Hydrometallurgical- pyrometallurgical process for the recovery of valuable metals from copper dust

Abstract

A hydrometallurgy-flame metallurgy combined process for recovering valuable metals from copper soot, comprising the following steps: (1) low-temperature sulfuric acid roasting and dehalogenation: mixing copper smelting soot containing bromine and chlorine with concentrated sulfuric acid, and then performing low-temperature roasting to obtain roasting slag; (2) leaching and separation: adding the roasting slag into water, stirring and leaching, and then filtering to obtain tin, lead, gold and silver-containing leaching residue and copper and zinc-containing leaching solution; (3) reduction smelting: adding carbonaceous reducing agent into the tin, lead, gold and silver-containing leaching residue, and then performing reduction smelting to obtain gold and silver-containing lead-tin alloy product; and (4) leaching solution treatment: adding elemental zinc into the copper and zinc-containing leaching solution, performing zinc replacement, and then filtering to obtain sponge copper product and zinc replacement solution. The process can realize efficient removal of bromine and chlorine in copper smelting soot, reduce environmental pollution, realize efficient separation of tin from copper and zinc, has high tin direct recovery rate, has simple process route, has high metal recovery rate, and is suitable for industrial application.

Description

Technical Field

[0001] This invention relates to a metallurgical process for recovering valuable metals, specifically a combined hydrometallurgical-pyrometallurgical process for recovering valuable metals from copper flue dust. Background Technology

[0002] The ash produced during copper smelting contains various valuable metals, such as tin, lead, copper, zinc, gold, and silver, as well as high levels of bromine and chlorine. Traditional treatment methods often involve direct pyrometallurgical processes, which suffer from low metal recovery rates, high energy consumption, and severe environmental pollution. In existing hydrometallurgical processes, the removal of bromine and chlorine and the efficient separation of valuable metals remain key technical challenges.

[0003] CN117660752A discloses a method for the recovery and harmless disposal of copper smelting flue dust. This method uses sulfation roasting and water leaching processes to obtain a copper-zinc containing solution, which is then further extracted by extraction and electrowinning to extract copper and zinc. However, when extracting copper and zinc from copper flue dust produced by traditional copper ore smelting, this method does not adequately consider the recovery of tin from tin-containing copper flue dust produced by secondary copper resource smelting, which can easily lead to a waste of tin resources.

[0004] CN109055722A discloses a method for separating bromine by roasting sodium nitrate in flue gas from circuit board co-smelting. This method uses sodium nitrate roasting to oxidize cuprous bromide and solidify bromine, thereby achieving the conversion and separation of bromine, lead, and tin-soluble substances. However, the resulting products, such as crude bromide, crude zinc sulfate, and crude lead sulfate, are all semi-finished products and require further processing.

[0005] CN109735700A discloses a method for recovering copper and zinc elements from copper flue dust using microwave reduction roasting-sulfuric acid leaching. This method involves microwave reduction roasting, sulfuric acid leaching of copper and zinc, zinc powder replacement, and electrolytic deposition of zinc sulfate to recover copper and zinc. However, this method has high energy consumption during the reduction process, and existing microwave reduction equipment has a small processing capacity and requires significant investment.

[0006] CN107523694A discloses a method for enhanced leaching of copper smelting ash through roasting. This method effectively leaches copper and zinc from copper smelting ash containing a high amount of metal sulfides through a process of primary sulfuric acid leaching, oxidative roasting, and secondary sulfuric acid leaching. However, the process involves oxidative roasting of the primary leaching residue followed by secondary leaching, repeatedly mixing pyrometallurgical and hydrometallurgical processes, resulting in a cumbersome and time-consuming operation.

[0007] In summary, there is an urgent need to find a hydrometallurgical-pyrometallurgical combined process for recovering valuable metals from copper smelting flue dust that can achieve efficient removal of bromine and chlorine from copper smelting flue dust, reduce environmental pollution, achieve efficient separation of tin from copper and zinc, have a high direct recovery rate of tin, a simple process route, a high metal recovery rate, and be suitable for industrial applications. Summary of the Invention

[0008] The technical problem to be solved by the present invention is to overcome the above-mentioned defects of the prior art and provide a hydrometallurgical-pyrometallurgical combined process for recovering valuable metals from copper smelting flue dust, which can achieve efficient removal of bromine and chlorine from copper smelting flue dust, reduce environmental pollution, achieve efficient separation of tin from copper and zinc, have a high direct recovery rate of tin, a simple process route, a high metal recovery rate, and is suitable for industrial applications.

[0009] The technical solution adopted by this invention to solve its technical problem is as follows: a hydrometallurgical-pyrometallurgical combined process for recovering valuable metals from copper flue dust, comprising the following steps: (1) Low-temperature sulfuric acid roasting dehalogenation: copper smelting flue dust containing bromine and chlorine is mixed with concentrated sulfuric acid and then roasted at low temperature to obtain roasting residue; (2) Leaching and separation: The roasted residue obtained in step (1) is added to water, stirred and leached, filtered to obtain leaching residue containing tin, lead, gold and silver and leaching solution containing copper and zinc; (3) Reduction smelting: Add carbonaceous reducing agent to the leaching residue containing tin, lead, gold and silver obtained in step (2) and carry out reduction smelting to obtain a lead-tin alloy product containing gold and silver. (4) Leachate treatment: Add elemental zinc to the copper and zinc leaching solution obtained in step (2) to perform zinc replacement, filter, and obtain sponge copper product and zinc replacement solution.

[0010] The inventive concept of this invention is as follows: Addressing the problems existing in the treatment of copper smelting flue dust containing bromine and chlorine in existing technologies, this invention aims to solve the problem of efficient direct recovery of valuable metals. To achieve a short-process treatment of copper smelting flue dust, it proposes a combined wet-pyrometallurgical process of "roasting-leaching-reduction smelting" to separate copper, lead, tin, zinc, gold, and silver from the flue dust. First, low-temperature sulfuric acid roasting dehalogenates the flue dust, effectively removing chlorine and bromine through copper persulfate roasting, converting copper and zinc into easily soluble zinc sulfate and copper sulfate. Further leaching allows copper and zinc to enter the leaching solution, while tin, lead, gold, and silver remain in the leaching residue, achieving efficient separation. Then, reduction smelting yields a lead-tin alloy containing gold and silver. Finally, zinc replacement yields sponge copper, forming a new process for economical, efficient, and green recovery of valuable metals from flue dust.

[0011] Preferably, in step (1), the mass ratio of the bromine- and chlorine-containing copper smelting flue dust to concentrated sulfuric acid is 1:0.2 to 1.0 (more preferably 1:0.3 to 0.8, and even more preferably 1:0.4 to 0.6). The function of concentrated sulfuric acid is to convert copper and zinc into easily soluble zinc sulfate and copper sulfate. If the amount of concentrated sulfuric acid used is too small, copper and zinc will not be fully converted, resulting in losses; if the amount of concentrated sulfuric acid used is too large, it will easily cause high pressure on the subsequent acidic wastewater treatment.

[0012] Preferably, in step (1), the main components and mass content of the bromine- and chlorine-containing copper smelting flue dust are: Cu 5-15%, Sn 1-5%, Pb 5-20%, Au 2-10 g / t, Ag 100-2000 g / t, Br 1-5%, Cl 5-10%. The bromine- and chlorine-containing copper smelting flue dust used in this invention is produced by smelting waste circuit boards.

[0013] Preferably, in step (1), the mass concentration of the concentrated sulfuric acid is 95-98%. During the aging process using concentrated sulfuric acid, a certain amount of heat is released, which is beneficial to improving the reaction rate and reaction efficiency.

[0014] Preferably, in step (1), the low-temperature roasting temperature is 200–300°C, and the time is 0.5–3.0 h (more preferably 1–2 h). When the temperature is above 300°C, a large amount of tin will be sulfated, and thus enter the leaching solution along with copper and zinc during the leaching stage, resulting in incomplete separation of tin from copper and zinc; when the temperature is below 200°C, a large amount of copper is difficult to completely enter the leaching solution during the leaching stage. The roasting tail gas is treated by alkaline absorption.

[0015] Preferably, in step (2), the solid-liquid ratio of the roasted residue to water is 1:3 to 5. Using a neutral or weakly acidic solution as the leaching agent, if the liquid-to-solid ratio is too low, the copper-zinc solution will easily become saturated, resulting in a decrease in the copper-zinc leaching rate. If the liquid-to-solid ratio is too high, it will cause high pressure in the subsequent wastewater treatment process.

[0016] Preferably, in step (2), the stirring leaching temperature is room temperature, the stirring speed is 400-800 r / min, the time is 0.5-3.0 h (more preferably 1-2 h), and the final pH value is 6-7. Under the stirring leaching conditions, the reaction rate can be increased while reducing stirring energy consumption.

[0017] Preferably, in step (3), the amount of the carbonaceous reducing agent is equivalent to 8-12% of the leaching residue containing tin, lead, gold, and silver.

[0018] Preferably, in step (3), the carbonaceous reducing agent includes coke and / or coal, etc.

[0019] Preferably, in step (3), the reduction smelting temperature is 1300–1600℃ (more preferably 1400–1550℃), and the time is 1–2 hours. During the reduction smelting process, the liquid metallic lead and tin reduced from the lead and tin compounds dissolve in each other at high temperatures to form a lead-tin alloy. Due to the trapping of precious metals by lead, gold and silver can simultaneously enter the lead-tin alloy. Under the reduction smelting conditions, it is beneficial to ensure the efficient reduction of tin and lead, promote the separation of impurities, and maintain the fluidity of the slag during the smelting process.

[0020] Preferably, in step (3), the equipment for reduction smelting includes an electric arc furnace or a blast furnace, etc.

[0021] Preferably, in step (4), the molar ratio of copper to elemental zinc in the copper-zinc leaching solution is 1:1.1 to 1.2. Based on the stoichiometric ratio of zinc to copper sulfate in the reaction being 1:1, a slight excess of zinc helps to ensure complete copper replacement.

[0022] Preferably, in step (4), the zinc replacement temperature is 50–70°C, and the time is 20–30 min. At this temperature, a sufficient reaction rate is ensured while avoiding excessively high temperatures that could lead to intensified side reactions, such as the violent reaction of zinc with acid producing a large amount of hydrogen gas, and also preventing energy waste. During the replacement process, elemental zinc (Zn)... 0 It loses electrons and is oxidized to zinc ions (Zn). 2+ ) enters the solution, while copper ions (Cu) in the solution 2+ The electrons are gained and reduced to metallic copper (Cu). 0 It will precipitate out.

[0023] Preferably, in step (4), an alkaline agent is added to the obtained zinc replacement solution to carry out a precipitation reaction, thereby obtaining zinc carbonate product.

[0024] Preferably, the molar ratio of zinc to alkali in the zinc replacement solution is 0.8–0.9:1. This ratio helps ensure complete zinc precipitation.

[0025] Preferably, the precipitation reaction is carried out at a temperature of room temperature to 60°C for 20 to 30 minutes. The precipitation reaction is a chemical precipitation and crystallization process aimed at removing dissolved zinc (Zn). 2+ The zinc carbonate product is converted into a solid state for recycling. Higher temperatures promote the formation of denser, more easily filtered zinc carbonate precipitates, such as basic zinc carbonate, and accelerate the reaction rate; however, excessively high temperatures may increase energy consumption. Limiting the precipitation reaction time is to prevent the endpoint pH from becoming too high, thus avoiding the conversion of zinc carbonate to zinc hydroxide or the formation of more difficult-to-filter colloids.

[0026] Preferably, the alkaline agent includes sodium carbonate and / or sodium bicarbonate, etc.

[0027] The beneficial effects of the process of this invention are as follows: (1) The process of the present invention can achieve efficient removal of bromine and chlorine from copper smelting flue dust, thereby reducing environmental pollution; (2) The process of the present invention achieves efficient separation of tin from copper and zinc through roasting-leaching control, thereby improving the direct recovery rate of tin; (3) The process route of the present invention is simple, the metal recovery rate is high, and it is suitable for industrial application. Detailed Implementation

[0028] The present invention will be further described below with reference to the embodiments.

[0029] The copper smelting flue dust containing bromine and chlorine used in the embodiments and comparative examples of this invention is produced by smelting waste circuit boards. The copper and zinc components in the flue dust exist in the form of soluble chloride salts, bromine salts, and oxides, which can be directly leached under water-soluble conditions. The remaining metals exist mostly in the form of insoluble oxides and alloy salts. The characteristics of this component directly affect the amount of roasted material, the amount of leaching residue, and the metal distribution pattern. The main components and mass content of copper smelting flue dust 1 containing bromine and chlorine are: Cu 5.65%, Sn 2.79%, Pb 5.10%, Zn 23.51%, Au 5.2 g / t, Ag 609.7 g / t, Br 1.12%, Cl 5.86%; the main components and mass content of copper smelting flue dust 2 containing bromine and chlorine are: Cu 11.67%, Sn 2.68%, Pb 18.95%, Zn 12.41%, Au 8.2 g / t, Ag 1135.6 g / t, Br 2.33%, Cl 7.01%; the raw materials or chemical reagents used in the embodiments and comparative examples of this invention, unless otherwise specified, were obtained through conventional commercial means.

[0030] Example 1 (1) Low-temperature sulfuric acid roasting dehalogenation: 1000g of copper smelting flue dust containing bromine and chlorine was mixed with 450g of concentrated sulfuric acid with a mass concentration of 98% and placed in a roasting furnace. The mixture was roasted at 200℃ for 2 hours to obtain 1085g of roasting residue (Br and Cl mass contents of 0.06% and 0.13%, respectively). The amount of concentrated sulfuric acid used can fully meet the requirements of halide decomposition and copper-zinc component sulfation in the flue dust. There is no problem of excessive sulfuric acid waste or incomplete reaction. It is suitable for low-temperature roasting process conditions. (2) Leaching and separation: 1085g of roasted residue obtained in step (1) was added to 5425mL of water (the solid-liquid ratio of roasted residue to water was 1:5). At room temperature and a stirring speed of 600r / min, the mixture was stirred and leached for 2.0h until the final pH value was 6.5. After filtration, leaching residue containing tin, lead, gold and silver and 5.40L of leaching solution containing copper and zinc were obtained (the concentrations of Cu, Zn, Cl and Sn were 9.85g / L, 43.48g / L, 0.52g / L and 0.0020g / L, respectively, and Au and Ag were not detected). (3) Reduction smelting: Add coke equivalent to 10% of its mass to the tin, lead, gold and silver leaching residue obtained in step (2), and send it into an electric arc furnace to carry out reduction smelting at 1430℃ for 1.0h to obtain a lead-tin alloy product containing gold and silver (the mass content of lead and tin is 61.58% and 33.68%, respectively). (4) Leachate treatment: In the copper and zinc leaching solution obtained in step (2), elemental zinc is added at a molar ratio of copper to elemental zinc of 1:1.15. Zinc replacement is carried out at 60°C for 25 min. After filtration, sponge copper product (Cu mass content is 92.11%) and zinc replacement solution are obtained. In a zinc replacement solution, sodium carbonate is added at a molar ratio of zinc to sodium carbonate of 0.85:1, and the precipitation reaction is carried out at room temperature for 25 minutes to obtain zinc carbonate product.

[0031] According to the test, in step (1), the removal rates of Br and Cl in the obtained roasted residue were 94.19% and 97.59%, respectively.

[0032] Calculations show that in step (2), the leaching rates of Cu, Zn, and Sn in the copper- and zinc-containing leachate are 94.14%, 99.87%, and 0.04%, respectively. This indicates that under the low-temperature sulfuric acid roasting conditions of 200 ℃, copper and zinc are fully sulfatized and enter the leachate, while tin, gold, and silver are mainly retained in the leaching residue, which is beneficial for the subsequent capture of gold and silver by the lead-tin alloy.

[0033] According to the test, in step (3), the direct recovery rate of lead and tin in the obtained lead-tin alloy product is 93.12% and 93.08%, respectively; in step (4), the direct recovery rate of Cu in the obtained sponge copper product is 94.03%.

[0034] Example 2 (1) Low-temperature sulfuric acid roasting dehalogenation: 1000g of copper smelting flue dust 2 containing bromine and chlorine was mixed with 420g of concentrated sulfuric acid with a mass concentration of 98% and placed in a roasting furnace. The mixture was roasted at 300℃ for 2h to obtain 1068g of roasting residue (Br and Cl mass contents of 0.08% and 0.14%, respectively). Based on the characteristics of high copper, high lead and high halogen components of flue dust 2, the amount of concentrated sulfuric acid was optimized to ensure efficient removal of halides and complete sulfation of copper and zinc. (2) Leaching and separation: 1068g of calcined residue obtained in step (1) was added to 4272mL of water (the solid-liquid ratio of calcined residue to water was 1:4). The mixture was stirred and leached for 2.0h at room temperature and a stirring speed of 700r / min until the final pH value was 6.5. The mixture was then filtered to obtain leaching residue containing tin, lead, gold and silver and 4.315L of leaching solution containing copper and zinc (the concentrations of Cu, Zn, Cl and Sn were 25.54g / L, 28.75g / L, 0.61g / L and 0.0014g / L, respectively, and Au and Ag were not detected). (3) Reduction smelting: Add coke equivalent to 10% of its mass to the tin, lead, gold and silver leaching residue obtained in step (2), and send it into an electric arc furnace to carry out reduction smelting at 1500℃ for 1 hour to obtain a lead-tin alloy product containing gold and silver (the mass content of lead and tin is 84.62% and 11.97%, respectively). (4) Leachate treatment: In the copper and zinc leaching solution obtained in step (2), elemental zinc is added at a molar ratio of copper to elemental zinc of 1:1.2. Zinc replacement is carried out at 65°C for 20 min. After filtration, sponge copper product (Cu mass content is 88.96%) and zinc replacement solution are obtained. In a zinc replacement solution, sodium carbonate is added at a molar ratio of zinc to sodium carbonate of 0.9:1, and the precipitation reaction is carried out at room temperature for 20 minutes to obtain zinc carbonate product.

[0035] According to the test, in step (1), the removal rates of Br and Cl in the obtained roasted residue were 96.33% and 97.87%, respectively.

[0036] Calculations show that in step (2), the leaching rates of Cu, Zn, and Sn in the copper and zinc leaching solution are 94.43%, 99.96%, and 0.02%, respectively; indicating that copper and zinc still maintain a high leaching rate under the low-temperature sulfuric acid roasting conditions of 300 ℃.

[0037] According to the test, in step (3), the direct recovery rate of lead and tin in the obtained lead-tin alloy product was 92.35% and 92.33%, respectively; in step (4), the direct recovery rate of Cu in the obtained sponge copper product was 93.18%.

[0038] Comparative Example 1 The only difference between this comparative example and Example 1 is that in step (1), 1000g of copper smelting ash containing bromine and chlorine is mixed with 450g of concentrated sulfuric acid with a mass concentration of 98% at room temperature without a roasting process, resulting in 1082g of acid-mixed ash; in step (2), the acid-mixed ash obtained in step (1) is directly leached and separated. The rest is the same as in Example 1.

[0039] Upon testing, the mass contents of Br and Cl in the acid-mixed flue ash obtained in step (1) were 0.86% and 2.14%, respectively, and the removal rates of Br and Cl were 16.92% and 60.49%, respectively. After leaching with a solid-liquid ratio of 1:5, the mass concentrations of Cu, Zn, Cl, and Sn in the 5.40L copper and zinc-containing leachate were 6.85g / L, 41.20g / L, 2.05g / L, and 0.96g / L, respectively. After calculation, the leaching rates of Cu, Zn, and Sn were 65.47%, 94.63%, and 18.65%, respectively.

[0040] According to the test, in step (3), the mass content of lead and tin in the gold and silver-containing lead-tin alloy product is 54.35% and 24.21% respectively, and the direct recovery rate of lead and tin is 83.80% and 68.23% respectively; in step (4), the mass content of Cu in the sponge copper product is 77.96%, and the direct recovery rate of Cu is 58.12%.

[0041] The above results fully demonstrate that without low-temperature roasting, simply mixing sulfuric acid with water cannot achieve efficient decomposition and removal of halogens. A large amount of chlorine and bromine impurities remain in the system. At the same time, a large amount of tin components dissolve into the leaching solution, making it impossible to achieve precise separation of copper and zinc from tin and lead. As a result, the product purity and direct metal recovery rate decrease significantly.

[0042] Comparative Example 2-1 The only difference between this comparative example and Example 1 is that in step (1), the calcination temperature is 400 °C. The rest is the same as in Example 1.

[0043] Upon testing, in step (1), the high temperature caused partial thermal decomposition of sulfuric acid, which hindered the decomposition reaction of halides. The mass contents of Br and Cl in the resulting roasted residue were 0.56% and 2.34%, respectively, and the removal rates of Br and Cl were 50.32% and 60.15%, respectively.

[0044] Upon testing, the mass concentrations of Cu, Zn, Cl, and Sn in the 5.37L copper and zinc leaching solution obtained in step (2) were 7.82 g / L, 31.56 g / L, 2.86 g / L, and 0.53 g / L, respectively. After calculation, the leaching rates of Cu, Zn, and Sn were 74.32%, 72.09%, and 10.25%, respectively.

[0045] According to the test, in step (3), the mass content of lead and tin in the gold and silver-containing lead-tin alloy product is 55.94% and 27.48% respectively, and the direct recovery rate of lead and tin is 69.35% and 62.28% respectively; in step (4), the mass content of Cu in the sponge copper product is 81.36%, and the direct recovery rate of Cu is 69.63%.

[0046] The results showed that when the calcination temperature was increased to 400 °C, the decomposition of sulfuric acid intensified, the effective sulfation reaction was insufficient, and the sulfation of copper and zinc was incomplete, resulting in a significant decrease in their leaching rate. At the same time, the high temperature also promoted the activation and dissolution of some sparingly soluble tin salts, which significantly increased the leaching rate of tin and worsened the metal separation effect.

[0047] Comparative Example 2-2 The only difference between this comparative example and Example 1 is that in step (1), the calcination temperature is 500 °C. The rest is the same as in Example 1.

[0048] According to the test, in step (1), the decomposition of sulfuric acid was intensified at high temperature, and the halogen removal efficiency was further reduced. The mass contents of Br and Cl in the obtained roasting residue were 0.67% and 3.12%, respectively, and the removal rates of Br and Cl were 35.09% and 42.23%, respectively.

[0049] Upon testing, the mass concentrations of Cu, Zn, Cl, and Sn in the 5.535L copper and zinc leaching solution obtained in step (2) were 6.15g / L, 25.28g / L, 3.42g / L, and 1.02g / L, respectively. After calculation, the leaching rates of Cu, Zn, and Sn were 60.25%, 59.52%, and 20.32%, respectively.

[0050] According to the test, in step (3), the lead-tin alloy product containing gold and silver has a lead content of 52.94% and a tin content of 23.10%, and the direct recovery rate of lead and tin is 65.28% and 52.07%, respectively; in step (4), the sponge copper product has a Cu content of 80.52% and a Cu direct recovery rate of 55.16%.

[0051] The results showed that when the calcination temperature was increased to 500 ℃, the degree of sulfation of copper and zinc continued to decrease, the leaching rate further decreased, the high-temperature activation effect greatly improved the water solubility of tin, the tin leaching rate doubled, seriously interfered with the separation process of copper and zinc and tin and lead, and the process adaptability was greatly reduced.

[0052] Comparative Examples 2-3 The only difference between this comparative example and Example 1 is that in step (1), the calcination temperature is 600 °C. The rest is the same as in Example 1.

[0053] According to the test, in step (1), the sulfuric acid was basically completely decomposed under ultra-high temperature conditions, and the halogen removal effect was extremely poor. The mass contents of Br and Cl in the obtained roasting residue were 0.71% and 3.53%, respectively, and the removal rates of Br and Cl were 31.2% and 34.6%, respectively.

[0054] Upon testing, the mass concentrations of Cu, Zn, Cl, and Sn in the 5.4L copper and zinc leaching solution obtained in step (2) were 4.88 g / L, 19.65 g / L, 3.95 g / L, and 1.73 g / L, respectively; and the leaching rates of Cu, Zn, and Sn were calculated to be 46.6%, 45.1%, and 33.58%, respectively.

[0055] According to the test, in step (3), the mass content of lead and tin in the gold and silver-containing lead-tin alloy product is 50.84% ​​and 18.50% respectively, and the direct recovery rate of lead and tin is 55.36% and 36.82% respectively; in step (4), the mass content of Cu in the sponge copper product is 78.02%, and the direct recovery rate of Cu is 40.12%.

[0056] The results showed that when the calcination temperature was increased to 600 ℃, the system had a large amount of residual halides and high impurity content. A large amount of tin dissolved into the leaching solution, the effective leaching ratio of copper and zinc was greatly reduced, and the various metal components could not be effectively separated, resulting in a complete deterioration of the overall process effect.

[0057] As can be seen from the above examples and comparative examples, when the calcination temperature increases from 200℃ to 600℃, the leaching rate of tin in the leaching process increases from 0.04% to 33.58%. Under high temperature conditions, a large amount of tin enters the leaching solution and is difficult to separate effectively from copper and zinc. At the same time, the copper and zinc leaching rate, halogen removal rate, and direct recovery rate of metal products all show a continuous downward trend.

[0058] In summary, the process of this invention uses low-temperature sulfuric acid roasting (200-300 ℃) to efficiently remove bromine and chlorine impurities and achieve selective sulfation leaching of copper and zinc, so that tin, lead, gold and silver are stably enriched in the leaching residue. After reduction smelting, zinc replacement and carbonate precipitation, the process accurately achieves the graded separation and recovery of copper and zinc from tin, lead, gold and silver. The metal recovery rate is high, the product purity is excellent, and the impurity removal effect is good, which has excellent prospects for industrial application and promotion value.

Claims

1. A combined hydrometallurgical-pyrometallurgical process for recovering valuable metals from copper flue dust, characterized in that, Includes the following steps: (1) Low-temperature sulfuric acid roasting dehalogenation: copper smelting flue dust containing bromine and chlorine is mixed with concentrated sulfuric acid and then roasted at low temperature to obtain roasting residue; (2) Leaching and separation: The roasted residue obtained in step (1) is added to water, stirred and leached, filtered to obtain leaching residue containing tin, lead, gold and silver and leaching solution containing copper and zinc; (3) Reduction smelting: Add carbonaceous reducing agent to the leaching residue containing tin, lead, gold and silver obtained in step (2) and carry out reduction smelting to obtain a lead-tin alloy product containing gold and silver; (4) Leachate treatment: Add elemental zinc to the copper and zinc leaching solution obtained in step (2) to perform zinc replacement, filter, and obtain sponge copper product and zinc replacement solution.

2. The hydrometallurgical-pyrometallurgical combined process for recovering valuable metals from copper flue dust according to claim 1, characterized in that: In step (1), the mass ratio of the copper smelting flue dust containing bromine and chlorine to concentrated sulfuric acid is 1:0.2-1.0; the main components and mass content of the copper smelting flue dust containing bromine and chlorine are: Cu 5-15%, Sn 1-5%, Pb 5-20%, Au 2-10g / t, Ag 100-2000g / t, Br 1-5%, Cl 5-10%; the mass concentration of the concentrated sulfuric acid is 95-98%; the low-temperature roasting temperature is 200-300℃, and the time is 0.5-3.0h.

3. The hydrometallurgical-pyrometallurgical combined process for recovering valuable metals from copper flue dust according to claim 1 or 2, characterized in that: In step (2), the solid-liquid ratio of the roasted residue to water is 1:3 to 5; the stirring leaching temperature is room temperature, the stirring speed is 400 to 800 r / min, the time is 0.5 to 3.0 h, and the final pH value is 6 to 7.

4. The hydrometallurgical-pyrometallurgical combined process for recovering valuable metals from copper flue dust according to any one of claims 1 to 3, characterized in that: In step (3), the amount of carbonaceous reducing agent is equivalent to 8-12% of the leaching residue containing tin, lead, gold, and silver; the carbonaceous reducing agent includes coke and / or coal; the reduction smelting temperature is 1300-1600℃ and the time is 1-2h; the equipment for reduction smelting includes an electric arc furnace or a blast furnace.

5. The hydrometallurgical-pyrometallurgical combined process for recovering valuable metals from copper flue dust according to any one of claims 1 to 4, characterized in that: In step (4), the molar ratio of copper to elemental zinc in the copper-zinc leaching solution is 1:1.1 to 1.2; the zinc replacement temperature is 50 to 70°C and the time is 20 to 30 minutes; an alkali is added to the obtained zinc replacement solution to carry out a precipitation reaction to obtain zinc carbonate product; the molar ratio of zinc to alkali in the zinc replacement solution is 0.8 to 0.9:1; the precipitation reaction temperature is room temperature to 60°C and the time is 20 to 30 minutes; the alkali includes sodium carbonate and / or sodium bicarbonate.