Method for recovering tin and copper resources from copper slag by two-step gas phase oxidation-reduction roasting

Abstract

A two-step method for recovering tin and copper resources from copper slag using a gas-phase oxidation-reduction roasting process relates to the field of secondary resource recycling. First, the copper slag is ground and mixed with a slagging agent, and then fed into a smelting furnace for selective oxidation smelting. This allows the gaseous oxidant to undergo a selective oxidation reaction with the tin, lead, copper, arsenic, and sulfur phases in the copper slag. Tin, arsenic, and copper enter the slag phase, while most of the lead is distributed in the crude lead phase. Sulfur is oxidized to sulfur dioxide and enters the gas phase, achieving selective oxidation slag formation and efficient separation of tin and copper. The smelting slag and the reducing agent are then evenly mixed in a certain proportion and fed into a rotary kiln for reduction roasting. Tin, arsenic, copper, and other components in the smelting slag undergo simultaneous reduction reactions. Arsenic is reduced to arsenic oxide, which volatilizes and enriches in the smoke, while tin and copper remain in the roasting slag. Reduction smelting produces a copper-tin alloy, achieving efficient recovery of tin and copper resources from the copper slag. The method of the present invention has the advantages of simple operation and low cost, and has good prospects for promotion and application.

Description

[0001] The invention belongs to the technical field of secondary resource recycling and relates to a method for recovering tin and copper resources in copper slag by a two-step process of gas phase oxidation-reduction roasting. Background Art

[0002] Copper dross is a byproduct formed during the copper removal process of crude lead pyrometallurgical refining. It is primarily composed of a mixture of impurity compounds and metal droplets entrained during slag removal. The morphology and composition of copper dross vary significantly depending on the slag removal method and equipment used. Lead and copper are the primary components of copper dross, along with smaller amounts of elements such as tin, arsenic, antimony, bismuth, and zinc. Lead sulfide, lead oxide, lead, copper sulfide, copper-arsenic alloys, copper-tin alloys, and lead-tin alloys are the primary phases of copper dross. From a metal recovery perspective, copper dross is a critical secondary resource in the lead smelting process. Failure to recycle it not only results in a waste of metal resources, but the heavy metals contained within it pose a serious threat to the environment and human health. Currently, copper dross treatment methods are primarily divided into wet processes and pyrometallurgical processes. Wet processes are categorized into acid leaching, ammonia leaching, and alkaline leaching, depending on the leaching agent. Acid leaching primarily uses acidic solutions such as sulfuric acid or hydrochloric acid to leach valuable metals, followed by purification through solvent extraction or electrowinning. This method is suitable for treating oxide slag and is effective for low-grade slag. However, it requires large amounts of acidic reagents, is prone to corrosion of equipment, and produces difficult-to-treat wastewater. Ammonia leaching primarily targets copper recovery from copper slag. It dissolves copper from the slag using copper-ammonia complex ions in ammonia and a stripping solution. This method includes leaching, extraction, stripping, and the production of cathode copper or copper sulfate crystals. While the process is stable, efficient, and minimizes mechanical losses, it also produces significant environmental pollution and high costs. Alkaline leaching, based on the principle that lead dissolves in alkaline solutions while copper does not, achieves copper separation and recovery. This method is also primarily used to treat slag containing impurities such as arsenic and antimony. Alkaline solutions (such as sodium hydroxide) can selectively leach these elements, preserving valuable metals. This method is suitable for raw materials with complex compositions and demonstrates good selective leaching results. However, this method is lengthy, requires high investment, and requires stringent technical skills, making it unsuitable for small and medium-sized enterprises. Currently, domestic and international factories primarily use pyrometallurgical processes to treat copper slag. Pyrometallurgical processes include the smelting method and the soda-iron method. The smelting method involves mixing copper slag with a flux (such as limestone or silica) in a specific proportion and then smelting it at high temperature. This adjusts the composition of the slag, promoting the enrichment and separation of valuable metals from the slag. This method consumes a lot of energy and is prone to secondary pollution. It cannot effectively recover the tin resources in the copper slag, and the recovered metals require further separation and recycling. The soda-iron method is currently the main process for treating copper slag. The soda-iron method involves smelting copper slag, coke, iron filings, and soda ash in a specific proportion to produce crude lead and matte.The addition of soda ash primarily lowers the melting points of slag and matte, forming copper matte and reducing the lead content in the matte and slag. Furthermore, soda ash can convert arsenic and antimony into sodium salts to form slag, removing some of the arsenic and antimony. Iron filings primarily reduce lead sulfide in copper slag to metallic lead, forming a crude lead phase. This, in turn, concentrates copper in the matte, improving the direct lead recovery rate. Coke is added to maintain a reducing atmosphere within the system, preventing sulfide oxidation and reducing lead oxide. While traditional copper slag treatment processes offer advantages such as high lead recovery rates, wide raw material adaptability, and mature processes, copper is primarily concentrated in the matte phase, containing significant amounts of copper, arsenic, and sulfur, hindering subsequent copper recovery. Furthermore, this process is unable to effectively recover tin from copper slag. With recent technological advancements, numerous researchers have proposed novel treatment methods for copper slag, such as vacuum metallurgy processes using vacuum-enhanced distillation and fractional condensation based on the boiling point differences between different metal compounds, and oxygen-enhanced smelting with the addition of varying proportions of pyrite. However, these methods are mainly aimed at the recovery of resources such as lead, copper and precious metals, and there is very little research on the recovery of tin resources and the separation of arsenic in copper slag. Summary of the Invention

[0003] In response to the above-mentioned problems, the present invention provides a method for recovering tin and copper resources from copper slag using a two-step gas-phase oxidation-reduction roasting method. The method fully utilizes the differences in the redox abilities of tin, lead, copper, arsenic, and sulfur phases in the copper slag, selectively oxidizes and converts the tin, lead, copper, arsenic, and sulfur phases using a gaseous oxidant as a medium, and efficiently enriches tin, arsenic, and copper in the slag. Arsenic is then separated by a subsequent slag-phase low-temperature reduction roasting method, and a copper-tin alloy is subsequently produced by conventional reduction smelting, thereby achieving efficient recovery of tin and copper resources from the copper slag. The method has the advantages of simple operation, low cost, and good prospects for promotion and application.

[0004] The purpose of the present invention can be achieved by the following technical solutions:

[0005] (1) Grinding the copper slag and slagging agent to a certain particle size and uniformly mixing them in a certain proportion;

[0006] (2) The mixed material is fed into a smelting furnace, and then a gaseous oxidant is introduced for selective oxidation smelting. During the process, key parameters such as smelting temperature, smelting time, and oxygen concentration are controlled so that the gaseous oxidant reacts with the tin, lead, copper, arsenic, and sulfur phases in the copper slag to undergo selective oxidation reactions. Tin, arsenic, and copper enter the slag phase, while most of the lead is distributed in the crude lead phase. Sulfur is oxidized into sulfur dioxide and enters the gas phase. The flue gas generated in the process is discharged into the atmosphere after waste heat recovery and flue gas treatment to meet standards.

[0007] (3) The resulting smelting slag and reducing agent are uniformly mixed in a certain proportion and then placed in a rotary kiln for reduction roasting. Key parameters such as the amount of reducing agent added, roasting temperature, and roasting time are controlled. During this process, the tin, arsenic, and copper components of the smelting slag undergo simultaneous reduction reactions. Arsenic is reduced to arsenic oxide, which volatilizes and accumulates in the smoke, while tin and copper remain in the roasting slag. The flue gas generated during the process is discharged into the atmosphere after waste heat recovery and flue gas treatment to meet standards.

[0008] (4) The resulting roasted slag and reducing agent are evenly mixed in a certain proportion and then placed in a reduction furnace for traditional reduction smelting. Key parameters such as the amount of reducing agent added, smelting temperature, and smelting time are controlled. During this process, the copper and tin oxides in the roasted slag are reduced to form a copper-tin alloy, achieving efficient recovery of tin and copper resources. The flue gas generated during the process is discharged into the atmosphere after waste heat recovery and flue gas treatment to meet standards.

[0009] Furthermore, the grinding particle size of the copper slag and the slag forming agent in step (1) is 20-300 mesh, and the tin and copper contents in the copper slag are between 0.1-6wt% and 10-30wt%.

[0010] Furthermore, in step (1), the slagging agent is iron oxide, and the addition amount is 2%-10% of the copper slag.

[0011] Furthermore, the oxygen concentration of the gaseous oxidant in step (2) is 15-30%.

[0012] Furthermore, in step (2), the selective oxidation smelting temperature is 1050-1300° C., and the smelting time is 75-180 minutes.

[0013] Furthermore, the reducing agent in step (3) is a carbon-based reducing agent such as coal powder, coke, waste cathode carbon, biochar, heavy oil, or a gaseous reducing agent such as carbon monoxide and natural gas. When it is a non-gaseous reducing agent, its addition amount is between 5% and 15% of the tin smelting slag. When it is a gaseous reducing agent, the concentration of carbon monoxide and natural gas in the gas phase is ≥30%.

[0014] Furthermore, the reduction smelting temperature in step (3) is 500-700°C, and the smelting time is 90-150 minutes.

[0015] Compared with the prior art, the present invention has the following characteristics:

[0016] 1) The present invention fully utilizes the differences in the redox capabilities of tin, lead, copper, arsenic, and sulfur phases in copper slag, selectively oxidizing and converting these phases using a gaseous oxidant, thereby efficiently enriching tin, arsenic, and copper in the slag. Arsenic is then separated through a subsequent low-temperature reduction roasting process of the slag phase, and a copper-tin alloy is subsequently produced using conventional reduction smelting. This ensures efficient recovery of lead while also achieving efficient recovery of tin and copper, resulting in outstanding resource efficiency.

[0017] 2) The present invention does not require the additional addition of iron powder and sodium carbonate, which greatly reduces the process cost, has good economic efficiency, is simple to operate, and has good prospects for promotion and application. BRIEF DESCRIPTION OF THE DRAWINGS

[0018] Figure 1 It is a process flow chart of the present invention. DETAILED DESCRIPTION

[0019] The present invention is described in detail below with reference to the accompanying drawings and specific embodiments. This embodiment is implemented based on the technical solution of the present invention, and provides a detailed implementation method and specific operation process, but the protection scope of the present invention is not limited to the following embodiments.

[0020] Example 1:

[0021] Combine Figure 1 The method for recovering tin and copper resources from copper slag by the two-step gas phase oxidation-reduction roasting method is specifically implemented as follows:

[0022] The copper slag from a lead smelter was used as the treatment object. Its lead content was 47.56%, copper content was 19.02%, tin content was 1.33%, arsenic content was 8.11%, and sulfur content was 1.46%. The above contents are all mass contents.

[0023] The processing flow is:

[0024] In the oxidation smelting stage, oxygen-enriched air is used as the oxidant, and iron oxide is used as the slagging agent. The iron oxide is ground to a 100-mesh particle size. Copper slag (ground to a 60-mesh particle size) and iron oxide are uniformly mixed and then placed in a reverberatory furnace, where oxygen-enriched air is introduced for oxidation smelting. The iron oxide addition level is controlled at 3 wt% (relative to the copper slag), the oxygen concentration in the gas is maintained at 17%, the smelting temperature is 1250°C, and the smelting time is 90 minutes. After the oxidation smelting is completed, crude lead and smelting slag are produced. The crude lead is discharged from the lead outlet of the reverberatory furnace, and the smelting slag is discharged from the slag outlet of the reverberatory furnace. The flue gas generated during the process is recovered through waste heat and treated to meet exhaust gas standards before being discharged into the atmosphere. In the low-temperature reduction roasting stage, the smelting slag is subjected to low-temperature reduction roasting in a rotary kiln using coke as the reducing agent. The coke addition level is controlled at 9% (relative to the slag), the reduction roasting temperature is controlled at 600°C, and the roasting time is controlled at 90 minutes. Arsenic-containing fumes emitted during the roasting process are collected and recovered by a dust collector. The calcined slag produced after the reduction roasting is discharged through the rotary kiln head. The flue gas generated during the process undergoes waste heat recovery and tail gas treatment to meet standards before being discharged into the atmosphere. The resulting calcined slag is mixed evenly with coke and then fed into a reduction furnace for traditional reduction smelting. The coke addition is controlled at 1200°C for a 90-minute reduction smelting time. The copper-tin alloy produced after the reduction smelting is discharged through the converter's molten metal outlet, achieving efficient recovery of tin and copper. The flue gas generated during the process undergoes waste heat recovery and tail gas treatment to meet standards before being discharged into the atmosphere. Chemical analysis of the oxidizing smelting slag and the reducing roasting slag revealed that the tin, arsenic, and copper contents in the oxidizing slag can be enriched to 3.97wt%, 18.54wt%, and 46.29wt%, respectively. The first stage of oxidation smelting allows for efficient enrichment of tin, arsenic, and copper from copper slag. The arsenic content in the reduction roasting slag is reduced to 0.76wt%, while the tin and copper contents reach 4.73wt% and 56.49wt% respectively. The second stage of low-temperature reduction roasting allows for efficient separation of arsenic from the smelting slag. After reduction smelting, the copper and tin contents in the smelting slag drop to 0.26wt% and 0.14wt%, respectively, achieving efficient recovery of tin and copper from the copper slag.

[0025] Example 2

[0026] Combine Figure 1 The method for recovering tin and copper resources from copper slag by the two-step gas phase oxidation-reduction roasting method is specifically implemented as follows:

[0027] The copper slag from a lead smelter was used as the treatment object. Its lead content was 54.08%, copper content was 18.19%, tin content was 0.93%, arsenic content was 9.77%, and sulfur content was 1.48%. The above contents are all mass contents.

[0028] The process flow is as follows: In the oxidation smelting stage, air is used as the oxidant and iron oxide as the slagging agent. The iron oxide is ground to an 80-mesh particle size. Copper slag (100-mesh particle size) and iron oxide are uniformly mixed and placed in a reverberatory furnace, where air is introduced for oxidation smelting. During the process, the iron oxide addition is controlled at 3.5wt%, the oxygen concentration in the gas is maintained at 19%, the smelting temperature is 1300°C, and the smelting time is 110 minutes. After the oxidation smelting is completed, crude lead and smelting slag are produced. The crude lead is discharged from the lead outlet of the reverberatory furnace, and the smelting slag is discharged from the slag outlet of the reverberatory furnace. The flue gas generated during the process is recovered through waste heat and treated to meet exhaust gas standards before being discharged into the atmosphere. In the low-temperature reduction roasting stage, a rotary kiln is used as the roasting furnace, and carbon monoxide is used as the reducing agent. The carbon monoxide concentration is controlled at 35%, the roasting temperature is 550°C, and the roasting time is 120 minutes. Arsenic-containing fumes emitted during the roasting process are collected and recovered by a dust collector. After the reduction roasting is complete, the slag is discharged through the rotary kiln head. The flue gas generated during the process is discharged into the atmosphere after waste heat recovery and exhaust gas treatment to meet standards. The resulting slag is evenly mixed with coke and then fed into the reduction furnace for traditional reduction smelting. The coke addition is controlled at 12% and the reduction smelting temperature is 1200°C for 90 minutes. The copper-tin alloy produced after the reduction smelting is discharged through the converter metal outlet, achieving efficient recovery of tin and copper. The flue gas generated during the process is discharged into the atmosphere after waste heat recovery and exhaust gas treatment to meet standards. Chemical analysis of the oxidation smelting slag and the reduction roasting slag revealed that the tin, arsenic, and copper contents in the oxidation smelting slag could be enriched to 2.68wt%, 21.74wt%, and 48.31wt%, respectively. This indicates that the first-stage oxidation smelting process allows for the efficient enrichment of tin, arsenic, and copper from copper dross. The arsenic content in the reduction roasting slag was reduced to 0.83wt%, while the tin and copper contents reached 3.51wt% and 57.22wt%, respectively. This indicates that the second-stage low-temperature reduction roasting process allows for the efficient separation of arsenic from the smelting slag. After reduction smelting, the copper and tin contents in the smelting slag were reduced to 0.58wt% and 0.19wt%, respectively, achieving efficient recovery of tin and copper from the copper dross.

[0029] Example 3

[0030] Combine Figure 1 The method for recovering tin and copper resources from copper slag by the two-step gas phase oxidation-reduction roasting method is specifically implemented as follows:

[0031] The copper slag from a lead smelter was used as the treatment object. Its lead content was 48.82%, copper content was 21.73%, tin content was 1.12%, arsenic content was 10.09%, and sulfur content was 2.19%. The above contents are all mass contents.

[0032] The process flow is as follows: In the oxidation smelting stage, air is used as the oxidant and iron oxide as the slagging agent. The iron oxide is ground to a 50-mesh particle size. The copper slag (40-mesh particle size) and iron oxide are evenly mixed before being placed in a reverberatory furnace and oxidized by air. The iron oxide addition is controlled at 4%, the oxygen concentration in the gas is maintained at 19%, the smelting temperature is 1250°C, and the smelting time is 90 minutes. After the oxidation smelting is completed, crude lead and smelting slag are produced. The crude lead is discharged from the lead outlet of the reverberatory furnace, and the smelting slag is discharged from the slag outlet of the reverberatory furnace. The flue gas generated during the process is recovered through waste heat and treated to meet exhaust gas standards before being discharged into the atmosphere. In the low-temperature reduction roasting stage, the smelting slag undergoes low-temperature reduction roasting in a rotary kiln using natural gas as the reducing agent. The natural gas partial pressure in the gas phase is controlled at 32%, the roasting temperature is 600°C, and the roasting time is 120 minutes. The slag produced after reduction roasting is discharged through the rotary kiln head, and the flue gas generated during the process is discharged into the atmosphere after waste heat recovery and tail gas treatment meet standards. The resulting slag is evenly mixed with coke and then placed in a reduction furnace for traditional reduction smelting. During the process, the coke addition is controlled at 12%, the reduction smelting temperature is 1200°C, and the smelting time is 90 minutes. The copper-tin alloy produced after the reduction smelting is discharged through the converter metal outlet, achieving efficient recovery of tin and copper. The flue gas generated during the process is discharged into the atmosphere after waste heat recovery and tail gas treatment meet standards. Chemical analysis of the oxidation smelting slag and the reduction roasting slag revealed that the tin, arsenic, and copper contents in the oxidation smelting slag could be enriched to 3.06wt%, 22.19wt%, and 47.71wt%, respectively. This indicates that the first-stage oxidation smelting process allows for the efficient enrichment of tin, arsenic, and copper from copper dross. The arsenic content in the reduction roasting slag was reduced to 0.92wt%, while the tin and copper contents reached 3.89wt% and 58.17wt%, respectively. This indicates that the second-stage low-temperature reduction roasting process allows for the efficient separation of arsenic from the smelting slag. After reduction smelting, the copper and tin contents in the smelting slag were reduced to 0.33wt% and 0.16wt%, respectively, enabling the efficient recovery of tin and copper from the copper dross.

[0033] The above description of the embodiments is intended to facilitate understanding and use of the invention by those skilled in the art. It will be apparent that those skilled in the art can readily make various modifications to these embodiments and apply the general principles described herein to other embodiments without requiring inventive effort. Therefore, the present invention is not limited to the above-described embodiments. Improvements and modifications made by those skilled in the art based on the disclosure of the present invention without departing from the scope of the present invention should be within the scope of protection of the present invention.

Claims

1. A method for recovering tin and copper resources from copper slag by a two-step gas phase oxidation-reduction roasting method, characterized in that the steps include: (1) grinding the copper slag and the slag-forming agent to a certain particle size and uniformly mixing them in a certain proportion; in step (1), the slag-forming agent is iron oxide, and the addition amount is 2%-10% of the copper slag; (2) The mixed material is fed into a smelting furnace, and then a gaseous oxidant is introduced for selective oxidation smelting. During the process, the key parameters of smelting temperature, smelting time, and oxygen concentration are controlled so that the gaseous oxidant reacts with the tin, lead, copper, arsenic, and sulfur phases in the copper slag to selectively oxidize the material. Tin, arsenic, and copper enter the slag phase, most of the lead is distributed in the crude lead phase, and sulfur is oxidized to sulfur dioxide and enters the gas phase. The flue gas generated during the process is discharged into the atmosphere after waste heat recovery and flue gas treatment to meet the standards; (3) The obtained smelting slag and reducing agent are uniformly mixed in a certain proportion, and then put into a rotary kiln for reduction roasting. The key parameters of reducing agent addition, roasting temperature and roasting time are controlled. During the process, the tin, arsenic and copper components in the smelting slag undergo reduction reactions simultaneously. Arsenic is reduced to arsenic oxide and volatilized and enriched in the smoke, while tin and copper are retained in the roasting slag. The flue gas generated in the process is discharged into the atmosphere after waste heat recovery and flue gas treatment to meet the standards; (4) The obtained roasting slag and reducing agent are evenly mixed in a certain proportion and then put into a reduction furnace for traditional reduction smelting. The key parameters of reducing agent addition, smelting temperature and smelting time are controlled. During the process, the copper and tin oxides in the roasting slag will be reduced to form copper-tin alloy, realizing the efficient recovery of tin and copper resources; the flue gas generated in the process is discharged into the atmosphere after waste heat recovery and flue gas treatment to meet the standards.

2. The method for recovering tin and copper resources from copper slag by a two-step gas phase oxidation-reduction roasting process according to claim 1, characterized in that: The grinding particle size of the copper slag and slag forming agent in the step (1) is 20-300 meshes.

3. The method for recovering tin and copper resources from copper slag by a two-step gas phase oxidation-reduction roasting process according to claim 1, characterized in that: The contents of tin and copper in the copper slag in step (1) are 0.1-6 wt% and 10-30 wt% respectively.

4. The method for recovering tin and copper resources from copper slag by a two-step gas phase oxidation-reduction roasting process according to claim 1, characterized in that: The oxygen concentration of the gaseous oxidant in step (2) is 15-30%.

5. The method for recovering tin and copper resources from copper slag by a two-step gas phase oxidation-reduction roasting process according to claim 1, characterized in that: In the step (2), the oxidation smelting temperature is 1050-1300° C., and the smelting time is 75-180 minutes.

6. The method for recovering tin and copper resources from copper slag by a two-step gas phase oxidation-reduction roasting process according to claim 1, characterized in that: In step (3), the reducing agent is coal powder, coke, waste cathode carbon, biomass charcoal, heavy oil carbon-based reducing agent or carbon monoxide, natural gas gas phase reducing agent. When the reducing agent is in non-gaseous phase, its addition amount is between 5% and 15% of the tin smelting slag. When the reducing agent is in gaseous phase, the concentration of carbon monoxide and natural gas in the gas phase is ≥30%.

7. The method for recovering tin and copper resources from copper slag by a two-step gas phase oxidation-reduction roasting method according to claim 1, characterized in that: The reduction roasting temperature in step (3) is 500-700° C., and the roasting time is 90-150 minutes.

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