WO2025044261A1 - Method for efficiently recovering sulfur and zinc-silver from high-sulfur residues obtained in zinc hydrometallurgy - Google Patents

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

This application claims the priority of the Chinese patent application filed with the China Patent Office 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 by reference into this application.

Technical Field

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

As a basic chemical raw material, sulfur is widely used in coatings, organic synthesis, acid production, medicine, food and other industries. Sulfur products currently mainly come from natural and recycled sulfur. Recycled sulfur has the advantages of high purity, few impurities and stable quality. my country has a large annual demand for sulfur, a small domestic sulfur production and a high degree of external dependence.

The hydrometallurgical zinc smelting process will produce a large amount of leaching slag, which is called high-sulfur slag or zinc-sulfur slag because of its high sulfur content. The sulfur grade is about 40-55%, mainly in the form of elemental sulfur, sulfide and sulfate, and contains valuable metal elements such as silver, lead, zinc and germanium. Therefore, comprehensive recovery of high-sulfur slag and development of sulfur and valuable metal recovery technology in high-sulfur slag are of great significance to alleviate the contradiction between sulfur supply and demand in my country, maximize the level of resource utilization, and reduce the environmental pressure caused by leaching slag storage. Traditional treatment methods are inefficient, cause waste of resources, and may cause secondary pollution. Therefore, how to efficiently and environmentally friendly recover sulfur and valuable metals (such as zinc and silver) from high-sulfur slag has become an important topic in the current metallurgical industry and environmental protection field. Patent 201410685240.X reports a method for recovering sulfur from zinc leaching sulfur-containing slag, which uses a high-pressure reactor to transform sulfur slag to improve sulfur recovery rate, but the cost is high and the operation is inconvenient. Patent 201310422340.9 first rapidly heats and pressurizes the oxygen leaching solution in the atmospheric oxygen-enriched direct hydrometallurgy of zinc, and then flashes it. After separating the elemental sulfur by hot filtration, the molten elemental sulfur is cooled by water into granular solid elemental sulfur, and the recovery rate of the sulfur element is difficult to guarantee. Patent 202210102498.7 crushes and screens the zinc-sulfur slag again, and performs secondary flotation to improve the recovery rate of sulfur concentrate, but does not consider the operating cost, the recovery of valuable metal elements, etc.

Currently, there are many studies on sulfur recovery 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.

Summary of the invention

In view of the above problems, the present application provides a method for efficiently recovering sulfur and zinc and silver from high-sulfur slag of hydrometallurgical 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 and silver recovery rate and high recovery cost.

In order to achieve the above purpose, this application adopts the following technical solutions:

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 reacting the obtained zinc ammonia complex and silver ammonia complex and entering 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 the solution no longer produces precipitation, 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.

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

Preferably, the inducer in step one is ferric sulfate, and the concentration of ferric sulfate in the bottom flow liquid is 0.01-0.5 mol/L.

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

The technical effect achieved by adopting step one 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%.

The reaction principle is as follows:
FeS ₂ +Fe ₂ (SO ₄ ) ₃ =3FeSO ₄ +2S

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

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

Preferably, the added mass of each agent in the inhibitor is the same, which is 100-300 g/ton of slag.

The technical effect achieved by adopting step 2 is: by controlling the reaction conditions and the added inhibitor, the sulfur recovery rate is increased to more than 95%; and more metals such as lead and zinc are introduced into the tailings.

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

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

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

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

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.

The technical effect achieved by adopting the above technical scheme is that the recovery rate of zinc in basic zinc carbonate is 65-90%.

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, cadmium, etc., it is listed as HW48 non-ferrous metal smelting waste, which is extremely harmful 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, silver, etc., so it is also one of the important raw materials for recovering sulfur and silver, zinc, and lead.

In order to efficiently separate the heavy metal phase and elemental sulfur in the sulfur-containing oxygen pressure leaching slag, and simultaneously directionally recover the zinc and silver valuable metals in the form of sulfate in the sulfur-containing oxygen pressure leaching slag, so as to facilitate the smelting process recovery, and at the same time overcome the technical problem of high recovery cost in the prior art, the technical concept of the present application is as follows:

Considering that part of the sulfur in the high-sulfur slag exists in the form of sulfide, if this part of the sulfur 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. In combination with 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 the existing process, lead and zinc entering the concentrate affecting sulfur purification, low zinc and silver recovery rate, and high recovery cost.

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

(1) By adding ferric sulfate inducer, the sulfide in the high-sulfur slag is converted into elemental sulfur, thereby improving the sulfur recovery rate and simplifying the operation.

(2) By controlling flotation conditions and adding inhibitors, the sulfur recovery rate is improved while inhibiting the transfer of zinc and lead to the concentrate, reducing the impact on subsequent sulfur recovery.

(3) The zinc and silver in the remaining slag are leached out in the form of zinc-ammonia and silver-ammonia complexes by adopting the principle of zinc-silver-ammonia complexation. 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. The metallic zinc is efficiently recovered, and waste treatment and energy conservation and emission reduction are achieved at the same time.

DETAILED DESCRIPTION

The technical solutions in the embodiments of the present application are described clearly and completely below. Obviously, the described embodiments are only part of the embodiments of the present application, not all of the embodiments. Based on the embodiments in the present application, all other embodiments obtained by ordinary technicians in this field without creative work are within the scope of protection of the present application.

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 art. The present application is further described below in conjunction with specific examples.

Example 1

Step 1: Take 1L of oxygen pressure leaching second stage underflow liquid, which contains high-sulfur slag, the pH value of the solution is 1, the temperature of the solution 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, and stir at a stirring speed of 100 rpm for 120 minutes.

Step 2: In the flotation stage, add an appropriate amount of sodium hydroxide to adjust the pH value of the solution to control the pH value at 2, blow air into the solution at an air flow rate of 1 L/min, add sodium sulfide + zinc sulfate + sodium sulfite inhibitors, and the dosage of each agent is 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 enter the tailings. The sulfur recovery rate is greater than 95%, and a concentrate and a slurry containing tailings are obtained respectively.

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 filtration residue.

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 be 0.5 mol/L, the reaction temperature is controlled to be 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.

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 recovery rate of silver is 60%.

Step 6: The final zinc ammonia 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 ammonia 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%.

Example 2

Step 1: Take 2L of oxygen pressure leaching second stage underflow liquid, which contains high-sulfur slag, the pH value of the solution is 3, the temperature of the solution 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, and stir at a stirring speed of 500 rpm for 30 minutes.

Step 2: In 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 4, air is blown into the solution, the air flow rate is 5L/min, ammonium sulfide + zinc sulfate + sodium sulfite inhibitors are added, the dosage of each agent is 300g/ton of slag, the flotation time is 15 minutes, and the sulfur recovery rate is improved by controlling the reaction conditions and adding inhibitors, and more metals such as lead and zinc enter the tailings. The sulfur recovery rate is greater than 95%, and a concentrate and a slurry containing tailings are obtained respectively.

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 filtration residue.

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 be 2 mol/L. The reaction temperature is controlled to be 60°C. The stirring speed is 100 rpm. 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.

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 recovery rate of silver is 90%.

Step 6: The final zinc ammonia 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 ammonia 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%.

Example 3

Step 1: Take 1.5L of oxygen pressure leaching second stage bottom flow liquid, which contains high-sulfur slag, the pH value of the solution is 2, the temperature of the solution is 70°C, the solid-liquid ratio (volume ratio) is 1:7, and the solution also contains a certain amount of zinc, silver and other valuable metals. Add ferric sulfate as an inducer to the solution at a concentration of 0.1mol/L to convert the iron disulfide in the high-sulfur slag into elemental sulfur and increase the content of elemental sulfur. Stir and stir at a speed of 0.1mol/L. The speed was 300 rpm and the reaction time was 60 minutes.

Step 2: Entering the flotation stage, adding an appropriate amount of sodium hydroxide to adjust the pH value of the solution to control the pH value at 3, blowing air into the solution, the air flow rate is 3L/min, adding ammonium sulfide + sodium hydroxide + zinc sulfate + sodium sulfite inhibitors, 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 enter the tailings slag, the sulfur recovery rate is greater than 95%, and the concentrate and the slurry containing the tailings slag are obtained respectively.

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 filtration residue.

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 be 1 mol/L. The reaction temperature is controlled to be 40°C. The stirring speed is 250 rpm. The reaction is carried out for 2 hours. The zinc and silver in the slag react with ammonia water and ammonium sulfate to generate zinc ammonia complex and silver ammonia complex, which enter the liquid phase.

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 recovery rate of silver is 75%.

Step 6: The final zinc ammonia 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 ammonia 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%.

Comparative Example 1

Take 1L of oxygen pressure leaching second stage underflow liquid, which contains high sulfur slag, the pH value of the solution is 1, the temperature of the solution is 80℃, and the solid-liquid ratio (volume ratio) is 1:5. It is directly sent to the flotation stage, with an air flow of 1L/min, a flotation time of 10 minutes, a stirring speed of 100rpm, a flotation temperature of 80℃, and a sulfur recovery rate of about 75%. The concentrate is sent to the hot filtration stage. The recovery of zinc and silver is not considered.

Although the above-mentioned 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 (12)

A method for efficiently recovering sulfur and zinc-silver from high-sulfur slag from hydrometallurgical zinc smelting, characterized in that it comprises the following steps: 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. The method according to claim 1 is characterized in that the pH value of the second-stage bottom flow liquid in step 1 is 1 to 3, the temperature is 60 to 80° C., and the solid-liquid ratio is 1:4 to 1:10. The method according to claim 1 is 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 to 0.5 mol/L. The method according to claim 1 is characterized in that the stirring speed in step 1 is 100 to 500 rpm, and the stirring time is 30 to 120 minutes. The method according to claim 1, characterized in that the pH value in step 2 is 1.5 to 4, and the air flow rate is 1 to 5 L/min. 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. The method according to claim 6 is characterized in that the mass of each agent added in the inhibitor is the same, which is 100-300g/ton of slag. The method according to claim 1, characterized in that the ammonia water in step 4 is The concentration 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. The method according to claim 1, characterized in that the temperature of the water bath in step 5 is 50-70°C. The method according to claim 1 is characterized in that the temperature of the reaction in step 6 is 20 to 40° C., and the pH value of the reaction solution in the reaction is 4 to 7. The method according to any one of claims 1 to 9, characterized in that the recovery rate of silver is 60 to 90%. 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 to 90%.