Method for recovering zinc and indium in low-quality zinc hypoxide

Abstract

A method for recovering zinc and indium in low-quality secondary zinc oxide comprises the steps that S1, part of secondary zinc oxide is subjected to alkali washing, and alkali washing residues and alkali washing liquid are obtained; and carrying out ammonia leaching on the residual secondary zinc oxide to obtain ammonia leaching liquid and ammonia leaching residues. S2, performing low-acid leaching on the alkali washing residues to obtain supernate and neutral leaching underflow; and washing the ammonia leaching residues to obtain washing liquid and washing residues. And S3, the neutral leaching underflow and the washing residues are mixed and subjected to acid leaching to obtain acid leaching liquid and acid leaching residues, and the acid leaching liquid is subjected to extraction and reverse extraction and then is subjected to zinc powder replacement to obtain sponge indium. S4, the ammonia leaching liquid is subjected to heavy removal and ammonia leaching side zinc electrodeposition, and waste electrolyte I and metal zinc I are obtained; and removing impurities from the supernate, and then carrying out alkali washing and side zinc electrodeposition to obtain waste electrolyte II and metal zinc II. According to the method disclosed by the invention, the secondary zinc oxide is divided into two parts, and the two parts are respectively subjected to alkali washing treatment and chlorine ammonia leaching, so that the advantages of the two treatment methods are effectively combined, the coordination capability of zinc, ammonia and chlorine is fully utilized, the effective separation of zinc and indium is realized, and the efficient recovery of zinc and indium is realized.

Description

Technical Field

[0001] The present invention relates to a method for recovering zinc and indium, in particular to a method for recovering zinc and indium from low-quality secondary zinc oxide, and belongs to the technical field of solid waste recovery. Background Art

[0002] Zinc-containing solid waste produced by industries such as steel and nonferrous metals produces a large amount of low-quality secondary zinc oxide after recycling. Zinc and indium have a high recycling value, but due to the high impurity content, especially high concentrations of chlorine, direct return to the zinc smelting system requires multiple stages of purification and impurity removal, which is a long process. Therefore, the current recycling of secondary zinc oxide is mainly used to recover low-priced products such as zinc sulfate and lithopone. To recover elemental zinc from secondary zinc oxide, it is usually necessary to go through alkaline washing-cascade leaching-multi-stage purification-electrodeposition. The purification process is long, especially with high requirements for impurities such as chlorine, resulting in high purification costs. Since zinc and ammonia have strong coordination abilities, the ammonia process can be used to recover zinc from low-quality zinc. However, since secondary zinc oxide also contains high-value indium, it is difficult to recover in the ammonia leaching process, making it difficult to utilize zinc and indium in zinc-containing solid waste.

[0003] Chinese patent CN102286759A discloses a method for producing electrolytic zinc from high-fluorine and chlorine subzinc oxide powder. The method uses industrial sulfuric acid and / or electrolytic residual liquid to directly leach the subzinc oxide powder, and the leachate is subjected to multi-stage purification to obtain a pure zinc sulfate solution, which is then electrolytically obtained to obtain zinc ingots. This method can achieve high-value recovery of zinc in subzinc oxide, but does not involve resource utilization of indium, and the process of iron removal and dechlorination is relatively long. Chinese patent CN117625989A discloses a method for treating subzinc oxide powder. The process includes multi-stage acid leaching and multi-stage purification and separation to recover valuable metals such as zinc, lead, and copper in the subzinc oxide, but the method does not involve the recovery of indium. Chinese patent CN115927873A discloses a method for preparing zinc oxide by wet treatment of subzinc oxide dust. The method uses ammonium chloride to leach the subzinc oxide, and uses zinc powder to perform multi-stage purification on the leachate, and finally adjusts the pH and crystallizes to obtain a zinc oxide product. This method utilizes the chloramine method to recover zinc from zinc suboxide to prepare zinc oxide products. The resulting product has a low value and cannot achieve the recovery of indium from zinc suboxide.

[0004] Because zinc suboxide contains valuable components such as zinc and indium, as well as high concentrations of impurities such as chlorine, conventional acid leaching requires alkaline washing to remove chlorine from the zinc suboxide, followed by multiple precipitation purification steps to obtain a relatively pure zinc sulfate solution before electrolysis. This process is lengthy, with complex iron and chlorine removal processes and high energy consumption. While the chloramine method can effectively simplify the purification process, it cannot recover the indium in the zinc suboxide, and the potassium, sodium, and other salts in the zinc suboxide cannot be recovered.

[0005] Therefore, there is an urgent need for a process that can simultaneously recover zinc, indium, and salts such as potassium and sodium from zinc suboxide. Summary of the Invention

[0006] To address the existing difficulty in recovering the higher-value zinc and indium in secondary zinc oxide, this paper proposes a method for recovering zinc and indium from low-quality secondary zinc oxide. Low-quality secondary zinc oxide (generally referring to zinc-containing materials with a grade of no more than 30%) is subjected to alkaline washing and chloramine leaching. Through a series of optimized steps, this method achieves efficient recovery of zinc, indium, and salts such as potassium and sodium. This method not only simplifies the purification process in traditional recovery processes, reduces energy consumption, and improves resource recovery efficiency, but also effectively recovers indium, a high-value component, offering both high economic value and environmental significance.

[0007] According to an embodiment of the present invention, a method for recovering zinc and indium from low-quality secondary zinc oxide is provided.

[0008] A method for recovering zinc and indium from low-quality secondary zinc oxide, the method comprising the following steps: S1. Alkali-washing a portion of the low-quality zinc oxide to obtain an alkaline washing residue and an alkaline washing liquid. Ammonia-leaching the remaining portion of the low-quality zinc oxide to obtain an ammonia leaching liquid and an ammonia leaching residue.

[0009] S2, the alkaline washed residue obtained in S1 is subjected to low-acid leaching to obtain a supernatant and a mid-leaching bottom flow. The ammonia leached residue obtained in S1 is subjected to water washing to obtain a water washing liquid and a water washed residue.

[0010] S3. The intermediate leaching bottom flow obtained in S2 is mixed with the water-washed slag and subjected to acid leaching to obtain an acid leaching solution and an acid leaching slag. The obtained acid leaching solution is further subjected to extraction and stripping to obtain an indium-containing liquid. The indium-containing liquid is then replaced by zinc powder to obtain sponge indium.

[0011] S4, the ammonia leaching solution obtained in S1 is subjected to weight removal and zinc electrowinning treatment on the ammonia leaching side in sequence to obtain spent electrolyte I and metallic zinc I. The supernatant obtained in S2 is first subjected to impurity removal treatment and then to zinc electrowinning treatment on the alkali washing side to obtain spent electrolyte II and metallic zinc II.

[0012] Preferably, the method further comprises: S5, the alkali wash obtained in step S1 is subjected to evaporation and crystallization to obtain potassium salt, sodium salt and dilute ammonia. Preferably, the spent electrolyte I obtained in S4 is mixed with the alkali wash and then subjected to evaporation and crystallization together.

[0013] S6: The water washing solution obtained in S2 and the spent electrolyte I obtained in S4 are recycled as ammonia leaching replenishment solution. The spent electrolyte II obtained in S4 is recycled as low-acid leaching replenishment solution.

[0014] S7. Return the acid leaching residue obtained in S3 to the rotary kiln for disposal.

[0015] Preferably, in step S1, the mass of the low-quality zinc oxide subjected to alkali washing accounts for 50% to 85% of the total mass of the low-quality zinc oxide, preferably 60% to 80%, and more preferably 65% ​​to 75%.

[0016] Preferably, in step S1, the liquid-to-solid ratio during alkali washing is 1.5-8 mL / g, preferably 2-6 mL / g. The temperature during alkali washing is 50-90°C, preferably 60-80°C. The pH during alkali washing is 7.5-11, preferably 8-10. The alkali used in alkali washing is one or more of sodium hydroxide, strong potassium hydroxide, and calcium hydroxide.

[0017] Preferably, in step S1, the leaching liquid used in the ammonia leaching treatment is an ammonium chloride solution.

[0018] Preferably, the concentration of the ammonium chloride solution is 3-8 mol / L, preferably 4-6 mol / L. The liquid-to-solid ratio of the ammonium chloride solution to zinc oxide is 7-12 mL / g, preferably 8-10 mL / g. The temperature of the ammonia immersion treatment is 30-50°C, preferably 35-45°C. The ammonia immersion treatment time is 2-4 hours, preferably 2.5-3.5 hours.

[0019] Preferably, in step S2, the leaching solution used for low-acid leaching is dilute sulfuric acid. The endpoint pH of the low-acid leaching is 4-6.5, preferably 5-6. The liquid-to-solid ratio of the low-acid leaching is 4-8 mL / g, preferably 5-7 mL / g. The temperature of the low-acid leaching is 40-80°C, preferably 50-70°C.

[0020] Preferably, in step S2, the liquid-to-solid ratio of water washing is 1.5-5 mL / g, preferably 2-4 mL / g.

[0021] Preferably, the acidic reagent used in the acid leaching in step S3 is sulfuric acid.

[0022] Preferably, the concentration of sulfuric acid is 110-200 g / L, more preferably 130-160 g / L.

[0023] Preferably, the liquid-to-solid ratio of the acidic reagent to the sum of the mass of the ammonia leaching residue after water washing and the intermediate leaching bottom flow is 3-8 mL / g, preferably 4-6 mL / g.

[0024] Preferably, the acid leaching time is 0.3 to 4 hours, preferably 0.5 to 2 hours.

[0025] Preferably, the extraction agent used is P204 (CAS Reg. No. 298-07-7) or P507 (CAS Reg. No. 238-865-3). The stripping agent used in the stripping is sulfuric acid or hydrochloric acid, preferably hydrochloric acid. The hydrogen ion concentration in the stripping agent is 1 to 3 mol / L. Preferably, the raffinate obtained from the extraction is returned to the low-acid leaching process described in step S2.

[0026] Preferably, the weight removal agent used in the weight removal in step S4 is zinc powder, and the solid-liquid ratio of the weight removal agent to the ammonia immersion solution is 1-5 g / L.

[0027] Preferably, the impurity removal treatment in step S4 is to perform iron removal treatment, chlorine removal treatment and heavy removal treatment on the supernatant in sequence.

[0028] Preferably, the iron removal agent used in the iron removal treatment is hydrogen peroxide, and the molar ratio of the iron removal agent to the iron content in the supernatant is 0.5-1.

[0029] Preferably, the dechlorination agent used in the dechlorination treatment is a cuprous compound (including but not limited to cuprous sulfate, cuprous oxide, etc.), and the molar ratio of the dechlorination agent to the chlorine content in the supernatant is 1-3.

[0030] Preferably, the weight-removing agent used in the weight-removing treatment is zinc powder, and the solid-liquid ratio of the weight-removing agent to the supernatant is 1-5 g / L.

[0031] Preferably, the endpoint acidity of the zinc electrowinning treatment on the ammonia immersion side and the zinc electrowinning treatment on the alkali washing side in step S4 is independently 120-250 g / L, preferably 150-200 g / L.

[0032] Preferably, the spent electrolyte I in step S6 is mixed with the dilute ammonia solution obtained in step S5 before use as an ammonia soaking solution. Preferably, the ammonia concentration of the mixed solution obtained after the spent electrolyte I and the dilute ammonia solution is 3 to 8 mol / L, preferably 4 to 6 mol / L.

[0033] In the present invention, low-quality secondary zinc oxide is divided into two parts and subjected to different treatments. The first part of the secondary zinc oxide is first subjected to an alkali wash to remove chlorine and some impurities. After the alkali wash, the purity of the secondary zinc oxide is improved, laying the foundation for subsequent efficient recovery. The second part of the secondary zinc oxide is treated using a chlorine-ammonia leaching method to effectively separate indium and zinc. At the same time, the intermediate leaching bottom flow obtained from the alkali wash is mixed with the ammonia leaching residue to improve the recovery rate of indium.

[0034] In the present invention, zinc has a strong coordination ability with ammonia and chlorine to form Zn(NH3) n n-2 and ZnCl n n-2Due to the presence of ammonia, the Cl2 produced during the electrolytic process will undergo a redox reaction with NH3 to generate N2 and Cl - Therefore, the secondary zinc oxide can be directly put into the leaching system without alkali washing and chlorination through the ammonia leaching system, and the chlorine in the secondary zinc oxide also helps the leaching of zinc.

[0035] In the present invention, in order to solve the problem of indium recovery during ammonia leaching, secondary zinc oxide is treated by acid leaching and ammonia leaching in a coordinated manner. The indium-containing slag produced by ammonia leaching is leached and recovered by the acid leaching process, and the insoluble zinc in the slag is recovered simultaneously. In order to avoid the introduction of indium-containing slag from ammonia leaching causing the acid leaching liquid-solid ratio to change and thus reduce the leaching rate, the percentage of the mass of low-quality secondary zinc oxide after alkali washing to the total mass of low-quality secondary zinc oxide is preferably controlled at 50% to 85% (for example, 50%, 51%, 53%, 55%, 57%, 60%, 52%, 64%, 65%, 66%, 67%, 68%, 69%, 70%, 71%, 72%, 73%, 74%, 75%, 78%, 80%, 82%, 84%, 85%) according to the slag rate and pH of the ammonia leaching slag. At this time, the comprehensive recovery cost of secondary zinc oxide can be effectively reduced. If the ammonia leaching ratio is too high, the comprehensive recovery rate of zinc will be reduced, and if the alkali washing ratio is too high, the comprehensive operating cost will be higher.

[0036] In the present invention, by further finely controlling the operating parameters of each step, such as the alkaline washing conditions, leachate composition, leaching endpoint pH, leachate-solid ratio and other parameters, the effective removal of chlorine is ensured, while avoiding excessive loss of valuable elements such as zinc and indium, thereby achieving a large amount of chlorine dissolution in the secondary zinc oxide, and effectively reducing the load of subsequent deep chlorine removal. The secondary zinc oxide after alkali washing will have iron and impurity metal elements dissolved in the leaching process, and needs to go through multiple stages of precipitation purification before entering the electrolytic process. After electrolytic deposition, the pH of the electrolyte is significantly reduced due to the precipitation of O2 at the anode, and the obtained acidic waste electrolyte can be returned to the leaching process.

[0037] In this invention, recovery efficiency and product quality are ensured by regulating temperature and time, impurity removal, zinc electrowinning on the ammonia immersion side, and the endpoint acidity of zinc electrowinning on the alkaline wash side. Furthermore, this method further improves the economic and environmental friendliness of the overall process by recycling resources such as spent electrolyte and dilute ammonia water, and the entire process produces no wastewater or waste residue.

[0038] Compared with the prior art, the present invention has the following beneficial effects: 1. The present invention provides a method for recovering zinc and indium from low-quality secondary zinc oxide. By dividing the secondary zinc oxide into two parts and performing alkali washing and chlorammonia leaching respectively, the advantages of the two treatment methods are effectively combined, the coordination capacity of zinc, ammonia and chlorine is fully utilized, and the effective separation of zinc and indium is achieved, thereby realizing the efficient recovery of zinc and indium.

[0039] 2. The present invention provides a method for recovering zinc and indium from low-quality secondary zinc oxide. By carefully controlling the operating parameters of each step, recovery efficiency and product quality are ensured. Furthermore, by recycling resources such as waste electrolyte and dilute ammonia, the overall process is further improved in terms of economic efficiency and environmental friendliness. The entire process does not generate wastewater or waste residue. BRIEF DESCRIPTION OF THE DRAWINGS

[0040] Figure 1 The present invention provides a schematic flow chart of a method for recovering zinc and indium from low-quality secondary zinc oxide. DETAILED DESCRIPTION

[0041] The technical solutions of the present invention are illustrated below with examples, and the scope of protection requested by the present invention includes but is not limited to the following embodiments. Example 1

[0042] A method for recovering zinc and indium from low-quality secondary zinc oxide, the method comprising: S1. 700 g of low-quality zinc oxide was alkali-washed with 2100 mL of NaOH solution having a pH of 9 to obtain about 592.54 g of alkali-washed residue and 2100 mL of alkali-washing liquid; 300 g of low-quality zinc oxide was ammonia-leached with 2700 mL of 5 mol / L ammonium chloride solution to obtain about 2700 mL of ammonia-leaching liquid and 85.73 g of ammonia-leached residue; S2. 592.54 g of the alkaline-washed residue obtained in S1 was subjected to low-acid leaching using 3555 mL of dilute sulfuric acid, with the leaching endpoint pH being 6 and the leaching temperature being 60°C, to obtain approximately 2600 mL of supernatant and 900 mL of a mid-leached bottom stream with a solid content of approximately 30% (i.e., the mid-leached bottom stream residue was approximately 900 mL); 85.73 g of the ammonia-leached residue obtained in S1 was washed with water at an ammonia-liquid-solid ratio of 3:1 to obtain approximately 250 mL of water washing liquid and 72.65 g of water-washed residue; S3. After mixing 900 mL of the bottom leaching stream obtained in S2 with 72.65 g of the water-washed slag, acid leaching was performed using 1000 mL of 150 g / L sulfuric acid for 1 h to obtain about 1800 mL of acid leaching solution and 33.02 g of acid leaching residue. The obtained 1800 mL of acid leaching solution was then extracted with P204 and stripped with 1.5 mol / L hydrochloric acid to obtain an indium-containing stripping solution. About 0.37 g of sponge indium was obtained by zinc powder replacement.

[0043] S4, the 2700mL ammonia immersion solution obtained in S1 was subjected to a weight removal treatment using 11.4g of zinc powder, and then an ammonia immersion side zinc electrowinning treatment was performed, with an endpoint acidity of 180g / L, to obtain approximately 2500mL of spent electrolyte I and 103.72g of metallic zinc I; the 2200mL supernatant obtained in S2 was sequentially subjected to an iron removal treatment (iron removal reagent is 30% hydrogen peroxide, with an amount of 35mL), a chlorine removal treatment (chlorine removal reagent is cuprous oxide, with an amount of 20g), and a weight removal treatment (weight removal reagent is zinc powder, with an amount of 10.5g), and then an alkali washing side zinc electrowinning treatment was performed, with an endpoint acidity of 180g / L, to obtain approximately 2100mL of spent electrolyte II and 247.36g of metallic zinc II; S5. Perform evaporation and crystallization treatment on 2100 mL of the alkaline washing solution obtained in step S1 and 1300 mL of the spent electrolyte I obtained in S4 to obtain about 102.48 g of potassium salt, 150.34 g of sodium salt and 1160 mL of dilute ammonia water.

[0044] S6. Recycle 250 mL of the water wash solution obtained in S2 and 1200 mL of the spent electrolyte I obtained in S4 as ammonia leaching replenishment solution, wherein the spent electrolyte I is mixed with the dilute ammonia water obtained in step S5 before use, and the mixing volume ratio is 3.5:1; recycle 2100 mL of the spent electrolyte II obtained in S4 as low-acid leaching replenishment solution; and / or S7. Return 33.02 g of the acid leaching residue obtained in S3 to the rotary kiln for disposal.

[0045] Example 1 was repeated, except that the parameters during the experiment were changed to conduct a single variable experiment. Example 2

[0046] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 65%. Example 3

[0047] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 75%. Example 4

[0048] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 60%. Example 5

[0049] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 80%. Example 6

[0050] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 50%. Example 7

[0051] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 45%. Example 8

[0052] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 90%. Example 9

[0053] Example 1 was repeated, except that the liquid-to-solid ratio of the ammonia leaching in step S1 was changed to 8.0 ml / g. Example 10

[0054] Example 1 was repeated, except that the liquid-to-solid ratio of the ammonia leaching in step S1 was changed to 10.0 ml / g. Example 11

[0055] Example 1 was repeated except that the liquid-to-solid ratio of the ammonia leaching in step S1 was changed to 7.0 ml / g. Example 12

[0056] Example 1 was repeated, except that the liquid-to-solid ratio of the ammonia leaching in step S1 was changed to 12.0 ml / g. Example 13

[0057] Example 1 was repeated, except that the liquid-to-solid ratio of the ammonia leaching in step S1 was changed to 6.0 ml / g. Example 14

[0058] Example 1 was repeated, except that the liquid-to-solid ratio of the ammonia leaching in step S1 was changed to 13.0 ml / g. Example 15

[0059] Example 1 was repeated except that the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was changed to 4.0 mol / L. Example 16

[0060] Example 1 was repeated except that the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was changed to 6.0 mol / L. Example 17

[0061] Example 1 was repeated except that the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was changed to 3.0 mol / L. Example 18

[0062] Example 1 was repeated except that the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was changed to 8.0 mol / L. Example 19

[0063] Example 1 was repeated except that the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was changed to 2.0 mol / L. Example 20

[0064] Example 1 was repeated except that the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was changed to 10.0 mol / L. Example 21

[0065] Example 1 was repeated except that the liquid-to-solid ratio of the low-acid leaching in step S2 was changed to 5.0 mL / g. Example 22

[0066] Example 1 was repeated except that the liquid-to-solid ratio of the low-acid leaching in step S2 was changed to 7.0 mL / g. Example 23

[0067] Example 1 was repeated except that the liquid-to-solid ratio of the low-acid leaching in step S2 was changed to 4.0 mL / g. Example 24

[0068] Example 1 was repeated except that the liquid-to-solid ratio of the low-acid leaching in step S2 was changed to 8.0 mL / g. Example 25

[0069] Example 1 was repeated except that the liquid-to-solid ratio of the low-acid leaching in step S2 was changed to 3.0 mL / g. Example 26

[0070] Example 1 was repeated except that the liquid-to-solid ratio of the low-acid leaching in step S2 was changed to 9.0 mL / g. Example 27

[0071] Example 1 was repeated except that the concentration of sulfuric acid in the acid leaching in step S3 was changed to 130 g / L. Example 28

[0072] Example 1 was repeated except that the concentration of sulfuric acid in the acid leaching in step S3 was changed to 160 g / L. Example 29

[0073] Example 1 was repeated except that the concentration of sulfuric acid in the acid leaching in step S3 was changed to 110 g / L. Example 30

[0074] Example 1 was repeated except that the concentration of sulfuric acid in the acid leaching in step S3 was changed to 200 g / L. Example 31

[0075] Example 1 was repeated except that the concentration of sulfuric acid in the acid leaching in step S3 was changed to 100 g / L. Example 32

[0076] Example 1 was repeated except that the concentration of sulfuric acid in the acid leaching in step S3 was changed to 220 g / L. Example 33

[0077] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 71%, the liquid-to-solid ratio of the ammonia leaching in step S1 was 9.1 ml / g, the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was 5.2 mol / L, the liquid-to-solid ratio of the low-acid leaching in step S2 was 5.7 mL / g, and the sulfuric acid concentration in the acid leaching in step S3 was 145 g / L. Example 34

[0078] Example 1 was repeated, except that the mass proportion of the low-grade secondary zinc oxide in the alkali washing in step S1 was changed to 68%, the liquid-to-solid ratio of the ammonia leaching in step S1 was 8.7 ml / g, the concentration of the ammonium chloride solution in the ammonia leaching in step S1 was 4.9 mol / L, the liquid-to-solid ratio of the low-acid leaching in step S2 was 6.2 mL / g, and the sulfuric acid concentration in the acid leaching in step S3 was 153 g / L.

[0079] The products obtained from the above experiment were tested, and the results are shown in Table 1.

[0080] Table 1 Test results of products in various examples

[0081] According to the above experimental results, the method for recovering zinc and indium from low-quality secondary zinc oxide provided by the present invention realizes the effective separation of zinc and indium, realizes the efficient recovery of zinc and indium, and ensures the recovery rate and purity by strictly limiting the process parameters of each step.

Claims

1. A method for recovering zinc and indium from low-quality secondary zinc oxide, characterized by: The method comprises the following steps: S1, alkali washing a portion of low-quality zinc oxide to obtain alkali washing residue and alkali washing liquid; and ammonia leaching the remaining portion of low-quality zinc oxide to obtain ammonia leaching liquid and ammonia leaching residue; S2, subjecting the alkaline washed residue obtained in S1 to low-acid leaching to obtain a supernatant and a mid-leaching bottom flow; and subjecting the ammonia leached residue obtained in S1 to water washing to obtain a water washing liquid and a water washing residue; S3, mixing the intermediate leaching bottom flow obtained in S2 with the water-washed slag and subjecting it to acid leaching to obtain an acid leaching solution and an acid leaching residue, extracting and stripping the obtained acid leaching solution to obtain an indium-containing solution, and replacing the indium-containing solution with zinc powder to obtain sponge indium; S4, subjecting the ammoniacal leaching solution obtained in S1 to weight removal and zinc electrowinning on the ammoniacal leaching side to obtain a waste electrolyte I and metallic zinc I; subjecting the supernatant obtained in S2 to an impurity removal treatment and then to an alkali-washing side zinc electrowinning treatment to obtain a waste electrolyte II and metallic zinc II; Preferably, the method further comprises: S5, performing evaporation and crystallization treatment on the alkaline washing solution obtained in step S1 to obtain potassium salt, sodium salt and dilute ammonia water; preferably, the spent electrolyte I obtained in S4 is mixed with the alkaline washing solution and then subjected to evaporation and crystallization treatment together; and / or S6, recycling the water wash solution obtained in S2 and the spent electrolyte I obtained in S4 as ammonia leaching replenishment solution; recycling the spent electrolyte II obtained in S4 as low-acid leaching replenishment solution; and / or S7. Return the acid leaching residue obtained in S3 to the rotary kiln for disposal.

2. The method according to claim 1, wherein: In step S1, the mass of the low-quality zinc oxide subjected to alkali washing accounts for 50% to 85% of the total mass of the low-quality zinc oxide, preferably 60% to 80%, and more preferably 65% ​​to 75%.

3. The method according to claim 1 or 2, characterized in that: In step S1, the liquid-to-solid ratio during alkali washing is 1.5-8 mL / g, preferably 2-6 mL / g; the temperature during alkali washing is 50-90° C., preferably 60-80° C.; and the pH during alkali washing is 7.5-11, preferably 8-10.

4. The method according to any one of claims 1 to 3, characterized in that: In step S1, the leaching liquid used in the ammonia leaching treatment is an ammonium chloride solution; Preferably, the concentration of the ammonium chloride solution is 3-8 mol / L, preferably 4-6 mol / L; the liquid-to-solid ratio of the ammonium chloride solution to the secondary zinc oxide is 7-12 mL / g, preferably 8-10 mL / g; the temperature of the ammonia immersion treatment is 30-50° C., preferably 35-45° C.; and the time of the ammonia immersion treatment is 2-4 h, preferably 2.5-3.5 h.

5. The method according to any one of claims 1 to 4, characterized in that: In step S2, the leaching solution used in the low-acid leaching is dilute sulfuric acid; the endpoint pH of the low-acid leaching is 4-6.5, preferably 5-6; the liquid-to-solid ratio of the low-acid leaching is 4-8 mL / g, preferably 5-7 mL / g; the temperature of the low-acid leaching is 40-80° C., preferably 50-70° C.; and / or In step S2, the liquid-to-solid ratio of water washing is 1.5-5 mL / g, preferably 2-4 mL / g.

6. The method according to any one of claims 1 to 5, characterized in that: The acidic reagent for the acid leaching in step S3 is sulfuric acid; The concentration of sulfuric acid is 110-200 g / L, preferably 130-160 g / L; The liquid-to-solid ratio of the acidic reagent to the sum of the mass of the ammonia leaching residue after water washing and the intermediate leaching bottom flow is 3-8 mL / g, preferably 4-6 mL / g; The acid leaching time is 0.3~4h, preferably 0.5~2h; Preferably, the extraction agent used is P204 or P507; the stripping agent used in the stripping is sulfuric acid or hydrochloric acid, preferably hydrochloric acid; the hydrogen ion concentration in the stripping agent is 1-3 mol / L; preferably, the raffinate obtained by the extraction is returned to the low-acid leaching described in step S2.

7. The method according to any one of claims 1 to 6, characterized in that: The weight removal agent used in step S4 is zinc powder, and the solid-liquid ratio of the weight removal agent to the ammonia immersion solution is 1-5 g / L.

8. The method according to any one of claims 1 to 7, characterized in that: The impurity removal treatment in step S4 is to sequentially perform iron removal, chlorine removal, and heavy removal treatment on the supernatant; Preferably, the iron removal agent used in the iron removal treatment is hydrogen peroxide, and the molar ratio of the iron removal agent to the iron content in the supernatant is 0.5 to 1; and / or The dechlorination agent used in the dechlorination treatment is a cuprous compound, and the molar ratio of the dechlorination agent to the chlorine content in the supernatant is 1 to 3; and / or The weight removal agent used in the weight removal treatment is zinc powder, and the solid-liquid ratio of the weight removal agent to the supernatant is 1-5 g / L.

9. The method according to any one of claims 1 to 8, characterized in that: The endpoint acidity of the zinc electrowinning treatment on the ammonia immersion side and the zinc electrowinning treatment on the alkali washing side described in step S4 is independently 120-250 g / L, preferably 150-200 g / L.

10. The method according to claim 9, characterized in that: The spent electrolyte I in step S6 is mixed with the dilute ammonia solution obtained in step S5 before use as an ammonia leaching solution; preferably, the ammonia concentration of the mixed solution obtained after the spent electrolyte I and the dilute ammonia solution is 3~8mol / L, preferably 4~6mol / L.

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