The Ana Paula processing route is designed around flotation of sulphides followed by atmospheric oxidation and cyanide leaching, with gravity recovery for a portion of the gold.
Report context
The Ana Paula Project 2023 Technical Report was prepared in accordance with National Instrument 43-101 and bears the date 09 March 2023, Revision 1. The process design described in the report was prepared for a proposed plant with nameplate capacity of 5,000 tonnes per day at a time when Heliostar had tasked M3 with the plant design. The report includes an overall process flowsheet, process design criteria, and descriptions of comminution, flotation, oxidation, leaching, carbon handling, cyanide destruction, water balance, tailings transport, and other processing areas. The design is presented as a proposed flowsheet based on metallurgical testwork and previous equipment purchases.
Processing route
Process description
Metallurgical tests and mineralogical analyses were performed on composites of the Ana Paula ore. The results indicate an average-hardness ore with a portion of the gold content being refractory due to encapsulation in iron sulphides. The rest of the gold can be liberated with normal grinding and recovered by gravity concentration or direct cyanidation.
The process flowsheet that was developed is based on the ore mineralogy and its response to metallurgical treatment. Run-of-mine ore is crushed and ground to 80 percent finer than 160 microns, processed by froth flotation to recover sulphides and free gold, followed by atmospheric oxidation of the sulphide concentrate, and cyanide leaching of the oxidized slurry. The report contains a simplified schematic of the overall process for the Ana Paula plant.
Process design criteria
The design basis uses a nameplate capacity of 5,000 t/d. The current mine plan is based on a 365-day calendar year, totalling 1,825,000 tonnes of ore per year.
The design assumes an overall mill availability of 92%, with the primary crusher at 75% and the pebble crusher at 85%. Availability is defined in the report as the estimated actual run time of equipment. Nomenclature and tracking parameters may vary from operation to operation.
The mass balance was developed using MetSim software. The process simulation assumed overall head grades and flotation recoveries of 2.24 g/t Au at 95% recovery and 6.89 g/t Ag at 89% recovery, with a flotation mass pull of 20%.
Comminution plant design
Primary crushing and pebble crushing were simulated using Metso's Bruno simulator. SAG milling was simulated using JKSimMet, and ball mill capacity was calculated using standard Bond equations, with and without taking credit for SAG mill slimes that bypass the ball mill.
Grindability parameters were measured for only four composites, each representing an ore type. With the limited number of data, the average or median hardness and the 80th percentile hardness were very close to each other. The A and b JK parameters were measured using the JK Rotary Breakage Tests and have not been calibrated with full drop-weight tests.
Primary crushing and coarse ore stockpile
ROM ore will be transported by 60-tonne haul trucks to the primary crusher and dumped into a dump hopper with an approximate 120-tonne live capacity (2 truckloads) through a stationary grizzly with an 800 mm opening. An apron feeder moves the ore from the hopper to another stationary grizzly with a 100 mm opening. The oversize reports to the primary crusher while the undersize reports directly to the transfer conveyor, bypassing the primary crusher.
The primary crusher is a 42" x 48" Kolberg-Pioneer jaw crusher, with a closed-side setting of 150 mm, powered by a 187 kW (250 hp) motor. The crushed ore is discharged to the transfer conveyor, which is also the stacking conveyor feeding the coarse-ore stockpile. A self-cleaning magnet followed by a metal detector remove tramp steel before stockpiling.
The coarse ore stockpile has a live capacity of 10,000 tonnes and a total capacity of 50,650 tonnes. The live capacity is equivalent to 2 days of SAG mill feed at nominal capacity. Crushed ore is reclaimed via a reclaim tunnel with three reclaim feeders (two operating and one standby) onto the SAG mill feed conveyor. Each feeder is 1,067 mm wide and 6.5 m long, powered by an 11 kW (15 hp) drive on variable frequency drives. The design reclaim rate is 226.4 t/h.
Dust suppression is accomplished by water sprays at the crusher dump hopper, jaw crusher, and at the discharge points of the feeders. A belt scale is included on the SAG mill feed conveyor after the feeders and before the point of addition of SAG mill balls.
Grinding and pebble crushing
The grinding circuit is a conventional SAG-mill, ball-mill, pebble-crushing system (SABC). The grinding line comprises one SAG mill, a pebble wash screen, one ball mill, one cyclone cluster, and a pebble crusher. The SAG mill is in a closed circuit with the screen and pebble crusher. The ball mill is in a closed circuit with the hydrocyclone cluster.
The SAG feed conveyor feeds ore to the SAG mill, 7.32 m diameter by 2.74 m effective grinding length (24 ft x 9 ft EGL), powered by a new 2,872 kW (3,850 hp) drive on VFD. The SAG mill product discharges to a pebble wash screen.
Pebbles separated by the SAG mill discharge screen are conveyed to the pebble crusher feed bin and crushed with a Metso HP100 cone crusher, or equivalent, set at a closed-side setting of 13 mm. The crushed pebbles are returned to the SAG mill via the SAG mill feed conveyor.
The undersize of the SAG mill discharge screen drops into the cyclone feed pump box. This constitutes fresh feed to the ball mill and mixes with the ball mill discharge and dilution water. The mixed slurry is pumped to a cluster of five 26-inch hydrocyclones (4 operating, 1 standby). Pumping is by a 260 kW (350 hp) Warman pump with a second pump installed as standby. Both motors are controlled by medium-voltage variable frequency drives.
The cyclone cluster underflow flows by gravity to the ball mill, 4.72 m diameter by 6.55 m length, driven by a fixed-speed 2,313 kW (3,100 hp) motor. The cyclone overflow is the product of the grinding circuit and is fed to the flotation circuit. The target grind size is 80 percent finer than 160 microns.
Gravity concentration
A split from the hydrocyclone overflow is processed for gold recovery by gravity concentration and intensive cyanidation. Gravity concentration is achieved using a centrifugal concentrator. The gravity concentrate is leached with cyanide in the presence of an oxidizer using an intensive leach package. The pregnant solution is sent to the same electrowinning circuit serving the oxidized concentrate leach circuit.
Based on testwork, approximately 40% of the gold is gravity recoverable. In the design, 36% of the circulating load is treated in the gravity concentration and intensive cyanidation section, which corresponds to approximately 20% of the gold being recovered by the gravity circuit.
The gravity concentrator is a centrifugal gravity concentrator, a Knelson KC-QS40 or equivalent, fed from the undersize of a vibrating screen (1.8 m x 4.8 m screening module). The machine is operated in batch mode on a set cycle to concentrate and then flush heavy materials to a downstream leach system. Tailing from the gravity operation is discharged to the ball mill feed chute with the cyclone underflow.
The gravity concentrate is fed to an intensive cyanidation package, a Consep Acacia CS200 or equivalent, through a reactor feed tank. The pregnant solution is pumped to the electrowinning cells that mainly process pregnant solution from the carbon-in-leach (CIL) process. Tailing from the intensive cyanidation reactor is pumped to the CIL process.
Flotation
Sulphides in the ore will be floated at the ore's natural pH using potassium amyl-xanthate (PAX) as collector, AERO 3418A as promoter, copper sulfate as activator, and F131A as frother.
The average laboratory rougher flotation time determined during bench-scale tests is 16 minutes. With a scale-up factor of 2 and 15% aeration at 30% solids, the design requires 36.8 minutes of plant residence time and a total volume of 377 cubic metres. This calls for 6 units of 70 cubic metre flotation cells.
Flotation of sulphides will be accomplished in a single rougher flotation stage. Cyclone overflow is first sent to a 41.2 cubic metre conditioning tank, then to a bank of six 70 cubic metre tank flotation cells. Each flotation cell mechanism is driven by a 93 kW (125 hp) motor through a gear reducer. Flotation air is supplied by a 70-kW (94-hp) blower, which can deliver 95 Nm3/min of air.
Flotation concentrates advance to the concentrate thickener and then to the regrind mill. The flotation tailing slurry is pumped to a flotation tailing thickener (28 m diameter high-rate thickener) to be thickened to 55% solids, in preparation for pumping to the tailing storage facility.
Concentrate thickening and regrind
Concentrate from the rougher flotation circuit is dewatered in a 10.5 m diameter high-rate thickener to a pulp density of 55% solids. Flocculant is added to the thickener feed. The withdrawal rate of settled solids is controlled by one of two underflow pumps, each driven by a 30 kW (40 hp) motor on variable frequency controller to deliver slurry at a nominal maximum rate of 65 cubic metres per hour. Underflow from the concentrate thickener is pumped using variable speed horizontal centrifugal slurry pumps to the regrind mill feed box.
Concentrate thickener overflow is pumped to the reclaim solution tank using two fixed speed horizontal centrifugal pumps, one operating and one standby, each driven by a 30 kW (40 hp) motor with a nominal capacity of 100 cubic metres per hour. The high-rate concentrate thickener is mounted on steel legs on foundations, with a concrete containment area to contain rain runoff and process spills.
From the regrind mill feed box, the thickened concentrate is pumped to the regrind mill with two variable speed feed pumps, one operating and one standby. Each pump is driven by a 45 kW (60 hp) motor to deliver a nominal maximum flowrate of 81 cubic metres per hour.
The concentrate regrind mill is a 900-kW tower mill with ceramic grinding media. It operates in open circuit while being monitored by an online particle size analyzer. The target grind of 80% finer than 25 microns is attained by controlling mill speed with a variable frequency drive. The regrind mill discharge is pumped using two horizontal centrifugal pumps, one operating and one standby, to the atmospheric oxidation feed box, with 30 kW (40 hp) drives.
Atmospheric oxidation
Atmospheric oxidation (AOX) of the sulphide concentrate is conducted in five agitated tanks. Each tank is 9 metres in diameter and 10 metres high (operating volume of 608 cubic metres), made of 2205 duplex stainless steel. Slurry is fed to each tank through a downcomer and overflows to the next tank or a pump box at the end of the series. Each agitator is powered by a 56-kW (75-hp) motor through a gear reducer. Oxygen is injected into each tank through fine-bubble spargers.
The reaction kinetics were optimized in the laboratory at around 75 degrees C. The reaction is exothermic, so the process is expected to be autothermic if the feed concentrate grade is kept at 10% sulphide sulphur or higher.
Carbon-in-leach (cyanidation)
The oxidized slurry flows to two neutralization tanks (6-m diameter, 7-m high, 185-cubic metre operating capacity) where lime is added to increase the pH to 10 to 10.5. The neutralized slurry is then pumped to a pre-leach thickener (13.7-m diameter) to increase the pulp density to 40 to 45% solids. Once thickened, slurry is pumped to the carbon-in-leach feed tank where it combines dilution water, sodium cyanide reagent feed, and other process streams, into the first CIL tank.
Cyanide leaching is achieved in six CIL tanks (9.8-m diameter, 9.8-m high, 696-cubic metre operating capacity) each equipped with 30-kW (40-hp) agitators with two narrow-blade hydrofoil impellers. The tanks are built with epoxy-coated mild steel. Air is delivered by a pipe under an inverted cone located directly below the agitator.
Based on leaching test results, a residence time of 48 hours is sufficient to achieve the target recovery for both gold and silver. After leaching, loaded activated carbon is sent to the carbon plant for stripping and electrowinning.
Carbon handling, elution, and electrowinning
Loaded carbon is first acid washed with a dilute solution of hydrochloric acid to remove scale, rinsed, and then pumped to the carbon stripping vessel. The carbon strip vessel is a pressure vessel with capacity to strip 5 tonnes of carbon per batch. The stripping process follows the pressure Zadra procedure developed by the US Bureau of Mines. It involves contacting a hot solution of cyanide and caustic (0.15% cyanide, 1.25% caustic) at a rate of 2 bed volumes per hour. The solution is introduced at the bottom of the carbon bed and overflows at the top through one or more cylindrical Johnson screens. The solution is preheated to 135 degrees C by heat exchangers. The strip vessel is kept at about 550 kPa to prevent boiling.
During stripping, gold and silver desorb from the activated carbon into the strip solution. This loaded strip solution is sent to electrowinning cells through a heat exchanger. In the electrowinning cells, gold and silver are deposited by electrolysis to a stainless-steel cathode. When enough is deposited, gold and silver are pressure washed off the cathodes and collected as sludge. The sludge is discharged to a tank and filtered through a plate-and-frame filter press.
The filtered residue is dried in retorts to remove and collect any mercury and smelted in a tilting furnace. Metallic gold and silver melt is poured into bar molds to produce doré bars.
Cyanide destruction
Residual weak-acid dissociable (WAD) cyanide in the leach tailing is destroyed by oxidation using oxygen from air and sodium metabisulfite. Milk-of-lime is added to maintain a slurry pH in the range of 8.0 to 8.5. The reaction is catalyzed by copper (5 ppm), which will need to be supplied if the ore does not contain enough cyanide-soluble copper.
Cyanide is oxidized first to cyanate, which eventually decomposes to carbon dioxide, ammonia, and nitrogen gas. The more stable iron cyanides are precipitated from solution as insoluble ferrocyanide compounds. The cyanide levels in solution are reduced to an environmentally acceptable level (less than 50 ppm WAD cyanide, per the Cyanide Code). The detoxified slurry is sampled prior to thickening.
The detoxification reactors are two agitated tanks operated in series, with a total volume of 315 cubic metres and a total residence time of approximately 3 hours. Slurry discharged from the detoxification circuit is pumped to the tailing thickener (28 m diameter) by two 56 kW (75 hp) horizontal centrifugal pumps, one operating and one standby.
Water balance and solution management
A water balance was developed as part of the mass balance model using MetSim modeling software. The estimated raw water requirement is 83.8 cubic metres per hour, of which 66.5 cubic metres per hour is for mill operations, 5 cubic metres per hour for potable water use, and the equivalent flowrate of 10 cubic metres per hour for mine dust control. Raw water supply comprises 31 cubic metres per hour from the well field, 52.8 cubic metres per hour from the rainfall diversion channel runoff, and 8.6 cubic metres per hour contained in the ore as moisture.
Well water use in the mill includes 11.6 cubic metres per hour for gland seal, 1.4 cubic metres per hour process water makeup, and 2.1 cubic metres per hour for crushing plant dust control. Fire protection water is also derived from well water. All runoff water is used as mill makeup water, introduced to the mill through the tailing storage facility and stored in the reclaim water tank with the tailing thickener overflow.
Mill operations use a total of 704.4 cubic metres per hour of water, of which 90.6% (638 cubic metres per hour) is water recycled from the TSF, the tailing thickener, the preleach thickener overflow, and the concentrate thickener overflow.
The reclaim solution tank (800 cubic metre capacity) functions as the process water tank. It is supplied with tailing pond reclaim solution, concentrate thickener overflow, pre-leach thickener overflow, tailing thickener overflow, and raw water. Process water is pumped by two horizontal centrifugal process water pumps, one operating and one standby, each powered by 150 kW motors, and is supplied to the grinding circuit, flotation circuit, concentrate oxidation circuit, process water tank as make-up, refinery scrubber, and flocculant systems.
Tailing slurry transport
Thickened tailing is discharged to a final tailing tank, from which the slurry is pumped using two fixed speed horizontal centrifugal pumps (56 kW, 274 cubic metres per hour) to the tailing storage facility (TSF). The tailing pipeline is a DN250/PN16 PE100 HDPE pipe, 2,700 m long, 250 mm bore, and will distribute tailing to Zone A spigots as well as to the dump spigot. This pipe connects to a 600 m long, 150 mm bore DN150/PN10 PE100 HDPE distribution header that will deposit tailing through Zones B and C spigots.
Solution from the pond reservoir is reclaimed by two 75 kW barge-mounted turbine pumps, one operating and one standby. The reclaim solution is pumped to the reclaim solution tank through a 700-m long DN225/PN20 PE100 HDPE pipe.
Sodium carbonate handling
Sodium carbonate is delivered to the site by trucks and off loaded to two 1700-tonne silo systems, providing enough storage capacity to supply 28 days of operation. Sodium carbonate is added as a solution to the regrind ball mill and to the oxidation tanks, diluted in an automatic dilution system located below the silos.
Mill power consumption
The average annual power consumption in the process plant is tabulated in the report. The estimated life of mine consumption totals 495.5 million kWh, which translates to about 35.1 kWh/tonne of ore processed.
Process control system
A central control room (CCR) is provided in the concentrator grinding facility core as the main operating control center. From the CCR, primary crushing, material handling systems, grinding and flotation, reagents, tailing, and utility systems are monitored and/or controlled. A computer room located adjacent to the CCR contains engineering workstations, a supervisory computer, historical trend system, management information systems server, programming terminal, network and communications equipment, and documentation printers. This is primarily used for Distributed Control System (DCS) development and support.
The DCS will use an Industrial Data Center (IDC), Programmable Logic Controllers (PLC) and Thin Clients or personal computers connected together with a fiber optic network using the Ethernet protocol. Interactive screens on monitors allow process control. A supervisory expert system will not be incorporated at this time.
Mobile equipment
Mobile equipment was included as part of the El Sauzal plant purchase and the Project capital cost estimate. Equipment listed includes a fire truck, ambulance, water truck, maintenance service trucks, telescopic crane, all-terrain crane, manlift, telehandler, mini loaders, and fork lifts.
Key reported parameters
| Parameter | Value | Basis |
|---|---|---|
| Nominal capacity | 5,000 t/d | Design |
| Annual throughput | 1,825,000 t/y | Design (365-day year) |
| General mill availability | 92% | Design |
| Primary crusher availability | 75% | Design |
| Pebble crusher availability | 85% | Design |
| Head grade Au | 2.24 g/t | Mass balance simulation |
| Head grade Ag | 6.89 g/t | Mass balance simulation |
| Flotation recovery Au | 95% | Mass balance simulation |
| Flotation recovery Ag | 89% | Mass balance simulation |
| Flotation mass pull | 20% | Mass balance simulation |
| Crushing work index | 13 kWh/t (assumed) | Design assumption |
| Primary crusher feed F80 | 500 mm | Simulated |
| Primary crusher product P80 | 126 mm | Simulated |
| SAG mill feed F80 | 126 mm | Design |
| SAG mill parameters (median) | A=54.58, b=0.806, ta=0.424 | JKRBT testwork, uncalibrated |
| Ore SG | 2.7 | Testwork |
| SAG mill power draw at target capacity | 2,274 kW (3,048 hp) | JKSimMet simulation |
| Ball mill feed F80 | 2,908 microns | Design |
| Ball mill product P80 | 160 microns | Design |
| Ball mill work index (average) | 18.3 kWh/t | Testwork |
| Bond abrasion index (average) | 0.1782 g | Testwork |
| Laboratory flotation time | 16 min | Testwork |
| Scale-up factor (flotation) | 2 | Design |
| Plant flotation time | 36.8 min | Design |
| Flotation cells | 6 x 70 m3 | Design |
| Regrind target | 80% passing 25 microns | Design |
| Regrind mill power | 900 kW tower mill | Design |
| AOX pulp density | 25-35% solids | Design |
| AOX minimum temperature | 75 deg C (167 deg F) | Testwork |
| Oxidation time | 24 h | Design |
| Sulphide-sulphur oxidation | ~100% | Design |
| Soda ash consumption | 120 kg/t conc | Design |
| Leach time | 48 h | Testwork |
| CIL pulp density | 40% solids | Design |
| Gold gravity recoverable | ~40% | Testwork |
| Gold recovered by gravity circuit | ~20% | Design |
| Live stockpile capacity | 10,000 t (2 days) | Design |
| Total stockpile capacity | 50,650 t | Design |
| SAG mill dimensions | 7.32 m dia x 2.74 m EGL | Equipment |
| SAG mill drive | 2,872 kW (3,850 hp) new | Design |
| Ball mill dimensions | 4.72 m dia x 6.55 m | Equipment |
| Ball mill drive | 2,313 kW (3,100 hp) fixed-speed | Equipment |
| Hydrocyclone classification | Not stated | Design |
Project website: https://www.heliostarmetals.com/projects/overview/
Technical qualifications
The report presents a proposed processing design based on available testwork and engineering assumptions. Values not stated in the source are not inferred.
