EA041234B1 - ACID BALANCE IN CHLORIDE HEAP LEACHING - Google Patents

Description

State of the art

The present invention relates to chloride leaching of copper ores.

Heap leaching of low grade copper sulfide ores using acid sulfate solutions with bacteria is a traditional industrial process used throughout the world.

Recently, chloride leaching processes have been developed that allow leaching copper sulfide ores, including chalcopyrite, to high expected productivity without the use of bacteria. WO 2014/030048 and WO 2015/059551 describe that the use of high chloride concentrations (100 to 230 g/l chloride) overcomes the passivation of chalcopyrite, allowing operation at a solution oxidation potential of more than 700 mV, in contrast to the SHE (Standard Electrode System ), which leads to an increase in the degree of leaching of chalcopyrite and the achievement of a higher degree of extraction of copper.

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. Heap leaching technologies use various configurations to optimize the copper concentrations in the respective loops. This is done to minimize the capital cost of building and operating solvent extraction facilities and to reduce copper losses. Heap leach flowsheets range from single settling systems to multiple settling systems that include settling tanks known as ILS (Intermediate Leach Solution). They are included in the overall process as an intermediate step to increase the copper content in the PLS (productive leaching solution).

In the leaching of copper sulfide minerals using highly concentrated salt solutions, a step of copper pre-solubilization and reduction of gangue acid consumption by means of a pre-leach step (described in WO 2016/094956) or a standing step (described in WO 2015/059551) is used, in which the ore and the solution is left in contact for a period of time to accelerate the dissolution of the reactive minerals in the ore prior to continuous irrigation of the dump. Then, in each case, a leach phase is carried out, which involves balancing the acid in the process flow and maintaining an acceptable lower acid concentration in the PLS compared to more traditional acid-sulfate systems. A reduced acid concentration is necessary to ensure a stable operation with minimal release of the solution from the flow sheet.

Definitions

The following expressions, as used in this specification, have the meanings indicated unless the context clearly dictates otherwise.

Ore and grade:

An ore is a type of rock that contains sufficient minerals with valuable elements, including metals, that can be economically extracted from the rock (ref. 1). Ore refers to mined material and not to a concentrate that is obtained by selectively separating and extracting ore minerals from the rock or gangue components of the ore. This material can be processed in the form in which it is mined, i. in the form of rock mass (ROM), or, after grinding, in the form of ground ore. The grinding of the ore increases the surface area of the ore minerals to be treated and increases the degree of recovery of the metals they contain in the heap leaching process.

Traditional agglomeration of ore:

In the heap leaching process, it is common practice to carry out a agglomeration of the crushed ore, followed by a drying (aging) step after the stockpile has been created. The crushed ore particles are mixed with the process solution and acid in the drum for agglomeration to a constant moisture content. The agglomeration step produces small particles that adhere to larger particles due to the formation of liquid bridges under the action of surface tension. The drying step after the spoil is created allows the acid added during the agglomeration of the ore to interact with the ore minerals and accelerates the dissolution of metals from the ore minerals during subsequent irrigation of the spoil. The traditional agglomeration process results in the formation of homogeneous particle agglomerates, in which small particles are agglomerated with larger particles throughout the heap (stack) to form a homogeneous and permeable ore layer, which accelerates the flow to irrigate the dump and dissolve metals in ore minerals. When processing ores in which it is difficult to obtain uniform and stable particle agglomerates, binders can be added during the agglomeration step to bind the particles together. Typical binders for acid leaching of ores are based on organic polymers such as polyacrylamides.

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Technical factors relevant to the present invention

The purpose of ore agglomeration in chloride heap leaching of copper:

In the proposed copper heap leaching process described below, a agglomeration process can be used to mix crushed ore particles with active leaching agents such as sulfuric acid, chloride, copper and iron. The mixing of reactants with the ore to create a surface film of solution around the ore particles is the main advantage of the agglomeration step in the proposed copper heap leaching process described herein. The formation of ore agglomerates and the interaction of acid with ore minerals, which occurs during the traditional agglomeration of ores, is of secondary importance. The ore agglomeration process in the proposed copper heap leaching process, described below, allows the reactants to be uniformly distributed in the solution film around all the ore particles. The solution film creates a large reactive surface area around the ore particles. After the creation of a pile of agglomerated ore and during the initial standing period after aeration, the mass transfer of oxygen from the air to the surface of the ore particles increases significantly due to the surface films surrounding the ore particles, and the dissolution of metals in ore minerals is significantly enhanced due to the rapid formation of Fe(III) oxidizers. and Cu(II) as a result of the interaction of dissolved oxygen with Cu(I). The increased surface area available for mineral oxidation and acid supply accelerates heat generation and increases the temperature of the ore during ore agglomeration.

Ore wetting cycle:

After completion of the ore agglomeration stage, chloride leaching may include five stages, namely: stages of (a) drying, (b) wetting, (c) irrigating, (d) standing or draining, and e) washing and final draining. The stage of wetting the ore is carried out to prevent the formation of channels for the movement of the irrigation solution through the dump. After drying, the ore is periodically irrigated with a raffinate solution or over a period of 10 to 50 days, typically 20 days. During this period, the irrigation time is gradually increased from 30 minutes to 9 hours, while adding a sequence of stages of drying or standing, as shown in table A. In table. A shows that for a 12-hour irrigation cycle on the first day, the dump is irrigated for 30 minutes in the morning, and then the ore is left to dry for 23.5 hours. From Days 2 to 5, the dump is irrigated for 30 minutes in the morning and again after 12 hours drying for the same period of time. This process is repeated with a corresponding increase in the duration of the cycle of irrigation/drying up to 20 days, when a 24-hour cycle is carried out, consisting of 12 hours of irrigation and 12 hours of drying.

Table A

Day Cycle Irrigation time x drying time cycle day number Total irrigation time (h) eleven 0.5 x 23.5 1 0.5 ² 0.5 x 11.5 2 1 ³ 0.5 x 11.5 2 1 ⁴ 0.5 x 11.5 2 1 15 0.5 x 11.5 2 1 To ⁶ 0.5 x 7.5 3 1.5 S 7 0.5 x 7.5 3 1.5 go with S ⁸ 0.5 x 7.5 3 1.5 ha ¹⁹ 0.5 x 7.5 3 1.5 § 10 0.5 x 3.5 6 3 O 1 11 0.5 x 3.5 6 3 S Q. w ¹² 0.5 x 2.5 8 4 With ¹³ 0.5 x 2.5 8 4 1 14 0.5 x 2.5 8 4 ¹⁵ 0.5 x 1.5 12 6 1 ¹⁶ 0.5 x 1.5 12 6 ¹⁷ 0.5 x 1.0 16 8 18 0.5 x 1.0 16 8 1 19 0.5 x 1.0 16 8 | 20 0.5 x 0.5 24 12

The inclusion of a wetting step in the overall process is important because it minimizes preferred raffinate liquor flows after reflux has started. It can also be used to replenish acid in the ore before irrigation; this is especially important for denser ores that do not form fine particles during the preparation stage and do not absorb sufficient moisture during the agglomeration stage. The acid is critical to the proposed process and must be provided in sufficient quantity to compensate for the consumption due to the dissolution of gangue minerals and copper minerals.

A complete five-stage heap leach cycle may include the following periods, which are given by way of example and not limitation:

initial drying of the ore or maturation stage - 45 days;

ore wetting - 20 days;

heap irrigation - 360 days;

drainage stage - 2-20 days;

- 2 041234 washing/washing stage - 20 days;

drainage - 30 days;

the total heap leaching cycle is 495 days.

The dumps are aerated during the heap leaching cycle, from the initial ore drying stage to the completion of the dump irrigation stage. Aeration may be continued during the flushing and final draining of the stockpile to create favorable conditions for high oxidation potential solutions in the flow sheet. The aeration rate is typically, but not limited to, 0.05 to 2.0 Nm ³ /h.m ² of blade area for a 6 meter blade, and preferably 0.1 to 0.4 Nm ³ /h.m ² and about 0.45 Nm ³ /h.m ² for leaching copper sulfide ores in an acid solution with a high chloride content. The rate of aeration is determined by the oxygen consumption of the leached minerals, as well as by the height or weight of the ore collected in the dump.

Acid control

In the high salinity leaching flowsheet, the process begins with acid agglomeration and continues with a drying and settling phase. During these time periods, acid is consumed as a result of the interaction of acid with the gangue, as well as the solubilization of copper in copper sulfide minerals such as chalcocite, covellite and chalcopyrite, in accordance with the general reaction, designated as reaction 1.

Cu _x Fe _y S _z + (x + y) H ₂ SO ₄ + (x + y) O ₂ -> xCuS0 ₄ + yFeS0 ₄ + (x + y) H ₂ O + zS Reaction 1 where x, y and z represent

Mineral X At Z chalcocyte 2 0 1 covellite 1 0 1 bornite 5 1 4 Chalcopyrite 1 1 2

During the drying period, the temperature rises and the concentrations of the reactants become high, so this period is favorable for leaching, since both factors contribute to the reaction kinetics. This reduces the total time it takes to leach the ore and saves capital and operating costs. In order for reaction 1 to proceed as completely as possible, it is necessary to provide a sufficient amount of acid, since it is a necessary reagent (according to reaction 1). In this drying/resting step, acid is also consumed in reactions with acid-soluble gangue minerals. The rate of these reactions also increases at elevated temperature. To accelerate the leaching of copper, sufficient acid is needed for these competing reactions to occur with the gangue.

At the end of the drying/resting period, typically (but not limited to) up to 50% of the copper expected to be dissolved in said process is solubilized and most of the readily soluble gangue minerals become leached. In the next leach phase, said copper, together with the remaining 50% leachable copper, is recovered from solution by solvent extraction (SX) coupled with electrowinning (EW). This process produces 1 mol of acid for every mol of copper deposited by electrowinning. The resulting acid is returned to the heap leach process via solvent extraction.

Email current h Reaction 2

Cu5O ₄ + H ₂ O ------ "Cu + H ₂ 5O ₄ + 7O ₂

Thus, during the leaching stage by reflux, in which all the solubilized copper is recovered, net positive acid formation can occur. In table. 1 shows the different acid rates for various types of minerals to leach 50% additional copper and acid production in the SX/EW stage for total soluble copper.

Table 1

If the residual mineral after drying is: Moles of acid per leach per ¹ X moles of Cu leached in the reflux phase (reaction 1) Moles of acid formed in the EW process for each mole of precipitated Cu (reaction 2) Net acid production, mol per mol precipitated Cu Covellite/chalcocite 0.5 1.0 0.5 bornite 0.6 1.0 0.4 Chalcopyrite 1.0 1.0 0.0

The result is that if all the residual copper is chalcopyrite, then the acid balance is neutral, otherwise, for all other copper sulfides, the balance is acid positive.

The neutrality of the balance for chalcopyrite in practice also does not correspond to reality. This is due to the fact that for every mole of copper leached from chalcopyrite, one mole of iron is also leached. For example, grades with a low initial content correspond to approximately 0.4% copper. If we take into account that about 30% of copper is supposedly solubilized at this stage of irrigation, then even for such a low grade, this is equivalent to 1.2 kg _Csc / t _{of ore} . The result is 1.1

- 3 041234 kg _Fe ( _II ) / t _{of ore} . The only output of soluble iron from this process is the residual moisture in the leached ore (leached residue) when it is removed from the leached heap and sent to a leached residue disposal site.

A typical moisture content is about 10% (v/w). This means that at equilibrium the iron content of the solution must be 11 g/l in order for the amount formed to correspond to the outgoing amount. The method described in WO 2015/059551 includes a washing step after the end of the leaching. The washing process typically captures 70% of the dissolved salts. This method is an optimization to reduce the amount of sodium chloride required for this process, and also assumes that the equilibrium concentration of iron should be 36 g/l, so that the amount formed is equal to the amount leaving the process. In testing, limited solubility was shown to account for the total soluble iron in this process, which equilibrated at about 5 g/l (see FIG. 1). After washing, the concentration in the waste is 1.5 g/l (or 0.15 g/t _{of ore} ). This means that out of 1.1 kg/t _ore of formed iron, 86% of the iron is expected to precipitate. Iron precipitates in the described scheme according to three equations of precipitation formation:

±Na ₂ SO ₄ + 3FeS0 ₄ + ³ -O ₂ + ^H ₂ O θ NaFe ^ SOMOH ^ + 1H ₂ SO ₄ Reaction 3 ±K ₂ SO ₄ + 3FeSO ₄ + |O ₂ + ^H ₂ O θ KFe ^ SO^OH^ + ^H ₂ SO ₄ Reaction 4

Na ₂ SO ₄ + FeS0 ₄ + |O ₂ + ^H ₂ O θ Aa ₂ Fe(5O ₄ ) ₂ (OH) Reaction 5

Iron(II) is converted to iron(III) and reacts to form precipitates in reactions 3, 4 and 5. Reactions 3 and 4 represent the formation of jarosites. These reactions result in net acid formation and assume an overall net acid formation due to the dissolution of chalcopyrite in the reflux phase of the heap leach when copper is recovered by a combination of solvent extraction (SX) and electrowinning (EW).

Reaction 5 is the formation of meta-sideronatrite. The overall reaction proceeds without acid formation. The separation factor between these three reactions is difficult to predict. It depends on both the equilibrium state and the rate of the reaction. In the case of potassium jarosite, the separation factor depends on how much potassium is leached into solution from other minerals that dissolve in the process (for example, from sericite).

For other ore, the amount of jarosite, which is formed during the processing of large samples of ore in barns (square columns 2 mx2 m) in the process of heap leaching with a high concentration of chlorides, is presented in table. 2.

Table 2. Jarosite in leached residue from ore leaching in an open loop test

Experimental Abmar Number and Ore Identifier Jarosite Percentage Measured in Leached Residue Barn 1 - A 0.5 Barn 2 - B 0.4 Barn 3 - C 1.4 Barn 4 - D 1.0 Barn 5 - E 1.0 Barn 6 - F 1.2 Average 0.9

The tests were carried out in an open circuit, i.e. the solution was passed once through the experimental column and not returned to the cycle. This usually results in weaker sludge formation because the iron concentration in the feed solution does not exceed the saturated solution concentration (about 5 g/l) and there is only a small amount of additional total dissolved salts (TDS) introduced into the column in one pass, so 0 9% jarosite is the minimum amount of iron that can precipitate out of solution.

0.9% jarosite corresponds to about 3.3 kg/t of precipitated iron _ore . This is -3 times the amount of iron obtained from pure chalcopyrite. Part of the produced iron is formed as a result of the leaching of iron-bearing minerals of gangue. In the case of jarosite precipitation, for every mole of iron (II) that oxidizes and precipitates, _1/2 mole of acid (H ₂ SO ₄ ) is formed; thus, the indicated amount of jarosite provides 3 kg/t acid _ore . The net effect is that the additional amount of iron formed from chalcopyrite is likely to be affected by jarosite to some extent and hence the acid balance shown in Table 1. 1 for chalcopyrite, instead of being zero, will actually be positive due to the acid formed as a result of the precipitation of iron leached in the form of jarosite.

It has been found that at a sufficiently high degree of copper dissolution in the initial drying phase, further dissolution of copper from disulfide minerals and its recovery by solvent extraction in combination with electrowinning (EW) is likely to lead to the formation of acid. Excess acid results from:

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1) the imbalance between the acid consumption for leaching additional copper mineral in the reflux phase compared to the total acid content for EW copper leached in the drying and reflux phase if the copper minerals are chalcocite and covellite;

2) the imbalance between the acid consumption for leaching additional copper mineral in the reflux phase compared to the total acid content for EW copper leached in the drying and reflux phase, as well as the precipitation of part of the iron (jarosite), if the copper mineral is bornite; And

3) jarosite precipitation if the copper mineral is chalcopyrite.

With regard to the ore mentioned with respect to FIG. 1, there is some difference for two reasons. The high consumption of acid on the waste rock negates the benefits achieved in the drying/holding stage, and said ore contains uranium. Uranium solubilizes at a high chloride content in solution (eg 150 g/l), but it is very difficult to remove it from solution due to the high chloride content and the high concentration of iron chloride complexes. This necessitates a pre-leach stage by irrigation. It takes at least 200 days to complete and also provides significant copper leaching as well as depletion of readily soluble gangue minerals. The second part of the high chloride reflux leach provides a high degree of copper dissolution, mainly due to the dissolution of secondary copper sulfides (covellite, bornite), which are the predominant copper minerals in the ore lumps. The formation of jarosite and the acid that is released from the leaching and electrowinning (EW) of the secondary copper sulfide minerals (listed in Table 1) provide a high degree of acid formation. Together with the reduction in acid-soluble minerals in the waste rock after pre-leaching, this results in a high acid content in the PLS, as predicted for a high chloride leach process. This makes it difficult to recover soluble copper, as described below.

Solvent copper extraction is sensitive to acid concentration. This is due to the fact that such extraction affects the equilibrium between copper and acid (H ⁺ ).

The load of copper on the organic phase under conditions in which the aqueous phase is brought into contact with the organic phase provides a high copper content with a low acid content. The copper is separated from the organic phase by bringing it into contact with an aqueous phase which has a high acid content. This sensitivity becomes more pronounced at very high chloride concentrations (100 to 150 g/l) due to the significantly increased proton (H ⁺ ) activity at higher chloride concentrations.

Under conditions with low chloride content, the optimal goal is to extract about 7 g/l Cu in one pass. This provides an organic flow to water flow ratio of approximately 1:1. The amount of extractant in the organic phase is such that the viscosity of the organic phase is moderate and effective phase separation can be achieved. This ensures high manufacturability, low losses of the organic phase (operating costs) and minimal capital costs. Under such conditions, the typical acidity of PLS is 2-4 g/l, and the acid concentration in the raffinate during acid transfer is 12-15 g/l.

When performing high chloride heap leaching (typically 100 g/l to 180 g/l), it is difficult to achieve a copper transfer value of 7 g/l due to acid transfer to the raffinate and increased proton activity. pH is a direct measure of proton activity. At a chloride content of 20 g/l and a free acid of 12 g/l, the measured pH is ~1, while at an equivalent acid content, but a chloride content of 150 g/l, the measured pH is ~ 0.2.

The target for copper carryover in a high chloride leaching process should be 5 g/L or less, combined with very low acid PLS. With a copper delta of 5 g/l, this means that the SX loop size must be 40% larger for the same copper transfer efficiency as for 7 g/l. This entails an increase in capital and operating costs. The problem described is rapidly exacerbated if the acidity in PLS increases with high chloride content.

The object of the present invention is to provide a method for chloride leaching of copper-bearing ores in which the concentration of acid in the process flow is controlled to achieve efficient recovery of copper from solution by a solvent extraction method (SX) and to minimize or eliminate solution washout from the loop, which leads to increased losses. copper and operating costs.

The essence of the invention

From the foregoing it is clear that it is important to control the acid concentration in the heap leach process flow in order to achieve efficient recovery of copper from solution in the solvent extraction process. In a heap leach process with a high chloride content (100 to 180 g/l), the effective recovery of copper from solution by solvent extraction depends on

- 5 041234 maintaining a low acid concentration in the PLS. It has been experimentally shown that the concentration of acid in the PLS that enters the solvent extraction process should be less than 10 g/l, preferably less than 7 g/l and more preferably less than 5 g/l.

The advantage is that the addition of acid during the ore agglomeration step is not limited, that the dissolution of copper minerals is not limited by the acid, and that the copper dissolution achieved during the drying of the ore is maximized.

The configuration of the heap leach flowsheet can be adapted, with leaching being carried out in multiple dumps or dump sections to maximize the amount of acid consumed by the residual acid soluble gangue minerals and to control the acid concentration in the PLS.

In the case of using two or more leaching stages, it is preferable that the leaching solution move in the opposite direction of loading the ore into the dumps and removing the ore from the dumps.

The above approach is based on the fact that slowly leaching gangue minerals such as hematite and sericite always remain. The dissolution rate of such minerals is usually proportional to the acid content. These minerals are present in large quantities, and the degree of their dissolution is low, so the amount of these minerals in the leaching cycle is not significantly reduced.

In a broad sense, the above object can be achieved by taking into account the nature of the acid-soluble minerals of waste rock and the amount of acid required to dissolve the copper mineral, and by changing the addition of acid in the agglomeration step accordingly. This makes it possible to maximize copper recovery in the drying step.

The amount of acid needed during agglomeration can be calculated from the reaction rate of each gangue mineral and the amount of acid needed to dissolve the contained copper minerals, i.e. the amount of added acid is optimally controlled to compensate for the consumption during the drying step.

The copper leached in the drying stage is fed to the PLS loop, and the acid recycled from copper extraction from the PLS in the SX stage is returned to the raffinate used for heap reflux during the leaching loop. This process can lead to an excess of acid in the raffinate, creating a positive acid balance. Excess acid can be removed from the dissolution of gangue minerals in the heap leach circuit by adapting the configuration of the heap leach flowsheet, in which leaching is carried out in several dumps or sections of the dump, using a countercurrent leach solution circulation method with respect to loading ore into dumps and removing ore from dumps .

Thus, the amount of acid added to the raffinate depends on the amount of acid added during agglomeration. Similarly, the amount of acid added during agglomeration depends on the amount of acid added to the raffinate to compensate for the expense of dissolving copper in the presence of gangue during the leaching cycle.

Dissolution of copper during agglomeration and during the subsequent soaking or drying period can be limited by the regulated amount of acid added during ore agglomeration, resulting in a higher consumption of acid in the reflux phase and avoiding the formation of excess acid during the leaching cycle after copper removal and acid recovery into the raffinate in the SX extraction step.

The first PLS harvested after the initial irrigation of the barns after drying the ore may have a high acid content (>10 g/l) and a high copper content (>20 g/l). In such a case, the first PLS can be used to agglomerate the ore of the following heaps before the drying phase is carried out. The values given are illustrative and non-limiting.

As an alternative to the processes described above, it is possible to reduce the amount of flushing at the end of the leach cycle to increase the amount of acid removed with the leached ore (leached residue). However, this is less preferred as the soluble losses of copper and sodium chloride increase. This entails an increase in lost profits and an increase in operating costs.

Brief description of graphic materials

Hereinafter, two embodiments of the present invention are described, respectively, with reference to FIGS. 2-7 and FIG. 8-12, which are accompanying graphic materials:

Fig. 2 is a generalized diagram of a reactor;

In FIG. 3 shows a typical heap leaching process;

In FIG. 4 is a schematic representation of a multi-stage heap leaching process carried out in accordance with the first embodiment of the present invention;

In FIG. 5 shows profiles of free acid versus time as PLS, ILS and raffinate parameters;

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In FIG. 6 is a diagram of a manufacturing process in accordance with the present invention with countercurrent circulation;

Fig. 7 is similar to FIG. 6, but with once-through circulation;

In FIG. 8 is a general flow diagram of a process carried out in accordance with a second embodiment of the present invention;

In FIG. 9 is a more detailed view of the process shown in FIG. 8;

Fig. 10 is a graphical representation of PLS concentration profiles showing acid, copper and iron concentrations as a function of PLS reflux rate at a reflux rate of 3 l/h.m ² ;

In FIG. 11 depicts an experimental reactor for determining the flow rate of acid on waste rock at a constant pH;

In FIG. 12 is a graphical representation of acid consumption per waste rock for a hypogene ore type as a function of pH versus time;

Fig. 13-27 refer to the flow charts used in the experimental work, as well as graphs representing the data obtained from the experimental work.

Description of Preferred Embodiments

Next, the first embodiment of the present invention will be described with reference to FIG. 2-7, and the second embodiment of the present invention is described below with reference to FIGS. 8-12.

Implementation Option 1: Multi-stage heap leaching operation

In FIG. 2 shows a typical reactor 10. From this reactor, the mass balance of reactant C can be determined. For simplicity, it is assumed that the volume flows in (12) and from (14) are equal. In addition, since real waste dumps are unexpectedly well mixed, it is assumed for simplicity that the reactant concentration in reactor 10 is the same as the reactor effluent (i.e., the homogeneously mixed reactor).

Below is the mass balance for reagent C: Intake = yield + losses in the reaction F. Ci = F.Co + rV;

where F is the flow rate; Ci is the inlet concentration; C _o is the outlet concentration; V is the volume of the reactor; and r is the rate of consumption of the acid, which in this case, as stated earlier, is proportional to the acid concentration C _o .

Thus, r = kC _o , where k is the reaction rate constant for the mineral in question.

This equation can be rewritten as follows: FC _i = FC _o + kC _o .V

The V/F ratio has a unit of time and is known as the mean residence time or τ.

This equation can be rewritten as follows:

Ci (1+kt) 1 1

The index k is constant and τ is constant for a fixed spoil layer at a fixed irrigation rate per unit area. This means that the percentage conversion (C _o /C _i ) of the incoming acid concentration is constant, regardless of the inlet concentration.

In FIG. 3 is a process flow diagram for a standard heap leach scheme 18 in which approximately 50% PLS is bypassed for CuSX optimization. In FIG. 3, one blade 20 is used. In contrast, in the process 24 shown in FIG. 4 use two blades 20A and 20B, respectively.

In both processes (FIGS. 3 and 4), the same amount of copper is transferred to the CuSX stage (30), as a result of this transfer of copper, the acid content increases by 7 g/l.

In the flow diagram shown in Fig. 3, the acid content of PLS 32 is y g/l. Then raffinate 34 contains y + 7 g/l of acid. Bypass stream 36 and raffinate 34 are combined again (38) to give an average acid concentration of (2y+7)/2. It is passed through dump 20, where the conversion is expected to be 50%. This conversion occurs regardless of acid concentration, as shown in Equation A.

In the equilibrium state, process 18 must be balanced, therefore: y (in PLS) = 50%. (2y+7)/2

Thus, y = 3.5 g/l of acid.

PLS in the example shown in FIG. 3 is divided in half. This is done for ease of comparison with the example shown in FIG. 4. If the PLS does not separate, then the acid concentration in the PLS remains the same. However, the increase in acid content in this raffinate is 3.5 g/L.

In the flow diagram shown in Fig. 4, the acid content of PLS (32) is also y g/l. This PLS is already % of the example shown in FIG. 3. This PLS (32) enters CuSX (30) and 7 g/l of acid goes into solution. Then the resulting solution (raffinate) 34 contains y + 7 g/l of acid. It is passed through a 20B dump where the conversion is expected to be 50%. Then it is passed through the dump 20A, where there is another 50% conversion.

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In the equilibrium state, the process (24) must be balanced, therefore: y (in PLS) = 50%.

50%. (y+7), and thus y = 2.3 g/l of acid.

If the process is extended to 3 steps, then y (in PLS) = 50%.50%.50% (y+10.5), and y=1.5 g/l acid.

Thus, by operating this process using two or more ILS stages 42, it is possible to significantly reduce the amount of acid in the PLS 46 entering the solvent copper extraction stage without adversely affecting the process. An example of the effectiveness of the proposed method can be seen from the change in free acid content in the oversalted section (from 100 to 150 g/l chloride) of the pilot plant shown in FIG. 5.

A multi-stage embodiment of the process according to the present invention may be configured to carry out the process in counter-current mode with the ore or co-current mode with it, as shown, respectively, in FIG. 6 and FIG. 7.

Counter-current operation has an advantage over once-through operation in that the final stage is to irrigate the dump with 20C raffinate (34). The raffinate has the lowest copper content and hence this minimizes soluble copper loss to leachate.

Implementation 2: Control and optimization of the amount of acid added during agglomeration

Acid control

In any of the schemes depicted in Fig. 3-7 (as examples), the acid in PLS (32) or raffinate (34) can be returned to the agglomerator to produce soluble copper, iron and acid. The acid is necessary for agglomeration of the ore to lower the pH during the drying/settling phase, as well as for the formation of an oxidizing agent for copper sulfide leaching (reaction 1). Iron and copper are needed as oxidizing agents. The effectiveness of oxidizing agents increases with concentration.

A typical simplified general flow diagram of the proposed process is shown in FIG. 8. It depicts a two step process in which the agglomeration solution is taken from the PLS.

In FIG. 9 is a more detailed view of the process shown in FIG. 8.

It has been found that when irrigating the ore layer after a period of standing and drying the ore, the first drainage solution or initial PLS contains exceptionally high concentrations of copper, iron and acid. An additional optimization of the proposed method is the use of the specified initial PLS or the first washing solution, first at the stage of ore agglomeration to increase the content of oxidants (soluble copper and iron) at the subsequent stage of drying / standing, and then in accordance with the acid control strategy, in the presence of excess acid, the amount fresh acid added to agglomerate the next ore may be reduced by this amount of acid. This helps to eliminate excess acid in the leaching phase.

This is demonstrated by the results presented in Fig. 10, for a 2m x 2m x 8m barn tested in a pilot plant used to treat hypogene ore with a copper content of 0.79% using 70% copper in the form of chalcopyrite with a crushed particle size such that about 80% of them pass through a 19 mm sieve .

The results shown in FIG. 10 demonstrate that the initial drainage solution or first PLS emerging from the Abmar ore layer (G10/2) contains exceptionally high concentrations of acid (80 g/l) (A), copper (55 g/l) (B) and iron (35 g/l) (C). This PLS solution can be effectively used to agglomerate a new batch of ore in place of the raffinate, with the advantage of using high concentrations of acid, copper and iron to accelerate the leaching of minerals, allowing the amount of fresh acid to be added to a given batch of ore to be reduced. The volume of PLS used depends on the required moisture content for agglomeration of the ore. A typical volume in this example is about 0.07 m ³ /t; which corresponds to the first 10 days of draining the PLS from the ore layer at an irrigation rate of 3 l/h.m ² . Iron in solution is oxidized to iron (III) in the process of ore agglomeration and further drying process under aeration conditions, which contributes to the leaching of copper minerals. High concentrations of copper accelerate the oxidation of iron(II) ions, increasing the oxidation potential of the solution, and increase the rate of leaching of copper minerals during drying. It can also be assumed that excess iron (III) in solution during this process will precipitate in the form of jarosites or in the form of meta-sideronatrite, as described in reactions 3, 4 and 5, respectively. This results in lower iron concentrations in the solution and maintains the iron content at the expected equilibrium concentration in the raffinate, typically, but not limited to, about 5 g/l in solutions with a chloride concentration of 150 g/l.

Gap Acid Rate (GAC) Determination and Acid Addition Control for Heap Leaching

The acid required for the chloride leach is added as fresh concentrated acid during ore agglomeration or as make-up acid for the raffinate liquor. Thus, when dealing with an ore agglomeration step, acid is provided by adding concentrated acid to the ore and/or only as acid in the raffinate solution. Addendum

- 8 041234 raffinate is adjusted to achieve the final moisture content specified for this agglomerated ore.

In the case of leaching the ore without agglomerating the ore, the acid is provided in the raffinate stream used to irrigate the dump.

The amount of acid needed is calculated from the acid consumption of the gangue minerals and from the amount of acid needed to leach the copper minerals. The rate of acid consumption during leaching of minerals depends on the concentration of acid in the solution and on temperature.

Experimental methods have been developed to determine the amount of acid used to leach gangue minerals, so for a particular type of ore it is possible to determine acid consumption per waste rock (GAC) as a function of solution pH or acid concentration and temperature. The results of these GAC tests are used to determine the model parameters that are used to calculate acid consumption for leaching waste rock minerals characterized by fast, medium and slow leaching. Thus, for a particular type of ore, the GAC can be determined and the addition of acid to the ore lump and to the raffinate can be adjusted to compensate for the consumption of acid without creating an excessive concentration of acid in the solution. The amount of acid needed to dissolve the copper can be calculated based on the grade of copper, the expected dissolution of copper determined in the tests, and taking into account the mineral composition of the ore and the reaction stoichiometry determined by the leaching reactions, for example, in accordance with reaction 1. The data obtained is then used to determining the amount of added acid required to agglomerate the ore and the concentration of acid that must be present in the raffinate to maintain mineral leaching, as well as to maximize copper dissolution and maintain optimum acid concentration in the PLS to ensure efficient copper extraction in the SX process.

Determination of the optimal amount of acid added during agglomeration

Copper-bearing ores contain many minerals. Sulfide minerals such as pyrite and chalcopyrite do not leach under low Eh conditions in the GAC test. Waste ore minerals leached by sulfuric acid typically include metal oxides and carbonates, as well as various potassium-iron(II)-magnesium-aluminum silicates. The reaction of an acid with a gangue mineral presumably results in stoichiometric dissolution of the mineral, for example:

Calcite: CaCO3 + 2H ⁺ -y Ca ²⁺ + CO ₂ + H ₂ O

Chlorite: Mg ₃ Fe ₂ AI ₂ Si30io (OH) ₈ + 16H ⁺ -y 3Mg ²⁺ + 2Fe ²⁺ + 2AI ³⁺ + 3SiO ₂ + 12H ₂ O

Feldspar: KAISi ₃ O ₈ + 4H ⁺ K ⁺ + Al ³⁺ + 3SiO ₂ + 2H ₂ O

Kaolinite: AI ₂ Si ₂ O ₅ (OH) ₄ + 6H ⁺ -y 2AI ³⁺ + 2SiO ₂ + 5H ₂ O

The chemical reaction between the acid and the gangue mineral in the ore releases ions into solution. If the solution is relatively dilute and precipitation of the sulfate mineral is negligible, then the amount of acid consumed in the reaction with gangue can be calculated from the sum of the total charge released in the form of ions, mathematically expressed as:

Δ[Η ⁺ ] = Δ[Κ ⁺ ] + 2 A[Fe ²⁺ ] + 2 A[Mg ²⁺ ] + 3 Δ[Al ³⁺ ] + · · - Δ[ΟΓ ] _(1} where Δ[Η ⁺ ] represents the spent amount of H ⁺ (protons) (mol/l), and Δ[Κ ⁺ ], Δ[Fe ²⁺ ], Δ[Α1 ³⁺ ], etc. represent changes in ion concentration (mol/ k) during the reaction.Sum all the cations released during the reaction.The calculation includes the change in the concentration of chloride A[C1'] to correct for the dissolution of NaCl in the ore, which releases Na ⁺ cations (and Cl ^- anions) into solution, but does not consume acid, as well as the possible leakage of KCl from the pH probe, which releases K ⁺ cations (and Cl ^- anions) into the solution, but does not consume acid.

The acid consumption is 49 Δ[Η ⁺ ] in r-H2SO4/l, or 49 Δ[Η ⁺ ] V/M in kf-H2SO ₄ /t ore, where V is the volume of the solution in the reactor and M is the mass of the ore ( kg). If the composition of the solution is known at different times during the reaction of the acid with the gangue, then the acid consumption can be calculated as a function of the reaction time.

The concentration of an acid in a solution during a reaction can be calculated from the total charge balance, mathematically expressed as

[H ₄ = 2[SOy] + m-[K ₄ -2[Fe ² 4-2[Mg ^! 4-3[A^ ⁽²⁾ where [H ⁺ ] is the concentration of protons or the total concentration of H (in the form of H ⁺ and HSO4) in solution (mol/l). The concentration of sulfate [SO4 ^2- ] is known from the concentration of sulfur [S] in solution.

The concentration of the acid in the solution is y 49 [H ⁺ ] in g-H ₂ SO ₄ /l. The acid concentration can be calculated as a function of the reaction time.

The pH of the solution in the reactor is kept constant during the reaction of the acid with the gangue by adding concentrated acid to the reactor. The amount of acid added to maintain the pH is recorded as a function of time and expressed as kg-H ₂ SO ₄ /t ore.

The amount of acid added to maintain a constant pH is not equal to the amount of acid used to react with the gangue. In fact, the amount of acid added to maintain a constant pH is usually greater than the amount of acid required to react with the gangue, and this is more true at low pH, which can be explained as follows.

- 9 041234

Acid H2SO4 dissociates in solution with the formation of acid chemical particles H ⁺ and HSO4 ^- , but only H ⁺ (due to its activity) affects the pH of the solution. Since ^HSO4 does not affect the pH, the amount of acid that needs to be added to maintain the pH is greater than the amount of acid used. The EQ3/6 geochemical model predicts that the HSO4 ^- content in H2SO4 solutions at pH1 and pH2 is approximately 40% and 20%, respectively. The release of cations into the solution during the reaction of the acid with the gangue is predicted to increase the percentage of HSO4 ^- . The results obtained suggest that at pH1 the amount of acid added to maintain a constant pH may be twice as much as the amount of acid required to react with the gangue.

As stated, the amount of acid added to maintain a constant pH is measured directly while the acid is reacting with the gangue. The amount of acid to be added can also be calculated from the composition of the solution, and such a calculation provides an opportunity to check the correctness of these data. The amount of acid added during a given reaction time is calculated as the sum of the amount of acid consumed in a given period, calculated by equation (1), and the difference between the acid concentrations in the solution at the end and at the beginning of this time period, which is calculated by equation (2 ), which can be expressed as follows:

Δ[Η ⁺ ] _ρΗ =2Δ[8Ο ₄ ² ] (3)

The amount of acid added to maintain a constant pH is 49 Δ[Η ⁺ ] _ρΗ n V/M in KΓ-H2SO4/Tru.

Equation (3) can be used to calculate the amount of acid added to the reactor from the change in sulfate concentration in the solution. This simple relationship is a consequence of the fact that the reaction between acid and gangue does not change the concentration of sulfate in the solution.

To create a model, the specific reaction rate for acid consumption on gangue minerals is expressed by the equation:

η _ _σ ^da GA , _σ d^Gal , _σ ^Ga oGal _Ί , '&Ga2 ι. '&Ga3 ι.

at at at (4) where RGa is the specific reaction rate (kg acid/t ore.s) expressed as the sum of three fast, medium and slow reaction elements, with gangue conversion a _Ga1 , a _Ga2 and a _Ga3 and final consumption of acid g _Ga1 , g _Ga2 and g _Ga3 (kg acid/t of ore), respectively. For each of the three types of acid consumption, the gangue conversion varies from zero to one, and the acid consumption varies from zero to g, depending on the reaction rate. The three types of acid flow presented can be considered as representative of different types of waste rock: for example, fast (calcite), medium (chlorite and biotite) and slow (kaolinite, feldspar and sericite).

The acid consumption per waste rock is determined by integrating the reaction rate represented by equation (4) over the elapsed reaction time:

GAC ⁼ g _Gal a _Gal + g _Ga2 a _Ga2 + g _Ga3 a _Ga 3 (3) where GAC is the acid consumption in kg acid/t ore.

For each of the three types of acid consumption, the rate of gangue transformation during the reaction is expressed by the Arrhenius equation in the following form:

where k _Ga is the rate constant ((l/mol) ^pN /s), EGa is the activation energy (cal/mol), dp is the ore particle size (mm), d0 is the characteristic ore particle size (taken as 10 mm), [H ⁺ ] is the concentration of protons (mol/l), T is the temperature (K), R is the universal gas constant (1.986 cal/mol.K) and np, n _H and n _a are power indicators for particle size, acid concentration and gangue conversion, respectively. As shown in equation (5), the waste rock acid consumption rate is expressed in units of proton concentration, not proton activity or pH.

The concentration of cations released into solution during the reaction of the acid with gangue is expressed as follows:

+ g _Go 2 ^a Ga2%2 ⁺ gGa3 ^a Ga3^C3)^^(^d^) where [C] is the cation concentration (mol/L) and Y _C1 , Y _C2 and Y _C3 are the cation release factors for each of the three acid-consuming material classes, expressed as the number of moles of cations released per mole of acid consumed (mol C/mol acid). For example, if calcite is the only rapidly reacting gangue material, then the release factor for Ca ²⁺ cations is YCa1 = 1. Similarly, if chlorite is the only material that reacts at a medium rate, then the release factors for Mg ²⁺ , Fe ²⁺ cations and Al ³⁺ are YMg2 = 0.375, Y _Fe2 = 0.25, and Y _Al2 = 0.25, respectively.

- 10 041234

The total mass of the element released during the interaction between the acid and the gangue materials must be less than the mass of this element in the ore:

%L1 +&L +g _M r _o )W _c /98<G _{c (8)} where Wc is the atomic weight of the element and Gc is the content (kg C/t of ore) of the element in the ore.

The cation release factors must comply with the charge balance condition: % + 2K _Fe(n) + 2T _Mg + ZUr _e(N) + ZU _A1 + · · · = 2 (9)

The model parameters in equations (4) - (7), namely k _Ga , E _Ga , g _Ga , Y, n _p , n _H and n _a , are determined by fitting the model of the calculated acid flow and solution composition to the data measured in tests GAC.

Experimental determination of GAC is carried out using the setup 100 shown in FIG. 11. Said plant includes a 5.5 liter reaction vessel 102 equipped with a 4.5 liter top outlet 104, into which 2750 g of ore sample 106 is charged (sample size is increased for larger fractions). The reaction vessel 102 is connected by a pipe to a nearby reactor 110 with a volume of 3 liters. The main part of the leaching agent (5300 ml) is introduced into this connected system and pumped by pump 116 from the adjacent reactor 110 into the reaction vessel 102 at a constant rate, while the return flow moves to the adjacent reactor 110 under the action of gravity.

After loading 5300 ml of the leaching agent into the system, said leaching agent is brought to a certain temperature using heaters 118 located under the reactors 102, 110. The pH and the final volume are adjusted to the desired initial values by adding 200 ml of iron-containing acid solution 124 to obtain a final volume of 5500 ml.

A reducing atmosphere is maintained above the surface of the reaction mixture by removing air from reactors 102 and 110 and introducing a stream of nitrogen 126 into reactors 102, 110. The reactors are hermetically sealed, allowing direct volumetric displacement of nitrogen from reactors 102, 110. Samples of the solution are taken from the adjacent reactor 110 at certain time intervals and analyzed for the content of certain elements using computer 130. Computer 130 records the pH and redox potential of the solution as a function of time.

In FIG. 12 is a graphical representation of the results obtained on the experimental setup 100 shown in FIG. 11. Acid consumption versus time is shown for pH1 (C), pH1.5 (B) and pH2 (A). It can be seen that the acid consumption largely depends on the pH of the solution, i.e. acid concentration in solution.

The reactor 100 can also be used to determine the effect of temperature and the effect of different particle sizes in ore samples on the rate of acid consumption.

Representative results obtained using the pilot reactor for the GAC method are shown in FIG. 12. The results obtained indicate a significant effect of the pH of the solution or the concentration of acid in the solution on the consumption of acid. It has been shown that it is possible to test the effect of temperature and various ore particle sizes to determine the acid consumption rate as a function of particle size and temperature.

The following description with reference to FIG. 13-23 relate to the tests carried out on the various aspects of the present invention described above.

Example 1 Description of the technical evaluation of an integrated pilot plant for carrying out the method according to the present invention.

An integrated pilot plant was assembled containing 9 bins and a solvent extraction unit to reproduce the industrial application of the method according to the present invention on a pilot scale. The described method is designed to process various low-grade chalcopyrite ores with dynamic race track style dump leaching operation. The race track style blade contains several sectors. New sectors are laid in the form of a stack, and the oldest, most leached sectors are removed from the bookmark. The bookmark is reused.

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

Three separate operating phases 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. 1A.

- 11 041234

Table 1A. Brief description of the work steps for the technical evaluation of a pilot plant

Parameter Stage 1 Stage II Stage III Ore particle size, passing 80% particles (mm) 19.05 19.05 19.05 Copper content (%) 0.49-0.75 0.38-0.79 0.3-0.6 Content of chalcocite CSR* (%) >10 10-20 10-20 Covellite content CSR* (%) 10-20 < 10 < 10 Chalcopyrite content CSR* (%) 40-75 65 -75 40-75 Pyrite(%) 2-6 2-4 3-7 Chlorite(%) 0.1 - 1 0-1 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 type of ORE 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 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. The empty barn was then prepared to receive the fresh, lumpy ore.

Phase I and Phase II included the start of operation and the collection of baseline data, as well as the construction of additional barns for closed-loop operation in Phase III.

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 new 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 process of the present invention described above.

The pilot plant was operated as a closed system, with carefully controlled outlets and inlets to simulate industrial heap leaching operation. That is, the inlets and outlets were adjusted so that the amount of fresh water added exactly corresponded to the loss of water to evaporation, the loss in the form of moisture in the leached residue (or in the residue after leaching), and the loss during the replacement of water in the process wash solution (if necessary to reduce impurity content). The technological scheme of the pilot plant is shown in Fig. 13. The main components and operations of the pilot plant are summarized below:

Refinery tank 1:

Storing 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 High Cu: 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. High copper content in acid agglomerated ore can improve

- 12 041234 dissolution of copper during the drying of the ore at the initial stage of maturation. The agglomerated ore is fed into barns (3A-I) as needed for loading.

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.

Extraction with 7 solvent (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, as described in the above description. 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.

The operating parameters for stage I, stage II and stage III are summarized in Table 1. 2A.

Table 2A. Pilot Plant Operating Parameters for Phase I, Phase II and Phase III Operations

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 a 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 period aging or drying days 45 45 Ore layer 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 200500 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. 14. Typical values for the balance of ore, acid and water are presented in table. 3.

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Table 3. Generalized mass balance of a pilot plant showing inflows and outflows for operation in stage II

Mass balance of pilot plant in stage II Incoming outgoing Input (kg/t) Output (kg/t) Ore and copper Total load of dry ore (t) - 9 barns 348.6 - Total copper load (0.5% CuT) (t) 1.7 5.00 Total copper recovered at 60% average recovery (t) 1.0 3.00 Total loss of copper in the balance (t) 0.7 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) 4183.2 12.0 Loss of acid in residual moisture at a moisture content of 10.5% (kg) 18.0 0.05 Total acid consumption (for waste rock and dissolution of metals) (kg) 6274.8 18.00 Amount of acid released at EW (kg) 1708.1 Amount of make-up acid added (kg) 401.4 1.15 Water balance Water for washing the organic phase at SX (m3) 4.3 The amount of flushing water at a flow rate of 0.1 m3/t (m3) 34.9 Moisture content in the residue at an average value of 10.5% (m3) 36.6 Water loss due to evaporation at 5.4 l/day (m3) 2.6 Released water (m3) zero

The configuration of the technological scheme of the pilot plant of the leaching zone is shown in Fig. 15. This configuration is an acceptable flowsheet because the ore used in the operation of the pilot plant was mainly chalcopyrite and had a low initial copper content. A key aspect of the flowsheet presented is the formation of a low acid raffinate during solvent extraction. The resulting low acid raffinate is distributed among all sectors (in this case barns). The low acid content minimizes waste rock leaching.

If the ore is of a higher grade, and especially if the copper content is primarily minor copper minerals such as covellite, chalcocite, or bornite, then a flow chart can be used to increase the acid consumption of the waste rock during the leaching process, as previously described. . Such a technological scheme is shown in Fig. 16.

The variant shown in Fig. 16 can be operated in the manner shown, or with solvent extraction carried out using the solution from the second 4 pits, and not from the first 4 pits, as shown in the drawing. The number of stages can be increased from two to three, or to any value that may be expected to be necessary.

Example 2. Acid balance

The above description indicates the need to maintain the balance of the acid throughout the process. Acid is required as a reactant in the agglomeration step to ensure copper dissolution in the drying phase. The acid is also needed as a reactant in the leach phase to dissolve the copper. The acid is formed during the electrolytic separation of dissolved copper with the formation of metallic copper. The resulting acid is returned to the process through the solvent extraction stage, and all of the acid enters the process in the leaching stage.

In the leaching phase of the described process, the amount of acid formed during EW must be compensated for by the amount of acid consumed in the reaction with gangue and copper sulfide minerals, as well as by the amount that exits with moisture after washing the leached ore.

In ideal operation, the amount of acid required for leaching and the amount of acid generated during EW are the same, and the leaching step in such operation is acid neutral. In this case, the consumption and intake of acid are balanced, which ensures that a balance of acid is achieved. It is also acceptable, if necessary, to add some acid in the leaching phase. The situation in which more acid is returned from the EW process during the leaching step than is used to leach the gangue and copper sulfide mineral. This is due to the fact that the operation is carried out in a closed circuit with a final washing stage, so excess acid will accumulate. Higher concentrations of acid in process solutions will adversely affect the solvent extraction process, as it requires a low acid content in the feed to facilitate the transfer of copper to the organic phase (proton/copper equilibrium).

In table. Figure 3 shows the balance between the barns for work at stage II from example 1. First,

- 14 041234 it can be seen that the amount of acid leaving the leaching process with moisture after washing the leached ore is extremely low (0.05 kg/t _ore ) and therefore can be neglected in the following description. Second, it can be seen that the leaching step is, on average, acid negative. During the leaching, a small amount of acid was added to adjust the pH (1.15 kg/t _ore ). This is ideal, and if the specified combination of ore fed into the process and treated in this way is observed, there will be no problems from an acid point of view.

In table. 4 shows acid consumption and copper dissolution for each barn.

Table 4. Data on copper and acid for each pit at stage II

Abmar code G5/2 G9/2 G10/2 G3/2 G6/2 G4/2 G1/2 G2/2 G8/2 Acid used in drying (kg/t ore) 10.19 9.07 12.84 12.62 11.84 11.53 12.25 12.26 11.89 Total acid (kg/t ore) 20.73 16.41 19.47 20.26 16.38 16.43 21.63 16.09 18.66 Drying leached copper eleven% 6% 7% 15% 8% 16% 18% 3% 9% Total leached copper 60% 55% 55% 75% 61% 65% 69% 56% 61% Initial copper content 0.62% 0.66% 0.79% 0.45% 0.37% 0.37% 0.59% 0.36% 0.43% Mass of acid used in drying (kg/t of ore) 10.19 9.07 12.84 12.62 11.84 11.53 12.25 12.26 11.89 Weight of acid used in leaching (kg/t of ore) 10.55 7.34 6.63 7.64 4.54 4.90 9.37 3.83 6.77 Weight of Cu leached during drying (kg/t of ore) 0.71 0.40 0.58 0.66 0.31 0.58 1.06 0.09 0.39 Mass of C leached in the cycle 2.99 3.26 3.74 2.72 1.96 1.82 3.02 1.94 2.22 leaching (kg/t of ore) Acid generated by EW during the leaching cycle (kg/t ore) 5.71 5.64 6.66 5.22 3.50 3.71 6.29 3.14 4.02 Is the leaching step acid positive? No No Yes No No No No No No Percentage of Cu leached upon drying relative to the total amount leached 19% eleven% 13% 20% 13% 24% 26% 4% 15%

The data presented in table. 4 show the amount of leached copper as well as the amount of acid used in the drying phase and the leaching phase for each barn. The table also shows the amount of acid generated and returned to the process at EW.

The data presented indicate that, in general, almost all pits are negative for acid during leaching, with the exception of the G10/2 pit. When commercially operating this particular ore in the manner indicated and over an extended period of time, acid will accumulate in the circuit. This operation would require either reducing the amount of acid fed to the agglomeration step to bring the loop to acid neutrality, or using the flow diagram shown in FIG. 16 instead of the flow chart used (FIG. 15) and this will increase the acid consumption per waste rock and return the process to acid neutral or acid deficient. The latter option is preferred as it allows the copper leaching performance to be maintained during the drying step.

There are two main differences between the ore in G10/2 and other ores. The first is the original content. This ore has a higher than average copper content and this, combined with acceptable dissolution, means that more acid than average will be formed in the EW stage. Secondly, the amount of acid used for waste rock and copper minerals during the leaching stage was only 6.6 kg/t. This is less than the amount likely to be needed to leach the 3.7 kg/t of copper formed over the period indicated. This suggests that the gangue minerals are of low acid reactivity and that acid forming compounds such as jarosite may have precipitated.

This example shows that regulation may be required to prevent the described process from becoming acidic, and that understanding the leaching behavior of gangue minerals is key to predicting acid consumption. The considerations described then need to be linked to the appropriate flowsheet configuration for a given leaching period to achieve the optimal value. The optimum value is usually the maximum amount of copper that can be leached since the value of the copper is substantially higher than the value of the acid.

Example 3 - Acid balance - interaction between the drying step and the final degree of copper dissolution

If additional acid is added to the step II agglomeration step, and if said acid provides a proportional increase in chalcopyrite leaching, it can be estimated at what point in time the acid balance for the step II barns will become acid positive.

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Table 5. Data on the content of copper and acid in individual pits at stage II based on a 30% increase in acid consumption at the stage of drying in terms of copper

Abmar code G5/2 G9/2 G10/2 G3/2 G6/2 G4/2 G1/2 G2/2 G8/2 Correction for acid consumption at the stage of drying thirty% thirty% thirty% thirty% thirty% thirty% thirty% thirty% thirty% Acid used in drying (kg/t ore) 13.24 11.79 16.69 16.41 15.40 14.99 15.93 15.94 15.45 Total acid (kg/t ore) 20.73 16.41 19.47 20.26 16.38 16.43 21.63 16.09 18.66 Drying leached copper 27% 19% 23% 42% 39% 46% 38% 36% 36% Total leached copper 60% 55% 55% 75% 61% 65% 69% 56% 61% Initial copper content 0.62% 0.66% 0.79% 0.45% 0.37% 0.37% 0.59% 0.36% 0.43% Mass of acid used in drying (kg/t of ore) 13.24 11.79 16.69 16.41 15.40 14.99 15.93 15.94 15.45 Weight of acid used in leaching (kg/t of ore) 7.49 4.2 2.78 3.85 0.98 1.44 5.70 0.15 3.21 Weight of Cu leached during drying (kg/t of ore) 1.70 1.28 1.82 1.89 1.46 1.70 2.25 1.28 1.55 Mass of Cu leached in the leaching cycle (kg/t of ore) 2.00 2.38 2.49 1.49 0.81 0.70 1.83 0.75 1.06 Acid generated by EW during the leaching cycle (kg/t ore) 5.71 5.64 6.66 5.22 3.50 3.71 6.29 3.14 4.02 Is the leaching step acid positive? No Yes Yes Yes Yes Yes Yes Yes Yes Percentage of Cu leached upon drying relative to the total 46% 35% 42% 56% 64% 71% 55% 63% 59% leached amount

In table. 5 shows that if the amount of acid consumed in the drying step is increased by 30%, then the total result will be such that 8 out of 9 barns will become acid positive in the leaching step. This will lead to the need to change the technological scheme to a multi-stage option. It is possible that in addition to changing the configuration of the technological scheme, it will be necessary to slightly reduce the amount of acid added during the drying stage in order to prevent this phenomenon and to limit the dissolution of copper.

Presented ores have a fairly high content of chalcopyrite. Chalcopyrite is known to leach more slowly than other copper sulfide minerals such as chalcocite and covellite. It can be assumed that at the end of the leaching stage, mainly the leaching of chalcopyrite occurs. Chalcopyrite leaching requires 4 protons (H ⁺ ) for each Cu ²⁺ released into solution. However, at EW, only two protons (H ⁺ ) are produced for each Cu ²⁺ deposited as copper metal (Cu). This means that if the final degree of copper leaching occurring in the leaching step is increased for these ores, then there is a point at which the process becomes acid negative again in the leaching, regardless of the more efficient copper dissolution achieved in the drying step at low grades. in the ore of an acid-consuming rock and when leaching is carried out in accordance with the flow diagram shown in FIG. 15.

In table. 6 shows the results of leaching several columns from the same ores. These columns performed better than barns, partly due to process conditions and partly due to more efficient solution distribution in the columns. The detailed reasons for this change in characteristics are irrelevant to the description presented. Significantly, these columns show that despite the higher degree of copper dissolution during the drying stage, they are all negative in acid during the leaching stage, including the ore used in the G10/2 barn (see column MK104), due to additional leaching reached during the leaching step and subsequent acid consumption as described in the previous paragraph.

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Table 6. Data on copper and acid content in individual leach columns for leaching columns from the same ores that were used in the stage II barn operation program

Column code MK11 7 mcu 2 ICJ 4 MK11 5 MK11 3 MK11 4 ICJ 4 MK11 9 MK12 0 Acid used in drying (kg/t ore) 10.3 10.6 14.5 13.1 8.4 11.6 14.5 12.5 12.5 Total acid (kg/t ore) 26.2 23.8 25.9 24.9 18.7 20.9 25.9 26.7 18.9 Drying leached copper 21% 26% 27% 29% 10% 12% 27% 36% 27% Total leached copper 77% 76% 68% 79% 51% 46% 68% 85% 74% Initial copper content 0.62% 0.66% 0.79% 0.45% 0.68% 0.81% 0.37% 0.59% 0.36% Mass of acid used in drying (kg/t of ore) 10.34 10.64 14.52 13.08 8.41 11.64 14.52 12.53 12.50 Weight of acid used in leaching (kg/t of ore) 15.85 13.18 11.40 11.87 10.27 9.23 11.40 14.22 6.36 Weight of Cu leached during drying (kg/t of ore) 1.29 1.74 2.11 1.31 0.69 0.94 0.99 2.11 0.98 Mass of Cu leached in the leaching cycle (kg/t of ore) 3.49 3.26 3.30 2.22 2.79 2.82 1.55 2.93 1.67 Acid generated by EW during the leaching cycle (kg/t ore) 7.37 7.72 8.35 5.46 5.38 5.79 3.91 7.78 4.09 Is the leaching step acid positive? No No No No No No No No No Percentage of Cu leached upon drying relative to the total amount leached 27% 35% 39% 37% 20% 25% 39% 42% 37%

Example 4: Extrapolation of data for barns in stage II

It is possible to extrapolate from the stage II barns to obtain information on the ideal separation between the copper dissolution in the drying stage and the leaching stage of these ores, providing maximum efficiency. The maximum target value for the dissolution of copper in these types of ores is ~85%. By achieving this value and maintaining the same requirements for acid consumption per waste rock, it is possible to estimate the total maximum percentage of copper leached during the drying stage, but while maintaining the neutrality of the process for acid during the leaching stage. In FIG. 17 shows that there is a close relationship between the dissolution of copper during the drying step and the initial grade of copper in the ore. At a low initial grade, all of the copper may be leached out in the drying step and the leaching step will remain acid neutral. This makes sense since the amount of EW formed is directly proportional to the mass of leached copper. Thus, with a low initial copper grade, only a small amount of acid is produced in the EW process, and this acid is consumed by the residual acid-soluble gangue minerals in the leaching step. As the initial grade increases, there is insufficient acid flow to the waste rock to compensate for the acid produced by the EW, and therefore more chalcopyrite needs to be leached in the leaching step.

Although the trend line is smoothed relative to the data obtained, the data points are scattered due to the variability in the composition of the gangue minerals in different ore samples, and therefore the same happens with the point of convergence of the acid balance. This is important because this fact highlights that knowledge and understanding of the leaching behavior of gangue minerals is necessary to optimize the process.

The presented graph will look different if the main copper minerals are bornite, chalcocite, or covellite. For these minerals, the number of protons (H ⁺ ) required for leaching is 2.4, 2 and 2, respectively. Thus, there is little or no difference between the amount of protons needed to leach copper and the amount of protons produced by EW. Thus, the proposed method will rely more on a constant flow of acid to the waste rock during the leaching step so that it does not become acid positive. Alternatively, a weaker copper leaching may be required during the drying step.

The trend line in the presented chart shifts upwards when using a multi-stage leaching process.

Example 5 Estimation of acid consumption per waste rock

The results in FIG. 18 show the total acid consumption during the operation of barns G5 (B) and G7 (A), which process two different types of ore, during the operation of a pilot plant in stage III. The mineral composition of the two types of ore loaded into these pits is presented in Table. 7. The results obtained demonstrate a higher content of biotite and chlorite in the type C ore (G5 pit in stage III) compared to the type D ore (G7 pit in stage III). Biotite and chlorite have been found to be major contributors to GAC as they typically leach at a faster rate than other silicate minerals.

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Table 7 Mineral Composition of Type C and D Ore Reflecting Mineral Composition of Waste Rock

Mineral Type C ore with low acid consumption Type D ore with very low acid consumption chalcocyte 0.13 0.08 covellite 0.08 0.06 Chalcopyrite 1.07 1.15 bornite 0.00 0.00 Other Si Minerals 0.13 0.11 Pyrite 6.25 2.93 Other sulfides 0.05 0.05 Fe oxides 0.53 0.15 Quartz 26.56 22.16 feldspars 24.29 28.29 Muscovite/sericite 18.51 19.85 Kaolinite/clays 10.36 23.38 Biotite 8.26 0.03 Chlorite 2.07 0.08 Frosite 0.21 0.00 Other sulfates 0.47 0.62 Other 1.01 1.09 Total 1.01 1.09

The total acid consumption includes the amount of acid to leach copper minerals and the amount of acid to leach gangue minerals (GAC). The consumption of acid for the leaching of copper minerals is shown in Fig. 19 along with a mass of leached copper. The copper recovery achieved over the same period of time is shown for reference in FIG. 20.

The acid consumption per waste rock (GAC) for a particular ore type and copper grade can be estimated by the method of the present invention by subtracting the acid consumption for copper recovery from the total acid consumption. The result obtained gives a net acid consumption (NAC), which in the case of the method according to the present invention is equal to the true consumption of GAC. Estimated acid consumption per waste rock for pits G5 and G7 in Stage III is shown in FIG. 21.

In the recovery step S1 of the pilot plant SX (shown in FIG. 13), copper is recovered from the loaded organic phase using a high acid spent electrolyte. Copper passes into a purified aqueous electrolyte solution, which is fed to the extraction of copper by electrolytic deposition. The copper in the organic phase is displaced with acid, and the resulting acid-rich separated organic phase is returned to the copper extraction stage E1, in which copper is extracted from the PLS as shown in FIG. 13. Thus, the SX process ensures that the acid is returned to the raffinate solution after copper is extracted from the PLS. The acid in the raffinate is used to leach additional copper minerals and to offset the GAC consumption during the leaching cycle.

Initially, the amount of acid added in the G7 ore agglomeration in Stage III (very low GAC type D ore) was 11.3 kg/t. In the case of pit G5 in stage III (low GAC type C ore), the amount of added acid in ore agglomeration was 10.6 kg/t. In the case of pit G7 in stage III, the initial concentration of acid in the solution was higher than in pit G5 in stage III due to the lower degree of GAC for type D ore during the drying stage. The higher initial concentration of acid for irrigating the ore in pit G7 in stage III causes an increase in the acid consumption rate. However, once the waste rock silicate minerals are depleted, the rate of acid consumption decreases. In the case of pit G5 in stage III, the initial acid concentration was relatively low and the initial flow rate was lower than in pit G7 in stage III, despite the higher GAC value for type C ore in pit G5 in stage III. However, when irrigated with acid raffinate, the rate of acid to waste rock consumption, GAC, was markedly higher for pit 5 in stage III, so that by the end of the leach at day 250, the GAC value for pit G7 was 11.8 kg/t compared to 12.9 kg/t for barn G5 at stage III. These trends are reflected in the respective PLS acid concentration profiles for both pits shown in FIG. 22 and the corresponding PLS pH profiles shown in FIG. 23.

The initial concentration of acid in the solution during irrigation depends on the amount of acid added during ore agglomeration and on the consumption of GAC in the ore. Too little addition limits the initial free acid available to leach copper minerals. Too much addition causes an excess of acid in the PLS during the leaching step, which adversely affects copper recovery during the solvent extraction process. Various variants of acid consumption and acid concentration in the solution presented in Figs. 21 and 22 are valid operating parameters for carrying out the method according to the present invention. The results presented demonstrate how little the effect of different ore GAC values and the amount of added acid on the acid concentration in PLS, and also show the importance of understanding the GAC in different types of leached ores and the amount of acid required to leach copper minerals.

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Example 6 Effect of Chloride Concentration on PLS Solution pH at Different Acid Concentrations

The effect of chloride concentration on solution pH as a function of acidity is shown in the results shown in FIG. 24. The results presented demonstrate that at a low chloride concentration (<5 g/l chloride), the free acid concentration is 4 g/l at a solution pH of 1.5. The corresponding pH at high chloride concentration (150 g/l) is approximately pH 0.8. The solvent extraction efficiency depends on the pH of the PLS. The ideal pH of PLS is approximately pH 1.5 to achieve maximum copper extraction efficiency in the SX stage. The efficiency of copper extraction deteriorates when the pH of the PLS is less than 1. Thus, it is understood that in PLS with a high content of chloride, the free acidity should be as low as possible, less than 10 g/l acid or preferably less than 7 g/l, more preferably less than 2 g/l. l and ideally 1 g/l.

Example 7. The effect of the amount of acid added during ore agglomeration on the extraction of copper at the stage of drying

A series of 1 m columns were tested and it was shown that, for example, for a certain type of ore, for which the dissolution of copper minerals in the drying stage is high and limited by the amount of added acid, increasing the amount of added acid for ore agglomeration promotes copper recovery. The results are shown in FIG. 25 where a fixed amount of salt added was used, 10 kg/t ore, and the amount of sulfuric acid added was varied from 10 to 21 kg/t ore. The stage of drying after agglomeration of the ore was carried out for 50 days. The initial copper dissolution, indicative of the amount of copper leached from the drying step, increased from about 30% to 80% as the amount of added acid increased. The overall copper recovery increased from 78 to 98% over a 70 day total leach period, with 20 days of raffinate reflux.

Example 8 Example of copper dissolution curves for different types of ores

The dissolution of copper by the high chloride (150 g/l chloride) heap leach method described herein has been demonstrated for various types of ores using the results presented in FIG. 26. The results are obtained for tests of 6 m high columns carried out by the leaching method with a high chloride content (150 g/l) described in the essence of the present invention. The amount of acid added to the ore during ore agglomeration, the total consumption of acid and NAC (total acid consumption minus the amount of acid used to leach copper, this acid is returned from SX to the raffinate after copper extraction) are presented in Table 8. The results obtained show a high degree recovery of copper, which can be achieved by the method according to the present invention at the stage of drying, if the amount of added acid is not limiting. The results also show the significance of the copper recovery in the drying step relative to the overall copper recovery achieved. The results obtained refer to low-grade ores with a high content of chalcopyrite (from 0.3 to 0.6% Cu) with a copper content in the raw material (CSR) in the form of chalcopyrite of 40-75%.

Table 8. Amount of added acid, acid consumption and copper recovery for various ores in a test of 6 m columns

Trial Amount of acid added during ore agglomeration (kg/t) Total acid consumption (kg/t) Net Acid Consumption (NAC) (kg/t) Extraction of copper at the stage of drying (%) Total copper recovery (%) A 11.3 23.0 16.2 50 85 IN 10.9 22.5 16.8 45 67 WITH 10.3 14.4 11.9 40 62 D 10.6 22.0 17.5 25 53

Example 9 - Moisture at agglomeration - influence on drying characteristics

The previous examples have discussed the optimization of the amount of acid added during the drying step and its implications for flowsheet optimization. It is important to note that it has been found that the way the acid is added also influences the drying characteristics.

Technological solutions are returned to the agglomeration of the ore for its wetting. The liquid volume in the ore bed is the volume of solution in the reactor in which the oxidizer is formed to leach the minerals. The amount of oxidizing agent formed per unit of time depends on the volume of the solution. The oxidizing agent is responsible for the oxidation of copper sulfide minerals, which leads to copper solubilization. If the amount of liquid available is limited, then insufficient oxidizing agent may be formed and this will limit the degree of leaching of the mineral. This effect is analogous to limiting the amount of acid added to the ore, whereby the amount of acid available is insufficient to meet the cost of dissolving metals, given that the acid is consumed by both the ore and the gangue minerals. However, the return of too much process solution for agglomeration can adversely affect

Claims (12)

the strength of the agglomerate and, consequently, the permeability of the heap for gas and liquid flows. In FIG. 27 shows the difference in leaching performance for a column fed with a limited amount of moisture (B) during the agglomeration step compared to a column fed with an ideal amount of moisture (A) during the agglomeration phase. Based on the presented trends, it becomes clear how insufficient moisture limits the dissolution of copper. The importance of this example is that although it has been shown that it is possible to change the copper leaching in the drying step by reducing the amount of acid or moisture (as in this case), the leaching phase is not as effective in terms of copper leaching and cannot catch up. Therefore, to optimize the proposed process, the optimal solution to ensure the maximum degree of copper recovery is to modify the technological scheme to counteract the excess amount of acid during leaching. Limiting the drying step should only be considered as a last resort. CLAIM 1. A method for controlling the acid balance in a heap leaching of copper ores using high chloride solutions, conducted at room temperature, where the chloride concentration is from 100 to 180 g/l, ensuring that the dissolution of copper in the agglomeration step is not limited by acid for maximizing the dissolution of copper in the drying step and thus increasing the overall recovery of copper, where the specified method includes: a) a agglomeration stage, which combines acid and process solutions with ore prior to stacking to form a pile; b) an aerated but non-irrigated drying step to leach out a portion of the copper in the ore in said heap, wherein the copper dissolution in the drying step is at least 30%; c) a flash leach step that leaches the rest of the copper minerals and recovers copper from the leach product in a solvent extraction step followed by an electrowinning step, wherein the total copper dissolution in the drying step and the spray leach step is greater than 40% and said heap contains a plurality of heap sections, in which the excess acid content in the irrigation leaching step is reduced by passing the leaching solution through the heap sections in the opposite direction relative to the loading of ore in the heap section and relative to the removal of said ore from the heap sections, which provides an increase in the consumption of acid for the dissolution of minerals waste rock is proportional to the number of irrigated sections of the heap, and the acid concentration in the productive leaching solution that enters the solvent extraction stage is less than 10 g/l to ensure efficient recovery of copper from the ductive leaching solution in the solvent extraction step. 2. The method according to claim 1, characterized in that the acid concentration in the productive leaching solution is less than 7 g/l. 3. The method according to claim 1, characterized in that the acid concentration in the productive leaching solution is less than 2 g/l. 4. Process according to claim 1, characterized in that the raffinate formed during the solvent extraction step has a copper content of less than 1 g/l. 5. Method according to claim 4, characterized in that said copper content is less than 0.5 g/l. 6. The method according to claim 1, characterized in that after completion of the stage of leaching by irrigation, washing of the heap is carried out. 7. The method of claim 1, wherein the amount of acid added to the process is calculated from the amount of acid consumed by the gangue minerals and from the amount of acid needed to leach the copper minerals. 8. The method according to claim 1, characterized in that the optimal amount of acid added to the process is determined by the content of acid-soluble gangue minerals and copper minerals in the ore. 9. The method according to claim 7, characterized in that the acid necessary for optimal leaching of copper is added to the ore at the stage of agglomeration. 10. The method according to claim 7, characterized in that the addition of acid to the raffinate solution used to irrigate the heap is carried out to compensate for the consumption of acid for leaching minerals in the stage of leaching by irrigation. 11. The method of claim 8, wherein the first leach solution collected during the initial irrigation of the heap after the drying step is added to the ore during the agglomeration step to utilize the acid in the first leach solution in a subsequent drying step. 12. The method according to claim 8, characterized in that the amount of acid added to the ore during the agglomeration step is determined by the consumption for dissolving copper minerals and gangue minerals during the drying step. -