CN114269954A

CN114269954A - Method for producing germanium concentrates from metallurgical residues

Info

Publication number: CN114269954A
Application number: CN202080053497.7A
Authority: CN
Inventors: M·G·阿库纳戈伊科拉; R·M·佩佐娅科特
Original assignee: Ecometales Ltd
Current assignee: Ecometales Ltd
Priority date: 2020-07-22
Filing date: 2020-07-22
Publication date: 2022-04-01
Also published as: PE20220264A1; JP2023542440A; WO2022018490A1; CA3143346A1

Abstract

A method of producing a germanium concentrate from a metallurgical residue, comprising: (i) leaching copper from a metallurgical residue with a first acid solution in order to obtain a first leach solution enriched in copper and iron and optionally arsenic, antimony and bismuth and a first leached sludge enriched in lead, silicon and germanium having a reduced copper and iron content and optionally a reduced arsenic content, (ii) leaching the first leached sludge, wherein the first leached sludge is treated with a first solution of sodium citrate in order to obtain a second leached sludge deficient in lead and enriched in lead, silicon and germanium(ii) alkaline leaching the second leached sludge, wherein alkali is added so as to form an alkaline leaching solution, so as to obtain a third leached sludge with reduced silicon and germanium content and a third leaching solution enriched 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 depleted in germanium and enriched in silicon, (v) rinsing the ion exchange column, wherein a rinsed fifth solution is fed, (vi) eluting the ion exchange column with an HCl solution so as to obtain a sixth elution solution enriched in germanium, (vii) distilling, wherein the sixth elution solution enriched in germanium is distilled so as to obtain a seventh solution in germanium and an eighth solution depleted in germanium, and (viii) hydrolyzing, wherein the seventh solution in germanium is contacted with an aqueous solution to produce the first GeO₂And (3) concentrating.

Description

Method for producing germanium concentrates from metallurgical residues

Technical Field

The invention relates to a method for producing germanium concentrates from metallurgical residues, in particular from residues containing copper, iron, lead and germanium and optionally containing elements such as arsenic, antimony and bismuth.

In a more particular aspect, the metallurgical residue is powder from a metal melting process.

In an even more particular aspect, the metallurgical residue is powder from a copper smelting process.

In an even more particular aspect, metallurgical residues or in particular smelted powders take into account materials that have undergone a leaching process such as sulphuric acid leaching.

In a broad sense, a germanium concentrate is understood to be a germanium-rich liquid fraction, mainly comprising germanium tetrachloride, or a solid germanium concentrate, which in more specific aspects may relate to germanium dioxide.

In the present disclosure, any metallurgical residue that has undergone a prior leaching process should be considered a sludge.

Prior Art

Copper leaching

The copper in the sludge mainly contains substances such as ferrite and/or CuFe₂O₄ZnFe as zinc₂O₄And the corresponding part of Fe₂O₄Spinel in its form. Leaching of these materials is based on temperature, acid concentration and residence time,synthetic ferrite leach studies of zinc, copper and cadmium were performed as described in the World academic of Science, Engineering and Technology, volume 9, 2015, 1592-. The results of this study show that ferrites dissolve better in HCl and H at elevated temperatures and high acid concentrations₂SO₄In (1).

At high acid concentrations, copper leaching was observed to have an asymptotic behaviour with respect to leaching temperature, which reached copper leaching yields of more than 90% in the temperature range between 85 and 90 ℃ once a 60 minute reaction time in a sulphuric acid medium had elapsed.

Lead leaching

World lead consumption in 2011 was greater than one million tons, with about 80% of the lead intended for use in the manufacture of lead acid batteries. These batteries contain a certain amount of Pb, PbO₂And PbSO₄Lead in the form of lead. The most traditional way of recovering lead is the pyrometallurgical route, which is characterized by the addition of reducing agents such as carbon powder, iron filings and sodium oxalate. Operation in an oven at temperatures greater than 1000 ℃ results in a high energy demand process He et al, Minerals 7, No. 6 (2017): 93.

On the other hand, the hydrometallurgical route for the recovery of lead allows working at reduced temperatures, reducing energy consumption and, instead, not producing gases that are not producing sulphur dioxide, characterized by gases that are harmful to the environment. The hydrometallurgical route uses a desulfurizing agent such as sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, sodium hydroxide, sodium citrate, acetic acid, sodium acetate, and the like. The purpose of these processes is to exchange sulfate ions for other anions in order to form insoluble salts. Once recovered, lead salts such as lead citrate may be calcined to produce lead oxide (Zrate-Guti e rrez and Lapidus, Hydrometallurgy 144(2014): 124-128).

Desulfurization with citrate

In the particular case of citrate, the citric acid and sodium citrate mixture favours the leaching of lead sulphate and the subsequent crystallisation of lead citrate.

Leaching of lead in citrate solutions

Dissolving lead alum at 20 DEG CThe product constant of degree is 6.31.10^-7Showing that PbSO₄The solubility is rather low. However, in the presence of a concentrated solution of citrate, lead forms a series of soluble complexes. In the presence of 0.12M Pb²⁺In a pH range of 4.6 to 11.5, a wide variety of citrate complexing species are present in the solution. At a pH below 4.6, mainly lead sulphate is present, and at a pH above 11.5 mainly lead hydroxide is present.

He et al, Minerals 7, phase 6 (2017):93 study lead leaching of pastes with a weight ratio of lead sulfate to water of 1:10 by adding 650g/L sodium citrate at 35 ℃. These conditions allow greater than 99% of the lead sulfate to be converted to lead citrate once the 60min reaction time has elapsed. The temperature was increased up to 95 ℃ and once the 60min reaction time had elapsed, a sodium citrate concentration of 300g/L allowed to obtain an efficiency of close to 99%. However, when citric acid was introduced into the mixture, a reduction in lead citrate production was observed. The pH optimized for the production of lead citrate is in the range of 6 to 7. The use of citric acid and ammonium reagents at a pH of 5.5 also results in increased lead leaching efficiency from lead acid batteries. In the pH range of 5.2 to 5.5, lead citrate trihydrate ([ Pb [)₃(C₆H₅O₇)₂]·[3H₂O]) The presence of (b) is reported as the major species. At higher pH in the range of 8 to 10, lead recovery as citrate is lower due to the formation of lead hydroxide. When the lead residue is rich in oxides such as PbO and PbO₂Leaching is carried out at 20 ℃ with citric acid reacting with lead (II) oxide and lead (IV) oxide in molar ratios of 1:1 and 4:1 for between 15 and 60min, to a leaching efficiency higher than 99% by weight, obtaining Pb (C)₆H₆O₇)·H₂O as the main substance (Sonmez and Kumar, Hydrometallurgy 95, stages 1-2 (2009), 82-86).

The slurry density is another important parameter for leaching lead with a citrate solution. In the range of 10 to 50g/L lead alum slurry, leaching with 1M, pH 7 sodium citrate solution at 600rpm and 25 ℃ reached higher lead extraction levels of 90 to 94% at a slurry concentration of 10 g/L. The greater the slurry concentration, the less lead is extracted.

Thus, the hydrometallurgical desulfurization process is affected by diffusion of citrate ions in the lead paste in the reactor due to the increased density of the lead paste. In this context, it is critical to design a reactor that maximizes mass transfer in the system.

Cambridge Enterprise Limited has developed a technology based on the use of citric acid to recover lead from lead waste (WO2008056125a 1). This technique essentially consists of treating the lead residue containing lead (II) oxide, lead (IV) oxide and lead sulphate with a citric acid solution, and it can alternatively be treated at a pH varying in the range 1.4 to 6 in combination with sodium citrate. Finally, hydrogen peroxide can be added in an alkaline environment as a reducing agent to facilitate the lead (IV) oxide leaching reaction to produce lead citrate (Sonmez and Kumar, Hydrometallurgy 95, stages 1-2 (2009), 82-86).

The present invention differs from the application (WO2008056125a1) in that the pH required for leaching varies from 5.33 to 8.8, in which range a pH equal to 7 is preferably used. In addition, the invention proposes to recycle the citric acid solution obtained after the precipitation step with sodium carbonate, so as to re-leach the outgoing metallurgical residues from the sulfuric acid leaching step.

Germanium leaching

Germanium is a metal that is widely used in the fields of optical fibers, infrared fibers, photovoltaic cells, and the aerospace, military, and other industries. Generally, germanium is not abundant in the crust of the earth, since it constitutes 1-7ppm of the crust, with an estimated total of 8600 tons. Germanium is commonly associated with copper, lead, zinc, and carbon deposits, however deposits with high germanium content are limited. Most of the germanium is recovered from lignite steam and lead zinc ore smelting undergoing the pyrometallurgical process. However, pyrometallurgical processes lose importance each time they create environmental problems associated with germanium (II) oxide and germanium sulfide volatility.

There are various methods of recovering germanium from electrolytic solutions of zinc, among which precipitation with tannic acid, distillation of germanium tetrachloride, flotation, activated carbon adsorption, precipitation, solvent extraction, and chelate resin adsorption are superior (US 455332). Germanium is usually germanium acid Ge (OH)₄Is present as the predominant species in the pH range of pH 1 to 8, but predominantly between pH 9 and 13The essential substance is GeO (OH)₃ ^1-And the predominant species at a pH greater than 13 is GeO₂(OH)₂ ^2-. The first dissociation constant of germanic acid is 4.9-10mol. L-1(log K)_Dissociation-9.31) (Wood and Iain, Ore geomology Reviews 28, stage 1 (2006): 57-102).

Acid leaching

Can use H₂SO₄H at a temperature between 40 and 60 ℃ at 100g/L₂SO₄The concentration is continued for 30min and the germanium is leached using a solid to liquid ratio of 1:4, and 78% germanium is recovered. At a higher temperature, about 85 ℃ and 150g/L of H₂SO₄Concentration and residence time of 1h, resulting in a mixed leaching of the different Metals with a germanium extraction of 92.7% (Rut legs et al Metals 5, stage 3 (2015): 1520-.

Patent application CN108486390A describes a method for separating germanium and gallium from germanium and gallium containing materials. In a first step, germanium and gallium material is added to 50 to 150g/L H in a ratio of 5 to 10% w/w₂SO₄And subsequently adjusting the pH to between 1 and 3. The leach solution is neutralized at pH 1 to 3 for subsequent addition of zinc powder to the neutralized liquor at a temperature of 40-80 ℃ in order to obtain a germanium concentrate and a liquid solution. To this liquid solution zinc powder is again added at a temperature of 40-80 ℃ in order to obtain gallium residues and a liquid solution.

Alkaline leaching

Patent application CN108300876A describes a method for leaching gallium and germanium from the slag of a method for obtaining zinc. In the first step, the slag is ground to a size of 50-100 microns and then 0.1-1mol/L H is added at a liquid-to-solid ratio of 4-10:1ml/g₂SO₄Solution at a leaching temperature of 25-80 ℃ of 100-Stirring at rpm for 0.25-4 h. Then carrying out solid-liquid separation to obtain leaching residue and H₂SO₄Adding 0.2-2mol/L hydrogen peroxide solution into the leaching solution at a liquid-solid ratio of 4-10:1mL/g, and adding 0.1-1mol/L NaOH to adjust the pH of the leaching solution to 5.0-8.0. This alkaline leaching is carried out at 25-80 ℃ with stirring at 100-600rpm for 0.25-4h in order to obtain a germanium-rich leaching solution. In the case of the present invention, in order to be able to leach germanium values, a sulphuric acid leach and a citric acid leach step are necessary in order to remove high levels of lead from the metallurgical residue and thus increase the germanium content of the residue due to mass loss of the metallurgical residue in the citric acid leach step. Furthermore, by creating a process such as one of the applications, the presence of lead in the metallurgical residue will lead to a greater consumption of soda ash, which will affect the germanium leaching yield, due to the conversion of lead sulphate to lead hydroxide.

Ion exchange

Following the leaching step, the germanium obtained must be concentrated by different alternatives, such as solvent extraction, precipitation with chelating agents or ion exchange. For the present invention, the ion exchange step is considered each time, since it is a step with a better recovery level and also allows to increase the germanium concentration ratio with respect to other metals present in the leaching solution, such as lead, aluminum, silicon and arsenic.

Patent US4525332 describes the adsorption of a germanium-containing solution in an ion exchange resin consisting of a polymer having functional groups selected from secondary, tertiary and quaternary ammonium groups, said ion exchange resin having a relative selectivity of germanium to antimony of 50:1, after which the germanium collected by the resin is eluted in an aqueous medium. The present invention utilizes resins containing N-methylglucamine as a functional group. Patent US4525332 does not specify that the feed solution may contain silicon. In the example of application of patent US4525332, the addition is carried out with an acid or slightly acid solution, instead of with a basic solution, as shown in the present invention. By being an element of the same family as germanium, silicon can interfere when using a recycled feed solution in the process. In the same way, depending on the resin used, with H₂SO₄Or sodium hydroxide, which is carried out as described herein and for tetrachloroThe subsequent step of germanium distillation necessarily involves elution with HCl and the opposite. Furthermore, the present invention differs from patent US4525332 in that the latter does not teach how to remove silicon from the alkaline solution in order to recycle the solution as leaching medium for metallurgical residues in order to obtain germanium.

Patent application GB933563A teaches treating an aqueous neutral or slightly acidic solution of germanium in an ion exchange resin containing hydroxyphenyl groups which collects the germanium, for subsequent elution thereof with a 7N hydrochloric acid solution for subsequent distillation, and 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, it is not obvious to neutralize the solution to effect the germanium addition in the resin, because silicon will precipitate and drag the germanium when the pH is lowered. In this case, once the solution has passed through the column and is depleted of germanium, it is necessary to feed and remove the silicon without prior precipitation. In addition, unlike patent application GB933563, in the present invention elution can be carried out with solutions having Ge concentrations less than 6N without evidence of significant loss of germanium.

Germanium distillation

Powell et al j.appl.chem.1951,541-551 teach leaching of smelted powder with HCl to produce germanium tetrachloride in situ which can be distilled at a temperature of 84 ℃. One of the problems with this process is the presence of arsenic trichloride because even if it boils at 130 ℃, its vapor pressure is high enough at 84 ℃ to distill with germanium tetrachloride. In this case the invention differs from the teachings of Powell in the sense that there is a series of prior leaching steps and ion exchange separations which prevent elements such as arsenic from being present in elevated concentrations which interfere with germanium distillation.

Patent US3102786 teaches the purification of germanium tetrachloride using a solution of HCl with a minimum concentration of 6N and maintaining the column in a continuous process at a temperature between 83 and 110 ℃. The present invention differs from US3102786 in that the distillation can be carried out at an acid concentration of less than 6N, allowing effective distillation of germanium at a distillation rate of more than 95%.

Patent US2811418 teaches a process for the purification of germanium tetrachloride, using a 12N concentration and chlorine-saturated HCL solution, allowing the mixture to separate into two phases, the heavier phase containing purified germanium tetrachloride. The present invention differs from US2811418 in that the distillation can be carried out at acid concentrations below 12N, in particular below 5N, allowing effective distillation of germanium with distillation rates exceeding 95% with minimal arsenic drag.

Germanium hydrolysis

Patent US3455645 discloses a process for producing amorphous germanium dioxide, characterized in that it precipitates germanium present in an aqueous solution, wherein the pH is at least 5 and at most 9. In particular, the experiment disclosed in US3455645 teaches the addition of germanium tetrachloride to a solution containing 10 parts NaOH/90 parts water until the pH is below 8, or preferably below 6. Patent US3455645 differs from the present invention in that the distilled solution obtained is sent directly to a cooled reactor, in which it has been verified that germanium dioxide precipitates without the need to control the pH to the value specified in US 3455645.

Hoffmann in Extracting and reflecting Germanium, Journal of Metals, 7.1987, 42-45 states that at HCl concentrations of less than 5.5, Germanium is found predominantly as germanic acid and 3 grams of water is sufficient to hydrolyze 1g of Germanium with a yield of 95%. In addition, it states that the precipitation is preferably carried out at temperatures close to 0 ℃ and that germanium dioxide can be used as a nucleation center for germanium precipitation. The present invention shows that improved germanium precipitation yields can be obtained without the need to use germanium dioxide as a seed to confirm germanium precipitation. The present invention demonstrates that germanium dioxide can be precipitated by using a sufficiently concentrated germanium solution, and that a low HCl concentration negatively affects the germanium dioxide precipitation process compared to an HCl concentration close to 3.7N (135g/L HCl).

Description of the drawings

FIG. I shows a flow chart of the method of the present disclosure.

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

Figure III shows a second distillation curve of germanium from the primary distilled solution. The second distillation refers to the distillate solution collected after the first distillation cycle.

Figure IV shows a third distillation curve of germanium from the twice distilled solution. The third distillation refers to the distillate solution collected after the second distillation cycle.

Description of the invention

Broadly, the present invention describes a method of producing germanium from metallurgical residues.

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

In a still more preferred embodiment, the present invention describes a method of producing a solid germanium concentrate.

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

In broad terms, the invention describes a process for producing a germanium concentrate from metallurgical residues, in particular from residues containing copper, iron, lead, silicon and germanium and optionally containing elements such as arsenic, antimony and bismuth, characterized in that it comprises:

step (i) of leaching copper from the metallurgical residue (1) using a first acid solution (2) so as to obtain a first leach solution (3) enriched in copper and iron and optionally arsenic, antimony and bismuth and a first leached sludge (4) having a reduced copper and iron content and optionally a reduced arsenic content and enriched in lead, silicon and germanium,

a step (ii) of leaching the first leached sludge (4), wherein said first leached sludge (4) is treated with a first solution (5) of a carboxylic acid salt, so as to obtain a second leached sludge (6) depleted in lead and a second leaching solution (7) enriched in lead,

a step (iii) of alkaline leaching the second leached sludge, in which alkali (8) is added in order to form an alkaline leaching solution in order to obtain a third leached sludge (9) with a reduced silicon and germanium content and a third leaching solution (10) enriched in germanium and silicon and optionally arsenic,

step (iv) of feeding the third leaching solution (10) enriched in germanium and silicon and optionally arsenic in an ion exchange column, wherein the germanium is captured by a resin, so as to obtain a fourth alkaline solution (11) depleted in germanium and enriched in silicon,

(vi) a step (v) of washing the ion exchange column with water (12) in which a fifth washing solution (13) of the column charge is obtained,

(vii) performing step (vi) of eluting the ion exchange column with HCl solution (14) so as to obtain a sixth elution solution (15) enriched in germanium,

a distillation step (vii) in which the sixth elution solution (15) enriched in germanium is distilled so as to obtain a seventh solution (16) of germanium and an eighth solution (17) devoid of germanium, and

(viii) a hydrolysis step, wherein a seventh solution (16) of germanium is contacted with an aqueous solution (18) to produce a first GeO₂A concentrate (19) and a ninth solution (20) lacking germanium.

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

In an even more preferred embodiment, the metallurgical residue has been subjected to a copper leaching process.

In an even more preferred embodiment, the metallurgical residue has been subjected to treatment with H₂SO₄And (4) leaching.

In a preferred embodiment, the metallurgical residues to be processed comprise the mineral substances chalcoalumite, covellite, CuOFe₂O₃Form of cuprochoxides, ZnOFe₂O₃Forms of gahnite, magnetite, iron (III) oxide, pyrite (pirita), scorodite, muscovite (mucovita), kaolinite, and lead (II) sulfate.

In an even more preferred embodiment, the copper contained in the metallurgical residue is as copper sulfate, chalcocite (calcosina), covellite and CuOFe₂O₃The form of cupferrospinel exists.

In an even more preferred embodiment, the metallurgical residue contains at least 50% of the copper as CuOFe₂O₃The form of the cupferrospinel exists.

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

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

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

In a preferred embodimentIn step (i), first H₂SO₄The solution may comprise H₂SO₄And/or refinery waste water.

In a preferred embodiment, in H₂SO₄The concentration is between 150 and 300g/L, more preferably H₂SO₄Step (i) was carried out at a concentration of 250 g/L.

In a preferred embodiment, step (i) is carried out at a temperature between 50 and 130 ℃, more preferably at a temperature of 85 ℃.

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

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

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

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

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

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

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

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.

In a preferred embodiment, in step (ii), the corresponding acid of the carboxylate is added for adjusting the pH.

In an even more preferred embodiment, in step (ii), citric acid is added for adjusting the pH.

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

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

In a preferred embodiment, in Mg (OH)₂The second base used in the leaching of step iv is chosen between KOH or NaOH.

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

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

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

In a preferred embodiment, step (iv) is carried out with a resin having nitrogen atom groups (N-donor groups).

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

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

In a preferred embodiment, step (v) is carried out by feeding the flushed fourth solution in a ratio of bed volumes between 5 and 15.

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

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

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

In another preferred embodiment, the fourth alkaline solution, which is devoid of germanium, is subjected to a silicon removal process.

In another preferred embodiment, slaked lime or aluminum sulfate is added during the silicon removal process.

In an even more preferred embodiment, hydrated lime is added during the silicon removal process.

In an even more preferred embodiment, slaked lime is added in a molar ratio of 1:1 relative to silicon contained in the fourth alkaline solution lacking germanium, so as to produce a regenerated alkaline solution and a solid consisting of calcium silicate.

In another preferred embodiment, the silicon removal step is carried out at a temperature between 20 and 90 ℃.

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

In another preferred embodiment, the distillation step (vii) is carried out at a temperature between 86.5 ℃ and 107 ℃ and at a bulb temperature between 86.5 and 108 ℃.

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

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

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

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

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

In an even more preferred embodiment, the contacting of the germanium solution concentrated in step (viii) with deionized water is performed at a temperature between 2 and 15 ℃.

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

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

In a preferred embodiment, the first leaching solution enriched in copper is routed to copper leaching of smelted powder.

In a preferred embodiment, the first leach solution enriched in copper is sent to an arsenic elimination (anatomiento) process.

In a preferred embodiment, the arsenic elimination process is selected from those that consider ferric arsenate production.

In an even more preferred embodiment, the arsenic elimination process is a scorodite production process.

Application examples

The following examples are to be considered as embodiments of the present invention and not to be considered as limitations of the present invention as various modifications thereof are covered within the claimed subject matter.

Sulfuric acid leaching

Examples 1 to 7

Preparation of a catalyst having H between 2550 and 2850g₂SO₄Is arranged in a 5L glass reactor, wherein the sludge previously subjected to the copper leaching process is added to a solids content of between 5 and 10% w/w. The mineralogical composition of the sludge is shown in table 1. The reactor was stirred at 300rpm for 3 to 6 hours at 85 ℃. Once the reaction time is over, the slurry is filtered in a Buchner system. The results are shown in Table 2.

TABLE 1 mineralogical composition of the sludge

Substance(s)	Unit of	Value of
			PbSO₄	％	12.84
PbS	％	0.1
			PbO	％	0.1
CuSO₄	％	2.54
			Cu₂S	％	0.63
CuS	％	4.02
			CuO	％	0.71
CuOFe₂O₃	％	15.09
			ZnOFe₂O₃	％	4.46
ZnS	％	2.94
			Fe₃O₄	％	4.74
Fe₂O₃	％	4.91
			FeS₂	％	6.32
Ag2S	％	0.1
			FeAsO₄*2H₂O	％	5.18
Bi₂O₃	％	0.59
			Sb₂O₃	％	0.5
KAl₃Si₃O₁₀(OH)₂	％	7.01
			Al₂Si₂O₃(OH)₄	％	2.92
Ge	g/ton	548

Table 2 examples 1 to 7Cu leaching results

Variables/examples	Unit of	1	2	3	4	5	6	7
									H₂SO₄Concentration of	g/L	150	250	150	250	250	150	250
Solids content	％w/w	5	5	15	15	15	20	20
									Time of leaching	h		6	6	6	3	6	6	6
Cu leaching yield	％	75.9	76.1	68.0	60	69.7	64.8	67.7

Examples 8 to 10

Preparation of 2550g of 250g/LH₂SO₄Solution, which was placed in a 4L autoclave, to which sludge previously subjected to a copper leaching process was added to a solids content of 15% w/w. The reactor was stirred at 300rpm for 1 to 6 hours at 130 ℃. Once the reaction time is over, the slurry is filtered in a Buchner system. The results are shown in Table 3.

Table 3 examples 8 to 10Cu leaching results

Variables/examples	Unit of	8	9	10
					Time of leaching	h		1	3	6
Cu leaching yield	％	75.9	76.1	82.0
					Loss of mass	％	35.0	41.0	42.0

Example 11

A refining waste water solution (Table 4) was prepared and its content was adjusted to H₂SO₄The concentration was adjusted to 250g/L, which was arranged in a 5L glass reactor, where 450g of sludge previously subjected to a copper leaching process was added in order to produce a slurry with 15% w/w solids. The reactor was stirred at 300rpm for 6 hours at 85 ℃. Once the reaction time is over, the slurry is filtered in a Buchner system. The results showed that the leaching yield of Cu was 72.0%, that of Fe was 62.0%, that of As was 71.5%, that of Zn was 57.0% and that the mass loss was 38.5%.

TABLE 4 refining waste Water composition

Citric acid leach

Example 12

A solution was prepared with 40L of distilled water to which 14kg of sodium citrate was added and the pH was adjusted to 7.0 with 800g/L of citric acid solution. Once the reagents were dissolved, 6kg of leached sludge was added as in example 3. The head mud (La borra de cabeza) had a Pb content of 15.4%. The leaching was carried out at 20 ℃ and stirred at 1,000rpm for 9 h. A leaching efficiency of 94% for Pb is obtained, so that a leached sludge with a mass reduction of 24% and a Pb content of 1.19% is obtained.

Examples 13 to 19

The solution was prepared with 2L of distilled water, with a sodium citrate concentration between 323 and 368g/L and a pH between 5.3 and 8.8. The pH was adjusted with a citric acid solution at 800 g/L. Once the reagents were dissolved, the sludge treated according to example 3 was added in a ratio of sodium citrate/g sludge comprised between 1.2 and 2.3 g. The head sludge has a Pb content between 15.0 and 15.1%. The leaching is carried out at between 30 and 60 ℃ and stirring is carried out between 500 and 700rpm for a time between 2 and 4 h. The results are shown in Table 5.

Table 5 examples 13 to 19 citric acid leach results

Alkaline leaching

Examples 20 to 28

The slurry is prepared with a sodium hydroxide solution having a concentration between 5.4 and 8.7% w/w and leached sludge having undergone successive copper and lead leaching processes, wherein the solids content is between 5.0 and 7.0% w/w. The slurry was placed in a 4L autoclave and heated at 600rpm to a temperature between 100 and 140 ℃ for between 1 and 6 hours. Once the leaching time is complete, the slurry is cooled and filtered in a buchner system. The results are shown in Table 6.

Table 6 results for examples 20 to 28

Examples 29 and 30

A slurry was prepared with 6230mL of water to which were added 420g of sodium hydroxide and 350g of leached sludge subjected to successive copper and lead leaching processes in order to obtain a NaOH concentration of 6.0% w/w and a solids of 5.0% w/w. The slurry was placed in a 10L glass reactor and heated at 90 ℃ for between 1 and 6 hours and stirred at 900 rpm. Once the leaching time is complete, the slurry is cooled and filtered in a buchner system.

Table 7 results of examples 29 and 30

Variables/examples	Unit of	29	30
				Residence time	h		1	6
Leaching yield
				Ge	％	78.1	82.0
Si	％	63.2	63.0

Ion exchange

Examples 31 to 38

12500mL of a solution derived from alkaline leaching of sludge having a Ge concentration of 28 to 34mg/L and a Si concentration of 7.2 to 8.6g/L was taken and passed through an ion exchange column having 400mL of a resin having nitrogen atom groups. The feed rate was 67mL/min to 133mL/min at a rate of 3.4 to 6.7 cm/min. The column was rinsed with 2000mL of water and no germanium elution was observed in any experiment. Germanium elution was performed with HCl at a rate of 99g/L over between 1 and 2 bed volumes. At a rate between 1 and 3.4cm/min, the elution flow was between 20 and 67 mL/min. Finally, the column was rinsed with 2000mL of water and no Ge resistance (arastratre) was observed in any experiment. The results are shown in Table 8.

TABLE 8 results from experiments 31 to 38

Silicon removal

Examples 39 to 41

2000 of the feed solution from the ion exchange column, having a Si concentration of 18.5g/L and a pH of 3.7, was taken, to which was added aluminum sulfate tetradecahydrate having an Al/Si molar ratio of 0.43 to 1.0. The mixture was stirred at 400rpm and at a temperature of 80 ℃ for 60 minutes. Once the reaction was complete, the slurry was filtered and Si removal efficiency of 74% was obtained, reducing the Si concentration to 2.8g/L and the Al concentration to 130mg/L, at a pH of 13.4.

TABLE 9 results for examples 39 to 41

Variables/examples	Unit of	39	40	41
					Ratio of Al to Si	h	0.43	0.70	1.00
Yield of Si precipitate	％	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

Examples 42 to 45

4200g of an IX (ion exchange) draw-off solution having 17.4g/L Si was taken at a temperature in the range from 20 to 70 ℃ in a 5L reactor, to which calcium hydroxide was added in a molar ratio in the range from 0.9 to 1.1mol Ca/mol Si, and stirring was continued at 300rpm for 60 minutes. Once the reaction was complete, the slurry was filtered through No. 42 filter paper.

TABLE 10 results for examples 42 to 45

Variables/examples	Unit of	42	43	44	45
						Ca/Si ratio	h	0.9	1.1	1.1	1.1
Temperature of	℃	70	20	50	70
						Yield of Si precipitate	％	82	70	83	95

Germanium distillation

Example 46

18000mL of a solution having 454mg/L Ge, 1039mg/L Pb, 294mg/L Al, 7mg/L As, and 150g/L HCl was taken As an output solution for elution from the ion exchange column. The solution was placed in a 20L distillation bulb and heated to 108 ℃. Once the bulb temperature reached 108 ℃, the distillate solution output was observed and kept in fractions between 300 and 1200 mL. The evaporated solution was condensed in a coil through which cold water was circulated at 5 ℃ and received in a jacketed collection cup through which water was circulated at 5 ℃. A total of 100% Ge was collected in the distillate, yielding a fraction of up to 1960mg/L Ge. 0.02% of the Pb present in the solution went into the distillate, whereas 0.19 and 50% of Al and As, respectively, were entrained by the distillate.

Example 47

14000mL of a solution having 1060mg/L Ge and 100g/L HCl was taken as a germanium distilled solution. The solution was placed in a 20L distillation bulb and heated to 108 ℃. Once the bulb temperature reached 108 ℃, the distillate solution output was observed and kept in 500mL fractions. The evaporated solution was condensed in a coil through which cold water was circulated at 5 ℃ and received in a jacketed collection cup through which water was circulated at 5 ℃. A total of 100% Ge was collected in the distillate, yielding a fraction of up to 4330mg/L Ge.

Example 48

16000mL of a solution having 2630mg/L Ge and 117g/L HCl was taken as the germanium distilled solution. The solution was placed in a 20L distillation bulb and heated to 108 ℃. Once the bulb temperature reached 108 ℃, the distillate solution output was observed and kept in 1000mL fractions. The evaporated solution was condensed in a coil through which cold water was circulated at 5 ℃ and received in a jacketed collection cup through which water was circulated at 5 ℃. A total of 100% Ge was collected in the distillate, yielding a fraction of up to 15000mg/L Ge.

Example 49

16000mL of a solution having 10400mg/L Ge and 150g/L HCl was taken as the germanium distilled solution. The solution was placed in a 20L distillation bulb and heated to 108 ℃. Once the bulb temperature reached 108 ℃, the distillate solution output was observed and kept in 1000mL fractions. The evaporated solution was condensed in a coil through which cold water was circulated at 5 ℃ and received in a jacketed collection cup through which water was circulated at 5 ℃. A total of 100% Ge was collected in the distillate, yielding a fraction of up to 25000mg/L Ge.

Germanium hydrolysis

Examples 50 to 57

250mL of a germanium solution having a concentration in the range of 4.1 to 24.8g/L Ge and an HCl concentration between 54 and 134g/L were taken, arranged in a 500mL jacketed reactor, and 5000mL of water mixed with 1 vol% ethylene glycol and cooled by cooling means at 1 ℃ were circulated through the reactor. The solution was mechanically stirred at 250rpm for 5h at a temperature between 2 and 3 ℃. The solution was filtered at the end of the process in 0.45 μm filter paper.

TABLE 11 results for examples 50 to 57

Variables/examples	Unit of	50	51	52	53	54	55	56	57
										Start of
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
Precipitate	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. Process for producing a germanium concentrate from metallurgical residues, in particular from residues containing copper, iron, lead and germanium and optionally arsenic, antimony and bismuth, characterized in that it comprises:

i. leaching copper from the metallurgical residue with a first acid solution so as to obtain a first leach solution enriched in copper and iron and optionally arsenic, antimony and bismuth and a first leached sludge enriched in lead, silicon and germanium having a reduced copper and iron content and optionally a reduced arsenic content,

leaching the first leached sludge, wherein said first leached sludge is treated with a first solution of a carboxylic acid salt, so as to obtain a lead-deficient second leached sludge and a lead-enriched second leach solution,

alkaline leaching the second leached sludge, wherein alkali is added to form an alkaline leaching solution, to obtain a third leached sludge having a reduced silicon and germanium content and a third leaching solution enriched in germanium and silicon and optionally arsenic,

feeding in an ion exchange column wherein germanium is captured by a resin to obtain a fourth alkaline solution depleted of germanium and enriched in silicon,

v. rinsing the ion exchange column with water, wherein a fifth solution of column feed rinse is obtained,

eluting the ion exchange column with HCl solution to obtain a sixth elution solution enriched in germanium,

distilling, wherein the sixth elution solution enriched in germanium is distilled so as to obtain a seventh solution of germanium and an eighth solution depleted in germanium, and

hydrolysis, wherein a seventh solution of germanium is contacted with an aqueous solution to produce a first GeO₂And (3) concentrating.

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

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

4. A method according to any one of claims 1, 2 or 3, characterized in that the metallurgical residue has undergone a copper leaching process.

5. According to claim4, characterized in that the metallurgical residue has undergone a treatment with H₂SO₄And (4) leaching.

6. A method according to any one of claims 1 to 3, characterized in that the metallurgical residues to be treated comprise the mineral substances ettringite, covellite, CuOFe₂O₃Form of cuprochoxides, ZnOFe₂O₃Forms of gahnite, magnetite, iron (III) oxide, pyrite, scorodite, muscovite, kaolinite and lead (II) sulfate.

7. Process according to claim 6, characterized in that the metallurgical residues contain copper as copper sulphate, chalcocite, covellite and CuOFe₂O₃The form of cupferrospinel exists.

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

9. A method according to any one of claims 1 to 7 or 1 to 8, characterized in that the lead contained in the metallurgical residue is present as lead (II) sulphate, galena or lead (II) oxide.

10. A method according to claim 9, characterised in that at least 95% of the lead is lead (II) sulphate.

11. The process according to any one of claims 1 to 10, characterized in that the first H from step i₂SO₄The solution may comprise H₂SO₄And/or refinery waste water.

12. The process according to any one of claims 1 to 11, characterized in that in H₂SO₄Step (i) was carried out at a concentration between 150 and 300 g/L.

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

14. The process according to any one of claims 1 to 13, characterized in that step (i) is carried out for a time between 3 and 12 hours.

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

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

17. The process 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 and 1M.

18. The method according to any one of claims 1 to 17, characterized in that in step (ii) the first leached sludge is fed to the sodium citrate solution in a mass ratio of 1: 9.

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

20. The process according to any one of claims 1 to 19, characterized in that step (ii) is carried out for a residence time of between 1 and 23 h.

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

23. The process according to any one of claims 1 to 22, characterized in that the pH adjustment in step (ii) is carried out with a citric acid solution of 600 to 900 g/L.

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

25. A process according to any one of claims 1 to 24, characterised in that the alkali used in the leaching of 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 in a ratio between 5 and 10% w/w with respect to the total mass of the alkaline leaching solution.

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 ℃.

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

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

30. The process according to any one of claims 1 to 29, characterized in that step (vi) is carried out by feeding the third leaching solution enriched in germanium and silicon in a ratio of bed volumes between 2 and 3.

31. A process according to any one of claims 1 to 30, characterised in that step (v) is carried out by feeding a fourth feed of flushing solution at a ratio of bed volume between 5 and 15.

32. The process according to any one of claims 1 to 31, characterized by step (vi) of elution with a solution of HCl.

33. The process according to any one of claims 1 to 32, characterized in that step (vi) of elution is carried out with a solution of HCl having a concentration between 2 and 8N.

34. Process according to any one of claims 1 to 33, characterized in that step (vi) is carried out by feeding the HCl solution in a ratio of bed volumes between 1 and 5.

35. The method according to any one of claims 1 to 34, characterized in that the fourth alkaline solution, which is devoid of germanium, is subjected to a silicon removal process.

36. A method according to any one of claims 1 to 35, characterized in that during the silicon removal process slaked lime is added.

37. Process according to any one of claims 1 to 36, characterized in that slaked lime is added in a molar ratio of 1:1 with respect to the silicon contained in the fourth alkaline solution devoid of germanium, so as to produce a regenerated alkaline solution and a solid consisting of calcium silicate.

38. The method according to any one of claims 1 to 37, characterized in that the silicon removal step is carried out at a temperature between 30 and 90 ℃.

39. The process according to any one of claims 1 to 38, characterized in that regenerated alkaline solution is recycled to step (iii) of alkaline leaching.

40. Process according to any one of claims 1 to 39, characterized by the fact that step (vii) of distillation is carried out using a temperature between 86.5 ℃ and 107 ℃ and a bulb temperature between 86.5 and 108 ℃.

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

42. The method of claim 41, wherein the seventh germanium solution is distilled 3 more times to produce a concentrated germanium solution and a distilled HCl solution.

43. Process according to any one of claims 1 to 42, characterized in that the distilled HCl solution is recycled in a preceding distillation step in order to increase the HCl concentration at the inlet of the distiller.

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

45. A process according to claim 44, characterized in that the germanium concentrate obtained in step (viii) is GeO₂。

46. The process according to any one of claims 1 to 45, characterized in that the contacting of the germanium solution concentrated in step (viii) with deionized water is carried out at a temperature between 2 and 15 ℃.

47. The method according to any one of claims 1 to 46, characterized in that the concentrated germanium solution sent to the hydrolysis step has a germanium concentration of between 8.1 and 24.8 g/L.

48. A process as claimed in any one of claims 1 to 47, characterized in that the concentrated germanium solution sent to step (viii) has a HCl concentration of between 55 and 135 g/L.

49. A method according to any one of claims 1 to 48, characterized in that the first leaching solution enriched in copper is routed to the copper leaching process of the smelted powder.

50. A method according to any one of claims 1 to 49, characterized in that the first leaching solution enriched in copper is routed to an arsenic elimination process.

51. The method according to claim 50, characterized in that the arsenic elimination process is selected from those considering ferric arsenate production.

52. A method according to claim 51, characterized in that the arsenic elimination process is a scorodite production process.