Tulsequah Chief Project – Feasibility Study Technical Report

The report presents a designed processing plant using conventional differential flotation to produce copper, lead, and zinc concentrates from the Tulsequah Chief ore.

Report context

The Tulsequah Chief Project Feasibility Study Technical Report, with an effective date of October 20, 2014 and report date of November 27, 2014, describes a processing plant designed to treat ore from the Tulsequah Chief deposit. The process design criteria and flowsheets were developed based on metallurgical test work results from historical and current test work programs, using industrial design factors. The plant is envisioned to accept primary crushed ore from an underground storage bin at a rate of approximately 1,100 tpd, operating 24 hours per day for 365 days per year with a plant availability of 90%. The crusher is planned to operate for 16 hours per day.

Processing route

Primary Crushing and Ore Storage

The crushing area is planned to consist of a grizzly, a 120 t dump pocket, a jaw crusher, a belt feeder, and dust collection system. A dump pocket capable of receiving ore from underground trucks is located on the top level of the crushing area. A vibrating grizzly feeder draws ore from the dump pocket and provides a constant feed to the jaw crusher. Crushed ore discharges to the storage bin feed conveyor. A dust collector collects dust generated in the crushing area, with dust discharged onto the storage bin feed conveyor. Maximum feed lump size to the grizzly is 500 mm. The primary crusher is a 30-inch by 40-inch jaw crusher in open circuit configuration with 110 kW installed power and closed side setting of 70 to 75 mm. The estimated product P80 is 100 mm. A fine ore surge bin has 1,250 tonnes live capacity. Crushed ore is reclaimed by belt feeders that feed the SAG mill feed conveyor.

Grinding Circuit

The grinding circuit uses a SAG mill followed by two stages of ball milling to produce a final cyclone overflow product of 80% passing 45 microns. Reclaimed ore feeds a 448 kW, 2.4 m diameter by 4.6 m long SAG mill driven by a variable speed motor, enabling power draw variation for circuit optimization under varying feed conditions. SAG mill discharge feeds a 2.4 m x 6.1 m vibrating screen with 10 mm deck aperture. Screen undersize discharges into the primary cyclones feed pump box, while oversize feeds a stockpile via a coarse material discharge conveyor and is reclaimed back to the SAG mill feed chute.

SAG discharge screen undersize slurry is collected in the primary cyclones feed pump box with primary ball mill discharge and gravity concentrator No. 1 tailings, then pumped to the primary cyclones for size classification. The primary cyclopac consists of eight 250 mm diameter cyclones, with five operating. Underflow feeds ball mill No. 1 and gravity concentrator No. 1. Cyclone overflow reports to the final grinding circuit. The overflow from the primary cyclopac flows by gravity to the secondary cyclones feed pump box, combined with secondary ball mill No. 2 discharge and gravity concentrator No. 2 tailings. Slurry is pumped to the secondary cyclopac of eight 250 mm diameter cyclones, five operating, with designated cyclones feeding ball mill No. 2 and gravity concentrator No. 2.

Ball mill No. 1 is a 2.9 m x 4.9 m closed circuit ball mill with 448 kW installed power and product size P80 of 95 microns. Ball mill No. 2 is a 2.9 m x 4.9 m closed circuit ball mill with 448 kW total installed power and product size P80 of 45 microns. The target particle size P80 of the secondary cyclopac overflow at 30% solids is 45 microns.

Gravity Gold Circuit

Designated cyclones from each ball mill circuit are equipped with gravity concentrators to recover gravity gold (electrum). Slurry from designated cyclones flows by gravity over concentrator trash screens, with oversize reporting to the cyclone feed pump box and undersize to 20-inch centrifugal gravity concentrators. The gravity concentrators each receive 100% of new feed to achieve a combined target of 41% gold recovery. Gravity gold concentrates report to intensive cyanide leach and electrowinning circuits to produce doré. The pregnant solution is pumped to an electrowinning circuit. The smelter is expected to run 2 batches per week producing 4.54 kg per batch. Leach circuit residue is pumped to the grinding circuit, and barren solution is used in the cyanide reagent mixing tank.

Copper Flotation

Secondary cyclone overflow flows by gravity to the copper conditioning tank where SMBS, A9810 and MIBC are added prior to rougher flotation. The slurry then gravitates to the rougher flotation circuit consisting of one bank of four 38 m³ conventional cells. Rougher concentrate froth feeds the copper cleaner flotation circuit. Copper rougher tailings and first cleaner tailings are combined and pumped to the lead flotation circuit conditioning tank.

The copper cleaner circuit comprises four 5.8 m³ first cleaner cells and four 5.8 m³ second cleaner cells. Rougher concentrate feeds the first cleaner cells. First cleaner concentrate flows by gravity to the second cleaner flotation circuit. Second cleaner concentrate flows by gravity to one of two copper thickeners. Two additional banks of four cleaner flotation cells and an additional copper thickener have been included to allow flexibility to produce low and high arsenic concentrates.

Copper Concentrate Dewatering

Copper concentrate dewatering uses two 4 m diameter high rate thickeners, each achieving 60% underflow density. Thickened slurry is transferred to an agitated concentrate stock tank providing 8 hours surge capacity, then fed to one horizontal pressure filter. Target concentrate moisture content is 8%. Filtered concentrate discharges onto a conveyor feeding the bagging system, which discharges into 2-tonne bags that are weighed and tagged before storage near the barge landing.

Lead Flotation

Copper flotation tailings are pumped to the lead flotation conditioning tank followed by one bank of four 38 m³ rougher cells. Rougher concentrate feeds the first of two stages of cleaning incorporating four 5.8 m³ cells each. The lead cleaner circuit comprises four 5.8 m³ first cleaner cells and four 5.8 m³ second cleaner cells. Second cleaner concentrate flows by gravity to the lead thickener. Lime, NaCN, ZnSO4, Cytec 3418a, and MIBC are added to enable flotation. Lead rougher tailings are delivered to the zinc flotation circuit.

Lead Concentrate Dewatering

A 3 m diameter high rate thickener thickens lead concentrates to 60% underflow density. Thickened slurry transfers to an agitated stock tank providing 8 hours surge capacity, then feeds a horizontal pressure filter. Target concentrate moisture content is 8%. Lead concentrate cake drops onto a conveyor feeding the bagging system for storage in 2-tonne bags.

Zinc Flotation

Lead circuit tailings are pumped to two zinc conditioning tanks. In the first tank lime is added, and in the second tank CuSO4, Cytec 7021 and MIBC are added to increase and control pH at 10.5, suppress lead and pyrite, and activate sphalerite. Conditioned slurry feeds the rougher flotation circuit consisting of two banks of four 38 m³ cells. Rougher concentrate froth feeds the cleaner circuit comprising two banks of four 5.8 m³ first cleaner cells followed by two banks of four 5.8 m³ second cleaner cells. First cleaner tailings combine with rougher tailings as pyrite flotation circuit feed.

Zinc Concentrate Dewatering

A 6.5 m diameter high rate thickener thickens zinc concentrates to 60% underflow density. Thickened slurry transfers to an agitated stock tank providing 8 hours surge capacity, then feeds a horizontal pressure filter. Target concentrate moisture content is 8%. Concentrate cake falls onto a conveyor feeding the bagging system for storage in 2-tonne bags.

Pyrite Flotation

Zinc rougher and zinc first cleaner tailings are fed to the pyrite flotation conditioning tank followed by one bank of four 38 m³ rougher cells. PAX and MIBC are added to enable pyrite flotation. Rougher concentrate gravitates to the pyrite thickener, and pyrite rougher tailings go to the final tailings thickener.

Pyrite Concentrate Dewatering

A 12 m diameter high rate thickener thickens pyrite concentrate to 50% underflow density. Thickened slurry transfers to the pyrite pond during the first two years of operation and reports to the paste plant after year 2.

Tailings Management

Pyrite flotation tailings report to the final tailings thickener. Thickener underflow is pumped to the paste backfill plant or collected in the final tailings pump box and pumped to the tailings management facility (TMF). Tailings not required for paste backfill are pumped to the tailings pond. A reclaim barge with two 75 kW pumps located on the TMF recovers process water for plant make-up. Excess process water is sent to the effluent treatment plant and then pumped to the plant as fresh make-up water or discharged to the environment.

Key reported parameters

Technical qualifications

The report states that process design criteria and flowsheets were developed based on metallurgical test work results from historical and current test work programs as described in Section 13, using industrial design factors. The flotation circuits are sized based on test work from ALS Project T0662 data, with laboratory retention time scaled up by two and a half times for roughers and four times for cleaners. Flotation cell launder design was based on lip loading of 200 kg/m/hr for roughers and 100 kg/m/hr for cleaners. The report notes that flocculant dosage has been estimated pending further test work. It is recommended that compressed air requirements be re-evaluated in the next stage of engineering once the filter supplier has been identified.

The report distinguishes between proposed design, historical operating data, and testwork results. The gravity gold circuit design criteria source is FLSmidth Knelson modelling and design. Concentrate dewatering and filtration design criteria are based on vendor data and recommendations. The grinding circuit design parameters including work indices and abrasion index are from testwork.

Source: Tulsequah Chief Project – Feasibility Study Technical Report, effective date October 20, 2014, report date November 27, 2014, Section 17 Recovery Methods.