WO2022018490A1 - Procedure for producing germanium concentrate from residues - Google Patents

Abstract

Procedure for producing germanium concentrate from metallurgical residues, comprising: (i) copper leaching with a first acid solution of the metallurgical residue, in order to obtain a first leaching solution rich in copper and iron, and a first leached sludge having a content reduced in copper and iron, (ii) leaching the first leached sludge wherein said first leached sludge is processed with a first solution of sodium citrate, in order to obtain a second leached sludge deficient of lead and a second leaching solution rich in lead, (iii) alkaline leaching of the second leached sludge, wherein a base is added in order to form an alkaline leaching solution, in order to obtain a third leached sludge having a content reduced in silicon and germanium and a third leaching solution rich in germanium and silicon, (iv) charging in an ion exchange column, wherein the germanium is captured by the resin, in order to obtain a fourth alkaline solution deficient of germanium and rich in silicon.

Description

Procedure for the production of germanium concentrate from metallurgical waste

field of invention

The invention relates to a process for the production of germanium concentrate from metallurgical waste, in particular from waste containing copper, iron, lead and germanium, and which may optionally contain elements such as arsenic, antimony and bismuth. .

In a more specific aspect, metallurgical waste is dust that comes from a metal smelting process.

In an even more specific aspect, metallurgical residues are powders that come from a copper smelting process.

In an even more specific aspect, metallurgical residues, or in particular foundry powders, contemplate materials that have already been subjected to leaching processes, such as sulfuric leaching.

In a broad aspect germanium concentrate can be understood as a liquid fraction rich in germanium, consisting mainly of germanium tetrachloride, or as a solid germanium concentrate, which in a more specific aspect can refer to germanium dioxide.

Within this description, all metallurgical waste that has been subjected to previous leaching processes will be treated as sludge.

State of the art Copper leaching

The copper in the fluff is found mainly forming ferrite and/or spinel-type species of the CuFe ₂ 0 _{4 form,} as well as zinc, ZnFe ₂ 0 ₄ and an important part of the iron FeFe ₂ 0 ₄ . The leaching of these species is a function of temperature, acid concentration, and residence time, as described in the study by BS Boyanov, et al. in World Academy of Science, Engineering and Technology, Vol 9, 2015, 1592-1598, who carried out a study on the leaching of synthetic zinc, copper and cadmium ferrites, evaluating the aforementioned variables. The results of this study show that ferrites dissolve better in HCI and H ₂ S0 ₄ , at high temperatures and high acid concentrations. At high concentrations of acid, it is observed that copper leaching has an asymptotic behavior with respect to the leaching temperature, which, after a reaction time of 60 min in sulfuric medium, reaches copper leaching yields greater than 90% for the temperature range. temperatures between 85 and 90°C.

lead leaching

Global consumption of lead in 2011 exceeded 10 million tons, of which around 80% of said lead was destined for the manufacture of lead-acid batteries. These batteries contain quantities of lead in the form of Pb, Pb0 ₂ and PbSC> ₄ . The most traditional way to recover lead is the pyrometallurgical route, which is characterized by the addition of a reducing agent, such as coal dust, scrap iron and sodium oxalate. The operation is carried out in furnaces at temperatures above 1000°C, which results in a process with high energy demand. He et al., Minerals 7, no. 6 (2017): 93.

On the other hand, the hydrometallurgical lead recovery route allows working at reduced temperatures, reducing energy consumption, and in turn does not produce sulfur dioxide, which is characterized as a gas that is harmful to the environment. The hydrometallurgical route makes use of desulfurization agents such as sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, sodium acetate, among others. The goal of these processes is to exchange the sulfate ion for other anions 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.).

Citrate desulfurization

In the particular case of the use of citrates, the mixture of citric acid with sodium citrate is beneficial for the leaching of lead sulfate and subsequent crystallization of lead citrate.

Leaching of lead in citrate solutions

The solubility product constant of anglesite at 20°C is 6.31 Ί O ⁷ , indicating that the solubility of PbSC> ₄ is very low. However, in the presence of concentrated citrate solutions, lead forms a series of soluble complexes. In solutions containing 0.12 M Pb ²⁺ , a wide variety of citrate complex species are present in solution over the pH range 4.6 to 11.5. At pH below 4.6 the presence of lead sulfate is predominant, while at pH above 11.5 the presence of lead hydroxide is dominant. He et al., Minerals 7, no. 6 (2017): 93, studied the leaching of lead from a paste with a weight ratio of lead sulfate and water of 1:10, by adding 650 g/L of sodium citrate at 35°C. These conditions made it possible to convert more than 99% of the lead sulfate to lead citrate after 60 min of reaction. Increasing the temperature to 95°C, at a sodium citrate concentration of 300 g/L, allowed an efficiency close to 99% to be obtained after 60 min of reaction. However, when citric acid was introduced into the mixture, a decrease in the production of lead citrate was observed. The optimum pH for lead citrate production was within the range of 6 to 7. At pH 5.5 using citric acid and ammonium agents high lead leaching efficiencies are also obtained from lead-acid batteries. Within the pH range 5.2 to 5.5, the presence of lead citrate trihydrate ([Pb3(C ₆ H507)2] [3H ₂ 0]) has been reported as the main species. At higher pH values within the range 8 to 10, the recovery of lead as citrate salt is lower, due to the formation of lead hydroxide. When lead residues are rich in oxides of the PbO and Pb0 _{2 type} , leaching is carried out with a molar ratio of citric acid to lead (II) and (IV) oxide of 1:1 and 4:1 to 20 °C between 15 and 60 min of reaction, reaching leaching efficiencies greater than 99% by weight, obtaining Rό(0 ₆ H ₆ q ₇ )Ή ₂ q as the main species (Sonmez and Kumar, Hydrometallurgy 95, no. 1- 2 (2009), 82-86.)

Pulp density is another important parameter for lead leaching with citrate solutions. Within the range of 10 to 50 g/L of anglesite pulp, leached with a 1 M sodium citrate solution, pH 7 at 600 rpm and 25°C, the highest lead extraction levels of 90 to 94% were reached. with a pulp concentration of 10 g/L. The higher the concentration of the pulp, the lower the amount of lead extracted.

From the foregoing, it follows that hydrometallurgical desulfurization processes are affected by the diffusion of citrate ions in the lead paste inside the reactor, due to the high density of the lead paste. In this context, the design of reactors that maximize the mass transfer in the system is essential.

Technology based on the recovery of lead from lead waste using citric acid has been developed by Cambridge Enterprise Limited (W02008056125A1). The technology basically comprises treating lead waste comprising lead(II) oxide, lead(IV) oxide and lead sulphate with a citric acid solution, and which can alternatively be treated in combination with sodium citrate at a pH that varies within the range of 1.4 to 6. It is eventually possible to add hydrogen peroxide in a basic environment as a reducing agent to accelerate the leaching reaction of lead (IV) oxide to produce lead citrate (Sonmez and Kumar, Hydrometallurgy 95, no 1-2 (2009), 82-86.). The present invention differs from the application (W02008056125A1) in that the pH required for leaching varies between 5.33 and 8.8, where a pH equal to 7 is preferably used. Additionally, the present invention proposes recirculating the citrate solution obtained after a precipitation stage with sodium carbonate, in order to be able to leach metallurgical residue from the sulfuric leaching stage again.

germanium leaching

Germanium is a metal that is widely used in the field of fiber optics, infrared optics, photovoltaic cells, aeronautics, aerospace and military industries, among others. In general, germanium is not an abundant metal in the earth's crust, since it constitutes between 1 - 7 ppm of the earth's crust, with a total amount estimated at 8600 tons. Germanium is usually found associated with deposits of copper, lead, zinc and coal, while deposits with high germanium content are limited. Most of the germanium is recovered from lignite vapors subjected to pyrometallurgical processes, and from the smelting of lead-zinc ores. However, pyrometallurgical processes have lost importance, since they generate environmental problems associated with the volatility of germanium (II) oxide and germanium sulfide.

There are various methods for the recovery of germanium from electrolytic zinc solutions, among which precipitation with tannins, germanium tetrachloride distillation, flotation, adsorption on activated carbon, precipitation, solvent extraction and adsorption on chelating resins (US 455332). Germanium is usually present in the form of germanic acid Ge(OH) ₄ , being the majority species within the range of pH 1 to 8, while between pH 9 to 13 the majority species is GeO(OH) ₃ ^{1 ·} and at pH greater than 13 the dominant species is Ge0 ₂ (0H) ₂ ² -. The first dissociation constant for germanic acid is 4.9-10 mol.L-1 (log Kd¡ss = -9.31) (Wood y lain, Ore Geology Reviews 28, no. 1 (2006): 57- 102.).

Ge(OH) _4< ®GeO(OH) ₃ + H ⁺ log K _d ¡ _S s = -9.31

Ge0(0H) ₃ -<®Ge0 ₂ (0H) ₂ ² + H ⁺ log Kd¡ss = -21 .9

acid leaching

Germanium can be leached using H ₂ S0 ₄ , at temperatures between 40 and 60°C, at a concentration of H ₂ S0 ₄ of 100 g/L for a time of 30 min and with a solid:liquid ratio of 1:4 , recovering 78% of the germanium. At higher temperatures, of the order of 85°C and at a concentration of H ₂ S0 ₄ of 150 g/L and a residence time of 1 h, a collective leaching of different metals, where germanium extraction was 92.7% (Rutledge et al., Metals 5, no. 3 (2015): 1520-1542.).

Patent application CN108486390A describes a process for separating germanium and gallium from a material containing germanium and gallium. In a first stage, the germanium and gallium material is added to a solution of 50 to 150 g/L of H ₂ S0 ₄ at a ratio of 5 to 10% p/p, and subsequently it is adjusted to a pH between 1 and 3. The leaching solution is neutralized to pH 1 to 3, then zinc powder is added to the neutralized liquid at the temperature of 40-80°C, to obtain germanium concentrates and liquid solution. Zinc powder is added again to this liquid solution at a temperature of 40-80 °C, to obtain gallium residues and a liquid solution.

alkaline leaching

Patent application CN108300876A describes a method for leaching gallium and germanium from slag from zinc production processes. In a first stage, the slag is ground to a size of 50-100 microns, then 0.1-1 mol/L of H ₂ S0 ₄ solution is added in a liquid-solid ratio of 4-10: 1 ml/g , at a leaching temperature of 25-80 °C, stirring at 100-600 rpm for 0.25-4 h. Liquid-solid separation is then carried out to obtain leaching residues with H ₂ S0 ₄ to which a solution of 0.2-2 mol/L of hydrogen peroxide is added in a liquid-solid ratio of 4-10: 1 mL/g, adding 0.1-1 mol/L NaOH to adjust the pH of the leaching solution to 5.0 - 8.0. This alkaline leaching is carried out at 25-80 °C, stirring at 100-600 rpm for 0.25-4 h, in order to obtain a germanium-rich leaching solution. In the case of the present invention, in order to be able to leach the germanium values, the sulfuric leaching and citric leaching stage is necessary to be able to remove the high lead content from the metallurgical residue, and thus increase the germanium content in the residue due to the loss of mass of the metallurgical residue in the citric leaching stage. In addition, the presence of lead in the metallurgical waste when generating a process such as the one in the application would result in a higher consumption of soda, due to the conversion of lead sulfate into lead hydroxide, which would impact the germanium leaching yield. .

ion exchange

After the leaching stage, the germanium obtained must be concentrated through different alternatives such as solvent extraction, precipitation with chelating agents or ion exchange. For the present invention, the ion exchange stage has been considered since it is the one that provides the best recovery levels, and in turn It allows to increase the concentration ratio of germanium versus other metals present in the leaching solution such as lead, aluminum, silicon and arsenic.

Patent US4525332 describes the adsorption of germanium-containing solutions on ion exchange resins consisting of a polymer having as functional groups selected from secondary, tertiary and quaternary ammonium groups, having a relative selectivity of germanium to antimony of 50:1, followed by an elution of the germanium captured by the resin in an aqueous medium. The present invention makes use of a resin that contains N-methylglucamines as a functional group. Patent US4525332 does not specify that the loading solution may contain silicon. In the application examples of patent US4525332, the charge is carried out with acid or slightly acid solutions, and not with alkaline solutions, as shown in the present invention. Silicon, being an element belonging to the same group as germanium, can interfere when recirculating the charging solutions used in the process. In the same way, the elution of the resin used in the examples of patent US4525332 is carried out with H ₂ S0 ₄ or sodium hydroxide, depending on the resin used, in contrast to the elution with HCl that is carried out in the present invention, and that it is necessary for the subsequent germanium tetrachloride distillation step. In addition, the present invention differs from patent US4525332 in that it does not teach how to remove silicon from alkaline solutions to recirculate the solution as a leaching medium for metallurgical residues to obtain germanium.

Patent application GB933563A teaches about the processing of neutral or slightly acidic aqueous solutions of germanium in ion exchange resins containing phenyl hydroxyl groups in which the germanium is captured, to then be eluted with hydrochloric solutions at 7 N, to later be distilled and the germanium tetrachloride hydrolyzed to produce germanium dioxide. Patent application GB933563A does not teach how to charge strongly alkaline solutions with the presence of silicon. As there is silicon in solution, the neutralization of the solution to carry out the loading of germanium in the resin is not obvious, since when the pH drops, the silicon will precipitate and drag the germanium. In this scenario, it is necessary to load without prior neutralization and remove the silicon once the solution has passed through the column and is depleted of germanium. Additionally, contrary to patent application GB933563, in this invention the elution can be carried out with solutions with a Ge concentration of less than 6 N without showing considerable loss of germanium. germanium distillation

Powell et al. J.Appl. Chem. 1951, 541-551, teaches leaching of foundry powders with HCl, to generate in-situ germanium tetrachloride which could be distilled at a temperature of 84°C. One of the problems with this method is the presence of arsenic trichloride, which, although it boils at 130°C, the vapor pressure of this compound at 84°C is high enough to co-distill with arsenic tetrachloride. germanium. In this sense, the present invention differs from the teachings of Powell in the sense that there are a series of previous leaching steps and separation by ion exchange that prevent elements such as arsenic from being present in high concentrations that interfere with the distillation of germanium.

Patent US3102786 teaches about a continuous process to purify germanium tetrachloride using an HCl solution with a minimum concentration of 6 N, and maintaining the column at a temperature between 83 and 110°C. The present invention differs from US3102786 in that the distillation can be carried out at acid concentrations lower than 6 N, allowing germanium to be distilled efficiently with distillation rates above 95%.

Patent US2811418 teaches about a germanium tetrachloride purification process, using a 12N concentration HCL solution and saturated with chlorine gas, allowing the mixture to separate into two phases, where the heaviest phase contains germanium tetrachloride purified. The present invention differs from US2811418 in that the distillation can be carried out at acid concentrations lower than 12 N, in particular lower than 5 N, allowing germanium to be distilled efficiently with distillation rates above 95%, minimizing entrainment. of arsenic.

germanium hydrolysis

Patent US3455645 discloses a process for the production of amorphous germanium dioxide, which is characterized by precipitating germanium present in an aqueous solution in which the pH is not less than 5 and not greater than 9. In particular, the experiments disclosed by US3455645 teaches about the addition of germanium tetrachloride to a solution containing 10 parts of NaOH per 90 parts of water to a pH of less than 8, or preferably less than 6. Patent US3455645 differs from the present invention in that solutions of distillation obtained are directly sent to a cooling reactor where the precipitation of germanium dioxide is verified, without the need to control the pH to the values indicated by US3455645. James E. Hoffmann in Extracting and Refining Germanium, Journal of Metals, June 1987, 42-45, points out that at HCl concentrations below 5.5, germanium is found mainly as germanic acid, and that 3 grams of water is sufficient to hydrolyze 1 g of germanium with a yield of 95%. Additionally, it states that it is preferable that the precipitation is carried out at a temperature close to 0°C, and that it is possible to use germanium dioxide to act as a nucleation center for the germanium precipitation. In the present invention it is shown that it is possible to obtain high yields of germanium precipitation without necessarily using germanium dioxide as a seed to verify germanium precipitation. The present invention demonstrates that using sufficiently concentrated solutions of germanium allow germanium dioxide to precipitate, and that low concentrations of HCl impair the germanium dioxide precipitation process compared to HCl concentrations close to 3.7 N (135 g/L of HCl).

Description of the figures

Figure I shows the process diagram of the method disclosed by the present invention.

Figure II shows the distillation curve of germanium from ion exchange solutions.

Figure III shows the second distillation curve of germanium from primary distillation solutions. Second distillation is understood as a distillate solution that has been collected after a first distillation cycle.

Figure IV shows the curve of the third distillation of germanium from secondary distillation solutions. Third distillation is understood as a distillate solution that has been collected after a second distillation cycle.

Description of the invention

In a broad aspect, the invention describes a process for the production of germanium from metallurgical waste.

In a preferred option, the invention describes a process for the production of germanium tetrachloride.

In an even more preferred option, the invention describes a process for producing a germanium concentrate in solid state. In an even more preferred option, the invention describes a process for producing technical grade germanium dioxide, with a concentration that varies between 60 and 70%.

In a broad aspect, the invention describes a process for the production of germanium concentrate from metallurgical waste, in particular from waste containing copper, iron, lead, silicon and germanium, and which may optionally contain elements such as arsenic, antimony and bismuth, characterized in that it comprises: a stage (i) of copper leaching with a first acid solution (2) of the metallurgical residue (1), to obtain a first leaching solution rich in copper and iron, and optionally arsenic, antimony and bismuth (3) and a first leached fluff having a reduced content of copper and iron, and optionally reduced in arsenic and enriched in lead, silicon and germanium (4), a stage (ii) of leaching the first leached fluff (4) wherein said first leached fluff (4) is processed with a first solution of a salt of a carboxylic acid (5), to obtain a second leached fluff depleted in lead (6) and a second a lead-enriched leaching solution (7), a step (iii) of alkaline leaching of the second leached fluff, where a base (8) is added to form an alkaline leaching solution, to obtain a third leached fluff having a reduced content of silicon and germanium (9), and a third leaching solution rich in germanium and silicon, and optionally arsenic (10), a step (iv) of loading the third leaching solution rich in germanium and silicon (10 ), and optionally arsenic in an ion exchange column, where the germanium is captured by the resin, to obtain a fourth alkaline solution depleted in germanium and rich in silicon (11), a stage (v) of washing the column of ion exchange, which is carried out with water (12), where a fifth washing solution is obtained from the column load (13), a stage (vi) of elution of the ion exchange column, which is carried out with a solution of HCI (14), to get a sext a germanium-rich elution solution (15), a distillation step (vii), wherein the sixth germanium-rich elution solution (15) is distilled to obtain a seventh germanium solution (16) and an eighth depleted solution of germanium (17), and a hydrolysis step (viii), wherein the seventh germanium solution (16) is contacted with a water solution (18) to produce a first germanium dioxide concentrate (19) and a ninth germanium-depleted solution (20) .

In a preferential option, the metallurgical residue to be processed is powder obtained by means of a metal smelting process or powder obtained by means of a copper smelting process.

In an even more preferential option, the metallurgical waste has been subjected to a copper leaching process.

In an even more preferential option, said metallurgical residue was subjected to leaching with H ₂ S0 ₄ .

In a preferential option, the metallurgical residue to be processed comprises the mineralogical species anglesite, coveline, copper spinels in the form Cu0Fe ₂ 0 ₃ , zinc spinels in the form ZnOFe ₂ C> ₃ , magnetite, iron oxide (III), pyrite, scorodite, mucovite, kaolinite and lead sulfate (II).

In an even more preferred option, the copper contained in the metallurgical residue is present as copper sulphate, chalcocite, covelin and copper spinels in the CuOFe2C>3 form.

In an even more preferred option, the copper contained in the metallurgical residue is present in at least 50% in the form of copper spinel in the CuOFe ₂ C> _{3 form} .

In a preferred option, the silicon contained in the metallurgical residue is present as muscovite and kaolinite.

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

In an even more preferred option, the lead is at least 95% as lead(II) sulfate.

In a preferred option, the first H2SO4 solution of step (i) may comprise H2SO4 and/or a refinery effluent.

In a preferred option, step (i) is carried out at a concentration of H2SO4 between 150 and 300 g/L, more preferably at a concentration of H2SO4 of 250 g/L. In a preferred option, step (i) is carried out at a temperature between 50 and 130°C, more preferably at a temperature of 85°C.

In a preferred option, step (i) is carried out for a time between 3 and 12 hours, more preferably at a residence time of 6 hours.

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

In a preferred option, in step (ii) of leaching, the carboxylic acid salt is sodium citrate.

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

In a preferential option, in stage (ii) the first leached fluff is fed to the sodium citrate solution in a mass ratio of 1:9.

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

In a preferred option, step (ii) is carried out for a residence time of between 1 and 23 h.

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

In a preferred option, in step (ii) the corresponding acid of the carboxylic acid salt is added for pH adjustment.

In an even more preferred option, in step (ii) citric acid is added for pH adjustment.

In an even more preferred option, the pH adjustment in stage (ii) is carried out with a citric acid solution of between 600 and 900 g/L.

In a preferred option, step (iii) is a germanium leaching step.

In a preferred option, the second base used in the leaching of step iv is selected from among Mg(OH)2, KOH or NaOH. In a preferential option, the base that is added in stage (iii) is added in a ratio of between 5 and 10% w/w with respect to the total mass of the alkaline leaching solution, more preferably in a ratio of 6, 0% w/w.

In a preferred option, the leaching reaction of stage (iii) is carried out at a temperature between 70 and 150°C, more preferably at a temperature of 130°C.

In a preferred option, the leaching reaction of stage (iii) is carried out during a residence time of between 1 to 12 hours, more preferably at a residence time of 3 hours.

In a preferred option, step (iv) is carried out with a resin with a group with a nitrogen atom (N-donor group).

In a preferred option, step (iv) is carried out by charging the third leach solution rich in germanium and silicon at a ratio of between 2 and 30 bed volumes.

In an even more preferred option, step (iv) is carried out by charging the third leach solution rich in germanium and silicon at a ratio of 10 bed volumes.

In a preferred option, step (v) is carried out by charging the fourth batch wash solution at a ratio of between 5 and 15 bed volumes.

In a preferred option, elution step (vi) is carried out with an HCl solution.

In an even more preferred option, elution step (vi) is carried out with an HCl solution with a concentration between 2 and 8 N, more preferably 6 N.

In a preferred option, step (vi) is carried out by charging the HCl solution at a ratio of between 1 and 5 bed volumes, more preferably at a ratio of 3 bed volumes.

In another preferred option, the fourth alkaline solution depleted in germanium is subjected to a silicon removal process.

In another preferential option, in said silicon removal process, slaked lime or aluminum sulfate is added.

In an even more preferred option, slaked lime is added to said silicon removal process. In an even more preferred option, the slaked lime is added in a 1:1 molar ratio with respect to the silicon contained in the fourth germanium-depleted alkaline solution, to generate a regenerated alkaline solution and a solid consisting of calcium silicate.

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

In an even more preferred option, the regenerated alkaline solution is recirculated to stage (iii) of alkaline leaching.

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

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

In an even more preferred option, the seventh germanium solution is redistilled 3 times to produce a concentrated germanium solution and a distilled HCl solution.

In a preferential option, the distilled HCl solution is recirculated to a previous distillation stage in order to increase the HCl concentration at the distiller inlet.

In another preferred option, the concentrated germanium solution is contacted with deionized water at a volumetric ratio of between 1:1 to 1:6 to precipitate the germanium as a germanium concentrate.

In an even more preferred option, said germanium concentrate is germanium dioxide.

In another even more preferred option, the concentrated germanium solution that is contacted with deionized water in step (viii) is carried out at a temperature between 2 and 15°C.

In another preferential option, the concentrated germanium solution sent to the hydrolysis stage has a germanium concentration between 8.1 and 24.8 g/L.

In another preferred option, the concentrated germanium solution sent to the hydrolysis stage has an HCl concentration between 55 and 135 g/L.

In a preferred option, the first copper-rich leach solution is sent to a process for leaching copper from smelter dusts.

In a preferential option, the first copper-rich leach solution is sent to an arsenic abatement process. In a preferential option, the arsenic abatement process is selected from those that contemplate the production of ferric arsenate.

In an even more preferred option, the arsenic abatement process is a scorodite production process. Application Examples

The following examples should be considered as modalities of the present invention, and in no case should they be considered as limiting it, since the different adaptations that can be made of it will be covered within the matter claimed by this invention. sulfuric leaching

Examples 1 to 7

Between 2,550 and 2,850 g of a sulfuric acid solution with a concentration between 150 and 250 g/L of H2SO4 were prepared, which was placed in a 5 L glass reactor, where sludge was added that was previously subjected to a process of copper leaching with a solid content between 5 and 10% p/p. The mineralogy of said sludge is presented in Table 1. The reactor was stirred at 300 rpm for 3 to 6 hours at 85°C. Once the reaction time was over, the pulp was filtered in a kitasate system. The results are presented in Table 2.

Table 1. Mineralogy of the sediments

Table 2. Leaching results of Cu examples 1 to 7

Examples 8 to 10 2,550 g of a solution of 250 g/L of H ₂ S0 ₄ were prepared, which were placed in a 4 L autoclave, where fluff was added that was previously subjected to a copper leaching process at a solids content of 15% p/p. The reactor was stirred at 300 rpm for 1 to 6 hours at 130°C. Once the reaction time was over, the pulp was filtered in a kitasate system. The results are presented in Table 3. Table 3. Leaching results of Cu examples 8 to 10

Example 11

A refinery effluent solution was prepared (table 4) to which the concentration of H ₂ S0 ₄ was adjusted to 250 g/L, which was placed in a 5 L glass reactor, where 450 g of sludge that was previously subjected to a copper leaching process to form a pulp with 15% p/p solids. The reactor was stirred at 300 rpm for 6 hours at 85°C. Once the reaction time was over, the pulp was filtered in a kitasate system. The results showed a Cu leaching yield of 72.0%, an Fe leaching yield of 62.0%, an As leaching yield of 71.5%, a Zn leaching yield of 57.0% and a mass loss of 38.5%.

Lixiviación cítrica Table 4. Refinery effluent composition

citrus leaching

Example 12

A solution was prepared with 40 L of distilled water to which 14 kg of sodium citrate were added and the pH was adjusted to 7.0 with a 800 g/L citric acid solution. Once the reagents had dissolved, 6 kg of leached fluff were added according to example 3. The top fluff had a Pb content of 15.4%. The leaching was carried out at 20°C and stirred at 1,000 rpm for a period of 9 h. A Pb leaching efficiency of 94% was obtained, obtaining a leached sludge that reduced its mass by 24% with a Pb content of 1.19%.

Examples 13 to 19

A solution was prepared with 2 L of distilled water with a concentration between 323 and 368 g/L of sodium citrate at a pH between 5.3 and 8.8. The pH was adjusted with a 800 g/L citric acid solution. Once the reagents were dissolved, the fluff processed according to example 3 was added in a ratio of between 1.2 and 2.3 g of sodium citrate/g of fluff. The top lint had a Pb content of between 15.0 and 15.1%. The leaching was carried out at between 30 and 60°C and agitated between 500 and 700 rpm for a period of between 2 and 4 h. The results are shown in Table 5. Table 5. Citrus Leaching Results Examples 13 to 19

alkaline leaching

Examples 20 to 28 A pulp was prepared with a sodium hydroxide solution with a concentration between 5.4 and 8.7% w/w and leached sludge subjected to consecutive copper and lead leaching processes with a solid content between 5, 0 and 7.0%w/w. The pulp was placed in a 4 L autoclave and heated at a temperature between 100 and 140°C for between 1 and 6 hours at 600 rpm. Once the leaching time was over, the pulp was cooled and filtered in a kitasato system. The results are shown in Table 6.

Table 6. Results examples 20 to 28

Examples 29 and 30

A pulp was prepared with 6,230 ml_ of water to which 420 g of sodium hydroxide and 350 g of leached fluff were added, subjected to consecutive copper and lead leaching processes, to obtain 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 at 90°C between 1 and 6 hours and stirred at 900 rpm. Once the leaching time was over, the pulp was cooled and filtered in a kitasato system.

Table 7. Results examples 29 and 30

ion exchange

Examples 31 to 38

12,500 mL of a solution from alkaline sludge leaching were taken, with concentrations of 28 to 34 mg/L of Ge and 7.2 to 8.6 g/L of Si, and passed through an ion exchange column with 400 mL of resin with a group with a nitrogen atom. The load flow was 67 mL/min to 133 mL/min at a velocity of 3.4 to 6.7 cm/min. The column was washed with 2000 mL of water, without observing germanium elution in any of the experiments. Germanium elution was performed with HCl at a concentration of 99 g/L passing between 1 and 2 bed volumes. The elution flow was between 20 and 67 mL/min at a speed between 1 and 3.4 cm/min. Finally, the column was washed with 2000 mL of water, without Ge drag being observed in any of the experiments. The results are presented in Table 8. Table 8. Results of experiments 31 to 38

Silicon removal

Examples 39 to 41 2,000 mL of the loading solution from the ion exchange column with a concentration of 18.5 g/L of Si and pH 13.7 were taken, to which aluminum sulfate tetradecahydrate was added in a ratio molar Al/Si of 0.43 and 1.0. The mixture was kept stirring at 400 rpm and at a temperature of 80°C for 60 minutes. Once the reaction was finished, the pulp was filtered and a Si removal efficiency of 74% was obtained, lowering the Si concentration to 2.8 g/L and Al to 130 mg/L with a pH of 13.4.

Table 9. Results examples 39 to 41

Examples 42 to 45

4,200 g of an IX discharge solution (acronym for ion exchange, in Spanish ionic exchange) with 17.4 g/L of Si were taken in a 5 L reactor at a temperature within the range of 20 to 70°C, to which calcium hydroxide was added at a molar ratio within the range of 0.9 to 1.1 mol Ca/mol Si, maintaining constant stirring at 300 rpm for 60 minutes. Once the reaction was complete, the pulp was filtered through No. 42 filter paper.

Destilación de germanio Ejemplo 46 Table 10. Results examples 42 to 45

Germanium distillation Example 46

18,000 ml_ of a solution with 454 mg/L of Ge, 1,039 mg/L of Pb, 294 mg/L of Al, 7 mg/L of As, and 150 g/L of HCl were taken as the elution output solution. of ion exchange columns. The solution was placed in a 20 L distillation flask and heated to 108°C. Once the temperature of the bulb reached 108°C, the distillate solution exited, which was stored in fractions of between 300 and 1,200 mL. The evaporated solution was condensed in a coil through which cold water at 5°C was circulated, and received in a collecting vessel with a jacket through which water at 5°C circulated. In the distillate, a total of 100% of the Ge was captured, obtaining fractions with up to 1,960 mg/L of Ge. 0.02% of the Pb present in the solution passed into the distillate, while 0.19 and 50% of the Al and As, respectively, were entrained by the distillate. Example 47

14,000 mL of a solution with 1,060 mg/L Ge and 100 g/L HCl were taken as germanium distillation solution. The solution was placed in a 20 L distillation flask and heated to 108°C. Once the temperature of the bulb reached 108°C, the output of the distillate solution was observed, which was stored in 500 mL fractions. The evaporated solution was condensed in a coil through which cold water at 5°C was circulated, and received in a collecting vessel with a jacket through which water at 5°C circulated. In the distillate, a total of 100% of the Ge was captured, obtaining fractions with up to 4,330 mg/L of Ge.

Example 48

16,000 mL of a solution with 2,630 mg/L Ge and 117 g/L HCl were taken as germanium distillation solution. The solution was placed in a 20 L distillation flask and heated to 108°C. Once the temperature of the bulb reached 108°C, the output of the distillate solution was observed, which was stored in fractions of 1,000 mL. The evaporated solution was condensed in a coil through which cold water at 5°C was circulated, and received in a collecting vessel with a jacket through which water at 5°C circulated. In the distillate, a total of 100% of the Ge was captured, obtaining fractions with up to 15,000 mg/L of Ge.

Example 49

16,000 mL of a solution with 10,400 mg/L Ge and 150 g/L HCl were taken as germanium distillation solution. The solution was placed in a 20 L distillation flask and heated to 108°C. Once the temperature of the bulb reached 108°C, the output of the distillate solution was observed, which was stored in fractions of 1,000 mL. The evaporated solution was condensed in a coil through which cold water at 5°C was circulated, and received in a collecting vessel with a jacket through which water at 5°C circulated. In the distillate, a total of 100% of the Ge was captured, obtaining fractions with up to 25,000 mg/L of Ge.

germanium hydrolysis

Examples 50 to 57

250 mL of a germanium solution with a concentration within the range of 4.1 to 24.8 g/L of Ge and a concentration of HCl between 54 and 134 g/L was taken, which were placed in a 500 mL reactor with jacket, through which 5,000 mL of water mixed at 1% by volume with ethylene glycol and cooled by a cooling equipment at 1 °C was circulated. The solution was allowed to stir mechanically at 250 rpm for 5 h at a temperature between 2 and 3°C. At the end of the process, the solution was filtered through a 0.45 pm filter paper. Table 11. Results examples 50 to 57

Claims

1. Procedure for the production of germanium concentrate from metallurgical waste, in particular, from waste containing copper, iron, lead and germanium, and which may optionally contain arsenic, antimony and bismuth, CHARACTERIZED because, it comprises: i. copper leaching with a first acid solution of the metallurgical residue, to obtain a first leaching solution rich in copper and iron, and optionally arsenic, antimony and bismuth and a first leached fluff that has a reduced content of copper and iron, and optionally reduced in arsenic and enriched in lead, silicon and germanium, ii. leaching of the first leached fluff wherein said first leached fluff is processed with a first solution of a salt of a carboxylic acid, to obtain a second leached fluff depleted in lead and a second leaching solution enriched in lead, iii. alkaline leaching of the second leached fluff, where a base is added to form an alkaline leaching solution, to obtain a third leached fluff having reduced silicon and germanium content, and a third leaching solution rich in germanium and silicon, and optionally arsenic, iv. loading into an ion exchange column, where the germanium is captured by a resin, to obtain a fourth alkaline solution depleted in germanium and rich in silicon, v. washing of the ion exchange column, which is carried out with water, where a fifth washing solution loading the column is obtained, vi. elution from the ion exchange column, which is carried out with a HCI solution, to obtain a sixth elution solution rich in germanium, vii. distillation, wherein the sixth germanium-rich eluting solution is distilled to obtain a seventh germanium solution and an eighth germanium-depleted solution, and viii. hydrolysis, wherein the seventh germanium solution is contacted with a water solution to produce a first germanium dioxide concentrate. 2. The procedure according to claim 1, CHARACTERIZED in that the metallurgical residue to be processed is powder obtained by means of a metal smelting process. 3. The process according to claim 2, CHARACTERIZED in that said powder obtained by means of a copper smelting process is smelting powder. 4. The process according to any of claims 1, 2 or 3, CHARACTERIZED in that the metallurgical residue has been subjected to a copper leaching process. 5. The process according to claim 4, CHARACTERIZED in that said metallurgical residue was subjected to leaching with H ₂ S0 ₄ . 6. The method according to any of claims 1 to 3, CHARACTERIZED because the metallurgical residue to be processed includes the mineralogical species anglesite, coveline, copper spinels in the form Cu0Fe ₂ 0 ₃ , zinc spinels in the form Zn0Fe ₂ 0 ₃ , magnetite, iron oxide (lll), pyrite, scorodite , mucovite, kaolinite and lead(ll) sulfate. 7. The process according to claim 6, CHARACTERIZED in that the copper contained in the metallurgical residue is present as copper sulfate, chalcocite, covelin and copper spinels in the form Cu0Fe ₂ 0 ₃ . 8. The method according to any of claims 1 to 7, CHARACTERIZED because the silicon contained in the metallurgical residue is present as muscovite and kaolinite. 9. The process according to any of claims 1 to 7 or 1 to 8, CHARACTERIZED in that the lead contained in the metallurgical residue is present as lead(ll) sulfate, galena or lead(ll) oxide. 10. The process according to claim 9, CHARACTERIZED in that the lead is in at least 95% as lead(ll) sulfate. 11. The process according to any of claims 1 to 10, CHARACTERIZED in that the first solution of H ₂ S0 ₄ of stage i can comprise H ₂ S0 ₄ and/or a refinery effluent. 12. The process according to any of claims 1 to 11, CHARACTERIZED because, step (i) is carried out at a concentration of H ₂ S0 ₄ between 150 and 300 g/L. 13. The method according to any of claims 1 to 12, CHARACTERIZED because, step (i) is carried out at a temperature between 50 and 130°C. 14. The method according to any of claims 1 to 13, CHARACTERIZED because, stage (i) is carried out for a time between 3 and 12 hours. 15. The method according to any of claims 1 to 14, CHARACTERIZED because, stage (i) is carried out at a concentration of solids between 5 and 20% p/p. 16. The method according to any of claims 1 to 15, CHARACTERIZED because, in stage (ii) of leaching, the carboxylic acid salt is sodium citrate. 17. The process according to any of claims 1 to 16, CHARACTERIZED because, in step (ii) the sodium citrate solution has a molar concentration of sodium citrate between 0.5 and 1 M. 18. The method according to any of claims 1 to 17, CHARACTERIZED because, in stage (ii) the first leached fluff is fed to the sodium citrate solution in a mass ratio of 1:9. 19. The method according to any of claims 1 to 18, CHARACTERIZED in that stage (ii) is carried out at a temperature between 20 and 60°C. 20. The process according to any of claims 1 to 19, CHARACTERIZED in that stage (ii) is carried out for a residence time of between 1 and 23 h. 21. The process according to any of claims 1 to 20, CHARACTERIZED in that step (ii) is carried out at a pH between 5.3 and 8.8. 22. The process according to any of claims 1 to 21, CHARACTERIZED because, in stage (ii) citric acid is added to adjust the pH. 23. The method according to any of claims 1 to 22, CHARACTERIZED because, the pH adjustment in stage (ii) is carried out with a citric acid solution of 600 and 900 g/L. 24. The method according to any of claims 1 to 23, CHARACTERIZED in that stage (iii) is a germanium leaching stage. 25. The method according to any of claims 1 to 24, CHARACTERIZED because the base used in the leaching of stage (iii) is sodium hydroxide. 26. The method according to any of claims 1 to 25, CHARACTERIZED because the base that is added in stage (iii) is added in a ratio of between 5 and 10% w/w with respect to the total mass of the alkaline leaching solution. 27. The method according to any of claims 1 to 26, CHARACTERIZED in that the leaching reaction of stage (iii) is carried out at a temperature between 90 and 140°C. 28. The method according to any of claims 1 to 27, CHARACTERIZED in that the leaching reaction of stage (iii) is carried out during a residence time of between 1 to 6 hours. 29. The method according to any of claims 1 to 28, CHARACTERIZED because, step (iv) is carried out with an ion exchange resin with a group with a nitrogen atom. 30. The method according to any of claims 1 to 29, CHARACTERIZED in that stage (iv) is carried out by loading the third leaching solution rich in germanium and silicon at a ratio of between 2 and 30 bed volumes. 31. The method according to any of claims 1 to 30, CHARACTERIZED in that step (v) is carried out by loading the fourth load washing solution at a ratio of between 5 and 15 bed volumes. 32. The process according to any of claims 1 to 31, CHARACTERIZED in that the elution step (vi) is carried out with an HCI solution. 33. The method according to any of claims 1 to 32, CHARACTERIZED because, the elution step (vi) is carried out with an HCl solution with a concentration between 2 and 8 N. 34. The method according to any of claims 1 to 33, CHARACTERIZED because, step (vi) is carried out by loading the HCl solution at a ratio of between 1 and 5 bed volumes. 35. The method according to any of claims 1 to 34, CHARACTERIZED because the fourth alkaline solution depleted in germanium is subjected to a silicon removal process. 36. The method according to any of claims 1 to 35, CHARACTERIZED because, in said silicon removal process, slaked lime is added. 37. The method according to any of claims 1 to 36, CHARACTERIZED because, the slaked lime is added in a 1:1 molar ratio with respect to the silicon contained in the fourth alkaline solution depleted in germanium, to generate a regenerated alkaline solution and a solid consisting of calcium silicate. 38. The method according to any of claims 1 to 37, CHARACTERIZED because, the silicon removal stage is carried out at a temperature between 30 and 90°C. 39. The method according to any of claims 1 to 38, CHARACTERIZED because the regenerated alkaline solution is recirculated to stage (iii) of alkaline leaching. 40. The method according to any of claims 1 to 39, CHARACTERIZED because, stage (vii) of distillation is carried out at a temperature of between 86.5 and 107°C and at a bulb temperature of between 86.5 and 108°C. 41. The method according to any of claims 1 to 40, CHARACTERIZED because, the seventh germanium solution is redistilled from between 1 to 5 times to produce a concentrated germanium solution and a distilled HCl solution. 42. The process according to claim 41, CHARACTERIZED in that the seventh germanium solution is redistilled 3 times to produce a concentrated germanium solution and a distilled HCI solution. 43. The method according to any of claims 1 to 42, CHARACTERIZED because the distilled HCl solution is recirculated to a previous distillation stage in order to increase the concentration of HCl at the distiller's inlet. 44. The method according to any of claims 1 to 43, CHARACTERIZED in that the concentrated germanium solution that is fed to step (viii) is contacted with deionized water at a volume ratio of between 1:1 to 1:6 to precipitate the germanium as a germanium concentrate. 45. The process according to claim 44, CHARACTERIZED in that said germanium concentrate obtained in step (viii) is germanium dioxide. 46. The method according to any of claims 1 to 45, CHARACTERIZED in that the concentrated germanium solution that is contacted with deionized water in step (viii) is carried out at a temperature between 2 and 15°C. 47. The method according to any of claims 1 to 46, CHARACTERIZED because the concentrated germanium solution sent to the hydrolysis stage has a germanium concentration between 8.1 and 24.8 g/L. 48. The method according to any of claims 1 to 47, CHARACTERIZED because the concentrated germanium solution sent to stage (viii) has a concentration of HCl between 55 and 135 g/L. 49. The method according to any of claims 1 to 48, CHARACTERIZED because, the first copper-rich leaching solution is sent to a process for leaching copper from foundry powders. 50. The method according to any of claims 1 to 49, CHARACTERIZED because the first copper-rich leaching solution is sent to an arsenic abatement process. 51. The process according to claim 50, CHARACTERIZED in that the arsenic abatement process is selected from those that contemplate the production of ferric arsenate. 52. The process according to claim 51, CHARACTERIZED in that the arsenic abatement process is a scorodite production process.