Method for efficiently extracting scattered metal indium from lead-zinc smelting soot

Abstract

The invention discloses a method for efficiently extracting scattered metal indium from lead and zinc smelting soot, and particularly relates to the technical field of scattered metal extraction. Step 2, a leaching stage: a, neutral leaching; b, low-acid leaching; c, high-acid leaching; step 3, an extraction and separation stage: a, primary extraction; b, primary reverse extraction; c, secondary extraction; d, secondary reverse extraction; step 4, a refining stage: a, replacement; and b, electrolytic refining. According to the method, the technical scheme that multi-section acid leaching and ultrasonic-assisted leaching are combined is adopted, the leaching rate of indium is remarkably increased, and the total leaching rate can reach 88%-92% and is increased by 8%-10% compared with a traditional method; and meanwhile, the leaching time is shortened from 4 hours to 2 hours by the ultrasonic-assisted technology, so that the energy consumption is reduced.

Description

Technical Field

[0001] This invention relates to the field of rare and dispersed metal extraction technology, specifically to a method for efficiently extracting the rare and dispersed metal indium from lead-zinc smelting flue dust. Background Technology

[0002] Indium is an important rare and dispersed metal widely used in high-tech industries such as electronics, semiconductors, and photovoltaics. Lead-zinc smelting flue dust is an important secondary resource of indium. During lead-zinc smelting, approximately 94% of indium enters the zinc ferrite lattice in an isomorphic form. After rotary kiln processing, the indium volatilization rate can reach 70%–90%, accumulating in zinc oxide dust. Therefore, recovering indium from lead-zinc smelting flue dust has significant resource utilization value and economic benefits.

[0003] Currently, traditional methods for extracting indium from lead-zinc smelting flue dust have many problems: low leaching efficiency, with indium leaching rates typically below 80%; poor separation performance, making it difficult to effectively separate indium from impurities such as iron, zinc, and cadmium; long process flow and high energy consumption; and serious environmental pollution, especially chloride ion emissions. Furthermore, existing extraction methods have low selectivity for indium and other metals, making it difficult to guarantee product purity and subsequent refining challenging. Summary of the Invention

[0004] The purpose of this invention is to provide a method for efficiently extracting the rare metal indium from lead-zinc smelting flue dust, so as to solve the problems mentioned in the background art.

[0005] To achieve the above objectives, the present invention provides the following technical solution: a method for efficiently extracting the rare and dispersed metal indium from lead-zinc smelting flue dust, comprising the following steps: Step 1, Preprocessing stage: a. Ball mill the lead-zinc smelting flue dust to control the particle size to be less than 74μm; b. For flue dust with a sulfur content >2.5 wt%, roasting should be carried out in a weak oxidizing atmosphere of 500-600℃, O2 volume concentration of 3-7%, and CO≤0.5%, with a heating rate of 4-6℃ / min and holding for 0.5-3 h; when the sulfur content of the flue dust is ≤2.5%, roasting should be omitted and the flue dust should be directly subjected to neutral leaching. Step 2, Leaching Stage: a. Neutral leaching: The pretreated soot is added to a sulfuric acid solution with a pH of 3.0-3.5 and leached at room temperature with stirring for 2 hours. The solid and liquid are separated to obtain a neutral leachate and a neutral leachate residue. b. Low acid leaching: Add the neutral leaching residue to a solution with a sulfuric acid concentration of 30-40 g / L, control the liquid-solid ratio to 5:1, stir and leach at 80-90℃, and use ultrasonic assisted leaching with a power of 300W for 2 hours. The solid and liquid are separated to obtain low acid leaching solution and low acid leaching residue. c. High acid leaching: Add the low acid leaching residue to a sulfuric acid solution with a concentration of 80 g / L, and leach at 80-90℃ for 4 hours to separate the high acid leaching solution and the high acid leaching residue. The high acid leaching solution is returned to the low acid leaching step for recycling. Step 3, Extraction and Separation Stage: a. Primary extraction: Using 10 vol% P204 as the extractant and sulfonated kerosene as the diluent, the low-acid leachate was subjected to a two-stage countercurrent extraction under the condition of A / O ratio of 2 / 1. The mixing time was controlled at 5 minutes, and the indium-loaded P204 organic phase and the primary raffinate were separated. b. Primary back-extraction: The indium-loaded P204 organic phase is back-extracted with 6 mol / L hydrochloric acid to obtain indium chloride back-extraction solution; c. Secondary extraction: A mixed extractant system of 20 vol% TBP + 15 vol% P350 is used to perform two-stage countercurrent extraction on the indium chloride back-extraction solution under the condition of A / O ratio of 3 / 2, to separate the indium-loaded mixed organic phase and the secondary raffinate. The secondary raffinate is returned to the primary back-extraction step for recycling. d. Secondary back-extraction: The indium-loaded mixed organic phase is subjected to three-stage countercurrent back-extraction with pure water to obtain a high-purity indium-rich solution; Step 4, Refining Stage: a. Displacement: The indium-rich solution is replaced with a zinc plate with a purity of not less than 99.9% at 50-60℃ to obtain sponge indium; b. Electrolytic refining: The sponge indium briquette is pressed and cast into a crude indium anode. An acidic electrolyte is used to control the pH at 2.0-2.5, the temperature at 25-30℃, and the current density at 80-100A / m² for electrolytic refining to obtain a refined indium product with a purity of over 99.99%.

[0006] Furthermore, in the pretreatment stage, the O2 volume concentration of the weak oxidizing atmosphere is 4-6%, the CO volume concentration is ≤0.3%, the heating rate is 5 ℃ / min, and the holding time is 2 h.

[0007] Furthermore, in the low-acid leaching step, the pulsed ultrasonic frequency is 22±1kHz, the power density is 18±2W / L, the duty cycle is 50%, the interval is 0.5s, and the ultrasonication and stirring are carried out synchronously.

[0008] Furthermore, in the high acid leaching step, the amount of high acid leaching solution returned accounts for 80-100% of the total acid content of the low acid leaching solution, and suspended solids are removed by pressure filtration before being returned.

[0009] Furthermore, in the secondary extraction step, the amount of secondary raffinate returned accounts for 60-80% of the volume of 6 mol / L hydrochloric acid required for the primary back-extraction, and any shortfall is supplemented with industrial hydrochloric acid.

[0010] Furthermore, germanium is recovered from the primary raffinate using a synergistic extraction process of P204 and Yw-100, and zinc and iron are recovered from each leachate.

[0011] Furthermore, Yw-100 is a hydroxyoxime-based copper extractant, with the active ingredient being 2-hydroxy-5-nonylacetophenone oxime (CAS 34138-71-1). Before use, it is diluted with sulfonated kerosene to a volume fraction of 10 vol%, and mixed with P204 at a volume ratio of 1:1.

[0012] Furthermore, in the replacement step, the replacement temperature is 55 ℃ and the zinc plate purity is ≥99.95%.

[0013] Furthermore, in the electrolytic refining step, the current density is 90 A / m², and the In content of the electrolyte is... 3+ With a concentration of 80-100 g / L and NaCl ≤ 0.1 g / L, the cathode yields indium with a purity ≥ 99.993%.

[0014] Furthermore, the method results in a chloride ion emission of ≤600 mg / kg dry flue ash, a reduction of ≥86% compared to traditional methods, a total indium leaching rate of 88%-92%, a total indium recovery rate of 85%-90%, and a refined indium product purity of ≥99.99%.

[0015] The technical effects and advantages provided by the present invention in the above technical solution are as follows: 1. This invention employs a technical solution that combines multi-stage acid leaching with ultrasonic-assisted leaching, which significantly improves the leaching rate of indium, with a total leaching rate of 88% to 92%, which is 8-10 percentage points higher than the traditional method. At the same time, the ultrasonic-assisted technology reduces the leaching time from 4 hours to 2 hours, thereby reducing energy consumption. 2. This invention develops a two-stage extraction and separation system, combining the advantages of P204 and TBP / P350 mixed extractants to achieve efficient separation of indium from various impurities in complex matrices. The first-stage extraction achieves an indium extraction rate of 99%, while the extraction rates for iron, zinc, and cadmium are only 1.6%, 0.6%, and 1.1%, respectively. The second-stage extraction again achieves a 99% indium extraction rate, significantly improving product purity. 3. By returning the high-acid leachate to the low-acid leachate process and returning the raffinate from the secondary extraction to the primary back-extraction process, the recycling of acid and hydrochloric acid was achieved, reducing chloride ion emissions by 86%, significantly reducing the impact on the environment, and achieving the goal of clean production. 4. The entire process of this invention achieves the synergistic recovery of zinc, germanium and other valuable metals, with a germanium recovery rate of up to 97%, which improves the comprehensive utilization rate of resources and increases the economic benefits of the process. 5. The process of this invention is stable and reliable, easy to operate, and the total recovery rate of indium can reach 85% to 90%, with a product purity of over 99.99%, showing good prospects for industrial application. Attached Figure Description

[0016] To more clearly illustrate the technical solutions in the embodiments of this application or the prior art, the drawings used in the embodiments will be briefly introduced below. Obviously, the drawings described below are only some embodiments recorded in this invention. For those skilled in the art, other drawings can be obtained based on these drawings.

[0017] Figure 1 This is a simplified process flow diagram of the present invention. Detailed Implementation

[0018] To enable those skilled in the art to better understand the technical solutions of the present invention, the present invention will be further described in detail below with reference to specific embodiments, but the present invention is not limited to these embodiments.

[0019] Example 1: Step 1, Pre-processing stage: The lead-zinc smelting flue dust was ball-milled to control the particle size to 70 μm. The flue dust contained 3.5% sulfur and was roasted at 550℃ in a weak oxidizing atmosphere with an O2 volume concentration of 5% and CO ≤ 0.3%. The heating rate was controlled at 5℃ / min and the holding time was 2 hours.

[0020] Step 2, Leaching Stage: a. Neutral leaching: The pretreated soot is added to a sulfuric acid solution with a pH of 3.2 and leached at room temperature for 2 hours to obtain a neutral leachate and a neutral leachate residue. b. Low-acid leaching: The neutral leaching residue is added to a solution with a sulfuric acid concentration of 35 g / L, and the liquid-solid ratio is controlled at 5:1. The leaching is carried out by stirring at 85℃. During the leaching process, pulsed ultrasound with a frequency of 22 kHz and a power density of 18 W / L is applied. The pulse waveform is a square wave with a duty cycle of 50% and an interval of 0.5s. The leaching is assisted for 2 hours, and the low-acid leachate and low-acid leaching residue are obtained by separation. c. High acid leaching: Add the low acid leaching residue to a solution with a sulfuric acid concentration of 80 g / L, leach at 85°C for 4 hours, and separate the high acid leaching solution and high acid leaching residue. Return the high acid leaching solution to the low acid leaching step for recycling.

[0021] Step 3, Extraction and Separation Stage: a. Primary extraction: Using 10 vol% P204 as the extractant and sulfonated kerosene as the diluent, a second-stage countercurrent extraction was carried out on the low-acid leachate under the condition of A / O ratio of 2 / 1. The mixing time was controlled at 5 minutes to separate the indium-loaded P204 organic phase and the primary raffinate. b. Primary back-extraction: The indium-loaded P204 organic phase is back-extracted with 6 mol / L hydrochloric acid to obtain indium chloride back-extraction solution; c. Secondary extraction: The indium chloride back-extraction solution is subjected to a two-stage countercurrent extraction using a mixed extractant system of 20 vol% TBP + 15 vol% P350 at a ratio of 3 / 2 to separate the indium-loaded mixed organic phase and the secondary raffinate. The secondary raffinate is then returned to the primary back-extraction step for recycling. d. Secondary back-extraction: The indium-loaded mixed organic phase is subjected to three-stage countercurrent back-extraction with pure water to obtain a high-purity indium-rich solution.

[0022] Step 4, Refining Stage: a. Displacement: The indium-rich solution is replaced with a zinc plate with a purity of 99.95% at 55°C to obtain sponge indium; b. Electrolytic refining: The sponge indium is pressed into briquettes and cast into a crude indium anode. An acidic electrolyte is used to control the pH value at 2.2, the temperature at 28℃, and the current density at 90A / m² for electrolytic refining to obtain refined indium products.

[0023] Testing showed that the total indium leaching rate in this embodiment was 91.2%, the total recovery rate was 88.5%, and the purity of the refined indium product was 99.993%. Chloride ion emissions were reduced by 86.3% compared to traditional processes.

[0024] Example 2: Step 1, Pre-processing stage: The lead-zinc smelting flue dust was ball-milled to control the particle size to 65 μm; the sulfur content of the flue dust was 2.0%, and no roasting treatment was required.

[0025] Step 2, Leaching Stage: a. Neutral leaching: The pretreated soot is added to a sulfuric acid solution with a pH of 3.0 and leached at room temperature for 2 hours to obtain a neutral leachate and a neutral leachate residue. b. Low-acid leaching: The neutral leaching residue is added to a solution with a sulfuric acid concentration of 30 g / L, and the liquid-solid ratio is controlled at 5:1. The mixture is stirred and leached at 80℃. During the leaching process, pulsed ultrasound with a frequency of 22 kHz and a power density of 18 W / L is applied. The pulse waveform is a square wave with a duty cycle of 50% and an interval of 0.5 s. The leaching is assisted for 2 hours, and the low-acid leachate and low-acid leaching residue are separated. c. High acid leaching: Add the low acid leaching residue to a solution with a sulfuric acid concentration of 80 g / L, leach at 80°C for 4 hours, and separate the high acid leaching solution and high acid leaching residue. Return the high acid leaching solution to the low acid leaching step for recycling.

[0026] Step 3, Extraction and Separation Stage: a. Primary extraction: Using 10 vol% P204 as the extractant and sulfonated kerosene as the diluent, a second-stage countercurrent extraction was carried out on the low-acid leachate under the condition of A / O ratio of 2 / 1. The mixing time was controlled at 5 minutes to separate the indium-loaded P204 organic phase and the primary raffinate. b. Primary back-extraction: The indium-loaded P204 organic phase is back-extracted with 6 mol / L hydrochloric acid to obtain indium chloride back-extraction solution; c. Secondary extraction: The indium chloride back-extraction solution is subjected to a two-stage countercurrent extraction using a mixed extractant system of 20 vol% TBP + 15 vol% P350 at a ratio of 3 / 2 to separate the indium-loaded mixed organic phase and the secondary raffinate. The secondary raffinate is then returned to the primary back-extraction step for recycling. d. Secondary back-extraction: The indium-loaded mixed organic phase is subjected to three-stage countercurrent back-extraction with pure water to obtain a high-purity indium-rich solution.

[0027] Step 4, Refining Stage: a. Displacement: The indium-rich solution is replaced with a zinc plate of 99.9% purity at 50°C to obtain sponge indium; b. Electrolytic refining: Sponge indium is pressed into briquettes and cast into crude indium anodes. Electrolytic refining is carried out using an acidic electrolyte, with the pH value controlled at 2.0, the temperature at 25℃, and the current density at 80A / m², to obtain refined indium products.

[0028] Testing showed that the total indium leaching rate in this embodiment was 89.5%, the total recovery rate was 86.2%, and the purity of the refined indium product was 99.991%. Chloride ion emissions were reduced by 85.8% compared to traditional processes.

[0029] Example 3: Step 1, Pre-processing stage: The lead-zinc smelting flue dust was ball-milled to control the particle size to 74 μm. The flue dust contained 4.5% sulfur and was roasted at 600℃ in a weak oxidizing atmosphere with an O2 volume concentration of 5% and CO ≤ 0.3%. The heating rate was controlled at 4℃ / min and the holding time was 3 hours.

[0030] Step 2, Leaching Stage: a. Neutral leaching: The pretreated soot is added to a sulfuric acid solution with a pH of 3.5 and leached at room temperature for 2 hours to obtain a neutral leachate and a neutral leachate residue. b. Low-acid leaching: The neutral leaching residue is added to a solution with a sulfuric acid concentration of 40 g / L, and the liquid-solid ratio is controlled at 5:1. The leaching is carried out by stirring at 90℃. During the leaching process, pulsed ultrasound with a frequency of 22 kHz and a power density of 18 W / L is applied. The pulse waveform is a square wave with a duty cycle of 50% and an interval of 0.5s. The leaching is assisted for 2 hours, and the low-acid leachate and low-acid leaching residue are obtained by separation. c. High acid leaching: Add the low acid leaching residue to a solution with a sulfuric acid concentration of 80 g / L, leach at 90°C for 4 hours, and separate the high acid leaching solution and high acid leaching residue. Return the high acid leaching solution to the low acid leaching step for recycling.

[0031] Step 3, Extraction and Separation Stage: a. First-stage extraction: The low-acid leachate was subjected to a second-stage countercurrent extraction using 10 vol% P204 as the extractant and sulfonated kerosene as the diluent, with a ratio of A / O of 2 / 1. The mixing time was controlled to be 5 minutes, and the indium-loaded P204 organic phase and the first-stage raffinate were separated. b. Primary back-extraction: The indium-loaded P204 organic phase is back-extracted with 6 mol / L hydrochloric acid to obtain indium chloride back-extraction solution; c. Secondary extraction: The indium chloride back-extraction solution is subjected to a two-stage countercurrent extraction using a mixed extractant system of 20 vol% TBP + 15 vol% P350 at a ratio of 3 / 2 to separate the indium-loaded mixed organic phase and the secondary raffinate. The secondary raffinate is then returned to the primary back-extraction step for recycling. d. Secondary back-extraction: The indium-loaded mixed organic phase is subjected to three-stage countercurrent back-extraction with pure water to obtain a high-purity indium-rich solution.

[0032] Step 4, Refining Stage: a. Displacement: The indium-rich solution is replaced with a zinc plate with a purity of 99.95% at 60°C to obtain sponge indium; b. Electrolytic refining: The sponge indium is pressed into briquettes and cast into a crude indium anode. An acidic electrolyte is used to control the pH value at 2.5, the temperature at 30℃, and the current density at 100A / m² for electrolytic refining to obtain refined indium products.

[0033] According to the test results, the total leaching rate of indium in this embodiment was 90.8%, the total recovery rate was 87.6%, the purity of the refined indium product was 99.992%, and the chloride ion emission was reduced by 86.5% compared with the traditional process.

[0034] The foregoing has only described certain exemplary embodiments of the present invention by way of illustration. Undoubtedly, those skilled in the art can modify the described embodiments in various ways without departing from the spirit and scope of the present invention. Therefore, the foregoing drawings and descriptions are illustrative in nature and should not be construed as limiting the scope of protection of the claims of the present invention.

Claims

1. A method for efficiently extracting the rare and dispersed metal indium from lead-zinc smelting flue dust, characterized in that, Includes the following steps: Step 1, Preprocessing stage: a. Ball mill the lead-zinc smelting flue dust to control the particle size to be less than 74μm; b. For flue dust with a sulfur content >2.5 wt%, roasting should be carried out in a weak oxidizing atmosphere of 500-600℃, O2 volume concentration of 3-7%, and CO≤0.5%, with a heating rate of 4-6℃ / min and holding for 0.5-3 h; when the sulfur content of the flue dust is ≤2.5%, roasting should be omitted and the flue dust should be directly subjected to neutral leaching. Step 2, Leaching Stage: a. Neutral leaching: The pretreated soot is added to a sulfuric acid solution with a pH of 3.0-3.5 and leached at room temperature with stirring for 2 hours. The solid and liquid are separated to obtain a neutral leachate and a neutral leachate residue. b. Low acid leaching: Add the neutral leaching residue to a solution with a sulfuric acid concentration of 30-40 g / L, control the liquid-solid ratio to 5:1, stir and leach at 80-90℃, and use ultrasonic assisted leaching with a power of 300W for 2 hours. The solid and liquid are separated to obtain low acid leaching solution and low acid leaching residue. c. High acid leaching: Add the low acid leaching residue to a sulfuric acid solution with a concentration of 80 g / L, and leach at 80-90℃ for 4 hours to separate the high acid leaching solution and the high acid leaching residue. The high acid leaching solution is returned to the low acid leaching step for recycling. Step 3, Extraction and Separation Stage: a. Primary extraction: Using 10 vol% P204 as the extractant and sulfonated kerosene as the diluent, the low-acid leachate was subjected to a two-stage countercurrent extraction under the condition of A / O ratio of 2 / 1. The mixing time was controlled at 5 minutes, and the indium-loaded P204 organic phase and the primary raffinate were separated. b. Primary back-extraction: The indium-loaded P204 organic phase is back-extracted with 6 mol / L hydrochloric acid to obtain indium chloride back-extraction solution; c. Secondary extraction: A mixed extractant system of 20 vol% TBP + 15 vol% P350 is used to perform two-stage countercurrent extraction on the indium chloride back-extraction solution under the condition of A / O ratio of 3 / 2, to separate the indium-loaded mixed organic phase and the secondary raffinate. The secondary raffinate is returned to the primary back-extraction step for recycling. d. Secondary back-extraction: The indium-loaded mixed organic phase is subjected to three-stage countercurrent back-extraction with pure water to obtain a high-purity indium-rich solution; Step 4, Refining Stage: a. Displacement: The indium-rich solution is replaced with a zinc plate with a purity of not less than 99.9% at 50-60℃ to obtain sponge indium; b. Electrolytic refining: The sponge indium briquette is pressed and cast into a crude indium anode. An acidic electrolyte is used to control the pH at 2.0-2.5, the temperature at 25-30℃, and the current density at 80-100A / m² for electrolytic refining to obtain a refined indium product with a purity of over 99.99%.

2. The method for efficiently extracting rare and dispersed metal indium from lead-zinc smelting flue dust according to claim 1, characterized in that: In the pretreatment stage, the O2 volume concentration of the weak oxidizing atmosphere is 4–6%, the CO volume concentration is ≤0.3%, the heating rate is 5 ℃ / min, and the holding time is 2 h.

3. The method for efficiently extracting rare and dispersed metal indium from lead-zinc smelting flue dust according to claim 1, characterized in that: In the low-acid leaching step, the pulsed ultrasonic frequency is 22±1kHz, the power density is 18±2W / L, the duty cycle is 50%, the interval is 0.5s, and the ultrasonication and stirring are carried out synchronously.

4. The method for efficiently extracting rare and dispersed metal indium from lead-zinc smelting flue dust according to claim 1, characterized in that: In the high acid leaching step, the amount of high acid leaching solution returned accounts for 80-100% of the total acid content of the low acid leaching solution, and suspended solids are removed by pressure filtration before being returned.

5. The method for efficiently extracting rare and dispersed indium from lead-zinc smelting flue dust according to claim 1, characterized in that: In the secondary extraction step, the amount of secondary raffinate returned accounts for 60-80% of the volume of 6 mol / L hydrochloric acid required for the primary back-extraction, and any shortfall is supplemented with industrial hydrochloric acid.

6. The method for efficiently extracting rare and dispersed metal indium from lead-zinc smelting flue dust according to claim 1, characterized in that: Germanium is recovered from the primary raffinate using a synergistic extraction process of P204 and Yw-100, and zinc and iron are recovered from each leachate.

7. The method for efficiently extracting rare and dispersed indium from lead-zinc smelting flue dust according to claim 6, characterized in that: Yw-100 is a copper extractant of hydroxamic acid, with the active ingredient being 2-hydroxy-5-nonylacetophenone oxime (CAS 34138-71-1). Before use, it should be diluted with sulfonated kerosene to a volume fraction of 10 vol% and mixed with P204 at a volume ratio of 1:

1.

8. The method for efficiently extracting rare and dispersed metal indium from lead-zinc smelting flue dust according to claim 1, characterized in that: In the replacement step, the replacement temperature is 55 ℃ and the zinc plate purity is ≥99.95%.

9. The method for efficiently extracting rare and dispersed metal indium from lead-zinc smelting flue dust according to claim 1, characterized in that: In the electrolytic refining step, the current density is 90 A / m², and the In content of the electrolyte is... 3+ With a concentration of 80-100 g / L and NaCl ≤ 0.1 g / L, the cathode yields indium with a purity ≥ 99.993%.

10. The method according to any one of claims 1-9, characterized in that: The method described above results in a chloride ion emission of ≤600 mg / kg dry flue ash, a reduction of ≥86% compared to traditional methods, a total indium leaching rate of 88%-92%, a total indium recovery rate of 85%-90%, and a refined indium product purity of ≥99.99%.

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