MXPA97006678A - Procedure for the electroextraction of mata de co - Google Patents

Abstract

A sulfur dioxide-free process for the production of high purity metallic copper from copper matte, where the copper matte is leached under oxidation conditions in a copper electrolyte electrolyte leaching assembly including one or more leaching reactors to produce a copper-rich electrolyte, and a copper cathode is produced in an electro-extraction assembly, which is physically uncoupled from the leach assembly and may include one or more cells. The procedure operates at ambient pressure and at temperatures lower than the boiling point

Description

PROCEDURE FOR ELECTROEXTRACCION DE MATA DE COPBRE FIELD OF THE INVENTION The present invention is a process in the fields of hydrometallurgy and electrometallurgy. Specifically, it relates to a sulfur dioxide-free process for continuous production of high purity metallic copper from copper matte in a sulfate-containing medium through leaching and electroextraction methods under ambient pressure and at a temperature below Boiling point.

BACKGROUND OF THE INVENTION Currently, commercial copper production is predominantly based on pyrometallurgical methods involving smelting, conversion, and electrorefining and to a lesser degree in hydrometallurgical methods involving leaching, solvent extraction and electroextraction. One of the serious problems inherent in those pyrometallurgical processes with respect to sulfide minerals that carry copper, is the production of large volumes of fugitive gases including sulfur dioxide gas. Particularly, the conversion and ignition refining operations result in the release of large volumes of these gases. The amount of sulfur dioxide gas discharged into the atmosphere has been strictly regulated by both federal and local governments. The regulations for the emission of sulfur dioxide have become much stricter, and the industry must both observe the regulations and incur associated costs, incur heavy penalties and costly, as in potentially facing plant strikes. In order to eliminate or minimize the emission of sulfur dioxide gas, procedures have been studied in the past, aimed at the conversion of sulfur sulfur to sulphate or elemental sulfur. The sulfate conversion route is expensive and can be technically difficult. A more desirable view is to convert sulfides to elemental sulfur, due to the highly salable value of elemental sulfur. In the last two decades, many researchers addressed this problem of SO2, focusing on the conversion of sulfur to elemental sulfur. As a result of various efforts, a number of hydrometallurgical processes were developed, such as the ELECTROSLURRY, Intec, Dextec and CLEAR processes, in attempts to recover sulfur sulfur in its elemental form. The Envirotech ELECTROSLURRY procedure, described in the U.S.A. Nos. 4,096,053 and 4,066,520, is a hydrometallurgical process based on sulphate, which is capable of recovering copper from the chalcopyrite concentrate, chalcocite concentrates, particles from the foundry fumes, and copper from cementation, where the sulfides are thrown towards the leaching residue. In the Envirotech process, leaching and electroextraction are carried out simultaneously in a mud electrolyte, and the catholyte and the anolyte are not separated with any type of diaphragm. This procedure provides a simple cell design, high current density electroextraction with no oxygen development and all unit operations under ambient pressure and at a temperature below the boiling point. However, in this process, the feed materials require a pre-treatment to remove any water-soluble impurities and to make the material treatable to be used as a feed for subsequent electroextraction. A substantial weakness of this procedure is in surprisingly high sulfur content in the copper cathode, due to the entrapment of sulfides and elemental sulfur. The Intec copper process is a completely hydrometallurgical process using a strong sodium chloride medium, where the copper sulphide is leached through the cupric ion and at an atmospheric pressure and at a temperature of 100 ° C to solubilize copper as cuprous ion to the solution. At the same time, sulfides and iron are thrown to the leaching residue as elemental sulfur and geotite, respectively. The leaching is followed by the purification of the solution and high current density extraction in a diaphragm cell, using a copper foil with holes as the cathode substance and a titanium mesh coated with RuO2 / lrO2 as an anode for produce copper granules. Copper granules may require further processing to be in an acceptable form for commercial transactions (ie, cathode, bar, wire). One of the most undesirable requirements for this procedure is the use of expensive ion exchange membranes such as diaphragms for electroextraction cells. Due to the potentially rapid growth of dendrites at the cathode in a chloride electrolyte, and due to the high cost and fragility of the ion exchange membranes, the electroextraction cells for this procedure can be very complicated. In addition, leaching solutions are highly sensitive to air, since oxygen in the air will oxidize cuprous ions. This can result in an added cost and complexity in the design and operation of the electroextraction cell. The Dextec Copper Procedure, described in the patent of E.U.A. No. 4,061,552, is similar to the Intec procedure, in some aspects. Both procedures are chlorine-based, operate under atmospheric pressure and at temperatures below 100 ° C, and provide high current density extraction to recover copper as granules or dust. However, in the Dextec Copper Procedure, both leaching and electroextraction occur simultaneously in a diaphragm cell, the purification of electrolyte taking place after leaching. The chalcopyrite particles are suspended in the mud anolyte and leached chemically and / or electrochemically to form cuprous ions, elemental sulfur and ferrous iron, which is more oxidized through oxygen and precipitated as Fe2O3. Concentrations with a high chlorine content are desirable to stabilize the cuprous ions, and a stream of air bubbles is necessary to keep the solid particles in suspension and favor the precipitation of Fe2O3. The mud anolyte is separated from the catholyte through a polypropylene filter cloth, through which the clean anolyte rich in cuprous ions is transferred to the catholyte. The copper is recovered at the cathode as copper powder, which is removed by mechanical vibration of the cathode at a frequency of 5 seconds for every 15 minutes. The Dextec Copper Procedure suffers from a complicated electroextraction cell design and a very strong loss of silver. The loss of silver is generally attributed to the fact that silver dissolves easily in the chlorine medium and then co-deposits with the copper in a copper cathode. In addition, the Dextec process requires a continuous flow of electrolyte from the anode compartment to the cathode compartment in order to avoid a substantial reduction in actual efficiency. The operating conditions are difficult to control, and the mass balance between copper removal in the catholyte and copper enrichment in the anolyte is difficult to maintain. The fatal problem with the copper product of this process is that it is not adequately pure and may contain unacceptable levels of silver, antimony and bismuth. The product will require further processing to be in an acceptable form and purity for commercial transactions. The CLEAR procedure of Duval has many aspects common to the most recent Intec procedure, since both procedures are based on a chlorine medium, both involve leaching of multiple stages with the chalcopyrite being converted to cuprous, both result in elemental sulfur and precipitation of ferric hydroxide, both have diaphragm electroextraction cells, and both produce copper granules as a final product. The CLEAR process also has the disadvantage of producing impure copper granules with high silver levels in the copper granules. Another disadvantage of the CLEAR process is the use of an autoclave for its leaching of 2a. stage under a pressure as high as -3.5 atm and at a temperature of up to 150 ° C.

OBJECTS OF THE INVENTION It is an object of the present invention to provide a simple, environmentally friendly and cost-effective hydrometallurgical process for replacing copper matte conversion and subsequent electrorefining procedures, which are normally associated with the pyro-metallurgical production of copper from the ore. sulfide carrying copper.

It is another object of the present invention to provide a hydrometallurgical process to be used in conjunction with the smelting of copper bearing sulfide minerals, where elemental sulfur is recovered as a by-product, thus eliminating the production of sulfur dioxide gas and fugitive emissions. associated with mat conversion and refining by ignition. Briefly, the present invention provides a process for the production of high purity metallic copper from copper matte comprising the steps of leaching the copper mat under oxidation conditions to obtain a pregnant leaching solution in a leach assembly including one or more leaching tanks, and electrolixing the copper of the pregnant leaching solution in one or more electroextraction cells, which are physically decoupled from the leaching assembly. Other objects and advantages will be apparent from the following description.

BRIEF DESCRIPTION OF THE DRAWINGS Figure 1 is a schematic illustration of a complete procedure according to the present invention. Figure 2 is a schematic illustration of a multi-stage leaching operation in a preferred embodiment of the present invention. Figure 3 is a schematic illustration of a multi-cell electroextraction operation in a preferred embodiment of the present invention.

DETAILED DESCRIPTION OF THE INVENTION The present invention provides a hydrometallurgical alternative to copper matte conversion processes and subsequent electrorefining of copper, which traditionally follow the smelting of copper bearing sulfide ores. The invention avoids the cost and harmful environmental emissions associated with the conversion of copper matte and refining by ignition. This invention also provides a simpler process chemistry, a simpler full-flow diagram, and a purer product than prior art hydrometallurgical attempts. The total procedure can be divided into eight stages: 1) crushing and grinding; 2) leaching; 3) solid-liquid separation; 4) electroextraction; 5) Bleeding current electrolyte purification; 6) removal of elemental sulfur; 7) recirculation of the mat with incomplete reaction; and 8) recovery of platinum group metal (PGM) and precious metals. The process is novel in several aspects, particularly in the stages of leaching and electroextraction. The feedstock is copper matte (some form of Cu Fe S), the final product of a casting process. For example, a sample of high grade copper mat treated with the present process, may contain about 5% chalcocite (Cu 2 S), about 3.3% magnetite (Fe 3 O 4), minor components such as iron, lead, nickel sulphides , bismuth, arsenic and antimony, etc., and a minute amount of precious metals, although other matte compositions may be used in the present invention. Normally, the mat is subsequently converted to remove the sulfur and iron. However, the present invention provides for the removal of these elements without the normal operation of conversion. Referring to the complete flow chart depicted in Figure 1, the copper matte is transported through line 11 to the crusher and mill 1, where the copper matte is crushed and milled to preferably 80% passing through. a mesh of 200, most preferably 80% through a mesh of 325 in particle size. A small particle size promotes rapid reaction and high copper recovery in the subsequent leaching step. Dry milling of this copper matte can be achieved in a ball mill or bar mill coupled with an overload cyclone. Nevertheless, other methods can be used, such as wet milling, which prevents the generation of heat and oxidation of copper sulphide, which can occur during dry milling. After grinding, the copper matte is fed through line 12 directly to leach to tank assembly 2, where Cu2S reacts with Fe3 + to form Cu +2 and S °, as represented by the Cu2S + reaction 4Fe3 + = 2Cu2 + + 4Fe + S °. magnetite (Fe3O4) and PGM metals, as well as gold and silver do not dissolve in the leach tank assembly 2, and therefore remain in the leach residue. After solid-liquid separation using a thickener or any other method known in the art, the pregnant leaching solution is passed via line 13 to a filter 3, which removes any possible transport of minute particles of reaction matte. incomplete and elemental sulfur, in order to secure a cleaning electrolyte via line 14 to the electroextraction cell assembly 4. The filter cake, filter 3, is returned through line 20 back to the leach tank 2. Within the electroextraction cell assembly 4, the cupric ions are reduced to densify the metallic copper on the cathodes, which are harvested via lines 41, 42, 43, and at the same time, the ferrous ions are oxidized to ions ferric on anodes in the assembly of cell 4, without presenting any development of hydrogen or oxygen on the cathode or anode. The regenerated ferric ions can be transferred via line 15 to the leach tank assembly 2 to dissolve more Cu2S copper in the copper matte. Although not necessary in the present invention, it is optional to remove, through line 16, an electrolyte bleed current to control the development of levels of impurities such as As, Bi, Sb, etc., in the electrolyte. After the impurities have been removed from the bleed stream in the purification means 5, using a technique such as ion exchange or solvent extraction, the purified bleed stream electrolyte is returned via lines 17 and 14 back to the electroextraction assembly 4. The removed impurities are discarded via line 51. A bleed stream can be removed at another point in the process, without departing from the present invention. The optionally purified bleed stream can be returned to the procedure. The leached residue generated in the leach tank assembly 2 is discharged through line 26 to unit 6 for the removal of elemental sulfur. The elemental sulfur can be recovered using a process such as heat fusion and filtration, or leaching of NH3-H2S, etc. Line 61 carries recovered elemental sulfur, which can be sold as a commercial product or can be recovered to SO2 or sulfuric acid. The remaining residue is also carried via line 27 to unit 7 for recovery of copper matte with incomplete reaction. This copper bush with incomplete reaction has undergone a partial decomposition and is greatly in the form of covellite (CuS), which can be recirculated via line 72 to the leach tank assembly 2.

The residue of unit 7 contains all PGM metals originally present, gold, silver, insoluble impurities such as magnetite (Fe3O) and gangue. This waste is fed via line 28 to unit 8 for the recovery of precious metals. The recovered precious metals are discharged via line 81 for additional separation, if necessary, or for sale as a mixture of impurities. The waste via line 82, after recovery of the precious metal, is mainly composed of unwanted materials and can be discarded. The leach tank assembly 2 of the present invention can be constructed so that leaching occurs either in one stage or in multiple stages in the process. In any case, the leaching occurs under ambient pressure and at temperatures below the boiling point. Through leaching, Cu2 + ions are released from Cu2S through chemical attack of Fe3 +, as expressed by the reaction Cu2S + 4Fe3 + = 2Cu2 + + S ° + 4Fe2 +. However, the actual leaching mechanism of the copper matte is staggered. Depending on the oxidation condition and especially the concentration of ferric ions or redox potential of the leach mud, leaching of the copper matte may end or may stop prematurely with the formation of intermediates. For example, leaching can result in digenite (Cut 8S) and covellite (CuS) intermediates, if the driving force is not large enough. It is believed that the decomposition of the copper matte is achieved through the following three steps: (a) 5Cu2S + 2Fe3 +? 5Cu? 8S + Cu2 + + Fe2 +; (2) 5Cu, .ßS + 8Fe3 +? 5CuS + 4Cu2 + + 8Fe2 +; and (3) 5CuS + 10Fe3 +? 5Cu2 + + 5S ° + 10Fe2 +. The first step of decomposition of Cu2S? Cu1 8S has the last resistance and can proceed very easily, even at a ferric concentration lower than 0.3 g / l. The second step of decomposition of Cu? .8S? CuS also has little resistance and can easily proceed to a ferric concentration in the order of 0.3 g / l. The third step of decomposition, CuS - > Cu2 + is the most difficult and probably will not be presented to acceptable regimens unless the iron concentration is greater than 3 g / l. In the leaching step of the present invention, preferably only the copper is dissolved in the Cu2 + form in the solution, and the elemental sulfur, PGM gold and silver metals are left in the leach residue. The additional oxidation of elemental sulfur to sulfate is negligible in this system, since the leaching sludge is highly acidic and is only moderately oxidizing. Due to the weak oxidation energy of ferric iron and the absence of any complexing agent, the PGM, gold and silver metals do not dissolve during copper leaching. Due to the unique mechanism of decomposition of the copper matte, it is preferred to have multi-stage flow-to-stream leaching. Multistage leaching, and preferably leaching of stages is preferred to promote high extraction of copper leaching. A preferred embodiment of the present invention is conducted as a continuous process employing a three stage leach assembly, as shown in Figure 2. In this preferred embodiment, the finely ground copper matte undergoes three step leaching under pressure. environment and at a temperature below the boiling point. The copper matte is leached in three leach tanks 21L, 22L, and 23L via the countercurrent flow. In the case of three stage leaching, it is believed that the reactions (1) and (2), mentioned above, occur in the leach tank of the 1st. stage 21L, reactions (2) and (3) are presented in the leach tank of 2a. step 22L, and only reaction (3) is presented in the leach tank of 3a. stage 23L. Ferric concentration is highest in the leach tank of 3a. stage 23L, and the lowest in the 1a leaching tank. stage 21L. Referring to Figure 2, the entire three-stage leach assembly includes the following components, inputs and outputs. The finely ground copper matte is added to the leach assembly via line 12 to the 1a leach tank. stage 21L. Copper matte of incomplete reaction, of the total process, is added to the leach assembly via line 72 to the second stage leach tank 22L. The regenerated ferrous ions from the electroextraction cells enter the leach assembly being fed via line 15 to the leach tank of 3a. stage 23L. The solid-liquid separation means, such as the thickeners 21S, 22S and 23S, separate a surplus and a lower flow from each of the leach tanks in the leach assembly. Line 13 represents the exit of clean pregnant leach solution from the leaching assembly to the electroextraction cells. Theoretically, according to the total leaching reaction (Cu2S + 4Fe3 + = 2Cu2 + + 4Fe2 + + S °), one mole of ferric produces only half a mol of cupric. Therefore, from the leaching point of view only, a high ferric concentration in the inlet stream 15 is desirable to the leach tank of 3a. stage 23L, since it will increase the rate of leaching of the copper matte and reduce the amounts of sludge and clear solutions, which will be managed within the same leach circuit and the amounts of inlet current via line 15 to the leach tank of 3a. stage 23L and the output current via line 13 of the thickener 1a. stage 21S. However, since the performance of the electroextraction varies inversely with the ferric concentration in the electrolyte, the ferric concentration in the input stream via the line 15 is determined entirely by the efficiency of the electroextraction. Ferric concentrations in the electrolyte less than 10 g / l are preferred to achieve a reasonably high current efficiency. To facilitate the above leaching reactions, the copper matte particles are preferably 200 mesh, most preferably 325 mesh and should be well suspended in the leach solution, possibly using mechanical agitation, recirculation or inert gas spraying. Referring to Figure 2, in the first stage leach tank 21L, the fresh finely ground copper matte, which is fed via line 12, and a small amount of filter cake containing incomplete reaction copper matte, which is added via line 20, are mixed with part of the lower flow stream via lines 21c and 21d from the thickener of 1a. stage 21S and the excess stream via line 22B from the thickener 2a. stage 22S. The feed rate of the fresh copper mat via line 12 is calculated according to the velocity of copper on the cathode in the electrowinning cells and the percentage of extraction of the copper during the leaching, the coppering speed must be equal to the total rate of copper leaching. The return of part of the lower flow stream via lines 21c and 21d from the thickener of 1a is preferable. stage 21S back to the 1a leaching tank. stage 21L, as each step of the ferrous solution passes through the leach tank and leaches only a small amount of copper from the copper matte. Each step results in a small increase in the concentration of Cu2 + in solution of the order of 1 to 5 g / l, which depends on the ferric concentration in the inlet solution via line 15. The residence time for the solid phase it depends on the speed of copper, the density of the pulp, the copper content in the original copper mat, the copper recovery required, the size of the leach reactor for each stage, and the number of leaching stages.

The residence time for the liquid phase depends on the cobrising speed, the redox potential of the electrolyte in the electroextraction cell, the redox potential of the sludge in the leach reactor of 1a. stage, the size of the leach reactor, and the number of leaching stages. The discharge sludge via line 21a from the leach tank of 1a. stage 21L to the first stage thickener 21S contains mainly CuS, Cu1 8S (in a small amount), FeSO4, Fe2 (SO) 3 (<0.3 g / l Fe3 +), CuSO, H2SO4 and small amounts of additives and impurities. Minimum elemental sulfur is formed in the 1a leach tank. stage 21L. The pulp density can vary from 5 to 70%. The pulp density is preferably about 40% for each stage. The recirculation process of the present invention ensures that the leaching can be carried out at a high pulp density, while minimizing the necessary size of the leach tanks. The excess stream via line 13 from the thickener of 1a. stage 21S contains CuSO4 (20 to 50 g / l Cu2 +), H2SO (20 to 200 g / l), FeSO4 (20 to 50 g / l Fe +), Fe2 (SO4) 3 (<; 0.3 g / l Fe3 +), small amounts of additives and trace amounts of impurities. Before this solution is pumped into the electroextraction cells, an extra filtration may be preferable to remove any transport of small solid particles, in particular, elemental sulfur. After filtration, this pregnant solution is introduced into the electroextraction cells.

The lower flow stream 21c of the thickener 1a. Stage 21S must contain very little solution. Otherwise, the leaching solution contained in the lower flow, filled with Fe3 +, is transferred in an undesirable direction to the next stage of leaching, thus reducing the rate of leaching and effecting the termination of the decomposition of the matte. copper. The lower flow stream is further divided into two sub-currents via lines 21d and 21e. The sub-stream via line 21e must carry a quantity of solids as required to perfect leaching while maintaining an apparent, specific pulp density in the 21L leach tank. The sub-stream via I line 21e only represents a small portion of the lower flow stream of line 21c from the thickener of 1a. stage 21S. A large portion of the lower flow stream is recirculated internally via line 21d back to the 21L leach tank. The leaching of 2a. stage is quite similar to leaching from 1a. stage. The apparent pulp temperature and density is the 2a leach tank. Stage 22L are similar to those of the 1a leach tank. stage 21L. However, in the 2a leaching tank. stage 22L, the ferric content is of the order of 3-4 g / l, which is more than 10 times greater than that in the 1a leaching tank. stage 21L. The Cu2S in the copper matte has been converted to Cu2 +, covelite (CuS) and also some elemental sulfur. The content of elemental sulfur in the solid phase can reach as much as 30 to 60% in the leach tank of 2a. stage. Also, it is preferred that the 2a leach tank. stage, instead of the 1a leaching tank. stage, receive via line 72, the recovered copper matte, without completely reacting, which is mainly in the form of covelite (CuS). Since the ferric concentration is higher in the 2a leach tank. stage compared to the 1a leach tank. stage, the decomposition of covelite to Cu + 2 can occur at an acceptable rate in the leach tank of 2a. stage. The sludge in the leach tank of 2a. stage 22L is discharged via line 22a to the thickener of 2a. stage 22S. As with the thickener of 1a. step 21S, the lower flow stream via the line 22c of the thickener 2a. Stage 22S must contain very little solution. The lower flow stream via line 22c is divided into sub-streams via lines 22d and 22e. The sub-stream via line 22e represents a small portion of the lower flow stream of line 22c, and must coincide equivalently with the sub-stream via line 21e. In this way, no solids develop in the system, the lower flow stream from the 2a thickener. Stage 22S goes to the 1a leach tank. stage 21L, since the ferric concentration in this solution is still too high to return to the electroextraction cell. The leaching of 3a. stage is similar to leaching from 1a. stage and 2a. stage in some aspects. However, the spent electrolyte from the electroextraction cells is added to the leach tank of 3a. stage 23L via line 15. The mud discharged via line 23a from the leach tank of 3a. stage 23L, it contains more than 75% by weight of elemental sulfur in the solid phase, and the total copper extraction reaches over 95%. The temperature in the leach tank of 3a. Stage 23L is similar to that in 1a leach tanks. stage and 2a. stage, 21L and 22L; however, the ferric content in the 23L leach tank is the highest among the leach tanks 21L, 22L and 23L, ranging from 4 to 10 g / l. The lower flow stream via line 23c of the thickener 3a. Stage 23S is also divided into two sub-currents via lines 23d and 23e. The sub-stream via the line 23e is advanced further towards the filter 24 to remove the excess solution contained in the lower flow. The solution recovered in the filter 24 is returned via lines 24a and 23f back to the leach tank of 3a. stage 23L. The filter cake obtained in the filter 24 is sent via line 24b to washing means 25. Two or three displacement washes may be suitable for releasing the solution trapped within the filter cake. During practice, the filter 24 and the washing means 25 can be combined in a single processing unit. To avoid precipitation of ferric hydroxide, the wash water added via line 29 is preferably slightly acidic. Part of the washing solution of the washing means 25 can be added to the leach tank of 3a. stage 23L via lines 25a and 23f. Alternatively, the wash solution of line 25a can be added to the leach tanks of 1a, 2a, and 3a. stage, 21L, 22L and 23L. The acid is introduced to compensate for acid loss, which occurs when electrowinning cells, leaching tanks and thickeners are exposed to open air. A certain amount of acid will be lost due to the absorption of oxygen followed by its reaction with the copper or ferrous matte. However, only a small amount of acid is preferably added via line 29 in order to significantly alter the acid balance throughout the process. Since water loss can occur through evaporation, if electrowinning cells, leach tanks and thickeners are not sealed, water introduced via line 29 must be added in order not to alter the equilibrium of the water. water throughout the system. The filter cake obtained after washing and draining at 25, is accelerated towards the sulfur removal means 6 (FIG. 1) - The separation of solid-liquid in 21S, 22S and 23S can be carried out using commercially available equipment such as a thickener or cyclone, etc. The separation of solid and liquid in leaching against current must be efficient; otherwise, the solids will be carried by line 23b from the thickener of 3a. Stage 23S to the leach tank of 2a. stage 22L, and via line 22b from the thickener of 2a. stage 22S to the leach tank of 1a. stage 21L. This solid phase movement is undesirable because, in the extreme case, all solids can go into the leach circuit of 1a. stage. Optionally, a filter or centrifuge can be installed between each stage of leaching to ensure that the flow of the solid phase is in the correct direction. To facilitate the sedimentation of the solid phase in the 21S, 22S and 23S thickeners, a small amount of flocculant can be added to the thickeners. The density of sludge pulp in the leach tanks can vary in the range of 5 to 70% (w / v), preferably to 40% (w / v). The pulp density is determined partly by the rate of leaching, and partly by the efficiency of the solid-liquid separation means. The leaching rate is preferably fast enough to reduce the ferric content in the inlet stream coming from line 15 down to almost zero in the outlet stream of line 13. A desirable leaching rate is achieved at temperatures on the scale of 60 ° C to the boiling point, and 80 ° C is preferred. The presence of a small amount of chlorine in the leaching sludge is also completely beneficial for the leaching rate. The amount of chlorine addition can be in the range of 0 to 200 ppm, preferably 75 ppm, which is again restricted through the performance of the electroextraction. However, the process herein will still work at much higher chlorine concentrations. In the present invention, the electroextraction can be performed in one or more electroextraction cells. Multiple cells can be used in parallel or in series. When a solution is recirculated, the multi-cell electroextraction promotes a higher total current efficiency, which is defined as the average current efficiency calculated based on the current efficiencies and the copper cathode weights in each cell. electroextraction. Figure 3 represents the electroextraction in a series of 3 cells, which can be used in the present invention. The polished or clear pregnant leaching solution, which does not contain any solid particles and an insignificant amount of ferric ions, is fed via line 14 as the feed for the electroextraction cell of 1a. stage 4a. The leftover electrolyte from the electroextraction cell of 1a. Stage 4a is used as the feed for the 2a electroextraction cell. step 4b, that is, this excess electrolyte is transferred via line 44 and added to the electroextraction cell 2a. Stage 4b. Similarly, the leftover electrolyte from the electroextraction cell of 2a. Stage 4b is used as the power for the 3a electroextraction cell. stage 4a. via line 45. The designs of the multiple electroextraction cells 4a, 4b and 4c can be exactly the same; however, they operate in different conditions, mainly differing in the selection of current density and recirculation speed of the electrolyte. With reference to Figures 1, 2 and 3, the excess electrolyte of the electroextraction cell of 3a. Stage 4c, is relatively high in ferric iron concentration on the scale of 3 to 10 g / l, it can be returned via line 15 back to the leach tank of 3a. stage 23L in the three stage leaching circuit. Copper cathodes are harvested via lines 41, 42 and 43, respectively, from the electroextraction cells 4a, 4b and 4c. Although the visual appearances of the copper cathodes produced in the electroextraction cells 4a, 4b and 4c may vary to a certain degree, their chemical compositions remain consistent on the same scale with each other. The purity of the cathodes produced in this method satisfies the specifications of ASTM Grade 1. In each electroextraction cell, preferably multiple cathodes and anodes are used with a gap between an anode and an adjacent cathode of the order of 1.27 to 5.08 cm, preferably 5.08 cm from the center. A suitable cathode substrate is titanium or another type of 316 stainless steel, and a suitable anode material is the Eltech's Activated Lead Electrode (ALE) anode (Eltech Activated Lead Electrode) or titanium anode coated with a precious metal oxide. (like RhO2 / lrO2) also known as a Dimensionally Stable Anode (DSA) (Dimensionally Stable Anode). The lead and lead alloy anodes used in conventional copper electroextraction, such as Pb-Cu-Sn-Ca, Pb-Ag and Pb-Sb-As, will also work; however, these lead anodes give a superior overpotential for ferrous oxidation.

The desirable cathode reaction in the reduction of cupric ion to metallic copper (Cu2 + + 2e = Cu). Hydrogen gas is not developed in the cathode of the present invention; however, the reduction of ferric to ferrous is thermodynamically favorable, and thus always occurs at the cathode along with the reduction of cupric ions. 3 + This side reaction of Fe Fe on the cathode forces the electroextraction of copper to run at a high current density in order to obtain a reasonably high current efficiency. This is due to the percentage of current consumed by the reduction of the ferric going down to a high current density, the iron reduction on the cathode being controlled by the mass transfer of Fe3 + to the cathode surface. This is preferable, since operating costs and capital are reduced by high productivity. With respect to the cathode reaction, the preferred deposit should be a coherent, soft copper cathode, rather than a cathode or dendritic granule. It has been found experimentally that in order to obtain a coherent copper cathode, additives (such as glue, thiourea, chlorine, etc.) are beneficial. In addition, the current density of the cathode must be below the limiting current density of the cupric reduction on the cathode. The chemical attack of the copper deposited on the cathode through the ferric ion in the electrolyte can be carried out in the present invention. This chemical attack through the ferric ion, if moderate, is useful, on the one hand to eliminate the formation of any dendrites or protrusions on the surface of the cathode and thus presents a smooth effect, but, on the other hand, if this chemical attack is too strong, excess amounts of deposited copper will be dissolved in the electrolyte resulting in a considerable loss of current efficiency. Since the redox potential is directly related to the ferric concentration, the ferric concentration in the electrolyte can be controlled online by checking the redox potential of the electrolyte. An appropriate ferric concentration must be selected in order to perfect the leaching of the copper out of the copper mat allowing still the production of a soft cathode, consistent with a reduced chemical attack. The desirable anode reaction is the oxidation of ferrous to ferric (Fe2 + = Fe3 + + e). Since the anode carries positive charges, the Fe2 + anodes are far from the anode surface due to the repulsive nature between the positive charges. Therefore, the ion migration of the Fe2 + ions provides a negative contribution to the total mass transfer of Fe2 + towards the anode surface. In such circumstance, the transfer mass of the Fe2 + ions is based on both diffusion and convection. The Fe2 + ions must be available on the anode surface to avoid the I anode competition reaction, that is, evolution of oxygen. The overpotential of the anode for ferrous to ferric oxidation should be minimized. The evolution of oxygen over the anode is preferably minimized or avoided in the present invention, since any evolution of gas will cause the formation of acid vapor and correspondingly deteriorate the environment of a tank housing. In addition, when oxygen gas develops, due to its slow reaction with ferrous (4Fe2 + + O2 + 4H + = 4Fe3 ++ 2H2O) in the electrolyte, acid was generated due to the evolution of oxygen (2H2O = 02 + 4H + + 4e), oxygen it is not completely consumed and the Fe3 + ions are not generated equivalently from silver copper on the cathode in the passage of oxygen gas from the surface of the anode to the open air. Thus, when this occurs, most of the oxygen gas developed at the cathode simply escapes into the air. Consequently, the Cu2 + ions will be eliminated and the acid will develop in the electrolyte. To prevent the evolution of oxygen at the anode and provide a coherent electroextracted cathode deposit at a high current efficiency, it is preferred to maintain a high recirculation rate. The upper limit of the recirculation rate is partially determined by the excessive loss of current efficiency due to the reduction of Fe3 + and its chemical etching of the copper cathode, and partly by the highly associated pumping cost. In a preferred embodiment of the present invention, as shown in Figure 3, recirculation of the electrolyte is provided through the pumps 47, 48 and 49 with a high volume of flow velocity and lower head. To reduce the cost of pumping, the electrowinning cells can be built higher, if the electrolyte must travel vertically, or they can be built longer, if the electrolyte should travel horizontally. The preferable operating conditions for 3-cell electroextraction, as shown in Figure 3, are current density of 500 to 1,500 A / m2, temperature of 60 to 90 ° C (although it is also acceptable up to the boiling point) , Fe3 + concentration from 1 to 10 g / l, Fe2 + concentration from 20 to 50 g / l, Cu2 concentration from 20 to 50 g / l, H2SO concentration from 0 to 200 g / l, chlorine concentration of 20 at 200 ppm, and linear velocity of recirculation of the electrolyte beyond the electrode of 2 to 20 cm / sec, with additions of certain additives, such as glue and thiourea, etc. At a current density of 1,000 A / m2, copper cathodes with a thickness of approximately 0.635 cm can be produced over a period of 2 days. Throughout the system, the electrolyte must be acidic enough to avoid any precipitation of ferric hydroxide. A high acid content reduces the voltage drop in the electrolyte in the electroextraction cells. Currently, it is believed that the upper limit is determined by the saturation of the electrolyte. A small amount of chlorine in the electrolyte improves copper cathode quality, current efficiency and copper extraction during leaching. If type 316 stainless steel is used as a cathode substrate, sting may occur at a high concentration of chlorine. The glue, thiourea, guar gum or other additives can be added in a suitable amount to ensure a smooth copper cathode. In a 3-cell electroextraction, such as that shown in FIG. 3, the redox potential, of ferric concentration, is lower in the electroextraction cell of 1a. stage 4a, and higher in the electroextraction cell of 3a. stage 4c. The copper current efficiency shows the opposite, that is, it is higher in the electroextraction cell of 1a. Stage 4a and lower in the electroextraction cell of 3a. stage 4c. If the electrolyte passes through the electroextraction cells 4a, 4b and 4c in series, varying inversely with the ferric concentration, that is, via lines 14, 44, 45 and 15 in Figure 3, and no electrolyte bleed currents are formed between them, only a redox potential of the electrolyte electrowinning cell can be effectively controlled by changing the input current via line 14, preferably, the redox potential or ferric concentration in the electrowinning cell of 3a. Stage 4c needs to be controlled. In view of the electroextraction alone, a lower redox potential or a low ferric concentration is preferred. However, since the surplus current from the electroextraction cell of 3a. Stage 4c returns to the leaching circuit, the ferric concentration must be high enough to completely leach the copper matte into Cu2 ions and elemental sulfur. Due to this contradictory nature for ferric concentration, a compromised selection of ferric concentration must be maintained.

A one-stage electroextraction operation in the present invention is also possible. In a one-stage electroextraction operation, it is desirable to have a long, single, horizontal electrowinning cell with multiple electrodes in parallel placed in series along the length of the cell. The Fe3 + is increased from 0.3 g / l to approximately 9 g / l along the length of the cell. Thus, the concentration of Fe3 varies continuously over the length of the cell as or will do the density efficiency. The various aspects of the present invention can be illustrated through the following examples: EXAMPLE 1 Effect of Chlorine and Thiourea Mode: Leaching of 1 stage in batches and electroextraction of 1 stage. Electrolyte: 35 g / L of Cu2 +, 35 g / L Fe (, ßtai), 110 g / l of H2SO4, Cl ", glue and / or thiourea Conditions: 80 ° C, redox potential = 0.425 volts vs. SCE; 6 hours long, 100 grams of high grade copper matte (-325 mesh) for each test, glue added at 2.75 kg / ton started at 16 ppm, EW electrolyte circulation at 7 GPM (7.9 cm / sec) 970 A / m2 cd for the cathode and the anode, electrode dimensions: 5.08 x 10.16 cm: thickness of the titanium cathode substrate 0.317 cm (0.15875 cm depressed), thickness of Pb, Sn, Ca, Cu anodes of 0.47625 cm; separation of 2.54 cm from center to center between the anode and the cathode.

Findings: (1) With the addition of thiourea together with the glue, the cell voltage increases from approximately 50 to 80 mV (2) When the concentration of CI "is equal to or is above 75 ppm, if the thiourea is or not added together with the glue, will not affect the current efficiency and the quality of copper cathodes. (3) However, when the concentration of CI "is on the scale of 25 to 50 ppm, the joint addition of thiourea together with the glue is beneficial. (4) With a concentration of CI "of zero in the electrolyte, the joint addition of thiourea together with glue is undesirable.

EXAMPLE 2 Effect of the Current Density Mode: Leaching of 1 stage in batches and electroextraction of 1 stage. Electrolyte: 35 g / L of Cu2 +, 35 g / L of Fe (totai), 110 g / l of H2SO4, 75 ppm of CI ', glue and / or thiourea Conditions: 80 ° C, redox potential = 0.425 volts vs. SCE; 6 hours long; high grade copper matte (-325 mesh); and for each test 100 grams for 1,000 A / m2 and 150 grams for 1,500 A / m2; glue added at 2.75 kg / ton with 16 ppm; EW electrolyte circulation at 7 GPM (7.9 cm / sec); electrode dimensions: 5.08 x 10.16 cm: thickness of the 0.317 cm titanium cathode substrate (0.15875 cm depressed); anode thickness of Pb, Sn, Ca, Cu of 0.47625 cm; separation of 2.54 cm from center to center between the anode and the cathode.

Findings: (1) Both cell voltage and current efficiency increase with increased density.

EXAMPLE 3 Effect of the Redox Potential of Electrolyte Electrolyte Mode: Leaching of 1 stage in batches and electroextraction of 1 stage. Electrolyte: 35 g / L Cu2 +, 35 g / L Fß (totai), 110 g / l H2SO4, 100 ppm CI ', glue and / or thiourea Conditions: 80 ° C, redox potential = 0.425 volts vs. SCE; high grade copper matte (-325 mesh); thiourea added at 0.185 kg / ton starting at 16 ppm; EW electrolyte circulation at 7 GPM (7.9 cm / sec); 970 A / m2 c.d. for the cathode and 1,720 A / m2 for the anode; electrode dimensions: 5.08 x 10.16 cm: thickness of the 0.317 cm titanium cathode substrate (0.15875 cm depressed); anode thickness of Pb, Sn, Ca, Cu of 0.47625 cm; separation of 2.54 cm from center to center between the anode and the cathode.

Findings: (1) The current efficiency with the redox potential in increment of the electroweak electroextraction due to more ferric ions is reduced at the cathode.

EXAMPLE 4 Temperature Effect Mode: Leaching of 1 stage in batches and electroextraction of 1 stage. Electrolyte: 35 g / L Cu2 +, 35 g / L Fe (, otai), 5.5 g / L Fe3 +, 110 g / L H2SO4, 75 ppm of CI ", glue and / or thiourea Conditions: 6 hours of duration, 100 grams of high grade copper matte (-325 mesh) for each test, glue added at 2.75 kg / ton started with 16 ppm; Thiourea added at 0.185 kg / ton started at 1.6 ppm; EW electrolyte circulation at 7 GPM (7.9 cm / sec) 970 A / m2 c.d. for the cathode and the anode; electrode dimensions: 5.08 x 10.16 cm: thickness of the 0.317 cm titanium cathode substrate (0.15875 cm depressed); anode thickness of Pb, Sn, Ca, Cu of 0.47625 cm; separation of 2.54 cm from center to center between the anode and the cathode.

Findings: (1) Between 60 and 80 ° C, the temperature has a small effect on current efficiency. However, it strongly affects the surface of the cathodes. (2) The cell voltage and power consumption always increase with the reduction of temperature.

EXAMPLE 5 Effect of Impurities Mode: Leaching of 1 stage in batches and electroextraction of 1 stage. Electrolyte: 35 g / L of Cu2 +, 35 g / L of Fe (tßtai), 110 g / l of H2SO4, 75 ppm of CI ", glue and / or thiourea Conditions: 80 ° C, redox potential = 0.425 volts vs. SCE 6 hours long, 120 grams of high grade copper matte (-325 mesh) for each test; glue added to 2. 75 kg / ton with 16 ppm; thiourea added at 0.185 kg / ton started at 1.7 ppm; EW electrolyte circulation at 6 GPM (7.1 cm / sec) 970 A / m2 c.d. for the cathode and the anode; Electrode dimensions: 5.08 x 10.16 cm: thickness of stainless steel substrate type 316 0. 317 cm (0.15875 cm depressed); anode thickness of Pb, Sn, Ca, Cu of 0.47625 cm; separation of 2.54 cm from center to center between the anode and the cathode.

Spectrographic Analysis of Emission of Copper Cathodes Findings: (1) The three impurities, Sb, As and Bi have no significant effect on the efficiency of cell voltage current and cathode surface morphology.

EXAMPLE 6 Effect of Circulation of Electrolyte in the Electroextraction Cell Mode: Leaching of 1 stage in batches and electroextraction of 1 stage. Electrolyte: 35 g / L Cu2 +, 35 g / L Fe totai), 110 g / l H2SO4, 75 ppm CI ", glue and / or thiourea Conditions: 80 ° C, redox potential = 0.425 volts vs. SCE; 6 hours in length, 120 grams of high grade copper matte (-325 mesh) for each test, glue added at 2.75 kg / ton started at 17 ppm, thiourea added at 0.185 kg / ton started at 1.7 ppm, 970 A / m2 cd for the cathode and anode, electrode dimensions: 5.08 x 10.16 cm: thickness of the 316 stainless steel substrate of 0.317 cm (0.15875 cm depressed), eltech anode thickness of 0.47625 cm, separation of 2.54 cm from center to center between the anode and the cathode.

Findings: (1) The current efficiency always increases with the reduction of electrolyte circulation in the electrowinning cell. (2) The slowest flow velocity, which still ensures a good, coherent, relatively smooth cathode without preferential dendrite growth is about 3.6 cm / sec.

EXAMPLE 7 Cobre Copper Cathodes (48-Hour Tests) Mode: Continuous 3-stage leaching and 1-stage electroextraction. Electrolyte - 35 g / L of Cu2 +, 35 g / L Fe (, 0t «i), 110 g / l of H2S04, 75 ppm of CI" Deanement v / o thiourea Conditions: 80 ° C, redox potential = 0.425 volts vs SCE, 48 hours long, 250 grams of high grade copper matte (-325 mesh) in each 1.6-L leach reactor: L1, L2, and L3; 970 A / m2 cd for the cathode and anode electrode dimensions: 5.08 x 10.16 cm: thickness of 316 stainless steel substrate of 0.317 cm (depressed: 0.79375 cm left and right, 0.635 c upper and lower), Eltech ALE anode thickness of 0.47625 cm; 2.54 cm from center to center between the anode and the cathode.

Analysis of Copper Matte Leach Leach Residues Findings: (1) A copper recovery of 95% can be easily obtained. Note: (1) Acc. Represents accumulated (or mixed) leaching residues, which were collected in a period of 48 hours.

EXAMPLE 8 Eltech ALE anode. { 48 hour tests) Mode: Continuous 3-stage leaching and 1-stage electroextraction. Electrolyte: 35 g / L of Cu2 +, 35 g / L of Fe (totai), 110 g / l of H2SO4, 75 ppm of CI ', glue and / or thiourea Conditions: 80 ° C, redox potential = 0.425 volts vs. SCE; 48 hours long; 250 grams of high grade copper matte (-325 mesh) in each reactor; 970 A / m2 c.d. for the cathode and the anode; Electrode dimensions: 5.08 x . 16 cm; glue added at 2.75 kg / ton started with 17 ppm; thiourea added at 0.185 kg / ton started with 1.7 ppm; electrolyte circulation EW 7 GPM (8.3 cm / sec); separation of 2.54 cm from center to center between the anode and the cathode.

Findings: (1) The ALE anode from Eltech provides approximately 42% energy savings, compared to the Pb-Ca-Sn-Cu anodes, which are used in the industry.

In summary, according to the present invention, an SO2-free process is provided for the extraction of copper. The process produces copper, in a system either batch or continuous, from a copper matte in an acid sulfate medium via leaching and electroextraction methods. Prior to leaching, the copper matte is milled at a particle size less than 325 mesh. Then, the copper matte is subjected to leaching through ferric sulfate in an acid solution containing cuprous and ferrous sulfates under atmospheric pressure and at a temperature lower than the boiling point. More than 95% of the copper originally present in the copper matte can dissolve in the solution during leaching. The leaching residue may contain incomplete reaction copper matte (less than 5%), elemental sulfur, PGM metals, gold and silver, and other impurities. The incomplete reaction copper matte, elemental sulfur, PGM metals, gold and silver in the leach residue can be completely recovered. Subsequent to solid-liquid separation after leaching, the pregnant solution is pumped into a forced-flow, non-diaphragm electroextraction cell, where high quality copper cathodes are produced at a density 3 to 4 times as high like that used in conventional copper electroextraction, and at the same time a ferric ion is generated from the oxidation of ferrous iron that occurs at the anode. Despite the current density of the electroextraction that is 3 to 4 times as high as that in conventional copper electroextraction, the cathode copper deposit is consistent and meets the market ASTM specifications for copper cathode grade 1 . There is no development at the anode, and the energy consumption of electroextraction is less than or equal to that in conventional copper electroextraction, that is, of the order of 1.2 kwh / kg. The spent electrolyte from the electroextraction is returned to the leach circuit. In the above specification, the present invention has been described with respect to the specific embodiments. These serve as examples to illustrate the invention instead of limiting its scope. Modifications can be made without departing from the broader teachings and scope of the invention.

Claims (32)

1. - A process for the production of high purity metallic copper from copper matte, comprising the steps of: leaching the copper matte under oxidation conditions in a ferric ion containing a copper sulfate electrolyte leaching assembly acid, including one or more reactors, to solubilize in copper in a leaching solution forming a copper-rich leaching solution and a residue containing elemental sulfur; and electroextracting the copper from the copper-rich leaching solution and regenerating ferric ions in an electroextraction cell assembly, including one or more electroextraction cells; wherein the electroextraction cell assembly is physically decoupled from the leach assembly.

2 - The method according to claim 1, wherein the steps are carried out at atmospheric pressure.

3. The process according to claim 1, wherein the steps are carried out at temperatures on the scale of 40 ° C to the boiling point.

4. The process according to claim 1, wherein the leaching is achieved through chemical attack of the ferric ions.

5 - The process according to claim 1, wherein the step of leaching the copper matte is carried out under a countercurrent flow of the copper matte and the leaching solution. 6 -.

6 - The method according to claim 1, wherein the electroextraction occurs in one or more electroextraction cells, which do not contain a diaphragm.

7. The method according to claim 1, wherein the current density in the electroextraction cell assembly is in the range of 500 to 2000 A / m2.

8. The method according to claim 1, wherein the linear recirculation speed of the electrolyte is in the range of 2 to 30 cm / sec.

9. The process according to claim 1, wherein one or more of glue, thiourea, guar, and chlorine are added to the electroextraction assembly.

10 - The process according to claim 1, wherein the chlorine is added to the leach tank assembly.

11. The method according to claim 1, wherein one or more of the electroextraction cells contains an activated Lead Electrode anode.

12. A high purity copper cathode product produced by the process of claim 1.

13 - The high purity copper cathode product according to claim 7, which satisfies the specifications for ASTM grade 1.

14 - A process for the continuous production of high purity metallic copper from copper matte, comprising the steps of: preparing a finely divided copper mat; leaching the copper matte under oxidation conditions in a ferric ion containing an acid copper sulfate electrolyte leach assembly, which includes one or more leach reactors, to solubilize the copper in a leach solution forming a rich leaching solution in copper and a residue that contains elemental sulfur; Separate the copper-rich leaching solution from the residue; electroextracting the copper from the copper-rich leaching solution and regenerating ferric ions in an electroextraction assembly including one or more electroextraction cells, wherein the electroextraction cell assembly is physically uncoupled from the leach tank assembly; recover an electrolyte rich in ferric ions from the electroextraction assembly; recover elemental sulfur from the leaching residue; and recover precious metals from the leach residue.

15. The process according to claim 14, wherein the steps are carried out at atmospheric pressure.

16. The process according to claim 14, wherein the steps are carried out at temperatures on the scale of 40 ° C to the boiling point.

17 - The method according to claim 14, where leaching is achieved through chemical attack of ferric ions.

18. The process according to claim 14, wherein the step of leaching the copper matte is carried out under an anti-current flow of the copper matte and the leaching solution.

19. The process according to claim 14, wherein it also comprises the steps of removing a stream of electrolyte rich in ferric ions and remove impurities from the stream, creating a purified electrolyte.

20. The process according to claim 19, wherein further comprises the step of transferring the purified electrolyte back to the electroextraction cell assembly.

21. The process according to claim 19, wherein the impurities are removed through a solvent extraction process.

22. The process according to claim 19, wherein the impurities are removed through an ion exchange process.

23. The method according to claim 19, wherein the removed impurities are one or more of As, Bi or Sb.

The method according to claim 14, wherein the electroextraction occurs in one or more electroextraction cells, which do not contain a diaphragm.

25. The method according to claim 14, wherein the current density in the electroextraction cell assembly is in the range of 500 to 2000 A / m2.

26. - The method according to claim 14, wherein the linear recirculation speed of the electrolyte is in the range of 2 to 30 cm / sec.

27. The process according to claim 14, wherein one or more of glue, thiourea, guar, and chlorine are added to the electroextraction assembly.

28. The process according to claim 14, wherein the chlorine is added to the leach tank assembly.

29. A high purity copper cathode product produced by the process of claim 14.

30.- The high purity copper cathode product according to claim 29, which satisfies the specifications for ASTM grade 1.

31. The process according to claim 1, wherein the copper matte is a high grade matte containing at least 60% Cu.

32. The method according to claim 14, wherein the copper matte is a high grade matte containing at least 60% Cu.