A method for cascade resource utilization of germanium, lead, zinc and iron in hard zinc slag

Abstract

The present invention proposes a stepwise resource recovery method for germanium, lead, zinc and iron in hard zinc slag, and belongs to the technical field of valuable element extraction in lead and zinc smelting waste. The method of the present invention includes 8 steps of reduction smelting, fuming volatilization, acid production, leaching, reduction precipitation of germanium, two-stage pyrolysis, chlorination distillation and rectification. The reduction smelting can recover the lead in the hard zinc slag in the form of metallic lead, the fuming volatilization can enrich zinc and germanium in the form of smoke dust, the iron is enriched in the water-quenched slag, the zinc and germanium are leached into the solution, and then tannin is used to precipitate germanium, which can enrich all the germanium in the tannin germanium slag, the tannin germanium is pyrolyzed to obtain different products, the washing of the pyrolysis product is subjected to chlorination distillation to obtain crude germanium tetrachloride, and the crude germanium tetrachloride is rectified to obtain refined GeCl 4 In the present invention, the lead direct recovery rate can reach 98%, the zinc direct recovery rate can reach 96.92%, the iron resource utilization rate can reach 99%, and the germanium direct recovery rate can reach 95.04%.

Description

Technical Field

[0001] The invention belongs to the technical field of valuable element extraction in lead-zinc smelting waste, and in particular relates to a cascade resource recovery method for germanium, lead, zinc and iron in hard zinc slag. Background Art

[0002] Germanium and its compounds have many excellent optoelectronic properties, such as partial transparency in the infrared electromagnetic spectrum, extended three-dimensional network former of glass germanium oxygen tetrahedron, low dispersion and high infrared refractive index, and are often listed as important basic raw materials in the fields of information communication, modern aviation and new energy. With the rapid development of the global digital economy and aerospace, the gap between supply and demand of germanium is expected to reach 48.87 tons by 2027. Establishing a new method for deep enrichment of germanium in existing raw materials is an important way to alleviate resource tension. The main raw materials for germanium extraction are zinc oxide smoke, lignite, hard zinc slag and germanium waste. Among them, the low recovery rate of germanium in hard zinc slag and the low recovery rate of other valuable elements have always been a problem that has plagued the germanium industry.

[0003] Regarding the recovery of valuable elements in hard zinc slag, the patent with application number CN201210306079.1 involves a method for recovering germanium and indium from germanium-containing materials produced in the lead-zinc smelting process. It uses neutral leaching of part of zinc, pre-oxidation and then leaching of germanium and indium to achieve separation and recovery of zinc, indium and germanium in germanium-containing materials, and can replace the existing chlorine oxidation leaching process of germanium, which is environmentally friendly, but the method does not study the recovery of other major valuable elements such as iron, lead and zinc. The patent with application number CN201510252678.3 involves a method for extracting scattered metal indium and germanium from germanium slag, which uses sulfuric acid leaching, oxidative roasting and chlorination distillation, and has a high germanium recovery rate, but the method also does not study the recovery of other major valuable elements such as iron, lead and zinc. Patent application number CN201510741534.4 relates to a mixed oxidant and a method for recovering germanium from hard zinc leaching slag using the mixed oxidant. CaCl2, NaClO, and Ca(ClO)2 are used as mixed oxidants to synergistically enhance ultrasonic leaching of low-priced germanium in hard zinc slag, and the recovery rate of germanium reaches 87%, but only germanium is recovered. Patent application number CN201810357797.9 relates to an oxidative roasting method for hard zinc vacuum smelting slag. After the vacuum smelting slag in an alloy state is mixed with an oxidant, microwaves are used to generate instantaneous high temperature, and the alloy is quickly and evenly oxidized into metal oxides to enhance hydrometallurgical leaching, but the recovery of other valuable metals in hard zinc slag is not studied. The patent with application number CN202011505745.5 relates to a method for recovering hard zinc materials. It uses a low-acid zinc leaching-high-acid germanium and silver leaching method to achieve efficient recovery of germanium, silver and lead in hard zinc slag. However, zinc and germanium in hard zinc slag are mainly present in a non-sulfuric acid leaching state. The patent with application number CN202310763551.2 relates to a method for extracting germanium from hard zinc. It first converts GeO2 into Na2GeO3 by alkaline roasting, and then enriches germanium in the slag after water leaching and neutralization. The direct recovery rate of germanium is high, but the extraction process will produce a large amount of wastewater, and the recovery of other valuable elements in hard zinc slag has not been studied.

[0004] In summary, the existing technology mainly strengthens the extraction of germanium from hard zinc slag by changing process factors, but does not involve the main valuable elements zinc, iron and lead in the slag. At present, a complete and effective treatment process for hard zinc slag has not been established. Summary of the invention

[0005] Aiming at the current problems of waste of lead, zinc and iron resources in hard zinc slag, low germanium extraction rate, large amount of HCl required for distillation, large amount of subsequent waste liquid and waste slag, etc., the present invention proposes a cascade resource recovery method for germanium, lead, zinc and iron in hard zinc slag. The present invention realizes the cascade efficient resource recovery of germanium, lead, zinc and iron in hard zinc slag through four steps of reduction for lead separation, fuming for iron separation, leaching for zinc separation and chlorination for germanium extraction.

[0006] To achieve the above object, the present invention provides a method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag, comprising the following steps:

[0007] (1) Reduction smelting: hard zinc slag is mixed with coke and reduced to obtain crude lead, reduction slag and flue gas;

[0008] (2) Fuming and volatilizing: fumigating and volatilizing the reduced slag to obtain water-quenched slag, zinc oxide fume and flue gas;

[0009] (3) Acid production: The flue gases produced by reduction smelting and fume volatilization are mixed to produce acid;

[0010] (4) Leaching: slurrying the zinc oxide fume to obtain leaching residue I and leaching solution I, and acid leaching the leaching residue I to obtain leaching solution II and lead slag;

[0011] (5) Reduction and precipitation of germanium: tannin is dispersed with a dispersant and then ultrasonically dissolved to obtain a tannin dispersion solution, which is then added to the leaching solution I to reduce and precipitate germanium, thereby obtaining tannin germanium and a germanium precipitation solution;

[0012] (6) Two-stage pyrolysis: the tannin germanium is subjected to two-stage pyrolysis to obtain coke and germanium concentrate, which are separated by flotation;

[0013] (7) Chlorination distillation: Chlorination distillation of the germanium concentrate to obtain crude GeCl4 and distillation residue;

[0014] (8) Distillation: distilling the crude GeCl4 to obtain refined GeCl4.

[0015] As described above, the method of the present invention comprises eight steps of reduction smelting, fuming volatilization, acid preparation, leaching, reduction germanium precipitation, two-stage pyrolysis, chlorination distillation and rectification. The reduction smelting can recover the lead in the hard zinc slag in the form of metallic lead (crude lead), the fuming volatilization can enrich zinc and germanium in the form of smoke dust, the iron is enriched in the water-quenched slag, the zinc and germanium are leached into the solution and then tannin is used to precipitate germanium, the zinc exists in the solution after the germanium precipitation and is used for back electrolysis, the germanium is completely enriched in the tannin germanium slag, the tannin germanium is pyrolyzed to obtain different products, the pyrolysis product washing is subjected to chlorination distillation to obtain crude germanium tetrachloride, and the crude germanium tetrachloride is rectified to obtain refined GeCl4.

[0016] Furthermore, in step (1), the germanium content in the hard zinc slag is 2000-5000 g / t, the zinc content is 10-20%, the iron content is 15-25%, the lead content is 20-30%, and the volatilization rate of lead in the reduction smelting process is 96-98%.

[0017] Further, the amount of coke added to the hard zinc slag is 11-15wt%. More preferably, the amount of coke added to the hard zinc slag is 11.32-14.02wt%.

[0018] Furthermore, in step (1), the reduction smelting temperature is 1000-1200° C., and the reduction smelting time is 90-120 min.

[0019] Furthermore, in step (2), the fuming volatilization temperature is 1100-1300° C., and the fuming volatilization time is 60-120 min.

[0020] Furthermore, in step (2), the reducing slag is mainly composed of ZnS, FeS, Fe2O3, Fe3O4 and germanate.

[0021] Furthermore, after the fuming volatilization in step (2), the zinc in the reduction slag is volatilized in the form of metallic zinc, and the germanate is volatilized in the form of GeO, and finally they are enriched in the dust collection system in the form of ZnO, zinc germanate or GeO2 to obtain zinc oxide fume.

[0022] Furthermore, after the fuming and volatilization in step (2), the water-quenched slag is subjected to magnetic separation to obtain CaSiO3 tailings and iron slag, and the iron slag is mainly composed of Fe3O4.

[0023] Furthermore, in step (3), the reducing smelting flue gas and the oxidizing volatilizing flue gas are combined and then enter the acid production process, and sulfuric acid is obtained by two conversions and two absorptions.

[0024] Furthermore, in step (4), the leaching temperature of the slurry leaching is 60-80°C, the leaching time is 10-30 min, the leaching liquid-solid ratio is (4-7) mL:1 g, and the leaching endpoint pH is 1-3.

[0025] Furthermore, in step (4), the leaching temperature of the acid leaching is 80-95°C, the leaching time is 90-180 min, the leaching liquid-solid ratio is (8-10) mL:1 g, and the leaching endpoint acidity is 20-40 g / L.

[0026] Furthermore, in step (5), the tannin is quinoa tannin, tara tannin, persimmon tannin or chestnut tannin, and the tannin is dispersed and dissolved in a dispersant under ultrasonic action to obtain a tannin dispersion solution, the dispersant is methanol, ethanol, acetone or ethylene glycol, the volume concentration of the dispersant is 0.5‰~2‰, and the concentration of the tannin dispersion solution is 5~20wt%.

[0027] Furthermore, the intensity of the ultrasound is 0.16-0.33 W / cm 2 , the dissolving pH is 2~4, and the dissolving temperature is 50~70℃.

[0028] Furthermore, in step (5), the total amount of tannin added is 12 to 20 times the mass of germanium in the leaching solution I, the pH of the reduction precipitation of germanium is 2 to 3, the precipitation time of germanium is 15 to 30 minutes, and the precipitation temperature of germanium is 55 to 65°C.

[0029] Furthermore, in step (5), in order to prevent the tannin from being oxidized by air and becoming ineffective during the germanium precipitation process, one or more of Na2SO3, NaHSO3 and Na2S2O3 are added as a reducing agent at a concentration of 0.5 to 1% by mass of the tannin during the germanium precipitation process.

[0030] Furthermore, in step (5), the solution after germanium deposition is returned to the zinc electrowinning system.

[0031] Furthermore, in step (6), the two-stage pyrolysis of tannin germanium includes stage I pyrolysis and stage II pyrolysis, the pyrolysis temperature of stage I pyrolysis is 350-500°C, and the pyrolysis time is 2-4 hours. In this stage, the gaseous products and liquid products generated by the pyrolysis of organic matter in the tannin germanium are separated and collected, and the gaseous products usually include hydrogen, methane, carbon monoxide, etc., and the liquid products usually include tar, oil, phenol, etc. The pyrolysis temperature of stage II pyrolysis is 600-800°C, and the pyrolysis time is 6-8 hours. In this stage, the semi-coke shrinks to form a mixture of coke and germanium concentrate.

[0032] Furthermore, in step (6), the mixture of coke and germanium concentrate is ground to 150-250 mesh, separated by flotation, and the coke is returned to the reduction smelting system.

[0033] Furthermore, in step (7), the concentration of the hydrochloric acid used in the chlorination distillation is 7-10 mol / L, and the liquid-to-solid ratio of the hydrochloric acid to the germanium concentrate is (3-7) mL: 1 g.

[0034] Furthermore, the chlorination distillation adopts an operation mode of premixing before distillation, the premixing liquid-solid ratio is (3~7)mL:1g, the premixing temperature is 60~80℃, the premixing time is 0.5~1h, and then chlorination distillation is carried out, the distillation temperature is 90~120℃, and the distillation time is 2~5h. The crude GeCl4 obtained is also mixed with HCl impurities, the volume ratio of GeCl4 to HCl is 90:10~95:5, and the distillation residue is mainly Ge oxide and PbCl2, etc.

[0035] Furthermore, in step (8), the distillation temperature is 85-95°C.

[0036] Furthermore, in step (8), chlorine gas is introduced into the crude GeCl4, and distillation is carried out by indirect heating with steam until no bubbles are generated to obtain refined GeCl4.

[0037] Compared with the prior art, the present invention has the following advantages and technical effects:

[0038] (1) The present invention separates lead by first reduction, and the lead volatilization rate is 98%. The lead grade in the crude lead after reduction reaches 99.5%. Then, the zinc is fumed and volatilized to separate from the iron. The iron is enriched in the water-quenched slag. After the obtained water-quenched slag is subjected to magnetic separation, the iron slag is mainly composed of Fe3O4, and the iron content reaches 70%. Then, zinc and germanium are leached, and the dispersed tannin is used to precipitate germanium, so that the germanium and zinc are effectively separated under low tannin consumption. The germanium content in the tannin slag reaches 7.30%, and the zinc content in the liquid after germanium precipitation reaches 17.33g / L. The tannin germanium is subjected to two-stage anoxic pyrolysis, which can also realize the efficient enrichment of germanium in the tannin germanium without CO2 discharge. The germanium content in the germanium concentrate reaches 60.72%. The GeO2 obtained by pyrolysis is collected in the form of refined GeCl4 by chlorination distillation, and the Ge(IV) grade reaches 33.87%.

[0039] (2) The present invention can effectively solve the problems existing in the current hard zinc slag treatment, such as the waste of lead, zinc and iron resources, the low germanium content in the slag, the large amount of HCl required for distillation, and the large amount of subsequent waste liquid and waste slag. The direct recovery rate of lead can reach 98%, the direct recovery rate of zinc can reach 96.92%, the resource utilization rate of iron can reach 99%, and the direct recovery rate of germanium can reach 95.04%. BRIEF DESCRIPTION OF THE DRAWINGS

[0040] The drawings constituting a part of the present application are used to provide a further understanding of the present application. The illustrative embodiments and descriptions of the present application are used to explain the present application and do not constitute an improper limitation on the present application. In the drawings:

[0041] Figure 1 This is a process flow chart of a method for cascade resource utilization of germanium, lead, zinc and iron in hard zinc slag in the present invention;

[0042] Figure 2 is the XRD pattern of the hard zinc slag used in Example 1;

[0043] Figure 3 EPMA image of the hard zinc slag used in Example 1;

[0044] Figure 4 The particle size and distribution of the pure tannin solution in Comparative Example 1 are measured;

[0045] Figure 5 The particle size and distribution of the tannin solution in Comparative Example 2 are measured;

[0046] Figure 6 The following are the results of measuring the particle size and distribution of the tannin dispersion solution in Example 3. DETAILED DESCRIPTION

[0047] Various exemplary embodiments of the present invention will now be described in detail. This detailed description should not be considered as limiting the present invention, but should be understood as a more detailed description of certain aspects, features, and embodiments of the present invention.

[0048] It should be understood that the terms described in the present invention are only for describing special embodiments and are not intended to limit the present invention. In addition, for the numerical range in the present invention, it should be understood that each intermediate value between the upper and lower limits of the scope is also specifically disclosed. Each smaller range between the intermediate value in any stated value or stated range and any other stated value or intermediate value in the described range is also included in the present invention. The upper and lower limits of these smaller ranges can be independently included or excluded in the scope.

[0049] Unless otherwise indicated, all technical and scientific terms used herein have the same meanings as those generally understood by those skilled in the art. Although the present invention describes only preferred methods and materials, any methods and materials similar or equivalent to those described herein may also be used in the implementation or testing of the present invention. All documents mentioned in this specification are incorporated by reference to disclose and describe the methods and / or materials associated with the documents. In the event of a conflict with any incorporated document, the content of this specification shall prevail.

[0050] It will be apparent to those skilled in the art that various modifications and variations may be made to the specific embodiments of the present invention description without departing from the scope or spirit of the present invention. Other embodiments derived from the present invention description will be apparent to those skilled in the art. The present invention description and examples are exemplary only.

[0051] The words “include,” “including,” “have,” “contain,” etc. used in this document are open-ended terms, meaning including but not limited to.

[0052] The embodiment of the present invention provides a method for the cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag. Figure 1 , including the following steps:

[0053] (1) Reduction smelting: hard zinc slag is mixed with coke and reduced to obtain crude lead, reduction slag and flue gas;

[0054] (2) Fuming and volatilizing: fumigating and volatilizing the reduced slag to obtain water-quenched slag, zinc oxide fume and flue gas;

[0055] (3) Acid production: The flue gases produced by reduction smelting and fume volatilization are mixed to produce acid;

[0056] (4) Leaching: slurrying the zinc oxide fume to obtain leaching residue I and leaching solution I, and acid leaching the leaching residue I to obtain leaching solution II and lead slag;

[0057] (5) Reduction and precipitation of germanium: tannin is dispersed with a dispersant and then ultrasonically dissolved to obtain a tannin dispersion solution, which is then added to the leaching solution I to reduce and precipitate germanium, thereby obtaining tannin germanium and a germanium precipitation solution;

[0058] (6) Two-stage pyrolysis: the tannin germanium is subjected to two-stage pyrolysis to obtain coke and germanium concentrate, which are separated by flotation;

[0059] (7) Chlorination distillation: Chlorination distillation of the germanium concentrate to obtain crude GeCl4 and distillation residue;

[0060] (8) Distillation: distilling the crude GeCl4 to obtain refined GeCl4.

[0061] In a preferred embodiment of the present invention, in step (1), the germanium content in the hard zinc slag is 2000-5000 g / t, the zinc content is 10-20%, the iron content is 15-25%, the lead content is 20-30%, and the balance is oxygen and impurities. The volatilization rate of lead in the reduction smelting process is 96-98%, and the lead grade of the crude lead obtained after the reduction smelting is 98.5-99.5%.

[0062] In a preferred embodiment of the present invention, the amount of coke added to the hard zinc slag is 11-15wt%, and more preferably, the amount of coke added to the hard zinc slag is 11.32-14.02wt%.

[0063] In a preferred embodiment of the present invention, in step (1), the reduction smelting temperature is 1000-1200° C., and the reduction smelting time is 90-120 min.

[0064] In a preferred embodiment of the present invention, in step (2), the fuming volatilization temperature is 1100-1300° C., and the fuming volatilization time is 60-120 min.

[0065] In a preferred embodiment of the present invention, in step (2), the reducing slag is mainly composed of ZnS, FeS, Fe2O3, Fe3O4 and germanate, and the iron content in the reducing slag is 25.73-35.52%, the zinc content is 17.15-28.41%, the germanium content is 3431-7104 g / t, and the lead content is 0.57-2.06%.

[0066] In a preferred embodiment of the present invention, after the fuming volatilization in step (2), the zinc in the reducing slag is volatilized in the form of metallic zinc, and the germanate is volatilized in the form of GeO, and finally they are enriched in the dust collecting system in the form of ZnO, zinc germanate or GeO2 to obtain zinc oxide fume, in which the zinc content of the zinc oxide fume is 72.27-76.34%, the germanium content is 1.47-1.92%, the iron content is 0.96-2.21%, and the lead content is 0.77-4.42%.

[0067] In a preferred embodiment of the present invention, after the fuming and volatilization in step (2), the water-quenched slag obtained has an iron content of 33.52-45.38%, a zinc content of 0.45-0.83%, a lead content of 0.80-1.37%, and a silicon-calcium compound content of 34.34-55.91%. After the water-quenched slag is subjected to magnetic separation, CaSiO3 tailings and iron slag are obtained. The iron slag is mainly composed of Fe3O4 and has an iron content of 50-70%.

[0068] In a preferred embodiment of the present invention, in step (3), the SO2 content in the reduction smelting flue gas is 1-2%, and the SO2 content in the oxidation volatilization flue gas is 1-2%. The rest is air and CO2, which are combined and then enter the acid production, and sulfuric acid is obtained by two conversions and two absorptions.

[0069] In a preferred embodiment of the present invention, in step (4), the leaching temperature of the slurry leaching is 60-80°C, the leaching time is 10-30 min, the leaching liquid-solid ratio is (4-7) mL: 1 g, and the leaching endpoint pH is 1-3.

[0070] In a preferred embodiment of the present invention, in step (4), the Ge(IV) content in the leaching solution I is 2.37-3.43 g / L, the Zn(II) content is 10.69-17.34 g / L, and the Fe(II) content is 41.31 mg / L-0.28 g / L. The germanium content in the lead slag is 7093-12225 g / t, the zinc content is 19.86-24.26%, the lead content is 12.25-30.40%, and the iron content is 14.44-17.35%.

[0071] In a preferred embodiment of the present invention, in step (4), the leaching temperature of the acid leaching is 80-95°C, the leaching time is 90-180 min, the leaching liquid-solid ratio is (8-10) mL: 1 g, and the leaching endpoint acidity is 20-40 g / L.

[0072] In a preferred embodiment of the present invention, in step (5), the tannin is quinoa tannin, tara tannin, persimmon tannin or chestnut tannin, and the tannin is dispersed and dissolved in a dispersant under ultrasonic action to obtain a tannin dispersion solution, the dispersant is methanol, ethanol, acetone or ethylene glycol, the volume concentration of the dispersant is 0.5‰~2‰, and the concentration of the tannin dispersion solution is 5~20wt%.

[0073] In a preferred embodiment of the present invention, the intensity of the ultrasound is 0.16-0.33 W / cm 2 , the dissolving pH is 2~4, and the dissolving temperature is 50~70℃.

[0074] Tannin mainly exists in a colloidal (molecular aggregated) state. The average molecular weight of tannin is 1250±60. The molecular weight tested in a 1% aqueous solution is 2500±125, which is equivalent to a dimer. The molecular weight is even higher in 10% and 20% aqueous solutions. The self-association of tannin will cause the tannin colloid particles to become unstable and precipitate. The present invention reduces the tannin colloid particle size through the action of ultrasound and a dispersant, reduces the instability and precipitation of tannin particles, and improves the utilization rate of tannin.

[0075] In a preferred embodiment of the present invention, in step (5), the total amount of tannin added is 12 to 20 times the mass of germanium in the leaching solution I, the pH of the reduction precipitation of germanium is 2 to 3, the precipitation time of germanium is 15 to 30 minutes, and the precipitation temperature of germanium is 55 to 65°C.

[0076] In a preferred embodiment of the present invention, in step (5), in order to prevent the tannin from being oxidized by air and becoming ineffective during the germanium precipitation process, one or more of Na2SO3, NaHSO3 and Na2S2O3 in an amount of 0.5 to 1% by mass of the tannin is added during the germanium precipitation as a reducing agent.

[0077] In a preferred embodiment of the present invention, in step (5), the germanium content in the tannin germanium is 4.60-7.30%, the zinc content is 18.78-37.32 g / t, the iron content is 7.26-59.50 g / t, the residual Ge(IV) content in the solution after germanium precipitation is 23.68-34.29 mg / L, the Zn(II) content is 10.68-17.33 g / L, and the Fe(II) content is 40.90 mg / L-0.27 g / L. The solution after germanium precipitation is returned to the zinc electrowinning system.

[0078] In a preferred embodiment of the present invention, in step (6), the two-stage pyrolysis of tannin germanium includes stage I pyrolysis and stage II pyrolysis, the pyrolysis temperature of stage I pyrolysis is 350-500°C, and the pyrolysis time is 2-4 hours. In this stage, the gaseous products and liquid products generated by the pyrolysis of organic matter in tannin germanium are separated and collected, and the gaseous products generally include hydrogen, methane, carbon monoxide, etc., and the liquid products generally include tar, oil, phenol, etc. The pyrolysis temperature of stage II pyrolysis is 600-800°C, and the pyrolysis time is 6-8 hours. In this stage, the semi-coke shrinks to form a mixture of coke and germanium concentrate.

[0079] In a preferred embodiment of the present invention, in step (6), the mixture of coke and germanium concentrate is ground to 150-250 mesh, then separated by flotation, and the coke is returned to the reduction smelting system. The germanium content in the germanium concentrate obtained by flotation is 54.34-60.72%, and the main impurity element is zinc, with a content of 2.48-3.02%.

[0080] In a preferred embodiment of the present invention, in step (7), the concentration of the hydrochloric acid used in the chlorination distillation is 7-10 mol / L, and the liquid-to-solid ratio of the hydrochloric acid to the germanium concentrate is (3-7) mL: 1 g.

[0081] In a preferred embodiment of the present invention, the chlorination distillation adopts an operation mode of premixing before distillation, the premixing liquid-to-solid ratio is (3~7)mL:1g, the premixing temperature is 60~80℃, the premixing time is 0.5~1h, and then chlorination distillation is carried out, the distillation temperature is 90~120℃, and the distillation time is 2~5h. The crude GeCl4 obtained is also mixed with HCl impurities, the volume ratio of GeCl4 to HCl is 90:10~95:5, and the distillation residue is mainly Ge oxide and PbCl2, etc.

[0082] In a preferred embodiment of the present invention, in step (8), the distillation temperature is 85-95°C.

[0083] In a preferred embodiment of the present invention, in step (8), chlorine gas is introduced into the crude GeCl4, and distillation is carried out by indirect heating with steam until no bubbles are generated, thereby obtaining refined GeCl4 with a germanium content of 33.87%.

[0084] The technical solution of the present invention is further illustrated by the following embodiments.

[0085] Example 1

[0086] The main components of the hard zinc slag in this embodiment are shown in Table 1;

[0087] Table 1 Main components of hard zinc slag

[0088] Ge Zn / wt.% Pb / wt.% Fe / wt.% Oxygen and impurities / wt.% 2000g / t 10 30 15 margin

[0089] The XRD pattern of the hard zinc slag used in Example 1 is shown in Figure 2 , EPMA diagram see Figure 3 It can be seen that lead and germanium exist in the same state, in the form of Pb2Ge4O7, Fe and Zn exist in the same form of ZnFe2O4, and there are three typical particles in the hard zinc slag: I-Pb2Ge4O7, II-ZnFe2O4, and III-Pb2Ge4O7 wrapped in ZnFe2O4.

[0090] A method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag, comprising the following steps:

[0091] (1) Reduction smelting: hard zinc slag is mixed with coke and subjected to reduction smelting to obtain crude lead, reduction slag and flue gas, wherein the amount of coke added is 14.02% of the mass of the hard zinc slag, the reduction smelting temperature is 1000°C, the reduction smelting time is 90 min, the lead volatilization rate is 96%, and the lead grade of the crude lead after reduction is 98.5%;

[0092] (2) Fuming and volatilization: The reducing slag obtained after reduction smelting is mainly composed of ZnS, FeS, ZnO, Fe3O4 and germanate. The iron content, zinc content, germanium content and lead content in the reducing slag are 25.73%, 17.15%, 3431g / t and 2.06% respectively. The reducing slag is subjected to fuming and volatilization to obtain water-quenched slag, zinc oxide fume and flue gas. The fuming and volatilization temperature is 1100℃ and the fuming and volatilization time is 60min. After fuming and volatilization, the zinc in the reducing slag is volatilized in the form of metallic zinc and the germanate is volatilized in the form of GeO. The zinc oxide fume contains 72.27% zinc, 1.47% germanium, 2.21% iron and 4.42% lead. The water-quenched slag contains 33.52% Fe, 0.45% Zn, 1.37% Pb and 49.95% silicon-calcium compounds. The water-quenched slag is subjected to magnetic separation to obtain CaSiO3 tailings and iron slag. The iron slag is mainly composed of Fe3O4 and has an iron content of 50%.

[0093] (3) Acid production: The SO2 content of the flue gas produced by reduction smelting is 1%, and the SO2 content of the flue gas produced by fuming volatilization is 1%. The rest is air and CO2. The reduction smelting flue gas and the oxidation volatilization flue gas are combined and then enter the acid production process for two conversions and two absorptions to obtain sulfuric acid, which is further used for subsequent acid leaching.

[0094] (4) Leaching: The zinc oxide dust obtained by fuming and volatilization was slurried to obtain leaching residue I and leaching solution I. The leaching temperature was controlled to be 60°C, the leaching time was 10 min, the leaching liquid-to-solid ratio was 4 mL: 1 g, the leaching endpoint pH was 1, and the Ge(IV) content, Zn(II) content and Fe(II) content in the leaching solution I were 3.43 g / L, 17.34 g / L and 0.28 g / L. The obtained leached residue I is subjected to acid leaching to obtain leachate II and lead slag. The acid leaching is controlled to have a leaching temperature of 80°C, a leaching time of 90min, a leaching liquid-solid ratio of 8mL:1g, and an acidity of 20g / L at the leaching end point. The lead slag obtained by the acid leaching has a germanium content of 7093g / t, a zinc content of 19.86%, a lead content of 30.40%, and an iron content of 14.44%. The lead slag can be reused for reduction smelting, and the leachate II is reused for slurry leaching.

[0095] (5) Reduction and precipitation of germanium: After the tannin is dispersed and dissolved, it is added to the leaching solution I for reduction and precipitation of germanium. The tannin is pentagonal tannin. Methanol dispersant is added during the tannin dissolution process, and ultrasound is used to enhance the dispersion and dissolution. The concentration of the obtained tannin dispersion solution is 5wt%, the dissolution pH is 2, the dissolution temperature is 50°C, the dispersant volume concentration is 0.5‰, and the dissolution ultrasound intensity is 0.16W / cm 2In order to prevent tannin from being oxidized by air and becoming ineffective during the germanium precipitation process, a Na2SO3 reducing agent of 0.5% of the mass of tannin is added during the germanium precipitation. The pH of the germanium precipitation is 2, the germanium precipitation time is 15 minutes, the germanium precipitation temperature is 55°C, and the total amount of tannin added is 12 times the mass of germanium in the leaching solution I. After the reduction precipitation of germanium, tannin germanium and the post-precipitation germanium liquid are obtained. The germanium content in the tannin germanium is 7.30%, the zinc content is 37.32g / t, and the iron content is 59.50g / t. The residual Ge(IV) content in the post-precipitation germanium liquid is 34.29mg / L, the Zn(II) content is 17.33g / L, and the Fe(II) content is 0.27g / L. The germanium precipitation liquid is returned to the zinc electrowinning system;

[0096] (6) Two-stage pyrolysis: tannin germanium is subjected to two-stage high-temperature, anoxic pyrolysis to obtain coke and germanium concentrate. The pyrolysis temperature of stage I is 350°C and the pyrolysis time is 2 hours. In this stage, the gaseous products and liquid products generated by the pyrolysis of organic matter in the tannin germanium are separated and collected. The gaseous products usually include hydrogen, methane, carbon monoxide, etc., and the liquid products usually include tar, oil, phenol, etc.; the pyrolysis temperature of stage II is 600°C and the pyrolysis time is 6 hours. In this stage, the semi-coke shrinks to form a mixture of coke and germanium concentrate; the coke and germanium concentrate are ground to 150 mesh, and then separated by flotation. The separated coke is returned to the reduction smelting system (step (1)). The germanium content in the germanium concentrate obtained by flotation is 54.34%, and the main impurity element is zinc, with a content of 2.78%;

[0097] (7) Chlorination distillation: The germanium concentrate is subjected to chlorination distillation to obtain crude GeCl4, distillation residue and residual liquid. The hydrochloric acid concentration of the chlorination distillation is 7 mol / L. The operation mode of premixing before distillation is adopted. The liquid-solid ratio of the premixing treatment is 3 mL:1 g. The premixing temperature is 60°C and the premixing time is 0.5 h. Subsequently, chlorination distillation is carried out. The distillation temperature is 90°C and the distillation time is 2 h. The crude GeCl4 obtained is also mixed with HCl impurities. The volume ratio of GeCl4 to HCl is 90:10. The distillation residue is mainly Ge oxide and PbCl2, which can be reused for reduction smelting, and the residual liquid is further reused for chlorination distillation.

[0098] (8) Distillation: Chlorine gas is introduced into the crude GeCl4 and distilled using steam for indirect heating. Refined GeCl4 is obtained by distillation at a temperature of 85°C. The distillation is carried out until no bubbles are generated. Refined GeCl4 with a germanium content of 33.87% is obtained. The residual liquid obtained by distillation is returned to the chlorination distillation.

[0099] In Example 1 of the present invention, the direct recovery rate of lead is 96%, the direct recovery rate of zinc is 93.98%, the resource recovery rate of iron is 97.5%, and the direct recovery rate of germanium is 92.07%.

[0100] Example 2

[0101] The main components of the hard zinc slag in this embodiment are shown in Table 2;

[0102] Table 2 Main components of hard zinc slag

[0103] Ge Zn / wt.% Pb / wt.% Fe / wt.% Oxygen and impurities / wt.% 3500g / t 18 23 20 margin

[0104] A method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag, comprising the following steps:

[0105] (1) Reduction smelting: hard zinc slag is mixed with coke and subjected to reduction smelting to obtain crude lead, reduction slag and flue gas, wherein the amount of coke added is 12.12% of the mass of the hard zinc slag, the reduction smelting temperature is 1150°C, the reduction smelting time is 120 min, the lead volatilization rate is 97%, and the lead grade of the reduced crude lead is 99%;

[0106] (2) Fuming and volatilization: The reduction slag obtained after reduction smelting is mainly composed of ZnS, FeS, ZnO, Fe3O4 and germanate. The iron content, zinc content, germanium content and lead content in the reduction slag are 29.85%, 26.87%, 5225g / t and 1.03% respectively. The reduction slag is subjected to fuming and volatilization to obtain water-quenched slag, zinc oxide fume and flue gas. The fuming and volatilization temperature is 1200℃ and the fuming and volatilization time is 90min. After fuming and volatilization, the zinc in the reduction slag is volatilized in the form of metallic zinc and the germanate is volatilized in the form of GeO. The zinc oxide fume contains 75.38% zinc, 1.50% germanium, 1.71% iron and 1.47% lead. The water-quenched slag contains 45.38% Fe, 0.83% Zn, 0.80% Pb and 34.34% silicon-calcium compounds. The water-quenched slag is subjected to magnetic separation to obtain CaSiO3 tailings and iron slag. The iron slag is mainly composed of Fe3O4 with an iron content of 60%.

[0107] (3) Acid production: The SO2 content of the flue gas produced by reduction smelting is 1.5%, and the SO2 content of the flue gas produced by fuming volatilization is 1.5%. The rest is air and CO2. The reduction smelting flue gas and the oxidation volatilization flue gas are combined and then enter the acid production process for two conversions and two absorptions to obtain sulfuric acid, which is further used for subsequent acid leaching.

[0108] (4) Leaching: The zinc oxide dust obtained by fuming and volatilization was slurried to obtain leaching residue I and leaching solution I. The leaching temperature was controlled to be 70°C, the leaching time was 20 min, the leaching liquid-solid ratio was 6 mL: 1 g, the leaching endpoint pH was 2, and the Ge(IV) content, Zn(II) content and Fe(II) content in the leaching solution I were 2.37 g / L, 12.19 g / L and 0.11 g / L. The obtained leached residue I is subjected to acid leaching to obtain leachate II and lead slag. The acid leaching is controlled to have a leaching temperature of 90°C, a leaching time of 135min, a leaching liquid-solid ratio of 9mL:1g, and an acidity of 30g / L. The lead slag obtained by the acid leaching has a germanium content of 7909g / t, a zinc content of 23.92%, a lead content of 15.59%, and an iron content of 17.35%. The lead slag can be recycled for reduction smelting, and the leachate II is recycled for slurry leaching.

[0109] (5) Reduction and precipitation of germanium: After the tannin is dispersed and dissolved, it is added to the leaching solution I for reduction and precipitation of germanium. The tannin is Tara tannin. During the dissolution of the tannin, an ethanol dispersant is added, and ultrasound is used to enhance the dissolution. The concentration of the obtained tannin dispersion solution is 10wt%, the dissolution pH is 3, the dissolution temperature is 60°C, the dispersant volume concentration is 1‰, and the dissolution ultrasound intensity is 0.20W / cm 2 In order to prevent tannin from being oxidized by air and becoming ineffective during the germanium precipitation process, a NaHSO3 reducing agent of 0.7% of the mass of tannin is added during the germanium precipitation. The pH of the germanium precipitation is 2.5, the germanium precipitation time is 25 minutes, the germanium precipitation temperature is 60°C, and the total amount of tannin added is 15 times the mass of germanium in the leaching solution I. After the reduction precipitation of germanium, tannin germanium and the post-precipitation germanium liquid are obtained. The germanium content in the tannin germanium is 5.98%, the zinc content is 31.08g / t, and the iron content is 29.07g / t. The residual Ge(IV) content in the post-precipitation germanium liquid is 23.68mg / L, the Zn(II) content is 12.17g / L, and the Fe(II) content is 0.11g / L. The germanium precipitation liquid is returned to the zinc electrowinning system;

[0110] (6) Two-stage pyrolysis: tannin germanium is subjected to two-stage high-temperature, anoxic pyrolysis to obtain coke and germanium concentrate. The pyrolysis temperature of stage I is 400°C and the pyrolysis time is 3 hours. In this stage, the gaseous products and liquid products generated by the pyrolysis of organic matter in the tannin germanium are separated and collected. The gaseous products usually include hydrogen, methane, carbon monoxide, etc., and the liquid products usually include tar, oil, phenol, etc.; the pyrolysis temperature of stage II is 700°C and the pyrolysis time is 7 hours. In this stage, the semi-coke shrinks to form a mixture of coke and germanium concentrate; the coke and germanium concentrate are ground to 200 mesh, and then separated by flotation. The separated coke is returned to the reduction smelting system (step (1)). The germanium content in the germanium concentrate obtained by flotation is 58.07%, and the main impurity element is zinc, with a content of 3.02%;

[0111] (7) Chlorination distillation: The germanium concentrate is subjected to chlorination distillation to obtain crude GeCl4, distillation residue and residual liquid. The hydrochloric acid concentration of the chlorination distillation is 8.5 mol / L. The operation mode of premixing before distillation is adopted. The liquid-solid ratio of the premixing treatment is 5 mL:1 g, the premixing temperature is 70°C, and the premixing time is 0.75 h. Subsequently, chlorination distillation is carried out. The distillation temperature is 105°C and the distillation time is 3.5 h. The crude GeCl4 obtained is also mixed with HCl impurities. The volume ratio of GeCl4 to HCl is 90:8. The distillation residue is mainly Ge oxide and PbCl2, which can be reused for reduction smelting, and the residual liquid is further reused for chlorination distillation;

[0112] (8) Distillation: Chlorine gas is introduced into the crude GeCl4 and distilled using steam for indirect heating. Refined GeCl4 is obtained by distillation at a temperature of 90°C. The distillation is carried out until no bubbles are generated. Refined GeCl4 with a germanium content of 33.87% is obtained. The residual liquid obtained by distillation is returned to the chlorination distillation.

[0113] In Example 2 of the present invention, the lead direct recovery rate is 97%, the zinc direct recovery rate is 94.96%, the iron resource recovery rate is 98%, and the germanium direct recovery rate is 94.05%.

[0114] Example 3

[0115] The main components of the hard zinc slag in this embodiment are shown in Table 3;

[0116] Table 3 Main components of hard zinc slag

[0117] Ge Zn / wt.% Pb / wt.% Fe / wt.% Oxygen and impurities / wt.% 5000g / t 20 20 25 margin

[0118] A method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag, comprising the following steps:

[0119] (1) Reduction smelting: hard zinc slag is mixed with coke and subjected to reduction smelting to obtain crude lead, reduction slag and flue gas, wherein the amount of coke added is 11.32% of the mass of the hard zinc slag, the reduction smelting temperature is 1200°C, the reduction smelting time is 150 min, the lead volatilization rate is 98%, and the lead grade of the crude lead after reduction is 99.5%;

[0120] (2) Fuming and volatilization: The reduction slag obtained after reduction smelting is mainly composed of ZnS, FeS, ZnO, Fe3O4 and germanate. The iron content, zinc content, germanium content and lead content in the reduction slag are 35.52%, 28.41%, 7104g / t and 0.57% respectively. The reduction slag is subjected to fuming and volatilization to obtain water-quenched slag, zinc oxide fume and flue gas. The fuming and volatilization temperature is 1300℃ and the fuming and volatilization time is 120min. After fuming and volatilization, the zinc in the reduction slag is volatilized in the form of metallic zinc and the germanate is volatilized in the form of GeO volatilize, and finally enrich in the dust collection system in the form of ZnO, zinc germanate or GeO2. The zinc content in the zinc oxide smoke is 76.34%, the germanium content is 1.92%, the iron content is 0.96%, and the lead content is 0.77%. The Fe content in the water-quenched slag is 55.91%, the zinc content is 0.45%, the lead content is 0.45%, and the silicon-calcium compound content is 21.17%. After the water-quenched slag is magnetically separated, CaSiO3 tailings and iron slag are obtained. The iron slag is mainly composed of Fe3O4, and the iron content is 70%;

[0121] (3) Acid production: The SO2 content of the flue gas produced by reduction smelting is 2%, and the SO2 content of the flue gas produced by fuming volatilization is 2%. The rest is air and CO2. The reduction smelting flue gas and the oxidation volatilization flue gas are combined and then enter the acid production process for two conversions and two absorptions to obtain sulfuric acid, which is further used for subsequent acid leaching.

[0122] (4) Leaching: The zinc oxide dust obtained by fuming and volatilization was slurried to obtain leaching residue I and leaching solution I. The leaching temperature was controlled to be 80°C, the leaching time was 30 min, the leaching liquid-solid ratio was 7 mL: 1 g, the leaching endpoint pH was 3, and the Ge(IV) content, Zn(II) content and Fe(II) content in the leaching solution I were 2.64 g / L, 10.69 g / L and 41.31 mg / L. The obtained leached residue I is subjected to acid leaching to obtain leachate II and lead slag. The acid leaching is controlled at a leaching temperature of 95°C, a leaching time of 180 min, a leachate-solid ratio of 10 mL: 1 g, and an acidity of 40 g / L. The lead slag obtained by the acid leaching has a germanium content of 12225 g / t, a zinc content of 24.26%, a lead content of 12.25%, and an iron content of 14.86%. The lead slag can be reused for reduction smelting, and the leachate II is reused for slurry leaching.

[0123] (5) Reduction and precipitation of germanium: After the tannin is dispersed and dissolved, it is added to the leaching solution I for reduction and precipitation of germanium. The tannin is chestnut tannin. Methanol dispersant is added during the tannin dissolution process, and ultrasound is used to enhance the dispersion and dissolution. The concentration of the obtained tannin dispersion solution is 20wt%, the dissolution pH is 4, the dissolution temperature is 70℃, the dispersant volume concentration is 2‰, and the dissolution ultrasound intensity is 0.33W / cm 2In order to prevent tannin from being oxidized by air and becoming ineffective during the germanium precipitation process, a Na2S2O3 reducing agent of 1% of the mass of tannin is added during the germanium precipitation. The pH of the germanium precipitation is 3, the germanium precipitation time is 30 minutes, the germanium precipitation temperature is 65°C, and the total amount of tannin added is 20 times the mass of germanium in the leaching solution I. After the reduction precipitation of germanium, tannin germanium and the post-precipitation germanium liquid are obtained. The germanium content in the tannin germanium is 4.60%, the zinc content is 18.78g / t, and the iron content is 7.26g / t. The residual Ge(IV) content in the post-precipitation germanium liquid is 26.44mg / L, the Zn(II) content is 10.68g / L, and the Fe(II) content is 40.90mg / L. The germanium precipitation liquid is returned to the zinc electrowinning system;

[0124] (6) Two-stage pyrolysis: tannin germanium is subjected to two-stage high-temperature, anoxic pyrolysis to obtain coke and germanium concentrate. The pyrolysis temperature of stage I is 500°C and the pyrolysis time is 4 hours. In this stage, the gaseous products and liquid products generated by the pyrolysis of organic matter in the tannin germanium are separated and collected. The gaseous products usually include hydrogen, methane, carbon monoxide, etc., and the liquid products usually include tar, oil, phenol, etc.; the pyrolysis temperature of stage II is 800°C and the pyrolysis time is 8 hours. In this stage, the semi-coke shrinks to form a mixture of coke and germanium concentrate; the coke and germanium concentrate are ground to 250 mesh, and then separated by flotation. The separated coke is returned to the reduction smelting system (step (1)). The germanium content in the germanium concentrate obtained by flotation is 60.72%, and the main impurity element is zinc, with a content of 2.48%;

[0125] (7) Chlorination distillation: The germanium concentrate is subjected to chlorination distillation to obtain crude GeCl4, distillation residue and residual liquid. The hydrochloric acid concentration of the chlorination distillation is 10 mol / L. The operation mode of premixing before distillation is adopted. The liquid-solid ratio of the premixing treatment is 7 mL:1 g, the premixing temperature is 80°C, and the premixing time is 1 h. Subsequently, chlorination distillation is carried out at a distillation temperature of 120°C and a distillation time of 5 h. The crude GeCl4 obtained is also mixed with HCl impurities. The volume ratio of GeCl4 to HCl is 95:5. The distillation residue is mainly Ge oxide and PbCl2, which can be reused for reduction smelting, and the residual liquid is further reused for chlorination distillation;

[0126] (8) Distillation: Chlorine gas is introduced into the crude GeCl4 and distilled using steam for indirect heating. Refined GeCl4 is obtained by distillation at a temperature of 95°C. The distillation is carried out until no bubbles are generated. Refined GeCl4 with a germanium content of 33.87% is obtained. The residual liquid obtained by distillation is returned to the chlorination distillation.

[0127] In Example 3 of the present invention, the direct recovery rate of lead is 98%, the direct recovery rate of zinc is 96.92%, the resource recovery rate of iron is 99%, and the direct recovery rate of germanium is 95.04%.

[0128] Comparative Example 1

[0129] Same as Example 3, except that the tannin dispersion solution is replaced by a pure tannin solution.

[0130] The preparation method of the pure tannin solution is as follows: dissolving pentagalloch tannin in water to obtain a pure tannin solution with a concentration of 5wt%.

[0131] The remaining steps are the same as in Example 3.

[0132] In comparative example 1 of the present invention, the direct recovery rate of lead is 98%, the direct recovery rate of zinc is 94.98%, the resource recovery rate of iron is 99%, and the direct recovery rate of germanium is 88.32%.

[0133] Comparative Example 2

[0134] The same as Example 3, except that the tannin dispersing solution is replaced by the tannin dissolving solution.

[0135] The preparation method of tannin solution is as follows: chestnut tannin is dissolved in methanol dispersant, and the solution is stirred to obtain a tannin dispersion solution with a concentration of 5wt%, a dissolution pH of 2, a dissolution temperature of 50°C, and a dispersant volume concentration of 0.5‰.

[0136] The remaining steps are the same as in Example 3.

[0137] In Comparative Example 2 of the present invention, the direct recovery rate of lead is 98%, the direct recovery rate of zinc is 95.95%, the resource recovery rate of iron is 99%, and the direct recovery rate of germanium is 93.12%.

[0138] The results of the determination of the particle size and distribution of the pure tannin solution in Comparative Example 1 are shown in Figure 4 The results of the determination of the particle size and distribution of the tannin solution in Comparative Example 2 are shown in Figure 5 The results of the determination of the particle size and distribution of the tannin dispersion solution in Example 3 are shown in Figure 6 ,Depend on Figure 4-Figure 6 It can be seen that the self-association of tannin will increase the tannin colloidal particle size to 3565nm, and the colloidal particle size will be reduced to 947nm after adding dispersant. The colloidal particle size will be reduced to 707nm under the condition of ultrasonic dispersion, thereby reducing the instability and precipitation of tannin particles and improving the utilization rate of tannin. Therefore, under the same tannin amount and germanium content conditions, the precipitation rate of germanium in Example 3 is increased, thereby improving the direct recovery rate of germanium. At the same time, the liquid germanium content is reduced after germanium precipitation, and the amount of neutralizer added in the subsequent electrolytic neutralization and impurity removal step is small, which reduces the neutralization and adsorption loss of zinc, thereby improving the direct recovery rate of zinc.

[0139] The above are only preferred specific implementations of the present application, but the protection scope of the present application is not limited thereto. Any changes or substitutions that can be easily thought of by a person skilled in the art within the technical scope disclosed in the present application should be included in the protection scope of the present application. Therefore, the protection scope of the present application should be based on the protection scope of the claims.

Claims

1. A method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag, characterized in that: The following steps are involved: (1) Reduction smelting: hard zinc slag is mixed with coke and reduced to obtain crude lead, reduction slag and flue gas; (2) Fuming and volatilizing: fumigating and volatilizing the reduced slag to obtain water-quenched slag, zinc oxide fume and flue gas; (3) Acid production: The flue gases produced by reduction smelting and fume volatilization are mixed to produce acid; (4) Leaching: slurrying the zinc oxide fume to obtain leaching residue I and leaching solution I, and acid leaching the leaching residue I to obtain leaching solution II and lead slag; (5) Reducing and precipitating germanium: dispersing tannin with a dispersant and then dissolving it by ultrasonication to obtain a tannin dispersion solution, which is then added to the leaching solution I to reduce and precipitate germanium, thereby obtaining tannin germanium and a germanium precipitation solution. In order to prevent the tannin from being oxidized by air and becoming ineffective during the germanium precipitation process, one or more of Na2SO3, NaHSO3 and Na2S2O3, which is 0.5-1% of the mass of the tannin, is added during the germanium precipitation process as a reducing agent; (6) Two-stage pyrolysis: the tannin germanium is subjected to two-stage pyrolysis to obtain coke and germanium concentrate, which are separated by flotation; (7) Chlorination distillation: Chlorination distillation of the germanium concentrate to obtain crude GeCl4 and distillation residue; (8) Distillation: distilling the crude GeCl4 to obtain refined GeCl4.

2. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: The amount of coke added to the hard zinc slag is 11-15wt%.

3. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (1), the reduction smelting temperature is 1000-1200° C., and the reduction smelting time is 90-120 min.

4. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (2), the fuming volatilization temperature is 1100-1300° C., and the fuming volatilization time is 60-120 min.

5. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (4), the leaching temperature of the slurry leaching is 60-80°C, the leaching time is 10-30 min, the leaching liquid-solid ratio is (4-7) mL: 1 g, and the leaching endpoint pH is 1-3.

6. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (4), the leaching temperature of the acid leaching is 80-95°C, the leaching time is 90-180 min, the leaching liquid-solid ratio is (8-10) mL:1 g, and the leaching endpoint acidity is 20-40 g / L.

7. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (5), the total amount of tannin added is 12 to 20 times the mass of germanium in the leaching solution I, the pH of the reduction precipitation of germanium is 2 to 3, the precipitation time of germanium is 15 to 30 minutes, and the precipitation temperature of germanium is 55 to 65°C.

8. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (6), the two-stage pyrolysis of tannin germanium includes stage I pyrolysis and stage II pyrolysis, the pyrolysis temperature of stage I pyrolysis is 350-500°C, the pyrolysis time is 2-4 hours, and the pyrolysis temperature of stage II pyrolysis is 600-800°C, the pyrolysis time is 6-8 hours.

9. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (7), the concentration of hydrochloric acid used in the chlorination distillation is 7-10 mol / L, and the liquid-to-solid ratio of the hydrochloric acid to the germanium concentrate is (3-7) mL: 1 g.

10. The method for cascade resource recovery of germanium, lead, zinc and iron in hard zinc slag according to claim 1, characterized in that: In step (8), the distillation temperature is 85-95°C.

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