Caspiche Project — 2013 Technical Report

The 2013 technical report presents process designs for oxide heap leaching and sulphide flotation-concentrator routes at the Caspiche Project, based on testwork and design criteria established for two nominal throughput options.

Report context

The Caspiche Project technical report, dated 2013, describes recovery methods for both oxide and sulphide mineralised material. The oxide plant was designed for nominal throughputs of 30 ktpd and 60 ktpd, employing crushing, heap leaching, and adsorption-desorption-recovery (ADR) with electrowinning and smelting. The sulphide plant was designed for nominal throughputs of 12 ktpd and 27 ktpd, employing crushing, grinding, flotation, concentrate handling, and a gold recovery plant including CIL and SART circuits. The report also includes a validation of the 2012 Prefeasibility Study (PFS) design for a larger Super Pit option.

Processing route

Oxide plant , crushing

For the 30 ktpd option, 136 t capacity mine trucks deliver run of mine (ROM) material to the primary crusher dump pocket. For the 60 ktpd option, 185 t capacity trucks are used. Material is fed to the primary crusher by a vibrating grizzly feeder that scalps off undersize material less than 150 mm, which reports directly to the primary crusher discharge conveyor. Oversize feeds the primary jaw crusher. The primary crusher product and grizzly undersize are conveyed to a double deck banana screen. Both screens’ oversize material feed the secondary cone crusher. Screen undersize, less than 51 mm, is conveyed with the secondary crusher undersize to the heap leach load out bin, where material is transported by mine trucks to the heap leach pad.

Oxide plant , heap leaching

Material delivered by truck is levelled and arranged in modules of 10 m height by 90 m long and 40 m wide. Leaching is conducted on a permanent valley fill pad. Gold and silver are recovered using a weak cyanide solution that irrigates the stacked material using drippers over a 130 day cycle. The leach process uses two-stage leaching: fresh material is initially leached with intermediate leach solution (ILS) containing a medium gold tenor, then ILS recovered from the irrigated material reports to the pregnant leach solution (PLS) pond and is pumped to the carbon in column (CIC) circuit. Barren solution exiting the CIC is pumped to the barren solution tank and used to irrigate partially leached material; the collected solution reports to the ILS pond. The irrigation system design uses an irrigation rate of 10 l/hr/m² and a 130 day irrigation cycle, with 65 days using barren solution and 65 days using intermediate solution. Modules are arranged in cells of 90 m width.

The heap leach pad is located 1 km west of the oxide pit at an elevation of 4,250 m.a.s.l. The drainage system consists of a network of perforated HDPE pipes running from east to west towards telescopic longitudinal collector pipelines placed at every 90 m. Perimeter collector pipelines are placed at the downstream end of each cell. The base liner system consists of a 0.80 m thick overliner, a 2.0 mm thick LLDPE geomembrane primary barrier, a secondary barrier of asphalt underliner, and 0.5 m thick basal improvement. The leach pad loading plan stacks material from lower to upper levels, with each platform 10 m high and a maximum piezometric height of 100 m.

Oxide plant , ADR, electrowinning and smelting

The adsorption plant consists of a counter current carbon in column (CIC) train of five columns for the 30 ktpd option and two trains of five columns for the 60 ktpd option. Loaded carbon is acid washed with 3% hydrochloric acid to remove fouling. Gold and silver are desorbed from carbon by elution using a strong cyanide/caustic solution at temperatures of 135–140°C in two carbon steel columns. The gold and silver bearing solution is recirculated through two electrowinning cells in parallel where gold and silver precipitate as sludge or on cathodes. The doré sludge is filtered and smelted to produce bullion doré. Stripped carbon is reactivated in a rotary kiln every four cycles.

Sulphide plant , crushing and grinding

For the underground mining option, mineralised material from the underground mine is rehandled at the mine portal and transported to the primary crushing facility by 150 t capacity mine trucks for the first three years. From year four to end of mine life, an in-mine sizer partially crushes mine feed which is conveyed through a 4,150 m long tunnel to the primary crusher. For open pit options, mineralised material is transported to the primary crusher by mine trucks. The crusher discharges onto an apron feeder and material is conveyed to the coarse material stockpile.

The grinding circuit is designed to deliver a P80 of 130 µm, based on metallurgical tests indicating low to medium rock hardness. A single SAG mill with a short trommel and pebble screen is used. The SAG mill discharge slurry passes through a vibrating double deck screen with a minimum aperture of 13 mm. Screen undersize flows into a header tank supplying the cyclone feed pump. Cyclone cluster overflow feeds the flotation circuit and underflow returns to the SAG mill. Pebble oversize is conveyed to the pebble crusher plant; the pebble crushing plant assumes a nominal 20% pebble production rate with a maximum of 35% for design purposes.

Sulphide plant , flotation

The flowsheet consists of rougher flotation, rougher concentrate regrind, three stages of cleaner flotation, and one stage of scavenger flotation. Rougher concentrate is reground to a P80 of 30 μm. The three outputs are: Cu-Au concentrate from third cleaner flotation, sent to concentrate handling; pyrite concentrate from scavenger cleaner tails, sent to the gold plant; and tailings from rougher flotation, sent to tailings handling.

For the 27 ktpd options, rougher flotation is fed with pulp at 35% solids and considers two lines of three conventional flotation cells each of 300 m³ for Phase I, increasing to four cells each for Phase II. For the 12 ktpd option, rougher flotation uses two lines of four flotation cells of 160 m³. Residence time is 45 minutes. Rougher concentrate grade is estimated at 4.5% Cu, with copper recovery expected at 94% and gold recovery at 85%.

The regrind circuit targets a cut size P80 of 30 μm using a cluster of cyclones operating in reverse closed circuit with an estimated circulating load of 150%. Cyclone overflow reports to first cleaner flotation. For the 27 ktpd option, the first cleaner stage comprises three conventional flotation cells each of 40 m³ in Phase I, increasing to four in Phase II. The 12 ktpd option uses three flotation cells each of 30 m³. Second and third cleaner stages are arranged in-line. Second cleaner concentrate copper grade is estimated at 20% with 78% Cu recovery. Third cleaner concentrate contains approximately 26% Cu and 67 g/t Au.

Scavenger cleaner flotation receives first cleaner tailings. For the 27 ktpd option, three cells each of 70 m³ are used in Phase I, increasing to four in Phase II. The 12 ktpd option uses three cells each of 50 m³. Residence time is 30 minutes. Scavenger cleaner concentrate grade is estimated at 4.5% copper with copper recovery estimated at 83.6%. Tails of this stage report to the gold recovery plant.

Sulphide plant , concentrate handling and treatment

Final Cu-Au concentrate at 19% solids with a characteristic size D80 of 30 µm is thickened to 60% solids in a High-Rate thickener. A typical thickening rate of 0.11 t/h/m² is considered. Concentrates are pressure filtered and washed to obtain 8–9% moisture. A typical filtering rate of 300 kg/h/m² was used. The filtered concentrate, with design criteria copper grade of 26% and gold content of 67.4 g/t, and overall recovery of 91.7% copper and 74.5% gold, is transported by trucks 140 km to the roasting plant.

The final concentrate is expected to contain an average arsenic grade of 2.5%, which must be reduced to less than 0.2% for commercial viability. The roasting process is performed in a FluoSolids roaster at around 700°C and a controlled atmosphere. The calcine produced is approximately 90% of the original concentrate weight. Gases containing arsenic sulphides and SO₂ are cooled to 370–400°C and fed to an electrostatic precipitator. Arsenic trioxide is precipitated and treated with lime and ferric ions to produce ferric arsenate. SO₂ is converted to SO₃ to produce 98% sulphuric acid.

Sulphide plant , gold recovery

The pyrite concentrate from cleaner scavenger tails is treated in a gold recovery plant comprising thickening and filtering, CIL, cold desorption, acid washing, hot desorption, electrowinning, smelting, SART, and carbon reactivation. Design gold recovery in the rougher concentrate is estimated at approximately 80%. Final gold recovery is approximately 60% of the gold head grade.

The SART process recovers copper and regenerates cyanide. SART feed solution is mixed with sulphuric acid and sodium hydrosulphide and feeds a precipitation reactor where copper sulphide is precipitated at pH 5. Barren solution rich in free cyanide reports to neutralisation where milk of lime raises pH to approximately 10.5, then the solution is reused in the CIL plant. SART copper recovery is 98% and NaCN recovery is 91%, obtained from metallurgical tests.

Tailings handling

For underground mining with paste backfill, rougher flotation tailings are thickened to 55% in a High-Rate thickener and pressure filtered to 20% moisture. For open pit mining, tailings are thickened to around 60% and conducted to the tailings storage facility.

2012 PFS design validation

The 2012 PFS validation confirms the Super Pit option at 150,000 tpd sulphide production with 889 Mt ore at 0.24% Cu, 0.58 g/t Au and 1.13 g/t Ag. The validation identified that a longer leach cycle from 80 days to 130 days is now recommended to achieve 80% gold recovery, incorporating an ILS pond. The leach pad design review identified additional liners, piping and other materials costs. No changes were made to the sulphide plant process circuit and equipment sizing.

The 2012 PFS design for the Super Pit option considered two 75,000 tpd gyratory primary crushers, two identical parallel grinding circuits of one SAG mill and two ball mills per circuit at 75,000 tpd capacity, rougher flotation with three pairs of lines of eight 300 m³ cells, regrind to P80 of 30 µm, and cleaner flotation with global copper recovery estimated at 97.7%. The final concentrate was estimated at approximately 25% copper. Heap leach operations for oxide material at 26 Mt/y and MacNeill ore at 12 Mt/y were described, with MacNeill ore processed through a SART plant.

Key reported parameters

Technical qualifications

The technical report presents proposed designs for two processing routes: an oxide heap leach plant and a sulphide concentrator plant. The oxide plant design is based on testwork that provided head grades of 0.404 g/t Au and 1.65 g/t Ag, a work index of 13 kWh/st, and a specific solid gravity of 2.6. Heap leach gold recovery of 80% and silver recovery of 40% were calculated from testwork. The sulphide plant design is based on head grades of 0.45% Cu and 1.3 g/t Au from mine planning, a work index of 15.0 kWh/st from testwork, and specific solid gravity of 2.56 from metallurgical tests. Flotation recovery values and concentrate grades were obtained from metallurgical tests and industry practice standards. The 2012 PFS design validation confirms the Super Pit option as a valid large project alternative, with minor changes including a longer leach cycle from 80 to 130 days and incorporation of an ILS pond. The report does not present historical operating data or project economics. This technical report is based on design criteria and testwork only.

Source: Caspiche Project 2013 Technical Report, Section 17 , Recovery Methods.