Method for efficiently recovering sulfur and zinc-silver from high-sulfur residues obtained in zinc hydrometallurgy

Abstract

The present application relates to the field of resource recovery, and discloses a method for efficiently recovering sulfur and zinc-silver from high-sulfur residues obtained in zinc hydrometallurgy. A large amount of high-sulfur residues are contained in second-stage underflow liquid obtained by performing oxygen pressure leaching. In the present application, a proper amount of iron sulfate is first added to serve as an inducing agent, iron disulfide in high-sulfur residues is converted into elemental sulfur, so as to increase the content of the elemental sulfur; in a flotation stage, reaction conditions are controlled and an inhibitor is added, so as to improve a recovery rate of sulfur, enable more metals such as lead-zinc to enter tailings, and provide a basis for further improvement of the purity of the sulfur; the tailings and hot filtered residues are mixed with filtrate to prepare a slurry, ammonia water is added to the slurry, a zinc-ammonia complex method is used to dissolve zinc-silver, then a silver mirror reaction is used to recover metallic silver, and finally a zinc-ammonia complex solution is used to absorb and capture carbon dioxide to prepare a basic zinc carbonate product. The present application is simple to operate, has good application prospects, and overcomes the problems of a low sulfur recovery rate in existing processes, an impact on sulfur purification caused by the lead-zinc entering concentrate, a low zinc-silver recovery rate and high recovery costs.

Description

A method for efficiently recovering sulfur and zinc-silver from high-sulfur slag from hydrometallurgical zinc smelting

[0001] This application claims priority to a Chinese patent application filed with the Patent Office of China on August 30, 2023, with application number CN202311100902.8 and invention name “A method for efficiently recovering sulfur and zinc-silver from high-sulfur slag from hydrometallurgical zinc smelting”, the entire contents of which are incorporated herein by reference. Technical Field

[0002] The present application relates to the technical field of resource recovery, and more specifically to a method for efficiently recovering sulfur and zinc-silver from high-sulfur slag from hydrometallurgical zinc smelting. Background Art

[0003] Sulfur, a basic chemical raw material, is widely used in coatings, organic synthesis, acid production, pharmaceuticals, and food industries. Currently, sulfur products primarily come from naturally occurring and recycled sources. Recycled sulfur offers advantages such as high purity, low impurities, and consistent quality. my country has a high annual demand for sulfur, but domestic production is low, resulting in a high degree of dependence on foreign sources.

[0004] The hydrometallurgical zinc smelting process produces a large amount of leach residue. Due to its high sulfur content, this residue is known as high-sulfur slag or zinc-sulfur slag. The sulfur grade is approximately 40-55%, primarily in the form of elemental sulfur, sulfides, and sulfates, and also contains valuable metals such as silver, lead, zinc, and germanium. Therefore, comprehensive recovery of high-sulfur slag and the development of technologies for recovering sulfur and valuable metals from high-sulfur slag are crucial for alleviating the imbalance between sulfur supply and demand in my country, maximizing resource utilization, and alleviating the environmental impact of leach residue storage. Traditional treatment methods are inefficient, wasteful, and potentially generate secondary pollution. Therefore, efficient and environmentally friendly recovery of sulfur and valuable metals (such as zinc and silver) from high-sulfur slag has become a major challenge in the metallurgical industry and environmental protection. Patent 201410685240.X reports a method for recovering sulfur from sulfur-containing zinc leaching slag. This method utilizes a high-pressure reactor to transform the sulfur slag to improve sulfur recovery, but this method is costly and inconvenient to operate. Patent 201310422340.9 rapidly heats and pressurizes the oxygen leaching solution in atmospheric oxygen-enriched direct hydrometallurgy, flashing it. After hot filtration to separate elemental sulfur, the molten sulfur is then water-cooled into granular solid elemental sulfur. This approach makes it difficult to guarantee sulfur recovery. Patent 202210102498.7 further crushes and screens the zinc-sulfur slag, and performs secondary flotation to improve the recovery of the sulfur concentrate. However, this approach fails to consider operating costs or the recovery of valuable metal elements.

[0005] Currently, there are many studies on the recovery of sulfur from high-sulfur slag, but the technology of low cost, high recovery rate and recovery of valuable metal resources is still a difficult problem that needs to be overcome.

[0006] Summary of the Invention

[0007] In response to the above problems, the present application provides a method for efficiently recovering sulfur and zinc-silver from high-sulfur slag in hydrometallurgy zinc smelting, which solves the above-mentioned problems of low sulfur recovery rate, the entry of lead and zinc into the concentrate affecting sulfur purification, low zinc-silver recovery rate and high recovery cost.

[0008] In order to achieve the above objectives, this application adopts the following technical solutions:

[0009] Step 1: adding an inducer to the bottom flow liquid of the second stage of oxygen pressure leaching, wherein the bottom flow liquid contains high-sulfur slag, stirring, and converting the iron disulfide in the high-sulfur slag into elemental sulfur;

[0010] Step 2: flotation stage, adding sodium hydroxide to the system obtained in step 1 to adjust the pH value of the underflow liquid, blowing air, adding inhibitors, and flotation for 8 to 30 minutes to obtain concentrate and slurry containing tailings;

[0011] Step 3: transporting the concentrate to a hot filtration section to recover high-purity sulfur, and returning the hot filter residue to the slurry containing tailings to obtain a slurry containing tailings and hot filter residue;

[0012] Step 4: adding ammonia water to the slurry containing tailings and hot filter residue, stirring at a certain temperature, and reacting the obtained zinc ammonia complex and silver ammonia complex and entering the liquid phase;

[0013] Step 5: filtering the system obtained in step 4 to obtain a solution containing zinc and silver, adding formaldehyde or acetaldehyde solution to the solution containing zinc and silver under water bath conditions until the solution no longer produces precipitation, and filtering to obtain silver and zinc ammonia complex solutions respectively;

[0014] Step 6: The zinc ammonia complex solution is used to capture and absorb waste gas containing carbon dioxide, react under certain conditions until the zinc ammonia complex solution no longer produces precipitation, and filter to obtain basic zinc carbonate.

[0015] Preferably, the pH value of the second-stage bottom flow liquid in step 1 is 1-3, the temperature is 60-80° C., and the solid-liquid ratio is 1:4-1:10.

[0016] Preferably, the inducer in step 1 is ferric sulfate, and the concentration of ferric sulfate in the bottom flow liquid is 0.01 to 0.5 mol / L.

[0017] Preferably, the stirring speed in step 1 is 100 to 500 rpm, and the stirring time is 30 to 120 minutes.

[0018] The technical effect achieved by adopting step 1 is that ferric sulfate converts iron disulfide in high-sulfur slag into elemental sulfur, thereby increasing the content of elemental sulfur and improving the recovery rate by 2 to 8%.

[0019] The reaction principle is as follows: FeS2+Fe2(SO4)3=3FeSO4+2S

[0020] Preferably, the pH value in step 2 is 1.5 to 4, and the air flow rate is 1 to 5 L / min.

[0021] Preferably, the inhibitor in step 2 is a combination of sodium sulfide, zinc sulfate and sodium sulfite, or a combination of ammonium sulfide, zinc sulfate and sodium sulfite, or a combination of ammonium sulfide, sodium hydroxide, zinc sulfate and sodium sulfite.

[0022] Preferably, the added mass of each agent in the inhibitor is the same, which is 100 to 300 g per ton of slag.

[0023] The technical effects achieved by adopting step 2 are: by controlling the reaction conditions and the added inhibitors, the sulfur recovery rate is improved to be greater than 95%; and more metals such as lead and zinc are allowed to enter the tailings.

[0024] Preferably, the concentration of the ammonia water in the slurry in step 4 is 0.5 to 2 mol / L;

[0025] The temperature is 20-60° C., the stirring speed is 100-400 rpm, and the stirring time is 1-3 hours.

[0026] Preferably, the temperature of the water bath in step 5 is 50-70°C.

[0027] The technical effect achieved by adopting the above technical solution is: the silver recovery rate is 60-90%.

[0028] Preferably, the reaction temperature in step six is ​​20-40° C., and the pH value of the reaction solution in the reaction is 4-7.

[0029] The technical effect achieved by adopting the above technical solution is: the recovery rate of zinc in basic zinc carbonate is 65-90%.

[0030] The hydrometallurgical zinc smelting process produces a large amount of leaching slag. Because the leaching slag often contains a variety of harmful elements such as lead, mercury, arsenic, and cadmium, it is listed as HW48 non-ferrous metal smelting waste, which poses great harm to the environment and human health. At the same time, the slag has a high sulfur content and contains a variety of valuable metals such as zinc, lead, and silver. Therefore, it is also one of the important raw materials for recovering sulfur and silver, zinc, and lead.

[0031] In order to efficiently separate the heavy metal phase and elemental sulfur in the sulfur-containing oxygen pressure leaching residue, and simultaneously directionally recover the zinc and silver valuable metals in the form of sulfates in the sulfur-containing oxygen pressure leaching residue, so as to facilitate the smelting process recovery and overcome the technical problem of high recovery cost in the existing technology, the technical concept of the present application is as follows:

[0032] Taking into account that part of the sulfur element in the high-sulfur slag exists in the form of sulfide, if this part of the sulfur element can be converted into sulfur, it will help to improve the sulfur recovery rate; because lead and zinc will affect the subsequent sulfur purification process, inhibitors are added in the flotation stage to allow more lead and zinc to enter the tailings; the slag and filtrate still contain valuable metals, and a step-by-step recovery method is proposed to recover zinc and silver. Combining the physical and chemical properties of silver and zinc, zinc and silver are first dissolved into the solution through a complex reaction, and then metallic silver is recovered through a silver mirror reaction, and metallic zinc is recovered through a carbonation process. The present application overcomes the problems of low sulfur recovery rate in existing processes, the entry of lead and zinc into the concentrate affecting sulfur purification, low zinc and silver recovery rate, and high recovery cost.

[0033] It can be seen from the above technical solution that compared with the prior art, the advantages of this application are:

[0034] (1) By adding ferric sulfate inducer, the sulfide in the high-sulfur slag is converted into elemental sulfur, which improves the sulfur recovery rate and is simple to operate.

[0035] (2) By controlling flotation conditions and adding inhibitors to improve sulfur recovery, the transfer of zinc and lead to the concentrate is inhibited, reducing the impact on subsequent sulfur recovery.

[0036] (3) The zinc, silver and ammonia complexation principle is adopted to leach the zinc and silver in the remaining slag in the form of zinc-ammonia and silver-ammonia complexes, and the metallic silver is recovered by the silver mirror method. The zinc-ammonia complex solution is then used to absorb and capture carbon dioxide to prepare a basic zinc carbonate product, which efficiently recovers metallic zinc and achieves waste treatment and energy conservation and emission reduction. DETAILED DESCRIPTION

[0037] The following is a clear and complete description of the technical solutions in the embodiments of this application. Obviously, the embodiments described are only part of the embodiments of this application, not all of them. Based on the embodiments in this application, all other embodiments obtained by ordinary technicians in this field without making creative efforts are within the scope of protection of this application.

[0038] Unless otherwise specified in the following examples, the reagents used are all common commercial products or are prepared by conventional means, and the equipment used are all conventional equipment in the field. The present application is further described below with reference to specific examples.

[0039] Example 1

[0040] Step 1: Take 1L of oxygen pressure leaching second-stage underflow liquid containing high-sulfur slag, the solution has a pH of 1, the solution temperature is 80°C, and the solid-liquid ratio (volume ratio) is 1:5. Add ferric sulfate as an inducer to the solution at a concentration of 0.05 mol / L to convert the iron disulfide in the high-sulfur slag into elemental sulfur. Stir at a speed of 100 rpm and react for 120 minutes.

[0041] Step 2: During the flotation stage, an appropriate amount of sodium hydroxide is added to adjust the pH value of the solution to 2. Air is blown into the solution at an air flow rate of 1 L / min. Sodium sulfide, zinc sulfate, and sodium sulfite inhibitors are added, with the dosage of each agent being 100 g / ton of slag. The flotation time is 10 minutes. By controlling the reaction conditions and adding inhibitors, the sulfur recovery rate is improved, and more metals such as lead and zinc are incorporated into the tailings. The sulfur recovery rate is greater than 95%, and a concentrate and a slurry containing the tailings are obtained, respectively.

[0042] Step 3: The concentrate obtained by flotation is sent to the hot filtration section to recover high-purity sulfur, and the residue after hot filtration is returned to the slurry containing tailings to obtain a slurry containing tailings and hot filter residue.

[0043] Step 4: Add ammonia water to the slurry containing tailings and hot filter residue. The concentration of ammonia water in the slurry is controlled to 0.5 mol / L, the reaction temperature is controlled to 20°C, the stirring speed is 400 rpm, and the reaction is carried out for 1 hour. The zinc and silver in the slag react with ammonia water and ammonium sulfate to form zinc ammonia complex and silver ammonia complex, which enter the liquid phase.

[0044] Step 5: Filter to obtain a solution containing zinc and silver, add formaldehyde solution to the solution in a 50°C water bath until the solution no longer produces precipitation, and filter to recover metallic silver. The silver recovery rate is 60%.

[0045] Step 6: The final zinc-ammine complex solution is used to capture and absorb waste gas containing carbon dioxide. The reaction temperature is controlled at 20°C and the pH value is 4.5. The zinc-ammine complex in the solution is converted into basic zinc carbonate precipitate, which is filtered to obtain the basic zinc carbonate product. The zinc recovery rate is 65%.

[0046] Example 2

[0047] Step 1: Take 2L of oxygen pressure leaching second-stage underflow liquid containing high-sulfur slag, the solution has a pH of 3, the solution temperature is 60°C, and the solid-liquid ratio (volume ratio) is 1:10. Add ferric sulfate as an inducer to the solution at a concentration of 0.5 mol / L to convert the iron disulfide in the high-sulfur slag into elemental sulfur. Stir at a speed of 500 rpm and react for 30 minutes.

[0048] Step 2: During the flotation stage, an appropriate amount of sodium hydroxide is added to adjust the pH value of the solution to 4. Air is blown into the solution at a flow rate of 5 L / min. Ammonium sulfide, zinc sulfate, and sodium sulfite inhibitors are added at a dosage of 300 g / ton of slag. The flotation time is 15 minutes. By controlling the reaction conditions and adding inhibitors, the sulfur recovery rate is improved, and more metals such as lead and zinc are incorporated into the tailings. The sulfur recovery rate is greater than 95%, and a concentrate and a slurry containing the tailings are obtained.

[0049] Step 3: The concentrate obtained by flotation is sent to the hot filtration section to recover high-purity sulfur, and the residue after hot filtration is returned to the slurry containing tailings to obtain a slurry containing tailings and hot filter residue.

[0050] Step 4: Add ammonia water to the slurry containing tailings and hot filter residue. The concentration of ammonia water in the slurry is controlled to 2 mol / L, the reaction temperature is controlled to 60°C, the stirring speed is 100 rpm, and the reaction is carried out for 3 hours. The zinc and silver in the slag react with ammonia water and ammonium sulfate to form zinc ammonia complex and silver ammonia complex, which enter the liquid phase.

[0051] Step 5: Filter to obtain a solution containing zinc and silver, add formaldehyde solution to the solution in a 70°C water bath until the solution no longer produces precipitation, and filter to recover metallic silver. The silver recovery rate is 90%.

[0052] Step 6: The final zinc-ammine complex solution is used to capture and absorb waste gas containing carbon dioxide. The reaction temperature is controlled at 40°C and the pH value is 7. The zinc-ammine complex in the solution is converted into basic zinc carbonate precipitate, which is filtered to obtain the basic zinc carbonate product. The zinc recovery rate is 90%.

[0053] Example 3

[0054] Step 1: Take 1.5L of the second-stage oxygen pressure leaching underflow solution, which contains high-sulfur slag. The solution has a pH of 2, a temperature of 70°C, a solid-to-liquid ratio (volume ratio) of 1:7, and contains a certain amount of valuable metals such as zinc and silver. Add ferric sulfate as an inducer to the solution at a concentration of 0.1 mol / L to convert the iron disulfide in the high-sulfur slag into elemental sulfur, thereby increasing the elemental sulfur content. Stir the solution at a speed of 300 rpm for 60 minutes.

[0055] Step 2: Entering the flotation stage, an appropriate amount of sodium hydroxide is added to adjust the pH value of the solution to control the pH value at 3, air is blown into the solution at an air flow rate of 3L / min, and ammonium sulfide, sodium hydroxide, zinc sulfate, and sodium sulfite inhibitors are added. The dosage of each agent is 200g / ton of slag. The flotation time is 15 minutes. By controlling the reaction conditions and adding inhibitors, the sulfur recovery rate is improved, and more metals such as lead and zinc are incorporated into the tailings. The sulfur recovery rate is greater than 95%, and a concentrate and a slurry containing tailings are obtained respectively.

[0056] Step 3: The concentrate obtained by flotation is sent to the hot filtration section to recover high-purity sulfur, and the residue after hot filtration is returned to the slurry containing tailings to obtain a slurry containing tailings and hot filter residue.

[0057] Step 4: Add ammonia water to the slurry containing tailings and hot filter residue. The concentration of ammonia water in the slurry is controlled to 1 mol / L, the reaction temperature is controlled to 40°C, the stirring speed is 250 rpm, and the reaction is carried out for 2 hours. The zinc and silver in the slag react with ammonia water and ammonium sulfate to form zinc ammonia complex and silver ammonia complex, which enter the liquid phase.

[0058] Step 5: Filter to obtain a solution containing zinc and silver, add formaldehyde or acetaldehyde solution to the solution in a 60°C water bath until the solution no longer produces precipitation, and filter to recover silver. The silver recovery rate is 75%.

[0059] Step 6: The final zinc-ammine complex solution is used to capture and absorb waste gas containing carbon dioxide. The reaction temperature is controlled at 30°C and the pH value is 5.5. The zinc-ammine complex in the solution is converted into basic zinc carbonate precipitate, which is filtered to obtain the basic zinc carbonate product. The zinc recovery rate is 80%.

[0060] Comparative Example 1

[0061] A 1L sample of underflow from the second stage of oxygen pressure leaching, containing high-sulfur slag, was taken. The solution had a pH of 1, a temperature of 80°C, and a solid-to-liquid ratio (volume ratio) of 1:5. It was fed directly to the flotation stage with an air flow of 1L / min, a flotation time of 10 minutes, an agitation speed of 100 rpm, and a flotation temperature of 80°C. The sulfur recovery rate was approximately 75%. The concentrate was then fed to the hot filtration stage. Zinc and silver recovery was not considered.

[0062] Although the above embodiment provides a detailed description of the present application, it is only a part of the embodiments of the present application rather than all the embodiments. People can also obtain other embodiments based on this embodiment without creativity, and these embodiments all fall within the scope of protection of the present application.

Claims

1. A method for efficiently recovering sulfur and zinc-silver from high-sulfur slag from hydrometallurgical zinc smelting, characterized in that: The following steps are involved: Step 1: adding an inducer to the second-stage underflow liquid of oxygen pressure leaching, wherein the underflow liquid contains high-sulfur slag, stirring, and converting the iron disulfide in the high-sulfur slag into elemental sulfur; Step 2: flotation stage, adding sodium hydroxide to the system obtained in step 1 to adjust the pH value of the underflow liquid, blowing air, adding inhibitors, and flotation for 8 to 30 minutes to obtain concentrate and slurry containing tailings, respectively; Step 3: transporting the concentrate to the hot filter section to recover high-purity sulfur, and returning the hot filter residue to the slurry containing tailings to obtain a slurry containing tailings and hot filter residue; Step 4: adding ammonia water to the slurry containing tailings and hot filter residue, stirring at a certain temperature, and the zinc ammonia complex and silver ammonia complex obtained by the reaction enter the liquid phase; Step 5: filtering the system obtained in step 4 to obtain a solution containing zinc and silver, adding formaldehyde or acetaldehyde solution to the solution containing zinc and silver under water bath conditions until no precipitation is generated, and filtering to obtain silver and zinc ammonia complex solutions respectively; Step 6: The zinc ammonia complex solution is used to capture and absorb waste gas containing carbon dioxide, react under certain conditions until the zinc ammonia complex solution no longer produces precipitation, and filter to obtain basic zinc carbonate.

2. The method according to claim 1, characterized in that The pH value of the second stage bottom flow liquid in step 1 is 1-3, the temperature is 60-80° C., and the solid-liquid ratio is 1:4-1:

10.

3. The method according to claim 1, characterized in that The inducer in step 1 is ferric sulfate, and the concentration of ferric sulfate in the bottom flow liquid is 0.01-0.5 mol / L.

4. The method according to claim 1, characterized in that: The stirring speed in step 1 is 100 to 500 rpm, and the stirring time is 30 to 120 minutes.

5. The method according to claim 1, characterized in that: In step 2, the pH value is 1.5-4, and the air flow rate is 1-5 L / min.

6. The method according to claim 1, characterized in that The inhibitor in step 2 is a combination of sodium sulfide, zinc sulfate and sodium sulfite, or a combination of ammonium sulfide, zinc sulfate and sodium sulfite, or a combination of ammonium sulfide, sodium hydroxide, zinc sulfate and sodium sulfite.

7. The method according to claim 6, characterized in that The added mass of each agent in the inhibitor is the same, which is 100-300g / ton of slag.

8. The method according to claim 1, characterized in that The concentration of the ammonia water in the slurry in step 4 is 0.5-2 mol / L; The temperature is 20-60° C., the stirring speed is 100-400 rpm, and the stirring time is 1-3 hours.

9. The method according to claim 1, characterized in that: The temperature of the water bath in step 5 is 50-70°C.

10. The method according to claim 1, characterized in that The reaction temperature in step six is ​​20-40° C., and the pH value of the reaction solution in the reaction is 4-7.

11. The method according to any one of claims 1 to 9, characterized in that: The silver recovery rate is 60-90%.

12. The method according to any one of claims 1 to 10, characterized in that: The recovery rate of zinc in the basic zinc carbonate is 65-90%.