Overview

Deposit type

• Vein / narrow vein
• Hydrothermal
• Breccia pipe / Stockwork

Summary:

The Detour Lake and West Detour deposits are considered to be examples of orogenic greenstone-hosted hydrothermal lode gold deposits.

Greenstone-hosted hydrothermal lode gold deposits are typical of the Abitibi Greenstone Belt, and in particular the gold deposits found along the Destor–Porcupine Fault Zone from Timmins, Ontario through to Destor, Québec. These deposit types are found in greenstone belts around the world and are responsible for a large proportion of past world gold production, including most of the Canadian gold production.

The majority of Archean orogenic greenstone-hosted lode gold deposits occur within volcano–plutonic domains, which are typically distributed along crustal-scale fault zones occurring along or in close proximity to terrane or sub-province boundaries (Card et al, 1989). Elongate belts of metavolcanic and some metasedimentary rocks containing subsidiary amounts of ultramafic to felsic intrusive rocks typically dominate these domains. The intrusive rocks will have typically been emplaced in multiple pulses throughout the geologic evolution of the area. Metamorphism within the belts is generally greenschist to lower amphibolite facies. The structure of the gold districts is characterized by the presence of multiple generations of structural fabrics indicating the presence of several periods of deformation.

Mineralization
There are two recognized episodes of gold mineralization at the Detour Lake and West Detour deposits.

The first episode consists of a wide and generally auriferous sulphide-poor quartz vein stockwork formed in the hanging wall of the Sunday Lake Deformation Zone. The sulphidepoor quartz vein stockworks observed in the hanging wall have sub-vertical north or south dips and are parallel to a series of east-west trending high strain zones. These veins form a weak stockwork and are boudinaged and/or folded. The second episode is a stage of gold mineralization overprinting the early auriferous stockwork, principally in the hanging wall of the Sunday Lake Deformation Zone, with a higher sulphide content. The sulphide-rich gold mineralization predominantly fills structural sites in deformed quartz veins, fractures and veins crosscutting the foliation fabric but also in pillow breccias and selvages. The distribution of sulphide-rich mineralization is strongly controlled by the geometry of kinematic orientation (i.e., pyrite and pyrrhotite concentrations have a shallow westerly plunge similar to the plunge of the main flexure zone in the Sunday Lake Deformation Zone at an angle of about 40° (in the area of the former Campbell pit), shallowing to approximately 10° further to the west).

Detour Lake
Dimensions: Mineralization at Detour Lake is hosted within a broad corridor. In mine grid coordinates the mineralization extends from 16,900E to 20,650E, 19300N to 20,650N and 5,500 m elevation to 6,300 m elevation. This corresponds to a strike extent of 3.75 km, a width of 1.35 km, and an approximate elevation range of 800 m.

Mineralization
Gold is associated with quartz–carbonate–pyrite–pyrrhotite ± tourmaline veins and/or disseminated to very local semi-massive sulphides in hydrothermally-altered wall rocks. There are two main mineralized zones, defined as hanging wall mineralization and footwall mineralization.

Hanging Wall
The hanging wall gold mineralization occurs in several different rock units within broad subvertical mineralized envelopes and splits into several sub-vertical domains sub-parallel to the orientation of the Sunday Lake Deformation Zone. The Main Zone was the largest gold-bearing mineralized zone exploited in the period 1983– 1987 and consists of gold mineralization occurring in the chert marker horizon or in quartz and quartz–carbonate vein systems splaying from the Sunday Lake Deformation Zone. In the Campbell pit area, <1 m thick quartz veins, with a frequency of greater than one vein per metre, were part of a series of sub vertically dipping east–west-trending highly-strained zones. Gold generally occurred as free gold with these veins.

West of section 19,620E, mineralization is commonly associated with increased biotite alteration, shearing, narrow quartz veining and minor pyrite or pyrrhotite. Local zones of strong brecciation with sulphide infilling have also been recognized along with minor chalcopyrite, telluride minerals and visible gold. Gold mineralization is associated with a series of sub-vertical to arcuate deformation zones characterized by enhanced strained fabrics, well defined open-space breccias, and to a lesser degree sheeted shear-hosted veins and extensional veins. The QK Zone further to the west is open down plunge and has not been tested below 800 m. This mineralization is associated with narrow parallel to sub-parallel quartz veins, quartz boudins and sulphide rich veins/breccias with adjacent silicification and potassic alteration envelopes.

Footwall Mineralization
The Talc Zone varies in width from 4–15 m and tends to be less continuous along strike. Gold in the Talc Zone is dominantly associated with pyrite, pyrrhotite, and minor chalcopyrite along foliation planes, narrow discrete shears or strain zones, and in irregular lenses. The zones also contain short deformed lenses or boudinaged quartz veins. In some cases, it appears that the mineralization is controlled by strong fault structures containing several centimetres of gouge material.

West Detour and North Pit Areas
Dimensions: The West Detour mineralization extends from mine grid sections 14,500E to 17,600E and 19,100N to 21,300N and between elevation 5,300 and 6,300 m. This corresponds to a strike extent of 3.1 km, a width of 2.2 km, and an approximate elevation range of 1 km.

Mineralization
The M Zone lies approximately 400–500 m north of the chert maker horizon and is a westerly-trending gold system that is spatially associated with the margins of the chlorite schist unit. The chlorite schist stratigraphic horizon and associated gold mineralization was traced by drilling for approximately 5 km. The footwall and hanging wall sequence of the chlorite schist in the M Zone is variably biotite altered with fairly abundant fractures and well-defined foliation (local veining) with associated pyrite, pyrrhotite, and rarely chalcopyrite. The mineralized zones appear lensoidal and plunge 20° west. Mineralized lenses vary from 5–50 m in true width. Mineralization hosted in the QK zone is associated with narrow veins that are parallel to sub-parallel to stratigraphy, quartz boudins, and sulphide-rich veins/breccias within silicified and potassic alteration envelopes up to 25 m in true width. Gold mineralization in the North Walter Lake Zone is within a relatively weak quartz vein stockwork with a low pyrite and pyrrhotite sulphide content. Mineralization is associated with moderate to strong shearing, potassic alteration of the mafic volcanic units, narrow quartz veining and minor pyrite and pyrrhotite.

Zone 58N
Dimensions: The Zone 58N mineralized system has been intersected over an east–west strike length of 450 m, from surface to a depth of 800 m, and the mineralized system remains open at depth. The mineralization extends from approximately 595,300E to 595,750E and 5,533,700N to 5,533,800N (UTM coordinates). The width of the mineralization is variable, ranging from 4 m to >100 m at the centre of the deposit. Infill drilling has demonstrated that the geology and mineralization dip subvertically at 75º to the south.

Mineralization
Visible gold is often present and occurs within coarse pyrite or as free gold within quartz. Sulphide mineralization associated with gold ranges from 0.5–5% pyrite with minor chalcopyrite, bismuth-tellurides, molybdenite and scheelite. Gold is found within and at the margins of quartz ± tourmaline ± carbonate stockwork-type veins that infill areas of brittle deformation. Visible gold occurs in nearly every drill hole that intersects mineraliza

Mining Methods

• Truck & Shovel / Loader

Summary:

The Detour Lake Mine uses conventional truck-shovel open pit mining. The mine is operated using an Owner-operator mining equipment and labour strategy. Excluding the muskeg, overburden/till top layer, all material must be blasted. Pioneering drilling and blasting is required in the overburden/rock contact. Additionally, during winter months free digging of overburden material is not possible due to frost.

The Detour Lake mine is a large open pit operation comprised of the Detour Lake Main Pit currently in operation and the planned North Pit.

Design Constraints
The design of each one of the three final pits (Detour Lake, West Detour and North Pit) is guided by the respective optimized shells. The final pit design incorporates the geotechnical parameters prescribed by Golder (Golder 2019, 2020).

The Detour Lake pit design incorporates a double ramp access for most of the LOM. The final ramp and principal access will be located in the north wall. The West Detour and North Pit were designed using a single ramp access.

All ramps are designed to accommodate the safe operation of 795CAT super class trucks. Typical road width is 33 m with a total length of 40 m to allow for safety berm and drainage. Ramp widths are adjusted at the bottom of the pit to a one-way lane. Other design parameters include a ramp slope (10%), minimum curve radius (39 m) and construction constraints (super elevation, crowning).

There are a total of nine phases designed for the LOM: five for the Detour Lake Main Pit, three for West Detour and one for North Pit.

Pit Final Dimensions
Detour Lake Main Pit: Width - 1,310 m; Length - 3,070 m; Depth - 606 m;
West Detour Pit: Width - 810 m; Length - 1,890 m; Depth - 302 m;
North Pit: Width - 460 m; Length - 640 m; Depth - 144 m.

Overburden and Pre-Stripping
Bedrock at both Detour Lake and West Detour is overlain by overburden composed of layers of muskeg, till, gravel, and sand. These overburden layers may be as thick as 40 m. At the start of each mining phase, this material must be stripped prior to the drilling and blasting of rock material. Depending on the material quality, weather and moisture content, this overburden material is typically ‘free-digging’ and does not require drilling and blasting. A portion of the till material (subject to quality control specifications) is stockpiled and used for tailings management area (TMA) construction purposes. The remaining overburden is stockpiled either in the main overburden stockpile or within designated areas of the WRSFs. A portion of this material will be rehandled for the reclamation of various waste rock stockpiles and/or TMA cells. Whenever possible, this overburden is directly placed for the progressive reclamation of ultimate WRSF slopes.

Mining Operations
The Detour Lake Mine uses conventional truck-shovel open pit mining. Excluding the muskeg, overburden/till top layer, all material must be blasted. Pioneering drilling and blasting is required in the overburden/rock contact. Additionally, during winter months free digging of overburden material is not possible due to frost. The mine operation also requires the management of old underground workings.

Underground workings records are considered of good quality. The mine has been operating in historical underground working areas for the past years; the reconciliation of historical recordings and field observation of these workings is very good. The assumption that stopes were backfilled using sand/rock filled has also been confirmed. There are procedures in place to ensure safe operations in these areas. The procedures define zones classified by the risk of ground instability.

The Detour Lake Mine has used a 12 m bench-height since operations commenced. A 6 m sub-bench was used to mine the areas requiring pioneering to improve operational conditions related to boulders and pinnacles of bedrock.

Starting in mid-2020 the mine transited to a 14.5 m bench height for areas to be primarily mined by rope shovels and to 7.25 m benches in areas to be mined using hydraulic shovels. The revised mine design has led to improvements in shovel productivities.

The mine operates 24-hr per day year-round on 12 hr shifts for all operational crews. Operational teams work on a seven-day in/seven-day out rotation. The mine has implemented a successful hot-seating process for its main production equipment (shovel, trucks and drills). Management/supervisory and support personnel work at site on different schedules.

Ore Mining and Stockpiles
Whenever possible, mined ore is delivered directly to the primary crusher in order to avoid unnecessary rehandling. When the mined ore tonnage exceeds the operating capacity of the crusher, the ore is placed on one of the ROM stockpiles for later mill feed.

The ROM stockpiles are located in close proximity to the primary crusher to allow for efficient feeding into the plant. Ore in ROM stockpiles is segregated into piles consisting of similar grade material to allow for controlled feeding as required. The ROM stockpiles provide additional buffering capacity to ensure that crusher feed is maximized while also limiting excessive tramming of loading equipment between ore and waste areas in the pit.

Waste Mining
Waste mining is generally allocated to the rope shovel fleet. The grade control process also delineates waste into PAG or NAG material. PAG material must be routed to the MRS3 or MRS1 WRSFs. Overburden material is generally mined by the hydraulic shovels.

Drilling and Blasting
Drilling and blasting activities are performed by Kirkland Lake Gold personnel. Dyno Nobel is the explosives supplier. Different objectives and geological/fractural information are considered when design the patterns and loading plans. Yields vary with the different objectives (wall control, fragmentation, higher fragmentation, reduction of overcasting material) and with areas in the pit.

Comminution

Crushers and Mills

Summary:

Crushing Circuit
The primary crushing system is a single stage, open circuit, primary gyratory crusher (60 x 113 inches). The crusher selection was based on a feed top size of 1,200 mm and a product P80 of 165 mm. The primary crusher has a capacity of 5,000 t/hr at a 138 mm close side setting. Assuming an 80% availability, the primary crusher has a capacity of 96,000 t/d. The live capacity of the feed and discharge hoppers of the gyratory crusher was designed for two truck loads each, assuming a nominal payload of 300 t. The crushed ore storage pile was designed with a live capacity corresponding to approximately 12 hours of crushing or 30,000 t, and an overall capacity (live plus dead) of 100,000 t.

Ore is reclaimed from the stockpile through two reclaim tunnels, one for each grinding line. Four apron feeders, two in each reclaim tunnel, discharge the crushed ore onto a belt conveyor that feeds a secondary cone crusher operated in open circuit. The secondary crusher is fed with the gyratory product with a P80 of 165 mm, and gives a product with a P80 of 50 mm. The secondary crusher product is conveyed directly to the SAG mill. The secondary crusher is equipped with a bypass chute to maintain high process plant availability. During maintenance of the secondary crusher, the bypass is put into place to feed the SAG mill directly from the stockpile.

Crusher Improvement Initiatives
The primary crusher is a key area where initiatives are planned to generate a finer product that will translate into a higher plant throughput.

The impact of fragmentation on throughput has been well documented and has a very high impact on throughput.

It has been established that each 50 mm reduction in the P80 equals about a 300 t/hr increase in plant throughput. In 2019, the Detour Lake Mine started a drill-and-blast optimization project that has resulted in higher throughput rates and better operational conditions. The initiative is ongoing as the mine aims to further optimize the process.

The choke feeding of the primary crusher is another key element to increase the plant throughput milling rates. It has been well established that the longer a crusher stays full, the finer the product will be (choke feeding).

The primary crusher choke feeding can be improved. The avenues to do this are to improve dispatching of trucks, better fragmentation to avoid bridging of the crusher, an excavator dedicated to the primary crusher re-feed, and improved rock breaker availability and performance.

One of the most important improvements associated with increases in the milling rate will be the addition of screens over the secondary crushers. The first consideration for efficient secondary crushing is to have the correct feed preparation. A secondary cone crusher achieves its best work when fines smaller than the desired crusher product are removed from the feed. This was not done during the initial design, and as a result, the secondary crushers receive direct feed from the primary crusher with all the fines still in the feed.

Grinding Circuit
The SAG mill operates in closed circuit with a pebble crusher while the ball mill operates in closed circuit with hydrocyclones.

The design circulating load from the cyclones to the ball mill is 250% of the SAG mill new feed. The pebble crushing circuit with a design discharge P80 of 13 mm processes the equivalent of up to 25% of the fresh feed.

Around 15% of the cyclone feed material is sent to the gravity recovery circuit. Six gravity concentrators with 48 inch bowls were selected (three per grinding line supporting a strategy of two units in operation at all time). A batch-intensive cyanidation system is used to process the gravity concentrate. The extraction performance of gold from the gravity concentrate by the intensive cyanidation system is 99%. The pregnant solution is pumped to a tank in the gold room followed by electrowinning in a dedicated cell.

Curved Pulp Lifters
As part of the initiatives to increase the milling rates, the SAG mill pulp lifters will be converted to curved from their current radial configuration. Evidence of severe backflow has been seen on the pulp lifters indicating the pulp lifters are not empty when they go back inside the charge.

Re-Feed System After the Secondary Crushers
Every time a crusher goes down, whether the crusher is primary, secondary, pebble, the plant throughput is impacted. The goal with the re-feed system is to generate 750 kt of -2.5” feed annually, which is the equivalent size of the secondary crusher discharge. This material would be re-fed onto the SAG mill feed conveyor during of crusher downtime events.

Pebble Crusher Variable Speed Drives
By its nature, the rate of pebbles a SAG mill generates is not uniform. The rate is influenced by the ore feed size, mill speed and load inside the mill. The result is that the power of the pebble crusher cannot be maximized due to the variation in feed to the pebble crushers. The crusher is gapped so it can receive any surge in pebbles. The average power is therefore maintained around 450 kW while the motor is rated for 750 kW. The pebble crusher is rarely choke fed.

By installing a variable speed drive (VFD) on the pebble crusher motors, the speed can be increased and decreased to maintain a set-point of 550 kW for power draw. This will enable the crusher to be choke fed and reduce the particle size coming out of the crusher. With a reduced particle size, the SAG mill throughput will increase.

Ore Blending
Ore blending is an avenue that the processing plant will use to optimize milling rates. With the open pit supplying more ore than what the mill can process for the next five years, talc ore can be stockpiled and processed opportunistically. That means that when the ore from the pit is soft, talc ore can be stockpiled and reclaimed when the ore will be harder.

Increase Plant Operating Time to 93%
The operating time for the processing plant is calculated when the ore is on the belt feeding the SAG mills. Operating time is typically 1–2% lower than the mechanical/electrical availability. Although plant operating time has improved since start-up, the design rate of 92% has not yet been achieved over an annual basis, instead averaging around 88–89%.

Processing

• Gravity separation
• Crush & Screen plant
• Intensive Cyanidation Reactor (ICR)
• Electric furnace
• Hydrochloric acid (reagent)
• Smelting
• Centrifugal concentrator
• Carbon re-activation kiln
• INCO sulfur dioxide/air process
• Agitated tank (VAT) leaching
• Carbon in pulp (CIP)
• Carbon adsorption-desorption-recovery (ADR)
• Elution
• Solvent Extraction & Electrowinning
• Cyanide (reagent)

Summary:

The process plant is based on a robust metallurgical flowsheet designed for optimum recovery with minimum operating costs. The flowsheet is based upon unit operations that are well proven in industry.

The milling operation uses a conventional crushing, grinding, gravity, cyanidation and carbon-in-pulp processing facility currently operating at approximately 24 million tonnes per year, with the Company targeting increasing this rate to 28 million tonnes per year late in the second half of 2024.

In 2022, the Company completed the construction and/or installation of the fourth Detox tank, the secondary crusher screens, 14 higher capacity Kemix© screens, and the 610 refeed system.

In 2023, the Company completed the construction of improvements to the elution system, and continued work on leach tanks and the construction on tailings facility cell 2 to an elevation of 301 metres. Major projects planned for 2024 including a further three metre lift on tailings facility cell 2, the construction of a four bay addition to the mine’s truck shop, the installation of a ball mill discharge grizzly, improvements to the 230kV substation, and the commissioning of a new Komatsu 4100 rope shovel and the secondary crusher variable frequency drive (VFD) and a leach tank improvements.

Leach and Carbon-in-PuIp
The gold recovery circuit selected was a “leach circuit followed by a CIP” circuit. The designed retention time for leaching is 29 hours. The leach feed size was designed at a P80 of 95 µm. Four new leach tanks are planned to be built to support the increase in throughput to 28 Mt/a. The CIP design selected uses a carrousel type system with an average design retention time of 80 minutes in this circuit. Upgrades to the pumpcells will support the 28 Mt/a throughput rate. The VSD and motors for the CIP tails pumps will be upgraded to 500 hp from 350 hp. The current adsorbtion profile and modeling support the planned increase in throughput rate.

Acid Wash, Stripping, Electrowinning, and Refining
The stripping system uses a modified version (i.e. ,on-line electrowinning and no pregnant solution tank) of the high-pressure Zadra process to recover the gold from the loaded carbon.

The circuit incorporates an acid wash stage and is designed to handle up to 10 t of carbon per day. The acid wash stage is not currently used and will be converted into a plastic removal vessel. The build up of calcium is not sufficient to warrant the use of the acid wash vessel.

The stripping circuit is designed to handle up to 10 t of carbon per stripping cycle (approximately 8–10 hours); 100% of the carbon is regenerated in two regeneration kilns designed to handle up to 20 t of carbon per day.

The electrowinning is done “in-line” with the stripping circuit. The flow of pregnant solution is split between three rows of two electrowinning cells. One dedicated electrowinning cell handles the pregnant solution from the ILR tank.

The refining equipment is designed to handle the gold from the stripping circuit and from the gravity recovery system. The electrowinning sludge is filtered, dried and mixed with fluxes, before being smelted using induction furnaces.

The current elution capacity is estimated to be 2.5 strips of 10 t per day. In 2020, 1.75 strips per day were completed. There is sufficient capacity to cover the increased throughput for the LOM.

Thickening
The feed to the leach circuit is maintained at a density of 50–55% solids by weight using one high rate thickener. The overflow of the pre-leach thickener is recycled to the process water tank. The pre-detox thickener is used to thicken the leach tails slurry to 55–60% solids and the overflow is recycled to the process water circuit.

The thickener capacity is over 90 kt/d and can support the increase throughput without any modifications. Current flocculent consumption is low at 10 g/t.

Water Supply

Summary:

Potable water for the Little Hopper Lodge is obtained from Little Hopper Lake, adjacent to the lodge, and processed through a water purification plant. This is adequate for the Detour Lake’s current and future needs.

Potable water for the Sagimeo Lodge, mine services facilities, site administration facility, and the processing plant is obtained from borehole wells close to the camp, and processed through a water purification plant.

Fresh water is pumped from East Lake. This water is primarily used in the processing plant for reagent mixing but is also used as wash water in the truck wash facility and water makeup for the fire water tank.

A decant tower system is used to reclaim and pump water from the TMA back to the process plant to satisfy process water requirements. For Cell#2 a barge reclaim water system will be used. Reclaim water is piped to the process plant. Depending on the water balance in the TMA, water from the open pit and various collection ditches around the WRSFs can also be directed to the plant.

Make-up water for the reagent mixing is sourced from East Lake when required.

Production

Operational metrics

Personnel

Job Title	Name	Email	Profile	Ref. Date
Consultant - Costs	Paul Fournier			Jul 26, 2021
Deputy General Manager	Colin Ashton			Apr 4, 2024
Health & Safety Superintendent	Douglas Brown			Apr 9, 2024
Mine Manager	Michelle Moore			Apr 9, 2024
Mine Operations Superintendent	Paul Pepper			Apr 9, 2024
Mobile Maintenance Manager	Louis Gendron			Apr 9, 2024
Mobile Maintenance Superintendent	Derek Buzzi			Apr 4, 2024
Plant Maintenance Manager	Raphael Boutin			Apr 4, 2024
Plant Maintenance Superintendent	Eric Saulnier			Apr 9, 2024
Sr. Maintenance Superintendent	Jason Thibodeau			Apr 9, 2024
Sustainability & Environmental Manager	Melissa Leclair	melissa.leclair@agnicoeagle.com		Apr 9, 2024
VP Operations	Andre Leite	andre.leite@agnicoeagle.com		Apr 9, 2024