MXPA97009727A - Hydrometalurgical extraction of metal assisted porclor - Google Patents

Abstract

The present invention relates to a process for the extraction of copper from a copper sulfide mineral or concentrate, which comprises the steps of: subjecting the ore or concentrate, together with a source of bisulfate or sulphate ions, to oxidation under pressure, at a temperature of about 115 ° C to about 160 ° C, in the presence of oxygen and of a solution containing halide ions, to obtain a resulting pressure oxidation paste; a liquid / solid separation step, to obtain a resulting pressure oxidation solution and a solid residue, and recover the copper from the oxidation solution under pressure, or the solid residue, characterized in that the pressure oxidation solution is recycled until the oxidation under pressure and the concentration of copper in the oxidation solution under pressure that is recycled, is maintained at a previously determined value

Description

HYDROMETALURGICAL EXTRACTION OF METHOD ASSISTED BY CHLORIDE FIELD OF THE INVENTION This invention relates to the hydrometallurgical treatment of metal ores or concentrates. In particular, it relates to the extraction of metals from minerals in the presence of halogen ions, such as chloride ions.

BACKGROUND OF THE INVENTION The hydrometallurgical treatment of copper sulphide minerals, such as chalcopyrite (CuFeS2), is problematic, due to the severe conditions required in a pressure oxidation step, for the effective leaching of copper from these minerals. , results in the oxidation of the sulphide in the sulphate mineral, resulting in the generation of large amounts of acid that requires costly neutralization. Attempts have been made to make the sulfide concentrate leach under relatively lighter conditions under which the sulfide is oxidized only in elemental sulfur, and not all the way to the sulfate. These attempts include the pretreatment of the concentrate before the oxidation step under pressure, to make the sulfide concentrate easy to leach, and the leaching of the concentrate in the presence of chloride ions, as described in the US Pat. United States of America Number 4,039,406. In this process, the copper values in the concentrate are transformed into a solid basic copper sulfate from which then the copper values must be subsequently recovered, as described in U.S. Patent Number 4,338,168. In the process described in U.S. Patent No. 4,039,406, a significant amount (20 to 30 percent) of the sulphide of the ore or concentrate is still oxidized in sulfate, resulting in an increased oxygen demand during leaching. under pressure, and the generation of sulfuric acid. This is particularly unfavorable for low grade concentrates, where the S / Cu ratio is high. In Canadian Patent Application Number 2,099,333, a process for the extraction of copper is described, whereby a sulfide concentrate is subjected to pressure leaching in the presence of chloride and sulfate ions, to produce an insoluble basic copper salt . The copper salt is removed by filtration, and the filtrate is recycled to oxidation under pressure. The present invention provides a process for the hydrometallurgical extraction of copper and other metals, in the presence of halogen ions, such as chloride and bromide in solution.

SUMMARY OF THE INVENTION In accordance with the invention, a process is provided for the extraction of copper from a copper sulphide mineral or concentrate, which comprises the steps of subjecting the ore or concentrate, together with an ion source. of bisulfate or sulfate, under pressure oxidation, at a temperature of about 115 ° C to about 160 ° C, in the presence of oxygen and of an acid solution containing halide ions, to obtain a resulting pressure oxidation paste; subjecting the paste to a liquid / solid separation step, to obtain a resulting pressure oxidation solution and a solid residue; and recovering the copper from the oxidation solution under pressure or the solid residue; characterized in that the pressure oxidation solution is recycled towards oxidation under pressure, and the concentration of copper in the pressure oxidation solution that is recycled is maintained at a predetermined value.

BRIEF DESCRIPTION OF THE DRAWINGS The invention will now be described by way of examples with reference to the accompanying drawings, in which: Figure 1 is a flowchart of a hydrometallurgical copper extraction process according to one embodiment of the invention, which is suitable for the treatment of high-grade copper concentrates or minerals. Figure 2 is a flow chart of a hydrometallurgical copper extraction process according to another embodiment of the invention, which is suitable for the treatment of minerals or copper concentrates of medium and lower grade. Figure 3 is a schematic illustration of a pressure vessel to illustrate another embodiment of the process for the treatment of concentrates with a high proportion of the sulfur to the metal. Figure 4 is a flow chart of a hydrometallurgical process for extracting metals from a copper-zinc sulfide concentrate, in accordance with another embodiment of the invention. Figures 5A and B show a flow diagram of a further embodiment of the process according to the invention, for the recovery of precious metals from a mineral or concentrate.

DETAILED DESCRIPTION OF THE PREFERRED MODALITIES The different embodiments of the process according to the invention can be used to treat a range of copper and other metal concentrates, wherein the degree of copper varies from low, that is, approximately 15%. copper or less, to a high degree, that is, approximately 35 percent copper or more. Broadly, the process comprises a step of oxidation under pressure that takes place in the presence of oxygen and an acid solution of halide ions, for example, chloride or bromide, and sulfate ions. More specifically, the process also includes an atmospheric leaching stage, one or more solvent extraction stages, and an electrolytic extraction stage. Different degrees of concentrate require different treatment in the oxidation stage under pressure, requiring different modes of operation. These operating modes are called Mode A, Mode B, and Mode C, respectively. In Mode A, which is effective when leaching high-grade copper ores, copper is not leached at the pressure oxidation stage. In Modes B and C, which are effective when leaching medium and low grade copper ores, the copper is leached at the pressure oxidation stage. Now each of the modes of operation will be described in turn.

Process Mode A Figure 1 is a flow chart of Mode A. The process comprises a step of oxidation under pressure 12, in a pressure oxidation vessel or autoclave, an atmospheric leaching stage 14, primary solvent extractant stages and secondary 16 and 18, respectively, and an electrolytic extraction stage 20. In the step of oxidation under pressure 12, all copper ores are converted to basic copper sulfate, CuSO 4"2Cu (OH)" The treatment is carried out with oxygen in the presence of an acid chloride solution. Oxygen, as well as HCl and H2S04 are introduced into the autoclave for this purpose. The temperature in the autoclave is approximately 130 ° C to 150 ° C, and the pressure is approximately 100 to 200 psig (1380 kPa). This is the total pressure that comprises the oxygen pressure plus the vapor pressure. The retention time is approximately 0.5 to 2.5 hours, and the process is normally carried out continuously in the autoclave. However, the process can also be done in a batch form, if desired. The solids content in the autoclave is maintained at approximately 12 to 25 percent, that is, 150 to 300 grams / liter of solids, determined by the heat balance and viscosity limitations. The paste produced in the autoclave is discharged through a series of one or more evaporation tanks 22 to reduce the pressure to atmospheric pressure, and the temperature to 90-100 ° C. The liquid part of the paste is referred to as the product solution from the pressure oxidation stage 12, and is indicated by the reference numeral 21. The paste from the evaporation tank 22 is filtered, as shown in 24, and the The resulting filter cake is washed thoroughly to remove the liquor as much as possible. The pressure oxidation filtrate from the filtration 24 is recycled to the pressure oxidation step 12, but there is a small purge of about 5 percent, as shown at 26. This purge 26 is determined by the concentration of the soluble metals in the concentrated mineral that can be dissolved during the oxidation step under pressure 12. The purge 26 is treated in 28 with lime to remove the metals, such as zinc and magnesium, as solid waste, which are present in the copper concentrate, and to counteract the accumulation of these metals in the pressure oxidation circuit. The pressure oxidation circuit is the circuit from the pressure oxidation stage 12 to the evaporation tank 22, until the filtration 24, until the purge 26, and back to the pressure oxidation stage 12. This is indicated by the reference numeral 23. The purge 26 is subject to a solvent extraction, as shown in 27, before the purge treatment 28. The extraction with solvent 27 is carried out by means of a suitable organic extractant to remove the copper from the purge 26. This extraction with solvent is associated with the solvent extraction stages 16 and 18, and will be referred to again when the last two stages of the extraction are described. extraction with solvent. Prior to the oxidation step under pressure 12, the copper concentrate is first subjected to a regrind, as shown at 30, to reduce the particle size to approximately 97 percent mesh minus 325, which corresponds to P80 ( spend 80 percent), 15 micras. The regrind 30 is made in the recycled solution from the purge treatment 28. Accordingly, the pulp from the purge treatment 28 is subjected to a liquid / solid separation, as shown in 32, and the solution is recycled to the regrind 30, and the residue of the zinc / magnesium purge is discarded, as shown at 17. The solution that is recycled to the regrind 30 is an alkaline chloride liquor at a pH of about 10. The use of this liquor minimizes the inlet of water into the pressure oxidation circuit 23, which is important to maintain the heat balance, and to preserve the chloride solution in the pressure oxidation circuit 23 as much as possible. As previously reported, the copper is not leached in the oxidation step under pressure 12, but is converted into a basic insoluble copper salt. The feed solution to the pressure oxidation stage 12, which is leach liquor that is being recycled from the filtration 24, is indicated by the reference numeral 25. Although there is copper present in the feed solution 25, there is no copper additional leachate, that is, the process is operated in such a way that the concentration of copper in the feed solution 25 to the oxidation step under pressure 12, is equal to the concentration of copper in the product solution 21 from the oxidation stage pressure 12. This is indicated as:? [Cu2 +] = 0. The feed solution 25 to the pressure oxidation stage 12 contains approximately 15 grams / liter of Cu, and 12 grams / liter of Cl, together with approximately 30 to 55 grams / liter of sulfuric acid. The acid is added in the form of filling H2S0 (normally 93 percent). The product solution 21 from the pressure oxidation stage 12 also contains approximately 15 grams / liter of Cu and 11 to 12 grams / liter of Cl, but is at a pH of approximately 3. Substantially no acid remains in the product solution 21 , since this is consumed all in the oxidation step under pressure 12 to form the basic copper salt. As referred to above, the liquid feed 25 to the pressure oxidation stage 12 is partially filled with the recycled filtrate to which H2SO4 is added.

The immediate effect of adding the acid to the filtrate is to increase the acidity of the filtrate that is fed to the autoclave for the pressure leaching step 12, but in a surprising way, it has been discovered that the most important effect is the addition of acid, or more specifically sulfate ions, which actually suppress the oxidation of the sulfur emanating from the concentrate in the oxidation step under pressure. Typically, the oxidation of sulfur that is experienced if no acid is added, is approximately 25 to 30 percent of the feed sulfur in the concentrate, as is the case with the process described in United States Patent Number 4,039,406. However, if acid is added, it has been discovered that the oxidation of sulfur in sulfate is reduced to about 5 to 10 percent. This improvement has substantial beneficial effects on the hydrometallurgical extraction process. The oxidation of sulfur in sulphate creates additional costs in various ways, such as the additional oxygen required for the reaction, the additional reagent required to neutralize the acid thus formed by oxidation, and heat removal due to oxidation must be available. of sulfur in sulphate, which is very exothermic. This really limits the production of the autoclave where the oxidation stage under pressure takes place 12.

It is believed that the reaction chemistry in the oxidation step under pressure 12 is altered by the addition of the acid as follows: Without Acid Addition: (1) 3CuFeS2 + 21/402 + 2H20? [CuS04 «2Cu (0H) 2] + 3 / 2Fe203 + 5S ° With Addition of Acid: (2) 3CuFeS2 + 15/402 + H20 + H2S04? CuS? 4 »2Cu (OH) 2 + 3 / 2Fe203 + 6S ° In both reactions, copper is precipitated in the form of a basic copper salt, which has been found to comprise mostly basic copper sulfate. In the first reaction, it appears that the basic copper sulfate sulfate is supplied by the oxidation of the feed sulfur in the concentrate, while, in the second reaction, it appears that it is supplied by the sulfate ions of the acid that is added to the the autoclave, obviating in this way the need for the oxidation of sulfur in sulphate. Therefore, in the second reaction, there is a net consumption of sulfate ions to form the basic copper salt. It has been experimentally discovered that the amount of sulfuric acid necessary to suppress the oxidation of sulfur is approximately 25 to 75 grams per liter, depending on the type of concentrate and the percentage of solids in the concentrate. In actual test work, there is more sulfur oxidation than what is predicted by any reaction. The first reaction predicts a sixth or 16.7 percent of the sulfur to be oxidized, while experimentally it is about 25 to 30 percent. With addition of acid, the experiments indicate that it oxidizes from about 2 to 16 percent sulfur in sulfate, instead of the zero oxidation that would be predicted if the second written reaction were the only reaction that took place. Accordingly, these reaction equations do not reflect exactly what is happening in the pressure leaching stage 12, but are only an approximation. The chloride is retained as much as possible in the pressure oxidation circuit 23, but typically about 3 to 10 percent of chloride is lost per pass to the solid product in filtration 24. Therefore, the chloride must be filled by addition of HCl or other chloride source, to provide 12 grams / liter of chloride in the feed solution 25. Losses of chloride are minimized by thorough washing of the solids from the oxidation stage under pressure 12 over the filter 24 The amount of wash water is limited by the requirement to maintain a water balance in the pressure oxidation circuit 23. The only water loss from the circuit 23 is in the vapor 9 from the evaporation step 22., and on the filter cake after evaporation 24. Accordingly, hence the need to use the recycled solution from the purge treatment 28 to form the concentrate paste in the milling step 30, and thereby minimize the fresh water inlet from the concentrate to the oxidation step under pressure 12. It has been discovered that it is convenient to maintain at least 15 grams / liter of Cu in the product solution 21 from the oxidation step under pressure 12, to counteract the loss of chloride in the form of solid basic copper chloride, CuCl2"3Cu (OH) 2, which may occur if insufficient copper is present in solution to allow basic copper sulfate to form: 4CuCl2 + 6H20? CuCl2- »3Cu (OH) 2 + 6HC1 (3) This reaction can be counteracted by the addition of sufficient acid to the autoclave during the oxidation step under pressure 12, to maintain at least enough copper in solution to satisfy the stoichiometric requirements of Cl as CuCl 2. For 12 grams / liter of Cl in solution, the stoichiometric amount of Cu is: 63-5 X 12 = 10.7 g / L Cu 71 Therefore, 15 grams / liter of Cu is a safe minimum to prevent a significant loss of chloride in the form of the basic copper salt. On the other hand, the concentration of copper in the product solution 21 from the oxidation stage at pressure 12, should be kept as low as possible, to counteract the formation of CuS, by the reaction of elemental sulfur with aqueous copper sulfate. This reaction may occur during the oxidation step under pressure 12, or in the paste after discharge from the autoclave, but before the filtration step 24: 3CuS? 4 (aqueous) + 4S ..oU + 4H20? 3CuS (s) + 4H2S04 (4) This reaction is particularly undesirable, because the CuS is insoluble under the dilute acid conditions of the atmospheric leach stage 14. Accordingly, the copper is not recovered, and this results in the loss of copper towards the final residue.

To counteract the formation of CuS, it is necessary to keep the concentration of copper in the product solution 21 as low as possible, that is, below 30 grams / liter for some concentrates. The tendency to the formation of CuS apparently is related to the type of concentrate that is being treated, with medium to high grade concentrates that are more susceptible to the formation of CuS. Therefore, although a high concentration of copper in the product solution 21 does not present a problem with the low grade concentrates, it can not be tolerated with the higher grade concentrates. As is known to date, high grade concentrates, ie, more than 35 percent copper, are best treated to produce as low a copper concentration in the product solution 21 as possible, ie less than 25 grams / liter of Cu. Given the need to maintain at least 15 grams / liter of Cu in solution in the pressure oxidation circuit 23, there is an optimum copper concentration range of 15 to 25 grams / liter of Cu for high grade concentrates. With medium grade concentrates, the upper limit can be stretched considerably, and for the low grade ore, the copper concentration does not play a significant role. The concentration of copper in the oxidation filtrate under pressure 29 can be controlled simply by adding the required amount of acid in the feed solution 25 to the oxidation step under pressure 12. It results more acid at a higher copper concentration, due to the dissolution of basic copper sulfate: CuS? 4- »2Cu (OH) 2 (s) + 2H2S04? 3CuS04 (aqueous) + 4H20 (5) The addition of about 1 gram / liter of acid results in an increase in copper concentration of about 1 gram / liter. The actual concentration of acid required is determined empirically by comparing the tests of the feed solution 25 with the step of oxidation under pressure 12 and the product solution 21 from the step of oxidation under pressure 12 to satisfy? [Cu2 +] = 0. The volume of solution in circuit 23, however, is determined by the heat balance. The percentage by weight of solids in the feed of the paste of the copper concentrate for the step of oxidation under pressure 12, can be varied according to taste. The weight of the concentrated solid fed to the oxidation stage under pressure 12 is determined by the amount of copper to be recovered. The weight of the solution is mainly determined by the heat equilibrium in the oxidation step under pressure 12.

The desired operating temperature in the oxidation stage under pressure 12, it is approximately 150 ° C, and the heat must be supplied largely by the heat of the reaction of the sulfide minerals with the high pressure oxygen in the autoclave. For high-grade concentrates, such as will be processed by Process Mode A that is currently being described, this means a relatively low S / Cu ratio, and consequently, a smaller heat production per ton of copper treated in the autoclave Much of the heat released is due to oxidation, not copper, but the other two main elements of the concentrate, iron and sulfur. If the degree of the concentrate is high, then the ratio of S / Cu and Fe / Cu is low, and therefore, there is a lower heat production. To reach the operating temperature from an initial temperature, say, 50 ° C to 80 ° C, which is typical for the pressure oxidation filtrate 29 that is recycled after filtration 24, it is necessary to control the amount of water that is must heat, since this is the main heat sink in the oxidation stage under pressure 12. It is impractical to cool or heat the paste inside the autoclave by indirect elements, such as by means of heating or cooling coils, due to the rapid formation of scale on all surfaces, particularly in heat exchangers, which leads to very poor heat transfer characteristics. Heating or direct cooling by steam or water injection is also impractical, due to water balance considerations. Accordingly, it is required that the water balance be maintained by balancing the production of heat from the heat of the reaction with the heat capacity of the feed materials, i.e., the feed solution 25 being recycled and the paste of the concentrate. The main variable that can be controlled here is the volume of the feed solution 25. This is one of the distinguishing characteristics between Modes A and B. In Process Mode B, which is still going to be described, where experience greater oxidation of sulfur, the heat release is much higher, expressed as heat per ton of copper product. Accordingly, it is possible to use more volume of solution in the feed 25 towards the step of oxidation under pressure 12. Once the volume of the solution is fixed, the acidity of the solution can be determined, since the total mass of acid is determined by the need to maintain? [Cu2 +] = 0. Typically, for a high grade concentrate, approximately 35 to 55 grams / liter of acid will be required. It has been discovered that it is beneficial to add small concentrations of certain surfactants that change the physical and chemical characteristics of liquid elemental sulfur (S °) in the autoclave during the oxidation stage under pressure 12. Surfactants, such as lignin sulfonate and quebracho, Aggregates to the feed solution of the oxidation under pressure in small quantities, that is, from 0.1 to 3 grams / liter, can reduce the viscosity of the liquid sulfur, and also change the chemistry in the autoclave. The additions of surfactants can reduce the oxidation of sulfur in ways that are not well understood, but are beneficial to the process. It is believed that this is due to the lower viscosity, which results in a lower tendency for liquid sulfur and solids to be kept up inside the autoclave thus reducing the retention time of these materials, and consequently, the reduced tendency for sulfur oxidation to occur. It has also been found that a more complete reaction of the copper ores takes place if surfactants are added, apparently due to the lower viscosity sulfur, which does not "wet" the unreacted sulphide minerals, and consequently allows the reaction until finished. The reaction (5) describes how the addition of sulfuric acid to the oxidation feed under pressure, will control the concentration of copper in the oxidation filtrate at pressure 29. The overall reaction for oxidation under pressure with the addition of sulfuric acid to a chalcopyrite mineral, is given by reaction (2) above. You can write a similar reaction using CuS04 as the source of sulfur ions instead of H2S04: 3CuFeS2 + 15/402 + 3H20 + 3 / 2CuS04? 3 / 2CuS04- »2Cu (OH) 2 + 3 / 2Fe203 + 6S ° (6) It is worth noting that there is 3/2 moles of sulfate required as copper sulfate in the reaction (6), compared with one mole of sulfuric acid in the reaction (2). Therefore, if CuS04 is to be used as the source of sulfate ions, instead of sulfuric acid, it is necessary to use 1.5 times as many moles of CuS04. To take this into account, the inventor has developed the concept of the Excess Sulphate Equivalent, which allows calculating how much acid to add to the feed solution to pressure oxidation 25, in order to reach an objective copper concentration, and still take Consider the reaction (6). By taking the reaction (6) into account, it is possible to calculate "a priori" the amount of acid required for a constant copper concentration in the oxidation filtrate at pressure 29. The concept of the Excess Sulphate Equivalent is useful: The Equivalent of Excess Sulfate is equal to the sulfate available in the feed solution of the oxidation under pressure for the formation of basic copper sulphate during the step of oxidation under pressure 12. The available sulfate is that which is in excess of a Base Level defined from CuS04 and CuCl2. The Base Level of CuS04 and CuCl2 is sufficient to support the chloride in solution at 12 grams / liter, in the form of CuCl2, and in addition, approximately 4.3 grams / liter of Cu as CuS04. The concentration of CuCl corresponding to 12 grams / liter of chloride in solution is 134.5 / 71 * 12 = 22.7 grams / liter of CuCl2, which contains 10.7 grams / liter of Cu in solution. The additional 4.3 grams / liter of copper, therefore, means a total of 15 grams / liter of copper combined as CuCl2 and CuS04 at the Base Level. The available sulfate is then the total sulfate as CuS04 minus the Base Level. For example, if the total copper concentration is 28 grams / liter in the pressure oxidation filtrate 29, then the available sulfate is 28 - 15 = 13 grams / liter of Cu * 98 / 63.5 = 20 grams / liter of H2S04 as sulfate available from CuS04. Then the Excess Sulphate Equivalent (ESE) is calculated from the available sulfate of CuS04 by dividing by 1.5: ESE =. { Sulfate Available as CuS04} /l.5 Therefore, in the example of 28 grams / liter of total copper concentration, or 20 grams / liter of available sulphate from CuS0, there is 20 / 1.5 = 13.3 grams / liter of Excess Sulphate Equivalent from CuS04. Finally, if the objective free acid equivalent is, say, 52 grams / liter of H2SO4 in the pressure oxidation feed solution 25, then the amount of acid required is 52 minus the Excess Sulphate Equivalent (13.3 grams / liter) or 38.7 grams / liter of H2SO4. This is the amount that must be added to the feed solution 25 in the oxidation step under pressure 12, to produce a constant copper concentration in the oxidation filtrate under pressure 29, that is, the Base Level of 15 grams / liter of Cu. Other reactions can be written using Fe2 (S04) 3 and ZnS04 as the source of sulfate ions, instead of H2S04. In the case of ZnSO4, it is assumed that zinc is hydrolysed in basic zinc sulfate, ZnS0 »3Zn (OH) 2, which is a basic salt of Zn analogous to basic copper sulfate. These reactions are given below as reactions (7) and (8). 3CuFeS2 + 15/402 + 2H20 + l / 3Fe2 (S04) 3? CuS04 »2Cu (OH) 2 + ll / 6Fe203 + 6S ° (7) 3CuFeS2 + 15/402 + 13 / 3H20 + 4 / 3ZnS04? (8) CuSO4- »2Cu (OH) 2 + 6S ° + Fe203 + l / 3. { ZnS04- »3Zn (OH) 2-» 4H20} The solids from the pressure oxidation stage 12 after the filtration 24 are treated in the atmospheric leaching stage 14 at a pH of about 1.5 to 2.0, using the raffinate from the extraction stage with primary solvent 16, which is acid, to dissolve the basic copper sulfate. The leaching 14 takes place at a temperature of about 40 ° C, during a retention time of approximately 15 to 60 minutes. The percentage of solids is typically from about 5 to 15 percent, or from about 50 to 170 grams / liter, although it is possible to operate the process outside this range. During the atmospheric leaching stage 14, the basic copper salts dissolve almost completely, with very little of the iron present in the concentrate going to the solution. Typically, the leach liquor produced after the liquid / solid separation 34, contains from about 10 to 20 grams per liter of copper, depending on the percentage of solids fed to leaching 14, with 0.1 to 1.0 grams / liter of iron , and from approximately 0.1 to 1.0 grams / liter of chloride. Much of this iron and chloride is derived from the raffinate of the feed 37, rather than from the oxidation solids under pressure, that is, they are recycled. Typically, they dissolve from approximately 0.1 to 0.2 grams / liter of iron and chloride per pass. It has been found that copper extraction is about 95 to 98 percent, based on the original feed to the pressure leaching stage 12. It has been found that the extraction of iron to the solution is less than about 1 per cent. hundred. The paste 31 from the atmospheric leaching stage 14, is difficult if not impossible to filter, but sits well. In view of the need to wash the solids from leaching very completely, therefore, the pulp 31 is pumped into a countercurrent settling wash circuit (CCD), symbolically indicated as a solid / liquid separation 34 in the Figure 1. In the countercurrent settling circuit 34, the solids are fed through a series of thickeners with washing water added in the opposite direction. By this method, the solids are washed and the methylated solution is removed. Approximately 3 to 5 thickeners (not shown) with a wash ratio (from water to solids) of about 5 to 7 are required to reduce liquor dropped to less than 100 ppm Cu in the final residue. The lower thickener flow from the last thickener is the final waste stream with approximately 50 percent solids. This can be treated for the recovery of precious metals, such as gold and silver, or it can be sent to the queue. The recovery of precious metals will be described later with reference to Figure 5. The main constituents of stream 35 are hematite and elemental sulfur, which can be recovered by flotation, if market conditions warrant it. The overflow of thickener from the first thickener is the product solution 33, which is fed to the extraction stage with primary solvent 16, as shown. As an example, this solution contains approximately 12 grams / liter of Cu, 1 gram / liter of Cl, and 0.5 grams / liter of Fe. The concentration of optical copper is determined by the capacity of the extraction stage with solvent 16 to extract the maximum copper of the solution 33. Since a fraction of about a third of the raffinate from the solvent extraction stage 16 is eventually neutralized, it is important to minimize the copper content of this raffinate.

Solvent extraction works best on diluted copper solutions, due to the fact that a concentrated copper solution results in a higher concentration of acid in the raffinate, which tends to lower the efficiency of the extraction. However, more concentrated solutions are cheaper to deal with from the point of view of the cost of capital, since the volume is lower. Above a certain point, however, the increased concentration does not reduce the size of the solvent extraction unit, since (i) there is a maximum organic load, and (ii) the water volume generally remains equal to the organic volume for the mixing purposes through aqueous recycling. Therefore, the total volume of organic extractant and aqueous solution is only determined by the volume of the organic extractant. The maximum organic load, and therefore, the volume of organic, is determined by the concentration and characteristics of the particular organic solvent selected. For the typical solvent, for example, the LIXMR reagent from Henkel Corporation, the maximum charge per pass at a concentration of 40 percent by volume in diluent is about 12 grams / liter of Cu. Therefore, the product solution 33 should also contain approximately 12 grams / liter of Cu. The copper is extracted from the product solution 33 from the overflow of thickener from the decanting to countercurrent in two stages of extraction in the step of extraction with primary solvent 16, to produce a raffinate 37 with approximately 20 grams / liter of acid free, and from approximately 0.3 to 1 gram / liter of Cu. Most of this raffinate 37 is recycled to the atmospheric leach stage 14, but from about 25 to 30 percent is a surplus for the acid requirements of the atmospheric leach stage 14, and must be neutralized. This surplus 121 is divided as shown in 36, and neutralized. The neutralization is carried out in two stages to maximize the recovery of copper, and to prevent possible environmental problems with the neutralization residue due to the copper content, that is, the copper not recovered from the raffinate 37 will precipitate on neutralization, and then it can be redissolved later, in a tail pond, for example. The neutralization of the first stage takes place at a pH of 2 to 3, as shown in 38, using limestone, which is very economical as a reactant, compared with lime. The neutralization product is filtered at 40, and the resulting solids are washed with water from the external source 45. The solids, which are mainly gypsum and iron hydroxides are discarded, as shown at 41. The filtrate 39 is sent to the extraction stage with secondary solvent 18 for recovery of residual copper values. The secondary solvent extraction 18 benefits from the primary neutralization 38, and results in a very low concentration of copper in the secondary raffinate 43, typically from about 0.03 to 0.06 grams / liter of Cu. As indicated by the dotted lines of Figure 1, the secondary solvent extraction stage 18 uses the same organic extractant as the primary solvent extraction circuit 16. This is also linked with the solvent extraction 27 of the filtrate purge of the oxidation under pressure 26. The organic extractant which is washed in 42 with wash water 122 from an external source 45, and which is separated in 44, is recycled to the extraction stage with secondary solvent 18, and then passes to the primary extraction stage 16. Separated organic 125 is divided to pass a portion thereof to solvent extraction 27. Raffinate from solvent extraction 27 is added to charged organic 123 from extraction with solvent 16 before washing 42. The wash water 47 from the wash 42 is passed to the pressure oxidation filter 24, to serve as a feed wash water on the filter 24. The wash filtrate results This is added to the pressure oxidation filtrate 29, thereby recovering the copper and chloride content from the extraction wash water with solvent 47.

The raffinate 43 from the extraction stage with secondary solvent 18, is again neutralized in a secondary neutralization step 46, this time at a pH of 10, and filtered at 48 to remove all the dissolved heavy metals, producing a solution 51 that it is used as the wash water in the countercurrent settling circuit 34, to wash the residue from the final leach 35. The solid residue from filtration 48 is discarded, as shown in 53. The separation of the charged organic and washing at 44 is effected by spent acid or electrolyte 55 from the electrolytic extraction stage 20, to obtain a solution of pure copper sulfate or the pregnant electrolyte 57, which is then passed to the electrolytic extraction stage 20, for the electrolytic extraction in the usual way. It can be seen that all solution streams in the process are recycled in this way, and there are no effluents of solution from the process. Only solid waste from the process is discarded.

Process Mode B Figure 2 is a flow chart of Mode B. The same reference numerals are used to indicate the stages or steps of the process that correspond to those of the previous modality of Figure 1. For example, the stage of oxidation under pressure again is indicated by 12, the atmospheric leaching stage by 14, the electrolytic extraction stage by 20, the evaporation tank by 22, the pressure oxidation filtration by 24, the purge treatment of the filtrate of the pressure oxidation 29 by the reference numeral 28, the milling step by the reference numeral 30, and the back-washing decanting circuit by the reference numeral 34. In this mode of the process, the pressure oxidation 12 is carried out both to oxidize and to leach in solution most of the copper contained in the feed concentrate. Typically, about 85 to 90 percent of the copper in the solution is leached, leaving only about 10 to 15 percent in the residue as the basic copper sulfate. The conditions of the oxidation step under pressure 12 in the autoclave are similar to those of Process Mode A, with the exception that the percentage of solids is lower, that is, from 150 to 225 grams / liter. In this process mode,? [Cu2 +] is typically from to 40 grams / liter of Cu, that is, the copper concentration is higher in the product solution 21 from the oxidation stage at pressure 12. The feed solution 25 to the pressure oxidation stage 12, typically contains 10 to 15 grams / liter of Cu, and 12 grams / liter of Cl, together with approximately 20 to 30 grams / liter of sulfuric acid. In this mode, sulfuric acid is not added to the oxidation step under pressure 12 from an external source, as is the case with the embodiment of Figure 1. In this mode, the acid is obtained from the recycle of the process, i.e. , by recycling the oxidation filtrate under pressure 29. The product solution 21 from the oxidation stage at pressure 12 contains approximately 40 to 50 grams / liter of Cu, and 11 to 12 grams / liter of Cl, at a pH from about 2 to 2.5. The copper leached to the product liquor 21 from the pressure oxidation stage 12, must be controlled to obtain the desired copper distribution between the liquor (from 85 to 90 percent) and the residue (from 10 to 15 percent). This distribution results in a small but significant amount of basic copper sulfate solids in the leach residue. The pH is a convenient indicator of the presence of basic copper sulfate, since it is a pH regulating agent. With the concentration of strong copper sulfate in solution, a pH range of 2 to 2.5 indicates basic copper sulfate. Below a pH of 2, almost all basic copper sulfate will dissolve, while above a pH of 2.5, too much basic copper sulfate is formed, and it is possible that insufficient copper is found in the solution 21. The primary control method is the amount of acid in the feed liquor 25 towards the oxidation step under pressure 12. In turn, the acid level is controlled by the degree of neutralization of the raffinate from the solvent extraction of the raffinate from the oxidation filtrate under pressure described below. Normally, about 25 to 50 percent of the acid must be neutralized, depending on the amount of acid required. The acid generated during the oxidation step under pressure 12 varies from one concentrate to another, and in accordance with the conditions employed. If the concentrate produces a large amount of acid during the oxidation step under pressure 12, then the feed solution will need less acid to achieve the desired result. The minimum copper (from the feed of the concentrate) that must go to the liquor 12 is approximately 10 percent. Below 10 percent, the pH falls sufficiently low so that the iron concentrations in the pressure oxidation filtrate 29 increase rapidly. Normally, the iron is about 10 to 50 ppm, but if the pH is below 2, and the basic copper sulfate disappears in the residue, then the iron can be increased up to 1 gram / liter very quickly. This is undesirable, because there are several impurity elements such as As and Sb that are only removed from the solution simultaneously with the hydrolysis of the iron. Therefore, the absence of iron in solution is a good guarantee of a low content of impurities in the filtering of the oxidation under pressure 29. Iron is also an impurity in itself, which should be avoided in the electrolytic extraction circuit 20 as possible. However, there is another factor that puts a maximum on the Cu in solution. Surprisingly, it has been discovered that certain concentrates actually leach more completely if the copper concentration is lower. It is believed that this is due either to the formation of secondary CuS, as described above, or to some other phenomenon related to poor oxidation characteristics of the primary mineral, chalcopyrite, in solutions of high copper concentration. It is found that the elemental sulfur, produced during the reaction in the oxidation step under pressure 12, can resort to or actually encapsulate the unreacted chalcopyrite particles, and hinder the access of the reagents. This results in poor copper recovery. The phenomenon is apparently accentuated by the high levels of Cu in solution. This can be overcome or mitigated by the use of surfactants as described above. The problem is more severe with some concentrates, particularly high grade, than others. Therefore, for these concentrates, it is desirable to limit the concentration of copper in the oxidation filtrate under pressure (i.e., more than about 95 percent) above all. To do this, it is necessary to have a substantial proportion of the copper as basic copper sulfate, that is, in a solid residue from the pressure oxidation stage 12, rather than the pressure oxidation filtrate. Typically, 20 to 40 percent of the copper can be reported in solids, if necessary, to keep the copper concentration low enough to obtain a copper recovery. Higher grade concentrates exhibit the problem of low copper recovery with a high copper content in solution. Therefore, an increasing proportion of copper should be reported in solids as the grade increases. Tests with three different concentrates illustrate this relationship: Conc. #% Cu H + / Cu Distribution of Cu. % Molar Liquor of PO Waste of Total PO Recovered 1 41 0.55 0 100 97.3 2 28 0.70 63 37 95.7 22 0.96 85 15 94.7 The molar ratio of H + / Cu refers to the H + in the feed acid, and to the Cu in the feed concentrate. The H + in the feed acid is taken as all the available protons on the complete dissociation of the acid, even when, under the existing conditions, the acid is not completely dissociated. The H + shown in the table is the optimal level found through experimentation, to give the best results. For concentrate # 1, which was a high grade concentrate, the chosen process is Mode A, where all the copper is reported to the lixiviation liquor 33 and? [Cu2 +] = 0. The ratio of H + / Cu is the found as necessary by experimentation, to give the desired result of? [Cu2 +] = 0. For concentrate # 2, a medium grade concentrate, Mode B was chosen, but with a substantial amount of copper reported to copper sulfate basic solid. This was achieved by keeping the ratio of H + / Cu sufficiently low, so that not all the copper dissolved in the liquor. For Concentrate # 3, a low grade concentrate, Mode B was also chosen, but in this case, the minimum amount of copper reported to the residue, by adjusting the H + / Cu ratio, was sufficiently high. The residue from the pressure oxidation stage 12 is leached 14 with the raffinate 37, returning from the extraction with solvent 16 which is dilute acid, to 3-10 grams / liter of H2SO4. Since most of the copper from the pressure oxidation stage 12 is reported to the pressure oxidation filtrate 29, and only a small fraction of the pressure oxidation residue, the resulting leaching liquor 31 from atmospheric leaching 14, is very diluted in copper. In turn, this produces a raffinate diluted 37 from the extraction with solvent 16. Typically, the atmospheric leach liquor 31 is 3 to 7 grams / liter of Cu, and 0.2 to 0.5 grams / liter of Fe. The resulting paste of atmospheric leaching stage 14 is difficult to filter, as was the case with Mode A. However, good liquid / solid separation and washing can be achieved, as before, using a series of thickeners in a decanting configuration countercurrent 34. Wash water 51 is provided by raffinate from extraction with solvent 16, which is neutralized, as indicated at 46. This is similar to Mode A. The only important difference is the lowest content of the solution 33, and the reduced volume. The solution 33 produced by the atmospheric leaching stage 14 is subjected to solvent extraction 16. The copper-containing solution 29 from the pressure oxidation stage 12 is subjected to a solvent extraction step. Accordingly, there are two solvent extraction operations, i.e., 16 and 50, which treat two different liquor streams 33 and 29, respectively. It is a feature of the process according to the invention that the organic extractant used to perform the solvent extraction operations is common for both solvent extractions 16 and 50. As shown in Figure 2, the separated organic 125 that comes from the common separation operation 44, is first introduced into the solvent extraction circuit 16, which has the weakest copper concentration in the aqueous feed stream 33, and therefore, needs the organic extractant to be as low as possible in the load, to be effective. The organic loaded 126 from the extraction with solvent 26, is then sent to the extraction with solvent 50, where it makes contact with the liquor with the highest concentration of copper 29. It is not necessary that the extraction with solvent 50 reaches a high proportion of extraction, because the raffinate 63 from this extraction is recycled to the oxidation step under pressure 12, as shown. On the other hand, the raffinate 37 from the extraction with solvent 16 is only partially recycled, and a part is neutralized 46, to remove excess acid from the circuit. Therefore, it is more important to achieve high copper recovery from solvent extraction 16.

Raffinate 37 from solvent extraction 16 is divided into 36 as in Mode A, with approximately one third 121 towards neutralization 46, and two thirds 120 recycled towards atmospheric leaching stage 14. An important difference of Mode A , is that the raffinate 37 from the extraction with solvent 16 is sufficiently low in copper, that is, less than 100 ppm, in such a way that it is not necessary to have a secondary solvent extraction step before neutralization 46, as was the case in Mode A. This is due to the lower concentration of copper and the volume of the solution, which allows solvent extraction 16 to be more efficient. The charged organic 65 produced by the two solvent operations 16, 50 in series, is washed in two stages in a countercurrent fashion, with a dilute acidic aqueous solution 122, as shown at 42. This is primarily to remove the aqueous solution placed from the loaded organic 65, and in particular, to reduce the chloride content before the organic reaches the separation at 44. The amount of wash water required is approximately 1 to 3 percent volume of organic. The resulting wash liquor 47 produced is recycled to the pressure oxidation stage 12. The washed organic 69 is separated at 44 with the spent electrolyte 55 from the electrolytic extraction stage 20, to provide a solution of pure copper or the pregnant electrolyte. 57 for electrolytic extraction in the usual manner. The raffinate 63 is divided into 70 into two portions 72, 74, as determined by the required molar ratio of H + / Cu. Portion 72 is recycled to pressure oxidation stage 12. Portion 74 is neutralized to a pH of 2 with limestone at 76, and filtered 78. The solid residue is washed and discarded, as shown at 80. The filtrate 82 is recycled with the portion 72 to form the feed solution 25 to the pressure oxidation stage 12. Therefore, a novel feature of the process is the use of a common organic to extract the copper from two aqueous feed liquors. separated. This provides considerable savings by lowering capital and operating costs in the solvent extraction circuits. It also allows the use of copious amounts of water in the atmospheric leaching back-stream decanting circuit, so that a good wash can be achieved on the final residue, and yet still recover the copper from this liquor diluted. It has been found that the degree of oxidation of sulfur that occurs in the oxidation stage at pressure 12, depends very much on the type of concentrate, such as the grade and mineralogy of the concentrate being treated, as well as the conditions of the stage. pressure oxidation 12. Certain concentrates exhibit a considerably higher sulfur oxidation, for example, from 25 to 30 percent, that is, an oxidation of the sulfur in the concentrate to obtain sulfate, and in fact it is particularly remarkable with the concentrates of low grade for less than about 28 percent Cu by weight. The inventor has discovered that the meaning of this variation is not so much in the copper grade itself, but in the copper / sulfur ratio in the concentrate. The main impurity elements in a copper concentrate are iron and sulfur, due to the fact that copper ores are generally composed of chalcopyrite, along with other minerals, particularly pyrite FeS2, or pyrrolite FeS. Process Mode B deals with the problem of oxidation of excess sulfur in the pressure oxidation stage 12, when using lower grade concentrates by deliberately dissolving 90 percent of the copper, and minimizing the formation of basic copper sulfate. The reaction for chalcopyrite is: CuFeS2 + 5/402 + H2S04? CuS04 + l / 2Fe203 + 2S ° + H20 (9) The filtrate 29 from the oxidation stage under pressure 12, therefore, contains high levels of copper sulfate and copper chloride, and this is treated in the solvent extraction step to produce a pure copper sulfate solution going to the electrolytic extraction stage 20. It has been discovered that there is a limit to the amount of sulfur oxidation that can be accommodated by Process Mode B. If the oxidation of sulfur is sufficiently high, and enough acid is generated during the oxidation under pressure, there will be a surplus of acid that remains after oxidation under pressure, even when no acid is added to the feed, such as in the form of acid raffinate. In this situation, not only all the copper in the concentrate will be converted to dissolved copper sulfate, but also some of the iron in the concentrate will be solubilized by the surplus acid, for example, as ferric sulfate. It is desirable that the iron in the concentrate be reported to the oxidation residue under pressure as stable hematite, Fe203, and not to the solution, where it should be separated from the copper. Typical concentrates have a Fe: Cu ratio of at least 1: 1, and therefore, efficient and complete removal of Fe at an early stage is an important aspect of the process. Other impurities, such as arsenic, antimony, etc., are also removed with iron by means of coadsorption or precipitation mechanisms.

It has been discovered, however, that some concentrates exhibit so much sulfur oxidation (acid generation) that the acid consumption capacity of the pressure oxidation is exceeded, and some iron is leached to the solution, even under the conditions of the Process Mode B. It is a goal of the process to produce a low iron liquor, typically with 0.05 grams / liter of Fe. Some concentrates that have been tested have produced pressurized oxidation liquors with 1.0 to 12.0 grams / liter of Fe In a similar way, the pH of the liquor of the oxidation under pressure is normally directed to be in the range of 2.0 to 3.5, which corresponds to less than 1 gram / liter of free acid, but the tested concentrates have produced oxidation liquors. under pressure with a pH on the scale of 1.2 to 2.0, which corresponds to 1 to 15 grams / liter of free acid. In accordance with the above, an additional modality of the process has been developed, that is, the Mode of Process C for the treatment of the previous concentrates, called "Mode C" concentrates. The Mode of Process C will now be described at once.

Process Mode C Mode C concentrates that exhibit a strong tendency toward sulfur oxidation, and consequently, toward acid generation, are those with a high S: Cu ratio, or more generally the S: M ratio , where M = base metals, such as Cu, Zn, Ni, Co, Pb, etc., but not including Fe, which does not consume acid. Nickel or copper / nickel concentrates can often be Mode C, because they are often of low grade, the S: M ratio often being about 2: 1 or higher. Some copper or copper / gold concentrates are also Mode C, and they are low grade, due to the high pyrite content. It has been found that some copper / zinc concentrates are also high in pyrite, and consequently, also of the Mode C type. In general, there is a correlation between the pyrite content (FeS2) and the tendency towards the behavior of the type of Mode C. However, there are also exceptions to this trend, since not all pyrites react in the same way. Some pyrites oxidize sulfur more easily than others. In contrast, pyrrhotite (Fe7S8) or iron-zinc ore spharelite (Zn, Fe) S, appear to result in much less sulfur oxidation, and therefore exhibit a behavior of Process Mode A or B. In this mode of the process, the object as with Modes A and B, is to reduce the oxidation of sulfur during oxidation under pressure by the addition of sulphate or sulfuric acid. The filtering of the oxidation under pressure resulting from the high concentration of acid, will have high levels of dissolved iron and low levels of dissolved copper. Accordingly, a neutralizing agent, such as lime or limestone, is added to neutralize the pressure oxidation paste prior to filtration. Process Mode C is essentially a special case of Process Mode B, with two important differences. First, all the raffinate 63 (Figure 2) is neutralized, before returning this current to the oxidation under pressure 12, that is, there is no division of the raffinate, one part being neutralized, and the other part of the neutralization being deviated. Second, the pressure oxidation paste before the filtration 24 of the leaching residue is subjected to an extra neutralization step, the "neutralization of the oxidation under pressure", to neutralize the excess of acid and precipitate any Fe in solution at this time. A convenient opportunity for the oxidation neutralization under pressure is in a conditioning tank after evaporation 22 to atmospheric pressure, when the pulp is at or near the boiling point of the solution, i.e., approximately 80 ° C at 95 ° C. However, there is an inherent problem with this, namely the undesired formation of gypsum deposits in the following operations of the downstream unit from neutralization, where the temperatures are lower. Gypsum deposits because the solubility of calcium sulfate is lower, cause the supersaturation of calcium sulfate once the temperature is reduced. If the neutralization described above is carried out on a paste of 80 ° C to 95 ° C, using limestone, then the resulting solution will be saturated with calcium sulfate at that temperature. If the resulting solution is subsequently cooled to 40-50 ° C for solvent extraction, then the solubility of calcium sulfate is markedly reduced, and consequently, there will be a slow precipitation of solid calcium sulfate, more possible the dihydrate form known as gypsum, CaS04 «2H20. It is well known that this plaster forms a tenacious incrustation in pipes, valves, tanks, etc., and causes severe operational problems in a commercial plant. This problem has been solved by performing the neutralization inside the autoclave or pressure vessel 300 (Figure 3), at the conclusion of the pressure oxidation step 12, when the temperature is from about 115 ° C to 160 ° C. It has been found that in this temperature scale, the solubility of calcium sulfate is equal to, or lower than, at the lower temperatures, where solvent extraction takes place. Accordingly, the saturation level for calcium sulfate produced during neutralization is equal to, or lower, than at any subsequent stage of the process, and supersaturation does not occur, due to a temperature drop. In this way, the problems of inlaying with plaster are eliminated. The pressure vessel 300 in the present example has five compartments 302. In order to achieve neutralization inside the pressure vessel 300, it has been found that it is preferable to use slaked lime paste, Ca (OH) 2, instead of limestone, CaCO3, as the active neutralizing agent, as indicated in 304 in Figure 3. The slaked lime prevents the formation of carbon dioxide gas, C02, which is coupled with the reaction of the limestone with acid. The CO 2 gas occupies a large volume of space inside the pressure vessel 300, otherwise necessary for oxygen, and effectively deactivates the oxidation reaction at the desired pressure. To use slaked lime in a continuous reaction vessel, it is necessary to pump it into pressure vessel 300 in the paste form customary with 10 to 20 percent solids in water, towards the last compartments 302 of vessel 300. Accordingly , it is beneficial to perform pressure oxidation in the first three or four compartments 302, and pumping the slaked lime paste to the latter, or the last except one compartment 302. The concentrate, H2SO4, chloride, and oxygen, are introduced into the autoclave 300, as indicated at 306, 307, and 308, respectively. The amount of slaked lime to be used is determined by the amount of acid and iron that will be neutralized, and the amount of copper that may need to be precipitated in the form of basic copper sulfate. In general, it is desirable to terminate the oxidation under pressure 12 without free acid, and virtually without iron in solution, ie, less than 10 ppm of Fe, and at a pH of about 2.5 to 4.0. As previously indicated, it is important to keep adding water to the system at a minimum. This also applies to the neutralization of pressure oxidation using Ca (0H) 2 (slaked lime). Normally, a solids content of about 10 to 20 percent is the maximum that can be tolerated before the viscosity of the slaked lime paste becomes too difficult to handle. This is particularly a problem when there is Ni present in the concentrate where the consumption of Ca (0H) 2 is high. This problem can be overcome by the addition of a viscosity modifier, such as caustic, potash, or lignosol. This effectively reduces the viscosity, so that a solids content of 30 percent or higher can be tolerated. The resulting paste (indicated at 309), now at a pH of 2.5 to 4.0, is evaporated in two stages at atmospheric pressure, and then filtered (24) (Figure 2). The filter cake is washed to remove the liquor put in (Cu, Cl) as much as practical. The filter cake which now contains solid calcium sulfate, produced at the temperature of oxidation under pressure, together with other solids, such as hematite, elemental sulfur, and basic copper sulphate, proceeds towards atmospheric leaching 14, where it is leached any copper precipitated as usual, at a pH of about 1.5 to 1.8, and the resulting residue is completely washed in the countercurrent settling circuit 34. The filtrate 29 from the oxidation filtration under pressure is treated as in the of process B for the removal of Cu by means of the extraction step with solvent 50, producing a raffinate 63, which then goes to neutralization 76, as before, and then recycles back to oxidation under pressure 12, but without division of raffinate 70, as indicated above. Although the oxidation step under pressure 12 is catalyzed by chloride, it does not use a strong chloride solution for example, in the preferred embodiment only about 12 grams / liter is needed, which will support approximately 11 grams / liter of Cu or Zn as the salt of respective chloride. If a higher concentration of metals is needed or produced, it is like the sulfate salt.

Accordingly, the solutions produced by the pressure oxidation step 12 are in general mixtures of the sulfate and chloride salts, and not pure chlorides. Tests were performed on a low grade sulfide mineral to investigate the effect of pressure oxidation neutralization. In a first test, which was operated according to Mode B without neutralization, the autoclave feed comprised 10.7 grams / liter of free acid, 12 grams / liter of Cu, and 12.5 grams / liter of chloride in solution. The resulting pressure oxidation filtrate after filtration was at a pH of 1.72, with a copper concentration of 48 grams / liter, and dissolved iron present in an amount of 2,350 ppm. The solid waste sent to atmospheric leaching 14 contained 2.0 percent Cu. In a second test, on the same low grade concentrate, which was operated according to Mode C, with a neutralization step in the autoclave, the autoclave feed comprised 16.0 grams / liter of free acid, 14 grams / liter of C, and 12 grams / liter of chloride in solution. The resultant pressure oxidation filtrate 29 was at a pH of 3.05, with a copper concentration of 42 grams / liter, and only 25 ppm of iron in solution. The solid residue contained 6.5 percent Cu. No problem was found with the precipitation of gypsum later in the stages of lower temperature of the process. In both tests, the oxidation of sulfur, that is, the oxidation of the sulfur in the concentrate to obtain sulphate, was approximately 27 to 30 percent. These tests illustrate that it is possible to control the pH by neutralizing in the autoclave, thus minimizing the iron content in the oxidation filtrate under pressure, without the problem of the precipitation of gypsum in the system. Referring to Figure 4, a mode of the process is shown which is suitable for the treatment of copper-zinc concentrates with from about 20 to 25 percent Cu, and from about 1 to 10 percent Zn. The process similar to Process Mode B of Figure 2, and again using equal reference numerals to indicate the corresponding steps. Figure 4 is less detailed than Figure 2. It has been found that good extraction of Zn can be achieved in the oxidation step under pressure 12, if enough acid is added to the feed solution to maintain the final pH of the pulp under about a pH of 2. Otherwise, the conditions are similar to those of the Cu concentrates that are treated by the process in Figure 2, that is, at 150 ° C to 200 psig (1400 kPa) of 02 , in 12 grams / liter of Cl.

In the Mode B type process, Cu is solubilized primarily during the oxidation step under pressure 12, and is extracted by solvent extraction Cu 50. This solvent extraction step 50 is operated in conjunction with the solvent extraction step of Cu 16, where the Cu is extracted from the leach liquor that comes from the atmospheric leaching stage 14, as described with reference to Figure 2 above. The solvent extractions 50, 16 produce a concentrated copper solution which is treated in an electrolytic extraction stage 20, as described above. The residue 35 from the atmospheric leaching stage 14 is treated for the recovery of sulfur and precious metals (39), as will be described below with reference to Figure 5. The raffinate 37 from the solvent extraction of Cu 16, is divided in two streams 120 and 121, in the ratio of 2/3 to 1/3, as in the embodiment of Figure 2. Stream 120 is recycled to atmospheric leaching 14, while stream 121 is neutralized 46, at a pH of about 4, and then subjected to a liquid / solid separation 48. Raffinate 63 from the solvent extraction of Cu 50, is subjected to neutralization 76 at a pH of 2 with limestone, and then subjected to a liquid / solid separation 78. The residue of solid gypsum 80 is discarded with the gypsum residue 53 from the liquid separation / solid 48. The liquid phase from the liquid / solid separation 78, is subjected, with the liquid phase from the liquid / solid separation 48, to the extraction with solvent 246, with a suitable zinc extractant, such as DEHPA, for produce an organic loaded with Zn. This organic stream is carefully purified from Cu, Co, Cd, Cl, etc., before being separated with the acid spent from a subsequent zinc electrolytic extraction step. Purification can be effected by purification of the charged organic using an aqueous solution of ZnSO4. The raffinate is recycled to the oxidation step under pressure 12. The raffinate from the solvent extraction of Zn is recycled to the oxidation step under pressure 12. Any traces of DEHPA left in the raffinate, will be subjected to the highly oxidizing conditions of the step of oxidation under pressure 12, to counteract the contamination of the copper extractant LIXMR with DEHPA. It has been found that contamination of the LIXMR reagent with DEHPA results in deterioration of the first reagent. The recovery of precious metals, such as gold and silver, will now be described with reference to Figures 5A and B. This process involves the treatment of the final waste stream 35 of Figures 1, 2, and 4.

The precious metals are not leached during the oxidation step under pressure 12, but remain in the remaining solid residue after the atmospheric leaching step 14. In order to facilitate the recovery of the precious metal, evaporation 22 is carried out from the step of oxidation under pressure 12 in two stages. The first stage is at a temperature slightly higher than the freezing point of the elemental sulfur, ie, from about 120 ° C to 130 ° C, with a corresponding vapor pressure of about 10 to 20 psig (7-14 kPa). The preference process is carried out in a continuous mode, the retention time of the first evaporation stage being approximately 10 to 30 minutes. The second evaporation stage is at atmospheric pressure, and from about 90 ° C to 100 ° C, with a retention time of at least 10 minutes again. This allows the elemental sulfur, which is still molten in the first evaporation stage, to become one of the solid phases, such as the stable orthorhombic crystalline phase. This procedure facilitates the production of clean crystals of elemental sulfur, which is important for the recovery of precious metals from the residue of leaching. The leaching residue 35, now produced by atmospheric leaching stage 14, contains, in addition to the precious metals, hematite, crystalline elemental sulfur, unreacted sulfides (pyrite), and any additional products that may result from the particular concentrate that is You are using, for example, gypsum and iron hydroxides. It is believed that the gold in residue 35 is largely untouched by the process, and more likely is in the native state. However, the silver is oxidized in the oxidation step under pressure 12, and is probably present as a silver salt, such as silver chloride or silver sulfate. It has been found that conventional cyanidation does not leach the gold well from residue 35. It is believed that this is due to the encapsulation of gold into mineral particles, such as pyrite. However, gold can be released by the oxidation under pressure of these minerals, referred to as "total oxidative leaching" or "pyrite leaching". In order to effect this leaching without oxidizing the elemental sulfur also contained in the residue 35, the process comprises the step of removing as much of the elemental sulfur as possible. First, by virtue of the evaporation of two stages, sulfur crystals of good quality are produced. Second, the leach residue 35 is subjected to foam flotation 402, to produce a sulfur-rich flotation concentrate 404, and a sulfur-depleted flotation tail 406. The tail 406 is subjected to a solid / liquid separation 408. to produce a liquid that is recirculated to a conditioning tank 410 upstream of the flotation passage 402, and a solid 412 that is sent to the total oxidative leaching stage 414. The flotation concentrate 404 is filtered (416), dry to a low humidity, and then melt at a casting step 418 of about 130 ° C to 150 ° C, to produce a paste 420 of liquid sulfur and solid mineral particles. The paste 420 is filtered (422) to remove the liquid sulfur, which is then cooled (424) to produce an elemental sulfur product 426. The cooled sulfur can be subjected to an optional sulfur purification step 425 to remove the impurities , such as selenium and tellurium thereof. The solid residue from filtration 422 is subjected to a hot sulfur extraction step 428 at 90 ° C with kerosene or other suitable extractant, such as perchlorethylene. The resulting hot paste is filtered (430) to produce a low sulfur residue (less than 5 percent elemental sulfur) 432, which is sent to total oxidative leaching 414. The hot filtrate is cooled (434) to reduce the Sulfur solubility, producing crystalline S °, which is filtered in 436, to return the kerosene that is recycled to the sulfur extraction step 428. A test was carried out in which 100 grams of atmospheric leaching residue 14 containing 25.1 percent elemental sulfur (S °), and 3 percent sulfur, through flotation 402, smelting 418, and extraction 428. This produced 73.8 grams of a desulfurized residue (feed material for the total oxidation leaching 414) containing 1.9 percent S ° and 4.1 percent sulfur, that is, a total of 6 percent total sulfur. The desulfurized residue contained 5.9 percent of the elemental sulfur (S °) in the original leach residue, that is, 94.1 percent was recovered to a pure elemental sulfur product. Total oxidative leaching 414 is performed at approximately 200 ° C-220 ° C and at 50-150 psig (345-1035 kPa) of partial pressure of oxygen, sufficient to completely oxidize all sulfur and metal compounds to the highest valences , respectively. Therefore, all the sulfur and pyrite are oxidized to obtain sulfate. Oxidation occurs under acidic conditions, such as acid being produced at the site. The reaction is highly exothermic, and in general the desired operating temperature can be achieved even with a cold feed pulp, provided that there is sufficient fuel present as sulfur in the solid feed. Typically, about 6 to 10 percent of total sulfur will be sufficient with a normal percentage of solids in the feed stock. After total oxidative leaching 414, the paste is subjected to neutralization 437 at a pH of 2 to 3 with limestone, and then filtered (438) to produce a solid residue containing precious metals, and a filtrate that is generally acidic. , and which may contain base metal values, such as copper, which can be extracted by an optional solvent extraction step 440, and sent to the extraction circuit with main solvent. The resulting raffinate is recycled to total oxidation leaching 414, as indicated at 442. Prior to cyanidation 444, solids from filtration 438 can be subjected to an optional lime boiling step 443, to facilitate the recovery of silver during 444 cyanidation, by decomposing silver jarosite compounds formed during total oxidative leaching 414. Precious metals are in the solids remaining after filtration 438. Now that pyrite and other encapsulating minerals have decomposed in the original concentrate, precious metals are susceptible to cyanidation 444.

In cyanidation step 444, the solids are leached with NaCN under alkaline conditions. In order to do this, the solids are formed in a paste with a cyanide solution to form a paste with 30 to 40 percent solids. Additional NaCN and slaked lime are added, as required, to maintain a minimum NaCN concentration of from about 0.2 to about 0.5 grams / liter of NaCN, with a pH of about 10. The temperature is ambient, and normally about 4 to 8 hours of retention time in continuous operation mode. Both gold and silver are reported in a high yield to the cyanide solution, and are typically recovered through the established pulp carbon circuit process, where activated carbon is added to the cyanide paste to absorb the precious metals, without the need for filtration. Charged coal, now rich in precious metals, is separated by sifting (445), and the auger pulp is discarded to the tail. Charged coal is treated by established methods to recover the precious metal content through an electrolytic leaching / casting process (447). The product is usually Dore metal that contains both gold and silver, which is sent to a 449 gold refinery for the final separation of the gold from the silver. Auger coal from a carbon 451 regeneration step after recovery of precious metals is recycled to the 444 carbon pulp circuit. The overall recovery of precious metals through the total process is generally more than 90 percent. percent, and under optimal conditions, approaches 99 percent. A test was performed where the sulfurized residue was processed in a total oxidative leaching 414 at 220 ° C for 2 hours under an oxygen pressure, and then depressurized and cooled to room temperature. The resulting paste was neutralized to a pH of 10 with limestone, and then filtered. The filter cake was then leached with a cyanide solution under conventional conditions, to leach the gold and silver. The extraction of gold after total oxidative leaching 414 and cyanidation 444 was 97 percent, with only a consumption of 1.0 gram / ton of NaCN. In comparison, the extraction of gold on a residue that had not been oxidized in total oxidative leaching 414 was only 34 percent, and the cyanide consumption was extremely high at 19.0 kilograms of NaCN / ton.

Claims (24)

1. A process for the extraction of copper from a copper sulphide mineral or concentrate, which comprises the steps of: subjecting the mineral or concentrate, together with a source of sulphate or bisulfate ions, to an oxidation under pressure, at a temperature of about 115 ° C to about 160 ° C, in the presence of oxygen and of an acid solution containing halide ions, to obtain a resulting pressure oxidation paste; subjecting the pulp to a liquid / solid separation step, to obtain a resulting pressure oxidation solution, and a solid residue; and recovering the copper from the oxidation solution under pressure, or the solid residue; characterized in that the oxidation solution under pressure is recycled until oxidation under pressure, and the concentration of copper in the oxidation solution under pressure that is recycled, is maintained at a previously determined value.

2. A process according to claim 1, wherein the concentration of copper in the pressure oxidation solution that is recycled is maintained at a value in the range of 10 to 15 grams / liter.

3. A process according to claim 1, characterized in that the concentration of copper in the solution that is recycled is controlled by subjecting the solution to an extraction with copper solvent, before being recycled, to produce a copper solution and a acid raffinate, and raffinate is recycled.

4. A process according to claim 3, which further comprises the step of subjecting at least a portion of the raffinate to a partial neutralization to lower the concentration of hydrogen ions in the raffinate, before recycling the raffinate.

5. A process according to claim 4, wherein the partial neutralization is carried out with limestone.

6. A process according to claim 4 or 5, wherein the concentration of copper in the oxidation solution under pressure resulting from the oxidation under pressure, is controlled by controlling the concentration of hydrogen ions in the raffinate that is is recycling, by means of partial neutralization of the raffinate.

7. A process according to claim 1, wherein the oxidation under pressure is carried out in the presence of a surfactant.

A process according to claim 1, which further comprises the step of lowering the viscosity of the liquid sulfur in the oxidation under pressure, by introducing a surfactant into the oxidation under pressure.

9. A process according to claim 8, wherein the surfactant is added to the oxidation solution under pressure which is recycled towards oxidation under pressure.

10. A process according to claim 9, wherein the surfactant is added in an amount of about 0.1 grams / liter to about 3 grams / liter.

11. A process according to any of claims 7 to 10, wherein the surfactant is selected from lignin sulfonate, quebracho, and its derivatives.

12. A process according to claim 1, wherein the halide is chloride or bromide.

13. A process according to claim 12, wherein the halide is chloride, and the concentration of the chloride ions in the acid solution is about 12 grams / liter.

A process according to claim 3, wherein the ore or concentrate also contains zinc together with copper, and further comprising the step of subjecting the raffinate from extraction with copper solvent to a solvent extraction, before recycle the raffinate, with a zinc extractant, to obtain a concentrated zinc solution for electrolytic extraction.

15. A process according to claim 14, which further comprises the step of subjecting the raffinate to neutralization before extraction with zinc solvent.

16. A process according to the preceding claims, wherein the ore or concentrate also contains precious metals, and which further comprises the steps of: removing the elemental sulfur from the solid residue resulting from the oxidation under pressure, to obtain a low residue in elemental sulfur; and subject the low residue in elemental sulfur to an oxidative leaching, to oxidize the elemental sulfur and the sulfide minerals present in the low sulfur residue, to produce a residue for the extraction of the precious metals from it.

A process according to claim 16, wherein the removal of sulfur comprises the steps of: subjecting the solid residue resulting from pressure oxidation, to foam flotation, to produce a sulfur-rich flotation concentrate, and a floating tail depleted of sulfur; and subjecting the flotation concentrate to sulfur extraction, with a sulfur extractant, to produce the low residue in elemental sulfur.

18. A process according to claim 17, wherein the sulfur-depleted flotation tail is subjected to a solid / liquid separation, to produce a liquid, which is recirculated towards the foam flotation, and a solid which is subjected to oxidative leaching.

19. A process according to claim 17, wherein the extraction of sulfur from the flotation concentrate is carried out at a temperature of about 90 ° C to 150 ° C.

20. A process according to claim 19, wherein the sulfur extractant is selected from the group consisting of kerosene and perchlorethylene.

21. A process according to claim 16, wherein the oxidative leaching is carried out at a temperature of about 200 ° C to 220 ° C, and at a partial oxygen pressure of about 500 to 1200 kPa under acidic conditions.

22. A process according to claim 16, wherein, prior to sulfur removal, the pressure oxidation paste is evaporated at atmospheric pressure in a two-stage evaporation, wherein the first stage is at a higher temperature than the freezing point of elemental sulfur.

23. A process according to claim 22, wherein the evaporation of the first stage is at a temperature of about 120 ° C to 130 ° C, and at a vapor pressure of about 170 kPa to about 240 kPa.

24. A process according to claim 22 or claim 23, wherein the evaporation of the second stage is at a temperature of about 90 ° C to 100 ° C, and at atmospheric pressure.