Hydrometallurgical process for the recovery of copper, lead or zinc

Abstract

A hydrometallurgical process for the treatment of polymetallic ores and sulphide concentrates of copper and zinc, and by-products of lead and zinc from smelting plants, treated independently and/or as mixtures thereof, which contain relevant amounts of lead, copper, zinc, iron, gold and silver, such as the matte-speiss mixture of lead foundries, and copper cements from the purification processes of electrolytic zinc plants. The process allows the recovery of metallic copper, zinc, copper as copper and zinc basic salts, which may be hydroxides, carbonates, hidroxysulphates or mixtures thereof; the production of stable arsenic residues; and the effective and efficient recovery of Pb, Au and Ag as a concentrate of lead sulphide and/or lead, Au, and Ag sulphate.

Claims

1. A hydrometallurgical process for treatment of polymetallic ores and sulfide concentrates of copper and zinc, by-products of lead and zinc from smelting plants, or mixtures thereof having a content of lead, copper, zinc, iron, arsenic, gold and silver, wherein the hydrometallurgical process comprises: grinding a raw material to a particle size of less than 44 microns; leaching the ground raw material by feeding to a solid-gas-liquid (SGL) reactor with oxygen of less than 30 psia and an acid solution so as to obtain a residue containing lead sulfate, silver and gold with elementary sulfur (PbSO.sub.4Ag/Au+S.sup.o) and a solution containing copper, zinc, arsenic and iron in a sulfuric acid medium; purifying the leaching solution by neutralizing residual acidity in the leaching solution; recovering gold and silver from the leaching residue by eliminating S.sup.o in the residue; precipitating copper or zinc basic salts from the purified solution with a neutralizing agent so as to obtain a solution containing magnesium sulfate and/or sodium sulfate with heavy metals, and basic salts of copper or zinc; purifying the precipitated solution containing magnesium sulfate and/or sodium sulfate so as to produce a diluted solution of sodium sulfate and/or magnesium sulfate and removing heavy metals therefrom; concentrating the diluted solution of sodium sulfate and/or magnesium sulfate so as to recover water and to obtain a saturated solution of magnesium sulfate and/or sodium sulfate; and crystallizing magnesium sulfate and/or sodium sulfate salts from the saturated solution of magnesium sulfate and/or sodium sulfate so as to obtain hydrated salts of magnesium sulfate and/or sodium sulfate respectively.

2. The hydrometallurgical process of claim 1, wherein the step of purifying the leaching solution comprises: eliminating arsenic as a stable residue of ferric arsenate.

3. The hydrometallurgical process of claim 2, wherein the step of eliminating comprises: neutralizing the free acidity of the leaching solution with a suspension of a neutralizing agent, the neutralizing agent selected from the group consisting of sodium hydroxide, sodium carbonate, magnesium oxide, magnesium hydroxide, and calcium hydroxide; and adjusting a pH of the neutralized leaching solution to a value of between 2 and 5 so as to stabilize the residue.

4. The hydrometallurgical process of claim 1 wherein the precipitated copper or zinc basic salts are selected from the group consisting of copper or zinc hydroxide, copper or zinc carbonate, copper or zinc hydroxy sulfate, and combinations thereof.

5. The hydrometallurgical process of claim 1, wherein the neutralizing agent used in the precipitation step is selected from the group consisting of sodium hydroxide, sodium carbonate, magnesium oxide, magnesium hydroxide and calcium hydroxide.

6. The hydrometallurgical process of claim 1, wherein the step of purifying precipitated solution containing magnesium sulfate and/or sodium sulfate comprises: adding a sodium sulfide solution to the precipitated solution so as to convert the heavy metals into respective sulfides thereof.

7. The hydrometallurgical process of claim 1, wherein the hydrometallurgical process is a batch process, wherein the step of leaching is in a reactor that contains a solution of high acidity of not less than 300 g/L of H.sub.2SO.sub.4 and in which a relation of iron II/As in solution is at least 2, wherein a pressure in the reactor is constant to a partial oxygen pressure of less than 30 psia, and the step of leaching being at a temperature of less than 100 C. and at a reaction time of less than 9 hours.

8. The hydrometallurgical process of claim 1, wherein an initial oxygen pressure in the step of leaching is from 5 to less than 30 psia.

9. The hydrometallurgical process of claim 1, wherein an initial solids concentration in the step of leaching is up to 500 g/L.

10. The hydrometallurgical process of claim 1, wherein the step of recovering gold and silver comprises: converting lead sulfate in the leaching residue to lead sulfide; and integrating the lead sulfide into a synthetic galena concentrate containing the gold and silver.

11. The hydrometallurgical process of claim 10, wherein the step of recovering the gold and silver further comprising: purging S.sup.o as a solution of Na.sub.2SO.sub.4.

12. The hydrometallurgical process of claim 1, wherein the step of recovering gold and silver comprises: dissolving a residue of PbSO.sub.4Ag/Au+S.sup.o in two countercurrent stages in a solution of sodium sulfide so as to eliminate the S.sup.o and to convert the lead sulfat into lead sulfide so as to produce a synthetic galena concentrate containing gold and silver.

13. The hydrometallurgical process of claim 1, wherein the step of recovering gold and silver comprises; dissolving the S.sup.o in tetrachloroethylene so as to produce a residue of PbSO.sub.4 with gold and silver, wherein the S.sup.o remains dissolved in the C.sub.2Cl.sub.4; separating the S.sup.o by cooling; and recovering the C.sub.2Cl.sub.4.

Description

BRIEF DESCRIPTION OF THE FIGURES

(1) FIG. 1 is a schematic diagram of the hydrometallurgical process for the recovery of Pb and Cu and/or Zn.

(2) FIG. 2 shows a graphical representation of the hydrometallurgical process for treating matte-speiss material (Cu.sub.2SCu.sub.3As) from a lead foundry, and copper cements from an electrolytic zinc plant.

(3) FIG. 3 corresponds to a graph indicating the leaching stage of the hydrometallurgical process for treating chalcopyrite-based copper concentrates.

(4) FIG. 4 is a diagram of the leaching stage of the hydrometallurgical process for treating sphalerite-based zinc concentrates.

(5) FIG. 5 corresponds to a graph showing the purification stage of the leaching solution for arsenical precipitation with MgO in the p hydrometallurgical process for treating matte-speiss material (Cu.sub.2SCu.sub.3As).

DETAILED DESCRIPTION OF THE INVENTION

(6) The hydrometallurgical process proposed for the recovery of Cu and Pb and/or Zn is schematically illustrated in the block diagram of FIG. 1, formed by the stages described below:

(7) Stage 1. Grinding (110)

(8) The raw material (101), which consists of polymetallic Cu and Zn ores, sulphide Cu and Zn concentrates, by-products of Pb foundries and by-products from Zn processing plants are subjected to dry grinding (110) to a particle size of less than 44 microns. Then, the material resulting from the grinding is sent for leaching (120).
Stage 2. Leaching (120) The material resulting from grinding (110) is fed to a solid-gas-liquid reactor (SGL) to be leached (120) at low pressure with oxygen (less than 25 psia) to obtain a residue containing lead sulphate, silver and gold with elemental sulphur (PbSO.sub.4Ag/Au+S.sup.o), which is sent for value recovery (130), and a solution containing either copper, zinc, arsenic and iron, or mixtures thereof in sulphuric acid medium, which in turn is sent for purification (140), to obtain an extraction of copper and zinc greater than 95% and extraction of arsenic of at least 80%. The operation is batch type, the reactor contains high acidity solution of not less than 300 g/l H.sub.2SO.sub.4, a ratio of iron II/metal in solution of less than 2, and a surfactant to regulate the surface tension, keeping the reactor pressure constant with partial oxygen pressure less than 30 psia, with agitation ensuring efficient contact between the solid-liquid-gas, at a temperature of less than 100 C., and reaction time of less than 9 hours, achieving a quasi-stoichiometric utilisation of oxygen and efficiency above 95%.

(9) The leaching operation has an initial solids concentration of 500 g/l, and is performed in a pressurised tank, where the initial partial oxygen pressure (Pp O.sub.2) ranges from 5 to 30 lb/in.sup.2.

(10) Stage 3. Purification of the Iron and Arsenic Leaching Solution (140)

(11) The purification of the iron and arsenic leaching solution, which further contains either copper and/or zinc or mixtures thereof, allows a purified solution of CuSO.sub.4 and/or zinc sulphate and magnesium sulphate to be obtained, which is sent to precipitation of basic salts (150), and elimination of As through a stable residue, FeAsO.sub.4 (142).

(12) For the removal of As content in the leaching solution, this is precipitated as ferric arsenate by neutralising the free acidity with a suspension of any of the neutralisers (141) such as sodium hydroxide (NaOH), sodium carbonate (Na.sub.2CO.sub.3), magnesium oxide (MgO), magnesium hydroxide [Mg(OH).sub.2] and/or calcium hydroxide [Ca(OH).sub.2], to an adjusted pH value of between 2 to 5, so as to ensure the chemical stability of the residue. The As is removed through a stable residue, FeAsO.sub.4 (142).

(13) Stage 4. Recovery of Gold and Silver Valuables (130a or 130b)

(14) There are two alternatives for removing the S.sup.o from the PbSO.sub.4Ag/Au+S.sup.o residue obtained from the leaching process (120): (a) A first alternative for the recovery of gold and silver valuables (130a), uses Na.sub.2S (131a). The S.sup.o in the PbSO.sub.4Ag/Au+S.sup.o residue obtained from leaching (120), is converted to polysulphides (Na.sub.xS.sub.Y) in a solution of sodium sulphide (Na.sub.2S) in two countercurrent stages, forming a solution of Na.sub.2SO.sub.4 (133a). Furthermore, the PbSO.sub.4 from the PbSO.sub.4Ag/Au+S.sup.o residue obtained from the leaching (120) is converted to lead sulphide (PbS), generating a synthetic galena concentrate rich in Au and Ag (132a) that is sent to the Lead Smelting Plant for subsequent processing. The excess sulphur is purged as a solution of Na.sub.2SO.sub.4 (133a), and the extraction rate of S.sup.o ranges from 95% to 99%. The recovery of Ag and Au valuables, contained mainly in the synthetic galena (lead sulphide, PbS), stands at around 99%, with S.sup.o content of less than 1%, and a reaction conversion of PbSO.sub.4 to PbS of over 99%. (b) A second alternative for the recovery of gold and silver valuables (130b), uses C.sub.2Cl.sub.4 (131a). The S.sup.o in the PbSO.sub.4Ag/Au+S.sup.o residue obtained from leaching (120), and the tetrachloroethylene (C.sub.2Cl.sub.4) form a solution (132b), subsequently separated by cooling the S.sup.o (133b) and the tetrachloroethylene is recovered for recycling back to the process (130b). Furthermore, a PbSO.sub.4 residue is generated containing Au and Ag (134b) that is sent to the Lead Smelting Plant for further processing.
Stage 5. Precipitation of Basic Salts (150) The purified solution of CuSO.sub.4 and/or zinc sulphate and magnesium sulphate coming from the purification of the leaching solution (140) is precipitated as copper and/or zinc basic salts (152) with a neutralising agent (151), such as preferably sodium hydroxide (NaOH), sodium carbonate (Na.sub.2CO.sub.3), magnesium oxide (MgO) and/or magnesium hydroxide Mg(OH).sub.2, resulting in a solution containing mainly magnesium sulphate (MgSO.sub.4) and/or sodium sulphate (Na.sub.2SO.sub.4) with traces of heavy metals (such as Cu, Cd, Co and Mn) to be purified subsequently (160), and, on the other hand, copper and/or zinc basic salts are obtained (152) which can be copper or zinc hydroxide [Cu(OH).sub.2 or Zn(OH).sub.2], copper or zinc carbonate (CuCO.sub.3 or ZnCO.sub.3), copper or zinc hydroxysulphate [Cu.sub.4SO.sub.4(OH).sub.6 or Zn.sub.4SO.sub.4(OH).sub.6], or mixtures thereof.
Stage 6, Purification of the Magnesium Sulphate or Sodium Sulphate Solution (160) The purification (160) of the magnesium sulphate or sodium sulphate solution with traces of heavy metals obtained from the precipitation of basic salts (150), generates diluted magnesium sulphate solution (MgSO.sub.4) and/or sodium sulphate solution (Na.sub.2SO.sub.4) and ensures the elimination of traces of heavy metals through the use of sodium sulphide (161) converting the heavy metals (such as Cu, Cd, Ca, and Mn) into their respective sulphides (162).
Stage 7. Water Recovery (170) For water recovery (170), the diluted magnesium sulphate or sodium sulphate solution obtained from the purification of the MgSO.sub.4 and/or Na.sub.2SO.sub.4 solution undergoes a concentration process that allows the recovery of water (171) in percentages above 70%, and resulting in a saturated solution of magnesium or sodium sulphate which subsequently undergoes crystallisation (180).
Stage 8. Crystallisation of the Magnesium Sulphate or Sodium Sulphate Salt (180)

(15) The saturated magnesium or sodium sulphate solution obtained in the water recovery (170), is sent a crystallisation process (180) to obtain MgSO.sub.4 salts (such as MgSO.sub.4.7H.sub.2O) or hydrated Na.sub.2SO.sub.4 salts (181).

(16) This invention is additionally described through the following examples that should not be considered to be limiting, which detail the preferred modalities.

EXAMPLE 1

Hydrometallurgical Process for Treating Matte-Speiss Material (Cu2S Cu3As) from a Lead Foundry, and Copper Cements from an Electrolytic Zinc Plant

(17) FIG. 2 shows the schematic block diagram of the hydrometallurgical treatment process of matte-speiss material (Cu.sub.2SCu.sub.3As) from a lead foundry, and copper cements from an electrolytic zinc plant, where each stage of the invention's hydrometallurgical process (FIG. 1) is renumbered for the specific application conditions of this example, as follows:

(18) Stage 1. Grinding (210)

(19) Matte-speiss material (211) containing 40.13% copper, 20.40% lead, 10.5% total sulphur, 6.73% iron and 4.22% arsenic is subjected to (210) dry grinding until obtaining a particle size P.sub.90 of 45 microns. Then, the resulting matte-speiss material is sent for leaching (220).

(20) Stage 2a. Leaching of the Matte-Speiss Material (220)

(21) A sample of 4.310 g of matte-speiss material from the grinding (210), with a particle size P90 of 46 microns, is mixed with an acid solution (221) containing 5 g/l of iron as iron sulphate, 18 g of a reactive surfactant, and 180 g/l of free acidity. The reactor (221) is closed and kept at a partial oxygen pressure of 12 lb/in.sup.2, the reaction temperature is 90 C. and it is allowed to react for 7 hours. Subsequently, the suspension is filtered and the residue is rinsed with water, obtaining 1.745 g of solids containing 0.79% copper; 39.81% lead; 2.15% silver; 0.96% iron; 3.08% arsenic and 12% elementary sulphur, which is sent to valuables recovery (230). The end solution contains 80 g/l copper; 12.98 g/l total iron; 10.04 g/l arsenic and 60 g/l free sulphuric acid, which is sent for purification (250). Table 1 shows the extraction of copper as a function of leaching time.

(22) TABLE-US-00001 TABLE 1 Extraction of copper as a function of leaching time TIME EXTRACTION (hr) (%) 0 0.0 0.5 48.6 1 67.3 1.5 74.0 2 80.8 3 88.6 4 94.9 5 95.5 6 96.7 7 99.2

(23) Other applications of this stage of leaching may be the leaching of concentrates of copper based on chalcopyrite (Example 2), or concentrates of zinc based on sphalerite (example 3), which are described below.

(24) The following steps, which are described below relating to Example 1, if required, can be applicable to leaching of copper concentrates based on chalcopyrite, as described in Example 2, as well as the leaching of zinc concentrates based on sphalerite, as described in Example 3.

(25) Stage 2b. Leaching of Copper Cements (240)

(26) A sample of 3.372 g of copper cement from the electrolytic zinc plant (241), containing 65.29% copper, 4.78% lead, 4.23% zinc, 1.92% cadmium, and 0.46% cobalt, is added to a solution composed of 24.3 litres of an acid solution (242) that contains 141 g/l of sulphuric acid, to be leached in another SGL reactor different to the reactor where matte-speiss material was leached. The reactor (241) is closed and kept at a partial oxygen pressure of 12 lb/in.sup.2, the reaction temperature is 90 C. and it is allowed to react for 2 hours. After the reaction time, the suspension is filtered and the residue is rinsed with water, obtaining 362 g of end solids containing 3.71% copper, 44.05% lead, 0.42% zinc, 0.09% cadmium and 0.014% cobalt, which is sent to valuables recovery (230). The end solution contains 100 g/l copper, 0.28 lead, 6.16 g/l zinc, 2.58 g/l cadmium and 0.66 g/l cobalt, which is sent to precipitation of basic salts (260). Table 2 shows the extraction of copper according to leaching time.

(27) TABLE-US-00002 TABLE 2 Copper extraction according to leaching time of the hydrometallurgical treatment process for copper cements TIME EXTRACTION (hr) (%) 5 49.8 15 54.7 30 68.5 45 83.0 60 81.9 75 95.4 90 98.0 120 99.6
Stage 3. Purification of the Leaching Solution for Arsenic Precipitation with Ca(OH).sub.2 (250)

(28) To a sample of 1 l of the end solution from the leaching of the matte-speiss material (220), containing 80 g/l copper, 12.98 g/l total iron, 10.04 g/l arsenic and 60 g/l free sulphuric acid, and pH=0.2, 2 ml of hydrogen peroxide is added (251), stirring slowly for 15 minutes, to ensure an oxidation-reduction potential greater than 0.77 V. After this time, 220 ml of a suspension of calcium hydroxide is added (252) containing 300 g/l of Ca(OH).sub.2, and/or to reach a pH value of 2.6 to 2.8, and allowed to react for 60 minutes. Following the reaction time, the suspension is filtered and the residue is rinsed with water, obtaining 137.34 g of end solids (252) with 0.60% copper, 5.83% iron and 5.68% arsenic. The end solution contains 70.86 g/l copper; 2.68 g/l total iron; 0.048 g/l arsenic and 0-16 g/l of free sulphuric acid, which is sent for precipitation of basic salts (260).

(29) Another application of the purification stage (250) of the end solution from the leaching of the matte-speiss material for arsenic precipitation can be the use of MgO as neutralising agent, rather than adding calcium hydroxide (251). This alternative corresponds to Example 4 described below.

(30) Stage 4. Recovery of Valuables (230)

(31) The solid obtained from the matte-speiss leaching (220) is combined with the final solid retrieved from the copper cement leaching (240) for the recovery of valuables (230). A sample of 244 g of the mixture of the solids obtained in the matte-speiss and copper cement leaching processes (220 and 240), containing 0.79% copper, 39.81% lead, 2.15% silver, 0.96% iron, 3.08% arsenic and 12% elemental sulphur, is leached with 0.810 l of a sodium sulphide solution (231) containing 49.172 g/l of sodium in sodium sulphide form, and allowed to react (231) for 1 hour at a temperature of 70-80 C. Following this reaction time, the suspension is filtered, obtaining 210 g of solids (232) containing 53.6% lead, 2.59% silver, 3.69% arsenic; and 0.01% elemental sulphur. Whereby the main type of lead is lead sulphide. The resulting solution (233) contains 31.02 g/l sodium; 44 g/l total sulphur and 1.89 g/l antimony.
Stage 5. Precipitation of Basic Salts (260) The end solution from the purification of the leaching solution (250) together with the end solution from the cement leaching process (240) go on to the basic salt precipitation stage. A sample of 21.65 l of the mixture of the end solutions obtained from the purification of the leaching solution (250) and the copper cement leaching process (240), with a content of 57 g/l copper, 2.71 g/l calcium, 2.38 g/l zinc, 1.32 g/l iron, 1.13 g/l sodium, 0.4 g/l cadmium and 0.23 g/l magnesium heated at between 70 and 80 C., to which 888 g of magnesium oxide is added (261) with a particle size P90 of 44 microns and/or until reaching a final pH of the suspension between 6.5 to 7.5 and allowed to react for 7 hours. The suspension is filtered and the residue is rinsed with water, obtaining 2.580 g of end solids containing 48% copper, 3.9% sodium, 1.85% zinc, 1.0% iron and 0.03% cadmium. The end solution contains 24 g/l magnesium, 1.83 g/l calcium, 1.02 g/l sodium, 0.16 g/l zinc and 0.07 g/l cadmium, which is sent for purification (270).
Stage 6. Purification of the Magnesium Sulphate Solution (270)

(32) A sample of 24 l of an end of solution magnesium sulphate obtained from the precipitation of basic salts (260), with a content of 24 g/l magnesium, 1.84 g/l calcium, 1.13 g/l sodium, 0.17 g/l zinc, 0.07 g/l cadmium and 0.05 g/l cobalt, to which 0.28 l of a solution of sodium sulphide is added (272) with a concentration of 83 g/l of Na.sub.2S is allowed to react for 60 minutes, after which time the suspension is filtered and the residue is rinsed with water, obtaining 10 g of solids (272) containing 34% zinc, 14% cadmium and 9.57% cobalt. The end solution contains 23.88 g/l magnesium, 1.87 g/l calcium, 1.67 g/l sodium, <1 ppm zinc, <1 ppm cadmium and <1 ppm cobalt, which is sent for water recovery and calcium removal (280).

(33) Stage 7. Water recovery and calcium removal (280)

(34) A sample of 24.28 l of an end solution of magnesium sulphate obtained from the purification of the magnesium sulphate solution (270), with a content of 23.88 g/l magnesium, 1.87 g/l calcium, 1.67 g/l sodium, <1 ppm zinc, <1 ppm cadmium and <1 ppm cobalt is taken to evaporation point until the magnesium sulphate saturation point is reached (77.9 g/l). The recovered water (281) amounts to 16.56 l. The resulting suspension is filtered and the residue is rinsed with water, obtaining 10 g of final solids (282) with 29.45% Ca. The end solution contains 77.9 g/l magnesium, 0.53 g/l calcium, 5.49 g/l sodium, <1 ppm zinc, <1 ppm cadmium and <1 ppm cobalt, which is sent for crystallisation (290).

(35) Stage 8. Crystallisation of the Magnesium Sulphate Heptahydrate Salt (290)

(36) A sample of 7.44 l of an end solution of magnesium sulphate obtained from the water recovery and calcium removal process (280), with a content of 77.9 g/l magnesium, 0.53 g/l calcium, 5.49 g/l sodium, <1 ppm zinc, <1 ppm cadmium and <1 ppm cobalt is subjected to a crystallisation process (290). The resulting suspension is filtered, obtaining 312 g of solids (291) in magnesium sulphate heptahydrate form with a purity of 99.95% containing 9.84% magnesium and 0.008 manganese, 0.098 sodium and 0.002 selenium. The end solution contains 46.6 g/l magnesium, 0.64 g/l calcium, 6.82 g/l sodium, <1 ppm zinc, <1 ppm cadmium and <1 ppm cobalt.

EXAMPLE 2

Leaching Stage of the Hydrometallurgical Process for Treating Chalcopyrite-Based Copper Concentrates

(37) FIG. 3 shows a block diagram of the grinding and leaching stages of the hydrometallurgical process for treating chalcopyrite-based copper concentrates, where:

(38) A sample of 999 g of a chalcopyrite concentrate (311) containing 19.80% copper, 10.20% zinc, 20.30% iron and 28.60% of total sulphur, is subjected to grinding (310) to obtain a particle size P80 of 15 microns, the resulting material is sent to leaching (320) where the sample is mixed with 16.5 l of a solution (321) containing 11.50 g/l iron as ferrous sulphate and 64.7 g/l free sulphuric acid. The reactor (321) is closed and kept at a partial oxygen pressure of 12 lb/in.sup.2, the reaction temperature is 80 C. and it is allowed to react for 8 hours, the redox potential during this reaction time is maintained between 400 and 500 my with respect to the Ag/AgCl electrode.

(39) Subsequently, the suspension is filtered (330) and the residue is rinsed with water, obtaining 402.1 g of end solids (331) with 4.80% copper, 2.1% zinc, 5.7% iron and 50.0% sulphur, and 16.5 l of an end solution (332) with 10.8 g/l copper, 5.1 g/l zinc, 21.7 g/l total iron (12.3 g/l as iron +3) and 12.3 g/l free sulphuric acid. Table 3 shows the extraction of zinc according to leaching time.

(40) TABLE-US-00003 TABLE 3 Extraction of copper according to leaching time, for the case of the treatment of chalcopyrite-based copper concentrates TIME EXTRACTION (hr) (%) 1 57.9 2 75.2 3 79.0 4 81.8 5 82.0 6 83.1 7 86.6 8 91.0

EXAMPLE 3

Leaching Stage of the Hydrometallurgical Process for Treating Sphalerite-Based Zinc Concentrates

(41) FIG. 4 shows a block diagram of the leaching stages of the hydrometallurgical process for treating sphalerite-based zinc concentrates, where:

(42) A sample of 262 g of a concentrate of zinc (411) containing 48.5% zinc, 12.39% iron and 34.6% of total sulphur, is subjected to grinding (410) to obtain a particle size P.sub.90 of 45 microns, the material retrieved is sent to leaching (420) where the sample is mixed with 239 g zinc ferrite (421) containing 19.8% zinc, 25% of total iron and 21.6% as iron (+3). This material mixture is added to a solution (421) composed of 0.4 l water, 0.043 l sulphuric acid at 98% purity and 3.070 l zinc sulphate solution containing 36.50 g/l zinc as zinc sulphate and 165.6 g/l free sulphuric acid.

(43) The reactor (421) is closed and kept at a partial oxygen pressure of 12 lb/in.sup.2, the reaction temperature is 90 C. and it is allowed to react for 14 hours, the redox potential during this reaction time is maintained between 400 and 500 my with respect to the Ag/AgCl electrode.

(44) Subsequently, the suspension is filtered (430) and the residue is rinsed with water, obtaining 125 g of end solids (431) with 0.7% zinc, 5.1% iron and 71.2% sulphur, and 3.5 l of an end solution (432) with 79.50 g/l zinc, 24.2 g/l total iron and 24 g/l free sulphuric acid. Table 4 shows the extraction of zinc as a function of leaching time.

(45) TABLE-US-00004 TABLE 4 Extraction of zinc as a function of leaching time, for the case of the treatment of sphalerite-based copper concentrates TIME EXTRACTION (hr) (%) 1 38.6 2 49.2 4 70.0 5 75.0 6 83.4 8 97.2 10 98.5 14 99.3

EXAMPLE 4

Purification Stage of the Leaching Solution for Arsenical Precipitation with MgO in the Hydrometallurgical Process for Treating Matte-Speiss Material (Cu2SCu3As)

(46) FIG. 5 shows the block diagram of the purification of the iron and arsenic leaching solution from the hydrometallurgical treatment leaching process of matte-speiss material (Cu.sub.2SCu.sub.3As), where:

(47) A sample of 1 l of the end solution from the leaching of the matte-speiss material (510), containing 73.12 g/l copper, 13.84 g/l total iron, 9.14 g/l arsenic and 60 g/l free sulphuric acid, is sent to purification (520), where 2 ml of hydrogen peroxide is added (521), stirring slowly for 15 minutes, to ensure an oxidation-reduction potential greater than 0.77 V. After this time, 50 g MgO is added (521) with a particle size of 350 mesh (less than 49 microns), with a magnesium content of 60% and 0.013% total iron, and/or until reaching a pH value of 2.6 to 2.8, and allowed to react for 60 minutes. Following the reaction time, the suspension is filtered (530) and the residue is rinsed with water, obtaining 70 g of end solids (252) with 9.64% copper, 17.61% iron and 11.84% arsenic. The end solution (532) contains 67 sodium; 0.06 g/l total iron and 0.002 arsenic.

(48) It may be seen that the above examples show some of the preferred modalities for implementing the invention, and it will be apparent to the person skilled in the art that a number of possible variations can exist to the process object of the present invention, based mainly, in the compositions of the raw material that will be processed; these variations, however, do not depart from the scope of this invention and should be considered to the light of the following claims.