Overview

Deposit type

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

Summary:

The Detour Lake and Zone 58N 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
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.

Seven dominant mineralized domains containing most of the high-grade mineralization at the Detour Lake deposit have been defined and characterized. Gold is associated with quartz–carbonate–pyrite–pyrrhotite ± tourmaline veins and/or disseminated to very local semi-massive sulphides in hydrothermally-altered wall rocks.

Domain 51 is spatially associated with the Chert Marker Horizon known as the Main Zone. The mineralized zone consists of gold mineralization occurring along the Chert Marker Horizon in quartz and quartz–carbonate vein systems splaying from the Sunday Lake Deformation Zone.

In the East end of the Detour deposit and coincident with the Chert Marker Horizon, various talc zones occur along the footwall. With widths from 4–15 m, the talc–chlorite mineralization tends to be less continuous along strike than mineralization along the hanging wall side of the Chert Marker Horizon. 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.

Domain 52 is associated with hanging wall mineralization occurring in several different rock units within broad sub-vertical mineralized envelopes, which may split into several sub-vertical domains sub-parallel to the Sunday Lake Deformation Zone.

The main underground vein corridors mined historically consisted of the Q zones (Q50- Q120), which splay off the northwestern side of the Chert Marker Horizon, and comprise steeply-dipping, parallel cm-to-tens-of-cm-wide quartz-sulphide veins, and veinlets in sheeted zones locally >10 m wide. Mineralization is commonly associated with increased biotite alteration, shearing, narrow quartz veining and minor pyrite or pyrrhotite.

Further to the west lies the “QK Zone” as denoted by Placer Dome. This mineralization is associated with narrow parallel to sub-parallel quartz veins, quartz boudins and sulphiderich veins/breccias with adjacent silicification and potassic alteration envelopes.

Further to the north and hanging wall to the Q-zones, Domain 53 is spatially associated with highly-strained corridors at the contacts between pillowed and massive mafic flows encompassing higher grade mineralized zones. These broad zones contain variable amounts of quartz and pyrite, and are controlled mainly by east-west trending, moderately north-dipping folds and shear structures which plunge at a shallow angle to the west.

Identified as “M Zone” by Placer Dome, Domain 55 encompasses high-grade areas towards the west end of the deposit, mostly following the west extension of the chloritic greenstone unit. Spatially, Domain 55 lies approximately 400– 500 m north of the Chert Marker Horizon and is a westerly-trending gold system that is spatially associated with the margins of the chloritic greenstone unit.

Domain 54 is located west of Domain 55. Both of these amphibole–chlorite schist stratigraphic horizons and associated gold mineralization were traced by drilling for approximately 5 km. The footwall and hanging wall sequence of the chlorite schist is variably biotite altered with abundant fractures, with a well-defined foliation which contains local quartz 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.

Domain 56 has similar characteristics to Domain 53, encompassing high-grade mineralized zones within partially to highly-strained corridors at the contacts between pillowed and massive mafic rocks. They appear to be semiparallel to the chloritic greenstone, and are predominantly narrow east-west subvertical to 65º north–northwest-dipping structures.

Domain 57 is spatially associated with hanging wall high-grade east–northeast to west–southwest mineralized zones. Gold mineralization is concentrated within potassic-altered mafic volcanic units containing relatively weak quartz vein stockwork with low pyrite and pyrrhotite content exhibiting moderate to strong shearing.

Zone 58N
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 along strike and at depth. The width of the mineralization is variable, ranging from 4 m to >100 m at the centre of the deposit.

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.

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, blasting of overburden material may occasionally be required due to frost.

Given the smaller dimensions of the North Pit, a smaller equipment fleet size will be used.

The mine production schedule forecasts a total 2,088 Mt to be mined over a period of 20 years (2024-2044). A total of 818 Mt of ore is planned to be milled over a period of 30 years (2024-2053); with the last nine years of production supported by long-term stockpile reclaim.

The optimized and variable cut-off strategy combined with 248 Mt of stockpile capacity, results in an average LOM stripping ratio of 1.76.

Design Constraints
The design of the final pits (Detour Lake and North Pit) is guided by the respective optimized shells.

The Detour Lake pit design incorporates a double ramp access for most of the LOM. The final ramp and principal access will be in the north wall.

All ramps are designed to accommodate the safe operation of 795/798 CAT 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).

Pit Phases
- Eight pit phases are planned over a 20-year mine life in the Detour Lake pit;
- One pit phase is planned over a six-year mine life in the North Pit.

Pit Final Dimensions
Detour Lake Main Pit: Width - 1,345 m; Length - 5,500 m; Depth - 638 m;
North Pit: Width - 429 m; Length - 636 m; Depth - 131 m.

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. During winter months, blasting of overburden material may occasionally be required 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. 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 either a seven-day in/seven-day out or a 14-day in/14-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
Mineralized zones previously delineated by core drilling are infilled using RC drilling, which is typically performed well ahead of production to decouple this step from the mining phase. The additional information is used to generate a short-term grade control model. Ore polygons based on different grade bins and material types are generated based on the grade control model. Blast movement technology is used to incorporate and adjust ore polygons for horizontal movement during blasting. The moved polygons limits are marked in the field to guide the excavation.

Whenever possible, mined ore is delivered directly to the primary crusher to minimize 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.

Additional stockpiling of low-grade ore exists on MRS4 and MRS5. This material is intended for reclaim after ex-pit mining has concluded.

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.

Underground project
The 2024 PEA mine plan assumes a five-year development period (2025–2029), a threeyear operational ramp-up (2030–2032), and 12 years of production with an average mill feed of 11,200 t/d.

The mining method is transverse longhole open stoping (LHOS). The level spacing will be 40 m and the stope panel length will be 20 m. Mining will be transverse with a minimum stope width of 8.0 m, with a planned equivalent linear overbreak slough dilution of 1.5 m on the hanging wall and 1.0 m on the footwall.

The underground mine will be divided into six mining regions according to their location in relation to the open pit, and mineralization will be directed to the crusher once it is operational.

Access to all underground workings will be via a main decline and conveyor ramp for material.

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 crushing system comprises a single-stage gyratory crusher in open circuit with an apron feeder underneath. At a 125 mm closed side setting, its production rate ranges from 4,200 to 4,500 t/oh. Assuming an 80% availability, the primary crusher has a capacity of 81-86 kt/d; enough for 28 Mt/year. 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 has a live capacity corresponding to approximately 44,000 t and an overall capacity (live plus dead) of 176,000 t.

Ore is reclaimed from the storage pile 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 new scalping screen above the secondary crushers. The undersize of the screen goes to the SAG mill feed conveyor, while the oversize feeds the secondary cone crusher operated and then discharged onto the SAG mill feed conveyor. The crushed pebbles are also discharged on the SAG mill feed conveyor as well.

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. A re-feed conveyor was installed to minimize the impact on production of having the secondary crushers down. Material of -1" is fed to the SAG mill feed conveyor with this system. This system has a capacity of 1,600 t/h.

Crusher Improvement Initiatives
Scalping screens have been added ahead of the secondary crushers.

Removal of fines has unlocked secondary crusher capacity but has also significantly reduced the volumetric feed rate to the crusher. As a result, choke feed conditions are not steadily achieved, which causes uneven mantle wear and lower power utilization. These elements impact size reduction and the availability of the secondary crushers. Improvements can be made to resolve these issues such as the addition of a variable frequency drive (VFD) to the crusher’s motor. This modification aims to regulate the crusher rotation speed in order to maintain power draw and choke feed conditions.

The 1,000 hp motors installed on the secondary crushers are smaller than the theorical 1,200 hp motor that could be installed on the Raptor XL1100. By increasing the power draw, a finer product could potentially be produced with same crushers.

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 mine started a drill-and-blast optimization project that has resulted in higher throughput rates and better operational conditions.

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 still 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.

The increased fineness coming from the open pit ore due to improved blasting techniques, caused some issues with the primary crusher as the fines would not flow well and cause power and hydraulic pressure spikes tripping the crusher. Using a Hi-Fines mantle/concave design for the crusher, together with High-Fines mantle liners, will raise the choke zone and fill up more the crushing chamber, which will distribute wear more evenly and generate a finer particle size, and result in reduced power draw and hydraulic pressure reduction. It is planned to install these items once the SAG mill grates are converted to 60 mm (from 70 mm) and the SAG discharge screen to 8 mm. These modifications are needed to handle the finer product generated by the Hi-Fines mantle/concaves.

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 90%. The pregnant solution is pumped to a tank in the gold room followed by electrowinning in a dedicated cell. Modifications are planned to convert the gravity circuit feed from cyclone feed to cyclone underflow. The higher concentration of gold in the cyclone underflow should lead to a higher gravity recovery.

The target P80 at the cyclone overflow is 130 µm, representing the historical value at which the plant has been running. It has been observed that over this value, recovery is impacted. To maintain the ore fineness of the cyclone overflow, the transfer size between the SAG and the ball-mill will have to be reduced. To do so, two modifications are planned to the circuit. First, the installation of smaller SAG discharge grates to keep more material inside the SAG mill and secondly, the reduction of the aperture of the SAG discharge screen panels from 9.5 mm to 8.0 mm. The advanced process control to be installed will also have some features that will control the final grind size.

In 2023, a grinding circuit survey was performed and from this the circuit was modelled using a software called JKSimmet. The simulations using this software have confirmed that at 28 Mt throughput rates and by doing such modifications to the grinding circuit, the 130 µm grind size can be maintained.

Pebble Crusher Variable Speed Drives
By nature, a SAG mill generates pebbles at an uneven rate. The rate is influenced by ore feed size, mill speed and the load inside the mill. 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 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 discharged from the crusher. With a reduced particle size, the SAG mill throughput will increase.

Ore Blending
Ore blending is an avenue that the process plant has used successfully to optimize milling rates. About 30% of the plant feed is blasted with high intensity and some of it is stockpiled so the daily blend is even at 30%. However, there are still room to do better and the new avenue that will be explored is machine learning to blend the ore.

Increase Plant Operating Time to 92%
The operating time for the process plant is calculated when the ore is on the belt feeding the SAG mills. The design rate of 92% has not yet been achieved over an annual basis, instead averaging around 90%. Once the milling rates have stabilized, it is expected that the 92% operating time will be achieved.

Processing

• Gravity separation
• Intensive Cyanidation Reactor (ICR)
• Centrifugal concentrator
• Smelting
• Carbon re-activation kiln
• INCO sulfur dioxide/air process
• Crush & Screen plant
• Electric furnace
• Agitated tank (VAT) leaching
• Carbon in pulp (CIP)
• Elution
• Carbon adsorption-desorption-recovery (ADR)
• 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
Assistant General Manager	Jean Chateauneu			Mar 30, 2025
Deputy General Manager	Colin Ashton			Mar 30, 2025
General Manager	Michelle Moore			Mar 30, 2025
Health & Safety Superintendent	Douglas Brown			Mar 30, 2025
Mine Operations Superintendent	Paul Pepper			Mar 30, 2025
Mobile Maintenance Manager	Louis Gendron			Mar 30, 2025
Mobile Maintenance Superintendent	Derek Buzzi			Mar 30, 2025
Plant Maintenance Manager	Raphael Boutin			Mar 18, 2025
Plant Maintenance Superintendent	Eric Saulnier			Mar 18, 2025
Sr. Maintenance Superintendent	Jason Thibodeau			Mar 18, 2025
Sustainability & Environmental Manager	Melissa Leclair	melissa.leclair@agnicoeagle.com		Mar 18, 2025
VP Operations	Andre Leite	andre.leite@agnicoeagle.com		Mar 18, 2025