La Cigarra Silver Project Process Design Balances Oxide Leaching with Sulphide Flotation

The PEA outlines a 6,000 t/d plant designed for flexibility, with separate circuits for oxide and sulphide ore and a shared Merrill–Crowe recovery stage.

The La Cigarra Silver Project is an advanced-stage project located in the State of Chihuahua, north-central Mexico, approximately 28 km northwest of Hidalgo del Parral. Owned 100% by Kootenay Silver Inc., the project is centered on a silver-dominant, polymetallic deposit targeting silver, gold, lead, and zinc. A Preliminary Economic Assessment (PEA) with an effective date of January 15, 2026, evaluates the potential for an open-pit operation. The project has no permanent mining or processing infrastructure on site, and the PEA assumes development of all required facilities, with a planned mill feed of approximately 6,000 t/d over a 15-year project life including a pre-strip phase.

Critical Data

Overview

The process plant design draws on metallurgical testing programs completed by G&T in 2011, BaseMet in 2015, and ALS Metallurgy in 2017. Those programs define two processing routes. Oxide material goes through conventional crushing, grinding, and classification followed by cyanide leaching. Silver and gold are then recovered from the pregnant solution using a countercurrent decantation (CCD) circuit and the Merrill–Crowe process. Sulphide material is treated by flotation to produce saleable lead and zinc concentrates, with a pyrite flotation step capturing silver and gold values for downstream leaching.

The plant is designed to handle oxide ore, sulphide ore, or a blend, depending on the mine plan. The plant can switch between processing oxide and sulphide material in campaigns, or process a mixture of the two. The core equipment train is shared, but separate thickeners and leach tank allocations support the different processing routes. Blended ore uses the same flotation facilities as the sulphide circuit.

The comminution circuit includes a primary jaw crusher, secondary and tertiary cone crushers, and a ball mill in closed circuit with cyclones. Design work indexes are 12.3 kWh/t for crushing and 16.9 kWh/t for the Bond ball mill work index. A crushed material rock management facility (RMF) with 6,000 tonnes live capacity provides surge between crushing and grinding, with reclaim belt feeders supplying the ball mill at a design rate of 326 t/h.

Key Process Stages

The crushing circuit reduces run-of-mine material from a nominal top size of 700 mm to a product size of P80 9.5 mm at a design throughput of 411 t/h. Feed is delivered by 100 tonne haulage trucks into a hopper with 150 tonnes surge capacity. A static grizzly with 700 mm apertures prevents oversize material from entering the circuit, and a rock breaker is available to clear blockages. The primary jaw crusher reduces material to P80 108 mm with a closed-side setting of 90 mm. A double-deck screen with 32 mm top deck and 14 mm bottom deck apertures operates in closed circuit with the secondary and tertiary cone crushers.

Grinding takes place in a single ball mill 6.0 m in diameter and 12.0 m long, with 7,800 kW installed power. The mill operates at 75% of critical speed with a circulating load of 400%. Water addition maintains slurry density at 70% solids. The mill discharge passes over a trommel and mill discharge screen before entering a pump box that feeds a cluster of 12 classification cyclones. The cyclone underflow returns to the ball mill, while overflow at P80 53 µm moves to a vibrating trash screen with 2.0 mm by 8.0 mm apertures.

For oxide material, the trash screen underflow flows by gravity to a 30 m diameter pre-leach thickener. Flocculant aids settling, and the underflow is thickened to 50% solids before entering the leach circuit. The leach circuit consists of seven tanks, each 16.0 m in diameter and 16.0 m high. Lime maintains pH between 10.5 and 11.0. Oxygen is sparged into each tank to maintain dissolved oxygen above 20 mg/L, which is intended to counter preg-robbing effects from carbon in the oxide material. Sodium cyanide solution is added at the head of the circuit and at staggered points along the train to maintain a concentration of about 1,000 mg/L. Total retention time is about 48 hours. The leached slurry exits the last tank into the CCD circuit.

The oxide CCD circuit uses four thickeners, each 30 m in diameter. Pregnant solution overflowing the first thickener is pumped to the Merrill–Crowe facility. Barren solution from precipitation feeds the last thickener. Flocculant is added to each thickener feed, with slurry feed diluted to about 30% solids and final underflow at about 50% solids. The theoretical washing efficiency is about 99.8%.

Sulphide material follows a different path after grinding. The slurry from the trash screen goes directly to flotation at a design feed rate of 326 t/h. The lead circuit uses a conditioning tank with two minutes retention, where reagents include lime, sodium cyanide and zinc sulphate as depressants, SIPX and Aerophine 3418A as collectors, and MIBC as frother. The lead rougher stage has five cells of 50 m³ each. Lead rougher concentrate represents about 4.0% of plant feed and is reground in a tower mill from P80 53 µm to P80 30 µm before two cleaner stages. The first cleaner has five 3 m³ cells and the second has one 1.5 m³ cell.

Lead flotation tailings go to the zinc circuit, which also has a two-minute conditioning stage. Reagents include copper sulphate as an activator, SIPX as collector, and MIBC as frother. The zinc rougher stage has five 50 m³ cells. Zinc rougher concentrate represents about 3.1% of plant feed and is reground to P80 19 µm before its two cleaner stages. The first cleaner stage has four 3 m³ cells and the second has one 1.5 m³ cell. Zinc first-cleaner tailings are routed to the leach circuit.

Pyrite flotation follows the zinc circuit. Six rougher cells of 50 m³ each recover pyrite concentrate, which represents about 12.8% of plant feed. This concentrate is reground in a tower mill to P80 15 µm and combined with the zinc first-cleaner tailings for cyanide leaching. The combined stream goes to a dedicated pre-leach thickener and then uses the existing oxide leach circuit, drawing on four of the seven leach tanks. The cyanide concentration required in the sulphide leach feed slurry is about 5,000 mg/L as sodium cyanide, with about 1,200 mg/L remaining in the residue.

The sulphide CCD circuit uses four thickeners, each 16 m in diameter. Pregnant solution reports to the shared Merrill–Crowe facility. Final thickener underflow at about 50% solids carries a cyanide concentration of about 1,200 mg/L CN_WAD. The sulphide leach circuit operates under the same conditions as the oxide circuit. When blended ore is processed, the reground pyrite concentrate is sent to the first two leach tanks, while the zinc first-cleaner tailings go to the remaining five tanks. The blended circuit uses the same CCD and detoxification facilities as the sulphide circuit.

Cyanide detoxification treats tailings before discharge to the tailings management facility (TMF). The oxide circuit uses two tanks, each 11.0 m in diameter and 12.0 m high. The sulphide circuit uses two smaller tanks, each 7.0 m in diameter and 8.0 m high. The detoxification process uses copper sulphate and sodium metabisulphite to generate sulphur dioxide, with lime maintaining pH at about 8.5. Oxygen is added to support the reaction. Retention time is a minimum of two hours. The target discharge concentration is less than 20 mg/L CN_WAD, down from a feed concentration of about 300 mg/L for oxide tailings and about 1,200 mg/L for sulphide tailings.

The Merrill–Crowe circuit recovers silver and gold from pregnant solution. The solution is clarified through filters pre-coated with diatomaceous earth. De-aeration takes place in a Crowe tower with vacuum pumping to reduce dissolved oxygen below 1 ppm. Zinc dust and lead nitrate are added to a mixing tank for precipitation. The resulting sludge is collected on a filter press, then dried and smelted. A retort furnace removes mercury that may be present in the precipitate. Smelting takes place in an induction furnace, with flux materials including borax and soda ash. The doré bars are cleaned, weighed, and stored in a secure safe until dispatch. The refinery includes gas monitors for hydrogen cyanide, hydrogen sulphide, and mercury vapour.

Additional Interesting Data and Summary

The lead and zinc second cleaner concentrates are the final flotation products. Each is thickened in a 3.0 m diameter thickener, with underflow at about 55% solids pumped to an agitated stock tank with eight hours storage. A filter press dewaters each concentrate to about 8% moisture. The dewatered concentrate is stored for about five days of production before being bagged and loaded into trucks for shipment to off-site smelters.

Water management is a key design element. Process water is recovered from the pre-leach thickener, the tailings thickener, the concentrate thickeners, and the TMF via a reclaim barge. Fresh water is drawn from the on-site San Felipe–Punta Estrella aquifer through borehole wells. A fresh- and fire-water tank with 1,200 m³ capacity holds water for half a shift, with a standpipe maintaining at least 600 m³ for two hours of fire water supply. Rainwater runoff is collected in earthen dams during the rainy season.

Air supply is split into three systems. Low-pressure air for flotation cells comes from three centrifugal air blowers. Instrument air is dried and stored in a dedicated receiver. High-pressure plant air is provided by air compressors. Oxygen for the leach and detoxification circuits is generated on site using vacuum swing adsorption technology with four units for redundancy.

The reagent handling facility is designed with spill containment for 110% of the largest tank content. Reagents include hydrated lime for pH control, SIPX and Aerophine 3418A as collectors, MIBC as frother, flocculant for settling, sodium cyanide for leaching, sodium metabisulphite for detoxification, copper sulphate as an activator, zinc sulphate as a depressant, and lead nitrate for precipitation. The cyanide preparation area is isolated with restricted access and hydrogen cyanide monitoring.

Key Processes

• Crushing: Primary jaw crusher with secondary and tertiary cone crushers in closed circuit with a double-deck screen to produce P80 9.5 mm.
• Grinding: Single ball mill with 7,800 kW installed power in closed circuit with 12 classification cyclones to produce P80 53 µm.
• Oxide Leaching: Seven leach tanks with 48 hours retention, cyanide concentration about 1,000 mg/L, and oxygen sparging above 20 mg/L.
• Sulphide Flotation: Lead, zinc, and pyrite rougher flotation circuits with regrind mills and two cleaner stages for lead and zinc.
• Concentrate Dewatering: Filter presses reduce moisture to about 8% before off-site shipment.
• Merrill–Crowe Recovery: Clarification, de-aeration, zinc dust precipitation, and smelting to produce doré bars.
• Cyanide Detoxification: Modified INCO procedure using copper sulphate, sodium metabisulphite, and oxygen to reduce CN_WAD to below 20 mg/L.
• Water Management: Reclaim from thickeners and TMF with fresh water from the San Felipe–Punta Estrella aquifer.

Source: Preliminary Economic Assessment for La Cigarra Silver Project, July 21, 2026. Project website: La Cigarra Silver Project