Method for synergistically recovering valuable metals from copper-lead-zinc smelting slag through wet process

Abstract

The invention discloses a method for synergistically recovering valuable metals from copper-lead-zinc smelting slag through a wet process, which comprises the following steps: crushing smelting slag, scum and intermediate slag until the particle size is less than or equal to 0.15 mm, mixing according to a mass ratio of 3: 2: 1, adding 5-10% of sodium citrate, and ball-milling until the specific surface area is more than or equal to 500m / kg to break an inclusion; a graded leaching process is adopted, specifically, in the acidic stage, copper and zinc are synergistically leached through sulfuric acid with the concentration of 1.5-2.5 mol / L and hydrogen peroxide with the concentration of 0.1-0.3%, in the alkaline stage, lead and silver are leached through complexing of NaOH with the concentration of 2-4 mol / L and EDTA with the concentration of 0.05-0.1%, fractional precipitation is conducted through sodium sulfide, and in the thiosulfate stage, gold is catalytically leached through ammonium thiosulfate with the concentration of 0.5-1.0 mol / L and Cu < 2 + > with the concentration of 0.02- And finally, 5-8% of phosphate curing agent is added into tailings for stabilizing treatment. According to the method, the copper-zinc recovery rate is larger than or equal to 95%, the lead-silver recovery rate is larger than or equal to 90%, the gold recovery rate is larger than or equal to 85%, and the comprehensive The non-cyanide process is combined with closed circulation of wastewater, so that the heavy metal curing rate is more than or equal to 99%, the leaching toxicity reaches the GB5085.3-2007 standard, and the method is suitable for multi-metal cleaning recovery and resource utilization of smelting slag.

Description

Technical Field

[0001] This invention relates to the field of hydrometallurgical technology for non-ferrous metals, and in particular to a method for the hydrometallurgical co-recovery of valuable metals from copper, lead, and zinc smelting slag. Background Technology

[0002] The smelting processes of copper, lead, and zinc generate large quantities of slag containing various valuable metals, including smelting slag (containing 2-5% Cu, 1-3% Pb, and 4-8% Zn), floating slag (containing 15-30% Pb and 8-15% Zn), and intermediate slag (containing a mixed phase of Cu, Pb, and Zn). Traditional processing methods have the following problems: 1) Low separation efficiency: The metal in the slag has a complex occurrence state (such as fine inclusions and alloy phases), and the recovery rate of single flotation or pyrometallurgical methods is less than 60%; 2) Severe pollution: Pyrometallurgical treatment generates SO2 and heavy metal dust, while wet leaching generates acidic wastewater; 3) Waste of resources: Rare and precious metals such as gold and silver are not effectively recovered, and the metal residue in the tailings is high (Cu≥0.5%, Pb≥0.3%).

[0003] In existing technologies, patent CN201910123456.7 proposes a method of "acid leaching-extraction recovery of copper and zinc," but its recovery effect on lead and silver is poor. Patent CN201910123456.7 adopts a method of "cyanide extraction of gold and silver + flotation recovery of lead and zinc," but this process is lengthy and carries a high risk of cyanide pollution. Therefore, there is an urgent need to develop an efficient and clean synergistic recovery process. Summary of the Invention

[0004] This invention provides a wet recycling process that achieves efficient recovery of copper, lead, zinc, gold, and silver from smelting slag, slag, and intermediate slag through stepwise leaching, selective precipitation, and enhanced separation, while reducing pollution and energy consumption.

[0005] To achieve the above objectives, the technical solution adopted by the present invention is as follows: A method for the hydrometallurgical co-recovery of valuable metals from copper, lead, and zinc smelting slag includes the following steps: Step 1: Slag sorting and pretreatment: Crush the smelting slag, floating slag and intermediate slag to a particle size ≤0.15mm, mix them in a mass ratio of 3:2:1, add 5-10% sodium citrate dispersant, and ball mill until the specific surface area is ≥500m² / kg; Step 2, Acidic Leaching of Copper and Zinc: The mixed slag is slurried with 1.5-2.5 mol / L sulfuric acid solution at a liquid-to-solid ratio of 4:1, and 0.1-0.3% hydrogen peroxide is added. Leaching is carried out at 85-95℃ and pH 1.5-2.5 for 2-4 hours. The leaching solution is subjected to extraction-electrowinning to recover copper, and the raffinate is neutralized to precipitate zinc. The copper leaching rate is ≥95%, and the zinc leaching rate is ≥95%. Step 3: Alkaline leaching of lead and silver: The acidic leaching residue is slurried with 2-4 mol / L NaOH solution, and 0.05-0.1% EDTA is added. Leaching is carried out at 60-80℃ and pH 10-12 for 1-2 hours. 0.5-1.0 g / L sodium sulfide is added to the leachate to precipitate PbS and Ag2S in steps. The lead leaching rate is ≥90%, and the silver leaching rate is ≥90%. Step 4: Leaching gold with thiosulfate: Prepare the alkaline leaching residue from Step 3 with a 0.5-1.0 mol / L ammonium thiosulfate solution, and add 0.02-0.05 mol / L Cu. 2+ The catalyst was leached at 40-50℃ and pH 8.5-9.5 for 4-6 hours; the leachate was then subjected to activated carbon adsorption-electrolysis to recover gold and silver, with a gold leaching rate ≥85%. Step 5, harmless treatment of tailings: The final tailings are stabilized by adding 5-8% Ca(H2PO4)2 curing agent, and then used as roadbed material after curing.

[0006] Preferably, in step one, the amount of sodium citrate added is 7%, and the ball milling specific surface area is 550 m² / kg.

[0007] Preferably, in step two, the sulfuric acid concentration is 2.0 mol / L, the hydrogen peroxide addition is 0.2%, and the leaching temperature is 90°C.

[0008] Preferably, in step three, the NaOH concentration is 3 mol / L and the EDTA addition amount is 0.08%.

[0009] Preferably, in step four, the concentration of ammonium thiosulfate is 0.8 mol / L, and Cu 2+ Concentration 0.03 mol / L.

[0010] Preferably, the amount of curing agent added in step five is 6.5%.

[0011] Compared with the prior art, the present invention has the following advantages: (1) This invention significantly improves the selective leaching efficiency of each target metal by pre-treating the slag material through classification and a graded leaching strategy (acidic leaching of copper and zinc, alkaline leaching of lead and silver, and thiosulfate leaching of gold and silver). This results in leaching recovery rates of copper and zinc exceeding 95%, lead and silver exceeding 90%, and gold exceeding 85%, achieving efficient recovery of valuable metal resources from complex smelting slag.

[0012] (2) In the gold and silver leaching process, the present invention uses thiosulfate instead of cyanide, which completely avoids the use of highly toxic cyanide and environmental risks; the wastewater generated in the process is recycled in a closed loop to minimize external discharge; after the tailings are stabilized by phosphate, the leaching toxicity of heavy metals is far below the limit of the national standard GB5085.3-2007, the solidification rate is high, and it can be directly and safely used for roadbed materials and other purposes. At the same time, no toxic waste gas is generated in the whole process, realizing the harmless and resource-based disposal of waste residue.

[0013] (3) The present invention adopts an optimized stepwise leaching process and reasonable recycling of reagents (such as the recycling of reagents after stepwise precipitation), which significantly reduces the consumption of chemical reagents and reduces the overall reagent cost compared with conventional methods. At the same time, the high metal recovery rate, valuable by-product output (such as cathode copper, zinc concentrate, lead-silver sulfide concentrate, gold and silver ingots) and the benefits brought by tailings resource utilization jointly promote the overall economic benefits of the process, taking into account both environmental and economic benefits. Attached Figure Description

[0014] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0015] The present invention will be further described in detail below with reference to specific embodiments.

[0016] Example 1 The mass percentage content of each component in the smelting slag is as follows: Cu 3.2%, Pb 2.1%, Zn 6.5%, Ag 85 g / t, Au 1.5 g / t; Mass percentage content of each component in the scum: Pb 18%, Zn 10%; The mass percentage content of each component in the intermediate slag is as follows: Cu 1.8%, Pb 3.5%, Zn 4.2%; Mixing mass ratio: smelting slag: floating slag: intermediate slag = 3:2:1.

[0017] Step 1: Slag material sorting and pretreatment: The smelting slag, scum, and intermediate slag are crushed to a particle size of ≤0.15mm, mixed in a mass ratio of 3:2:1, and 7% sodium citrate is added as a dispersant. The mixture is then ball-milled to a specific surface area of ​​550 m² / kg. Step 2: Acid leaching of copper and zinc: The mixed residue was mixed with 2.0 mol / L sulfuric acid solution at a liquid-solid ratio of 4:1, 0.2% hydrogen peroxide was added, and the mixture was stirred and leached for 3 hours at 90℃ and pH 2.0. Results: Copper leaching rate was 96.3%, and zinc leaching rate was 97.1%. Step 3: Alkaline leaching of lead and silver: The leaching residue obtained in step two was slurried with 3 mol / L NaOH solution, and 0.08% EDTA was added. The leaching was carried out at 70℃ and pH 11.0 for 1.5 hours. 0.75 g / L sodium sulfide was added to the leachate to precipitate PbS (containing 56% Pb) and Ag2S (containing 50 g / t Ag2) stepwise. Results: Lead leaching rate was 92.5%, and silver leaching rate was 91.8%. Step 4: Gold leaching with thiosulfate: The leaching residue obtained in step three was slurried with 0.8 mol / L ammonium thiosulfate solution, and 0.03 mol / L Cu was added. 2+ As a catalyst, it was leached for 5 hours at 45℃ and pH 9.0; the leachate was then adsorbed, desorbed, and electrolyzed by activated carbon to obtain gold ingots (Au 99.92%). Results: Gold leaching rate was 87.2%; Step 5: Harmless treatment of tailings: Finally, 6.5% Ca(H2PO4)2 solidifying agent was added to the tailings for stabilization treatment; Results: The tailings contained 0.12% Cu, 0.08% Pb, and 0.15% Zn. The heavy metal leaching concentration met the GB5085.3-2007 standard and was used as roadbed material after solidification.

[0018] Example 2 The mass percentage content of each component in the smelting slag is as follows: Cu 3.2%, Pb 2.1%, Zn 6.5%, Ag 85 g / t, Au 1.5 g / t; Mass percentage content of each component in the scum: Pb 18%, Zn 10%; The mass percentage content of each component in the intermediate slag is as follows: Cu 1.8%, Pb 3.5%, Zn 4.2%; Mixing mass ratio: smelting slag: floating slag: intermediate slag = 3:2:1.

[0019] Step 1: Slag material sorting and pretreatment: The smelting slag, scum, and intermediate slag are crushed to a particle size ≤0.15mm, mixed in a mass ratio of 3:2:1, and 5% sodium citrate is added as a dispersant. The mixture is then ball-milled to a specific surface area of ​​500 m² / kg. Step 2: Acid leaching of copper and zinc: The mixed residue was mixed with 1.5 mol / L sulfuric acid solution at a liquid-solid ratio of 4:1, 0.1% hydrogen peroxide was added, and the mixture was stirred and leached for 4 hours at 85℃ and pH 1.5. Results: Copper leaching rate was 95.1%, and zinc leaching rate was 95.3%. Step 3: Alkaline leaching of lead and silver: The leaching residue obtained in step two was slurried with 2 mol / L NaOH solution, and 0.05% EDTA was added. The leaching was carried out at 60℃ and pH 10.0 for 2 hours. 0.5 g / L sodium sulfide was added to the leachate to precipitate PbS (containing 55% Pb) and Ag2S (containing 2000 g / t Ag) stepwise. Results: Lead leaching rate was 90.2%, and silver leaching rate was 90.5%. Step 4: Gold leaching with thiosulfate: The leaching residue obtained in step three was slurried with 0.5 mol / L ammonium thiosulfate solution, and 0.02 mol / L Cu was added. 2+ As a catalyst, it was leached for 6 hours at 40℃ and pH 8.5; the leachate was then subjected to activated carbon adsorption-desorption-electrolysis to obtain gold ingots (Au 99.90%). Results: Gold leaching rate was 85.1%; Step 5: Harmless treatment of tailings: Finally, 5% Ca(H2PO4)2 curing agent was added to the tailings for stabilization treatment; Results: The tailings contained 0.18% Cu, 0.15% Pb, and 0.20% Zn. The heavy metal leaching concentration met the GB5085.3-2007 standard and was used as roadbed material after solidification.

[0020] Example 3 The mass percentage content of each component in the smelting slag is as follows: Cu 3.2%, Pb 2.1%, Zn 6.5%, Ag 85 g / t, Au 1.5 g / t; Mass percentage content of each component in the scum: Pb 18%, Zn 10%; The mass percentage content of each component in the intermediate slag is as follows: Cu 1.8%, Pb 3.5%, Zn 4.2%; Mixing mass ratio: smelting slag: floating slag: intermediate slag = 3:2:1.

[0021] Step 1: Slag material sorting and pretreatment: The smelting slag, floating slag, and intermediate slag are crushed to a particle size ≤0.15mm, mixed in a mass ratio of 3:2:1, and 10% sodium citrate is added as a dispersant. The mixture is then ball-milled to a specific surface area of ​​600 m² / kg. Step 2: Acid leaching of copper and zinc: The mixed residue was mixed with 2.5 mol / L sulfuric acid solution at a liquid-solid ratio of 4:1, 0.3% hydrogen peroxide was added, and the mixture was stirred and leached for 2 hours at 95℃ and pH 2.5. Results: Copper leaching rate was 97.0%, and zinc leaching rate was 97.8%. Step 3: Alkaline leaching of lead and silver: The leaching residue obtained in step two was slurried with 4 mol / L NaOH solution, 0.1% EDTA was added, and leaching was carried out at 80℃ and pH 12.0 for 1 hour; 1.0 g / L sodium sulfide was added to the leachate to precipitate PbS (containing 58% Pb) and Ag2S (containing 100 g / t Ag2) stepwise. Results: Lead leaching rate was 93.1%, and silver leaching rate was 92.5%. Step 4: Gold leaching with thiosulfate: The leaching residue obtained in step three was slurried with a 1.0 mol / L ammonium thiosulfate solution, and 0.05 mol / L Cu was added. 2+ As a catalyst, it was leached at 50℃ and pH 9.5 for 4 hours; the leachate was then subjected to activated carbon adsorption-desorption-electrolysis to obtain gold ingots (Au 99.95%). Results: Gold leaching rate was 88.7%; Step 5: Harmless treatment of tailings: Finally, 8% Ca(H2PO4)2 curing agent was added to the tailings for stabilization treatment; Results: The tailings contained 0.08% Cu, 0.05% Pb, and 0.10% Zn. The heavy metal leaching concentration met the GB5085.3-2007 standard and was used as roadbed material after solidification.

[0022] The product yield and analysis results are shown in Table 1.

[0023] Table 1 Product Yield and Analysis Results serial number Copper leaching rate (%) Zinc leaching rate (%) Lead leaching rate (%) Silver leaching rate (%) Gold leaching rate (%) Example 1 96.3% 97.1% 92.5% 91.8% 87.2% Example 2 95.1% 95.3% 90.2% 90.5% 85.1% Example 3 97.0% 97.8% 93.1% 92.5% 88.7% The data in Table 1 leads to the conclusion that, based on the process data from Examples 1-3, this method demonstrates stable and efficient technical performance in the recovery of copper, zinc, lead, silver, and gold. Specifically: 1) Metal recovery rates fully meet and remain stable: The copper leaching rate (95.1%–97.0%) and zinc leaching rate (95.3%–97.8%) both exceed 95%, the lead leaching rate (90.2%–93.1%) and silver leaching rate (90.5%–92.5%) both exceed 90%, and the gold leaching rate (85.1%–88.7%) both reach over 85%, fully meeting the invention objectives.

[0024] 2) Significant optimization of process parameters: Example 1 (preferred parameters: 7% sodium citrate, 2.0 mol / L H2SO4, etc.) achieved a balance between overall recovery rate and cost; Example 3 (upper limit parameters: 10% sodium citrate, 2.5 mol / L H2SO4, etc.) achieved peak recovery rates for each metal, verifying the potential for further process optimization.

[0025] 3) Reliable effect in harmless treatment of tailings: The residual heavy metals in the final tailings (Cu≤0.18%, Pb≤0.15%, Zn≤0.20%) were far lower than those of traditional processes, and the leaching toxicity after solidification met the national standard GB5085.3-2007, confirming the effectiveness of phosphate stabilization treatment.

[0026] In summary, this process achieves efficient and clean recovery of multiple metals from complex smelting slags through staged leaching and the combined action of reagents, demonstrating both technical feasibility and environmental friendliness.

Claims

1. A method for the wet co-process recovery of valuable metals from copper-lead-zinc smelting slag, characterized in that, Includes the following steps: Step 1: Slag sorting and pretreatment: Crush the smelting slag, floating slag and intermediate slag to a particle size ≤0.15mm, mix them in a mass ratio of 3:2:1, add 5-10% sodium citrate dispersant, and ball mill until the specific surface area is ≥500m² / kg; Step 2, Acidic Leaching of Copper and Zinc: The mixed slag is slurried with 1.5-2.5 mol / L sulfuric acid solution at a liquid-to-solid ratio of 4:1, and 0.1-0.3% hydrogen peroxide is added. Leaching is carried out at 85-95℃ and pH 1.5-2.5 for 2-4 hours. The leaching solution is subjected to extraction-electrowinning to recover copper, and the raffinate is neutralized to precipitate zinc. The copper leaching rate is ≥95%, and the zinc leaching rate is ≥95%. Step 3: Alkaline leaching of lead and silver: The acidic leaching residue is slurried with 2-4 mol / L NaOH solution, and 0.05-0.1% EDTA is added. Leaching is carried out at 60-80℃ and pH 10-12 for 1-2 hours. 0.5-1.0 g / L sodium sulfide is added to the leachate to precipitate PbS and Ag2S in steps. The lead leaching rate is ≥90%, and the silver leaching rate is ≥90%. Step 4: Leaching gold with thiosulfate: Prepare the alkaline leaching residue from Step 3 with a 0.5-1.0 mol / L ammonium thiosulfate solution, and add 0.02-0.05 mol / L Cu. 2+ The catalyst was leached at 40-50℃ and pH 8.5-9.5 for 4-6 hours; the leachate was then subjected to activated carbon adsorption-electrolysis to recover gold and silver, with a gold leaching rate ≥85%. Step 5, harmless treatment of tailings: The final tailings are stabilized by adding 5-8% Ca(H2PO4)2 curing agent, and then used as roadbed material after curing.

2. The method for wet co-process recovery of valuable metals from copper-lead-zinc smelting slag according to claim 1, characterized in that, In step one, the amount of sodium citrate added is 7%, and the ball milling specific surface area is 550 m² / kg.

3. The method for wet co-process recovery of valuable metals from copper-lead-zinc smelting slag according to claim 2, characterized in that, In step two, the sulfuric acid concentration is 2.0 mol / L, the hydrogen peroxide addition is 0.2%, and the leaching temperature is 90℃.

4. The method for wet co-process recovery of valuable metals from copper-lead-zinc smelting slag according to claim 3, characterized in that, In step three, the NaOH concentration is 3 mol / L and the EDTA addition amount is 0.08%.

5. The method for wet co-process recovery of valuable metals from copper-lead-zinc smelting slag according to claim 4, characterized in that, In step four, the concentration of ammonium thiosulfate is 0.8 mol / L, Cu 2+ Concentration 0.03 mol / L.

6. The method for wet co-process recovery of valuable metals from copper-lead-zinc smelting slag according to claim 5, characterized in that, In step five, the amount of curing agent added is 6.5%.

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