Process for producing germanium concentrate from metallurgical residues

Abstract

Process for the production of germanium concentrate from metallurgical residues, comprising: (i) leaching of copper from the metallurgical residues with a first acid solution, so as to obtain a first leach solution rich in copper and iron and optionally arsenic, antimony and bismuth, and a first leached sludge having reduced copper and iron content and optionally reduced arsenic content and being rich in lead, silicon and germanium, (ii) leaching of the first leached sludge, wherein the first leached sludge is treated with a first solution of sodium citrate, so as to obtain a second leached sludge lacking lead and a second leach solution rich in lead, (iii) alkaline leaching of the second leached sludge, wherein a base is added so as to form an alkaline leach solution, so as to obtain a third leached sludge having reduced silicon and germanium content and a third leach solution rich in germanium and silicon and optionally arsenic, (iv) feeding in an ion exchange column, wherein germanium is captured by a resin, so as to obtain a fourth alkaline solution lacking germanium and rich in silicon, (v) rinsing of the ion exchange column, wherein a fifth solution of column feeding rinse is obtained, (vi) elution of the ion exchange column with a solution of HCl, so as to obtain a sixth elution solution rich in germanium, (vii) distillation, wherein the sixth elution solution rich in germanium is distilled so as to obtain a seventh solution of germanium and an eighth solution lacking germanium, and (viii) hydrolysis, wherein the seventh solution of germanium is brought into contact with an aqueous solution to produce a first GeO2 concentrate.

Description

Technical Field

[0001] The present invention relates to a method for producing germanium concentrate from metallurgical residues, particularly from residues containing copper, iron, lead and germanium and optionally elements such as arsenic, antimony and bismuth.

[0002] In a more specific sense, metallurgical residues are powders derived from the metal smelting process.

[0003] In an even more specific sense, metallurgical residues are powders from the copper smelting process.

[0004] In even more specific terms, metallurgical residues or, in particular, smelting powders, are considered materials that have undergone leaching processes such as sulfuric acid leaching.

[0005] In a broad sense, germanium concentrate can be understood as a liquid fraction rich in germanium, mainly containing germanium tetrachloride, or solid germanium concentrate, which in a more specific sense may involve germanium dioxide.

[0006] In this disclosure, any metallurgical residues that have undergone a previous leaching process should be considered as sludge.

[0007] Background Art and Invention Content

[0008] Copper leaching

[0009] The copper in the sludge mainly comprises substances such as ferrite and / or CuFe2O4, ZnFe2O4 as zinc, and spinel in the form of FeFe2O4 as a corresponding fraction of iron. The leaching of these substances is based on temperature, acid concentration, and residence time, as described in the study by B.S. Boyanov et al. in World Academy of Science, Engineering and Technology, Vol. 9, 2015, 1592-1598, which conducted a synthetic ferrite leaching study of zinc, copper, and cadmium, evaluating the aforementioned variables. The results of this study show that ferrites dissolve better in HCl and H2SO4 at elevated temperatures and high acid concentrations.

[0010] At high acid concentrations, copper leaching exhibited asymptotic behavior with respect to leaching temperature, achieving a copper leaching yield greater than 90% within a temperature range between 85 and 90 °C once a 60-minute reaction time had elapsed in the sulfuric acid medium.

[0011] Lead leaching

[0012] In 2011, global lead consumption exceeded ten million tons, of which approximately 80% was intended for the manufacture of lead-acid batteries. These batteries contain lead in the forms of Pb, PbO2, and PbSO4. The most traditional method for recovering lead is the pyrometallurgical route, characterized by the addition of reducing agents such as carbon powder, iron filings, and sodium oxalate. This operation takes place in an oven at temperatures exceeding 1000°C, resulting in a high-energy-demand process. (He et al., Minerals 7, Vol. 6 (2017): 93).

[0013] On the other hand, hydrometallurgical pathways for lead recovery allow operation at lower temperatures, reducing energy consumption and eliminating the production of sulfur dioxide, a harmful gas. Hydrometallurgical pathways use desulfurizing agents such as sodium carbonate, ammonium carbonate, sodium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, and sodium acetate. The purpose of these processes is to exchange other anions with sulfate ions to form insoluble salts. Once recovered, lead salts such as lead citrate can be calcined to produce lead oxide (Zárate-Gutiérrez and Lapidus, Hydrometallurgy 144(2014):124-128).

[0014] Desulfurization using citrate

[0015] In the special 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.

[0016] Leaching lead in citrate solution

[0017] The solubility product constant of lead alum at 20℃ is 6.31·10⁻⁶. -7 This indicates that the solubility of PbSO4 is quite low. However, in the presence of concentrated citrate solution, lead forms a series of soluble complexes. In a 0.12 M Pb... 2+ In solutions containing this compound, a variety of citrate complexes exist in the pH range of 4.6 to 11.5. Below pH 4.6, lead sulfate is predominantly present, while above pH 11.5, lead hydroxide is predominantly present.

[0018] He et al., Minerals 7, Vol. 6 (2017): 93, studied lead leaching with a lead sulfate to water weight ratio of 1:10 using a paste with 650 g / L sodium citrate at 35 °C. These conditions allowed for greater than 99% conversion of lead sulfate to lead citrate after a 60 min reaction time. Increasing the temperature up to 95 °C allowed for near 99% efficiency after a 60 min reaction time at a sodium citrate concentration of 300 g / L. However, a decrease in lead citrate production was observed when citric acid was introduced into the mixture. The optimal pH for lead citrate production was in the range of 6 to 7. Increased lead leaching efficiency was also obtained from lead-acid batteries at a pH of 5.5 using citric acid and ammonium reagent. In the pH range of 5.2 to 5.5, the presence of lead citrate trihydrate ([Pb3(C6H5O7)2]·[3H2O]) was reported as the dominant substance. At higher pH ranges of 8 to 10, lead recovery to citrate was lower due to the formation of lead hydroxide. When lead residues are rich in oxides such as PbO and PbO2, leaching is carried out at 20°C with citric acid reacting with lead(II) oxide and lead(IV) oxide in molar ratios of 1:1 and 4:1 for 15 and 60 min, respectively, achieving a leaching efficiency of over 99% by weight, and obtaining Pb(C6H6O7)·H2O as the main substance (Sonmez and Kumar, Hydrometallurgy 95, Vol. 1-2 (2009), 82-86).

[0019] Slurry density is another important parameter for lead leaching with citrate solution. In the range of 10 to 50 g / L lead alum slurry, leaching with 1M, pH 7 sodium citrate solution at 600 rpm and 25°C achieves a relatively high lead extraction level of 90 to 94% at a slurry concentration of 10 g / L. Higher slurry concentrations result in less lead extraction.

[0020] Therefore, hydrometallurgical desulfurization methods are affected by the diffusion of citrate ions in the lead paste within the reactor due to the increased density of the lead paste. In this context, the key is to design a reactor that maximizes mass transfer within the system.

[0021] Cambridge Enterprise Limited has developed a technology for recovering lead from lead waste using citric acid (WO2008056125A1). This technology essentially involves treating lead residues containing lead(II), lead(IV), and lead sulfate with a citric acid solution, alternatively combined with sodium citrate at a pH ranging from 1.4 to 6. Ultimately, hydrogen peroxide can be added as a reducing agent in an alkaline environment to promote the leaching reaction of lead(IV) oxide, thereby producing lead citrate (Sonmez and Kumar, Hydrometallurgy 95, Vol. 1-2 (2009), 82-86).

[0022] This invention differs from application (WO2008056125A1) because the pH required for leaching varies from 5.33 to 8.8, with pH 7 being preferred within this range. Furthermore, this invention proposes recycling the citric acid solution obtained after the sodium carbonate precipitation step to leach the metallurgical residues from the sulfuric acid leaching step again.

[0023] Germanium leaching

[0024] Germanium is a metal widely used in optical fibers, infrared fibers, photovoltaic cells, and the aerospace and military industries. Typically, germanium is not abundant in the Earth's crust, constituting 1-7 ppm, with an estimated total of 8,600 tons. Germanium is usually associated with copper, lead, zinc, and carbon deposits; however, deposits with high germanium content are limited. Most germanium is recovered from lignite steam and lead-zinc ore smelting via pyrometallurgical processes. However, pyrometallurgical processes lose their importance whenever they generate environmental problems associated with the volatility of germanium(II) oxide and germanium sulfide.

[0025] Various methods exist for recovering germanium from zinc electrolyte solutions, with tannic acid precipitation, germanium tetrachloride distillation, flotation, activated carbon adsorption, precipitation, solvent extraction, and chelating resin adsorption being particularly prominent (US 455332). Germanium typically exists as germanic acid Ge(OH)4, which is the predominant form in the pH range of 1 to 8; however, between pH 9 and 13, GeO(OH)3 is the most predominant form. 1- Furthermore, the dominant substance at pH values ​​greater than 13 is GeO2(OH)2. 2- The first dissociation constant of germanic acid is 4.9–10 mol·L⁻¹ (log K). 解离 =-9.31)(Wood and Iain, OreGeology Reviews 28, Vol. 1 (2006):57-102).

[0026]

[0027] acid leaching

[0028] Germanium can be leached using H2SO4 at a concentration of 100 g / L for 30 min at a temperature between 40 and 60 °C with a solid-liquid ratio of 1:4, recovering 78% of the germanium. At higher temperatures, approximately 85 °C, with an H2SO4 concentration of 150 g / L and a residence time of 1 h, mixed leaching of different metals was produced, with a germanium extraction rate of 92.7% (Rutledge et al., Metals 5, Vol. 3 (2015): 1520-1542).

[0029] Patent application CN108486390A describes a method for separating germanium and gallium from germanium- and gallium-containing materials. In a first step, germanium and gallium materials are added to a solution of 50 to 150 g / L H2SO4 at a ratio of 5 to 10% w / w, and the pH is subsequently adjusted to between 1 and 3. The leaching solution is neutralized at pH 1 to 3 for subsequent addition of zinc powder to the neutralized liquid at a temperature of 40-80°C to obtain a germanium concentrate and a liquid solution. Zinc powder is then added again to this liquid solution at a temperature of 40-80°C to obtain a gallium residue and a liquid solution.

[0030] alkaline leaching

[0031] Patent application CN108300876A describes a method for leaching gallium and germanium from slag used in methods for obtaining zinc. In a first step, the slag is ground to a size of 50-100 micrometers, and then a 0.1-1 mol / L H₂SO₄ solution is added at a liquid-to-solid ratio of 4-10:1 mL / g, and the mixture is stirred at 100-600 rpm for 0.25-4 h at a leaching temperature of 25-80°C. Solid-liquid separation is then performed to obtain the leaching residue and H₂SO₄, to which a 0.2-2 mol / L hydrogen peroxide solution is added at a liquid-to-solid ratio of 4-10:1 mL / g, and 0.1-1 mol / L NaOH is added to adjust the pH of the leaching solution to 5.0-8.0. This alkaline leaching is carried out at 25-80°C with stirring at 100-600 rpm for 0.25-4 h to obtain a germanium-rich leaching solution. In the case of this invention, in order to leach the useful germanium components, sulfuric acid leaching and citric acid leaching steps are necessary to remove high levels of lead from the metallurgical residues and thus increase the germanium content in the residues due to mass loss of the metallurgical residues during the citric acid leaching step. Furthermore, the presence of lead in the metallurgical residues, through the method described in one of the applications, will lead to greater consumption of soda ash as lead sulfate is converted to lead hydroxide, which will affect the germanium leaching yield.

[0032] Ion exchange

[0033] Following the leaching step, the obtained germanium must be concentrated using various alternatives, such as solvent extraction, precipitation with a chelating agent, or ion exchange. For the purposes of this invention, an ion exchange step is considered each time because it offers better recovery levels and also allows for a higher concentration of germanium relative to other metals present in the leaching solution, such as lead, aluminum, silicon, and arsenic.

[0034] US Patent 4525332 describes the adsorption of a germanium-containing solution in an ion exchange resin composed of a polymer having functional groups selected from secondary ammonium, tertiary ammonium, and quaternary ammonium groups, the ion exchange resin having a relative selectivity of germanium to antimony of 50:1, followed by elution of the germanium collected through the resin in an aqueous medium. This invention utilizes a resin containing N-methylglucosamine as a functional group. US Patent 4525332 does not specify that the feed solution may contain silicon. In the application examples of US Patent 4525332, the feed is carried out with an acidic or slightly acidic solution instead of an alkaline solution, as shown in this invention. Silicon can interfere when a recycled feed solution is used in this method, as it belongs to the same group as germanium. In the same manner, depending on the resin used, elution with H₂SO₄ or sodium hydroxide is carried out with the resin used in the examples of US Patent 4525332, in contrast to the elution with HCl carried out herein and necessary for the subsequent steps of germanium tetrachloride distillation. Furthermore, this invention differs from patent US4525332 because the latter does not teach how to remove silicon from an alkaline solution so that the solution can be recycled as a leaching medium for metallurgical residues to obtain germanium.

[0035] Patent application GB933563A teaches the treatment of a neutral or slightly acidic aqueous solution of germanium in an ion exchange resin containing hydroxyphenyl groups for subsequent elution with a 7N hydrochloric acid solution for subsequent distillation, and the hydrolysis of germanium tetrachloride to produce germanium dioxide. Patent application GB933563A does not teach how to feed a strongly alkaline solution in the presence of silicon. Because silicon is present in the solution, neutralizing the solution to allow for germanium feeding in the resin is not obvious, as silicon will precipitate and drag germanium down when the pH is lowered. In this case, once the solution has passed through the column and germanium is lacking, it is necessary to feed and remove silicon without prior precipitation. Furthermore, unlike patent application GB933563, in this invention, elution can be performed with a solution containing a Ge concentration less than 6N without evidence of significant germanium loss.

[0036] Germanium distillation

[0037] Powell et al., J. Appl. Chem. 1951, 541-551, taught the use of HCl to leach smelted powders to produce germanium tetrachloride in situ, which could be distilled at 84°C. One problem with this method was the presence of arsenic trichloride, because even when it boils at 130°C, its vapor pressure is high enough at 84°C to distill it together with germanium tetrachloride. In this case, the present invention differs from Powell's teachings in that it involves a series of prior leaching steps and ion exchange separations that prevent elements such as arsenic from being present in elevated concentrations that would interfere with germanium distillation.

[0038] US3102786 teaches the purification of germanium tetrachloride using an HCl solution with a minimum concentration of 6N, and a continuous process maintaining the column at a temperature between 83 and 110°C. This invention differs from US3102786 because distillation can be carried out at acid concentrations less than 6N, allowing for efficient distillation of germanium with a distillation rate exceeding 95%.

[0039] US Patent 2811418 teaches a method for purifying germanium tetrachloride using a 12N concentration and chlorine-saturated HCl solution, allowing the mixture to separate into two phases, with the heavier phase containing purified germanium tetrachloride. This invention differs from US 2811418 because distillation can be carried out at acid concentrations below 12N, particularly below 5N, allowing for efficient distillation of germanium with a distillation rate exceeding 95%, minimizing arsenic drag.

[0040] Germanium hydrolysis

[0041] US Patent 3455645 discloses a method for producing amorphous germanium dioxide, characterized by precipitating germanium present in an aqueous solution, wherein the pH is at least 5 and at most 9. Specifically, US 3455645 discloses experimental teachings to add germanium tetrachloride to a solution containing 10 parts NaOH / 90 parts water until the pH is below 8, or preferably below 6. US Patent 3455645 differs from the present invention because the resulting distilled solution is directly fed to a cooling reactor, where germanium dioxide precipitation has been confirmed, without the need to control the pH to the value specified in US 3455645.

[0042] James E. Hoffmann, in *Extracting and Refining Germanium*, *Journal of Metals*, July 1987, pp. 42-45, states that germanium is primarily found as germanic acid at HCl concentrations less than 5.5 N, and that 3 g of water is sufficient to hydrolyze 1 g of germanium with a yield of 95%. Furthermore, he states that precipitation is preferably carried out at temperatures close to 0 °C, and that germanium dioxide can be used as a nucleation site for the germanium precipitate. This invention demonstrates that an improved germanium precipitation yield can be obtained without using germanium dioxide as a seed crystal to confirm the precipitation. This invention proves that germanium dioxide can be precipitated using a sufficiently concentrated germanium solution, and that low HCl concentrations negatively impact the germanium dioxide precipitation process compared to HCl concentrations close to 3.7 N (135 g / L HCl). Attached Figure Description

[0043] Figure I shows a flowchart of the method disclosed in this invention.

[0044] Figure II shows the distillation curve of germanium from the ion-exchange solution.

[0045] Figure III shows the second distillation profile of germanium from the first distillation solution. The second distillation refers to the distillate solution collected after the first distillation cycle.

[0046] Figure IV shows the tertiary distillation profile of germanium from the secondary distillation solution. The tertiary distillation refers to the distillate solution collected after the second distillation cycle. Detailed Implementation

[0047] In a broad sense, this invention describes a method for producing germanium from metallurgical residues.

[0048] In a preferred embodiment, the present invention describes a method for producing germanium tetrachloride.

[0049] In a still more preferred embodiment, the present invention describes a method for producing solid germanium concentrate.

[0050] In a still more preferred embodiment, the present invention describes a method for producing technical grade germanium dioxide with a concentration range of 60-70%.

[0051] In a broad sense, this invention describes a method for producing germanium concentrate from metallurgical residues, particularly from residues containing copper, iron, lead, silicon, and germanium, and optionally containing elements such as arsenic, antimony, and bismuth, characterized in that it comprises:

[0052] The step (i) involves leaching copper from the metallurgical residue (1) using a first acid solution (2) to obtain a first leaching solution (3) rich in copper and iron and optionally arsenic, antimony and bismuth, and a first leached sludge (4) rich in lead, silicon and germanium with reduced copper and iron content and optionally reduced arsenic content.

[0053] Step (ii) involves leaching the first leached sludge (4) with a first carboxylate solution (5) to obtain a second leached sludge (6) lacking lead and a second leaching solution (7) rich in lead.

[0054] The step (iii) of alkaline leaching of the second leached sludge, wherein alkali (8) is added to form an alkaline leaching solution to obtain a third leached sludge (9) with reduced silicon and germanium content and a third leaching solution (10) rich in germanium and silicon and optionally arsenic.

[0055] Step (iv) involves feeding a third leaching solution (10) rich in germanium and silicon, and optionally arsenic, into an ion exchange column, wherein germanium is captured by the resin to obtain a fourth alkaline solution (11) that is germanium-deficient and rich in silicon.

[0056] Step (v) involves rinsing the ion exchange column with water (12), thereby obtaining a fifth rinse solution (13) for the column feed.

[0057] Step (vi) involves eluting the ion exchange column with HCl solution (14) to obtain a sixth elution solution (15) rich in germanium.

[0058] The distillation step (vii) involves distilling the germanium-rich sixth elution solution (15) to obtain a germanium-rich seventh solution (16) and a germanium-deficient eighth solution (17), and

[0059] The hydrolysis step (viii) involves contacting a seventh solution (16) of germanium with an aqueous solution (18) to produce a first GeO2 concentrate (19) and a ninth solution (20) lacking germanium.

[0060] In a preferred embodiment, the metallurgical residue to be treated is a powder obtained through a metal smelting process or a powder obtained through a copper smelting process.

[0061] In an even more preferred embodiment, the metallurgical residue has undergone a copper leaching process.

[0062] In an even more preferred embodiment, the metallurgical residue has been leached with H2SO4.

[0063] In a preferred embodiment, the metallurgical residue to be processed comprises the minerals pyrite, indigo chalcopyrite, copper-iron spinel in the form of CuOFe2O3, zinc spinel in the form of ZnOFe2O3, magnetite, iron oxide (III), pyrite, spodumene, mucovita, kaolinite, and lead sulfate (II).

[0064] In an even more preferred embodiment, the copper contained in the metallurgical residue exists as copper sulfate, calcosina, indigo chalcopyrite, and copper-iron spinel in the form of CuOFe2O3.

[0065] In an even more preferred embodiment, at least 50% of the copper in the metallurgical residue is present in the form of copper-iron spinel in the form of CuOFe2O3.

[0066] In a preferred embodiment, the silicon contained in the metallurgical residue exists as muscovite and kaolinite.

[0067] In another preferred embodiment, the lead contained in the metallurgical residue is present as lead(II) sulfate, galena, or lead(II) oxide.

[0068] In an even more preferred embodiment, at least 95% of the lead is lead(II) sulfate.

[0069] In a preferred embodiment, the first H2SO4 solution in step (i) may contain H2SO4 and / or refining wastewater.

[0070] In a preferred embodiment, step (i) is carried out when the H2SO4 concentration is between 150 and 300 g / L, more preferably at a H2SO4 concentration of 250 g / L.

[0071] In a preferred embodiment, step (i) is performed at a temperature between 50 and 130°C, more preferably at a temperature of 85°C.

[0072] In a preferred embodiment, step (i) is performed for a duration between 3 and 12 hours, more preferably for a duration of 6 hours.

[0073] In a preferred embodiment, step (i) is carried out at a solid concentration between 5% and 20% w / w, more preferably at a solid concentration of 15% w / w.

[0074] In a preferred embodiment, in the leaching step (ii), the carboxylate is sodium citrate.

[0075] In a preferred embodiment, the sodium citrate solution in step (ii) has a molar concentration of sodium citrate between 0.5 and 1 M.

[0076] In a preferred embodiment, in step (ii), the first leachable sludge is added to the sodium citrate solution at a mass ratio of 1:9.

[0077] In a preferred embodiment, step (ii) is carried out at a temperature between 20 and 60°C, more preferably at 40°C.

[0078] In a preferred embodiment, step (ii) is performed for a stay time between 1 and 23 hours.

[0079] In a preferred embodiment, step (ii) is carried out at a pH between 5.3 and 8.8, more preferably at a pH of 7.0.

[0080] In a preferred embodiment, in step (ii), the corresponding acid of the carboxylate is added to adjust the pH.

[0081] In an even more preferred embodiment, in step (ii), citric acid is added to adjust the pH.

[0082] In an even more preferred embodiment, the pH adjustment in step (ii) is performed using a citric acid solution with a concentration between 600 and 900 g / L.

[0083] In a preferred embodiment, step (iii) is a germanium leaching step.

[0084] In a preferred embodiment, the second alkali used in the leaching of step iv is selected from Mg(OH)2, KOH, or NaOH.

[0085] In a preferred embodiment, the alkali added in step (iii) is added at a ratio between 5% and 10% w / w relative to the total mass of the alkaline leaching solution, more preferably at a ratio of 6.0% w / w.

[0086] In a preferred embodiment, the leaching reaction of step (iii) is carried out at a temperature between 70 and 150°C, more preferably at a temperature of 130°C.

[0087] In a preferred embodiment, the leaching reaction in step (iii) is carried out for a residence time between 1 and 12 hours, more preferably for a residence time of 3 hours.

[0088] In a preferred embodiment, step (iv) is performed using a resin having nitrogen atom groups (N donor groups).

[0089] In a preferred embodiment, a third leaching solution rich in germanium and silicon is fed at a bed volume ratio between 2 and 30 to carry out step (iv).

[0090] In an even more preferred embodiment, a third leaching solution rich in germanium and silicon is fed at a ratio of 10 bed volumes to carry out step (iv).

[0091] In a preferred embodiment, the fourth solution of the feed rinse is fed at a ratio of 5 to 15 bed volumes to perform step (v).

[0092] In a preferred embodiment, the elution step (vi) is performed using an HCl solution.

[0093] In a preferred embodiment, the elution step (vi) is performed using an HCl solution with a concentration between 2 and 8N, more preferably 6N.

[0094] In a preferred embodiment, step (vi) is performed by feeding an HCl solution at a bed volume ratio between 1 and 5, more preferably at a bed volume ratio of 3.

[0095] In another preferred embodiment, the fourth alkaline solution lacking germanium undergoes a silicon removal process.

[0096] In another preferred embodiment, quicklime or aluminum sulfate is added during the silicon removal process.

[0097] In an even more preferred embodiment, quicklime is added during the silicon removal process.

[0098] In an even more preferred embodiment, quicklime is added at a molar ratio of 1:1 relative to the silicon contained in the fourth alkaline solution lacking germanium, in order to produce a regenerated alkaline solution and a solid composed of calcium silicate.

[0099] In another preferred embodiment, the silicon removal step is performed at a temperature between 20 and 90°C.

[0100] In an even more preferred embodiment, the regenerated alkaline solution is recycled to the alkaline leaching step (iii).

[0101] In another preferred embodiment, the distillation step (vii) is carried out at a temperature between 86.5°C and 107°C, and at a ball temperature between 86.5°C and 108°C.

[0102] In another preferred embodiment, the seventh germanium solution is redistilled 1 to 5 times to produce a concentrated germanium solution and a distilled HCl solution.

[0103] In an even more preferred embodiment, the seventh germanium solution is redistilled three times to produce a concentrated germanium solution and a distilled HCl solution.

[0104] In a preferred embodiment, the distilled HCl solution is recycled in the previous distillation step in order to increase the HCl concentration at the distiller inlet.

[0105] In another preferred embodiment, the concentrated germanium solution is contacted with deionized water at a volume ratio between 1:1 and 1:6 to precipitate germanium as a germanium concentrate.

[0106] In an even more preferred embodiment, the germanium concentrate is germanium dioxide.

[0107] In an even more preferred embodiment, the germanium solution concentrated in step (viii) is contacted with deionized water at a temperature between 2 and 15°C.

[0108] In another preferred embodiment, the concentrated germanium solution fed to the hydrolysis step has a germanium concentration between 8.1 and 24.8 g / L.

[0109] In another preferred embodiment, the concentrated germanium solution fed to the hydrolysis step has an HCl concentration between 55 and 135 g / L.

[0110] In a preferred embodiment, a copper-rich first leaching solution is fed into a copper leaching process for smelting powder.

[0111] In a preferred embodiment, a copper-rich first leaching solution is fed to an arsenic removal (abatimiento) process.

[0112] In a preferred embodiment, the arsenic removal process is selected from those considering the production of ferric arsenate.

[0113] In an even more preferred embodiment, the arsenic removal process is the onion stone production process.

[0114] Application Examples

[0115] The following embodiments should be considered as implementations of the present invention and should not be considered as limitations on the present invention, as various modifications that may be made thereto should be covered within the scope of the subject matter claimed in the present invention.

[0116] Sulfuric acid leaching

[0117] Examples 1 to 7

[0118] A sulfuric acid solution with H₂SO₄ concentrations between 150 and 250 g / L was prepared at a concentration between 2550 and 2850 g, and arranged in a 5 L glass reactor. Sludge previously subjected to a copper leaching process was added to the reactor until the solids content was between 5% and 10% w / w. The mineralogical composition of the sludge is shown in Table 1. The reactor was stirred at 300 rpm for 3 to 6 hours at 85 °C. Once the reaction time was complete, the slurry was filtered through a Büchner system. The results are shown in Table 2.

[0119] Table 1 Mineralogical composition of sludge

[0120] substance unit value <![CDATA[PbSO4]]> ％ 12.84 PbS ％ 0.1 PbO ％ 0.1 <![CDATA[CuSO4]]> ％ 2.54 <![CDATA[Cu2S]]> ％ 0.63 CuS ％ 4.02 CuO ％ 0.71 <![CDATA[CuOFe2O3]]> ％ 15.09 <![CDATA[ZnOFe2O3]]> ％ 4.46 ZnS ％ 2.94 <![CDATA[Fe3O4]]> ％ 4.74 <![CDATA[Fe2O3]]> ％ 4.91 <![CDATA[FeS2]]> ％ 6.32 <![CDATA[Ag2S]]> ％ 0.1 <![CDATA[FeAsO4*2H2O]]> ％ 5.18 <![CDATA[Bi2O3]]> ％ 0.59 <![CDATA[Sb2O3]]> ％ 0.5 <![CDATA[KAl3Si3O 10 (OH)2]]> ％ 7.01 <![CDATA[Al2Si2O3(OH)4]]> ％ 2.92 Ge g / ton 548

[0121] Table 2 Cu leaching results from Examples 1 to 7

[0122] Variables / Examples unit 1 2 3 4 5 6 7 <![CDATA[H2SO4 concentration]]> g / L 150 250 150 250 250 150 250 Solid content %w / w 5 5 15 15 15 20 20 Leaching time h 6 6 6 3 6 6 6 Cu leaching yield ％ 75.9 76.1 68.0 60 69.7 64.8 67.7

[0123] Examples 8 to 10

[0124] 2550 g of a 250 g / L H₂SO₄ solution was prepared and placed in a 4 L pressure vessel, to which sludge from a previous copper leaching process was added until the solids content was 15% w / w. The reactor was stirred at 130 °C and 300 rpm for 1 to 6 hours. Once the reaction time was complete, the slurry was filtered through a Büchner system. The results are shown in Table 3.

[0125] Table 3. Cu leaching results in Examples 8 to 10

[0126] Variables / Examples unit 8 9 10 Leaching time h 1 3 6 Cu leaching yield ％ 75.9 76.1 82.0 quality loss ％ 35.0 41.0 42.0

[0127] Example 11

[0128] Prepare a refined wastewater solution (Table 4), adjust its H2SO4 concentration to 250 g / L, and arrange it in a 5 L glass reactor. Add 450 g of sludge previously subjected to a copper leaching process to produce a slurry with 15% w / w solids. Stir the reactor at 85 °C and 300 rpm for 6 hours. Once the reaction time is complete, filter the slurry using a Büchner system. The results show that the leaching yield of Cu is 72.0%, Fe is 62.0%, As is 71.5%, and Zn is 57.0% with a mass loss of 38.5%.

[0129] Table 4 Composition of Refining Wastewater

[0130]

[0131]

[0132] Citric acid extraction

[0133] Example 12

[0134] A solution was prepared using 40 L of distilled water, to which 14 kg of sodium citrate was added, and the pH was adjusted to 7.0 with an 800 g / L citric acid solution. Once the reagent was dissolved, 6 kg of leached sludge was added as in Example 3. The sludge (La borra decabeza) had a Pb content of 15.4%. Leaching was carried out at 20 °C with stirring at 1,000 rpm for 9 h. A Pb leaching efficiency of 94% was obtained, resulting in leached sludge with a mass reduction of 24% and a Pb content of 1.19%.

[0135] Examples 13 to 19

[0136] A solution was prepared using 2 L of distilled water, wherein the concentration of sodium citrate was between 323 and 368 g / L, and the pH was between 5.3 and 8.8. The pH was adjusted with an 800 g / L citric acid solution. Once the reagent was dissolved, the sludge treated according to Example 3 was added at a ratio of sodium citrate / g sludge between 1.2 and 2.3 g. The head sludge had a Pb content between 15.0 and 15.1%. Leaching was carried out between 30 and 60 °C and stirred at 500 and 700 rpm for a duration between 2 and 4 h. The results are shown in Table 5.

[0137] Table 5 Results of Citric Acid Extraction in Examples 13 to 19

[0138] Variables / Examples unit 13 14 15 16 17 18 19 Sodium citrate: sludge ratio g:g 2.3 2.3 2.3 2.3 1.2 2.3 2.3 Sodium citrate concentration g / L 350 350 350 350 323 368 368 pH 8.8 8.8 8.8 5.3 5.6 5.3 5.3 temperature ℃ 30 40 60 40 40 40 40 Stir rpm 500 500 500 500 500 500 500 Duration of stay H 4 2 2 4 4 4 4 Pb in the head ％ 15.1 15.1 15.1 15.1 15.1 15.0 15.0 Pb in residue ％ 2.1 2.2 1.6 0.6 0.9 0.7 0.9 Pb leaching yield ％ 90 89 92 97 96 97 96 quality loss ％ 24 25 24 26 28 32 29

[0139] alkaline leaching

[0140] Examples 20 to 28

[0141] Slurries were prepared using sodium hydroxide solutions with concentrations between 5.4% and 8.7% w / w and leached sludge that had undergone continuous copper and lead leaching processes, with solid content between 5.0% and 7.0% w / w. The slurries were placed in a 4 L autoclave and heated to temperatures between 100 and 140 °C at 600 rpm for 1 to 6 hours. Once the leaching time was complete, the slurries were cooled and filtered using a Büchner system. The results are shown in Table 6.

[0142] Table 6 Results of Examples 20 to 28

[0143]

[0144]

[0145] Examples 29 and 30

[0146] A slurry was prepared using 6230 mL of water, with 420 g of sodium hydroxide and 350 g of leached sludge from a continuous copper and lead leaching process added to obtain a NaOH concentration of 6.0% w / w and a solids content of 5.0% w / w. The slurry was placed in a 10 L glass reactor and heated at 90 °C for 1 to 6 hours with stirring at 900 rpm. Once the leaching time was complete, the slurry was cooled and filtered using a Büchner system.

[0147] Table 7 Results of Examples 29 and 30

[0148] Variables / Examples unit 29 30 Duration of stay h 1 6 Leaching yield Ge ％ 78.1 82.0 Si ％ 63.2 63.0

[0149] Ion exchange

[0150] Examples 31 to 38

[0151] A 12,500 mL alkaline leaching solution derived from sludge, with a Ge concentration of 28–34 mg / L and a Si concentration of 7.2–8.6 g / L, was passed through an ion exchange column with a 400 mL resin containing nitrogen-containing groups. The feed flow rate was 67 mL / min to 133 mL / min at a rate of 3.4–6.7 cm / min. The column was washed with 2000 mL of water, and no germanium elution was observed in any experiment. Germanium elution was performed with HCl at a rate of 99 g / L in bed volumes between 1 and 2. The elution flow rate was between 20 and 67 mL / min at a rate between 1 and 3.4 cm / min. Finally, the column was washed with 2000 mL of water, and no Ge resistance was observed in any experiment. The results are shown in Table 8.

[0152] Table 8 Results of Experiments 31 to 38

[0153]

[0154] Silicon removal

[0155] Examples 39 to 41

[0156] Take 2000 μL of feed solution from an ion exchange column, with a Si concentration of 18.5 g / L and pH 3.7, and add aluminum sulfate tetradecylhydrate with an Al / Si molar ratio of 0.43 to 1.0. The mixture is stirred at 400 rpm and 80 °C for 60 minutes. Once the reaction is complete, filter the slurry to obtain a Si removal efficiency of 74%, reduce the Si concentration to 2.8 g / L and the Al concentration to 130 mg / L, and the pH to 13.4.

[0157] Table 9 Results of Examples 39 to 41

[0158] Variables / Examples unit 39 40 41 Al / Si ratio h 0.43 0.70 1.00 Si precipitation yield ％ 74 98 99 Si concentration after precipitation g / L 2.8 0.41 0.12 Al concentration after precipitation g / L 0.13 1.00 3.5 pH after precipitation 13.4 13.4 12.5

[0159] Examples 42 to 45

[0160] In a 5L reactor, 4200g of IX (ion exchange) effluent solution with 17.4g / L Si was taken and added to it at a molar ratio of 0.9 to 1.1 mol Ca / mol Si, and stirred continuously at 300 rpm for 60 minutes. Once the reaction was complete, the slurry was filtered through 42-gauge filter paper.

[0161] Table 10 Results of Examples 42 to 45

[0162] Variables / Examples unit 42 43 44 45 Ca / Si ratio h 0.9 1.1 1.1 1.1 temperature ℃ 70 20 50 70 Si precipitation yield ％ 82 70 83 95

[0163] Germanium distillation

[0164] Example 46

[0165] A solution containing 454 mg / L Ge, 1039 mg / L Pb, 294 mg / L Al, 7 mg / L As, and 150 g / L HCl was used as the elution output from the ion exchange column. The solution was arranged in a 20 L distillation bulb and heated to 108 °C. Once the bulb reached 108 °C, the distillate output was observed and maintained in fractions between 300 and 1200 mL. The evaporated solution was condensed in a coil through which cold water was circulated at 5 °C and collected in a jacketed collection cup through which water was circulated at 5 °C. A total of 100% Ge was collected from the distillate, yielding a fraction with a maximum concentration of 1960 mg / L Ge. 0.02% of Pb present in the solution entered the distillate; however, 0.19% and 50% of Al and As, respectively, were entrained by the distillate.

[0166] Example 47

[0167] A 14000 mL solution containing 1060 mg / L Ge and 100 g / L HCl was prepared as the germanium distillation solution. The solution was placed in a 20 L distillation bulb and heated to 108 °C. Once the bulb reached 108 °C, the distillate solution output was observed and maintained in a 500 mL fraction. The evaporated solution was condensed in a coil through which cold water circulated at 5 °C and collected in a jacketed collection cup through which water circulated at 5 °C. A total of 100% Ge was collected from the distillate, yielding a fraction with a maximum concentration of 4330 mg / L Ge.

[0168] Example 48

[0169] Take 16000 mL of a solution containing 2630 mg / L Ge and 117 g / L HCl as the germanium distillation solution. Arrange the solution in a 20 L distillation bulb and heat to 108 °C. Once the bulb temperature reaches 108 °C, observe the distillate output and maintain it in a 1000 mL fraction. The evaporated solution is condensed in a coil through which cold water is circulated at 5 °C and collected in a jacketed collection cup through which water is circulated at 5 °C. Collect a total of 100% Ge in the distillate, yielding a fraction with a maximum concentration of 15000 mg / L Ge.

[0170] Example 49

[0171] Take 16000 mL of a solution containing 10400 mg / L Ge and 150 g / L HCl as the germanium distillation solution. Arrange the solution in a 20 L distillation bulb and heat to 108 °C. Once the bulb temperature reaches 108 °C, observe the distillate solution output and maintain it in a 1000 mL fraction. The evaporated solution is condensed in a coil through which cold water is circulated at 5 °C and collected in a jacketed collection cup through which water is circulated at 5 °C. Collect a total of 100% Ge in the distillate, yielding a fraction with a maximum concentration of 25000 mg / L Ge.

[0172] Germanium hydrolysis

[0173] Examples 50 to 57

[0174] Take 250 mL of a germanium solution with a concentration ranging from 4.1 to 24.8 g / L Ge and an HCl concentration between 54 and 134 g / L, and place it in a 500 mL jacketed reactor. Circulate 5000 mL of water, mixed with 1% vol% ethylene glycol and cooled at 1 °C, through the reactor. Mechanically stir the solution at 250 rpm for 5 h at a temperature between 2 and 3 °C. At the end of the process, filter the solution through 0.45 μm filter paper.

[0175] Table 11 Results of Examples 50 to 57

[0176] Variables / Examples unit 50 51 52 53 54 55 56 57 start Ge g / L 4.1 4.1 8.3 8.3 16.5 16.5 24.8 24.8 HCl g / L 54 134 54 134 54 134 54 134 Sediment g s / p s / p s / p 1.1 2.7 4.6 4.5 7.8 Ge ％ - - - 69 66 66 63 68

Claims

1. A method for producing germanium concentrate from metallurgical residues containing copper, iron, lead, and germanium, and optionally arsenic, antimony, and bismuth, characterized in that it comprises: i. Leach copper from metallurgical residues with a first acid solution to obtain a first leaching solution rich in copper and iron, and optionally arsenic, antimony, and bismuth, and a first leached sludge with reduced copper and iron content, and optionally reduced arsenic content, and rich in lead, silicon, and germanium. ii. Leaching the first leached sludge, wherein the first leached sludge is treated with a first solution of carboxylate to obtain a second leached sludge lacking lead and a second leaching solution rich in lead. iii. Alkaline leaching of the second leached sludge, wherein alkali is added to form an alkaline leaching mixture to obtain a third leached sludge with reduced silicon and germanium content and a third leaching solution rich in germanium, silicon, and optionally arsenic. iv. Feed is added to an ion exchange column, wherein germanium is captured by the resin to obtain a fourth basic solution that is germanium-deficient and silicon-rich. v. Rinse the ion exchange column with water to obtain the fifth solution for column feeding and rinsing. vi. Elute the ion exchange column with HCl solution to obtain a germanium-rich sixth elution solution. vii. Distillation, wherein the sixth elution solution, rich in germanium, is distilled to obtain a seventh solution rich in germanium and an eighth solution lacking germanium, and viii. Hydrolysis, in which a seventh solution of germanium is contacted with an aqueous solution to produce a first GeO2 concentrate.

2. The method according to claim 1, characterized in that the metallurgical residue to be treated is a powder obtained from a metal smelting process.

3. The method according to claim 2, characterized in that the powder obtained by the copper smelting process is a smelted powder.

4. The method according to any one of claims 1 to 3, characterized in that the metallurgical residue has undergone a copper leaching process.

5. The method according to claim 4, characterized in that the metallurgical residue has been leached with H2SO4.

6. The method according to any one of claims 1 to 3, characterized in that the metallurgical residue to be treated comprises the minerals galena, indigo chalcopyrite, copper-iron spinel in the form of CuOFe2O3, zinc spinel in the form of ZnOFe2O3, magnetite, iron oxide (III), pyrite, spathite, muscovite, kaolinite, and lead sulfate (II), galena, or lead oxide (II).

7. The method according to claim 6, characterized in that the copper contained in the metallurgical residue exists as copper sulfate, chalcocite, indigo chalcopyrite, and copper-iron spinel in the form of CuOFe2O3.

8. The method according to any one of claims 1 to 7, characterized in that the silicon contained in the metallurgical residue exists as muscovite and kaolinite.

9. The method according to any one of claims 1 to 8, characterized in that the lead contained in the metallurgical residue exists as lead(II) sulfate, galena, or lead(II) oxide.

10. The method according to claim 9, characterized in that at least 95% of the lead is lead(II) sulfate.

11. The method according to any one of claims 1 to 10, characterized in that the first acid solution from step i may contain H2SO4 and / or refining wastewater.

12. The method according to claim 11, characterized in that step (i) is carried out at an H2SO4 concentration between 150 and 300 g / L.

13. The method according to any one of claims 1 to 12, characterized in that step (i) is carried out at a temperature between 50 and 130°C.

14. The method according to any one of claims 1 to 13, characterized in that step (i) lasts for a period of 3 to 12 hours.

15. The method according to any one of claims 1 to 14, characterized in that step (i) is carried out at a solid concentration between 5% and 20% w / w.

16. The method according to any one of claims 1 to 15, characterized in that in the leaching step (ii), the carboxylate is sodium citrate.

17. The method according to any one of claims 1 to 16, characterized in that in step (ii) the sodium citrate solution has a molar concentration of sodium citrate between 0.5 M and 1 M.

18. The method according to any one of claims 1 to 17, characterized in that in step (ii), the first leachable sludge and sodium citrate solution are added to the sodium citrate solution at a mass ratio of 1:

9.

19. The method according to any one of claims 1 to 18, characterized in that step (ii) is carried out at a temperature between 20 and 60°C.

20. The method according to any one of claims 1 to 19, characterized in that step (ii) is performed for a stay time between 1 and 23 hours.

21. The method according to any one of claims 1 to 20, characterized in that step (ii) is carried out at a pH between 5.3 and 8.

8.

22. The method according to any one of claims 1 to 21, characterized in that citric acid is added in step (ii) to adjust the pH.

23. The method according to claim 22, characterized in that the pH adjustment in step (ii) is performed using a citric acid solution of 600 to 900 g / L.

24. The method according to any one of claims 1 to 23, characterized in that step (iii) is a germanium leaching step.

25. The method according to any one of claims 1 to 24, characterized in that the alkali used in the leaching in step (iii) is sodium hydroxide.

26. The method according to any one of claims 1 to 25, characterized in that the alkali added in step (iii) is added at a ratio between 5% and 10% w / w relative to the total mass of the alkaline leaching mixture.

27. The method according to any one of claims 1 to 26, characterized in that the leaching reaction of step (iii) is carried out at a temperature between 90 and 140°C.

28. The method according to any one of claims 1 to 27, characterized in that the leaching reaction in step (iii) lasts for a residence time between 1 and 6 hours.

29. The method according to any one of claims 1 to 28, characterized in that step (iv) is carried out using an ion exchange resin having nitrogen atom groups.

30. The method according to any one of claims 1 to 29, characterized in that a third leaching solution rich in germanium and silicon is fed to the resin at a bed volume ratio between 2 and 3, thereby carrying out step (iv).

31. The method according to claim 1, characterized in that step (v) is performed by flushing the ion exchange column with water at a bed volume ratio between 5 and 15.

32. The method according to any one of claims 1 to 31, characterized in step (vi) of elution with an HCl solution of concentration between 2N and 8N.

33. The method according to any one of claims 1 to 32, characterized in that the HCl solution is fed to the resin at a bed volume ratio between 1 and 5, thereby carrying out step (vi).

34. The method according to any one of claims 1 to 33, characterized in that the fourth alkaline solution lacking germanium and rich in silicon undergoes a silicon removal process.

35. The method according to claim 34, characterized in that quicklime is added during the silicon removal process.

36. The method according to any one of claims 1 to 35, characterized in that quicklime is added at a molar ratio of 1:1 relative to the silicon contained in a fourth alkaline solution lacking germanium and rich in silicon, so as to produce a regenerated alkaline solution and a solid composed of calcium silicate.

37. The method according to any one of claims 34 to 36, characterized in that the silicon removal step is performed at a temperature between 30 and 90°C.

38. The method according to claim 36, characterized in that the regenerated alkaline solution is recycled to the alkaline leaching step (iii).

39. The method according to claim 38, characterized in step (vii) by using a temperature between 86.5°C and 107°C for distillation.

40. The method according to any one of claims 1 to 39, characterized in that the seventh solution of germanium is distilled 1 to 5 times to produce a concentrated germanium solution and a distilled HCl solution.

41. The method according to claim 40, characterized in that the seventh solution of germanium is distilled three times to produce a concentrated germanium solution and a distilled HCl solution.

42. The method according to any one of claims 40 to 41, characterized in that the distilled HCl solution is recycled in the preceding distillation step in order to increase the HCl concentration at the inlet of the distiller.

43. The method according to any one of claims 1 to 41, characterized in that the concentrated germanium solution added in step (viii) is contacted with deionized water at a volume ratio between 1:1 and 1:6 to precipitate the germanium dioxide concentrate.

44. The method according to any one of claims 40 to 43, characterized in that the contact between the concentrated germanium solution in step (viii) and deionized water is carried out at a temperature between 2 and 15°C.

45. The method according to any one of claims 40 to 44, characterized in that the concentrated germanium solution sent to the hydrolysis step has a germanium concentration between 8.1 and 24.8 g / L.

46. ​​The method according to any one of claims 40 to 45, characterized in that the concentrated germanium solution sent to step (viii) has an HCl concentration between 55 and 135 g / L.

47. The method according to any one of claims 1 to 46, characterized in that a first leaching solution rich in copper and iron, and optionally arsenic, antimony and bismuth, is fed into a copper leaching process of smelting powder.

48. The method according to any one of claims 1 to 47, characterized in that a first leaching solution rich in copper and iron, and optionally arsenic, antimony and bismuth, is fed to an arsenic removal process.

49. The method according to claim 48, characterized in that the arsenic removal process is selected from the production process of ferric arsenate.

50. The method according to claim 49, characterized in that the arsenic removal process is a styrofoam production process.

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