A technical report dated March 16, 2018, presents conceptual and historical designs for a 15,000 tpd oxide cyanidation plant and a 140,000 tpd flotation concentrator, with testwork and transition schedules for processing saprolite and hard rock materials.
Report context
This technical report for the Siembra Minera Project (Project #2832), prepared for Gold Reserve Inc., is dated March 16, 2018. The report documents a Preliminary Economic Assessment (PEA) that includes conceptual plant designs completed by Samuel Engineering for the oxide cyanidation plant, and historical feasibility study and detailed design work completed by Aker Kvaerner (2005) and SNC-Lavalin (2006–2007) for the flotation concentrator. The report presents processing routes for oxide saprolite, sulphide saprolite, and hard rock materials over a planned production schedule.
Processing route
Oxide cyanidation plant (conceptual design)
The conceptual 15,000 tpd oxide cyanidation plant design was completed by Samuel Engineering. Saprolite is excavated, loaded onto trucks, and transported to a stockpile adjacent to a saprolite crushing plant located near the mine. A front-end loader feeds material into the crusher feed bin, which is equipped with a static grizzly to remove large debris or rocks that could cause problems in the double roll crusher. Oversize material is rejected and stockpiled; if gold grade is low it is periodically moved to a waste dump, or to a separate stockpile if it may be economically processed in the future.
Mineralized material from the feed bin is fed to the crusher using an apron feeder. Crusher discharge is transferred by conveyor to a vibrating screen where debris that could plug a pump is removed. Screen undersize discharges into a mix tank where it is mixed with water to prepare a slurry that is pumped through a high density polyethylene (HDPE) pipeline to a surge tank located in the grinding circuit at the plant site, approximately five to six kilometres away.
At the oxide plant, slurry is pumped to a ball mill for nominal size reduction. Ball mill discharge is pumped to a cyclone cluster; material with a particle size P80 of 50 μm reports to the cyclone overflow and discharges into a pre-leach thickener. Coarser material reports to the cyclone underflow and is returned by gravity to the ball mill feed chute for further size reduction.
A fraction of the cyclone underflow is fed to the gravity concentration circuit, which uses batch centrifugal concentrators. The concentrators alternate operation to simulate continuous processing. Tailings from the gravity concentrators are returned to the grinding circuit. Concentrate from the centrifugal concentrators is fed to a storage tank in the gold room, then transferred to a feed tank and the primary gravity concentration table to further separate gold from gangue material. Concentrate from the primary gravity table is collected and stored before being fed to the secondary gravity table for final cleaning and upgrading. Secondary gravity table concentrate is collected and stored in a decant tank where free water is removed. Tailings from the gravity tables are returned to the grinding circuit.
The pre-leach thickener separates solids from liquid to produce optimum slurry density for the leach circuit. Overflow is collected and pumped to the process water tank. Underflow is pumped to the cyanide leach circuit, which consists of six tanks operated in series providing 18 hours retention time. Lime slurry is added to maintain pH between 10.0 and 11.0. Slurry discharging from the leach circuit is fed to a carbon-in-pulp (CIP) circuit consisting of six tanks operated in series providing eight hours retention time. Activated carbon is advanced counter-currently to the slurry flow to recover dissolved gold by adsorption. Slurry handling and carbon advancement uses Kemix pump cell technology.
Loaded carbon is treated using the Anglo-American Research Laboratory (AARL) method. The design includes an industry standard cold strip method to remove copper from activated carbon; the report notes that optimization work completed by Crystallex indicates copper adsorption can be mitigated by management of cyanide in the leach circuit, including maintaining proper cyanide concentrations and staged additions. Loaded carbon is transferred from the CIP circuit to the acid wash vessel where it is soaked in three percent hydrochloric acid solution, then rinsed and transferred to the elution column where it may undergo a cold strip with caustic-cyanide solution to remove copper if needed. The carbon is then soaked in a circulating hot solution of two percent sodium hydroxide and two to three percent sodium cyanide to remove gold and silver. After elution, carbon is rinsed with hot fresh water.
Pregnant solution is pumped to a single electrowinning cell equipped with stainless steel cathodes. After the electrowinning cycle, cathodes are washed to remove metal-bearing sludge. Sludge is filtered and dried. Dewatered sludge and gravity concentrates are transferred to pans placed in a mercury retort to remove mercury and dry materials prior to smelting. Dried concentrate is mixed with flux and transferred to an induction furnace to produce doré.
Stripped carbon is transferred to a dewatering screen and fed to a carbon regeneration kiln. Particulates from the kiln are collected in a wet scrubber. Carbon discharges from the kiln to a carbon quench tank, then is pumped to a carbon sizing screen. A single tailings thickener increases slurry solids density and recovers solution for recycling. Thickened underflow is processed in a sulphur dioxide (SO2)–air cyanide detoxification circuit before being discharged by gravity to the tailings management facility (TMF) via an 11 km HDPE pipeline.
Flotation concentrator (historical design with minor modifications)
Aker Kvaerner completed a plant design to support a feasibility study in 2005. SNC-Lavalin completed minor modifications during detailed design in 2006 and 2007, including: changing the leach circuit from CIP to CIL, increasing the size of the SAG mills from 10.97 m diameter to 11.58 m diameter, and adding intensive cyanide leach to process gravity recovered gold.
Hard rock is crushed in two gyratory crushers (1,473 mm by 1,905 mm) operating in parallel near the open pit mine. Crusher discharge falls into hoppers; variable speed apron feeders transfer crushed material to primary crusher discharge conveyors, which transfer to an overland stockpile feed conveyor. The stockpile feed conveyor transports crushed hard rock and sulphide saprolite containing greater than 0.07% Cu to the crushed material stockpile,an elongated stockpile fed by a tripper conveyor. Apron feeders reclaim crushed material and transfer it to the semi-autogenous grinding (SAG) mill feed conveyors.
The numbers of equipment pieces from the SNC-Lavalin basic engineering design were doubled in most cases, with an optimized plant layout for stockpiles and feeders developed by Samuel Engineering. Four parallel grinding lines are included, each consisting of one SAG mill and two ball mills. Water is added to the SAG mill feed hoppers to create slurry density of approximately 70% solids by weight. Slurry discharges from SAG mills through trommel screens; oversize is directed to a pebble collection conveyor. Undersize discharges to cyclone feed sumps, each receiving discharge from one SAG mill and two ball mills. Slurry is pumped to hydrocyclones; cyclone overflow is the grinding circuit product designed to produce a particle size P80 of 100 μm. Cyclone underflow discharges to ball mill feed chutes, with a portion fed to the gravity gold recovery circuit.
Each of the four grinding lines includes a gravity gold recovery circuit with two gravity scalping screens and two centrifugal concentrators (batch concentrators shut down every four hours to flush concentrate). Tailings from gravity recovery are returned to ball mill feed chutes. Concentrate from centrifugal concentrators is processed in an intensive cyanide leach reactor.
Cyclone overflow flows by gravity into rougher flotation conditioning tanks. Four rougher flotation lines, each with two banks of flotation cells providing 960 m³ of capacity for 20 minutes retention time, operate at neutral pH. Rougher tailings are collected and pumped to a tailings thickener feed tank. Rougher concentrate is pumped to regrind cyclone pump boxes, then to regrind cyclones for classification. Cyclone overflow has a particle size P80 of 37 μm; underflow flows to vertical regrind mills. Four parallel regrind circuits are provided.
Four stages of cleaner flotation improve concentrate grade, operated at pH between 11.5 and 12.0 to reject pyrite. First cleaner flotation includes two parallel circuits providing 390 m³ capacity for 10 minutes retention time. First cleaner concentrate progresses through second and third cleaner flotation circuits. First cleaner tailings flow by gravity to the cleaner scavenger circuit (two parallel circuits providing 260 m³ capacity). Cleaner scavenger concentrate is returned to the regrind cyclone pump box; cleaner scavenger tailings are pumped to the cyanide leach circuit.
Second stage cleaner flotation consists of two parallel circuits providing approximately 102 m³ capacity. Third stage cleaner flotation consists of two parallel circuits providing approximately 51 m³ capacity. Fourth stage cleaner flotation uses four pairs of column flotation cells operating in parallel. Tailings from each cleaner stage are returned to the feed of the previous cleaner stage.
Final flotation concentrate is pumped to two 9 m diameter concentrate thickeners operating in parallel. Overflow is pumped to the process water pond. Underflow is transferred to concentrate holding tanks, then filtered in automated horizontal plate filter presses to produce a target moisture concentration of 8% solids by weight. Final concentrate is stockpiled and trucked to a port facility for overseas transport.
First cleaner scavenger tailings are leached in a CIL circuit. A trash screen removes trash; underflow flows into one of two 30 m diameter pre-leach thickeners operating in parallel. Two parallel CIL circuits, each with six agitated tanks (13.6 m diameter by 14.0 m high), are provided. Lime slurry increases pH to between 10.0 and 10.5; sodium cyanide is added at the circuit start. New or re-activated carbon is added to the sixth tanks and advanced counter-currently. Loaded carbon is removed from tank one and sent to the adsorption desorption recovery (ADR) circuit.
Slurry from the CIL circuit flows to the cyanide destruction circuit where the SO2–air process reduces WAD cyanide concentration to less than 0.6 mg/L. Discharge is pumped to the TMF.
Two parallel ADR circuits are provided. Loaded carbon flows across a loaded carbon screen to a surge bin, then to an acid wash tank where hydrochloric acid removes inorganic contaminants. Carbon is washed with fresh water and neutralized with dilute sodium hydroxide, then transferred to elution columns with 6,500 kg capacity. The AARL elution process is used: carbon is pre-soaked in hot solution containing cyanide and sodium hydroxide, then elution continues until four bed volumes of solution are collected. Pregnant solution is pumped to two electrowinning cells with stainless steel mesh cathodes operating in series. Sludge is removed, filtered, dried, mixed with fluxes, and smelted to produce doré. After elution, carbon is reactivated in a regeneration kiln, quenched, transferred to carbon attrition tanks, and sized before return to the CIL circuits.
Rougher flotation tailings are dewatered in tailings thickeners to produce underflow slurry density of 55% solids by weight, then pumped to the TMF. Thickener overflow reports to the process water pond.
The plant design includes all reagent handling facilities, utilities, and auxiliary facilities. The report notes that trade-off studies were completed at the time of the feasibility study and detailed design, but conditions and assumptions made over twenty years ago were much different than current conditions; the report recommends re-evaluation of some studies using current metal prices, equipment sizes, and costs.
Plant transitions and reconfiguration
The production schedule initially processes oxide saprolite through the 15,000 tpd cyanide leach plant. The crushing and screening plant feed is approximately 10% higher, assuming some material rejection due to oversize or rock material. Starting in year 7, the majority of oxide saprolite is depleted and sulphide saprolite containing low copper concentrations is also fed to the plant. In years 9 and 10, only low copper sulphide saprolite is fed.
In year 4, the flotation concentrator is commissioned. Feed includes sulphide saprolite with higher copper concentration and a combination of high and low copper hard rock material at a nominal rate of 140,000 tpd, though actual feed rate is somewhat higher due to the presence of sulphide saprolite.
In year 11, hard rock with suitable copper grades for acceptable flotation concentrates diminishes. The plant is re-configured to process less material through the flotation plant and additional material through the oxide leach plant: feed to the flotation concentrator is reduced to approximately 105,000 tpd, and tonnage to the oxide leach plant is increased to 35,000 tpd. Low copper hard rock material is ground in the existing flotation plant milling circuit, and the leach plant is expanded to accommodate higher tonnage. The ball mill in the oxide leach plant, sized only to process saprolite, can be decommissioned or used to grind saprolite pumped from the open pit mine to the oxide leach plant.
Key reported parameters
| Parameter | Unit | Design/Historical Basis |
|---|---|---|
| Oxide cyanidation plant feed rate | tpd | 15,000 (conceptual design) |
| Flotation concentrator feed rate | tpd | 140,000 (historical design; actual higher with sulphide saprolite) |
| Oxide plant grinding product size (P80) | μm | 50 (design) |
| Oxide leach retention time | hours | 18 (design, six tanks) |
| Oxide CIP retention time | hours | 8 (design, six tanks) |
| Oxide leach pH range | – | 10.0 to 11.0 (design) |
| Flotation grinding product size (P80) | μm | 100 (design) |
| Regrind product size (P80) | μm | 37 (design) |
| Rougher flotation retention time | minutes | 20 (design, 960 m³ capacity) |
| Rougher flotation pH | – | Neutral (design) |
| Cleaner flotation pH range | – | 11.5 to 12.0 (design) |
| First cleaner retention time | minutes | 10 (design, 390 m³ capacity) |
| First cleaner scavenger capacity | m³ | 260 (design) |
| Second cleaner capacity | m³ | 102 (design) |
| Third cleaner capacity | m³ | 51 (design) |
| Concentrate thickener diameter | m | 9 (design) |
| Concentrate target moisture | % solids by weight | 8 (design) |
| Pre-leach thickener diameter | m | 30 (design) |
| CIL tank dimensions | m | 13.6 diameter × 14.0 high (design) |
| Elution column capacity | kg carbon | 6,500 (design) |
| WAD cyanide limit (destruction circuit) | mg/L | <0.6 (design) |
| Tailings thickener underflow density | % solids by weight | 55 (design) |
| Slurry density at SAG mill feed | % solids by weight | Approximately 70 (design) |
| Oxide plant pipeline distance | km | 5 to 6 (design) |
| TMF pipeline length (oxide plant) | km | 11 (design) |
| SAG mill diameter (original design) | m | 10.97 (historical) |
| SAG mill diameter (modified design) | m | 11.58 (SNC-Lavalin modification) |
| Primary gyratory crusher size | mm | 1,473 × 1,905 (historical design) |
| Year 4 reconfiguration flotation feed rate | tpd | 140,000 nominal (plan) |
| Year 11 reconfiguration flotation feed rate | tpd | Approximately 105,000 (plan) |
| Year 11 reconfiguration oxide leach feed rate | tpd | 35,000 (plan) |
| Saprolite crushing plant feed allowance | % | Approximately 10% higher (design for oversize rejection) |
Project website: https://goldreserve.bm/mixed-company
Technical qualifications
The oxide cyanidation plant design is conceptual, completed by Samuel Engineering to support a PEA. The flotation concentrator design is based on a feasibility study completed by Aker Kvaerner (2005) and minor modifications during detailed design by SNC-Lavalin (2006–2007). The report notes that trade-off studies completed over twenty years prior were based on conditions and assumptions much different than current conditions, and recommends re-evaluation using current metal prices, equipment sizes, and costs. The report expresses an opinion that copper adsorption on carbon can be mitigated by cyanide management, based on optimization work completed by Crystallex. No actual operating data or testwork results for recovery or metallurgical performance are presented in the processing sections.
*Source: Gold Reserve Inc. – Siembra Minera Project, Project #2832, Technical Report NI 43-101 – March 16, 2018, Section 17 – Recovery Methods.*
