Method for producing germanium concentrates from metallurgical residues - Patents.com

Abstract

A method for producing a germanium concentrate from a metallurgical residue includes (i) copper leaching the metallurgical residue with a first acidic solution to obtain a first leaching solution rich in copper and iron, and optionally rich in arsenic, antimony, and bismuth, and a first leached sludge having a copper- and iron-reduced content, optionally arsenic-reduced content, and a lead-enriched content, silicon, and germanium; (ii) leaching the first leached sludge with a first solution of sodium citrate to obtain a second leached sludge depleted in lead and a second leaching solution rich in lead; and (iii) adding a base to produce an alkaline leaching solution and alkaline leaching the second leached sludge to obtain a third leached sludge having a reduced content of silicon and germanium. (iv) packing an ion exchange column and capturing germanium with a resin to obtain a fourth alkaline solution that is germanium-depleted and silicon-rich; (v) washing the ion exchange column to obtain a fifth column pack wash solution; (vi) eluting the ion exchange column with an HCl solution to obtain a sixth eluted solution that is germanium-rich; (vii) distilling the sixth germanium-rich eluted solution to obtain a seventh germanium solution and an eighth germanium-depleted solution; and (viii) contacting the seventh germanium solution with a water solution to hydrolyze it and produce a first germanium dioxide concentrate.

Description

Detailed Description of the Invention

[0001] [Field of the Invention] The present invention relates to a method (procedure) for producing germanium concentrates from metallurgical residues, in particular from residues containing copper, iron, lead and germanium, and which may optionally contain elements (components) such as arsenic, antimony and bismuth.

[0002] In a more particular embodiment, the metallurgical residue is a powder obtained from a metal smelting process.

[0003] In a more particular embodiment, the metallurgical residue is a powder obtained from a copper smelting process.

[0004] In a more specific embodiment, metallurgical residue (or, in particular, smelter powder) contemplates material that has already been subjected to a leaching process, such as sulfuric acid leaching.

[0005] In a broad aspect, a germanium concentrate may be understood as a germanium-rich liquid fraction contained primarily in germanium tetrachloride or contained in a solid germanium concentrate, and in a more specific aspect, the liquid fraction may be associated with germanium dioxide.

[0006] In this disclosure, metallurgical residues that have been subjected to a prior leaching process are considered as sludge.

[0007] [Latest technology] <Copper leaching> The copper in the sludge is mainly composed of species such as ferrite and / or spinel in the form CuFe2O4, zinc as ZnFe2O4, and the associated part of iron, i.e. FeFe2O4. The leaching of these species is based on temperature, acid concentration, and residence time, as described in a study by BS Boyanov et al. (World Academy of Science, Engineering and Technology, Vol 9, 2015, 1592-1598). BS Boyanov et al. carried out a study on the leaching of synthetic ferrites of zinc, copper, and cadmium, evaluating each of the variables mentioned above. The results of this study show that at high temperatures and high acid concentrations, ferrites are well dissolved by hydrochloric acid and H2SO4.

[0008] At high acid concentrations, copper leaching has been observed to have an asymptotic behavior with respect to leaching temperature: after a reaction time of 60 minutes in sulfur medium, copper leaching yields of over 90% are reached at temperatures of 85-90 °C.

[0009] <Lead leaching> The world lead consumption in 2011 was over 10 million tons, of which about 80% was for the production of lead-acid batteries. These batteries contain lead amounts in the form of Pb, PbO2, and PbSO4. The most traditional method of recovering lead is pyrometallurgy, characterized by the addition of reducing agents (e.g. carbon powder, iron scrap, and sodium oxalate). This operation is carried out in ovens at temperatures above 1000 °C. As shown in He et al., Minerals 7, no. 6 (2017): 93, this leads to a high demand for energy processes.

[0010] On the other hand, hydrometallurgical processes to recover lead allow operation at lower temperatures and with lower energy consumption. Hydrometallurgical processes then do not produce sulfur dioxide, a gas that is harmful to the environment. Hydrometallurgical processes use desulfurizing agents such as sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, and sodium acetate. The purpose of these processes is to exchange sulfate ions with other anions to form insoluble salts. Once recovered, lead salts such as lead citrate can be calcined to produce lead oxide (Zarate-Gutierrez and Lapidus, Hydrometallurgy 144 (2014): 124-128).

[0011] <Desulfurization using citrate> In the specific case of using citrate, a mixture of citric acid and sodium citrate is beneficial for the leaching of lead sulfate and the subsequent crystallization of lead citrate.

[0012] Lead leaching in citrate solutions The solubility product constant of anglesite at 20°C is 6.31·10 -7 This indicates that the solubility of PbSO4 is significantly reduced. However, in the presence of concentrated citrate solutions, lead forms a series of soluble complexes. 2+ In solutions having a pH of 4.6 to 11.5, a wide variety of citrate complex species are present. Below a pH of 4.6, lead sulfate dominates, while above a pH of 11.5, lead hydroxide dominates.

[0013] He et al., Minerals 7, no. 6 (2017): 93, studied lead leaching for a paste with a lead sulfate to water weight ratio of 1:10 by adding 650 g / L sodium citrate at 35°C. These conditions allowed the conversion of more than 99% of the lead sulfate to lead citrate after 60 minutes of reaction time. Increasing the temperature to 95°C at a sodium citrate concentration of 300 g / L allowed an efficiency of nearly 99% to be achieved after 60 minutes of reaction time. However, a decrease in the production of lead citrate was observed when citric acid was introduced into the mixture. The optimum pH for the production of lead citrate was in the range of 6-7. A high lead leaching efficiency was also obtained from acid and lead-acid batteries at pH 5.5 with citric acid and ammonium agents. In the pH range of 5.2-5.5, the presence of lead citrate trihydrate ([Pb3(C6H5O7)2]·[3H2O]) was reported as the main species. In the higher pH range of 8-10, the recovery of lead as citrate is lower due to the formation of lead hydroxide. When the lead residue is rich in oxides such as PbO and PbO2, leaching is carried out at 20 °C for 15-60 min with molar ratios of citric acid to lead(II) and (IV) oxides of 1:1 and 4:1. In this case, leaching efficiencies of more than 99 wt.% are reached, yielding Pb(C6H6O7)·H2O as the main species (Sonmez and Kumar, Hydrometallurgy 95, no. 1-2 (2009), 82-86).

[0014] Pulp density is another important parameter for lead leaching with citrate solutions. In the range of 10-50 g / L anglesite pulp, a higher level of lead extraction of 90-94% was achieved at 10 g / L pulp consistency for a leaching solution of 1M sodium citrate solution, pH 7, at 600 rpm and 25°C. At higher pulp consistency, the amount of lead extracted was less.

[0015] Therefore, the hydrometallurgical desulfurization process is influenced by the diffusion of citrate ions in the lead paste within the reactor due to the high density of the lead paste, so it is important to design the reactor to maximize mass transfer in the system.

[0016] A technology based on lead recovery from lead waste using citric acid has been developed by Cambridge Enterprises Limited (WO2008056125A1). Essentially, the technology involves treating lead residues, including lead(II) oxide, lead(IV) oxide, and lead sulfate, with citric acid solution. Alternatively, the lead residues may be treated in combination with sodium citrate at pHs ranging from 1.4 to 6. Finally, hydrogen peroxide can be added as a reducing agent in a basic environment to accelerate the lead(IV) oxide leaching reaction to produce lead citrate (Sonmez and Kumar, Hydrometallurgy 95, no. 1-2 (2009), 82-86).

[0017] The present invention differs from the above application (WO2008056125A1) in that the pH required for leaching ranges from 5.33 to 8.8. In the present invention, a pH equal to 7 is preferably used. Furthermore, the present invention proposes recirculating the citrate solution obtained after the precipitation step with sodium carbonate to re-leaching the output metallurgical residue from the sulfuric acid leaching step.

[0018] <Germanium leaching> Germanium is a metal that is widely used in the fields of optical fiber, infrared fiber, solar cells, aviation, aerospace, and military industries. In general, germanium constitutes 1-7 ppm of the Earth's crust, with a total estimated amount of 8600 tons. For this reason, germanium is not abundant in the Earth's crust. Normally, germanium is associated with deposits of copper, lead, zinc, and carbon. However, deposits with high germanium content are limited. Most of the germanium is recovered from lignite (lignite) vapours subjected to pyrometallurgical processes and from lead-zinc smelting. However, pyrometallurgical processes are losing importance as they give rise to environmental problems related to the volatility of germanium(II) oxide and germanium sulfide.

[0019] There are various methods for recovering germanium from zinc electrolyte solutions, among which tannin precipitation, germanium tetrachloride distillation, flotation, activated carbon adsorption, precipitation, solvent extraction, and chelating resin adsorption are the main ones (US455332). Usually, germanium is present in the form of germanic acid Ge(OH)4, the most predominant species in the pH range 1-8. Meanwhile, at pH 9-13, the most predominant species is GeO(OH)3. 1- At pH levels above 13, the predominant species is GeO2(OH)2. 2- The first dissociation constant of germanic acid is 4.9 to 10 mol.L. -1 (logK diss =-9.31) (Wood and Iain, Ore Geology Reviews 28, no. 1 (2006): 57-102). [ka] [ka]

[0020] <Acid leaching> Germanium may be leached using H2SO4 at a temperature of 40-60°C for a period of 30 minutes, a H2SO4 concentration of 100 g / L, and a solid:liquid ratio of 1:4, in which case 78% of germanium can be recovered. At higher temperatures (on the order of 85°C), a H2SO4 concentration of 150 g / L, and a residence time of 1 hour, collective leaching of the various metals occurs. In this case, the extraction rate of germanium was 92.7% (Rutledge et al., Metals 5, no. 3 (2015): 1520-1542).

[0021] Patent application CN108486390A describes a method for separating germanium and gallium from germanium-gallium-containing material. In the first (first) step, germanium-gallium material is added to a 50-150 g / L H2SO4 solution in a ratio of 5-10% w / w. The pH is then adjusted to 1-3. The leaching solution is neutralized at pH 1-3. Zinc powder is then added to the neutralized solution at a temperature of 40-80°C. This gives a germanium concentrate and a liquid solution. Zinc powder is added again to the liquid solution at a temperature of 40-80°C. This gives a gallium residue and a liquid solution.

[0022] <Alkaline leaching> Patent application CN108300876A describes a method for leaching gallium and germanium from the slag of a process for obtaining zinc. In a first step, the slag is ground to a size of 50-100 microns. Then, a 0.1-1 mol / L H2SO4 solution is added with a liquid-solid ratio of 4-10:1 ml / g and a leaching temperature of 25-80°C for 0.25-4 h (hours) while stirring at 100-600 rpm. Then, a liquid-solid separation is carried out to obtain the leaching residue with H2SO4. In this case, a 0.2-2 mol / L hydrogen peroxide solution is added to the H2SO4 in a liquid-solid ratio of 4-10:1 mL / g. Then, 0.1-1 mol / L NaOH is added to adjust the pH of the leaching solution to 5.0-8.0. To obtain a germanium-rich leaching solution, the alkaline leaching is carried out at 25-80°C, 100-600 rpm, stirring for 0.25-4 h. In this case of the present invention, a sulfuric acid leaching step and a citric acid leaching step are necessary to make the germanium value leaching possible. This allows the removal of high levels of lead in the metallurgical residue, which in turn increases the germanium content in the residue due to the loss of mass in the citric acid leaching step. Furthermore, the presence of lead in the metallurgical residue from some processes (e.g. one process in the present application) can lead to an increase in soda consumption due to the conversion of lead sulfate to lead hydroxide. This can affect the germanium leaching yield.

[0023] <Ion exchange> After the leaching step, the germanium obtained needs to be concentrated by various techniques such as solvent extraction, precipitation with chelating agents or ion exchange. In the context of the present invention, an ion exchange step has always been considered, since it has a better recovery level and also makes it possible to increase the germanium concentration ratio with respect to other metals present in the leaching solution (e.g. lead, aluminum, silicon and arsenic).

[0024] Patent US4525332 describes the adsorption of a germanium-containing solution in an ion exchange resin consisting of a polymer with functional groups selected from secondary ammonium, tertiary ammonium and quaternary ammonium groups. The functional groups have a relative selectivity of germanium to antimony of 50:1, prior to elution of the germanium collected by the resin in an aqueous medium. The invention uses a resin containing N-methylglucamine as functional group. Patent US4525332 does not specify that the loading solution may contain silicon. In the application examples in patent US4525332, the loading is carried out with an acidic or slightly acidic solution. That is, the loading is carried out without an alkaline solution, unlike the examples of the invention. Silicon, being an element belonging to the same group as germanium, may interfere with the recycling of the loading solution used in the process. Similarly, the leaching of the resins used in the examples of patent US4525332 is carried out with H2SO4 or sodium hydroxide, depending on the resin used (as opposed to the leaching with hydrochloric acid in this specification), and requires a subsequent germanium tetrachloride distillation step. Furthermore, patent US4525332 does not teach how to remove silicon from the alkaline solution in order to recycle the solution as a leaching medium for metallurgical residues to obtain germanium. In this respect, the present invention differs from patent US4525332.

[0025] Patent application GB933563A teaches the treatment of an aqueous, neutral or slightly acidic solution of germanium in an ion exchange resin containing hydroxyphenyl groups from which germanium is recovered. The solution is subsequently eluted with a 7N hydrochloric acid solution and subsequently distilled, hydrolyzing germanium tetrachloride to produce germanium dioxide. Patent application GB933563A does not teach how to carry out the loading of a strong alkaline solution in the presence of silicon. Due to the presence of silicon in the solution, neutralization of the solution to carry out the loading of germanium in the resin is not self-evident. This is because if the pH is lowered, silicon will precipitate and may drag out the germanium. In this scenario, it is necessary to carry out the loading without prior precipitation and remove the silicon when the solution passes through the column and there is a shortage of germanium. Moreover, unlike patent application GB933563, in the present invention, the elution may be carried out with a solution having a Ge concentration of less than 6N, without any evidence of significant deficiency of germanium.

[0026] <Distillation of Germanium> Powell et al., J. Appl. Chem. 1951, 541-551, teaches the steep leaching of smelted powder with HCl to produce in situ germanium tetrachloride which can be distilled at a temperature of 84° C. One problem with this method is the presence of arsenic trichloride, because even if arsenic trichloride boils at 130° C., its vapor pressure at 84° C. is high enough to co-distill with germanium tetrachloride. In this sense, the present invention differs from the teachings of Powell in that there is a series of pre-leaching steps and ion exchange separations to suppress components such as arsenic, which are present in high concentrations that prevent the distillation of germanium.

[0027] Patent US3102786 teaches a continuous process for purifying germanium tetrachloride using a minimum concentration of 6N HCl solution and maintaining the column at a temperature of 83-110° C. The present invention differs from US3102786 in that the distillation can be carried out at acid concentrations below 6N and makes it possible to distill germanium in an efficient manner with a distillation rate of more than 95%.

[0028] Patent US2811418 teaches a process for purifying germanium tetrachloride, using an HCL solution with a concentration of 12N and saturated with chlorine gas, which allows the mixture to be separated into two phases, the heavier phase containing the purified germanium tetrachloride. The present invention differs from US2811418 in that the distillation can be carried out at acid concentrations lower than 12N, in particular lower than 5N, and allows the distillation of germanium in an efficient manner with a distillation rate of more than 95%, while minimizing arsenic drag.

[0029] <Germanium hydrolysis> Patent US3455645 discloses a method for the production of amorphous germanium dioxide, which involves the precipitation of germanium present in an aqueous solution having a pH of at least 5 and at most 9. In particular, the experiments disclosed by US3455645 teach the addition of germanium tetrachloride to a solution containing 10 NaOH x 90 water, with a pH of less than 8, preferably less than 6. Patent US3455645 differs from the present invention in that the resulting distilled solution is sent directly to a cooled reactor, in which the precipitation of germanium dioxide is verified. In the present invention, there is no need to control the pH to the pH value described in US3455645.

[0030] James E. Hoffmann, Extracting and Refining Germanium, Journal of Metals, June 1987, 42-45, states that at HCl concentrations below 5.5, germanium is found mainly as germanic acid, and 3 grams of water is sufficient to hydrolyze 1 g of germanium with a yield of 95%. The document also states that the deposition is preferably carried out at temperatures close to 0° C., and that germanium dioxide can be used to act as a nucleation center for germanium deposition. The present invention shows that high germanium deposition yields can be obtained without the need to use germanium dioxide as a seed to verify germanium deposition. The present invention shows that germanium dioxide can be deposited by using a sufficiently concentrated solution of germanium, and that low HCl concentrations have a detrimental effect on the germanium dioxide deposition process, compared to HCl concentrations close to 3.7N (135 g / L HCl).

[0031] [Drawing Description] FIG. 1 shows a process diagram of the method disclosed by the present invention.

[0032] FIG. 2 shows the distillation curve of germanium obtained from the ion-exchange solution.

[0033] Figure 3 shows the second distillation curve of germanium obtained from the first (primary) distilled solution. The second distillation refers to the distilled solution collected after the first distillation cycle.

[0034] Figure 4 shows the third distillation curve of germanium obtained from the secondary distilled solution. The third distillation refers to the distilled solution collected after the second distillation cycle.

[0035] [Description of the invention] Broadly speaking, the present invention describes a method for producing germanium from metallurgical residues.

[0036] In one preferred option, the present invention describes a method for producing germanium tetrachloride.

[0037] In a further preferred option, the present invention describes a method for producing a solid phase germanium concentrate.

[0038] In a further preferred option, the present invention describes a method for producing technical grade germanium dioxide having a concentration in the range of 60-70%.

[0039] In one broad aspect, the present invention describes a method for producing germanium concentrate from metallurgical residues, in particular from residues containing copper, iron, lead, silicon, and germanium (which may optionally contain components such as arsenic, antimony, and bismuth), comprising: step (i): subjecting the metallurgical residue (1) to copper leaching using a first acidic solution (2), a first leaching solution (3) rich in copper and iron, and optionally rich in arsenic, antimony, and bismuth; obtaining a first leached sludge (4) having a copper and iron reduced content, optionally arsenic reduced content, and a content rich in lead, silicon and germanium; step (ii): treating the first leached sludge (4) with a first solution (5) of carboxylate to leach the first leached sludge (4), A second leached sludge depleted in lead (6); obtaining a second leaching solution (7) rich in lead; - step (iii): adding a base to generate an alkaline leaching solution, and alkaline leaching the second leached sludge; a third leached sludge (9) having a reduced content of silicon and germanium; obtaining a third leaching solution (10) rich in germanium and silicon, and optionally rich in arsenic; - step (iv): loading the third leaching solution (10), rich in germanium and silicon, and optionally rich in arsenic, into an ion exchange column to capture germanium by the resin and obtain a fourth alkaline solution (11), depleted in germanium and rich in silicon; - step (v): washing the ion exchange column with water (12) to obtain a fifth solution (13) of column packing wash; - step (vi): eluting the ion exchange column with an HCl solution (14) to obtain a sixth elution solution (15) enriched in germanium; step (vii): distilling the sixth eluted solution (15) enriched in germanium, a seventh solution of germanium (16); obtaining an eighth solution (17) deficient in germanium; and step (viii): contacting said seventh solution (16) of germanium with a solution (18) of water for hydrolysis, a first germanium dioxide concentrate (19); a ninth solution (20) depleted in germanium; Contains:

[0040] In one preferred option, the metallurgical residue to be treated is a powder obtained from a metal smelting process or a powder obtained from a copper smelting process.

[0041] In a further preferred option, said metallurgical residue is subjected to a copper leaching process.

[0042] In a further preferred option, the metallurgical residue is subjected to leaching with H2SO4.

[0043] In one preferred option, the metallurgical residue to be treated contains the mineral species anglesite, coberlin, cuprospinel in the form of CuOFe2O3, zinc spinel in the form of ZnOFe2O3, magnetite, iron(III) oxide, pyrite, scorodite, mucovite, kaolinite, and lead(II) sulfate.

[0044] In a further preferred option, the copper contained in said metallurgical residue is present as copper sulfate, chalcosine, coberlin and cuprospinel in the form CuOFe2O3.

[0045] In a further preferred option, the copper contained in said metallurgical residue is present at least 50% in the form of cuprospinel of the CuOFe2O3 form.

[0046] In one preferred option, the silicon contained in the metallurgical residue is present as muscovite and kaolinite.

[0047] In another preferred option, the lead contained in the metallurgical residue is present as lead(II) sulfate, galena, or lead(II) oxide.

[0048] In a more preferred option, at least 95% of the lead is present as lead(II) sulfate.

[0049] In one preferred option, the first H2SO4 solution in step (i) may comprise H2SO4 and / or refinery wastewater.

[0050] In one preferred option, step (i) is carried out at a H2SO4 concentration of 150-300 g / L, more preferably at a H2SO4 concentration of 250 g / L.

[0051] In one preferred option, step (i) is carried out at a temperature between 50 and 130°C, more preferably at a temperature of 85°C.

[0052] In one preferred option, step (i) is carried out for a residence time of 3 to 12 hours, more preferably for a residence time of 6 hours.

[0053] In one preferred option, step (i) is carried out at a solids concentration of 5-20% w / w, more preferably at a solids concentration of 15% w / w.

[0054] In one preferred option, the carboxylate in leaching step (ii) is sodium citrate.

[0055] In one preferred option, in step (ii), the sodium citrate solution has a molarity of sodium citrate between 0.5 and 1 M.

[0056] In one preferred option, in step (ii), the first leached sludge is fed to a sodium citrate solution in a mass ratio of 1:9.

[0057] In one preferred option, step (ii) is carried out at a temperature between 20 and 60°C, more preferably at 40°C.

[0058] In one preferred option, step (ii) is carried out for a residence time of from 1 to 23 hours.

[0059] In one preferred option, step (ii) is carried out at a pH of 5.3 to 8.8, more preferably at pH 7.0.

[0060] In one preferred option, in step (ii) an acid corresponding to the carboxylate is added to adjust the pH.

[0061] In a further preferred option, in step (ii), citric acid is added to adjust the pH.

[0062] In a further preferred option, the pH adjustment in step (ii) is carried out using a 600-900 g / L citric acid solution.

[0063] In one preferred option, step (iii) is a germanium leaching step.

[0064] In one preferred option, the second base used in leaching step iv is selected from Mg(OH)2, KOH, or NaOH.

[0065] In one preferred option, the base added in step (iii) is added in a ratio of 5-10% w / w, more preferably 6.0% w / w, relative to the total mass of the alkaline leaching solution.

[0066] In one preferred option, the leaching reaction in step (iii) is carried out at a temperature between 70 and 150°C, more preferably at a temperature of 130°C.

[0067] In one preferred option, the leaching reaction in step (iii) is carried out for a residence time of 1 to 12 hours, more preferably for a residence time of 3 hours.

[0068] In one preferred option, step (iv) is carried out using a resin carrying nitrogen atom groups (N-donor groups).

[0069] In one preferred option, step (iv) is carried out by charging said third leaching solution, rich in germanium and silicon, in a ratio of 2 to 3 bed volumes (bed volume, flow rate factor).

[0070] In a further preferred option, step (iv) is carried out by charging said third leaching solution enriched in germanium and silicon in a ratio of 10 bed volumes.

[0071] In one preferred option, step (v) is carried out by loading said fourth loading wash solution at a ratio of 5 to 15 bed volumes.

[0072] In one preferred option, the elution in step (vi) is carried out with an HCl solution.

[0073] In a further preferred option, the elution in step (vi) is carried out using an HCl solution having a concentration between 2 and 8N, more preferably with an HCl solution having a concentration of 6N.

[0074] In one preferred option, step (vi) is carried out by charging the HCl solution in a ratio of 1 to 5 bed volumes, more preferably in a ratio of 3 bed volumes.

[0075] In another preferred option, said fourth alkaline solution, which is depleted in germanium, is subjected to a silicon removal process.

[0076] In another preferred option, hydrated lime or aluminum sulfate is added in the silicon removal process.

[0077] In a further preferred option, hydrated lime is added to the silicon removal process.

[0078] In a further preferred option, hydrated lime is added in a 1:1 molar ratio to the silicon contained in the fourth alkaline solution depleted of germanium to produce a regenerated alkaline solution and a solid consisting of calcium silicate.

[0079] In another preferred option, the silicon removal step is carried out at a temperature between 20 and 90°C.

[0080] In a further preferred option, the regenerated alkaline solution is recycled to the alkaline leaching in step (iii).

[0081] In another preferred option, the distillation in step (vii) is carried out at a temperature between 86.5°C and 107°C and a bulb temperature between 86.5°C and 108°C.

[0082] In another preferred option, the seventh germanium solution is redistilled 1-5 times to produce a concentrated germanium solution and a distilled HCl solution.

[0083] In a further preferred option, the seventh germanium solution is redistilled three times to produce a concentrated germanium solution and a distilled HCl solution.

[0084] In one preferred option, distilled HCl solution from a previous distillation step is recycled to increase the HCl concentration at the inlet to the still.

[0085] In another preferred option, the concentrated germanium solution is contacted with deionized water in a volume ratio of 1:1 to 1:6 to precipitate germanium as a germanium concentrate.

[0086] In a further preferred option, said germanium concentrate is germanium dioxide.

[0087] In a further preferred option, in step (viii), the concentrated germanium solution is contacted with deionized water at a temperature between 2 and 15°C.

[0088] In another preferred option, the concentrated germanium solution sent to the hydrolysis step has a germanium concentration of 8.1 to 24.8 g / L.

[0089] In another preferred option, the concentrated germanium solution delivered to the hydrolysis step has an HCl concentration of 55-135 g / L.

[0090] In one preferred option, the copper-rich first leaching solution is sent to a copper leaching process to smelt the powder.

[0091] In one preferred option, the copper-rich first leaching solution is sent to an arsenic removal process (arsenic reduction process).

[0092] In one preferred option, the arsenic removal process is selected from among those intended to produce ferric arsenate.

[0093] In a further preferred option, the arsenic removal process is a scorodite production process.

[0094] [Example of application] The following examples should be considered as embodiments of the present invention, and they should not be considered in any way as limiting the embodiments, since various adaptations (modifications) that can be performed to the embodiments are included within the scope of the subject matter claimed by the present invention.

[0095] [Sulfuric acid leaching] <Examples 1 to 7> 2.550-2.850 g of sulfuric acid solution with H2SO4 concentration of 150-250 g / L was prepared. The solution was placed in a 5 L glass reactor. Sludge previously subjected to a copper leaching process was added to a solid content of 5-10% w / w. Table 1 shows the mineralogy of the sludge. The reactor was stirred at 300 rpm at 85°C for 3-6 hours. After the reaction time was over, the pulp was filtered by a Buchner system. Table 2 shows the results. [Table 1] [Table 2]

[0096] <Examples 8 to 10> 2,550 g of 250 g / L H2SO4 solution was prepared. The solution was placed in a 4 L autoclave. Sludge previously subjected to a copper leaching process was added to a solid content of 15% w / w. The reactor was stirred at 300 rpm at 130°C for 1-6 hours. After the reaction time was over, the pulp was filtered by a Buchner system. Table 3 shows the results. [Table 3]

[0097] <Example 11> A refinery wastewater solution was prepared (Table 4). The H2SO4 concentration of the solution was adjusted to 250 g / L. The solution was placed in a 5 L glass reactor. 450 g of sludge previously subjected to a copper leaching process was added to produce a pulp with a solid content of 15% w / w. The reactor was stirred at 300 rpm for 6 hours at 85°C. After the reaction time was over, the pulp was filtered by a Buchner system. The results showed a Cu leaching yield of 72.0%, Fe leaching yield of 62.0%, As leaching yield of 71.5%, Zn leaching yield of 57.0%, and a mass loss (mass defect) of 38.5%. [Table 4]

[0098] [Citric acid leaching] <Example 12> A solution was prepared with 40 L of distilled water. 14 kg of sodium citrate was added to the solution and the pH was adjusted to 7.0 with 800 g / L of citric acid solution. After the test solution (reaction solution) had dissolved, 6 kg of leached sludge was added according to Example 3. The head sludge had a Pb content of 15.4%. Leaching was carried out at 20°C and stirred at 1,000 rpm for 9 hours. A Pb leaching efficiency of 94% was obtained. A leached sludge was thus obtained with a Pb content of 1.19% and a mass reduction of 24%.

[0099] <Examples 13 to 19> A solution was prepared using 2 L of distilled water with sodium citrate at a concentration of 323-368 g / L with a pH of 5.3-8.8. The pH was adjusted with 800 g / L of citric acid solution. After the reagent solution had dissolved, sludge treated according to Example 3 was added at a ratio of 1.2-2.3 g of sodium citrate per gram of sludge. The head sludge had a Pb content of 15.0-15.1%. Leaching was carried out at 30-60°C and stirred at 500-700 rpm for 2-4 hours. The results are shown in Table 5. [Table 5]

[0100] [Alkaline leaching] <Examples 20 to 28> Pulps were prepared using sodium hydroxide solutions with concentrations ranging from 5.4 to 8.7% w / w and leached sludge from a continuous copper and lead leaching process with a solid content of 5.0 to 7.0% w / w. The pulps were placed in a 4 L autoclave and heated to a temperature of 100 to 140°C at 600 rpm for 1 to 6 hours. After the leaching time was over, the pulps were cooled and filtered in a Buchner system. The results are shown in Table 6. [Table 6]

[0101] <Examples 29 to 30> A pulp was prepared by adding 420 g of sodium hydroxide and 350 g of leached sludge from a continuous copper and lead leaching process to 6,230 mL of water. This resulted in a concentration of 6.0% w / w NaOH and 5.0% w / w solids. The pulp was placed in a 10 L glass reactor and heated to 90°C and stirred at 900 rpm for 1-6 hours. After the leaching time was completed, the pulp was cooled and filtered in a Buchner system. [Table 7]

[0102] [Ion exchange] <Examples 31 to 38> A 12,500 mL solution with Ge concentration of 28-34 mg / L and Si concentration of 7.2-8.6 g / L was collected from alkaline leaching of the sludge. The solution was passed through an ion exchange column containing 400 mL of resin with nitrogen atom groups. The charge flow was 67-133 mL / min at a rate of 3.4-6.7 cm / min. The column was washed with 2000 mL of water, but no germanium elution was observed in any of the experimental cases. Germanium elution was performed with 1-2 bed volumes of HCl at a rate of 99 g / L. The elution flow was 20-67 mL / min at a rate of 1-3.4 cm / min. Finally, the column was washed with 2000 mL of water, but no Ge drag was observed in any of the experimental cases. The results are shown in Table 8. [Table 8]

[0103] [Silicon Removal] <Examples 39 to 41> 2,000 mL of charge solution obtained from the ion exchange column with a Si concentration of 18.5 g / L and a pH of 3.7 was collected. To this solution was added aluminum sulfate tetradecahydrate with Al / Si molar ratios of 0.43 and 1.0. The mixture was stirred at 400 rpm and at a temperature of 80° C. for 60 minutes. After the reaction was completed, the pulp was filtered, and a Si removal efficiency of 74% was obtained. Thus, the Si concentration could be reduced to 2.8 g / L and the Al concentration to 130 mg / L at pH 13.4. [Table 9]

[0104] <Examples 42 to 45> 4,200 g of IX (ion exchange) effluent solution with 17.4 g / L Si was collected in a 5 L reactor at a temperature of 20-70°C. Calcium hydroxide was added to the solution in a molar ratio range of 0.9-1.1 mol Ca / mol Si and stirred continuously at 300 rpm for 60 min. After the reaction was completed, the pulp was filtered through a No. 42 filter paper. [Table 10]

[0105] [Distillation of Germanium] <Example 46> A solution of 18,000 mL containing 454 mg / L Ge, 1,039 mg / L Pb, 294 mg / L Al, 7 mg / L As, and 150 g / L HCl was collected as the output solution of the ion exchange column elution. The solution was placed in a 20 L distillation balloon and heated to 108 °C. When the bulb temperature reached 108 °C, the output of the distillate solution was observed. The output was kept in fractions of 300-1,200 mL. The evaporated solution was condensed in a coil with circulating cold water at 5 °C and the solution was received in a jacketed collection cup with circulating water at 5 °C. A total of 100% of Ge contained in the distillate was collected. Thus, fractions with up to 1,960 mg / L Ge were obtained. 0.02% of the Pb present in the solution was transferred to the distillate. Meanwhile, 0.19 and 50% of Al and As, respectively, were dragged by the distillate.

[0106] <Example 47> 14,000 mL of a solution containing 1,060 mg / L Ge and 100 g / L HCl was collected as germanium distillate solution. The solution was placed in a 20 L distillation balloon and heated to 108° C. When the bulb temperature reached 108° C., the production of distillate solution was observed. The production was kept in fractions of 500 mL. The evaporated solution was condensed in a coil with circulating cold water at 5° C. and the solution was received in a jacketed collection cup with circulating water at 5° C. A total of 100% of the Ge contained in the distillate was collected. Thus, fractions with up to 4,330 mg / L Ge were obtained.

[0107] <Example 48> 16,000 mL of a solution containing 2,630 mg / L Ge and 117 g / L HCl was collected as germanium distillate solution. The solution was placed in a 20 L distillation balloon and heated to 108° C. When the bulb temperature reached 108° C., the production of distillate solution was observed. The production was kept in fractions of 1.000 mL. The evaporated solution was condensed in a coil with circulating cold water at 5° C. and the solution was received in a jacketed collection cup with circulating water at 5° C. A total of 100% of the Ge contained in the distillate was collected. In this way, fractions with up to 15,000 mg / L Ge were obtained.

[0108] <Example 49> 16,000 mL of a solution containing 10,400 mg / L Ge and 150 g / L HCl was collected as the germanium distillate solution. The solution was placed in a 20 L distillation balloon and heated to 108° C. When the bulb temperature reached 108° C., the production of a distillate solution was observed. The production was kept in fractions of 1.000 mL. The evaporated solution was condensed in a coil with circulating cold water at 5° C. and the solution was received in a jacketed collection cup with circulating water at 5° C. A total of 100% of the Ge contained in the distillate was collected. In this way, fractions with up to 25,000 mg / L Ge were obtained.

[0109] [Hydrolysis of germanium] <Examples 50 to 57> 250 mL of germanium solution was taken, with Ge concentration in the range of 4.1-24.8 g / L and HCl concentration of 54-134 g / L. The solution was placed in a 500 mL jacketed reactor, in which 5000 mL of water in a 1% volume mixture of ethylene glycol was circulated and cooled to 1° C. by a chiller. The solution was mechanically stirred at 250 rpm at a temperature of 2-3° C. for 5 hours. At the end of the process, the solution was filtered through a 0.45 μm filter paper. [Table 11] [Brief description of the drawings]

[0110] [Figure 1] 1 shows a process diagram of the method disclosed by the present invention. [Diagram 2] 1 shows a distillation curve of germanium obtained from an ion exchange solution. [Diagram 3] 2 shows the second distillation curve of germanium obtained from the first distillation solution. The second distillation refers to the distillation solution collected after the first distillation cycle. [Figure 4] 3 shows the third distillation curve of germanium obtained from the second distillation solution. The third distillation refers to the distillation solution collected after the second distillation cycle.

Claims

1. 1. A method for producing a germanium concentrate from metallurgical residues, the residues containing copper, iron, lead, silicon and germanium, and optionally arsenic, antimony and bismuth, comprising: i. copper leaching the metallurgical residue using H2SO4; a first leaching solution rich in copper and iron, and optionally rich in arsenic, antimony, and bismuth; obtaining a first leached sludge having reduced copper and iron, optionally reduced arsenic, and rich contents of lead, silicon and germanium; ii. leaching the first leached sludge with a first solution of sodium citrate; a second leached sludge that is lead-depleted; and a second leaching solution rich in lead; iii. Adding a base to the second leached sludge to generate an alkaline leaching solution, and alkaline leaching the second leached sludge; a third leached sludge having a reduced content of silicon and germanium; a third leaching solution rich in germanium and silicon, and optionally rich in arsenic; iv. loading the third leaching solution from step (iii) into an ion exchange column and capturing germanium by resin to obtain a fourth alkaline solution depleted in germanium and rich in silicon; v. washing the ion exchange column with water to obtain a fifth solution of column packing wash; vi. eluting the ion exchange column with an HCl solution to obtain a sixth eluted solution enriched in germanium; vii. Distilling the sixth eluted solution rich in germanium to obtain a seventh solution of germanium; an eighth solution deficient in germanium; and viii. contacting the seventh solution of germanium with a solution of water to hydrolyze and produce a first germanium dioxide concentrate; The method includes:

2. 2. The method of claim 1, wherein the metallurgical residue to be treated is a powder obtained by a metal smelting process.

3. The method of claim 2, wherein the metallurgical residue to be treated is a powder obtained by a copper smelting process.

4. 4. The method according to claim 1, wherein the metallurgical residue has been subjected to a copper leaching process prior to step (i).

5. The metallurgical residue to be treated contains the mineral species anglesite, coberlin, CuOFe ２ O ３ Cuprospinel in the form of ZnOFe ２ O ３ 4. The method of claim 1, wherein the inorganic filler includes zinc spinel in the form of zinc sulphate, magnetite, iron(III) oxide, pyrite, scorodite, mucovite, kaolinite, and lead(II) sulphate.

6. The copper contained in the metallurgical residue is present in the form of copper sulfate, chalcone, cobalt, and CuOFe. ２ O ３ 6. The method of claim 5, wherein the cuprospinel is in the form of cuprospinal.

7. 7. The method according to claim 1, wherein the silicon contained in the metallurgical residue is present as muscovite and kaolinite.

8. 8. The method according to claim 1, wherein the lead contained in the metallurgical residue is present as lead(II) sulfate, galena or lead(II) oxide.

9. 9. The method of claim 8, wherein at least 95% of the lead is present as lead (II) sulfate.

10. Step (i) is ２ SO ４ The method according to any one of claims 1 to 9, carried out at a concentration of 150 to 300 g / L.

11. The method according to any one of claims 1 to 10, wherein step (i) is carried out at a temperature of from 50 to 130°C.

12. 12. The method of any one of claims 1 to 11, wherein step (i) is carried out for a period of from 3 to 12 hours.

13. 13. A method according to any one of claims 1 to 12, wherein step (i) is carried out at a solids concentration of 5 to 20% w / w.

14. 14. The method according to any one of claims 1 to 13, wherein in step (ii) the sodium citrate solution has a molarity of sodium citrate of 0.5 to 1 M.

15. 15. The method according to any one of claims 1 to 14, wherein in step (ii) the first leached sludge is provided to the sodium citrate solution in a mass ratio of first leached sludge:sodium citrate solution of 1:

9.

16. The method according to any one of claims 1 to 15, wherein step (ii) is carried out at a temperature of from 20 to 60°C.

17. 17. The method of any one of claims 1 to 16, wherein step (ii) is carried out for a residence time of from 1 to 23 hours.

18. 18. The method of any one of claims 1 to 17, wherein step (ii) is carried out at a pH of 5.3 to 8.

8.

19. 19. The method of any one of claims 1 to 18, wherein in step (ii) citric acid is added to adjust the pH.

20. 20. The method of claim 19, wherein the pH adjustment in step (ii) is carried out using a citric acid solution of 600 and 900 g / L.

21. 21. The method of any one of claims 1 to 20, wherein step (iii) is a germanium leaching step.

22. 22. The process of any one of claims 1 to 21, wherein the base used in step (iii) is sodium hydroxide.

23. 23. The method according to any one of claims 1 to 22, wherein the base added in step (iii) is added in a ratio of 5 to 10% w / w, relative to the total mass of the alkaline leaching solution.

24. 24. The method according to any one of claims 1 to 23, wherein the leaching reaction in step (iii) is carried out at a temperature of 90 to 140°C.

25. 25. The method according to any one of claims 1 to 24, wherein the leaching reaction in step (iii) is carried out for a residence time of from 1 to 6 hours.

26. 26. The method of any one of claims 1 to 25, wherein step (iv) is carried out using an ion exchange resin having nitrogen atom groups.

27. 27. The method according to any one of claims 1 to 26, wherein step (iv) is carried out by loading said third leaching solution rich in germanium and silicon in a ratio of 2 to 3 bed volumes.

28. 28. The method of any one of claims 1 to 27, wherein step (v) is carried out with the water charged at a ratio of 5 to 15 bed volumes.

29. 29. The method of any one of claims 1 to 28, wherein the elution in step (vi) is carried out using an HCl solution.

30. 30. The method according to any one of claims 1 to 29, wherein the elution in step (vi) is carried out with an HCl solution having a concentration of 2 to 8 N.

31. 31. The method according to any one of claims 1 to 30, wherein step (vi) is carried out by charging the HCl solution at a ratio of 1 to 5 bed volumes.

32. 32. The method of any one of claims 1 to 31, wherein the fourth alkaline solution depleted in germanium is subjected to a silicon removal process.

33. 33. The method of claim 32, wherein hydrated lime is added in the silicon removal process.

34. to produce a regenerated alkaline solution and a solid comprising calcium silicate; 34. The method of any one of claims 1 to 33, wherein hydrated lime is added in a molar ratio of 1:1 to the silicon contained in the fourth alkaline solution that is depleted in germanium.

35. The method of any one of claims 1 to 34, wherein the silicon removal conditions are carried out at a temperature of 30 to 90°C.

36. 36. The method of any one of claims 1 to 35, wherein the regenerated alkaline solution is recycled to the alkaline leaching in step (iii).

37. 37. The process according to any one of claims 1 to 36, wherein the distillation in step (vii) is carried out at a temperature of 86.5 to 107°C and a bulb temperature of 86.5 to 108°C.

38. 38. The method of any one of claims 1 to 37, wherein the seventh germanium solution is redistilled 1 to 5 times to produce a concentrated germanium solution and a distilled HCl solution.

39. 39. The method of claim 38, wherein the seventh germanium solution is redistilled three times to produce a concentrated germanium solution and a distilled HCl solution.

40. 40. The method of any one of claims 1 to 39, wherein distilled HCl solution from a previous distillation step is recycled to increase the HCl concentration at the inlet to the still.

41. To precipitate germanium as a germanium concentrate, 41. The method of any one of claims 1 to 40, wherein the concentrated germanium solution provided in step (viii) is contacted with deionized water in a volume ratio of concentrated germanium solution:deionized water of 1:1 to 1:

6.

42. 42. The method of claim 41 , wherein the germanium concentrate obtained in step (viii) is germanium dioxide.

43. 43. The method of any one of claims 1 to 42, wherein in step (viii) the concentrated germanium solution is contacted with deionized water at a temperature of 2 to 15°C.

44. 44. The method of any one of claims 1 to 43, wherein the concentrated germanium solution delivered to the hydrolysis step has a germanium concentration of 8.1 to 24.8 g / L.

45. 45. The method of any one of claims 1 to 44, wherein the concentrated germanium solution delivered to step (viii) has an HCl concentration of 55 to 135 g / L.

46. 46. ​​The method of any one of claims 1 to 45, wherein the copper-rich first leaching solution is delivered to a copper leaching step of smelted powder.

47. 47. The method of any one of claims 1 to 46, wherein the first copper-rich leaching solution is sent to an arsenic removal process.

48. 48. The method of claim 47, wherein the arsenic removal process is selected from among those intended to produce ferric arsenate.

49. 49. The method of claim 48, wherein the arsenic removal process is a scorodite production process.

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