Stirred cyanide leaching

Stirred Cyanide Leaching

5.5.1 Principle Flowchart of Stirred Cyanide Leaching

Stirred cyanide leaching is generally suitable for gold-bearing materials with a grinding particle size of less than 0.3 mm. Its principle flowchart is shown in Figure 5-15. Compared with percolation cyanide leaching, stirred cyanide leaching has the advantages of smaller plant area, shorter leaching time, higher degree of mechanization, higher gold leaching rate, and stronger adaptability to raw materials.

Figure 5-15, Flowchart of stirred cyanide leaching

To improve gold recovery and shorten cyanide leaching time, amalgamation, gravity separation, or flotation is often used to recover coarse gold particles before cyanide leaching. Gravity separation or flotation is also commonly used to remove large amounts of gangue and harmful impurities that hinder cyanide leaching, obtaining gold concentrate. The gold concentrate is then regrinded before cyanide leaching. When the gold-bearing ore contains large amounts of clay, ochre, shale, and has a high content of fine gold particles, the flotation index is low. In such cases, the gold-bearing ore can be ground in a gold-removing solution (lean solution) and then subjected to whole-sludge cyanide leaching. Therefore, the gold-bearing material obtained through agitated cyanide leaching can be raw ore, amalgamation tailings, gravity separation tailings, flotation gold concentrate, gold-copper mixed concentrate separated from flotation, gold-bearing ferrous ore concentrate, and gold-bearing ferrous ore slag, etc.

Stirred Cyanide Leaching Tank

In stirred cyanide leaching for gold extraction, finely ground gold-containing material and cyanide leaching agent are continuously stirred and aerated in a stirred tank to complete the gold leaching. Based on the stirring principle and method, stirred leaching tanks can be divided into three types: mechanically stirred leaching tanks, air-stirred leaching tanks, and combined air-mechanical stirred leaching tanks.

5.5.2.1 Mechanically Stirred Leaching Tank

In a mechanically stirred leaching tank, the stirring of the slurry is accomplished by a high-speed rotating mechanical agitator. The mechanical agitator can be a propeller, impeller, or turbine type. The propeller-type stirred leaching tank shown in Figure 5-16 is a commonly used type of mechanically stirred leaching tank in production. It mainly consists of a tank body, a vertical shaft with a propeller, a central slurry receiving pipe 1, a circulation pipe, a cover plate 6, a slurry feed pipe 8, and a discharge pipe 9. As the propeller rotates rapidly, the slurry in the tank flows into the central slurry receiving pipe through the branch pipes, forming a vortex. Air is drawn into the vortex, bringing the oxygen content in the slurry to saturation. As the propeller rotates, it pushes the slurry entering the receiving pipe to the bottom of the tank, then returns from the bottom and rises along the tank wall, re-entering the central slurry receiving pipe through the circulation pipe, thus achieving multiple cycles of slurry circulation and continuously drawing air into the slurry.

Figure 5-16 Propeller-type stirred leaching tank

1—Pulp receiving pipe; 2—Branch pipe;

3—Vertical shaft; 4—Propeller; 5—Support; 6—Cover;

7—Flow channel; 8—Feed pipe; 9—Discharge pipe

Mechanically stirred leaching tanks provide uniform and intense agitation of the slurry, continuously drawing air into it to increase oxygen content. This facilitates restarting after shutdown, preventing the slurry from obstructing the propeller.

In production practice, several vertical compressed air pipes are sometimes inserted into the tank, or air lifters are installed on the inner (outer) wall of the tank to further enhance oxygen content and agitation intensity.

Besides propeller-type stirred leaching tanks, impeller-type and other mechanically stirred leaching tanks are also used in production.

5.5.2.2 Air-Stirred Leaching Tank

Air-stirred leaching tanks achieve uniform and intense agitation of the slurry through the pneumatic action of compressed air. The structural diagram of an air-stirred leaching tank is shown in Figure 5-17. Internationally, this type of tank is often referred to as a Pachuca or Brown air-stirred leaching tank. The upper part of the leaching tank is a tall cylinder, and the bottom is a 60-degree cone. It mainly consists of the tank body, central pipe, compressed air pipe, feed pipe, and discharge pipe. During operation, the slurry enters the tank through the feed pipe. The compressed air pipe leads directly to the lower part of the central tube. The compressed air rises in bubbles within the central tube, causing the density of the slurry inside the central tube to be lower than that in the annular space outside the central tube. This results in the slurry continuously rising inside the central tube and continuously falling in the annular space outside the central tube, achieving slurry circulation. Adjusting the compressed air pressure and flow rate regulates the agitation intensity of the slurry.

The air-stirred leaching tank can operate intermittently or continuously. In intermittent operation, the slurry is discharged through the lower discharge pipe at the end of leaching. In continuous operation, the slurry is discharged through the upper discharge pipe and enters the next leaching tank.

The compressed air-stirred leaching tank provides very strong agitation of the slurry, bringing the oxygen content in the slurry close to saturation.

Figure 5-17 Air-stirred leaching tank (tower)

1—Central pipe; 2—Feed pipe; 3—Compressed air pipe; 4—Lower discharge pipe; 5—Upper discharge pipe; 6—Tank body

A mixed-agitation leaching tank is a circular tank equipped with an air lift and mechanical rake in the center, or with air lifts around the perimeter and a slurry circulation pipe and a spiral agitator in the center. Therefore, it is also called a rake-type agitator leaching tank. Rake agitators are used in various types abroad, including the Dow, Denver, and Worman models. Figure 5-18 shows a mixed-agitation leaching tank widely used in gold processing plants. It mainly consists of a tank body, air lift pipe, rake, vertical shaft, flow channel, and transmission device. The slurry enters the tank through the feed inlet located at the top. Within the tank, it settles in layers towards the bottom. The concentrated slurry at the bottom, aided by the rotation of the rake (1-4 rpm), gathers towards the hollow lift pipe inlet. Under the action of compressed air, the concentrated slurry rises along the air lift pipe and overflows into two perforated channels. It then flows back into the tank through the perforations of the channels, which rotate with the vertical axis, ensuring a uniform distribution of the slurry within the tank. The leached slurry is continuously discharged from the outlet opposite the feed inlet, thus achieving continuous operation.

This type of stirred leaching tank has a low profile, no sediment at the bottom, a high gold leaching rate, and low cyanide consumption.

Figure 5-18 Rake-type mixer leaching tank

1—Air lift pipe; 2—Rake; 3—Flow channel; 4—Vertical shaft; 5—Horizontal frame; 6—Transmission device

Currently, my country mainly uses the SJ type double-impeller leaching mixing tank, which is a mechanical and air-mixed mixing leaching tank. The impeller speed is relatively low, low-pressure air is injected through the hollow shaft, resulting in uniform aeration, stable ore flow, good mixing effect, low power consumption, and rubber-lined impellers for long service life.

5.5.3 Mixed Leaching Operation Method

Mixed leaching can be divided into continuous stirred cyanide leaching and intermittent stirred cyanide leaching according to its operation method. In continuous stirred leaching, the slurry flows downstream through several stirred leaching tanks connected in series. When the slurry cannot flow by gravity, it can be pumped. Generally, the slurry should flow by gravity to minimize the number of pumping operations and reduce power consumption. In intermittent stirred leaching, the slurry is fed into several parallel stirred leaching tanks. At the end of leaching, the slurry is discharged into a storage tank, and another batch of slurry is fed into the stirred leaching tanks for leaching.

Continuous stirred cyanide leaching is more commonly used in gold concentrators. Intermittent stirred cyanide leaching is only used in some small concentrators, when processing certain refractory gold ores, or when a fresh cyanide solution is required for each leaching stage.

Continuous stirred cyanide leaching is usually carried out in 3 to 6 stirred leaching tanks connected in series.

Generally, the leaching tanks are installed in a stepped configuration, allowing the slurry to pass through each tank evenly and continuously. The time for the slurry to pass through each leaching tank should be equal to or slightly longer than the leaching time required under those leaching conditions. Compared to intermittent leaching, continuous stirred leaching (CSL) shortens loading and unloading time, has a larger equipment processing capacity, allows for continuous solid-liquid separation of the leached slurry, eliminates the need for storage tanks, reduces plant area and power consumption, and facilitates process automation and improved working conditions. This operating method can be used in most gold concentrators.

Intermittent stirred leaching is generally used in gold concentrators with a processing capacity of less than 100 tons/day. Large concentrators often require several or even a dozen leaching tanks operating in parallel. The leached slurry is sent to a storage tank for batch solid-liquid separation. The slurry in the storage tank must be continuously stirred to prevent mineral particle settling.

The liquid-to-solid ratio during stirred cyanide leaching is typically 1:1. When leaching and washing are carried out in the same leaching tank, wash water can be added at the end of leaching to dilute the slurry to a liquid-to-solid ratio of 3:1.

5.5.4 Solid-Liquid Separation and Washing of Leaching Slurry Stirred cyanide leaching slurry requires solid-liquid separation to obtain a clear, precious solution for gold precipitation. To improve gold recovery, the solid fraction after solid-liquid separation must be washed to recover as much gold-containing solution as possible. In production, decantation, filtration, and fluidized bed methods can be used for solid-liquid separation and washing of the leaching slurry.

5.5.4.1 Decantation Method Most gold concentrators in my country use decantation for solid-liquid separation and washing of the leaching slurry. Abroad, this method is mainly used in North America. Decantation is divided into intermittent decantation and continuous decantation.

A. Intermittent Decanting Method

Intermittent decanting is commonly used for solid-liquid separation of intermittent cyanide leaching pulp and can be carried out in a settling tank or thickener. When using a settling tank, after the leaching pulp is clarified, the clarified gold-bearing solution is discharged using a siphon with a float and sent to settle the gold. The remaining concentrated pulp is returned to the stirred leaching tank for washing with a dilute sodium cyanide solution. The pulp is then sent back to the settling tank for clarification. This process is repeated several times until the gold content in the wash liquid is trace. When using a thickener for solid-liquid separation, the leaching pulp is sent to the thickener for clarification. The overflow is sent to settle the gold, and the underflow is returned to the stirred leaching tank for washing with a dilute sodium cyanide solution. The pulp is then sent back to the thickener for clarification. This process is repeated several times until the gold content in the overflow is trace. The gold content in the washing solution is usually low, so a stepwise concentration method can be used for washing. The first washing solution, after adjusting the sodium cyanide concentration and pH, is used as the leaching agent for the next batch of gold-containing material. The second washing solution is used as the first washing agent for the next batch of raw material leaching slurry, and so on, increasing the concentration step by step. The final washing solution uses clean water as the washing agent.

Intermittent decanting has a long operation time, requires a large volume of solution, and occupies a large area. Currently, it is rarely used in industry, only in cyanide plants with small processing capacities.

B. Continuous Countercurrent Decanting Method

Industrially used continuous decanting methods are mostly countercurrent decanting methods (called CCD process), where the leaching slurry and washing solution move in opposite directions. The operation is carried out in several single-layer or multi-layer thickeners connected in series.

a. Single-layer Thickener Continuous Countercurrent Washing

Several single-layer thickeners connected in series can be used for solid-liquid separation and continuous countercurrent washing of the stirred leaching slurry. A typical process is shown in Figure 5-19. The single-layer thickener continuous countercurrent washing method has the advantages of simple operation, high washing rate of dissolved gold, and easy automation. However, the equipment occupies a large area and the slurry needs to be pumped multiple times.

Figure 5-19 Typical Flowchart of Continuous Countercurrent Tilting Washing

Therefore, many cyanide plants tend to adopt a continuous countercurrent washing process using multi-stage thickeners.

To improve concentration efficiency and reduce equipment footprint, in the late 1970s, the Elandsrand gold mine in South Africa and Silver King Mining and Houston International Minerals in Nevada, USA, successively adopted a new type of high-efficiency thickener. Elandsrand used it for thickening the overflow of the hydrocyclone, while Silver King and Houston used it for countercurrent decant washing of the leaching slurry. The slurry is first degassed and then mixed before entering the thickener (Figure 5-20). The flocculant is added to the mixer in at least three stages to mix with the slurry. The mixer is equipped with an impeller to ensure uniform distribution of the flocculant in the slurry. The slurry mixed with the flocculant flows into the lower tank of the mixer. The lower tank of the mixer is equipped with radial inclined plates to increase the concentration area. The concentrated underflow is raked out by a rake mechanism. The precipitated solution is filtered through the flocs in the compression layer and discharged through the upper overflow chute. The high-efficiency thickener is small in size and highly efficient, up to 10 times more efficient than conventional thickeners, requiring less investment. However, it is more sensitive to changes in ore particle size and properties.

Figure 5-20 Schematic cross-sectional view of a high-efficiency thickener

1—Rake arm drive device; 2—Mixer drive device; 3—Flocculant addition pipe;

4—Mixer; 5—Rake arm; 6—Feeding pipe; 7—Overflow chute; 8—Settled sand discharge pipe; 9—Degassing system

b. Multi-layer thickener continuous countercurrent washing

The structure of a multi-layer thickener is largely similar to that of a center-driven single-layer thickener, except that multiple (usually 2-5) single-layer thickeners are stacked together. A mud seal device (interlayer gate) is used between layers to prevent overflow (clarified liquid) from the lower layer from reaching the upper layer.

In my country, gold beneficiation plants often use two-layer or three-layer thickeners. Figure 5-21 shows the principle of continuous countercurrent washing for a three-layer thickener. The rake frames of each thickener are fixed on the same vertical shaft, which is driven by an electric motor to rotate. Interlayer gates are installed between layers, and the overflow pipes of each layer are connected to the washing liquid tank. As the rake slowly rotates, the upper underflow can smoothly drain to the next layer, but the clarified liquid from the lower layer cannot enter the upper layer. The suspensions between the layers are interconnected and can maintain relative stability through hydrostatic equilibrium. Therefore, the overflow port of the lower layer of clarified liquid in the washing tank must be higher than the liquid surface of the upper layer. Let the height difference be Δh, and its value can be calculated using the following formula:

Inter-floor gate (cross-section)

Figure 5-21 Schematic diagram of continuous countercurrent washing principle of three-layer thickener

1—Central shaft; 2—Rake frame; 3—Inlet; 4—Discharge port;

5—Washing liquid pipe; 6—Overflow pipe; 7—Washing liquid pipe; 8—Overflow pipe; 9—Washing liquid pipe; 10—Overflow tank; 11—Washing liquid tank

In the formula, △h—the height of the lower overflow outlet above the upper liquid surface;

h—the height of the upper underflow;

p—the density of the concentrated underflow;

Po—the density of the clarified liquid.

During operation, the leaching slurry enters the uppermost layer of the multi-layer thickener through the feed pipe. The underflow passes through each layer sequentially and is finally discharged through the bottom outlet. The washing liquid (usually a gold-removing solution) enters the bottom thickener through the washing liquid pipe to wash the underflow discharged from the upper layer. Its overflow enters compartment II of the washing liquid tank along the overflow pipe, and then enters the second thickener through pipe 7 to wash the underflow discharged from the first thickener. The overflow of the second thickener enters compartment I of the washing liquid tank through the overflow pipe; then enters the first thickener through washing liquid pipe 9 to wash the slurry. The overflow discharged from the first thickener is the gold-bearing precious solution. Figure 5-22 shows the solution equilibrium diagram of continuous countercurrent washing using a three-layer thickener in a cyanide plant in my country. The gold-containing material was ground to 88%-0.037 mm, the feed concentration was 27.8%, the discharge concentration was 57.72%, and the washing rate of dissolved gold was 98.86%.

Figure 5-22 Continuous countercurrent washing process of a three-layer thickener in a cyanide plant

Filtration Method

Solid-liquid separation and washing in filtration can be done continuously or intermittently. Continuous operation commonly uses cylindrical vacuum filters and disc vacuum filters, while intermittent operation often uses frame vacuum filters and filter presses. Continuous filtration is more common in gold concentrators.

Intermittent filtration is generally used for solid-liquid separation of difficult-to-filter cyanide slurries, allowing for longer washing of the filter cake. However, its processing capacity is lower and it requires a larger area, generally used in gold concentrators with smaller processing capacities.

The concentration of agitated cyanide leaching slurry is relatively low, typically around 30%. To improve the processing capacity and filtration efficiency of the filter, the cyanide leaching slurry is generally first concentrated and dewatered, with the concentrated underflow then sent to the filter. Flocculants can be added during concentration to achieve an underflow concentration of over 55% before filtration. The filtered cake must be washed multiple times to recover the mechanically entrained gold-containing solution. Generally, it is first washed with a dilute sodium cyanide solution, followed by washing with clean water. A two-stage filtration and washing method is usually used to wash the filter cake. Figure 5-23 shows the filtration and washing process of a cyanide plant using a concentrator and two filters.

Figure 5-23 Three-stage filtration and washing process of a cyanide plant

The washing rate of dissolved gold at this plant is 98.27%.

Some cyanide processing plants in South Africa have adopted 60m² and 120m² vacuum belt filters. The vacuum belt filter mainly consists of a frame, drive wheel, tail wheel, annular belt, padding belt, vacuum chamber, and air box (Figure 5-24). The annular belt has many transverse drainage grooves, with the bottom of the grooves sloping from both sides to the middle. There are longitudinal through holes in the middle of the belt, with one hole in each drainage groove for collecting and discharging the filtrate. The filtrate flows into the vacuum chamber below the belt. Filter cloth is laid on the belt, which is driven by a transmission device along the drive wheel and tail wheel. The upper working part of the belt moves along the air cushion, only pressing against the stationary vacuum chamber within the vacuum range, dragging across the vacuum chamber. To avoid wear, a narrow, polished padding belt is placed under the belt. The padding belt moves together with the annular belt and can be quickly replaced when worn. The slurry is fed from above by a distributor onto the filter cloth tightly attached to the belt. Vacuum filtration forces the solution along the belt’s drainage grooves and through the drainage holes into a vacuum chamber, and then into a storage tank. The belt filter can be divided into three zones: a suction filtration zone, a washing zone, and a drying zone. After passing through the drying zone and the drive wheel, the filter cloth separates from the conveyor belt. The filter cake is unloaded by the discharge roller. The filter cloth belt passes through a jet washer, tension wheel, and automatic spacing adjustment system, aligning with the conveyor belt at the tail pulley for continuous automated operation. A turbidimeter is installed to measure the turbidity of the filtrate, allowing for real-time monitoring of the filter cloth’s working condition. Belt filters can perform multi-stage filtration and washing, do not require filter cake slurrying, have high processing capacity and efficiency, and allow for easy filter cloth replacement. However, they require significant initial investment, have high maintenance costs, and are relatively complex to operate.

Figure 5-24 Schematic diagram of belt filter structure

1—Slurry cylinder and slurry distributor; 2—Wash water distributor; 3—Drive wheel;

4—Continuous belt; 5—Continuous filter cloth; 6—Tail wheel

Intermittent filtration generally uses plate and frame filter presses, but automatic plate and frame filter presses can also be used for solid-liquid separation and washing of cyanide slurry.

5.5.4.3 Fluidized Bed Method

Fluidized bed solid-liquid separation and washing are often carried out in a fluidized bed washing column (tower). The fluidized bed washing column (tower) is a tall, slender hollow cylinder (Figure 5-25). It is mainly used to remove mineral sands from the leaching slurry and to wash the mineral sands. It consists of three parts: an expansion chamber, a column body, and a conical bottom. A feed cylinder is located in the center of the expansion chamber, allowing the leaching slurry to enter the expansion chamber smoothly and evenly. The washing liquid is fed in from the interface between the washing section and the compression section, and is evenly distributed across the column cross-section by a distribution device. In the washing section, the ore and washing liquid move countercurrently. The gold-bearing solution and fine mineral particles in the slurry are discharged from the upper overflow weir along with the wash water, and then filtered to obtain a clear gold-bearing solution. The ore settles downwards through the expansion chamber and undergoes countercurrent washing in the washing section, forming a fluidized bed with a thinner upper layer and a thicker lower layer. The washed ore settles into the compression section. In the compression section, the ore is compressed and concentrated, descends in a moving bed state, and finally discharged from the bottom of the column.

Figure 5-25 Schematic diagram of the principle and structure of the washing column

Clarification of Gold-Bearing Solutions

Gold-bearing solutions must be clarified before entering the displacement gold precipitation process to improve the efficiency of the process and reduce reagent consumption.

The gold-bearing solution obtained from the leaching slurry after solid-liquid separation and washing contains a small amount of sludge and difficult-to-settle suspended solids. These impurities, when introduced into the displacement gold precipitation process, will contaminate the zinc surface, reduce the gold precipitation rate, and consume cyanide in the solution. Currently, frame clarifiers are widely used for clarifying gold-bearing solutions, followed by filter presses. Small mines may use sand filter boxes and sedimentation tanks.

A sand filter box consists of a filter medium (filter cloth, canvas, or burlap sacks) laid on a false bottom. A 120-150 mm thick gravel layer and a 60 mm thick fine sand layer are then placed on top of the filter medium. Generally, two sand filter boxes are used for periodic replacement. The fine sand should be replaced when cleaning the sand filter box. Like sedimentation tanks, sand filters have low production efficiency and poor clarification effect, but their simple structure often allows them to be used in conjunction with frame clarifiers.

Filter cloths used in gold-containing solution clarification are frequently clogged by carbonates, sulfides, or mineral sludge. To eliminate these harmful effects, an intermediate storage tank is usually not installed between filtration and clarification to shorten the contact time between the gold-containing solution and air, reducing the amount of carbon dioxide dissolved in the solution. Clarification equipment should be cleaned regularly, and the filter cloth should be washed with 1%–1.5% hydrochloric acid to remove calcium carbonate precipitates.

Figure 5-26 Flat-bottomed air mixing tank

1—Splash guard; 2—Gate valve; 3—Central air lift pipe;

4—Stainless steel support pipe; 5—Conical air inlet pipe; 6—Concrete base; 7—Observation hole; 8—Feed chute; 9—Discharge chute; 10—Overflow pipe; 11—Drain pipe

Example of Agitation Cyanide Gold Extraction Application: The old cyanide plant at the Sint Helena gold mine, which began operation in 1951, was rebuilt due to increased ore volume, severe plant corrosion, outdated equipment, and declining gold recovery. The new plant’s design incorporated grinding practices from the old plant since 1956 and the Leslie gold mine since 1963, as well as experiences from the Kinross concentrator since 1967 in reducing production costs and saving labor. The new plant was completed and put into operation in September 1976, with a monthly processing capacity of 300,000 tons of ore. The new plant eliminates waste rock sorting and crushing operations; the ore is brittle, and ore from the mine is directly fed into the autogenous mill or fed with 100mm steel balls. A closed-loop system consisting of an autogenous mill and a Krebs hydrocyclone grinder is used to grind ore to -200 mesh. The slurry is then fed into ten flat-bottomed air-stirred tanks (Figure 5-26) for cyanidation. The slurry is circulated via an air lifter in the center of the tank, surrounded by six high-pressure air nozzles to agitate the slurry and prevent sedimentation. The leached slurry is filtered using a vacuum filter. The precious liquor is clarified using an upflow sand filter box and deoxidized using a Klaus degassing tower. The gold mud obtained from zinc powder displacement leaching is dewatered using a horizontal belt filter and then roasted in a tunnel kiln. After roasting and drying, the gold mud is smelted in a large electric arc furnace to shorten the smelting time. Dust is collected using a baghouse dust collector.

The New Occidental gold mine produces fine-grained gold-bearing silicified slate and sandstone, containing 8.7 g/t of gold, approximately 1% chalcopyrite, and 1-2% pyrrhotite. To reduce cyanide consumption and improve gold recovery, lead oxide is added during cyanide leaching in addition to careful control of the leaching alkalinity. The processing flow is shown in Figure 5-27, with a processing capacity of 8000 tons/month. The ore is crushed to 9.5 mm, and then ground to 90%-200 mesh using lime, cyanide, and lead oxide in a two-stage closed-circuit grinding process. The overflow from the leaching slurry is concentrated and then sent to four Deverow stirred tanks connected in series for cyanidation for 30 hours. The overflow from the concentrated leaching slurry and the washing liquid from the underflow are returned to the grinding operation. The overflow during leaching contains 0.06% NaCN and 0.007% CaO. The reagent consumption was: NaCN 1.14 kg/ton, CaO 1.4 kg/ton, PbO 300 g/ton. The overflow (gold-containing solution) from the concentrated ore slurry was clarified, deoxidized, and subjected to zinc powder displacement to produce gold-containing precipitate. After roasting, an oxidant was added for smelting, producing high-copper alloy gold ingots. These ingots were then smelted with sulfur to remove some copper, yielding an alloy gold ingot containing 86% gold, 4% silver, and 10% base metals (mainly copper). The gold recovery rate was 91%. Although the plant has ceased production, its production experience has certain practical significance.

Figure 5-27 Flowchart of the New Oxytontal Gold Mine

After 30 hours, the overflow from the leaching slurry concentration and the wash liquid from the underflow are returned to the grinding operation. The overflow during leaching contains 0.06% NaCN and 0.007% CaO. The reagent consumption is: NaCN 1.14 kg/ton, CaO 1.4 kg/ton, PbO 300 g/ton. The overflow (gold-containing solution) from the grinding slurry concentration is clarified, deoxidized, and subjected to zinc powder replacement to produce gold-containing precipitate. After roasting, an oxidant is added for smelting to produce high-copper alloy gold ingots. These ingots are then smelted with sulfur to remove some copper, producing alloy gold ingots containing 86% gold, 4% silver, and 10% base metals (mainly copper). The gold recovery rate is 91%. Although the plant has ceased production, its production experience has certain practical significance. 5.5.6.2 Flotation and Cyanide Gold Extraction

A gold concentrator in my country processes pyrite-bearing quartz ore. The main metallic minerals are pyrite, pyrrhotite, sphalerite, galena, chalcopyrite, magnetite, argentite, and native gold. Gangue minerals are mainly quartz, sericite, plagioclase, dolomite, amphibole, and kaolin. Native gold is distributed in rounded, elongated, and irregular shapes within chalcopyrite, pyrite, and quartz. The raw ore contains approximately 10 g/t of gold and 0.1% copper.

The plant employs a combined flotation and cyanide gold extraction process. After mixed flotation and separate flotation, the ore yields gold-copper concentrate and gold-sulfur concentrate. The gold-copper concentrate contains 500–1000 g/t of gold and 4% of copper, with a gold recovery rate of approximately 50% and a copper recovery rate of approximately 50%. The gold-bearing sulfur concentrate contains 80-100 g/t of gold, less than 0.1% copper, and 35-40% sulfur, with a gold recovery rate of approximately 40%. The gold-copper concentrate is sent to a smelter for further processing, while the gold-bearing sulfur concentrate is used for on-site gold production via stirred cyanidation, as shown in Figure 5-28. The gold-bearing sulfur concentrate is regrinded to 98%-0.074 mm and has a pH of 11. After concentration (with the addition of coagulant No. 3), xanthate, No. 2 oil, and soluble salts are removed, resulting in an underflow concentration of 25%-30%. The underflow is then sent to five tandemly connected mechanically stirred tanks for cyanidation leaching. Several compressed air pipes are also inserted into the tanks for aeration. The sodium cyanide concentration during leaching is 0.1%-0.12%, the pH is 10, and after 24 hours of leaching, the gold leaching rate is 94.2%. The leached slurry is sent to a three-layer thickener for solid-liquid separation and continuous countercurrent washing. The cyanide tailings contain approximately 2 grams of gold per ton and 35%–40% sulfur. After natural drying in a tailings pond, it is sold as sulfur concentrate. The overflow (precious solution) from the three-layer thickener contains more than 10 grams of gold per cubic meter. After clarification in a sand filter, it is sent to a zinc wire displacement bath for gold precipitation. Lead acetate is added during the displacement process, achieving a gold precipitation rate of 99.5%. The gold-free solution (lean solution) contains 0.1–0.05 grams of gold per cubic meter. A portion of this solution is returned to the bottom layer of the three-layer thickener for washing, while the remainder is treated with bleaching powder and discarded.

Figure 5-28 Gold-bearing sulfur concentrate from a beneficiation plant in China

Stirred cyanidation process flow

The gold mud obtained from gold precipitation contains 2%~5% gold. After washing with sulfuric acid, drying and calcining, nitrates and borax are added as flux and smelted in a converter to produce a gold alloy containing 40%~50% gold. The gold smelting recovery rate reaches 98%, and the total gold recovery rate is 85%. The material consumption (kg/ton) for processing gold-bearing sulfur concentrate at this plant is: sodium cyanide 7.22, lead acetate 0.36, zinc wire 5.17, bleaching powder 6.95. Due to the low copper content of the raw ore, the plant has now eliminated flotation separation. The mixed flotation concentrate is regrinded and then directly cyanidated to extract gold, using a two-leaching and two-washing process. The cyanidation leaching slurry from the second stage is sent to a three-layer thickener for solid-liquid separation and countercurrent washing. The overflow is returned to the first stage of cyanidation leaching, and the underflow is vacuum filtered to obtain sulfur concentrate for sale. The first stage of cyanide slurry undergoes solid-liquid separation and countercurrent washing in a three-stage thickener. The resulting precious liquor, after clarification in a frame clarifier, is sent to a zinc powder displacement gold precipitation process. The resulting gold mud is dried and sent to a converter for smelting. The lean liquor obtained from displacement and the filtrate from the underflow vacuum filtration of the second-stage three-stage thickener, except for a portion returned for use as washing liquid, are regenerated using sulfuric acid acidification to recover cyanide. After absorption with sodium hydroxide solution, an alkaline sodium cyanide solution containing 20.35% sodium cyanide and 0.67% sodium hydroxide is obtained, which is returned to the cyanide leaching process, saving a significant amount of sodium cyanide.

Another gold beneficiation plant in my country processes quartz ore containing gold and pyrite. The main metallic minerals are pyrite, magnetite, limonite, molybdenite, and native gold, while the gangue minerals are mainly quartz, carbonates, plagioclase, and sericite. The raw ore contains 6 grams of gold per ton, with fine gold grains closely associated with pyrite. The plant employs a combined process of flotation and stirred cyanidation for gold extraction (Figure 5-29). The ore is flotated to obtain a gold-bearing sulfur concentrate, containing 117.5 g/t of gold, 43.17 g/t of silver, 0.18% copper, 0.05% molybdenum, 25.26% sulfur, and 1.54% carbon, with a gold recovery rate of 96%. After concentration, filtration, and reagent removal, the gold-bearing sulfur concentrate is continuously regrinded in two stages to a particle size of 98%-0.044 mm, and then fed into four mechanically stirred leaching tanks equipped with aeration devices for cyanidation leaching. Lime is used as a protective alkali and added to the first-stage regrinding mill. The leaching pH is 10, the sodium cyanide concentration is 0.1%, the leaching time is 36 hours, the pulp concentration is 25%–30%, and the gold leaching rate reaches 94%–96%. The leaching slurry undergoes solid-liquid separation and continuous countercurrent washing using a double-layer thickener and a filter, achieving a gold washing rate of over 98%. The gold-containing solution, with a gold content of 20 g/m³, is clarified in a sand filter with the addition of lead acetate, and then sent for zinc wire displacement precipitation. The displacement precipitation time is 90 minutes, achieving a gold displacement precipitation rate of over 99%. The gold mud is smelted and cast into ingots to obtain alloyed gold. The gold-removing solution contains approximately 0.06 g/m³ of gold and is generally discarded after treatment with bleaching powder or chlorine gas. The filter cake (cyanide tailings) after filtration and washing is sent to a molybdenum flotation system, where 8000 g/ton of water glass, 80 g/ton of kerosene, and 60 g/ton of No. 2 oil are added to produce molybdenum middlings containing 13% molybdenum. After oxidative roasting, the molybdenum middlings undergoes chemical treatment to produce ammonium molybdate and tailings. The tailings contained approximately 300 grams of gold per ton, mainly due to the loss of gold due to adsorption of some of the carbonaceous matter in the gold-bearing sulfur concentrate. The tailings were returned along with the gold-bearing sulfur concentrate for cyanide leaching to recover the gold. The molybdenum flotation tailings contained approximately 25%–30% sulfur and were sold to a sulfuric acid plant. The material consumption (kg/ton) for processing the gold-bearing sulfur concentrate was: sodium cyanide 8; lime 12; lead acetate 3; zinc wire 2–3. Later, the plant underwent technical upgrades; the zinc wire replacement gold precipitation was replaced with zinc powder replacement gold precipitation. After clarification, the gold-containing solution was sent to a deaerator for deoxidation, followed by zinc powder replacement gold precipitation. This increased the gold content in the gold mud, reduced zinc and flux consumption, and improved the overall gold recovery rate.

Figure 5-29 Flowchart of a gold beneficiation plant in my country for gold extraction by stirred cyanide treatment