphism. Given the metamorphic grade, it is challenging to identify the rock’s alteration as it has been largely, if not totally, overprinted by the metamorphism. A description of the most commonly observed types of alteration are presented in Table 7-1 (see also Figure 7-13).

The bulk of the mineralization at the West Branch is hosted within the GDI. The GDI typically shows a zonation from a barren garnet-amphibole assemblage at its margins to an auriferous quartz-biotite-ankerite-pyrite-pyrrhotite assemblage and back into the barren garnet-amphibole assemblage. This zonation likely reflects a metamorphosed alteration assemblage with the garnet-amphibole assemblage representing a chlorite alteration precursor and the biotite-quartz-sulfides representing a quartz-sericite precursor.

a) Isoclinally folded sediments that are representative of the host rocks for all the Piment deposits, Tasiast Mine, pen for scale. b) Strongly foliated meta-diorite from West Branch typical of the Greenschist Zone. c) Silica flooding and extensive sulphide mineralization, mainly pyrrhotite, parallel to bedding in the metaturbidites at Piment. d) Visible gold hosted in a quartz–carbonate vein, in drill core taken from the West Branch deposit.

Recent ore control (and pit mapping) identified a high grade quartz-carbonate-chlorite-tourmaline-gold vein which is coincident with an interpreted Tasiast splay (locally termed the central fault). The vein has been mapped over several benches striking 330-340 degrees and dipping at approximately 55-60 degrees to the east (Figure 7-14 and Figure 7-15).

Quartz-carbonate veins sets occur sub-parallel and oblique to foliation and range in style from boudinaged, buckled, folded to planar. The veins formed in extensional and/or Riedel shear orientations and were progressively folded, rotated, locally boudinaged and partially or wholly transposed parallel to the foliation. Density of veining is typically higher in the GDI (averaging between 2% to 5%) than in the meta-basalt (<2%). Quartz-carbonate veins also observed locally within FVC that envelops the GDI and within the footwall meta-sedimentary units.

Gold occurs as both microscopic grains and coarse visible gold. When observed in hand specimen, grains are commonly spatially associated with hairline fractures in quartz veins and margins of sulphide minerals. The majority of the quartz veins containing coarse visible gold cut the foliation at a slightly oblique angle and mainly dip gently to the east. The majority of gold grains occur along the margin of the gangue and ore minerals, with 98% of the calculated volume/mass of the grains occurring in liberated and partially liberated forms. By volume/mass calculations, the majority (greater than 70%) of the volume is associated with the coarser (plus 75 μm) size fraction. Encapsulated gold grains are rarely observed. When present they are predominantly a very fine grain size. Semi-quantitative scanning electron microscopy analysis of gold grains indicated low silver (less than 15%) and trace iron (less than 3%) content.

Tasiast gold deposits are hosted in Archaean volcanic-sedimentary sequences that have been deformed and metamorphosed to lower amphibolite peak metamorphic grade. Mineralization is both structurally and lithologically controlled, epigenetic in style and was coincident with early stages of post-peak metamorphic retrograde Greenschist P-T conditions.

The Tasiast gold deposits fall into the broad category of orogenic gold deposits. The regional geological setting and deposit features at Tasiast are similar to other well-known Archaean lode gold deposits hosted along greenstone belts in granitoid-greenstone terranes.

Examples of analogue terranes of similar ages to the Aouéouat belt include the Kaapvaal craton in South Africa, the Pilbara craton in Australia and the Wyoming craton in the USA. The Aouéouat belt also shares many similarities with gold-rich Archaean terranes, such as the Yilgarn in Australia and the Abitibi in Canada.

Exploration Activities have been undertaken by TMLSA, its precursor companies (e.g. NLSD), consultants and contractors (e.g., geophysical surveys).

The Coordinate System used on site is a mine grid, a truncated UTM Zone 28 North grid system; the UTM Easting is shifted by -400,000m and UTM Northing is shifted by -2,200,000 m. The Original Control has been set out by IPH Engineering and ten control points are set out across the mine. Surveyors use a differential GPS for surveying at the mine (Trimble DGPS TSC3, Scanner MAPTEK I-Site 8820, and Trimble Drone UX5HP).

Numerous phases of geological and regolith mapping have been undertaken during the life of the project, and range from regional (1:100,000) to prospect (1:1,000) scale. Work was completed by the BRGM, SNIM, NLSD, Defiance, Red Back and Kinross. Mapping was facilitated by good outcrop, RC, and DD drilling, high resolution satellite imagery and detailed airborne geophysical data. Results were used to identify areas of alteration, structural complexity, quartz-carbonate veining, and sulphide outcrop that warranted additional work.

A total of 20,524 surface samples have been collected by Kinross since it started operating the project, including soil samples (40%) and rock samples (60%) that cover a surface area of approximately 1000 km2. In addition, 299 auger drill samples were collected during 2016. Soil samples were collected by a contractor and supervised by Kinross staff. The sample grid was generally west-east with lines spacing at 800 m and sample spacing at 200 m. The geochemical sampling includes exposed geology as well as areas covered by sand; in which the bedrock was sampled with auger drilling. Accordingly, the geochemical dataset has the potential to identify areas of prospective mineralization, otherwise blind from surface mapping. Surface exploration geochemistry samples were analyzed for gold and multi-element geochemistry.

To complement the surface exploration geochemistry, a geochemical study is being conducted to characterize the geochemical footprint of proven economic mineralization. The study so far indicates that the dominant gold pathfinders are As, S, Ag, and Te. In addition, the lithogeochemical footprint of the known deposits is being characterized to detect the chemical attributes of rocks that are better chemical traps to gold mineralization. Collectively, the geochemistry of the known deposits is used to evaluate mineralization potential and to delineate areas with target and pathfinder metal anomalies for follow-up exploration such as infill surface geochemistry, geophysics, and ultimately drilling.

Airborne magnetic-radiometric surveys were completed by NLSD (2000-2001) and Red Back (2007). These surveys were mainly used to map out lithological formations and major structures. In 2008-2009, Red Back completed a helicopter-borne electromagnetic (VTEM) survey. In 2011, Kinross completed airborne magnetic and radiometric surveys over the complete license package. This survey overlapped the previous survey and generated a higher resolution version. In 2013, Kinross completed ground based gravimetric and induced polarization (PDIP/IP) surveys. The gravity survey covered the complete license package (Figure 9-1) while the IP surveys covered only specific prospect areas; South West Branch South, Fennec, C67, C68 and Morris. In 2014, a comprehensive review was completed by a consulting geophysicist, along with some reprocessing of the magnetic data.

Excavation of trenches as an exploration technique has been very successful at Tasiast. In total, 445 trenches for 100,527 m have been completed across the Tasiast lands. Historically, trenches were completed manually, and more recently, trenches are completed using an excavator. The standard, excavated trench dimension is approximately 2 m wide and not more than 1.5 m deep and typically sampled every 2 m along the full length of the trench.

Numerous petrographic and gold deportment studies have been completed by TMLSA and predecessors on Tasiast in 2006, 2010, 2011, 2012, and 2017. In 2010, Red Back submitted 10 core samples from West Branch for a petrological and mineralogical study. Results from the work indicated significant pyrrhotite mineralization developed along foliation planes and associated with accessory magnetite, chalcopyrite and pyrite (Strashimirov, 2010). Further petrological studies were carried out for Kinross in 2010-2017, including work by Leitch (2010) Larson (2011), Pollard (2011) and Panterra Geoservices (2012 and 2017) which concluded that Quartz veins are pre- and/or syn-tectonic and folded or transposed into the dominant foliation and Pyrrhotite is the dominant sulphide mineral in many samples. A mineralogical (gold) characterization study of five samples was completed by Blake (2011a, b). The main conclusions were that coarse gold forms a significant component of the total gold content in the samples and that native gold grains encapsulated within their host (gangue/ore minerals) are relatively uncommon and often exhibit a very fine grain size.

The Tasiast project area has significant exploration potential to delineate additional resources both around the Tasiast mine (near mine exploration) and within the wider district (generative exploration). Exploration targeting and target ranking at Tasiast incorporates all available data sets including; satellite imagery (Worldview-2), airborne geophysical data (high resolution aeromagnetics and VTEM), reconnaissance scale geological prospecting, regional-district-target scale geological mapping, surface soil and auger sampling (gold and multi-element geochemical data), trenching, reconnaissance style shallow RC drilling on fences and conventional reverse circulation/diamond drilling to define resources.

Exploration efforts to date have discovered additional prospects, gold deposits and mineral resources along strike to the North and to the South of the main Tasiast mine area (West Branch and Piment-Prolongation), and generally along the Aouéouat (Tasiast) belt. The deformed greenstone rocks to the west (Imkebdene-Kneiffissat) of the Aouéouat belt are notable in that they host quartz-carbonate veins with anomalous gold values, however to date no significant deposits or mineral resource have been defined. To the immediate north of the Tasiast operation (5-12 km) and within the Guelb El Ghaîcha license, a cluster of deposits referred to as “North Mine Satellites” have been outlined, these are Fennec, C67 and C68. These gold deposits currently host approximately 0.5 Moz Au and are part of the near-mine resource growth strategy. Further north of the Tasiast operation (12-25 km) and within the Imkebdene and Tmeimichat licenses are another cluster of gold deposits referred to simply as “Morris”, these are Tef, Askaf, Central, NE, N1 and N2. This large area saw extensive exploration drilling from 2012 to 2014, which resulted in the discovery of several small deposits best described as narrow, high grade vein systems. Most of these deposits are open to depth down plunge.

Beyond 25 km from the Tasiast operation, within the northern extents of the Imkebdene and Tmeimichat licenses, are several gold exploration prospects that are pending follow-up exploration and drilling, of note are C23, Kneiffissat and Grindstone. These prospects have significant surface geochemical footprints and are considered highly prospective.

To date, 15,360 drill holes (14,286 RC, 840 diamond core and 234 RC-DD) for an aggregate total of 1,683,635 meters have been completed within the three mining licenses that constitute the Tasiast project area; Guelb El Ghaîcha, Imkebdene and Tmeimichat (Figure 10-1, Table 10-1). Drilling activities peaked in 2011 during the West Branch resource definition program. Drilling activities were conducted by various drilling contractors and supervised by geological staff of the Project operator. Where programs are referred to by company name, that company was the Project manager at the time of drilling and was responsible for the collection of data. (Table 10-1).

Drill programs were completed primarily by contract drill crews, supervised by geological staff of the Project operator. Where programs are referred to by company name, that company was the Project manager at the time of drilling and was responsible for the collection of data.

Between 1999 and 2000, Normandy LaSource Development Ltd. (NLSD) completed 339 RC holes for 28,039.69 m (including nine RAB holes) and 46 diamond core holes for 5,396.27 m at the Piment deposit area. Drilling was initially undertaken along 200 m spaced east-west sections at 50 m hole spacing, to 50-100 m depth. Drilling methods were predominantly RC with lesser core drilling (HQ; 63.5 mm core diameter) that included RC pre-collars with diamond core tails (NQ; 47.6 mm diameter core).

No drilling was completed by Newmont during the period that it held the Tasiast property (as part of its acquisition of Normandy).

From March to April 2003, Midas drilled 84 RC holes for 7,898 m and 29 DD holes for 2,908.4 m at the Piment deposit area. Diamond core drilling was completed mainly with HQ3 core diameter (61.1 mm) and PQ3 core diameter (83 mm) for seven geotechnical holes and three metallurgical holes. In addition, 4 RC pre-collars with diamond core tails for a total of 1,417.2 m were completed to test down-dip extensions.

Defiance completed 225 RC drill holes for 19,121 meters at the Piment deposit area. For the most part, the drilling program focused on infilling existing NLSD drill holes along 25 meters spaced, east-west drill fences.

Between 2004 and 2007, Rio Narcea completed 246 holes for 24,024 meters mostly aimed at extending the Piment deposit northwards towards what is now referred to as Prolongation and in addition completed sterilization drilling over planned waste dumps and tailings storage facility.

Following the acquisition of the Project in 2007, Red Back commenced an infill program of RC drilling to fully define and grow the mineral resources at and around the Piment deposit. In early 2009, step back drilling to the south of Piment, discovered what is now known as the West Branch deposit. For the remainder of 2009 and into 2010, RC drilling was ramped up to test the resource potential of West Branch. In late 2009, diamond core drilling was increased added to the continuation of West Branch mineralization beyond the depth penetration limits of the RC rigs. A small RC rig was used to conduct shallow (40 m) RC drilling on district targets along reconnaissance style fences. In total, Red Back completed 5,857 drill holes for 522,844 meters.

Upon closing of the Red Back acquisition in 2010, Kinross further accelerated drilling activities, by 2011, a total of 23 drill rigs were operating on site. From 2010 to 2012, drilling was primarily aimed at resource and reserve growth of the West Branch deposit. In addition, drilling activities to support mining studies were completed such as geotechnical, hydrogeological and metallurgical. 2013-2015 drilling shifted focus to the northern licenses; Tmeimichat and Imkebdene licenses with an aim to define resources that could be used in a study to support the conversion of both licenses from prospecting to exploitation. 2016-2018 drilling refocused back on the Guelb El Ghaîcha license and continued to test near-mine exploration targets. In total, Kinross has completed 8,530 drill holes for 1,071,985 meters (71% reverse circulation, 17% diamond core and 12% combination RC-DC).

For the Red Back and Kinross RC drill programs, a geologist first described (logged) the drill cuttings (chips) and then placed a representative sample into pre-labelled plastic RC chip boxes. The logging data was recorded directly in digital format at the rigs into the Database System. Data recorded included drill hole ID, sample number and depth, oxidation state, colour, presence of water, lithology, texture, structure, alteration, presence and type of quartz carbonate veining, and presence, type and abundance of sulphide minerals. Prior to 2009, diamond core logging geologists recorded geological and geotechnical descriptions on separate, hard copy log sheets and then input to Microsoft® Excel files. In 2009, the diamond core logging methodology was converted to the current system of digitally recording geological information via a note book or tough book computer into a Fusion database which was replaced with Acquire in 2018. Diamond core logging collected Rock Quality Designation (RQD), lithology, oxidation, alteration, sulphide mineralogy, structure, and veining. All diamond core holes have been photographed in the entirety, with digital camera and are stored in site server.

Pre-Red Back, drill collars were surveyed upon completion, using a Geodimeter 510 total station instrument. During the Red Back and Kinross drilling programs between 2006 to 2012, drill hole collars were surveyed immediately after completing the holes or later, initially with electronic distance measurement (EDM) and differential GPS. Once completed, the Cartesian coordinates were digitally recorded and emailed to the database manager to be imported into the database and from 2013, the survey data was imported directly into the Database. From 2013 to date, exploration drill collars are surveyed exclusively with a differential GPS, operated by trained staff with oversight by Tasiast Survey dpt. The drill collar locations are collected in the local grid system (Chapter 9) and includes a comprehensive array of permanent and semi-permanent survey stations, which have been checked for internal consistency by numerous EDM traverse closures and numerous comparisons with differential GPS data. Kinross completed an internal audit (re-survey) in 2013 using a differential GPS with 87% of all the project holes identified and validated.

Prior to 2010, most of the drilling was completed by shallow RC and did not include down-hole survey due to the complexities of surveying RC drill holes. Where diamond core drilling was completed (typically in deeper drill holes) and in selective cases of RC drilling, Humphries gyroscope, Maxibore and Reflex single shot down-hole survey tools were used. From 2010 to 2013 Kinross used three different contractors to complete down-hole surveys; ABIM solution (Year 2010 with SPT004 NS GYRO instrument and measurements were taken every 5m), WELL FORCE International (Years 2010-2012 with Gyro and MEMS instruments, and measurements were taken every 10m), and SEMM logging (Years 2011-2013 with SPT Gyro 07 and SPT Gyro 109 instruments, and measurements were taken every 5m). From 2014 to 2017, the down-hole surveys were completed by trained Kinross staff using MEMS and North Seeking Gyro 103 (measurements were taken every 10m). In 2018, down hole surveys were completed with Reflex EZ Gyro, operated by drilling companies (measurements were taken every 10m for core holes and 24m for RC holes). Considering the complete project database; 60% of all drill holes have down-hole survey data (58% of RC, 99% of RC-DD combination and 77% of diamond core drill holes).

Prior to 2013 total RC sample weights were not collected routinely however based on selective, available data, RC recoveries were determined to be acceptable. From 2013 onwards, total sample weights samples were routinely collected and confirm that recoveries are good. Recovery data for diamond core holes was collected from all Red Back and Kinross drill programs. Based on 17,718 measurements the average total recovery from core runs (in both oxide and fresh) is 98% and the RQD is greater than 93%. Measurements from downhole depths below 50 m (approximate oxide-fresh boundary) returned values of 99% and 95% for total recovery and RQD, in comparison to shallower depths where total recovery is 87% and RQD averages 43%.

Both the Piment and West Branch deposits dip eastward at moderate angles (approximately 40°-60°). In consideration of the deposit geometry Exploration and Resource definition drilling at Piment and West Branch was inclined at approximately 60° towards azimuth 270° (drilling east to west). The Piment and shallow portion of West Branch deposits were initially drilled along 50m spaced sections at approximately 50m drill hole spacing. Infill drilling was later completed along 25 meter sections to approximately 25m drill hole spacing. Deeper drilling (down dip and down plunge at West Branch) was completed along 50m spaced section with approximately 75m hole spacing.

Geomechanical drilling campaigns for collection of rock mass characteristics and ground-water conditions were conducted across the mine lease to identify potentially suitable locations for mine infrastructure. The various geotechnical studies completed, based upon the progress of pit development, include the following:

Hydrogeological drilling was conducted in the area of the water bore field, to identify sufficient water for processing. Large diameter core holes (typically PQ) were completed to collect samples for metallurgical test work.

In the opinion of the QP, the quantity and quality of the lithological, geotechnical, collar and down hole survey data collected in exploration and infill drill programs are sufficient to support mineral resource and mineral reserve estimation as follows:

As the geochemical and trench data have been superseded by information from drilling and mining operations, these sample types are not discussed further in the Technical Report. This information is not relied on for use in geological modelling or resource estimations.

Little information has been kept or is available regarding drilling procedures used in this drilling by NLSD.

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

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

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

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

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

At the core cutting facility the drill core boxes were stacked in ascending order so as to avoid sampling mix-ups. The core was cut on the orientation line marked by the geologist and the right hand side of the core (looking down hole) was placed in a numbered calico bag.

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

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

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

After September 2013, RC samples with a nominal weight from 36 kg to 40 kg (each 1 m) were collected in a large plastic bag, tightly clamped onto the cyclone base and reduced in the field by 50/50 manual riffle splitters. About 6 kg to 8 kg weight samples were placed into pre-numbered sample bags to dispatch to the laboratory. Every 20 samples, a field duplicate was collected as part of the quality assurance/quality control (QA/QC) procedure.

For diamond drilling, core was transported from the drill rigs to the core facility where geological and geotechnical core logging was completed. The geologist marked one-metre intervals and orientation lines (bottom of hole) along the core axis for core cutting. A record was made at the core facility of the sample identity numbers and corresponding intervals. At the core cutting facility the drill core boxes were stacked in ascending order so as to avoid sampling mix-ups. The core was cut on the metre and orientation lines and the left hand side of the core looking down-hole is placed in a numbered plastic bag with sample ticket.

Once the core for a drill hole was cut, sampled and bag sealed, the core was then stored in a secure area (either locked 40 ft shipping container or fenced off area) for sample dispatch.

All the sampling processes for RC and diamond drilling were handled under TMLSA’s chain of custody.

The results from 1,699 bulk density determinations completed by NLSD at Tasiast during previous drilling programs are available. The origin of the sample, its borehole number and sample depth was entered as an individual MS Access database file into NLSD’s project database. However, information on the sample size/length, lithology and oxidation state was not recorded in the NLSD database. The bulk density measurement for each NLSD sample was derived by using the Weight in Air/Weight in Water (Archimedes) method. The oxidized core samples were sealed with molten wax and re-weighed to determine the weight of the paraffin coating, prior to weighing in water. The bulk density determinations were done on short (5 cm), half core specimens, taken at close intervals. The NLSD bulk density data were collected from one core hole in the Piment South area and from 13 core holes from the Piment Central area.

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

From 2008 to December 2011, Red Back and TMLSA completed 24,702 specific gravity determinations of bulk density using the Archimedes method. The samples were selected to provide a representative suite of densities covering all major lithology types and from all oxidation levels.

Initial Red Back and TMLSA density determinations were done using wax-coated samples for both oxide and primary material. This procedure was changed by using uncoated core samples for only primary material to speed up the test work. Duplicate tests with one-wax coated samples for every lithology per hole were done to evaluate bias between the data pairs. About 650 duplicate tests were done up to December 2011. Initial analysis of the check samples showed very good correlation between the uncoated and coated density values. A total of 90% of the dataset shows a difference of 1% variability between the samples pairs (coated and uncoated).

Sample preparation was undertaken on site by NLSD staff during their drill programs. Analytical laboratories used were the BRGM laboratory in Orleans, France and the OMAC laboratory in Ireland. QA/QC was undertaken by Genalysis Laboratories in Perth, Australia, and SGS Laboratories in France. Laboratory accreditations at the time are not known; all analytical laboratories were independent of NLSD.

During Defiance’s RC and diamond drill programs, the analytical work was carried out by SGS Analab in Kayes, Mali and by Abilab located in Bamako, Mali. Analab is an ISO accredited laboratory whereas Abilab is not ISO accredited. The laboratories were independent of Defiance.

Following Red Back’s acquisition of the Tasiast deposit in August 2007, an on-site SGS Analab sample preparation facility became operational. Prior to that time, samples had been prepared on site by the previous owner’s technical crew members under supervision of senior geological staff. All drill samples since 2007 have been prepared and analysed under contract by SGS on site and by SGS Analab in Kayes, Mali, SGS Analab in Morila, Mali, and SGS in Ouagadougou, Burkina Faso. Laboratories were independent of Red Back. The two SGS laboratories hold ISO9000 accreditations.

In December 2010, SGS constructed and commissioned a mobile sample preparation facility in Nouakchott, Mauritania, and selected samples were submitted to the facility for preparation. In late 2011, a new on-site SGS preparation and assay laboratory was commissioned at Tasiast, with a capacity of up to 2,000 samples per day. In mid-2012, TMLSA stopped sending exploration samples to the SGS Tasiast preparation laboratory due to quality control concerns. Due to the large volumes of the samples and turnaround time issues, TMLSA started sending samples to nine different accredited laboratories outside the country for sample preparation and assays. These were Actlabs Burkina Faso, ALS Johannesburg, ALS Kumasi, ALS Loughrea, ALS Romania, ALS Vancouver, SGS Kayes, SGS Morila, and SGS Ouaga. In April 2013, ALS Chemex took over the Tasiast laboratory facilities and initiated carrying on the sample preparation and analytical services. All drill hole samples have been prepared and assayed by the ALS Tasiast laboratory.

Midas, Defiance and Rio Narcea RC drill sample preparation involved the entire RC calico sample bag, which was oven-dried for 24 hours and then weighed before pulverizing the entire 2 kg to 3 kg subsample using a Labtecnics LM5 mill. Each core sample was crushed to -10 mm in a jaw crusher, and the entire sample was pulverized to P90 (90% passing) at 75 µm using a Labtecnics LM5 mill. Barren dune sand was used to clean the bowls after every sample. The pulverized material was sampled using a spatula, and two 120 g pulp sub-splits were taken; one packet was prepared for shipment to the assay laboratory and one packet remained on site for future reference. Blanks of dune sand and certified reference standard were then inserted with the field samples.

Sample pulp shipments were conducted on a weekly basis. The samples were transported in secured wood boxes to Nouakchott, where Mauritanian Custom inspected the shipment and released the proper documentation for exportation. The boxes included a sample submission sheet prepared by the laboratory manager. Samples were then shipped by airfreight to SGS Analab.

At SGS Tasiast and SGS Nouakchott, the entire RC and core sample was oven-dried for 24 hours in a cleaned metal dish, weighed and then crushed to 75% passing at 2 mm. At SGS Tasiast, a 1.5 kg subsample was split using a Jones riffle splitter, and pulverized in a Labtech Essa LM2 ring pulverizer using a 2 kg bowl to 85% passing at 75 µm. At SGS Nouakchott, the sample was split once using a Jones riffle splitter and pulverized in a Labtech Essa LM2 ring pulverizer using a 2 kg bowl to 85% passing at 75 µm. Both laboratories took a 200 g subsample for gold (Au) fire assay.

At the SGS Tasiast laboratory and relocated mobile sample preparation facility, the procedure for sample analysis remains unchanged. However, subsample size at the Tasiast laboratory has been increased to 2 kg, to improve the precision of results.

For RC and core samples processed by SGS Analab in Kayes, samples were stockpiled in a secure area within the Tasiast core facility, and collected by a truck contracted by either Analab or TMLSA for shipment to Kayes. The samples were enclosed and secured in a large tarpaulin and transported directly from the site to the laboratory. The entire core or RC sample was oven-dried for 24 hours and then weighed before pulverization. Samples were crushed to 75% passing 2 mm, and two 1.5 kg subsamples were split using a Jones riffle splitter and pulverized in a Labtech Essa LM2 ring pulverized using a 2 kg bowl to 85% passing at 75 µm. These two pulps were recombined before being subsampled (200 g) for an Au fire assay.

After ALS Chemex took over the laboratory facilities at Tasiast in early 2013, a few changes were introduced at the sampling preparation stage, including:

Check analysis of 74 Phase 1 samples showed no significant variations between roasted and non-roasted samples (Guibal et al., 2003).

All of the sample pulps from the Midas, Defiance and Rio Narcea drill programs were analysed for gold using a 50 g fire assay with an atomic absorption spectroscopy (AAS) finish at both laboratories. The Analab 50 g fire assay/AAS method (FA50) has a lower detection limit of 0.005 g/t Au; Abilab’s lower detection limit is 0.010 g/t Au.

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

For Red Back and TMLSA samples, sample pulps were analysed for gold using a 50 g fire assay with an AAS finish with a detection limit of 0.01 g/t Au. Results higher than 5 g/t Au were re-analyzed by fire assay technique and gravimetric finish. In 2012, TMLSA began gravimetric finishes for gold above 5 g/t, and also began screened metallic fire assays.

Most of the documented QA/QC cited by SRK in 2003 on NLSD samples related to measurements of the analytical errors through pulp duplicates, where two analytical methods (aqua regia: AAS and Fire assaying: FA) are compared (Guibal et al., 2003). No significant problem was detected. In early 2003 a total of 429 pulp samples, collected by staff from TMLSA and representing close to 10% of the mineralized samples within the wireframed resources, along with 54 standards (of values 0.5, 1.66 and 3.22 g/t Au) and 18 blanks were re-assayed by Genalysis. SRK noted that the Genalysis results compared well with the database and standards and blanks were assayed within acceptable limits.

For the Defiance and Rio Narcea drill programs, a total of 21,686 RC sample pulps, including field duplicates, blanks and standards, and 904 diamond drill hole core sample pulps, including field duplicates, blanks and standards, were shipped in 16 batches, of which 14 went to Analab and two went to Abilab. Included within these sample batches were a total of 774 field duplicate samples, each one being a second split from a 1 m interval field sample bag, and 1,136 preparation duplicates, each one being a second split from the pulverized RC and core sample at the preparation laboratory.

The analytical QA/QC program implemented by Defiance was monitored by the routine submission of commercial SRMs purchased from Gannet Holdings Pty. Ltd. of South Perth, Western Australia. SRMs were inserted at every 20th sample and an internally prepared coarse blank sand inserted at every 10th sample within the RC and core sample stream. Field duplicates were collected by the field geologist after the completion of each RC hole and the number of field duplicates on a per RC hole basis was dependent on the length of the hole or equivalent to every 20th sample. Preparation duplicates were selected for every 20th sample number in a sequence and submitted as a separate sample number series on a per batch basis.

In 2011, TMLSA engaged an independent consultant to provide a regular review of the QA/QC data. Issues identified during the early review in September 2011 (Heberlein, 2011), such as switched standards and standard identification, have been corrected and control actions implemented.

Additional actions implemented to address other recognized issues, such as duplicate precision, include the following:

In 2012, a QA/QC team was established and became in charge of the sampling protocols, sample transport, sample tracking and reporting.

All drill hole and geological data in Tasiast has been collected and stored using a variety of software and systems. Not much information exists regarding the way the data was stored at the very beginning prior to Red Back, but a data analysis suggests that the data was collected using spreadsheets and then imported to the 3D software which was used at the time. During this process much of the information regarding assays, such as laboratories and methods used, was lost.

In 2007, Red Back implemented a Relational Database Management System called Fusion, which was developed by Century Systems and later sold to Datamine. The System Architecture for Fusion consists of three different components: A Central Database, Fusion Remote, and Local Databases, but it can be implemented with only two of these components: Central and Locals. Red Back implemented with the second option as shown in Figure 11-1.

After Kinross acquired Red Back Mining in September of 2010, Kinross continued using Fusion as it was used by Red Back. In 2012, Kinross migrated the Tasiast System Architecture to the Central – Fusion Remote – Locals Model, shown in Figure 11-1b, which had already been implemented by Kinross Gold globally. The Central Database (structure owner) was located in Toronto. The sites (Brazil, Chile, Tasiast, Russia) hosted in their local server the Remotes, which were the data owners. In Tasiast, Exploration and Grade Control were sharing the same Remote database. However, in 2014 Technical Services Tasiast separated them in two different databases: Exploration still would have the architecture shown in Figure 11-1b and Grade Control would go back to the architecture shown in Figure 11-1a, with the Central being located in Tasiast.

In January 2018, Kinross replaced Fusion with acQuire, another Relational Database Management System based in SQL. The Exploration Fusion Remote Database was disconnected from the Central database in Toronto and both Fusion databases at Tasiast (Grade Control and Exploration) were migrated and combined into a single database in acQuire. Since acQuire has more robust business rules than Fusion, when there were conflicting data (for example orphan samples, two records in the assay table for the same interval and method), the conflicting records were not migrated.

In October 2018, during a validation process of the assay records in acQuire, some issues that compromised the integrity of the data were encountered. The causes of the issues had different origins that had not been detected before. Some of the issues found are listed below:

The rebuild of the assay database was done by re-importing all the lab certificates that were available and then performing the QA/QC to pass or reject assays using the procedure described in the sections above in this same chapter. The re-import of certificates for holes in the West Branch area was completed in March 2019, then the reimport continued with the Piment and Prolongation areas and finishing with the Northern satellites. For records where no Lab certificates were found, the data was migrated from the previous database using a qualifier to denote that the data was not imported from certificates during this phase.

Sample pulps are returned from the laboratory in plastic vials or sealed paper envelopes, and these are stored in sealed containers at site. The majority of historic coarse reject samples were not stored, but TMLSA has commenced storing selected mineralized coarse reject material. The remaining half of the drill core is well stored in stacked wooden trays referenced by hole identification number and interval length. Some core intervals have been totally sampled for metallurgical or check (umpire) sampling.

Following TMLSA’s acquisition of the Project in September 2010, all drill samples collected are under direct supervision of Project staff from the drill rig and remained within the custody of staff up to the moment the samples were delivered to laboratory or placed on contracted trucks for delivery to the Mali laboratory. Samples, including duplicates, blanks and certified reference materials are delivered daily from the drill rig to a secure storage area within the fenced Tasiast core facility.

Chain of custody procedures consist of filling out sample submittal forms that are sent to the laboratory with sample shipments to make certain that all samples are received by the laboratory.

In the opinion of the QP, the sampling methods are acceptable, meet industry-standard practice, and are adequate for mineral resource and mineral reserve estimation and mine planning purposes, based on the following:

A number of verification checks have been performed on data collected from the Project, either in support of technical reports, or as part of the Project feasibility study.

A number of external consultants and consultancies have reviewed Project data, and made recommendations for future work.

SRK (Guibal, 2003) reviewed the data available in 2003, as part of supporting documentation for the acquisition of Midas by Geomaque, and commented:

ACA Howe inspected Defiance’s sample preparation facility, and considered the facility to be reasonably well equipped and maintained, in accordance with acceptable industry standards (Leroux and Puritch, 2003; Leroux et al., 2007).

Midas collected a total of 429 pulp samples of known NLSD drilled mineralized zones in early 2003 (Hyde, 2003). Midas inserted blanks and standards and submitted this sample batch to Genalysis. The Genalysis results compared well with the NLSD assays and the standards and blanks inserted by Midas assayed within acceptable limits.

A comparison of RC and core duplicate samples indicated no major bias across all grade ranges. Howe considered that the degree of scatter shown in graphed data was acceptable for resource estimation purposes. No bias occurred towards the higher-grade original or repeat assays.

A total of 134 one metre interval RC samples from six of Defiance’s RC drill holes and 27 core pulp samples from two Defiance core drill holes were submitted by Howe to ALS Chemex Laboratories in Mississauga, Ontario for check analysis. Upon review of the results, Howe was of the opinion that its independent check assay results confirmed the presence of gold mineralization at Tasiast.

Defiance selected mineralized intersections from 30 RC holes covering the four mineralized areas of the Piment Zone, which were sent to Canada for metallurgical test work. SNC-Lavalin (SNC) (Demers et al., 2004) reviewed the drill hole information on the geological sections prepared by Howe and combined the sampled intersections of several drill holes to obtain nine samples considered to be more or less representative for the various mineralized zones and their high and low gold grades. These samples were sent to SGS Lakefield in Ontario, Canada; an ISO/IEC 17025 accredited laboratory for assay. The comparison of the assay results of the initial samples and those from Lakefield was acceptable and showed a reasonable correlation.

SNC representatives collected eight samples of RC drilling chips that had previously been assayed by Analab. These samples were sent to Lakefield for assay. Results showed that gold was present in the indicated mineralized zones even though the correlation was rather erratic due to the statistically low number of samples.

Red Back conducted an analysis of the available, historical QA/QC data from Defiance and Rio Narcea as part of the February 2008 resource update comparing all historical data with data generated by Red Back as at February 2008.

Review of the blank, duplicate and SRM submissions in 2009 and 2010 (Stuart, 2009; Stuart, 2010) indicated no significant errors or biases in the analytical data. Prior to late 2009 the majority of the field duplicate analyses completed were from non-Greenschist mineralization styles, e.g., Piment iron-formation and West Branch footwall. A total of 16,907 (2009) and 15,929 (2010) QA/QC samples were blindly inserted as part of the routine sample preparation and were submitted for analysis. Red Back concluded that the QA/QC data reported was of industry accepted standards and the assay data was considered reliable for inclusion in the December 2008, 2009 and 2010 resource estimations.

In April 2013, SRK conducted a review of the analytical quality control procedures and results for the Tasiast gold project in Mauritania (Chartier, 2013). The objective of the review was to provide an independent analysis of the sampling procedures and a review of the analytical quality control results for the data to be used in resource estimation.

SRK visited the Project site from October 11 to 15, 2013. SRK also visited a third-party preparation laboratory operated by SGS Minerals in Nouakchott, Mauritania. The purpose of the site visit was to audit project technical data and collect all relevant information for the compilation of the Sample Preparation, Analyses, and Security and Data Verification sections of a technical report. SRK was given full access to relevant data and conducted interviews with Kinross personnel to obtain information on past exploration work and understand the procedures used to collect, record and analyze historical and current exploration data.

SRK reviewed the field procedures and analytical quality control measures used in the Tasiast gold project. In SRK’s opinion, Kinross personnel used care in collecting and managing field and assay exploration data. The sample preparation, security and analytical procedures used by Kinross are consistent with generally accepted industry best practices, and are therefore adequate to support the mineral resource estimation.

As part of database migration from Fusion to acQuire, the QA/QC data from 2007 to 2017 for the Guelb El Ghaîcha area, which contains West Branch pit, all Piment and Prolongation pits, and northern satellite deposits (C67, C68 and Fennec), were reviewed in preparation for the 2019 model update. The purpose was to validate all QC assays available in the database since Red Back’s acquisition of the project in 2007 and to ensure that only assays with acceptable QC results are being used for resource estimation. A total of 16 different assay laboratories were involved for the sample analysis in this period. Table 12-1 presents the summary of quality control samples reviewed this time. A total of 44,003 standard assays and a total of 12,261 blank assays were exported from the Tasiast acQuire database for review. Duplicate assays were not included this time because the purpose of the review was to select the assays with acceptable QC results for the resource estimation – duplicate assays present only precision information on each sampling/subsampling stage, which does not provide the pass/fail criteria.

The exported standard and blank assays were grouped and assessed by Lab Job number which is equivalent to sample shipment number. The Lab Job number may contain multiple assay batches. The pass/fail assessment was performed for 5,302 Lab Jobs. The pass/fail criteria applied were ±3 x standard deviation for standards and <=0.05 g/t Au for blanks. A total of 48 different standards and two types of blanks (barren sand and barren pegmatite) were used as QC samples. Among these 48 standards, 16 historical standards that were used from 2007 to 2011 do not have any information on the actual standard code or their standard deviations. A constant standard deviation of 3.33% had been used for these historical standards in QC check and this caused significant number of standard failures due to challenging control limit of 10.0% (3 x 3.33%). A decision was made to re-assign a constant standard deviation of 6.67% for these historical standards considering the acceptable industry practice standard deviation level of 5-7% for gold standard. This resulted in reducing the number of unnecessary QC failures. The average certified standards deviation of all standards including the historical ones was 5.6%, which is well within the acceptable industry practice level mentioned above.

Numerous occasions of QC sample swap have been identified especially during 2008-2012 drilling campaigns when there were challenging number of drills running spontaneously with limited trained work force. The total number of swaps identified was 906, which is equivalent to 1.6% of the all QC samples reviewed. Most of the swaps were easily identifiable in QC chart by the presence of cluster(s) of noticeably different grade value(s) than the certified value (Figure 12-1).

The number of QC sample swaps by year is summarized in Figure 12-2. The number of swaps has decreased to a normal level since the start of the 2013 drilling campaign. The identification of swaps resulted in lowering the QC failure rate significantly.

In a normal drill campaign, the standards and the blanks that failed QC control were requested to be re-analyzed where a material impact on the reported results was considered. In this case, the failed ones were simply excluded from the resource estimation due to unavailability of the original pulps.

An average accuracy of 98% was achieved for the internationally accredited standards (see Table 12-2). This implies that 98% of the total standard samples submitted to the various laboratories reported within acceptable limit of ±3 standard deviation. The blank samples submitted to the various laboratories reported the similar results - 98% of the samples below 0.05 g/t Au.

According to Heberlein (2013), measurable improvement was observed in the QA/QC procedures since his involvement in the Project, especially in duplicate assays. The improvement has resulted in a measurable increase in the overall precision of the analytical results. Early precision estimates of field duplicate results (containing the total error of sampling, preparation and analysis) showed unacceptably high values for both core and RC duplicate samples. The initial analysis of duplicate results in 2011 determined precisions above 85% range for drill core and above 80% range for RC chips. Improvements to sampling and sample preparation procedures, particularly at the on-site laboratory (TML) have brought the duplicate precision down to the 45% range, which is reasonable for the nugget style of mineralization at Tasiast

In 2010 and 2011 TMLSA also twinned three RC holes from the Greenschist zone. In 2010 two twin holes were completed on the lower portion of the Greenschist zone to test for due diligence and variability between RC and core drilling. Results from the work returned a strong to acceptable correlation between mineralized intervals in the RC and core holes. The purpose of the 2011 hole that drilled shallow in the Greenschist zone is to improve understanding of any potential sampling differences between RC and core material. Geological logs between the core and RC drilled in the shallow portion of the zone retuned similar data with lithological units within expectations.

The process of data verification for the Project has been performed by TMLSA, Red Back, and personnel of precursor companies, and external consultancies contracted by those companies.

The QP has reviewed the reports and is of the opinion that the data verification programs undertaken on the data collected from the Project adequately support the geological interpretations, the analytical and database quality, and therefore support the use of the data in mineral resource and mineral reserve estimation.

Data used to support mineral resource and mineral reserve estimates have been subjected to validation, using built-in software program triggers that automatically check data for a range of data entry errors. Verification checks on surveys, collar coordinates, lithology, and assay data have also been conducted. The checks are appropriate, and consistent with industry standards.

Ongoing sample preparation and analytical work is recommended to obtain more acceptable precision from the duplicate samples.

The Tasiast mineralization is free-milling and amenable to gold extraction by simple cyanide leaching. The existing mill has been operating since 2008, initially treating oxide banded iron mineralization (BIM) hosted ore yielding a typical gold recovery of 93%. Gold recovery from fresh ore, which forms an increasing portion of the mill feed since 2010, varies between 91% and 93%. A proportion of the gold is coarse and responds well to gravity concentration. Gold mineralization is associated with structurally controlled faults and shears, quartz-veining and silica-flooding. Gold grains observed in the exploration core holes are seen in isolated grains in quartz veins and are closely associated with pyrrhotite. The mineralization has relatively low levels of sulphides, approximately 1% to 5% S, predominantly represented by pyrrhotite and to lesser extents pyrite, arsenopyrite, and chalcopyrite. Other metal contents are low such as silver approximately 1 ppm to 2 ppm, copper approximately 100 ppm, arsenic approximately 10 ppm and very low levels of mercury, less than 0.3 ppm Hg.

The bulk of the metallurgical test work has been done to evaluate the optimum process for the West Branch ore which has become the major source to the processing plant. Major metallurgical sampling campaigns were conducted on the West branch mineralized zone and test work to optimize cyanide addition rate and grinding tests were completed.

Four major metallurgical sampling campaigns were conducted on the West Branch mineralized zone as follows:

A program of waste rock sampling and characterization was also undertaken with core samples selected to represent all rock lithologies and depths.

Test work was conducted by multiple laboratories and the results from the different laboratories were comparable.

Work was carried out by Ammtec, SGS and JKTech to determine the comminution characteristics primarily of West Branch samples. Tests were performed to assess the variation in comminution parameters and confirm grinding energy requirements for the deeper ore.

All of the samples were checked for their correct lithologies and split into separate lithologies for analysis. The mine plan by lithology is shown in Table 13-1 and comminution parameters obtained from laboratory testwork are in Table 13-2. The majority of the ore will be from the fresh granodiorite intrusives (GDI) and fresh banded iron (BIM) lithologies, with reclamation of the low grade stockpiles in 2024-2025 and again near the end the mine life.

The semi-autogenous grinding (SAG) mill test work indicates that the ore becomes progressively harder at depth. A typical relationship indicating the trend of increasing SAG mill grinding energy (SAG power index [SPI]) requirement with depth is shown in Figure 13-1.

2 A x b and ta are parameters in the JKTech Drop Weight Test model. Acronyms for the other parameters in this table are listed in Section 2.5 List of Abbreviations.

The current mine plan does not include mining to below a depth of 500 m, so the effective hardness increase is minimized.

Extensive metallurgical testing was completed on West Branch samples, twinned hole samples and deeper level variability samples. In general, test work indicated that the ore was amenable to gravity recovery and cyanide leaching, resulting in selection of a flow sheet similar to that of the existing plant. Some of the key parameters that resulted from the test work are:

The grinding test work results show that gold extraction increases with a finer grind size. Gold dissolution kinetics were enhanced at the finer 80% passing (P80) grind sizes of 90 µm and 75 µm. At the selected grind of 90 µm, test work indicates that the majority of leaching is complete at around 24 hours, as shown in Figure 13-2. Historically, the operating plant has shown 18 hours to be the optimal leach time, which was selected for design. The improved kinetics relative to the test work are likely the result of grinding in process water containing cyanide recycled from the tailings thickener and gravity recovery circuit, which removes coarse slower leaching gold.

The cyanide addition rate has been optimized to a low addition rate. Test work results indicate that cyanide consumption rate as low as 0.5 g/L is possible. Operationally, a cyanide addition rate of 0.7 kg/t is used for CIL.

A compilation of all the relevant tests done, limited to those samples within the currently defined resource, produced the recoveries shown in Figure 13-3.

The figure shows that all of the selected samples leached well, that oxygen enhancement improved leach rate, and that recoveries are between 84% and 94% at a grade of approximately 2 g/t. The recoveries are predominantly above 86%, with a few exceptions that, from a metallurgical perspective, gives high confidence that all the sampled parts of the orebody are amenable to gravity and cyanidation. Comparative tests using oxygen (dissolved oxygen, D.O. = 15-20 mg/L) vs. air (D.O. = 6-8 mg/L) indicate that oxygen increases gold extraction in the range of 0.5% to 1.4%, depending on cyanide concentration.

The trend line made up of representative samples indicates a relationship between head grade and gold recovery, with higher recoveries achieved at higher gold head grades, as expected. The mathematical relationship developed was used to estimate recovery based on the ore grade obtained from the mine plan.

AMMTEC performed flocculent screening tests on ground composite samples of West Branch ore using seawater obtained near Perth. Magnafloc MF336 flocculent was selected for subsequent settling tests to optimize flocculent consumption and develop thickener sizing criteria.

Thickening characteristics of deeper level variability samples were determined through Outotec test work in 2010, FLSmidth test work in 2011, SGS Lakefield test work in 2013 and FLSmidth test work in 2013. Outotec investigated the dynamic settling characteristics and determined the thickener sizing criteria. In 2011, FLSmidth conducted sedimentation and rheology testing. SGS conducted dynamic settling tests on a number of composite samples that had been prepared for leaching test work in 2012 and 2013. Based on test work a unit rate of 0.45 m2/t/d was selected for design.

Acid rock drainage (ARD) testing was completed on leach residue generated from the GDI samples in the AMMTEC 2011 follow-up test work program to simulate plant tailings. Results indicated that the leach residues do not have potential acid generating characteristics, but have significant acid consuming capacity (likely due to the carbonate content of each ore composite).

In 2011, a waste rock material characterization program was conducted by URS Scott Wilson and supported by Kinross Tasiast and SRK Consulting. During the study, 154 samples were collected from exploration drilling core of different lithologies to assess the ARD potential. Study results showed that waste rock typically exhibits a significant residual neutralization potential for all the lithologies investigated.

The study results, coupled with the favorable arid climate, lack of surface water and very limited groundwater (no viable groundwater aquifer exists) and a Materials Management Plan indicate low potential for ARD or metal leaching to develop.

The Mineral Resource statement for Tasiast comprises estimates for West Branch, Piment, Prolongation and the satellite deposits Fennec, C67, C68, and C69. Figure 14-1 illustrates the location of the satellite deposits relative to West Branch, Piment and Prolongation.

The Mineral Resources are stated in accordance with the definitions in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101). Mineral resources have an effective date of December 31, 2018. The West Branch resources were updated in 2019 by Kinross Technical Services to support the Feasibility study work completed in connection with the Tasiast 24k expansion.

The Mineral Resources were reported below the December 31, 2018 mined surface or topographic surface, using cut-off grades based on a gold price of $1,400/oz, and are constrained using Lerchs-Grossman pit shells. The Mineral Resources are classified as Measured, Indicated or Inferred primarily based on drill spacing and geological continuity. Table 14-1 shows the classified Mineral Resources exclusive of reserves.

The main components of the stated Mineral Resource are West Branch (WB), and Piment and Prolongation (PP), which is where all open pit mining has taken place to date. These deposits constitute 92% of the total resource. The remaining 8% of the resource is reported for the satellite projects Fennec, C67, C68W, C68C and C69. Table 14-2 shows the classified Mineral Resources inclusive of reserves by deposit. Table 14-3 lists the cut-off grades used.

The following sections detail the dataset and methodology used for the preparation of the Mineral Resource estimate. The Piment and Prolongation model remains unchanged from the 2013 Feasibility Study (FS). The West Branch model was updated in 2019 to support engineering studies. The satellite deposits do not provide material contributions to the overall resources, and the associated datasets and block models are therefore not discussed in detail in this report. Unless otherwise mentioned, all discussions in this section pertain to West Branch, Piment and Prolongation only.

The Tasiast database is maintained in a acQuire database management system implemented in 2018. Before 2018, data were stored in a Century Systems platform. The data were exported as .csv files.

The dataset for the West Branch, Piment and Prolongation models contained 4,750 drill holes, of which 90% are RC holes and 6.7% were diamond drill holes and 3.3% were RC pre-collars with core tail. The previous resource estimate contained 4,343 holes, so 407 additional holes were available for the models discussed in this report. A summary of the dataset used in this estimate is shown in Table 14-4.

* to avoid edge effects from the resource estimation process, drill holes were exported from the database with a 200 m overlap on either side of the initial model boundary of 72,325N.

The West Branch resource estimate was prepared by Kinross Technical Services. Geologic and estimation domains were prepared using a combination of Micromine and Leapfrog Geo software. Micromine was used to prepare assay data for geostatistical analysis and construct the block model. Isatis software was used for geostatistical analysis of raw and composited drill hole assay data, variography and to estimate metal grades.

The .csv files were imported into Micromine and checked in 3D for inconsistencies. No errors were detected. In the export, missing assays value are reported as -999. Assay values below detection limit are reported as half detection limit.

Most of the drilling is reverse circulation (72% of total samples) while the remainder is either diamond core (6% of total samples) or diamond with reverse circulation pre-collar (22% of total samples). To assess the impact of volume variance on the assay values for the different sample support, drilling data were paired (5.0 m and 10.0m) and QQ plots of paired data were constructed. Both sample types display the same distribution and therefore, both were merged into one dataset for estimation.

The WB drill hole dataset contains many missing assays of various interval lengths. To allow for consistent treatment of the missing intervals, in particular at the compositing stage, all assays were classified by the database team based on length. After review, it was decided to assign the background value of 0.0025 g/t (half the lowest detection limit seen at WB) to all missing assays with interval lengths less or equal to 2.0 m. A length threshold of 2.0 m was chosen because the majority (98%) of assays are of equal length and measure 1.0 m. Statistics of the raw length-weighted assays before and after addition of the latter missing intervals were reviewed. The resulting dilution is not material, and the background value assignment will ensure more consistency in compositing length and will avoid numerous smaller composite intervals.

Four sets of wireframes were used for estimation: lithology, redox, structure and mineralized domains. Geology wireframes were generated in Micromine based on sectional interpretations every 25.0 m for lithology and 100.0 m for structure. Redox solids and grade shells used as mineralized domains were built via implicit modeling in Leapfrog Geo. All wireframes are based on logging information and pit mapping data where available.

Table 14-5 summarizes the codes used in the block models and provides a brief description of each lithology.

Lithology, structure and grade were assessed for use as estimation domains. Key tools to evaluate the domains were exploratory data analysis, contact analysis and variography.

The estimation domains were derived from lithology and indicator kriging. The indicator shapes were used to constrain mineralization hosted in the footwall sediments. Table 14-6 describes the mineralization domains.

In addition to the estimation domains mentioned above, two additional domains were created for compositing and capping (Table 14-7).

The Upper and lower transition material were modelled together on the basis that visual logging cannot discriminate these materials consistently, and that their metallurgical differences are not considered material. Table 14-8 summarizes the redox codes that were used to code the redox variables in both block models.

Figure 14-2 and Figure 14-3 illustrate the lithology and mineralization domains in plan and section view.

The majority of the assay intervals (98%) are of equal length and measure 1.0 m. To reduce the variability and ensure same support, assays were composited using a 4.0 m compositing length. Composites honoured the mineralized domain boundaries as presented in Table 14-6. The composites were not broken down on mineralization domains 1041 and 1061 to include some dilution on the edges of the domains. The small intervals (< 2.0 m) resulting from the compositing were merged with the preceding composite.

Table 14-9 and Table 14-10 summarize statistics by mineralized domain for length-weighted raw data and composited data.

The capping analysis was completed on composited data (4.0 m). The capping was assessed in two parts:

(i)       Theoretical – capping was assessed via several methods: log probability plot, histogram and disintegration analysis. Then, the different outcomes were averaged. In parallel, capping analysis was completed on the raw assays. The output statistics, of the 4 m composites capped on raw assays, were compared to the 4 m composites capped on composites. Both distributions are similar, which gives confidence in the capping values chosen. For each domain, the 3D distribution of the cut values was reviewed to ensure that the capped samples were not clustered and represented true outliers.

(ii)       Reconciliation with Ground Truth model - The capping values in part (i) were later reviewed and homogenized to the values presented in Table 14-11. As a result, 929 gold assays were capped. These values represent 0.79% of the entire sample population.

The capping was implemented by capping assay grades to the selected values after compositing. The raw assays statistics are presented in Table 14-11. The results of composting are presented in Table 14-12.

Contact profiles were generated to confirm the grade interpolation limits along the domain contacts. Contact plots were generated between all domains.

Based on the analysis, all contacts were treated as hard boundaries, with the exception of the contact between the Sediments FW and Sediments FW – indicator solids (1041 and 1061) that was treated as semi-soft. That is, blocks within the indicator solids were estimated using composites from the indicator solids and surrounding Sediments FW domain. Whereas, blocks within the Sediments FW domain were estimated using composites from that same domain only.

Variography was performed for each domain individually with the exception of domains 1041 and 1061. Both were modelled with domain 3 to ensure enough data support. Variography was completed in Isatis. Experimental variograms were calculated from the capped declustered composites and the modelling was conducted on non-transformed data. The variograms have been modelled with as few structures as possible – a nugget and 2 spherical structures. The nugget was obtained from down hole variograms of the 4 m composites. The results are presented in Table 14-13 and Table 14-14. For each domain, the orientation of the variography was checked visually against the mineralization’s orientation.

The West Branch block model was constructed in the Tasiast Mine grid system and is a regular (X=10, Y=10, Z=10) non-rotated block model. The model was built using a combination of Micromine and Isatis for wireframe coding, geostatistics and estimation. Elevation of the WB model is using the mining datum, such that Z = Z UTM real elevation + 5,000 m.

PP 2013 model and WB 2019 model overlap between northing 73,510 m and 73,500 m. The WB 2019 model overwrites the PP model in that area.

Gold grades hosted within the GDI (domain 1) were estimated via Localized Uniform Conditioning (LUC).

The LUC method is a two-stage technique. UC provides the probability distribution of the SMU within the panels but not their individual spatial location within the panels. Then, the UC grade-tonnage curves are localized onto each SMU within the given panel.

Based on drill spacing, the panel size was chosen as 40x40x10 m (XYZ). The WB SMU is 10x10x10 m (XYZ)

Table 14-15 presents the interpolation strategy used for estimating the GDI mineralization (domain 1).

The remaining domains were estimated via Ordinary Kriging (OK). Table 14-16 summarizes the interpolation strategy that was used. High yield restriction was implemented to reduce the influence of high grade mineralization, known to be erratic, in the sediment domains (2, 3, 4 and 1061, 1041). Table 14-17 shows the parameters used.

After interpolation, estimated gold fields for all six domains were merged into one unique gold field and dyke dilution was applied using the fraction of the blocks intersected by the dyke solids with half the lowest detection limit seen at WB (0.0025 g/t).

A total of 26,940 density values have been collected for Tasiast. Density has been assigned in the block model based on lithology and oxidation. Statistics for each of the density domains are shown in Table 14-18.

The Mineral Resources are classified under the categories of Measured, Indicated and Inferred, in accordance with CIM Definition Standards (CIM, 2014). Classification of the resources reflects confidence in grade continuity, as a function of many factors, including primarily assay data quality, QA/QC procedures, quality of density data, and sample spacing relative to geological and geostatistical observations regarding the continuity of mineralization. Where available, reconciliation of the resource against production data was also used.

Measured resources were defined based on a nominal drill spacing of 35.0 m. Indicated resources were categorized based on a nominal drill spacing of 70-80 m down to 60.0 m where only sediment mineralization was present. The remaining interpolated blocks within 150.0 m of the nearest composite were classified as Inferred resources. Blocks estimated using data beyond 150.0 m were not classified as mineral resources.

Note: In 2019, the lower and upper transition domains were merged into one domain for the WB model update.

Plan View looking down at 5,005 m RL (mining convention) North Looking section at 71,535N. The resource pit outline is shown for reference as a grey line. The codes are: 1 Measured, 2 Indicated, 3 Inferred and 4 unclassified.

Declustered data via nearest neighborhood estimation were compared with the block model. Results are shown in Table 14-19. The difference between the declustered data mean and the estimated capped grade is 1% for the GDI mineralization (which contains the bulk of the metal). For the secondary domains – 2, 3 and indicators 1041 and 1061 – the difference is more significant and the estimate systematically understates the overall grade. This is an expected result and reflects the estimation methodology implemented. These domains have typically underperformed when mining and a more restrictive estimation methodology has been applied through High Yield Restriction. This method limits the spreading of HG and better predicts the erratic nature of the mineralization. The 11% difference for domain 4 applies to a very low grade. The comparisons were completed over a volume well informed by drilling.

Interpolated block grades, resource classification, geological interpretation outlines and drill hole composite intersections were verified on screen in plan and in vertical section. No issues were identified in the visual validations. For areas of dense drilling the estimated blocks are directly comparable to the composite values.

Notes: Red lines are for visualization only and represent 2.0 g/t HG shell (domain 10) for capping purposes. Orange line represents the GDI lithology solids (domain 1).

Swath plots are commonly used as a block model validation tool as they provide a graphical comparison of key modeling outputs such as the number of composites, composite grades, raw grades, and block grades for various interpolation methods.

Swath plots (or trend reproductions) were produced for the block model and the declustered composites by domain. There are no issues with the swath plots, i.e., the composites and the estimates match in easting, northing and level.

The LUC estimate was compared (by domain) to the Discrete Gaussian Model (DGM)[5] and the OK estimate on SMU support (Figure 14-7). The DGM was prepared in Isatis using Point Anamorphosis (based on the capped declustered composites), the variogram model and an SMU assumption of 10x10x10 m.

The grade-tonnage curve comparisons between the LUC and DGM are within ± 5% for tonnes and grades at all cut-off between 0.0 and 2.5 g/t. As expected, the UC corrected the smoothing seen in the OK estimate and results are close to the “true” distribution (DGM).

Comparison of recoverable tonnage and average grade of gold between LUC (red lines) and direct kriging estimate at the SMU support (black lines) against the “True distribution” DGM (green lines).

The best method of validation is to compare the resource model to production. West Branch production history and the grade control practices routinely implemented there via reverse circulation drilling allowed reconciliation of the WB model against grade control data.

The WB resource model was coded by the end of month production shape for the two-year period January 2017- December 2018, and was reconciled against the grade control (GC) model for that period. In addition, it was reconciled against ground truth models built in parallel with the resource model using grade control data. The WB resource model was also compared to previous iterations of the resource model from 2013 and 2016. Figure 14-8 illustrates the evolution of resource modelling and shows the comparison against production data. The findings are as follows:

In summary, the 2019 resource model for WB aligns well with the grade control model at the economic cut-off and within the core of the high-grade mineralization. On the edges of the high-grade mineralization, it is expected that the WB resource model may underperform. This will be mitigated by grade control drilling the area well in advance of production and by using the GC model for short term planning.

The Piment and Prolongation resource estimate was prepared by Kinross Technical Services and T. Maunula & Associates Consulting Inc.. Geologic and estimation domains were prepared using Gemcom GEMS 6.4.1 desktop software. Sage was used for variography.

Before conducting statistical analyses, all data were imported into GEMS software (for Piment and Prolongation) and a check on the database was performed to search for any obvious errors, such as negative values and overlapping sample intervals.

A visual check of the drilling against the most recent topographic surface revealed that the majority of collars are set to the surface. However, some drill holes were noted above the topographic surface, and this was deemed a consequence of poor topographic resolution and the variable elevation arising from the mining operations.

Assay values less than the detection limit were assigned a value half of the lower detection limit value, which depending on the laboratory, was commonly 0.003 g/t Au or 0.005 g/t Au.

Three sets of wireframes were used for estimation: lithology, redox and mineralized domains. Lithology and Redox wireframes were generated in GEMS based on sectional interpretations every 25.0 m. All wireframes were based on logging information and pit mapping data where available.

The mineralized domains were the basis for grade estimation, the lithological wireframes were used for density, and the redox surfaces were used to assist with density assignment.

Table 14-20 summarizes the codes used in the block models and provides a brief description of each lithology.

Lithology, structure and grade were assessed for use as estimation domains. Key tools to evaluate the domains were exploratory data analysis, contact analysis and variography.

The mineralization domains were constructed using a 0.1 g/t Au threshold value. This value defines contiguous grade envelopes around the vein and structural styles of mineralization. Table 14-21 describes the mineralization domains.

Table 14-22 summarizes the redox codes that were used to code the redox variables in the Piment and Prolongation block models.

Figure 14-9 and Figure 14-10 illustrate the lithology and mineralization domains in plan view and section for both deposits.

The capping analysis was completed on raw length weighted assays. It included a visual review of the probability plots, a statistical assessment of the 97% and 99.9% percentiles, and decile analysis using the Parrish method. A summary of the analysis and the recommended capping levels is shown in Table 14-23. As a result, 154 gold assays were capped. These values represent 0.17% of the entire sample population. Capping was applied because a visual review of the outliers did not identify a spatial constraint or sufficient drill density to constrain the overestimation of the outliers.

Upon examination of the raw sample length statistics, a composite length of 2.0 m was chosen. The composites honoured the mineralized domain boundaries as presented in Table 14-24 and Table 14-25.

For PP, no soft boundaries were identified and only a single firm contact relationship was identified, as most of the domains were not in contact with adjoining domains. All contacts in PP were implemented as hard boundaries.

Variogram models were developed using SAGE2001 software. Directional sample correlograms were calculated along horizontal azimuths of 0, 30, 60, 120, 150, 180, 210, 240, 270, 300 and 330 degrees. For each azimuth, sample correlograms were also calculated at dips of 30 degrees and 60 degrees, in addition to horizontal correlograms. Lastly, a correlogram was calculated in the vertical direction (-90 degrees). The model was fitted to reflect geological knowledge and grade continuity of the deposit.

All conventions follow those of the Cartesian coordinate system. For example, the Z axis will be vertical with values increasing upward, if the system of axes is oriented so that:

A positive dip angle is measured upwards from the horizontal, whereas a negative dip angle is measured downwards from the horizontal.

The order and direction of the rotations around the three axes are given by the following (in each case the direction is given by the right-hand rule):

Table 14-26 summarizes the variogram results for the capped 2.0 m Au composites for PP. All correlograms were spherical models consisting of a nugget and two structures.

The Piment and Prolongation model was constructed in the Tasiast Mine grid system and is a regular (X=5, Y=5, Z=5) non-rotated block model. The model was built in GEMS for wireframes coding, geostatistics and estimation. Elevation of the PP model is using the UTM real elevation.

Piment and Prolongation block model was estimated via Ordinary Kriging (OK). Nearest Neighbour (NN) and Inverse Distance Squared (ID2) was also used to aid to validation.

For the ID2 and OK interpolation methods, two passes were used. For the first pass, a minimum of seven composites and a maximum of 18 composites were used and a constraint of three composites per drill hole was imposed. This had the combined effect of estimating all blocks with a minimum of two drill holes. For the second pass, the same composite selection parameters were maintained, with the exception of using a minimum of four composites.

For the NN estimate, one pass was used, which reflected the second pass interpolation parameters used for ID2.

Lithology and state of oxidation were used to determine the appropriate density values from Table 14-18. These values were then assigned to the domains in the block model using simple manipulation scripts in Gemcom.

The Measured resources were defined based on a nominal drill spacing of 25.0 m. Indicated resources were categorized based on a nominal drill spacing of 50.0 m, and the remaining interpolated blocks within 120.0 m of the nearest composite were classified as inferred resources. Blocks estimated using data beyond 120.0 m were not classified as mineral resources.

Plan view looking down at 5m RL and looking north section at 75,035N. The codes are as follows: 1 Measured, 2 Indicated, 3 Inferred and 4 Unclassified. The resource pit outline is shown for reference as a grey line in section.

The validation checks described for West Branch were also used for the Piment and Prolongation model.

The Mineral Reserve for the Tasiast open pit mine was estimated using a planning model derived from the 2019 resource models for West Branch (WB) and FS2013 resource model for Piment (PM), as discussed in Section 14.

The 24 kt/d pre-feasibility reserves, effective December 31, 2018, were estimated using the mine planning block model and applying a gold price of $1200/oz, a CIL processing cost of $23.18/t, selling costs of $63.84/oz and a base mining cost of $2.30/t, excluding incremental haulage. The reserve estimate includes material contained within the final pit design that can be extracted and processed economically. Reported reserves are solely based on the Measured and Indicated mineral resource classifications which correspond to Proven and Probable reserves classifications shown in Table 15-1.

An economic pit shell generated at a gold price of $1200/oz, with cost criteria, metallurgical recoveries, geologic and geotechnical considerations guides the final pit design. The economic pit shell used to define the final pit limits was created using Datamine’s NPV Scheduler software (NPVS). NPVS uses the Lerchs-Grossman (LG) algorithm to define blocks that can be mined at a profit. The program then creates an economic shell based on the following information:

The Mineral Reserve Estimate was prepared using the December 31, 2018 topography and the parameters detailed in Table 15-2.

Mineral Reserves are stated within an ultimate pit design at cut-off grades that are based on the process type, operating costs and metallurgical recovery.

Slope parameters based on geotechnical considerations were applied to the pit design along with ramps and geotechnical catch benches, and subsequently used to generate overall slope angles. The overall slope angles used in pit optimization are shown in Table 16-3 and Table 16-4.

Gold recovery is determined by ore type and process method. The gold recovery is calculated from the information in Table 15-3 where gold grade is expressed in grams per tonne (g/t).

The mine operating costs used for pit optimization include ongoing major mine equipment capital costs. The mine equipment sustaining capital was used in the economic model to simulate mine capital expenditures when generating the economic pit.

The top-down discount method was used during pit optimization. This is a procedure based on multiplying the block value by a discount factor that is a function of the annual cost of capital, an estimate of the average annual vertical advance rate of mining, and the relative depth of the block. This method simulates the actual mine plan discounted cash flow that is burdened with up front stripping costs and aids in the selection of a higher value pit.

John Sims, AIPG Certified Professional Geologist, has certified that, to the best of his professional judgment as a QP (as defined under NI 43-101), the Mineral Reserve and Resource estimates have been prepared in compliance with NI 43-101, including the CIM Definition Standards incorporated by reference, and conform to generally accepted mining industry practices.

The results of the economic analysis to support Mineral Reserves represent forward-looking information that is subject to a number of known and unknown risks. These uncertainties and other factors may cause actual results to differ materially from those presented here. Areas of uncertainty that may materially impact mineral reserve estimation include:

The Tasiast gold mine is located in northwestern Mauritania, approximately 300 km north of the capital Nouakchott and 160 km east-southeast of the port city of Nouâdhibou. The Tasiast Permit Area is in the Inchiri and Dakhlet Nouâdhibou Districts.

The mine site is located in a remote part of the Sahara desert, consisting largely of flat barren plains covered by stony surface with some sand and soil, interspersed with occasional sand dunes and upstanding outcrops of bedrock. The average elevation is approximately 130 masl (metres above sea level). Vegetation is sparse and consists primarily of grasses and occasional acacia trees. The climate is Saharan with an average rainfall of 30 - 40 mm per year, most of which falls from July to September. The climate is hot most of the year ranging from 10°C to 45°C and experiences strong prevailing NE-SW winds from the Sahara and occasional reverse SW-NE winds from the Atlantic.

The main ore hosting lithology is a Granodiorite Intrusive (GDI) with lesser contributions from the FVC (Felsic Volcaniclastic) and BIM (Banded Iron Magnetite). The orebody strikes at about 350° and dips easterly at about 50°, true width is about 40 m.

Ore and waste rock is mined in 10 m benches by conventional open pit methods primarily from the West Branch pit (Figure 16-1). Tasiast currently operates a load and haul fleet of 46 Cat 793D (220 t) trucks, five Komatsu 785 (92 t) and five Cat 6060 shovels plus two RH340B excavators. Blasting techniques, including presplit and buffer hole blasting, are employed to protect the pit walls. The grinding circuit produces a product size of 80% passing 90 microns which is processed in a conventional CIL circuit to produce gold bullion. Gold recovery averages 93%. Tailings slurry from the CIL process is currently pumped to the tailings storage facility 4 (TSF4).

Commercial production of gold at Tasiast began in January 2008, and 2,283 k oz has been produced up to the end of 2018.

Ore and waste rock is mined by conventional open pit methods from two pits (West Branch and Piment). Prior mining has taken place in West Branch, Piment and several other smaller pits at Tasiast. From late 2010 when Kinross acquired the property to the end of 2018, a total of 595 million tonnes of material have been mined from various pits, including 53 million tonnes in 2016, 75 million tonnes in 2017 and 87 million tonnes in 2018.

The current mill operates at approximately 15 kt/d. Ore is fed directly from the mine and stockpile to the primary crusher. Sub-grade material is stockpiled adjacent to the ROM (Run of Mine) pad for later treatment.

Cut-offs are based on the net block value and cut over grades applied during scheduling. Applying cut-over grades during scheduling ensures that the highest-value materials are routed to the CIL process over time. Lower-grade materials are routed to stockpiles. All materials below the cut-off are sent to waste destinations. The grades and potential destinations used for strategic planning are shown in Table 16-1.

Loading and hauling requirements will be met by maintaining the current existing operating mine fleet to support the processing expansion case to 24 kt/d. Estimates of future equipment utilization is based on current operating practices, general and site experience, and takes into consideration:

The existing shovel and haul truck fleets will be used for the duration of mining and no replacement of this equipment is anticipated. Equipment life has been projected from actual operating hours, with estimates of future usage based on the mine plan. The current mining fleet at Tasiast is shown in Table 16-2.

The Tasiast final pit designs consist of a series of pits that extend along a strike length of approximately 8 km. The configuration of the mining area is shown in Figure 16-2. Only West Branch is actively being mined during 2019.

Historically, the project has been broken into two geotechnical zones where the Piment Zone is north of approximately 72,000 N and the West Branch Zone is south of 72,000 N. Overall pit slope angles and inter-ramp angles for the Piment Zone were initially determined by Scott Wilson Mining UK (Scott Wilson) in 2009 and subsequently optimized by Stacey Mining Geotechnical Ltd (Stacey) in 2011. The slope angles that are applied at Piment are shown in Table 16-3.

Golder Associates, the engineer of record for Tasiast, generated the slope design parameters for West Branch as summarized in Table 16-4. These slope parameters are based on the assumption that appropriate practices in dewatering or depressurization, drilling and blasting and movement monitoring are effectively implemented and carried out.

Kinross Technical Services reviewed and recompiled Rock Mass Rating (RMR) data in 2017. The data indicated that the rock mass quality in the Lower Transition zone is in the same range as that in the Fresh zone as shown in Figure 16-3. After consultation with Stacey Mining Geotechnical Ltd, a modification to the Bench Face Angle (BFA) for the Hanging Wall was recommended and the proposed BFA was 70o with a calculated Inter-ramp Angle (IRA) of 52o.

The basis for the final pit design is an optimized economic shell generated using the NPVS software package. The economic pit shell used to define the final pit limits was created using Datamine’s NPV Scheduler software (NPVS). NPVS uses the Lerchs-Grossman (LG) algorithm to define blocks that can be mined at a profit. Cut-off grade and pit limits developed in NPVS were defined using the criteria outlined in Table 16-5. Adjusted overall slope angles were used to define the slopes in NPVS. This adjustment is used to address the placement of ramps within the mined area. Only measured and indicated resources were used to define this limit. Sensitivity analyses were carried out comparing the affects of high and low ranges for various inputs on ore tonnes and contain gold oz. Inputs tested were gold price, mining cost, processing cost, recovery and slopes. See Figure 16-4 and Figure 16-5 for the sensitivity spider charts.

Inter-ramp slopes – Inter-ramp slopes are based on geotechnical recommendations outlined in Subsection 7.2 and vary according to the slope sector involved. Inter-ramp slope angles range from 31° to 55° based on the criteria. The inter-ramp angles and the bench face angles were adhered to for the pit and phase designs. Catch benches for the design vary based on bench face angles and ultimate bench height. The overall design slopes include access ramps and follow the same criteria used in the LG cone calculation.

Bench height – The design operating bench height is 10m. The final pit walls in West Branch will be triple benched where it is permissible, resulting in a bench height of 30m with intervening catch benches. Piment will be double benched, resulting in a 20 m bench height with the appropriate catch benches.

Minimum mining width – A phased approach was taken as an optimization strategy to improve the mining sequence. Efforts were made to maximise mining width where possible. Where mining widths indicated by the selected LG shells were too narrow to safely or effectively mine, the phase walls were pushed out to the final pit design.

Access – Dual lane haul roads for the current design are 32.5 m wide in West Branch. Ramps are designed on a 10% grade and do not exceed that grade. Intersections and switchback curves are designed without grade (flat) wherever possible. Two main haul roads exit the West Branch. These haul roads split to minimise haul times between different destinations. A third ramp exits the pit at the northern end of the WB pit. Haul road profiles are shown in Figure 16-6. Haul road dimensions are shown in Table 16-6.

Detailed pit design work was completed using MineSight software. This MineSight Pit Design Tool uses design parameters contained within the 3D block model to ensure compliance to the geotechnical parameters. Final designs have remained unchanged from those designs used to report mineral reserves as at December 31 2018 in March 2019.

A cross-section of the West Branch pits is shown in Figure 16-7. This section illustrates how the pushback sequencing targets high ore zones while minimizing the waste stripping requirements during the early pushbacks.

Waste rock is used for haul road and tailings dam construction as needed. The existing road network is well developed, and requires continued maintenance. Additional roads will also be required throughout the life of the mine. These roads will be constructed using the current mining and support mining fleets.

Dumped waste material comprises weathered and unweathered rock. Blasted weathered rock is finely graded, including clay-silt fines. The unweathered rock is strong and massive and, when blasted, is coarsely graded, including boulders that can require secondary blasting prior to loading.

As the climate is arid and there is no permanent surface water and very limited groundwater, there is low potential for acid rock drainage (ARD). However, any potential ARD issue will be mitigated by ensuring that material that is identified as potentially acid forming (PAF) will not be dumped on the outer shell of the waste dumps.

The dump design is based on 20 m high lifts with a maximum overall effective slope of 2H:1V (27° Overall slope angle). The maximum dump height is currently limited to 100m total vertical height. The dumps will be accessed by 35 m wide dual-lane ramps at a 8% gradient. Track bulldozers will be used to assist the haulage fleet to facilitate proper dump construction, including grading the top of each lift away from the pit to direct any rainwater run-off and placing coarsely-graded, unweathered rock on final dump faces.

The waste dumps are located within the footprints of previous studies and require no modification to the current permitting.

There are two waste dumps along the east and west sides of the Piment Pit and two waste dumps along the east and west side of the West Branch pit (Figure 16-2). Based on the 24k t/d study mine plan, the waste dumps will require 400 Mt of capacity after 2019. The current permitted waste dumps have excess capacity relative to the current mine requirement.

Phased pushbacks were developed to optimize the mining sequence (Figure 16-8). Five pushbacks were developed for the West Branch pit, WB1 and WB2 have been completed, and WB3 will be completed in Q3 2020. Piment consists of two phases located around the current Piment Central Pit.

The mining rate was optimized to feed the CIL plant at the rates outlined in the project schedule. The plant expands from the current capacity of approximately 15 kt/d (Phase One expansion completed in 2018) to 21 kt/d at the end of Q4 2021, and 24 kt/d at the end of Q2 2023. The forecast mining rate is 50 Mt/a and 55 Mt/a in 2022 and 2023 respectively increasing to approximately 79 Mt/a in 2024 (Figure 16-9). This mining profile delays waste mining and leverages fleet capacity in 2024. Material is reclaimed from stockpile as required.

The distribution of equipment across the pits, phases and benches results in variable production rates using the same fleet. As such, the total material moved varies year over year (Table 16-8). The mining schedule was optimized with a maximum of 45 Caterpillar 793 haul trucks and a maximum of 5 Komatsu 785 haul trucks.

The life of mine production schedule is shown in Figure 16-10 and Table 16-9. Mining of the West Branch pit continues throughout 2026. Mining in Piment is resumed in 2028 and ends in 2031. The 24 kt/d expansion has an overall strip ratio of 6.5:1.

The process plant currently operates at 15 kt/d, with the plant expected to reach 21 kt/d in Q4 2021 and then 24 kt/d in mid 2023. Table 16-10 shows the process plant feed schedule by year.

Once mining operations have been completed in 2031, the CIL plant will continue processing the low-grade stockpiles that will have developed over the course of mining.

The stockpiles will be used in conjunction with the operating cut-off grade strategy to add value to the project and aid in sequencing the pit. Generally, the stockpiles will be reclaimed during periods where the material mined in the pit is of lower grade than the stockpiled material. The high-grade stockpile will be reclaimed first, followed by the medium-grade stockpile. The low-grade and marginal stockpiles will be consumed at the end of the mine life. Stockpile evolution over LoM is presented in Figure 16-11.

Selection of the mining equipment at Tasiast is based on the current mining fleet on site. The larger fleet of equipment includes Caterpillar 6060 excavators (also referred to as RH340s) paired with Caterpillar 793D haul trucks. The smaller fleet consists of Komatsu PC1250 excavators in a backhoe configuration. Komatsu HD785 trucks are matched with the PC1250 excavators and will be used at West Branch in narrower sections of pit cutbacks, at the bottom of each pit stage and in the Piment phases.

Payload sizes were estimated for the study, based on analysis of actual payloads. Payloads estimates were not varied by rock type or time, and represent the average payload size expected, see Table 16-11.

The LoM schedules provide details regarding the tonnages moved by material type and destination (i.e. waste dumps, dump leach pads or primary crusher). During LoM scheduling, haulage analyses were performed using Hexagon MineSight software to estimate the required number of truck hours.

Haulage speed estimates used the speeds displayed in Table 16-12. These speeds have been benchmarked from other Kinross sites and adjusted to as necessary to reflect operating conditions at Tasiast. These speeds were applied for both Cat 793 and Komatsu 785 haul trucks.

Complete cycle time estimates were calculated for each discrete pit cut to each required material destination. This cycle time estimate includes travel time, loading time, shovel spotting time, and dumping time. Loading time estimates are calculated using dig-rate estimates for the different primary loading units used in the mine plan. Variable dig rates are applied in oxide, transition, and fresh materials. These dig rates are used to calculate load times for paired trucks. Dig rate estimates are derived from current rates and are consistent with the site’s budget estimates. Dig rate estimates are shown in Table 16-13 through Table 16-16.

A fixed 1.5 minutes allocation for dump time is applied to each haul cycle. The fixed non-travel time is based on current site data. Truck requirements and productivity by period were calculated based on the quantity of material moved and the cycle times associated with each material. The breakdown of the fixed cycle times is shown in Table 16-17.

As the mine sequence progresses, the pit becomes deeper resulting in longer travel times to reach the pit exit point. At the same time, the primary waste dump destinations will be filled, and longer routes will be required to reach their tops. The crusher location will remain fixed during the project. The net of these effects over time is that the average cycle time generally increases over the mine life. Figure 16-12 shows the truck fleet requirement over the LoM (including stockpile reclaim). Figure 16-13 shows CAT-793 operating truck hours by year.

Estimation for the loading equipment requirements is based on the material movement specified in the LoM schedule and the estimated productivity rates. The productivity rates for the loading units are based on loading cycle times, bucket capacity, bucket fill factors and historical productivity rates.

The fleet size was determined based on anticipated operating hours for the loading equipment and machine life estimates provided by equipment manufacturers, along with benchmarking data. The total fleet requirement was estimated by applying the percentage of mechanical availability and usage to the operating fleet requirements. Loading fleet requirements by year are shown in Figure 16-14.

Blasting requirements determine the drill-hole pattern size (i.e. burden and spacing) which provides the information to estimate the metres of drilling and hours required to achieve planned production. The plan assumes that the Sandvik DR580 drills will be used predominantly for buffer holes, trim shots and presplit drilling. The larger Bucyrus SKFX and Atlas Copco PV 235 drills will be used for waste and ore production drilling.

The blast design parameters and pattern designs match current practices at the site, as detailed in Table 16-18 and Table 16-19 for Small (DR580) and large (SKFX and PV 235) drills respectively. Penetration rates for each drill by material type are illustrated in Table 16-20.

The fleet size was determined based on estimated drilling hours, machine-life estimates provided by equipment manufacturers and benchmarking data. The total fleet requirements were estimated by applying percent mechanical availability and utilization to the operating fleet requirements.

The operating plan for the different drilling equipment was provided to a zero based economic model where operating costs were applied based on site performance to date and reasonable improvement projects. Blasting costs were developed, bottom up using blast designs outlined in Table 16-18 and Table 16-19 and quantities of presplit and buffer drill holes.

The mine equipment fleet will need various support equipment for constructing and maintaining roads, waste storage dumps and dump leach pads. Support equipment fleet requirements depend on infrastructure maintenance requirements, and the number of shovels and excavators operating in the pit. Typically, these estimates were based on operating experience and benchmarking data.

Costs and equipment hours for support tasks were not calculated in detail. Annual auxiliary equipment hours were estimated based on historical performance and benchmarking of operations with comparable material movement rates and mining equipment fleets.

A modular mining dispatch system with high-precision GPS is used for haul truck dispatching. Slope monitoring is carried out with portable slope radar and survey equipment.

The current mine operation at Tasiast is owner-operated and applies conventional open-pit operational practices, with drilling, blasting, loading, hauling, support and administrative functions. The mine operates 365 scheduled days per year and 24 hours per day, primarily divided into two 12-hour shifts per day for mine operations and mine maintenance.

The mine organization includes functional groups for mine operations (drilling and blasting, loading and hauling), maintenance, mine technology and technical services (Table 16-21). The mine is staffed to support all operational, safety and environmental requirements. Mining-related functional groups are organized under the mine manager or technical services manager. The mine manager is allocated functional groups for mine operations, maintenance and mine technology. Among the functional groups responsible for mine operations, drilling and blasting are managed together, as are loading and hauling. The technical services manager oversees functional groups for technical services, including mine planning, survey, geology and geotechnical services. The mine manager and technical services manager collaborate to manage mine operations, with each reporting to the operations director at Tasiast.

Mining cost estimates for the expansion include appropriate staffing levels for mine operations, based on present staffing levels for the existing organization and future mine plan requirements. Expansion requirements for equipment operators and mechanics, along with the requisite supervisors and support equipment operators, have been projected in line with the vehicle operating hours required to realize planned tonnage as per the expansion case mine plan. Expansion mining cost estimates have been benchmarked against mines similar in scope to the expanded Tasiast mine.

Table 16-21 shows staffing levels for the mining area by functional group. The expansion of the plant requires an increase in personnel from 2021 to 2022 to accommodate the increase in mined tonnage. Based on this mine plan staffing levels are predicted to remain constant through until 2031 when mining is finished in West Branch. Table 16-22 presents the reduction of expatriate personnel at site.

The existing Tasiast CIL plant has proven capable of processing approximately 15 kt/d. Kinross has decided to increase the existing CIL plant capacity in stages through a de-bottlenecking exercise. The modifications to achieve and initial target of 21 kt/d and later 24 kt/d form the basis of this project and are detailed below.

Currently, two waste dumps along the east and west sides of the existing Piment Pit and the west side of West Branch pit are used to dispose of waste rock.

A brackish water bore field (known as “Sondage") provides raw water for the mine, processing plant, dump leach and camp. The site potable water requirements are met by three reverse osmosis plants.

For the initial 21 kt/d expansion, pumps at the existing Sondage pump station P1 will be upgraded. Pump stations, P2 and P4, will also be added along the existing 400DN PVC pipeline. For the 24 kt/d expansion, two additional pump stations, P3 and P5, will be added on the existing 400DN PVC pipeline as well. Each pump station will consist of a tank, operational and standby pumps and ancillaries. Several additional wells will be added at the Sondage in order to provide additional capacity and flexibility for the operation.

Internal process water recovery from the thickener overflow, reclaimed water from the tailings storage facility and recycled grey water from the camp are used to reduce the overall plant raw water requirements.

The existing CIL processing plant produces the majority of the gold shipped from Tasiast. The remaining gold is produced from the dump leach operations via a dedicated ADR plant. All produced gold is in the form of bullion and is transported regularly to a refinery for final refining and sale.

Unit gold recovery estimates for the existing plants are based on metallurgical test work and a review of historical performance. The recovery varies by mineralization type, lithology and grade. Recoveries by treatment method are presented in Table 17-1.

The dump leach operation is nearing the end of its life. Rinsing of the pads is expected to begin in early 2020 and take approximately two years to complete before closure.

Previously, the DLP was designed to process up to 11 Mt/a of low-grade oxide mineralization using two dump leach operations: one pad (Piment) with five cells separated by a raised berm for solution drainage control, and the other pad (West Branch) with 8 cells. The design of each pad allowed for three 10 m lifts for a final stack height of 30 m. All solution collection ponds are plastic-lined with installed leak detection systems and bird netting protection. The gold-containing (pregnant) solution produced by the dump leach operations is pumped to a dedicated ADR plant to recover the gold and reactivate the carbon.

Three ponds adjacent to each of the two pads are provided for storage and management of the barren, intermediate and pregnant solutions. At each of the two pads, barren solution is pumped from the barren pond to irrigate the heap. The first drainage collection of “intermediate” solution is collected in a pond, then pumped to another section of the pad to contact freshly placed ore. The resulting drainage is captured in the pregnant solution pond adjacent to each pad and pumped at a controlled fixed rate to the ADR plant for gold recovery. Make-up water is added to the systems from the raw water pipeline system connecting the bore field with the CIL mill. The lime added during stacking maintains a minimum pH.

The pregnant solution produced by the dump leach operations is received in a single train of six carbon columns at the ADR plant, as seen in Figure 17-1. The pregnant solution flows by gravity from column to column, contacting carbon in each column as it rises from the bottom to overflow. The activated carbon granules adsorb the dissolved gold from solution. The solution leaving the last column is barren solution and is pumped to the dump leach operations for re-use. When the carbon in the first column loads with gold to the target level, the first column’s contents are advanced by airlift to a screen and discharged to an acid wash vessel. Sequentially, each following column is advanced counter-current to solution flow.

6 Head grade vs. recovery relationship develop from testwork. Au recovery = 1.2579*ln(head grade, g/t) + 91.838

After acid washing, the carbon is advanced to the elution vessel, where the same AARL process as in the Plant elution system is used to elute gold from the carbon. The gold rich “pregnant” solution is stored in a surge tank prior to electrowinning. After elution of the gold, the barren carbon is re-activated thermally in a diesel fired rotary kiln, stored in a surge bin and eventually returned to the column train via the last tank in series.

Gold is recovered from the pregnant solution by electrowinning onto stainless steel wire wool cathodes. The gold is washed from the cathodes after each elution cycle, dried in an oven, mixed with fluxes and melted in a crucible furnace, all within the gold room attached to the ADR plant. Doré bars of 94% or higher purity are produced and transported to a commercial refinery for further refining and sale.

The existing Tasiast CIL plant was recently expanded in 2018 to 12 kt/d. However, the operation has proven itself capable of processing approximately 15 kt/d. The process schematic for the existing CIL plant is shown in Figure 17-2.

Mined ore above plant feed cut-off grade is transported from the open pits to the plant by truck and either deposited onto the ROM pad or directly dumped into a Gyratory Crusher. The primary crushing facility is shown in Figure 17-3. The material can be blended according to grade and competency.

Crushing of the mineralized material takes place in a single stage; a primary gyratory crusher that reduces rock to less than 210 mm. The rock is conveyed to a coarse ore stockpile shown in Figure 17-4 that uses three apron feeders to feed the 40 ft SAG mill.

The SAG mill works in closed circuit pebble crushing with a circulating load of up to 40%. The SAG mill is also in closed circuit with a cyclone cluster. Cyclone overflow is pumped to two, parallel Ball Mills. The Ball Mills are in closed circuit with cyclones. The target grind size is 80% passing 90 µm. Cyclone overflow containing solids of the required size flows to linear trash screens prior to entering the leaching circuit. Coarser solids are returned to the mill for further grinding, with a bleed to a gravity circuit to capture liberated, coarse gold particles.

Two Knelson concentrators installed within the Ball Mill circuits recover the coarse gravity recoverable gold. The concentrate is collected and treated separately in an Intensive Leach Reactor where the gold is dissolved and the resulting pregnant solution is pumped to the gold room for electrowinning and smelting.

The leaching circuit at Tasiast comprises four leach tanks and six CIL (carbon-in-leach) tanks where the dissolved gold is recovered on activated carbon. Material that exits the grinding circuit has an approximate slurry density of 50% solids, by weight. The slurry gravitates to the first of the four agitated leach tanks via launders. Lime is added to increase the slurry pH to 10.2, and then dilute sodium cyanide solution is added to maintain a fixed cyanide concentration. Oxygen gas is injected into the tanks to enhance gold leaching kinetics.

Activated carbon granules contact slurry in the CIL tanks to adsorb the dissolved gold from solution. Carbon that has achieved the target gold content, typically around 2,500 g/t, is termed "loaded" and is transferred daily to the elution circuit for recovery of the gold on a batch basis. After each such transfer, carbon in the remaining tanks is advanced counter-current to the slurry flow on a sequential basis, and fresh or “barren” carbon is added to the last CIL tank to maintain the carbon inventory.

After maximising gold recovery from the solution and ore particles in the CIL process, the resulting slurry flows via carbon retention screens to the thickener, where the solids are settling to achieve a density of approximately 60% solids and residual solution is returned to the process.

Loaded carbon recovered from the CIL slurry by screening is first water washed to remove entrained ore particles and then washed with hydrochloric acid solution, in a dedicated acid wash vessel, to remove inorganics from the carbon surfaces. The acid-washed carbon after being neutralized is transferred to the elution pressure vessel. To recover gold from the loaded carbon, batches of approximately 7.5 tonnes of carbon are subjected to a high pressure and temperature, stripping process, called elution. Tasiast uses the Anglo American Research Laboratories (AARL) strip process. A hot caustic and cyanide rinse under pressure removes the gold from the carbon and into solution. After gold removal, the "barren" carbon is transferred to a regeneration kiln for thermal reactivation of the carbon. Reactivated carbon is returned to the last CIL tank. Gold is recovered from the caustic solution by electrowinning onto stainless steel wire wool cathodes in electrowinning cells, located within the gold room. The gold is removed as a sludge by pressure washing the cathodes at intervals. The sludge is dried and mixed with fluxing materials and charged to a diesel-fired crucible furnace. After melting, the slag is poured and followed by pouring the gold into bullion moulds. Bullion or “doré” contains approximately 94% or higher gold content together with a minor content of silver. The doré bars are transported by a security firm to a commercial refinery for further purification and sale.

Tailings from the CIL process plant are treated with a ferrous sulfate solution to detoxify the residual cyanide before being pumped to a tailings storage facility (TSF). A new facility (TSF4) was constructed and commissioned in 2018. After settling of the solids, a portion of the contained water drains to a collection area within the storage basin, from there it is returned to the plant process water system. Solids are retained in the TSF.

The existing plant has proven capable of processing approximately 15 kt/d. As a result, Kinross has decided to increase the existing plant capacity in stages through a de-bottlenecking process. The modifications to achieve the new target of initially 21 kt/d and later to 24 kt/d form the basis of this project and are detailed below.

A simplified overall process flow diagram for the initial 21 kt/d expansion is illustrated in Figure 17-6 and 24 kt/d in Figure 17-11.

The design capacities for the crushing plant and process plant use 70% and 92% effective operating time, respectively. Based on test work and operating experience, the key nominal design criteria for the major process circuits are summarized in Table 17-2. It should be noted that design factors accounting for higher gold grades were applied when required to match the mine plan. The plant design life is 15 years.

The Tasiast Expansion Project is a brownfield expansion with the new processing sections located within the current plant boundaries. Figure 17-5 shows the 24 kt/d CIL Plant Layout.

The environmental and plant conditions will be extreme and will require careful selection of construction materials to suit the 15-year design life of the facility. Erosive process slurries will cause wear to plant equipment and piping. Strong acids and bases, cyanide and the use of brackish ground water as process water and wash water will expose most of the wetted equipment to corrosion conditions. Reference documents for guiding the selection of construction materials were provided in appendices in the 2014 Feasibility Study Report by Hatch on the Tasiast Expansion Project.

All wet areas of the process plant will be bunded with a containment volume equal to 110% of the volume of the largest tank in the containment area, or 25% of the total combined tank volume in the case of hazardous materials.

Areas with specific requirements (design code requirements or incompatibility of solutions, such as hydrochloric acid and sodium cyanide) will be provided with separate containment.

The CIL plant and Dump Leach operations are both designed and operated to International Cyanide Management Code (ICMC) standards..

The primary crushing facility shown in Figure 17-3 ore storage shown in Figure 17-4 will both remain unchanged.

The grinding circuit will consist of the existing SAG mill and two existing Ball Mills operating in closed circuit, each with a hydrocyclone cluster.

The gravity recovery and intensive leach circuits will consist of two XD-48 centrifugal gravity concentrators with feed-scalping screens, concentrate hoppers and two skid-mounted intensive leach units. The equipment will be located on the east side of the grinding area (Figure 17-7).

A portion of the hydrocyclone underflow from the grinding area will be directed to two scalping screens to remove coarse particles.

Gravity concentrate from the centrifugal concentrators will be batch processed in the intensive leach circuit.

At the completion of the batch leach cycle, the resulting gold-rich pregnant solution will be pumped to the existing ILR pregnant-solution tank in the vicinity of the gold room. One gravity concentrator and both ILR units are already owned by Kinross and will be relocated.

The existing linear trash screens will remain unchanged. The oversize rock particles and trash from the screens will be collected in a bunker and periodically picked up and trucked to the tailings area for disposal. The undersize slurry from the screens will be combined in a new feed box. Various streams feeding the current distribution box, (lime, spillage etc.), will be re-piped to the new feed box. A pipe launder from the new box will transfer the slurry to either of the two new leach tanks

The leach tanks will be mechanically agitated tanks (18 m diameter by average 17.5 m high) similar to those installed in the “Phase 1” (12 kt/d) project.

Slurry will gravitate to the first leach tank and will be leached in a weak cyanide solution to dissolve gold. The leach circuit will increase gold concentration in the solution before contact with activated carbon in the CIL circuit. The leach circuit has been designed to provide approximately 9 hours leach retention time, with an additional 9 hours in the CIL tanks for a total of 18 hours. Oxygen will be sparged at the bottom of the leach tanks to speed kinetics.

Leached ore slurry from the last leach tank will flow by gravity to the CIL circuit. Dissolved gold and silver will be adsorbed onto activated carbon particles in the CIL tanks. The CIL circuit has been designed to provide approximately 9 hours of slurry retention time. To achieve this, the existing pre-leach tank will be modified to a CIL tank. This will involve the addition of two, 8.5 m2 area, NKM-type interstage screens and a new carbon transfer pump (Figure 17-8). The existing 7 m2 screens in the existing CIL tanks will be modified to 8.5 m2 units.

Slurry from the last CIL tank will gravitate to a new CIL safety carbon screen feed box in the new tailings thickening area. The CIL safety carbon screen feed box will distribute the slurry between two vibrating screens for the recovery of fine carbon. Carbon recovered on the carbon safety screens will gravitate to bulk bags. Underflow slurry will gravitate through a tailings sampler and will be pumped using a new pump box and pumps to the new tailings thickener.

The tails thickener is used to recover water from the CIL tailings slurry to reduce plant feed water requirements as well as reduce cyanide detoxification requirements. The thickener will have a diameter of 40 m, with underflow density controlled at approximately 60% solids, by weight. The layout of the tailings thickener is shown in Figure 17-9 below.

Tailings from the CIL process plant will continue to be treated with a ferrous sulfate solution to detoxify the residual cyanide before being pumped to the tailings storage facility. The slurry will be pumped from the tailings thickener using the existing tailings line. An intermediate tails pumping booster station (TP3) will be required in order to maintain allowable pipeline pressures. This will consist of two-stage, in-line, duty and standby pumps, complete with a gland service water tank and pumps. A spillage sump pump will discharge spillage into a dedicated tank, from which it will be pumped directly to the TSF. A dedicated pump on the gland service water tank will supply flushing water to the spillage tank.

The tailings pipeline and the return water line are contained in a trench connected to five existing emergency ponds. In case of emergency or a prolonged shut down period, tailings slurry can be discharged into these ponds by operating the manual valves provided to prevent line blockage.

The tailings pipeline design includes flushing points at regular interval along the length of the line to the TSF for manual line flushing. This is achieved by connecting a flushing hose between the slurry line and the duty tailings return water line.

A new tailings storage facility (TSF5) will be constructed adjacent to the existing TSF4. Both facilities will eventually be joined by a common wall and become one large tailings containment, referred to as TSF4/5.

There will be no changes to the existing carbon elution circuit. However, two additional electrowinning cells and associated ancillaries will be installed within the gold room.

The current oxygen generation plant has three VSA oxygen generation modules. These provide the oxygen gas sparged into the bottom of the leach tanks. A fourth VSA module, booster compressor and air receiver will be added to increase the oxygen plant capacity.

The current sodium hydroxide facility will be demolished to provide a site for new the CIL area tower crane. The existing cyanide facility will be converted to a new sodium hydroxide system. Existing sodium hydroxide pumps will be relocated. One existing cyanide transfer pump will be retained to provide caustic to the gravity recovery circuit.

A new sodium cyanide facility will be constructed in the current lime bag storage area (Figure 17-10). Sodium cyanide will be supplied in 1 t boxed bags as solid briquettes and dissolved in treated water to make a 20% w/w solution in the 50 m3 mixing tank. Sodium hydroxide will be added to safely dissolve the solid sodium cyanide in a high pH solution. During dissolution, the solution will be maintained at a pH greater than 12 to avoid volatilization of hydrogen cyanide gas. The cyanide will be pumped into the existing ring-main. A dedicated pump will service the gravity recovery area.

The primary crushing facility shown in Figure 17-3 ore storage shown in Figure 17-4 will both remain unchanged.

The grinding circuit will consist of the existing SAG mill and two existing Ball Mills operating in closed circuit, each with a hydrocyclone cluster.

One additional leach tank will be added for 24 kt/d. This leach tank will be a mechanically agitated tank (18 m diameter by average 17.5 m high) similar to that installed in the 21 kt/d expansion. The pipe launder from the trash screen underflow box will be extended to this new leach tank.

Slurry will gravitate to the first leach tank and will be leached in a weak cyanide solution to dissolve gold. The leach circuit will increase gold concentration in the solution before contact with activated carbon in the CIL circuit. The leach circuit has been designed to provide approximately 10 hours of leach retention time, with an additional 8 hours in the CIL tanks for a total of 18 hours. Oxygen will be sparged at the bottom of the leach tanks to speed kinetics.

Leached ore slurry from the last leach tank will flow by gravity to the CIL circuit. Dissolved gold and silver will be adsorbed onto activated carbon particles in the CIL tanks. The CIL circuit has been designed to provide approximately 8 hours of slurry retention time. A third, 7m2, interstage screen will be added to each CIL tank.

Tailings from the CIL process plant will continue to be treated with a ferrous sulfate solution to detoxify the residual cyanide before being pumped to the tailings storage facility (TSF).

When upgrading to 24 kt/d, a second tailings pumping booster station (TP2) will be required. This will consist of single-stage, in-line, duty and standby pumps, complete with a gland service water tank and pumps. A sump pump will discharge spillage into the dedicated tank at TP3, from which it will be pumped directly to the TSF.

The project schedule is designed to incrementally alleviate each bottleneck in order of priority to first achieve 21 kt/d, and then advance to 24 kt/d. Commissioning will begin during the final stages of construction in each area. As the construction of a section of the plant is completed, the section is handed over to commissioning. Having both construction and commissioning underway at the same time minimizes the delay between final construction completion and the start-up of the plant. Each project area will have a detailed commissioning and tie-in plan prepared to minimize operational downtime.

Raw water for the Tasiast site is from a water supply bore field, which is located 64 km west of the mine, and draws from a brackish aquifer using a system of 43 wells. Individual well yields range from 340 to 1,000 m3/d as determined during pump testing completed in June 2015. Individual wells within three separate well areas are combined in a manifold for each area and fed to a primary pumping station located at a facility referred as the Sondage. Water from the Sondage is transported to site via pipelines with booster stations downstream. Currently, the bore field is capable of supplying up to 22,000 m3/d of raw water. However, the existing pipelines are limited to approximately 16,000 m3/d of raw water to the site based on the capacity of the pumping system.

The Tasiast permit, issued May 7, 2017 by the Ministry of Hydraulics and Sanitation, allows abstraction at a maximum rate of 30,000 m3/d through to December 31st, 2034. Modeling done by Piteau Associates, and calibrated against monitoring data, confirms the Sondage can be operated at this maximum rate and continue to meet the permit conditions.

Work by Schlumberger Water Services (SWS) to evaluate the abstraction capacity of the bore field concluded that the field can provide the water required for the expansion. SWS has also indicated that potential water supply issues may be identified early through proper monitoring, providing a lead time of at least three years to design and implement remedial measures.

The project infrastructure required for the proposed 24 kt/d expansion will include the addition of booster stations along the existing PVC pipeline. The new booster stations will increase the total pumping capacity of the Sondage raw water system to approximately 20,000 m3/d. Several new wells are also planned to provide additional capacity and flexibility to the Sondage operation.

Reverse osmosis (RO) water treatment plants and storage basins/tanks are located at the mine site. Saline water produced from the RO plant is used to water the haul roads or used in processing. Potable water is produced from RO water following additional disinfection steps. Potable water is also used for domestic purposes at the Tasiast site.

The Phase 1, TTV and Phase 1B power plants are all interconnected to and able to feed onto the 33 kV distribution system to supply the required site loads. The above power supply totals 42 MWe, which is more than the approximately 30 MW required for approximately 15 kt/d production rate while providing redundancy. As the production is ramped up, the power requirement will also increase. In order to meet this load and to reduce the dependency on higher price LFO fuel (compared to HFO fuel), a rental modular power plant is currently being constructed at site to be operational by January 2020. This modular power plant will consist of 8 MAN 9L21 engines (HFO fired with LFO back-up) x 1.6 MW each for a plant capacity of approximately 13 MW. The rental approach was selected to meet the immediate need as it has a relatively quick deployment time (about six months) while a long term solution was being evaluated and contracted.

For the long term need for 24 kt/d, there will be a new Phase Two power plant located adjacent to the Phase 1B plant, with an expected in-service date of Q1 2021. This will consist of 4 Wärtsilä W20V32TS simple cycle, medium speed, reciprocating engines (HFO as primary fuel and LFO as back-up) each of approximately 10 MW capacity. They will be located in a totally enclosed new engine hall with an overhead crane for service and maintenance. The peak load at 24 kt/d will be approximately 50 MW. With the addition of the Phase Two power plant, the combined site HFO generation capacity will be approximately 65 MW (Ph 2, existing Ph 1 B, existing MaK, but without the modular rental units).

The combined plants will provide the ability to meet the required net peak power demand at any expected ambient condition, while accounting for equipment fouling, ageing, power plant parasitic loads as well as spinning reserve requirements. The combined facility will have N-1 redundancy, meaning that the facility will still be able to meet the maximum site power demand with the loss of availability of one generator set (considering both scheduled and forced maintenance). For further backup, sufficient LFO engines would be kept in service. It is the intent to phase out the 13 MW rental power plant after the Phase Two power plant is in-service.

The power plants will operate 24 hours per day, seven days per week and 365 days per year as an islanded operation. At this time, there are no plans to connect to the national electric grid, although that could be an option in the future.

Waste from plant and equipment maintenance, construction, offices, kitchens and accommodation is processed at the waste management facility where materials are sorted for reuse, recycling, or incineration. Composters are also used in the camp to process food waste into compost for use in tree planting initiatives.

Sewage is collected and pumped to the wastewater treatment plant with treated effluent recycled back into the process or reused in road watering or vegetation projects. In remote locations septic tanks and leach beds are used.

The TTV is sized to accommodate a workforce of 3,540 personnel. It includes various facilities, such as clinic, laundry, kitchen and dining areas, gymnasiums, recreational rooms and various sports playgrounds.

Kinross typically establishes refining agreements with third-parties for refining of doré. Kinross’s bullion is sold on the spot market or as doré, by marketing experts retained in-house by Kinross. The terms contained within the refining contracts and sales contracts are typical and consistent with standard industry practice, and are similar to contracts for the supply of bullion and doré elsewhere in the world.

Current mine operations and the expansion project are based on the formal approval of a number of Environmental Impact Assessment (EIA) studies completed before and since mine commissioning in 2007.

For all project areas, environmental baseline conditions have been determined by reviewing existing published data, previous EIAs, satellite imagery and environmental reporting undertaken for the mine. Where appropriate, existing data for project areas was supplemented by primary data collected through environmental baseline surveys. Field-level baseline surveys were completed for project areas, including air quality, archaeology, flora, fauna, marine, water quality, traffic, and socioeconomics.

The baseline conditions formed the basis to assess the project through a series of EIAs and Environmental Impact Notices (EINs). The environmental assessments used applicable Mauritania legislation, the International Finance Corporation Performance Standards, the International Cyanide Management Code and Kinross Health, Safety, Environment and Social Management Systems for project design and management, mitigation strategies and performance monitoring. The environmental assessments determined appropriate mitigation and management where impacts could not be avoided through project design.

A review of waste rock geochemistry to determine the potential for acid rock drainage concluded that the rock has excess neutralizing capacity. Given the excess neutralizing capacity and the very low precipitation at Tasiast, acid rock drainage is not anticipated.

The Tasiast facilities operate under an environmental management system (EMS) that specifies activities to be planned and implemented by the mine’s environmental management team. The EMS incorporates the project design and management, mitigation strategies and performance monitoring commitments outlined in the environmental assessments, applicable legislation and specific permit requirements.

An element of each EIA prepared for the Tasiast mine site is a preliminary reclamation and closure plan and associated cost estimate. The preliminary reclamation and closure plan outlines the measures that will be taken to reclaim and close the proposed activities assessed in each EIA. The preliminary reclamation and closure cost estimate forms the basis of the financial assurance. The current financial assurance for the existing operation is approximately $6.2 million. In 2016, Tasiast submitted an updated closure estimate of $37 million which included the 12 kt/d expansion. With the proposed 24 kt/d expansion the estimated closure costs increase to $45.4 million. The Government of Mauritania has requested an independent review of the closure estimate. Upon completion of the independent review, Tasiast will put additional financial assurance in place. At least two years before entering closure, a detailed reclamation and closure plan must be submitted to the appropriate ministries for approval.

In addition to the exploitation permit No. 229 C2 (Section 4.2) and the adjacent exploitation and exploration permits, all other necessary permits for exploiting the Tasiast mine complex have been granted by the relevant Mauritanian authorities. A Phase 3 EIA for “off-site” sea water supply was approved following submission of a Phase 3 addendum. A subsequent EIA was approved to allow receipt of pre-assembled equipment at a beach landing and transportation to site. In addition, following discussion with the Government, an addendum to the Phase Two EIA was submitted and approved that described the project optimization through incremental increases in production and relocation of certain infrastructure. This addendum was approved by the Ministry of Environment in February 2016 and the Ministry of Mines in March 2016. . An application is pending with Mauritanian authorities for the installation of additional wells at the Sondage. The key permits are shown in Table 20-1.

Mauritania is divided into 12 wilayahs (regions), one district (Nouakchott), 53 moughataas (counties) and 208 communes (municipalities).

The mine site is located in the Inchiri wilayah, which has a very low population density. The wilayah includes the Akjoujt moughataa and two main municipalities, Akjoujt and Bennichab, Akjoujt being the administrative capital with a population of approximately 8,500. The wilayah is administered by a council, directed by a governor (wali) who reports to the Minister of Interior. The basic administrative unit, the moughataa, is directed by a Prefect (Hakem) who exercises his power under the authority of the governor.

Inchiri is the least populated wilayah in the country, with the nomadic way of life being a key feature making up 20% of the total population. There tends to be a small number of nomadic people in the vicinity of the Tasiast mine. The mine itself is located 80 km northeast from the nearest permanent community of Chami.

The nearest industries to the site are in the towns of Chami, Boulanour, Akjoujt and Bennichab, which are respectively 80 km southwest, 120 km northwest, 150 km east-southeast and 130 km southeast from the mine site.

There are no permanent settlements within the vicinity of the Tasiast mine. However, within 30 km of the Tasiast mine, a number of isolated families have set up structures and reside, predominantly within three communities. Residents practice animal husbandry and other subsistence forms of livelihood. There are also nomadic groups that occasionally transit the area.

The incremental capital cost estimates of the 24 kt/d throughput expansion includes, but is not limited to the following scope:

The Tasiast debottlenecking project assumes the total capital costs will be distributed over four years, approximately $85M of which will be spent through 2020, decreasing to $30M, $20M and $15M per year in 2021 to 2023, respectively.

The annual sustaining and “non-sustaining capitalized stripping” cost estimate is summarized by year in Table 21-3. Per World Gold Council guidance, the capitalized portion of waste movement is considered non-sustaining life of mine, as it exposes long-term ore sources. It is excluded from Table 21-1 and the calculation of the All-In Sustaining Cost metric.

The Tasiast “life of mine” operating costs are split into four primary categories: Mining, Processing, Site Administration, and Other. See Table 21-4 for a summary of the basis of estimate for these categories.

Processing includes both the 24 kt/d CIL Mill and existing Dump Leach operations. Note the dump leach operation is nearing the end of its life. Rinsing of the pads is expected to begin in early 2020 and take approximately 2 years to complete before closure.

·     Defining a haulage network (specific to the detailed mine plan) and generating truck hours based on travel distance, speed and fixed non-travel time

o   Headcounts – fitted to the scale of the mine (i.e. fewer operator and non-operator positions would be required as mining rate decreases)

o   Maintenance costs, calculated from a zero-based maintenance model (tracks and schedules maintenance events for each piece of equipment at site by operating hours)

o   Other inputs, such as tire life and drill consumables – based on existing site strategy and experience

The following costs are allocated by department and based on actuals adjusted for changes in mining headcount and rates:

Estimation methodology varied by cost component, but primarily built from first principles, relying on a combination of:

Major categories include the following, which collectively result in a processing cost estimate for the expansion scenario:

The following costs are allocated by department and based on actuals adjusted for changes in processing headcount and rates:

The following costs are allocated by department and based on actuals adjusted for changes on headcount:

The economics of the Tasiast Debottlenecking Project were evaluated using a real (non-escalated), after-tax discounted cash flow (DCF) model on a 100% project equity (unlevered) basis. Unless otherwise stated, all economic parameters are shown on an absolute basis (not incremental to existing operations). Production, revenues, operating costs, capital costs and taxes were considered in the financial model. The main economic assumptions are a US$1,200/oz gold price and a 5% discount rate.

The valuation date for the financial analysis was set for January 1, 2020. All cash flows assumed for the purposes of this study are from this date onward.

The cash flow analysis was used to estimate the economics of the 24 kt/d debottlenecking project. This scenario assumes that 21 kt/d is achieved by end of Q4 2021, with full 24 kt/d capacity at end of Q2 2023.

Dump leach is not considered a significant source of production beyond early 2020. Rinsing for two years and decommissioning planned to be completed by end of 2022.

Existing Phase 1B, Phase 1A “MAK units” and the new 40MW power plant will provide 65MW of HFO capacity.

Spending before January 1, 2020, is treated as a sunk cost and is not considered in the analysis, except for opening balances for tax depreciation.

Sustaining capital costs: portions were derived using zero-based costing where possible. Provisional estimates for LOM values were made otherwise.

The results of the financial analysis, with sensitivities to gold price and discount rate assumptions, are shown in Table 22-2, based on a real discount rate of 5%, and an oil price of $55/bbl. Annual life-of-mine cash flows are shown in Table 22-3.

Tasiast is viewed as a long-term strategic asset for Kinross, located in a district that is believed to have significant future potential. The phased expansion project is believed to provide an opportunity to capitalize on the full potential of the operation and to solidify Tasiast as a low cost, long life asset within the company’s portfolio.

The project economics, as stated at a 5% discount rate and a $1,200 base case gold price, are robust and offer significant potential.

Kinross is confident in the technical and economic assessment presented in this Technical Report. However, the results of this Technical Report are subject to many risks including, but not limited to: commodity and foreign exchange assumptions (particularly relative movement of gold and oil prices), unanticipated inflation of capital or operating costs, significant changes in equipment productivities, geotechnical assumptions in pit designs, ore dilution or loss, throughput and recovery rate assumptions, availability of financing and changes in modelled taxes.

It is recommended that Kinross proceed with the project to incrementally increase throughput capacity at Tasiast from approximately 15,000 t/d throughput to 24,000 t/d.

Blake, C., 2011a. Mineralogical characterisation of five gold-bearing samples from the Tasiast mine, Mauritania for Kinross Gold Corporation. Internal unpublished report (February, 2011).

Blake, C., 2011b. Mineralogical characterisation of seven gold-bearing composite samples from the Tasiast Mine, Mauritania for Kinross Gold Corporation. Internal unpublished report (September, 2011).

Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2014. CIM Standards for Mineral Resources and Mineral Reserves, Definitions and Guidelines: Canadian Institute of Mining, Metallurgy and Petroleum, May 10, 2014.

Chartier, D., 2013. Draft of Sampling and Data Verification Sections for Tasiast Feasibility Study. Memo dated December 4, 2013.

Demers, P., Gauthier, D., Kroon, A.S., and Lafleur P-J., 2004. Technical Report on the Tasiast Gold Project, Islamic Republic of Mauritania: unpublished technical report prepared by SNC Lavelin for Defiance Mining Corporation Limited, effective date 27 May 2004.

Fabre, J., 2005. Géologie du Sahara occidental et central. Série/Reeks: Tervuren African Geosciences Collection, MRAC Tervuren, Belgique, 572 p.

Golder Associates, 2004. Preliminary Pit Slope Design, Tasiast Project, Mauritania, Report 03-117-073. Unpublished Technical Report, June 2004.

Golder Associates, 2014. Geotechnical Study in Support of West Branch Expansion, Kinross Tasiast, Report 12514150422.502/A.0. Unpublished Technical Report, June 2014.

Guibal, D., Uttley, P., de Visser, J., and Warren, M., 2003. Independent Technical Assessment Report on the Tasiast Project, Mauritania; Report Prepared for Midas Gold plc. and Geomaque Explorations Ltd: unpublished technical report prepared by SRK Consulting for Midas Gold plc. and Geomaque Explorations Ltd, effective date 4 March 2003.

Heberlein, D., 2011. QAQC update – laboratory duplicate results (Oct-Dec, 2011). Internal unpublished report (October, 2011).

Heberlein, D., 2013. QAQC Update (August 2012). Internal report prepared for Tasiast Mauritanie Ltd S. A.; a division of Kinross Gold Corp. Dated September 4, 2013.

Heron, K., Jessell, M., Benn, K., Harris, E., and Crowley, Q.G., 2016. The Tasiast deposit, Mauritania. Ore Geology Reviews 78: 564–572.

Hyde, R., 2003. Tasiast Gold Project, Review of Sampling Procedures and QAQC Analysis for Midas Gold plc.: RSG Global Report No. TAS003, 42 pages (unpublished).

Leitch, C., 2010. Petrographic report on 18 samples from Archean greenstone au deposit. Internal unpublished report (October, 2010).

Leroux, D.C. and Puritch, E., 2003. Technical Report and Resource Estimation on the Tasiast Gold Project Islamic Republic of Mauritania for Defiance Mining Corporation: unpublished technical report prepared by ACA Howe for Defiance Mining Corporation, effective date 10 October 2003.

Leroux, D.C., Roy, W.D., and Orava D., 2007. Technical Report on the Tasiast Gold Project Islamic Republic of Mauritania for Red Back Mining Inc.: unpublished technical report prepared by ACA Howe for Red Back Mining Inc., effective date 20 July 2007.

Panterra, 2012. Petrographic Study of the Tasiast Deposit, Mauritania, West Africa, November 7, 2012.

Panterra, 2017. Petrographic Report on the Tasiast West Branch Pit, Mauritania, West Africa, Feb 18, 2017.

Pollard, P., 2011. Hydrothermal Alteration and Mineralization in the Lower West Branch Zone, Tasiast Gold Mine, Mauritania: unpublished technical report prepared for Kinross Gold.

Schlumberger, 2014. Hydrogeological Characterisation of the West Branch Pit, Tasiast, Mauritania. Technical Report Ref 52116/R1. January 2014.

Schofield, D., Horstwood. M.S.A., Pitfield, P.E.J., Crowley, Q.G., Wilkinson, A.F. & Sidaty, H.C.O., 2006. Timing and kinematics of Eburnean tectonics in the central Reguibat Shield, Mauritania. Journal Geological Society London, 163, 549-560.

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

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

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

Strashimirov, S., 2010. Petrological and mineralogical studies of 10 samples from Mauritania. Internal unpublished report (August, 2010).

Stuart, H., 2008. Technical Report on the Tasiast Gold Mine Islamic Republic of Mauritania for Red Back Mining Inc.: unpublished technical report prepared for Red Back Mining Inc., effective date 24 May 2008.

Stuart, H., 2009. Technical Report on the Tasiast Gold Mine Islamic Republic of Mauritania for Red Back Mining Inc.: unpublished technical report prepared for Red Back Mining Inc., effective date 8 May 2009.

Stuart, H., 2010. Technical Report on the Tasiast Gold Mine Islamic Republic of Mauritania for Red Back Mining Inc.: unpublished technical report prepared for Red Back Mining Inc., effective date 6 September 2010.

The effective date of this Technical Report entitled “Kinross Gold Corporation, Tasiast Project, Mauritania, NI 43-101 Technical Report” is October 31, 2019.

Reference is made to the technical report, dated October 31, 2019, entitled “Tasiast Project, Maritania, National Instrument 43-101 Technical Report”, which the undersigned prepared for Kinross Gold Corporation (the “Technical Report”).

Pursuant to Section 8.3 of National Instrument 43-101 – Standards of Disclosure for Mineral Projects, the undersigned hereby consents to the public filing of the Technical Report by Kinross Gold Corporation with the Canadian Securities Commissions.

I hereby consent to the inclusion of the Tasiast Project, Mauritania, National Instrument 43-101 Technical Report (“Technical Report”), effective date October 31, 2019 in the report on Form 6-K dated October 31, 2019 to be filed by Kinross Gold Corporation.

I also hereby consent to the incorporation by reference of the Technical Report into the Registration Statements on Form S-8 (Registration No. 333-217099 filed on April 3, 2017; and Registration Nos. 333-180824, 333-180823 and 333-180822 filed on April 19, 2012) and the Registration Statement on Form F-10 (Registration No. 333-223457 filed on March 6, 2018) as amended.

Ang Gu Kueh Machine Suppliers for Sale

Receive full access to all new and archived articles, unlimited portfolio tracking, e-mail alerts, custom newswires and RSS feeds - and more!

Stuffed Pastry Machine, Maamoul Machine, Automatic Encrusting Machine - Papa,https://www.papaindustrial.com/