EA041170B1 - WATER BALANCE IN CHLORIDE HEAP LEACHING - Google Patents

Description

State of the art

The present invention relates to a process for leaching copper ores using a high concentration of chloride.

Bioleaching of low-grade copper sulfide ores using acid-sulfate solutions with bacteria is a traditional industrial method used throughout the world.

More recently, chloride leaching methods have been developed to allow leaching copper sulfide ores, including chalcopyrite, to high expected productivity without the use of bacteria. WO2015/059551 and WO2014/030048 describe the use of high chloride concentrations to overcome chalcopyrite passivation, allowing operation at solution oxidation potentials greater than 700 mV, in contrast to the standard electrode system (SHE), resulting in increased leaching and higher leaching. degree of extraction of copper.

A high chloride concentration increases the rate of formation of oxidants such as copper ions (Cu(II)) and iron ions (Fe(III)), in the presence of an oxygen-containing gas. These oxidizers then react with the copper sulfide minerals, releasing the copper into solution.

In practice, heap leaching of copper ores in chloride solutions on an industrial scale is limited by requirements that, if ignored, limit the economic feasibility of the process. A particular limitation in heap leaching using high chloride concentrations (100-230 g/l chloride) is the loss of chloride from the loop to the solution contained in the leach residue in the form of moisture.

The aim of the present invention is to at least partially solve the aforementioned problem, i.e. maximizing the amount of water available and suitable for washing the leached solids to ensure chloride recovery and reduce chloride losses in said process.

The essence of the invention

The present invention provides a method for maximizing the amount of water available for washing (or washing) the leached ore residue in a high chloride heap leach, comprising the step of using process make-up water to wash the leached ore in a dump to displace a chloride-containing aqueous solution from the leached ore.

The water balance refers to the difference between the volume of water supplied to the system and the volume of water leaving the system.

Process make-up water, which contains little or no chloride, is used to displace the chloride-containing aqueous solution, which is then recovered and recycled to the heap leach cycle, thereby reducing the amount of make-up salt added to the process.

Soluble copper in the leached ore or a chloride-containing aqueous solution can be released into the wash water, increasing the overall copper recovery.

To maximize the amount of process make-up water available for flushing and to prevent accidental washout and associated loss of chloride and copper, methods for increasing the amount of process make-up water and using process make-up water can be implemented as follows:

a) Water scarcity can be created by increasing evaporation from the leaching process, for example by using aeronautical sprinklers or by avoiding the installation of evaporation reducing devices such as covers or reflective sheets over pools, i.e. leaving said devices open to the weather;

b) the ingress of rainwater into the dump can be prevented;

c) the volume of wash water in the solvent copper extraction loop can be reduced by applying a multi-stage countercurrent wash approach using loaded organic solvent; And

d) the consumption of make-up water to compensate for leakage during electrolytic separation can be reduced due to efficient washing of the loaded organic phase, so this is related to the above.

The chloride concentration in the organic phase can be reduced to less than 50 ppm.

The washing step can be carried out for 5 to 100 days. The preferred duration is 20 days.

Residual soluble copper that remains in the leached ore residue and is recovered in the washing step can be recycled to the leaching step and then can be recovered in the solvent extraction step and the electrowinning step.

The amount of process make-up water in the washing step may be from 0.05 to 0.35 m ³ /t ore and preferably from 0.1 to 0.15 m ³ /t ore.

- 1 041170

The amount of chloride and copper precipitated in the aqueous solution displaced in the process make-up water washing step can be expressed as a percentage of the total content in the residual ore before washing, and can be described as washing efficiency. Washing efficiency is typically 40 to 85% for the specified wash volume (i.e. 0.05 to

0.35 m ³ / t of ore). The required proportion (or ratio) is in the order of 60%.

The chloride content may be in the specified range and is usually 150 g/l; but this is only an approximate, but not limiting value.

Brief description of graphic materials

The present invention is further described by way of examples with reference to the accompanying drawings.

In FIG. 1 is a schematic representation of a typical heap leaching process.

In FIG. 2 is a black box diagram of the inputs and outputs of the heap leaching process shown in FIG. 1.

In FIG. 3 is a schematic representation of the solvent extraction process and the electrowinning process, which form part of the process depicted in FIG. 1.

In FIG. 4 shows an optimized heap leaching process according to the present invention.

In FIG. 5A is a wash efficiency curve that graphically illustrates the percentage recovery of soluble ions in the wash step as a function of the amount of wash water added per ton of ore.

In FIG. 5B graphically shows the reduction in required make-up salt addition in kg/L as a function of the amount of make-up water added per ton of ore.

In FIG. 6 is a graphical representation of an optimized copper solvent extraction during a laden organic solvent wash using a countercurrent loop that forms part of the process depicted in FIG. 4.

In FIG. 7A, 7B, and 7C, respectively, represent various countercurrent washing circuits that can be used in the washing process depicted in FIG. 6 and which are used to reduce the chloride concentration in the electrolyte entering the subsequent electrowinning process.

In FIG. 8 is a flow diagram of an integrated pilot plant used to process low grade chalcopyrite ore.

In FIG. 9 is a simplified diagram illustrating the inflows and outflows for the pilot plant mentioned in relation to FIG. 8.

In FIG. 10 shows copper recovery curves versus irrigation duration for several piles at various stages of pilot plant operation.

In FIG. Figure 11 shows a curve of rinsing efficiency (salt recovery) versus the amount of water added per barn in a pilot plant during the second phase of operation.

In FIG. 12 is a graphical representation of the amount of salt needed for make-up, as a function of the amount of rinse water added, for the barn mentioned in relation to FIG. eleven.

In FIG. 13 shows the accumulation of impurity metals (aluminum, magnesium and iron) and sulfates in the raffinate at the pilot plant, as a function of time.

Description of Preferred Embodiments

Among the accompanying graphics, FIG. 1 shows a general heap leaching process 10 that includes a single stage heap leaching. In general, this chloride heap leaching process is used using a dumping method known as a race track. In a race track dump, ore 12 is subjected to agglomeration (13) using acid 14 and salt 16 and deposited in dump 18 in a pool (not shown). The dump (stack) 18 is irrigated with an aqueous leaching solution 20 for a set period of time. The solution 20 is fed from above the dump 18. The copper released from the ore 12 passes into the leach solution 22, which drains from the dump 18 into the leach solution pool 24.

As a result of the combination of stage 26 extraction with an organic solvent and stage 28 electrolytic recovery of the extraction of metallic copper 29 from the solution.

At the end of the leaching period, irrigation is stopped and the leached ore is left to drain the solution. The leached ore residue 30 (also referred to as leach residue) and the entrained aqueous solution are removed and sent to a waste rock dump.

In FIG. Figure 2 shows various input and output streams for heap 18. Input streams include:

a) ore 12: a mined mineral containing a valuable mineral or metal, in this case copper;

b) acid 14: usually sulfuric acid added to ore 12 to maintain an acidic environment and to provide acid to interact with gangue minerals, and

- 2 041170 also for the dissolution of the copper mineral and the formation of Fe(III) and Cu(II) oxidizers, as well as for the release of soluble copper and other metal ions into the leaching solution 22;

c) rain or precipitation 32: i.e. rain falling on the heap 18 and in the respective pools, which subsequently combines with the solution circulating in said process;

d) Air 34: Air is fed into said process to provide oxygen as an oxidizing agent to form Cu(II) and Fe(III) and to leach the sulfide mineral;

e) process make-up water 36: this water is used to make up the deficiency if the amount of water in the outgoing stream is greater than the amount of water in the incoming stream; process make-up water is added to compensate for evaporative loss of water (42); loss of water in the form of moisture in the leached ore residue (30) and loss of water in the wash stream 45 coming from the leaching circuit, in the right amount and at the right time;

f) water 38 for washing the solvent used for copper extraction: this water is used during solvent extraction to remove chloride trapped in the organic phase to prevent the transfer of chloride to the electrolyte entering the next stage of electrowinning;

g) make-up water 40 to compensate for losses during electrolytic separation: this water is used to make up for losses of electrolyte withdrawn from the circuit in order to control the content of impurities in it; And

h) salt 16: for the introduction of chloride;

i) water 37: includes make-up water 36; wash water 38; and make-up water 40 for electrolytic separation.

The output streams shown in FIG. 2 are as follows:

a) leached ore (leached residue) 30, which contains moisture in the form of an entrained solution 31 in the leached ore residue 30 (ie, leached residue). Moisture is the residual aqueous solution remaining in the leached ore 30 after the irrigation with the leach solution 20 is stopped, the solution drains from the leached ore 30, and the productive leaching solution 22 is collected;

b) evaporation 42 from the heap and associated pools, i.e. loss of water to the atmosphere;

c) copper 29, i.e. target metal extracted from ore 12; And

d) flush stream 45, which is used as needed to remove contaminants from the loop or to reduce the acid concentration in the leach solution 20. Wash stream 45 may be supplied from any suitable stream or solution pool in the heap leach process 10 to remove impurities. Preferably, said stream should contain a low concentration of copper to prevent significant losses of copper in the heap leaching process 10 .

The air 34 that is fed into the heap 18 exits through the sides and over the top of the heap as air streams (not shown) that have a reduced oxygen content.

In FIG. 3 is a simplified process flow diagram of stage 26 of copper solvent extraction and stage 28 of the electrowinning process. The product leach solution 22X is fed from pool 24 through an organic solvent extraction unit 44, which includes an extraction process 46, then through a washing process 58, and then to a stripping process 48.

During the extraction process 46, dissolved copper ions are captured from the leach product 22X into the organic phase 51 and then the copper depleted solution 50 is fed into the raffinate pool 52 for recycling to the dump 18 during irrigation. During the washing process 58, the organic phase 51 is washed using wash water 38 with a low content of chloride, which is contained in a small amount in the aqueous solution entrained in the organic phase. After use, wash water 38X is recycled to pool 24 PLS.

During the stripping process 48, the dissolved copper is transferred from the organic phase 51 to the electrolyte 54, which is fed to the electrolytic separation stage 56 in the process 28, where the dissolved copper ions are reduced and then removed in the form of metallic copper 29. The spent electrolyte 55 is returned to steaming process 48. The electrowinning wash solution 35 may be recycled from the electrowinning step 56 to the heap leaching process 10, preferably to the PLS pool 24.

The only water leaving the process 10 shown in FIG. 2 is moisture 31 in leached ore 30, vapor phase 42, and wash stream 45, with liquor removed from the loop to limit buildup of impurities in process liquors or to reduce acid concentration. The moisture content of the leached ore 30 may be 5 to 25% v/w, or more commonly 8 to 20%, depending on the characteristics of the ore material.

The chloride in the leach solution 20 in process 10 may come from the ore 12 itself, but typically must be added in the form of a salt 16. This is because most ores typically contain very little soluble chloride. For comparison, the amount of chloride required for the leaching process 10 is 100-230 g/l in the leaching solution 20.

The amount of chloride added to carry out heap leaching should be equal to the amount of loss. The only mechanism by which the chloride-containing salt is removed from the described process is the formation of precipitated salts and chloride-containing moisture in the leached heap residue, i. in the leached residue 30. Flushing outside the circuit (without returning to the internal cycle) contributes to an increase in the loss of chloride, acid and copper, as well as the removal of dissolved metals and salt impurities from the circuit. Typically, high copper solution flushes are not carried out outside the loop. Thus, washing is used only as a last resort.

The loss of chloride must be compensated. The chloride may be added in the form of sodium chloride or in the form of a similar salt. Salt must be purchased and brought to the industrial site. In general, under normal conditions, there is usually a net reduction in solids in the ore of about 5% due to the loss of gangue minerals and ore from the leaching process, which compensates for the small gain due to precipitation of salts. In short, it should be noted that the precipitation of Fe increases the mass, since the precipitated iron salts contain sulfate ions, hydroxide ions, and Na or K, so the mass of the precipitate is greater than the mass of metallic iron and leached sulfide (due to the addition of S and oxygen from solution ). The weight gain due to precipitation compensates for the weight loss due to the dissolution of copper and gangue minerals, but a total net weight loss of about 5% can generally be expected.

The cost of salt and its transportation depends on the location of the industrial site. Costs can be as low as $2.5/t or as high as $40/t (values shown are examples only). The initial capacity of a high chloride heap leach is typically 10,000,000 tons of ore per year and can be increased substantially. Thus, the cost of acquiring and transporting salt can be significant.

In FIG. 4 depicts a process 10A according to the present invention, which is similar in many respects to the process shown in FIG. 1. Water 37A includes make-up water 36A, wash water 38A, and make-up water 40A for electrolytic separation.

Salt 16A includes the salt added at the start of the process and additional salt (make-up salt) required to maintain the required chloride concentration in the process solution, such as solution 20A.

Make-up water 36A is fed into the leached ore residue 30A to wash the ore and to displace the entrained high chloride solution into the wash solution stream 60A. Make-up water 36A has a low chloride content and wash solution stream 60A has a high chloride content. Wash solution stream 60A is fed to the leach solution pool 24A or any other suitable pool in the heap leach scheme. Thus, a substantial proportion of the original salt content is retained in said process.

After irrigating with aqueous leach solution 20A, it is preferable to allow a rest period of 5 to 50 days, typically 20 days, before flushing dump 18A with make-up water 36A to allow the entrained irrigate solution to run off dump 18A to achieve a residual moisture content of 8 to 20%, usually 10%. Flushing the stockpile after runoff improves flushing efficiency.

The flush period for supplying make-up water 36A to the leached ore residue 30A is 5 to 100 days, preferably 10 to 50 days, or typically 20 days.

Process 10A is associated with salt savings, but an additional benefit is that some of the residual soluble copper remaining in the leached ore 30A is returned to process 10A, where it can then be recovered using a solvent extraction step 26A and an electrowinning step 28A.

In FIG. 5A is a wash efficiency curve calculated from the displacement of soluble copper from the leached ore. The percentage of displaced copper refers to any soluble species, including chloride, using copper as an indicator due to the accuracy of the analysis of the solution. The curve shown shows the likely operating range R where a balance between process make-up water (36A) per tonne of ore and soluble ion recovery (expressed as a percentage) is expected to be achieved. The specified range may vary depending on the circumstances. The R range is 0.05 to 0.35 m ³ /t ore and preferably 0.1 to 0.15 m ³ /t ore. The curve in Fig. 5A illustrates that the optimum washing efficiency in the indicated R range is 50 to 70%. The target recovery when adding wash water of 0.1 m ³ /t is 60%.

In FIG. 5B shows the decrease in make-up salt addition in kg/t as a function of the addition of wash water 36A. The reduction in the amount of make-up salt, which is directly related to the release of salt, is achieved by washing the ore residue with the specified amount of make-up water 36A as wash water.

An effective way to maximize the amount of make-up water 36A available for flushing is to reduce the amount of water fed to the process with other streams. Water enters the specified process with rain 32A, acid 14A, as a result of the reaction with acid, with

- 4 041170 wash water 38A for solvent extraction, and make-up water 40A to compensate for losses during electrolytic separation. Evaporation of 42A results in water deficiency.

To maximize the amount of water that can be supplied to the washing stage, the following factors can be considered. The amount of wash water 38A can be reduced by using the multi-stage countercurrent approach 70 shown in FIG. 6, which allows more make-up water 36A to be used to wash the leached ore 30A (or leached residue).

Process 70 includes a first step 70A, a second step 70B, and a third step 70C. You can add more stages, but this will reduce profits and increase capital costs. Optimum performance for process 70 is expected to be achieved using 3 steps. Loaded organic phase 51 is fed from the extraction process 46 to a mixer 72a for mixing with the first intermediate spent wash water 38Xb with a high chloride content from the settling tank 74b in the second stage 70B of the previous cycle, and then fed to the settling tank 74a for separation into the washed organic phase 49a and the spent wash water 38Xc with a high chloride content.

The organic phase 49a is fed to mixer 72b (in the current second stage 70B) to be mixed with the second intermediate waste wash water 38Xa with a high chloride content (obtained from the waste wash water in the third stage 70C of the previous cycle), and then to the settler 74b for separation into intermediate the high chloride organic phase 49b; and the intermediate waste wash water 38Xb (which is fed to the first stage 70A of the next cycle).

Organic phase 49b is fed to mixer 72c (current third stage 70C) to be mixed with fresh wash water 38A and then to settler 74c to be separated into washed organic phase 49c and high chloride intermediate waste wash water 38Xa (which is fed to second stage 70B next cycle). The washed organic phase 49c is fed into the stripping process 48.

Solvent extraction plants use mixers/settlers to contact and separate the aqueous and organic phases, respectively. These two liquid phases are combined in a mixer to obtain a dispersion in which the aqueous liquid forms a continuous phase and the organic liquid is dispersed therein as a dispersed phase. The aqueous/organic phase dispersion is fed to a settling tank where the two liquids are separated into an organic phase and an aqueous phase. The separation is not ideal and each phase contains some other liquid trapped in it.

The obtained organic phase 51, which comes from the extraction or loading stage in the solvent extraction process 46, is fed to the washing stage 58.

In FIG. 7A, 7B and 7C, respectively, show a one-stage countercurrent washing circuit, a two-stage countercurrent washing circuit and a three-stage countercurrent washing circuit used to reduce the chloride content in the organic phase 51.

The organic phase 51, which is fed to the washing stage 58, contains the entrained product leach solution, which typically has a chloride content of about 150,000 ppm. It is usually necessary to reduce this value to about 50 ppm. before desorption of copper into electrolyte 54.

In a single wash step 58A, as shown in FIG. 7A, the organic phase 51 from the extraction step 46 is combined in the mixer with the low chloride aqueous solution 38A. The washing is not ideal, and some of the entrapped chloride ends up with the entrapped solution in the final organic phase 49a.

In FIG. 7B depicts two cycle countercurrent washing steps 58A, 58B. The resulting organic phase 49b has a lower content of entrapped chloride than the organic phase 49a obtained in the single wash step 58A. However, when performing a high chloride leach, a two cycle wash step may not be sufficient to reduce the chloride content of the resulting organic phase unless a large amount of wash water is used.

Process 10A uses a three-stage countercurrent washing circuit 58A, 58B, 58C (FIG. 7C). In this scheme, the chloride concentration in the organic phase 49c coming from the last stage 58C is reduced to below 50 ppm, while the wash stream contains only about 1/75 of the organic phase obtained in the extraction stage 46.

The backwash water content of the multi-cycle countercurrent wash is reduced to less than one-third of that used in a traditional single-stage wash circuit. The saved water can be used as wash water 36A as shown in FIG. 4, and then can be used to wash the leached ore. This maximizes the washing efficiency in terms of chloride recovery and reduces the loss of soluble copper in the leached ore.

- 5 041170

The following description relates to experimental work carried out to test the described principles of the present invention.

Example 1 Pilot Evaluation - Integrated Pilot Plant.

An integrated pilot plant was assembled containing 9 barns and a solvent extraction unit to reproduce the application of the method according to the present invention on a pilot scale. Said plant is configured to process various low grade chalcopyrite ores with dynamic race track style dump leaching operation. It is a dump consisting of several sectors, with new sectors laid on top, and older, most leached sectors, removed from the bookmark, which is then reused.

Each barn has a cross-sectional area of 4 m ² , a working height of 7.5 m and an overall height of 10 m and contains approximately 40 tons of ore. Barns were operated to simulate industrial operation in 9 sectors. The proposed method is dynamic with periodic removal of the leached residue from the leached pit and replacement of the leached residue with fresh ore in the free pit.

Three separate phases of operation were carried out. Various compositions of the major copper sulfide minerals and gangue minerals contained in the ore samples were tested, and a summary of the operation is summarized in Table 1. 1.

Table 1. Technological stages of work

P Parameter Ore particle size, passing 80% of particles (mm) Copper content (%) Chalcocyte content CSR* (%) Covellite content CSR* (%) Chalcopyrite content CSR* (%) Pyrite (%) Chlorite(%) [ semi-industrial Stage 1 19.05 0.49-0.75 >10 10-20 40-75 2-6 0.1 - 1 oh installation Stage II 19.05 0.38-0.79 10-20 < 10 65 -75 2-4 0-1 Stage III 19.05 0.3-0.6 10-20 < 10 40-75 3-7 0-9 Biotite(%) 0-0.3 0-0.03 0-2 Kaolinite (%) 8-16 10-45 10-30 Stage description exploitation Launch and collection of initial data Experimental change of leaching conditions and ore type Closed loop operation

*CSR - copper content in the raw material (total percentage of copper contained in the mineral).

The ore samples were crushed in a 3-stage crushing cycle. The crushed ore was subjected to agglomeration before being loaded into barns. In the process of agglomeration, sodium chloride (salt), acid and raffinate (or other process solution containing copper, iron and acid) were added to the ore. Then the agglomerated ore was placed in an empty barn; dried (kept) for some time, then irrigation was started. At the end of the irrigation cycle, each barn was drained, sprayed with rinsing water, drained again, and finally unloaded. Then the empty barn was prepared to receive a fresh batch of agglomerated ore.

Phases I and II relate to the start of operation and data collection.

In the III phase of operation, low-grade ores were used with a copper content of only 0.3% Cu and with a copper content bound in chalcopyrite up to 75% (CSR of chalcopyrite 75%). The work cycle consisted of 45 days of drying period, 20 days of wetting period, 360 days of irrigation, 20 days of washing and 30 days of draining (475 days in total), after which the samples were unloaded and processed.

The phase III operation schedule is designed to load and unload the barn every 45 days, which is a model of industrial dynamic operation of backfill and industrial fluids. This shutdown was necessary to obtain equilibrium impurity concentrations and identify possible problems during operation. The work was carried out in a closed system with repeated supply of wash water to obtain a raffinate intended to compensate for the raffinate lost due to residual moisture and evaporation. The water feed rate was maintained at 0.11 m ³ /ton of ore. Fresh make-up water was used to flush out organic compounds added to remove chlorides and to flush out residual ore in accordance with the method of the present invention.

The pilot plant was operated as a closed system, carefully regulating incoming and outgoing

- 6 041170 outflows for simulating an industrial heap leaching process, i. e. fresh water additions were limited to compensate for evaporative water loss, moisture loss in the leached ore (or leached residue) and to replace water lost during process brine washing (if necessary to reduce impurities). The technological scheme of the semi-industrial plant is shown in Fig. 8. Below is a summary of the main components and processes of the pilot plant.

Raffinate tank 1. Storage of the return raffinate from the solvent extraction unit 7 (SX). Provides separate 1 m ³ 2A-I raffinate feed tanks to each 3A-I barn. Make-up acid can be added to the raffinate tank as needed to meet the process flow rate of acid in said process.

Productive leach solution (PLS) 4 and 5. The PLS solution from each barn (3A-I) is collected in 1 m ³ PLS tanks 4A-I. PLS is fed from tanks 4A-I to storage tank 5 PLS.

PLS 6 with high Cu content. The first PLS harvested after the initial irrigation of the barns after drying the ore has a high copper content. The original PLS is collected in tank 6. Make-up acid can be added to high copper PLS tank 6. Tank 6 for high copper PLS can be filled with SX wash water.

Agglomeration of 8 ore. The ore is agglomerated in a sinter drum. The raffinate from tank 1 is added to the ore. Acid and solid salt are added to achieve the required amount of added acid and total salt, as needed. High copper grade PLS 6 can be used for ore agglomeration to provide direct return of copper and acid (contained in high Cu grade PLS) to ore agglomeration. The high content of copper in acid agglomerated ore can improve the dissolution of copper during the drying of the ore in the initial curing stage. The agglomerated ore is fed into barns (3A-I) as needed for loading.

The ore is leached in barns 3A-I. The agglomerated ore is loaded into barns. It is left to dry (the initial standing period). Then irrigation begins. Irrigation is started slowly to ensure wetting of the ore. Irrigation is accomplished by feeding the low copper raffinate from the solvent extraction process to the top of the barns. The solution is distributed over the surface of the barn using a drip system. In each barn, the solution passes down through the ore. At the same time, air is supplied in the lower part of each barn. Copper is solubilized by the combined action of the acid in the process solution and oxygen in the air. Irrigation solution enters the base of the barn. Now it has an increased copper content and is called PLS (productive leaching solution). It is collected and fed into the solvent extraction process.

After leaching of the ore, irrigation of the ore with raffinate is stopped. A layer of ore is left to drain off the liquid, and then the ore is washed with fresh water to recover the trapped chloride (salt) and dissolved copper. Then the washed leached ore residue is unloaded from the barn to waste.

Solvent Extraction 7 (SX): Copper is recovered from PLS by solvent extraction. Copper enters the organic phase in two extraction steps, E1 and E2. The loaded organic phase is then washed with water in two washing steps, L1 and L2. Washing is necessary to remove the entrapped aqueous solution so that the chloride content of the purified electrolyte after separation of the loaded organic phase is <50 ppm. After the washing step, separation of the loaded organic phase in S1 is carried out to recover copper to obtain a purified electrolyte 7b. The spent electrolyte for separating the loaded organic phase is fed from the storage tank 7c. If necessary, make-up acid can be added to the spent electrolyte in the storage tank 7c to increase the acid concentration to the amount necessary to complete the separation of the loaded organic phase. The purified electrolyte is removed from the copper recovery circuit, and the spent electrolyte is returned to the circuit. The washing efficiency of the loaded organic phase can be improved by increasing the number of washing steps to three (see FIG. 7C). Increasing the washing efficiency reduces the amount of wash water required to wash the loaded organic phase, which allows more water to be washed in the leached ore residue, which increases the recovery of chloride and copper by displacing the solution retained in the ore residue.

- 7 041170

Table 2. Pilot Plant Operating Parameters for Stage I, Stage II and Stage III

Parameter Units Stage 1 Stages II and III Barn irrigation area (per barn) m ² 4 4 Ore layer height m 5.4 7.5 Loading ore (on the barn) T 36 44 Acid in agglomerated ore kg/t 12 12 Target moisture content in agglomerated ore % wt. 8.5 8.6 Bulk density of ore t/m ³ 1.65 1.60 Initial standing or drying period days 45 45 Ore bed temperature °C 15-30 15-30 Aeration rate Nm ³ /h/m ² 0.13 0.13 and 0.325 Application rate of raffinate by period - within 9 hours at 6 l/h/ ^m2 within 12 hours at 6 l/h/ ^m2 Cycle operation leaching days Irrigation 200-500 20 days of low irrigation (wetting the ore), 360 days of irrigation, 20 days of washing cycle. Operation of the technological scheme - Closed with SX Closed circuit with SX and flushing stage Raffinate solution for closed circuit - 5 g/l Cu, 15 g/l Fe, 8 g/l H ₂ SO ₄ , 150 g/l CI 0.5 g/l Cu, 1-3 g/l Fe, 8 g/l H2SO4, 150 g/l CI

A simplified mass balance diagram for a semi-industrial dump, showing incoming and outgoing flows, is shown in Fig. 9. Typical values for the balance of ore, acid, salt and water are presented in table. 3. Results in table. 3 show that the amount of salt recovered from washing the ore at a process target of 0.11 m ³ /t water is 9 kg/t, indicating a significant savings in terms of required make-up salt. For the ore processed in stage II, the total amount of copper recovered, including copper recovered from ore washing (as shown in Example 2), is 60% of the copper content of the original raw ore.

- 8 041170

Table 3. Generalized mass balance of a semi-industrial plant, showing incoming and outgoing flows

Mass balance of pilot plant in phase II Internal recycling Incoming outgoing Input (kg/t) Output (kg/t) Ore balance Total load of dry ore (t) 9 barns 348.6 - Total copper load (0.5% CuT) (t) 1.743 5.00 Total copper recovered at 60% average recovery (t) 1.046 3.00 Total loss of copper in the balance (t) 0.697 2.00 Total residue without payload (t) 348.6 - acid balance Amount of acid added for agglomeration at a rate of 12 kg/t (kg) 4183 12.0 Loss of acid in residual moisture at a moisture content of 10.5% (kg) 18 0.05 Total consumption of acid (for the dissolution of putrid rock and metals) at a consumption of 13.2 kg/t NAC, kg) 4616 13.24 Amount of make-up acid added (kg) 451 1.29 Salt balance Total salt added: raffinate + solid make-up salt at a rate of 15 kg/t (kg) 5229 Amount of solid make-up salt (kg) 2092 6,000 The amount of salt released during the washing of the ore residue with a washing efficiency of 60% (kg) 3137 Loss of salt in residues (kg) = amount of make-up salt 2092 6,000 Water balance Water for washing the organic phase at SX (m3) Amount of washing water at a flow rate of 0.1 mS/t (m3) Moisture content of the residue at an average value of 10.5% (m3) 4.308 34.86 36.6 Water loss due to evaporation at 5.4 l/day (m3) 2.57 Released water (m3) zero The amount of added raffinate at a flow rate of 0.1 m3/t (m3)) 33.5

In table. 3 shows that the amount of wash water for solvent extraction was 4.3 m ³ for 348 tons of ore processed in said process during this period. The amount of wash water used to wash the ore after leaching was ~8x greater than for 34.9 m ³ , so there is very little range in this case to improve ore wash by further reducing solvent extraction wash. However, solvent extraction wash performance at this wash volume was unsatisfactory for a two-stage configuration. The chloride concentration in the entrapped solution in the loaded organic phase after washing was about 600 ppm. The amount of wash water available for solvent extraction can be increased, but this will adversely affect the amount of water that can be used to wash the leached ore, and is likely not to significantly reduce the chloride content of the washed organic phase. This can result in increased loss of soluble copper and increased salt consumption. Another option, as described above, is to increase the number of solvent countercurrent washing steps to 3 or more steps (FIG. 7C). This will provide a significant reduction in the chloride content of the organic phase after washing without the need to increase the amount of washing water.

This example demonstrates the relationship between overall process water balance and SX flush stage configuration and its effect on overall plant performance. This example demonstrates that for efficient washing of the loaded organic phase, several washing steps are necessary to minimize the volume of added washing water and to achieve the desired chloride content in the purified electrolyte. The result obtained corresponds

- 9 041170 principles described above in relation to maximizing the amount of water available for washing the residue.

Example 2 Wash Efficiency Curves and Recovery of Copper and Chloride.

Copper Recovery: The results shown in FIG. 10 show the copper recovery and concentration profile of copper in PLS (or effluent after irrigation) versus time for the operation of two pits, pit 1 in Stage I (ore A; 0.5% Cu and chalcopyrite with a CSR of 60%) and barn 5 in stage II (ore B; 0.65% Cu and chalcopyrite with a CSR of 75%), operating a pilot plant.

In barn 1 at stage I, the copper content in the irrigation solution or in the raffinate was 5 g/l. The results show that for the treated ore sample, the copper recovery at the end of the leaching cycle was 59%. The leached residue was allowed to drain and then washed with water to recover the trapped dissolved copper. The displacement of copper from the residual ore by washing with water increased the copper recovery to 70% as shown in FIG. 10.

In barn 5 at stage II, the copper content in the irrigation solution or in the raffinate was 0.5 g/l. The results show that for the treated ore sample, the copper recovery at the end of the leaching cycle was 59.5%. The leached residue was allowed to drain and then washed with process water to recover the trapped dissolved copper. Displacing the copper in the water wash increased the copper recovery to 60% as shown in FIG. 10.

The reason why the copper recovery in pit 1 increased by 11% and in pit 5 only increased by 0.6% is because pit 1 was irrigated with a solution with a high copper content (5 g/l), and therefore , a significant amount of soluble copper remained in the pit at the end of the leaching. Barn 5 was irrigated with a solution containing 1/10 of the concentration of copper in the solution used for barn 1. Consequently, the amount of soluble copper remaining in the barn before washing was much lower.

Proper operation of the solvent extraction process can reduce the loss of soluble copper by providing a low copper raffinate that is used for irrigation.

Chloride recovery: As stated, proper operation of the solvent extraction process can reduce the loss of soluble copper by providing a low copper raffinate that is used for irrigation. But this does not apply to chloride. This process requires an increased chloride content in the irrigation solution (~150 g/l) to maximize the leaching of chalcopyrite. Residual chloride that remains trapped in the leached ore (leached residue) that is removed from the heap must be replenished by adding sodium chloride or an alternative chloride source (HCl, MgCl, etc.).

The efficiency of the wash with respect to chloride recovery is shown in FIG. 11. The given data are obtained from the results of washing barn 5 with process water at stage II. The results obtained demonstrate an increase in the release of chloride with an increase in the volume of wash water used. In order to maintain the water balance of the technological scheme of a semi-industrial plant, it is necessary that the amount of water used to wash the rest of the ore is limited so that the loss of water in this process is less than the amount of fresh water supplied. Water is lost as moisture in the residual ore removed from the pilot plant through evaporation, and may also be removed when the solution is removed from the circuit.

The other fresh water stream entering the loop is fresh water used in the washing step of the solvent loaded organic phase in the extraction to remove entrapped chloride.

The maximum volume of water available for washing the residual ore is in the order of 0.1 m ³ /t ore. The results shown in the example shown in FIG. 11 indicate that with this volume of wash water per ton of ore, the wash (or chloride recovery) efficiency is about 60%.

The chloride recovery achieved by washing with water can also be expressed as the equivalent amount of chloride that must be added to compensate for the loss of chloride in moisture in the residual ore. The amount of make-up salt (NaCl) for the washing curve of barn 5 in step II shows that at a wash rate of 0.1 m ³ /t, the amount of make-up salt (NaCl) is about 18 kg/t, as shown in FIG. 12.

Example 3. Accumulation of impurities.

Washing is believed to result in chloride retention and a slight increase in copper recovery. On the negative side, unwanted compounds that are formed as a result of the interaction of acid with gangue minerals are also retained in this process.

The interaction of gangue minerals with acid leads to the dissolution of metals such as Ca, Si, Al, Mg, Fe and K. Solutes accumulate in solution until they reach such a concentration that the mass of dissolved substances in the washed ore after leaching will be equal to wt -

Claims (9)

solutes, which is less than the mass of any substances re-precipitated after leaching. The accumulation of impurities in the raffinate of the pilot plant during operation in a closed circuit with washing is shown in Fig. 13. The concentration of aluminum and magnesium (dissolved by acid dissolution of silicate gangue minerals such as biotite and chloride) increases markedly during closed circuit operation and reaches concentrations of approximately 10 g/l and 5 g/l, respectively. The total concentration of iron is 2-4 g/l. The total sulfate concentration is approximately 80 g/l. In general, the accumulation of such compounds together with the addition of sodium chloride leads to the formation of saturated solutions. As is known, sulfate precipitates from solution in the form of insoluble iron jarosites (jarosite Na or K, for example), in the form of meta-sideronatrite (Na ₂ Fe ³⁺ (SO ₄ ) ₂ (OH) - (H ₂ O)], and also at low temperatures and high degrees of saturation, possibly in the form of sodium sulfate, aluminum sulfate and/or magnesium sulfate Precipitation of Na, Al or Mg or Fe sulfates depends on temperature (increased precipitation at low temperature) and acidity, with precipitation jarosite and meta-sideronatrite are enhanced at low acidity values, i.e. at higher pH values of the solution. Precipitation of iron sulphates and sulphates of other metals can cause problems during the operation of the circuit, in particular at low ambient temperatures (<20°C) and with increased salt precipitation. Possible problems include: deposition of sulfate salts on the drip lines used to irrigate the heap, causing clogging of the lines and inability to effectively irrigate the heap; sedimentation of salts in the pools with solutions, reducing the effective volume; and sedimentation in the layers of the heap, causing clogging of the pore space and disruption of the flow and distribution of the solution in the ore layer, with a possible decrease in the degree of copper recovery. The amount of sediment depends on the degree of washing. There may be a point in this process where the optimization of chloride recovery is offset by the costs associated with process problems due to accumulation and precipitation of undesired compounds. In this case, it may be necessary to reduce the degree of flushing. CLAIM 1. A method for maximizing the amount of water available for washing in a high chloride heap leach followed by organic solvent extraction, comprising the step of using process make-up water in an amount of 0.05 to 0.35 m ³ /t of ore to wash the leached ore residue into dump, which ensures the displacement of the chloride-containing aqueous solution from the leached ore, and the amount of process make-up water available for washing is increased by using, after extraction of the leached solution with an organic solvent, a multi-stage countercurrent washing circuit loaded with organic solvent to reduce the concentration of chloride in the organic phase supplied from the previous extraction process, while the concentration of chloride in the organic phase is reduced to a value of less than 50 ppm. 2. The method according to claim 1, characterized in that the amount of process make-up water at the washing stage is from 0.1 to 0.15 m ³ /t of ore. 3. The method of claim 1, comprising the step of increasing the amount of process make-up water available for flushing by at least one of the following: (1) increase the evaporation of water from the leaching circuit; (2) prevent rainwater from entering the dump; and (3) reducing the consumption of make-up water to compensate for losses in electrolytic separation through efficient washing of the loaded organic phase. 4. The method according to claim 1, characterized in that the displaced chloride-containing aqueous solution is recovered and returned to the heap leach process, reducing the amount of make-up salt added to the heap leach. 5. The method according to claim 1, characterized in that the washing step is carried out for 5 to 100 days. 6. Method according to claim 5, characterized in that said period is 20 days. 7. The method according to claim 1, characterized in that copper is isolated from the chloride-containing aqueous solution displaced in the washing step, which is returned to the leaching, and then recovered in the solvent extraction step and the electrowinning step. 8. The method according to claim 1, characterized in that the ratio of chloride separated from the chloride-containing aqueous solution, displaced at the stage of washing with process make-up water, to its total content in the residual ore before washing is from 40 to 85%. 9. Method according to claim 8, characterized in that said ratio is 60%. -