Process for recovering valuable metals from high-chlorine zinc slag

Abstract

The invention relates to a process for recovering valuable metals from high-chlorine zinc slag. The process comprises the following steps: roasting the high-chlorine zinc slag to obtain low-chlorine zinc slag and zinc-containing smoke dust; the method comprises the following steps: mixing low-chlorine zinc slag and zinc-containing smoke dust, performing first-stage leaching, and separating to obtain filtrate A and filter residue A; performing zinc extraction, zinc reverse extraction and electrolysis on the filtrate A to obtain cathode zinc; a sulfuric acid solution is added into the filter residues A for second-stage leaching, and second-stage acid leaching liquid is obtained through separation; adding iron powder into the second-stage pickle liquor to carry out replacement reaction, and separating to obtain filtrate B; performing indium extraction on the filtrate B to obtain a first organic phase and a first water phase; the first organic phase is subjected to indium reverse extraction and zinc-indium replacement reaction, and sponge indium is obtained; the first water phase is subjected to impurity removal, concentration and crystallization to obtain zinc sulfate. According to the method disclosed by the invention, the high-chlorine zinc slag is roasted to volatilize most of chlorine, and the filtrate A and the filter residue A are sequentially treated after one-stage leaching, so that the comprehensive resource utilization of the high-chlorine zinc slag is realized, high-purity zinc sulfate and sponge indium are obtained, and the recovery rate of zinc and indium is high.

Description

TECHNICAL FIELD

[0001] The present application relates to the technical field of metallurgy, and in particular to a process for recovering valuable metals from high-chlorine zinc slag. BACKGROUND

[0002] In recent years, the utilization of metal ore resources has been increasing year by year. In order to alleviate the shortage of primary ore resources, more and more attention has been paid to the research and development of secondary resources.

[0003] High-chlorine zinc slag is a high-risk pollutant and also an important secondary resource. It is mainly derived from gas ash (flue dust) in the process of steel smelting and lead-zinc smelting. The main component is zinc oxide, and it also contains a large amount of valuable metals such as chlorine, lead, copper, germanium, and indium. It has high recycling value, but faces many difficulties such as low comprehensive utilization rate and difficult recovery. The chlorine in high-chlorine zinc slag is very harmful, and usually needs to be roasted to remove chlorine. However, after roasting, the phase will change, causing zinc and indium germanium to be wrapped together to form a difficult-to-dissolve substance, thereby affecting the leaching rate of zinc and the recovery of indium.

[0004] It is very important to realize the recovery of main metal components in high-chlorine zinc slag by adopting a reasonable and simple process to reduce environmental pollution and improve resource utilization level. SUMMARY

[0005] The purpose of the present application is to overcome the above technical deficiencies and provide a process for recovering valuable metals from high-chlorine zinc slag, which solves the technical problem of low recovery rate and product purity caused by the difficulty in recovering the main components zinc and indium in high-chlorine zinc slag in the prior art.

[0006] To achieve the above technical purpose, the technical solution provided by the present application is as follows: In a first aspect, the present application provides a process for recovering valuable metals from high-chlorine zinc slag, comprising the following steps: S1, roasting high-chlorine zinc slag to obtain low-chlorine zinc slag and zinc-containing dust; S2, mixing the low-chlorine zinc slag and the zinc-containing dust and then subjecting them to one-stage leaching to separate a filtrate A and a filter residue A; subjecting the filtrate A to zinc extraction, zinc back extraction, and electrolysis to obtain cathode zinc; S3, adding a sulfuric acid solution to the filter residue A to perform two-stage leaching, and separating a two-stage acid leaching solution; S4, adding iron powder to the two-stage acid leaching solution to perform a displacement reaction, and separating a filtrate B; S5, subjecting the filtrate B to indium extraction to obtain a first organic phase and a first aqueous phase; S6, subjecting the first organic phase to indium back extraction and zinc-indium displacement reaction to obtain sponge indium; and subjecting the first aqueous phase to impurity removal and concentration crystallization to obtain zinc sulfate.

[0007] Compared with the prior art, the present application has the following advantages: The invention firstly performs high-chlorine zinc slag roasting treatment at high temperature to volatilize most of the chlorine to obtain low-chlorine zinc slag and zinc-containing smoke dust, and then performs a leaching step, and then sequentially treats the obtained filter residue A and filter residue A to obtain products such as zinc, bismuth, indium, cadmium, iron slag, cobalt-nickel slag, etc., thereby realizing comprehensive resource utilization of the high-chlorine zinc slag, and in particular, obtaining high-purity zinc sulfate and sponge indium with a purity of more than 99%, and a high recovery rate of zinc and indium of more than 90%. BRIEF DESCRIPTION OF THE DRAWINGS

[0008] Figure 1 This is a schematic diagram of a process for recovering valuable metals from high-chloride zinc slag according to the present invention; Figure 2 Schematic diagram of the process flow of Examples 1-3 of the present invention. DETAILED DESCRIPTION

[0009] In order to make the purpose, technical solutions and advantages of the present invention more clearly understood, the present invention will be further described in detail below with reference to the accompanying drawings and embodiments. It should be understood that the specific embodiments described herein are only used to explain the present invention and are not intended to limit the present invention.

[0010] Regarding the low zinc recovery rate and low product purity of high-chlorine zinc slag, see Figure 1 The present invention provides a process for recovering valuable metals from high-chloride zinc slag, comprising the following steps: S1, high-chlorine zinc slag is roasted to obtain low-chlorine zinc slag and zinc-containing smoke; S2, low-chlorine zinc slag and zinc-containing smoke are mixed and leached in one stage to separate filtrate A and residue A; filtrate A is subjected to zinc extraction, zinc stripping and electrolysis to obtain cathode zinc; S3, adding sulfuric acid solution to the filter residue A for secondary leaching, and separating to obtain secondary acid leaching liquid and lead-containing residue; S4, adding iron powder to the second-stage acid leaching solution to perform a replacement reaction to separate sponge bismuth and filtrate B; S5, extracting the filtrate B with indium to obtain a first organic phase and a first aqueous phase; S6, the first organic phase is subjected to indium stripping and zinc-indium replacement reaction to obtain sponge indium; the first aqueous phase is subjected to impurity removal and concentrated crystallization to obtain zinc sulfate.

[0011] Preferably, in step S1, the zinc content in the high-chlorine zinc slag is 30-60%, and the chlorine content is 10-30%.

[0012] It is understandable that high-chloride zinc slag also contains small amounts of valuable metals such as Pb, Cu, Cd, Co, Ni, and In.

[0013] Preferably, in step S1, the conditions for the calcination treatment include: a calcination temperature of 800-1200° C. and a calcination time of 1-2 hours.

[0014] Preferably, in step S2, the filtrate A is subjected to zinc extraction, zinc stripping and electrolysis to obtain cathode zinc, specifically comprising: after the filtrate A is subjected to zinc extraction, a second organic phase and a second aqueous phase are separated; the second organic phase is subjected to zinc stripping with a sulfuric acid solution to obtain a zinc-lean organic phase and a zinc-rich solution; the zinc-rich solution is subjected to electrolysis to obtain cathode zinc and electrolytic waste solution; in step S2, the electrolytic waste solution is used for the first-stage leaching.

[0015] Further preferably, the conditions for the zinc extraction include: the extractant A comprises P204, the diluent A comprises kerosene; the volume ratio of the extractant A to the diluent A is 1: (3-9); the consumption of the extractant A is 1-2 kg / t, and the extraction time is 20-30 min.

[0016] Further preferably, the conditions for the zinc stripping include: the mass concentration of the sulfuric acid solution is 15-25%, the consumption of the sulfuric acid solution is 3-5 kg / t (the second organic phase), and the reaction time is 20-30 min.

[0017] Further preferably, the conditions for the electrolysis include: the voltage is 1.8-2.2 V, the anode is a carbon electrode, and the cathode is a high-purity zinc plate.

[0018] Further preferably, the conditions for the first-stage leaching include: the liquid-solid ratio (1.5-4.5):1, the leaching temperature is 60-90°C, the leaching time is 1-2 h, and the leaching end-point pH value is 5.0-5.5.

[0019] It should be noted that in the present application, the liquid-solid ratio is a mass ratio, and is the mass ratio of the solution at the current link to the solid material, for example, the liquid-solid ratio of the first-stage leaching, i.e. the mass ratio of the first-stage leaching solution (electrolytic waste solution) to the solid material (low-chlorine zinc residue and zinc-containing dust).

[0020] Preferably, in step S3, the conditions for the second-stage leaching include: the mass concentration of the sulfuric acid solution is 15-25%, the liquid-solid ratio (1.5-4.5):1, the leaching temperature is 70-90°C, the leaching time is 1-2 h, and the leaching end-point pH value is 1.5-2.5.

[0021] Preferably, in step S4, the consumption of the iron powder is 1-2 kg / t, and the reaction time of the displacement reaction is 20-30 min.

[0022] Preferably, in step S5, the conditions for the indium extraction include: the extractant B comprises P204, the diluent B comprises kerosene; the volume ratio of the extractant B to the diluent B is 1: (3-9); the consumption of the extractant B is 0.5-1 kg / t, and the extraction time is 20-30 min.

[0023] Preferably, in step S6, the step of subjecting the first organic phase to indium stripping and zinc-indium displacement reaction to obtain sponge indium specifically comprises: S611, the first organic phase is back-extracted by indium to obtain an indium-depleted organic phase and an indium-rich liquid; S612, zinc powder is added to the indium-rich liquid to perform a zinc-indium displacement reaction to obtain sponge indium and a filtrate C.

[0024] Further preferably, in step S611, the back-extraction conditions of indium include that the back-extraction agent comprises a hydrochloric acid solution with a mass concentration of 35% or more, the amount of the back-extraction agent is 3-5 kg / t (of the first organic phase), and the back-extraction time is 20-30 min.

[0025] It can be understood that, in order to reduce the cost better, the indium-depleted organic phase can be returned to the indium extraction step for reuse.

[0026] Further preferably, in the zinc-indium displacement reaction of step S612, the amount of zinc powder added to the indium-rich liquid is 1-5 kg / t, and the zinc-indium displacement reaction time is 20-30 min.

[0027] Further preferably, the method further comprises: S613, ammonium bicarbonate is added to the filtrate C to perform zinc precipitation, and the separation obtains a filtrate D and a filter residue D; the filtrate D is combined with the ammonium chloride solution obtained by absorbing the chlorine-containing flue gas with ammonium bicarbonate in the roasting treatment of step S1, and ammonium chloride is obtained by concentration and crystallization; S614, the filter residue D is dissolved in a sulfuric acid solution to obtain a first solution; and zinc sulfate is obtained by concentration and crystallization of the first solution.

[0028] Further preferably, in the zinc precipitation treatment of step S613, the amount of ammonium bicarbonate is 3-5 kg / t, and the zinc precipitation time is 20-30 min.

[0029] Further preferably, in step S614, the amount of the sulfuric acid solution used to dissolve the filter residue D is 3-5 kg / t; the mass concentration of the sulfuric acid solution is 10-35%; and the dissolution time is 20-30 min.

[0030] Further preferably, in the roasting treatment of step S1, the amount of ammonium bicarbonate used to absorb the chlorine-containing flue gas is 70-100 kg / t.

[0031] Preferably, in step S6, the impurity removal includes removal of iron, cadmium and cobalt-nickel, specifically including: S621, hydrogen peroxide is added to the first aqueous phase to remove iron, and the separation obtains an iron residue and a filtrate E; S622, zinc powder is added to the filtrate E to remove cadmium, and the separation obtains sponge cadmium and a filtrate F; S623, the filtrate F is subjected to cobalt-nickel removal, and the separation obtains a cobalt-nickel residue and a second solution, and the second solution is subjected to concentration and crystallization to obtain zinc sulfate.

[0032] It can be understood that the first solution can be mixed with the second solution after the first water phase is removed of impurities, and concentrated and crystallized together to obtain zinc sulfate.

[0033] Further preferably, in the iron removal of step S621, the dosage of hydrogen peroxide is 4-6 kg / t, and the reaction time is 20-30 min. The commercially available hydrogen peroxide with a preferred concentration of 30% is used.

[0034] Further preferably, in the cadmium removal of step S622, the amount of zinc powder added to the filtrate E is 1-3 kg / t, and the reaction time for cadmium removal is 20-30 min.

[0035] Further preferably, in step S623, the cobalt and nickel removal includes zinc powder antimony salt purification method, zinc powder arsenic salt purification method, beta-naphthol cobalt removal method or xanthate cobalt removal method, the dosage of the purification agent for cobalt and nickel removal is 2-4 kg / t, and the reaction time for cobalt and nickel removal is 20-30 min. Among them, the above methods for removing nickel and cobalt are commonly used methods, and the methods not specifically limited are conventional methods.

[0036] The main advantages of the present application are: 1. Harmless treatment and recovery of chlorine in the slag. The chlorine in high-chlorine zinc slag is highly hazardous, and the present application realizes the precipitation and solidification of chlorine by high-temperature volatilization and carbon-ammonium absorption, and produces ammonium chloride products.

[0037] 2. Purification of zinc. The leaching of zinc slag produces a large amount of harmful ions, which affects the recovery purity of zinc. The present application processes the filter residue A through a multi-stage purification process to remove impurities other than zinc, and the final leaching solution mainly contains zinc sulfate, which is directly concentrated and crystallized to obtain zinc sulfate monohydrate; cathode zinc is obtained by zinc extraction and zinc back extraction, and the two can effectively recover zinc from high-chlorine zinc slag.

[0038] 3. Comprehensive recovery of valuable metals in flue dust. The present application comprehensively recovers a variety of valuable metals, including cathode zinc, bismuth sponge, indium sponge, cadmium sponge, lead slag, iron slag, cobalt and nickel slag.

[0039] 4. Recycling of waste liquid. The electrolytic waste liquid obtained by zinc electrolysis is a mixed liquid containing a small amount of zinc and a certain amount of sulfuric acid discharged from the electrolytic cell, which is used for one-stage leaching. The waste liquid is not discharged, and the economic recycling of waste liquid is realized.

[0040] The present application will be further described in detail below in conjunction with specific examples. These examples are only used to illustrate the present application and not to limit the scope of the present application. The experimental methods not specified in the following examples are generally carried out according to conventional conditions in the art or according to the conditions recommended by the manufacturer; the raw materials, reagents, etc. used, if not specifically stated, are considered to be raw materials and reagents that can be obtained through conventional market or commercial channels.

[0041] Example 1 Referring to Figure 2 A process for recovering valuable metals from high-chlorine zinc residue, specifically comprising the following steps: S1, high-temperature roasting: high-temperature roasting treatment of high-chlorine zinc residue (smelting dust raw material), roasting temperature is 900℃, roasting time is 2h, most of the chlorine is volatilized, low-chlorine zinc residue, zinc-containing dust and chlorine-containing flue gas are obtained; wherein the chlorine-containing flue gas is absorbed with ammonium bicarbonate, the ammonium bicarbonate consumption is 70kg / t, and the ammonium chloride solution is obtained.

[0042] S2, obtaining filtrate A and cathode zinc; S201, first stage leaching: mixing low-chlorine zinc residue and zinc-containing dust and leaching with electrolytic waste liquid in the process flow, liquid-solid ratio of leaching liquid is 2:1, leaching temperature is 70℃, leaching time is 1h, leaching endpoint pH is 5.2, pressure filtration to obtain filtrate A and residue A; S202, obtaining cathode zinc: Zinc extraction: the filtrate A is extracted with extractant (P204) and diluent (kerosene) in a volume ratio of 1:4, the extractant consumption is 2kg / t, the reaction time is 30min, and the second organic phase and the second aqueous phase are obtained after separation; Zinc stripping: using 20wt% sulfuric acid solution for zinc stripping of the second organic phase, the sulfuric acid solution consumption is 5kg / t, the reaction time is 30min, and the zinc-depleted organic phase and zinc-rich liquid are obtained; Electrolysis: electrolysis of zinc-rich liquid, electrolytic zinc voltage is 1.9V, anode is carbon electrode, cathode is high-purity zinc plate, cathode zinc and electrolytic waste liquid are obtained.

[0043] S3, second stage leaching: adding sulfuric acid solution to residue A for leaching, sulfuric acid solution concentration is 20%wt, consumption is 70kg / t, liquid-solid ratio is 2:1, leaching temperature is 80℃, leaching time is 1h, hydrogen peroxide can be added as needed during the leaching process to ensure that the leaching endpoint pH is 1.5, pressure filtration to obtain second stage acid leaching liquid and lead-containing residue.

[0044] S4, replacing bismuth: adding iron powder to the second stage acid leaching liquid obtained in step S3, iron powder consumption is 2kg / t, reaction time is 30min, after replacement, pressure filtration to obtain sponge bismuth and filtrate B; S5, indium extraction: using P204 as extractant for indium extraction of filtrate B, the volume ratio of extractant (P204) to diluent (kerosene) is 1:5, the extractant consumption is 1kg / t, the reaction time is 30min, and the first organic phase and the first aqueous phase are obtained.

[0045] S6, obtaining sponge indium and zinc sulfate; S611, indium stripping: stripping the first organic phase obtained in step S5 with hydrochloric acid, with a consumption of 5 kg / t of 38% concentrated hydrochloric acid and a reaction time of 30 min to obtain an indium-depleted organic phase and an indium-rich solution; S612, replacing indium: adding zinc powder to the indium-rich solution of step S611, with a zinc powder consumption of 3 kg / t and a reaction time of 30 min, and performing filter pressing after replacement to obtain sponge indium and filtrate C; S613, zinc precipitation: ammonium bicarbonate was added to the filtrate C of step S612 to precipitate zinc, with an ammonium bicarbonate consumption of 5 kg / t and a reaction time of 30 min. Filtrate D and filter residue D were obtained by pressure filtration. The ammonium chloride solution obtained in step S1 and the filtrate D obtained in S613 were combined and concentrated and crystallized to obtain an ammonium chloride product. S614, dissolution: adding the filter residue D to a sulfuric acid solution to dissolve it to obtain a first solution; wherein the concentrated sulfuric acid consumption is 5 kg / t, and the reaction time is 30 min; S621, iron removal: hydrogen peroxide (concentration 30%) was added to the first aqueous phase obtained in step S5 to remove iron, with a hydrogen peroxide consumption of 6 kg / t and a reaction time of 30 min. Filtrate E and iron slag were obtained by filter pressing; S622, cadmium removal: zinc powder was added to the filtrate E at a zinc powder consumption of 3 kg / t, and the reaction time was 30 min. After displacement, filtration was performed to obtain sponge cadmium and filtrate F; S623, second-stage purification: add cobalt-nickel purifier (zinc powder: arsenic salt mass ratio of 1:0.2) to the filtrate F to remove cobalt and nickel, the cobalt-nickel purifier consumption is 4 kg / t, the reaction time is 30 min, and filter press to obtain cobalt-nickel slag and a second solution; the first solution obtained in step S614 and the second solution obtained in step S623 are concentrated and crystallized together to obtain zinc sulfate monohydrate product.

[0046] By adopting the above process, the smelting dust raw material contains Zn 38.25% and Cl 13.59% (low-content metals are not counted), and zinc sulfate monohydrate with a purity of 99.25% can be obtained. The total recovery rate of zinc (calculated based on the total content of zinc in cathode zinc and zinc sulfate monohydrate) is 91.64%.

[0047] Example 2 Reference Figure 2 A process for recovering valuable metals from high-chloride zinc slag comprises the following steps: S1. High-temperature roasting: high-chlorine zinc slag is roasted at a high temperature of 1000°C for 1.5 hours to volatilize most of the chlorine, thereby obtaining low-chlorine zinc slag, zinc-containing smoke and chlorine-containing flue gas; the chlorine-containing flue gas is absorbed by ammonium bicarbonate at an ammonium bicarbonate consumption of 80 kg / t to obtain ammonium chloride solution.

[0048] S2, obtaining filtrate A and cathode zinc; S201, one-stage leaching: mix low-chlorine zinc residue and zinc-containing dust, and use electrolytic waste solution and second-stage acid leaching solution in the process to leach, with a leaching liquid-to-solid ratio of 2:1, a leaching temperature of 60°C, a leaching time of 2h, and a leaching end-point pH of 5.0, to obtain filtrate A and filter residue A through pressure filtration; S202, obtaining cathode zinc: Zinc extraction: the filtrate A is extracted with an extractant (P204) and a diluent (kerosene) in a volume ratio of 1:6, with an extractant consumption of 1.5kg / t, a reaction time of 25min, and a second organic phase and a second aqueous phase obtained through subsequent separation; Zinc stripping: zinc is stripped from the second organic phase using a 20wt% sulfuric acid solution, with a sulfuric acid solution consumption of 4kg / t, a reaction time of 25min, and a zinc-depleted organic phase and a zinc-rich solution obtained; Electrolysis: the zinc-rich solution is electrolyzed, with an electrolytic zinc voltage of 2.0V, a carbon electrode as the anode, and a high-purity zinc plate as the cathode, to obtain cathode zinc and electrolytic waste solution.

[0049] S3, second-stage leaching: the filter residue A is leached with a sulfuric acid solution, with a sulfuric acid solution concentration of 20%wt, a consumption of 90kg / t, a liquid-to-solid ratio of 2:1, a leaching temperature of 80°C, a leaching time of 1h, and a leaching end-point pH of 2.5, to obtain second-stage acid leaching solution and lead-containing residue through pressure filtration.

[0050] S4, bismuth displacement: iron powder is added to the second-stage acid leaching solution obtained in step S3, with an iron powder consumption of 1.5kg / t, a reaction time of 25min, and sponge bismuth and filtrate B obtained through subsequent pressure filtration.

[0051] S5, indium extraction: the filtrate B is extracted with P204 as the extractant, with an extractant (P204) and diluent (kerosene) volume ratio of 1:8, an extractant consumption of 0.8kg / t, and a reaction time of 25min, to obtain a first organic phase and a first aqueous phase.

[0052] S6, obtaining sponge indium and zinc sulfate; S611, indium stripping: the first organic phase obtained in step S5 is stripped with concentrated hydrochloric acid, with a concentrated hydrochloric acid consumption of 4kg / t, a reaction time of 25min, and an indium-depleted organic phase and indium-rich solution obtained; S612, indium displacement: zinc powder is added to the indium-rich solution of step S611, with a zinc powder consumption of 5kg / t, a reaction time of 25min, and sponge indium and filtrate C obtained through subsequent pressure filtration; S613, zinc precipitation: ammonium bicarbonate is added to the filtrate C of step S612 for zinc precipitation, with an ammonium bicarbonate consumption of 4kg / t, a reaction time of 25min, and filtrate D and filter residue D obtained through pressure filtration; the ammonium chloride solution obtained in step S1 and the filtrate D obtained in step S613 are combined and concentrated to obtain an ammonium chloride product through crystallization. S614, dissolving: the filter residue D is added into the sulfuric acid solution to dissolve to obtain a first solution; wherein the consumption of concentrated sulfuric acid is 4 kg / t, and the reaction time is 25 min; S621, iron removal: hydrogen peroxide (concentration 30%) is added into the first aqueous phase obtained in step S5 to remove iron, the consumption of hydrogen peroxide is 5 kg / t, the reaction time is 25 min, and the filter cake is obtained by pressure filtration; S622, cadmium removal: zinc powder is added into the filter liquor E, the consumption of zinc powder is 2 kg / t, the reaction time is 25 min, and sponge cadmium and filter liquor F are obtained by pressure filtration after displacement; S623, two-stage purification: cobalt-nickel purifying agent (zinc powder, arsenic salt mass ratio 1:0.2) is added into the filter liquor F to remove cobalt and nickel, the consumption of cobalt-nickel purifying agent is 3 kg / t, the reaction time is 25 min, and cobalt-nickel residue and second solution are obtained by pressure filtration; the first solution obtained in step S614 and the second solution obtained in step S623 are concentrated and crystallized together to obtain zinc sulfate monohydrate product.

[0053] By using the above process, the smelting dust raw material containing Zn 43.67% and Cl 15.32% (low content metal is not counted) can obtain zinc sulfate monohydrate with a purity of 99.52%, and the total recovery rate of zinc (calculated according to the total content of zinc in cathode zinc and zinc sulfate monohydrate) is 93.61%.

[0054] Example 3 Reference Figure 2 A process for recovering valuable metals from high-chlorine zinc residue, specifically comprising the following steps: S1, high-temperature roasting: high-temperature roasting treatment is performed on the high-chlorine zinc residue, the roasting temperature is 1200℃, the roasting time is 1h, most of the chlorine is volatilized, low-chlorine zinc residue, zinc-containing dust and chlorine-containing flue gas are obtained; wherein the chlorine-containing flue gas is absorbed by ammonium bicarbonate, the consumption of ammonium bicarbonate is 100 kg / t, and ammonium chloride solution is obtained.

[0055] S2, obtaining filter liquor A and cathode zinc; S201, one-stage leaching: the low-chlorine zinc residue and the zinc-containing dust are mixed and leached by using the electrolytic waste liquid and the two-stage acid leaching liquid in the process flow, the leaching liquid solid ratio is 2:1, the leaching temperature is 70℃, the leaching time is 2h, the leaching endpoint pH is 5.5, and the filter cake A and the filter liquor A are obtained by pressure filtration; S202, obtaining cathode zinc: Zinc extraction: the filter liquor A is extracted by the extractant (P204) and the diluent (kerosene) with a volume ratio of 1:5, the consumption of the extractant is 1.0 kg / t, the reaction time is 20 min, and the second organic phase and the second aqueous phase are obtained by separation; Zinc stripping: the second organic phase is subjected to zinc stripping with 20wt% sulfuric acid solution, the consumption of sulfuric acid solution is 3kg / t, the reaction time is 20min, and a zinc-lean organic phase and a zinc-rich solution are obtained; Electrolysis: the zinc-rich solution is subjected to electrolysis, the electrolytic zinc voltage is 2.1V, the anode is a carbon electrode, and the cathode is a high-purity zinc plate, and a cathode zinc and an electrolytic waste solution are obtained.

[0056] S3, two-stage leaching: the filter residue A is added with a sulfuric acid solution for leaching, the concentration of the sulfuric acid solution is 20%wt, the consumption is 100kg / t, the liquid-solid ratio is 2:1, the leaching temperature is 80℃, the leaching time is 2h, hydrogen peroxide can be added in an appropriate amount during the leaching process according to the need, the leaching end-point pH is ensured to be 2.0, and pressure filtration is performed to obtain a two-stage acid leaching solution and a lead-containing residue.

[0057] S4, bismuth displacement: iron powder is added to the two-stage acid leaching solution obtained in step S3, the consumption of iron powder is 1kg / t, the reaction time is 20min, and pressure filtration is performed after displacement to obtain sponge bismuth and a filtrate B.

[0058] S5, indium extraction: the filtrate B is subjected to indium extraction with P204 as an extractant, the volume ratio of the extractant (P204) to the diluent (kerosene) is 1:5, the consumption of the extractant is 0.5kg / t, the reaction time is 20min, and a first organic phase and a first aqueous phase are obtained.

[0059] S6, sponge indium and zinc sulfate are obtained; S611, indium stripping: the first organic phase obtained in step S5 is subjected to stripping with hydrochloric acid, the consumption of concentrated hydrochloric acid is 3kg / t, the reaction time is 20min, and an indium-lean organic phase and an indium-rich solution are obtained; S612, indium displacement: zinc powder is added to the indium-rich solution of step S611, the consumption of zinc powder is 1kg / t, the reaction time is 20min, and pressure filtration is performed after displacement to obtain sponge indium and a filtrate C; S613, zinc precipitation: ammonium bicarbonate is added to the filtrate C of step S612 for zinc precipitation, the consumption of ammonium bicarbonate is 3g / t, the reaction time is 20min, pressure filtration is performed to obtain a filtrate D and a filter residue D; the ammonium chloride solution obtained in step S1 and the filtrate D obtained in S613 are combined and concentrated to obtain an ammonium chloride product by crystallization; S614, dissolution: the filter residue D is added with a sulfuric acid solution for dissolution to obtain a first solution; the consumption of concentrated sulfuric acid is 3kg / t, and the reaction time is 20min; S621, iron removal: hydrogen peroxide (concentration 30%) is added to the first aqueous phase obtained in step S5 for iron removal, the consumption of hydrogen peroxide is 4kg / t, the reaction time is 20min, and pressure filtration is performed to obtain a filtrate E and an iron residue; S622, cadmium removal: zinc powder is added to the filtrate E, the consumption of zinc powder is 1kg / t, the reaction time is 20min, and pressure filtration is performed after displacement to obtain sponge cadmium and a filtrate F; S623, two-stage purification: cobalt-nickel purifying agent (zinc powder, arsenic salt mass ratio of 1:0.2) was added to the filtrate F to remove cobalt and nickel, the cobalt-nickel purifying agent consumption was 2 kg / t, the reaction time was 20 min, and pressure filtration was used to obtain cobalt-nickel residue and a second solution; the first solution obtained in step S614 and the second solution obtained in step S623 were concentrated and crystallized together to obtain zinc sulfate monohydrate product.

[0060] After the above process treatment, the smelting dust raw material containing Zn 46.73% and Cl 17.25% (low content of metal not counted) can obtain zinc sulfate monohydrate with a purity of 99.38%, and the total recovery rate of zinc (calculated according to the total content of zinc in cathode zinc and zinc sulfate monohydrate) is 93.12%.

[0061] Comparative Example 1 The difference from Example 2 is only that step S1 is removed, and the high-chlorine zinc residue is directly used as the raw material for leaching in step S2; other steps and conditions are the same as those in Example 2.

[0062] Comparative Example 2 The difference from Example 2 is only that the leaching end point pH value in step S201 is adjusted to 4.0; other steps and conditions are the same as those in Example 2.

[0063] Comparative Example 3 The difference from Example 2 is only that the leaching end point pH value in step S201 is adjusted to 6.5; other steps and conditions are the same as those in Example 2.

[0064] Comparative Example 4 The difference from Example 2 is only that the one-stage leaching is removed, and the two-stage leaching is directly performed; other steps and conditions are the same as those in Example 2.

[0065] Comparative Example 5 The difference from Example 2 is only that step S4 is removed, and the two-stage acid leaching solution obtained in step S3 is directly used for indium extraction; other steps and conditions are the same as those in Example 2.

[0066] The sponge indium and zinc sulfate obtained in the above Example 2 and Comparative Examples (the smelting dust raw material contains Zn 43.67% and Cl 15.32%) were tested and the results were counted, and the results are shown in Table 1.

[0067] Table 1 Test results of sponge indium and zinc sulfate obtained in Example 2 and Comparative Examples

[0068] From the results in Table 1, it can be found that after direct leaching in Comparative Example 1, the purity of zinc sulfate and sponge indium both decreased significantly, which is mainly due to the influence of chlorine in high-chlorine zinc residue.

[0069] The terminal pH value of the first leaching in the comparative example 2 is 4.0, and the terminal pH value of the first leaching in the comparative example 3 is 6.5, the purity of the obtained zinc sulfate and sponge indium decreases, and the recovery rate of zinc and indium also decreases, thus indicating that too high or too low pH value is not conducive to the comprehensive recovery of indium and zinc; therefore, the terminal pH value of the first leaching is preferably adjusted to 5.0-5.5.

[0070] The first leaching is removed in the comparative example 4, resulting in a decrease in the yield of the obtained zinc sulfate and sponge indium, indicating that the two-stage leaching is conducive to improving the yield.

[0071] The step of replacing bismuth is removed in the comparative example 5, resulting in a decrease in the purity of the obtained zinc sulfate and sponge indium, indicating that removing bismuth before indium extraction is conducive to improving the purity of the product.

[0072] In summary, the high-chlorine zinc residue is first treated by high-temperature roasting in the present application, so that most of the chlorine is volatilized and absorbed by ammonium bicarbonate, obtaining low-chlorine zinc residue, zinc-containing dust and ammonium chloride solution; the low-chlorine zinc residue and the zinc-containing dust are subjected to the first leaching by using the electrolytic waste liquid and the second acid leaching liquid in the process flow, and pressure filtration is performed to obtain filtrate A and filter residue A; the filtrate A and the filter residue A are sequentially treated, so that cathode copper, sponge bismuth, sponge indium, sponge cadmium, lead residue, iron residue, cobalt-nickel residue, ammonium chloride, zinc sulfate monohydrate and other products can be obtained, realizing the comprehensive utilization of the high-chlorine zinc residue, wherein the purity of the zinc sulfate monohydrate is ≥99%, and the total recovery rate of zinc is ≥90%.

[0073] The specific embodiments of the present application described above do not constitute a limitation on the scope of protection of the present application. Any various other corresponding changes and modifications made according to the technical concept of the present application shall be included in the scope of protection of the claims of the present application.

Claims

1. A process for recovering valuable metals from high-chlorine zinc slag, characterized in that: The following steps are involved: S1, high-chlorine zinc slag is roasted to obtain low-chlorine zinc slag and zinc-containing smoke; S2, the low-chlorine zinc slag and the zinc-containing smoke are mixed and subjected to a leaching step to separate and obtain filtrate A and filter residue A; the filtrate A is subjected to zinc extraction, zinc stripping and electrolysis to obtain cathode zinc; S3, adding sulfuric acid solution to the filter residue A for two-stage leaching, and separating to obtain a two-stage acid leaching liquid; S4, adding iron powder to the second-stage acid leaching solution to perform a replacement reaction, and separating to obtain a filtrate B; S5, extracting the filtrate B with indium to obtain a first organic phase and a first aqueous phase; S6, the first organic phase is subjected to indium stripping and zinc-indium replacement reaction to obtain sponge indium; the first aqueous phase is subjected to impurity removal and concentrated crystallization to obtain zinc sulfate.

2. The process for recovering valuable metals from high-chloride zinc slag according to claim 1, wherein: In step S1, the zinc content of the high-chlorine zinc slag is 30-60%, and the chlorine content is 10-30%; The calcination conditions include: a calcination temperature of 800 to 1200° C. and a calcination time of 1 to 2 hours.

3. The process for recovering valuable metals from high-chlorine zinc slag according to claim 1, wherein: In step S2, the filtrate A is subjected to zinc extraction, zinc stripping and electrolysis to obtain cathode zinc, which specifically includes: separating the filtrate A after zinc extraction to obtain a second organic phase and a second aqueous phase; zinc stripping the second organic phase with a sulfuric acid solution to obtain a zinc-depleted organic phase and a zinc-rich liquid; and electrolyzing the zinc-rich liquid to obtain cathode zinc and electrolysis waste liquid. The first stage leaching is carried out using the electrolytic waste liquid.

4. The process for recovering valuable metals from high-chloride zinc slag according to claim 3, wherein: The zinc extraction conditions include: extractant A includes P204, diluent A includes kerosene; the volume ratio of extractant A to diluent A is 1:(3-9); the consumption of extractant A is 1-2 kg / t, and the extraction time is 20-30 min; The zinc stripping conditions include: a mass concentration of the sulfuric acid solution of 15-25%, a consumption of the sulfuric acid solution of 3-5 kg / t, and a reaction time of 20-30 min; The electrolysis conditions include: a voltage of 1.8 to 2.2 V, a carbon electrode as the anode, and a high-purity zinc plate as the cathode.

5. The process for recovering valuable metals from high-chloride zinc slag according to claim 1, wherein: In step S2, the conditions for the first stage leaching include: liquid-solid ratio (1.5-4.5):1, leaching temperature at 60-90°C, leaching time at 1-2 hours, and pH value at the leaching endpoint at 5.0-5.5; In step S3, the conditions for the second-stage leaching include: a mass concentration of the sulfuric acid solution of 15-25%; a liquid-to-solid ratio of (1.5-4.5):1, a leaching temperature of 70-90°C, a leaching time of 1-2 hours, and a leaching endpoint pH of 1.5-2.5; In step S4, the consumption of the iron powder is 1-2 kg / t, and the reaction time of the replacement reaction is 20-30 min.

6. The process for recovering valuable metals from high-chloride zinc slag according to claim 1, wherein: In step S5, the conditions for indium extraction include: extractant B includes P204, diluent B includes kerosene; the volume ratio of extractant B to diluent B is 1:(3-9); the consumption of extractant B is 0.5-1 kg / t, and the extraction time is 20-30 min.

7. The process for recovering valuable metals from high-chloride zinc slag according to claim 1, wherein: In step S6, the step of obtaining sponge indium by stripping indium from the first organic phase and performing zinc-indium replacement reaction specifically includes: S611, stripping the first organic phase with indium to obtain an indium-depleted organic phase and an indium-rich liquid; S612, adding zinc powder to the indium-rich solution to perform a zinc-indium replacement reaction to obtain sponge indium and filtrate C.

8. The process for recovering valuable metals from high-chloride zinc slag according to claim 7, wherein: In step S611, the conditions for the indium stripping include: the stripping agent includes a hydrochloric acid solution with a mass concentration of 35% or more, the amount of the stripping agent is 3-5 kg / t, and the stripping time is 20-30 min; In the zinc-indium replacement reaction of step S612, the amount of zinc powder added to the indium-rich solution is 1-5 kg / t, and the zinc-indium replacement reaction time is 20-30 min; Also includes: S613, adding ammonium bicarbonate to the filtrate C to precipitate zinc, separating to obtain a filtrate D and a filter residue D; combining the filtrate D with the ammonium chloride solution obtained by absorbing the chlorine-containing flue gas with ammonium bicarbonate in the roasting treatment in step S1, and concentrating and crystallizing to obtain ammonium chloride; the amount of ammonium bicarbonate used is 3-5 kg / t, and the zinc precipitation time is 20-30 min; S614, adding sulfuric acid solution to the filter residue D to dissolve it to obtain a first solution; the first solution is concentrated and crystallized to obtain zinc sulfate; wherein the amount of the sulfuric acid solution is 3-5 kg / t; the mass concentration of the sulfuric acid solution is 10-35%; and the dissolution time is 20-30 min.

9. The process for recovering valuable metals from high-chloride zinc slag according to claim 1, wherein: In step S6, impurity removal includes iron removal, cadmium removal, and cobalt and nickel removal, and specifically includes: S621, adding hydrogen peroxide to the first aqueous phase to remove iron, and separating to obtain iron slag and filtrate E; S622, adding zinc powder to the filtrate E to remove cadmium, and separating to obtain sponge cadmium and filtrate F; S623, removing cobalt and nickel from the filtrate F, separating to obtain cobalt-nickel slag and a second solution, and concentrating and crystallizing the second solution to obtain zinc sulfate.

10. The process for recovering valuable metals from high-chloride zinc slag according to claim 9, characterized in that: In the iron removal in step S621, the amount of hydrogen peroxide used is 4-6 kg / t, and the reaction time is 20-30 min; In the cadmium removal in step S622, the amount of zinc powder added to the filtrate E is 1-3 kg / t, and the reaction time for the cadmium removal is 20-30 min; In step S623 , the removal of cobalt and nickel includes a zinc powder antimony salt purification method, a zinc powder arsenic salt purification method, a β-naphthol cobalt removal method, or a xanthate cobalt removal method.

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