CN116768246A - A method for efficiently separating lithium and aluminum from aluminum electrolyte waste residue and enriching lithium - Google Patents

Abstract

The invention belongs to the technical field of solid waste treatment, and provides a method for efficiently separating and enriching lithium from lithium and aluminum in aluminum electrolyte waste residues, which comprises the following steps: mixing aluminum electrolyte waste residue and aluminum sulfate, and performing primary low-temperature roasting and secondary high-temperature thermal decomposition to obtain a roasting material and sulfur-containing flue gas; introducing sulfur-containing flue gas into an alkali solution to obtain a sulfate solution; leaching the roasting material with water, and filtering to obtain water leaching slag and sodium-potassium-containing lithium sulfate solution; and (3) repeating the water leaching process for a plurality of times on the lithium sulfate solution containing sodium and potassium to obtain the lithium-rich sulfate solution. The invention ensures the high-efficiency separation of lithium and aluminum in the aluminum electrolyte waste residue, realizes the recycling of aluminum fluoride, obtains the lithium-rich sulfate solution, and provides convenience for the subsequent high-efficiency recovery of high-added-value lithium salt products. The method has the advantages of simple process flow, low equipment requirement and strong practicability.

Description

A method for efficiently separating lithium and aluminum from aluminum electrolyte waste residue and enriching lithium

Technical field

The present invention relates to the technical field of solid waste treatment, and in particular to a method for efficiently separating lithium and aluminum from aluminum electrolyte waste residue and enriching lithium.

Background technique

my country's diaspore-type bauxite contains alkali metal impurities such as lithium, which are enriched in the aluminum electrolyte through the "alumina-aluminum electrolysis" production process, seriously affecting the electrolytic production of aluminum. Lithium-rich aluminum electrolyte waste residue is a unique non-traditional lithium resource produced by the development and utilization of lithium-containing diaspore-type bauxite. The existence form and mineral structure of lithium are not yet clear, and existing lithium extraction technology cannot be applied. At present, aluminum electrolyte waste residue is mainly stored and disposed, which not only poses the risk of fluorine pollution, but also cannot effectively utilize the strategic resource lithium.

Among the existing technologies, CN113981232A discloses a direct leaching and recovery method of aluminum sulfate for lithium elements in aluminum electrolyte waste residue. The method is to leach lithium-containing aluminum electrolyte waste residue through aluminum sulfate solution, and then obtain lithium carbonate through alkaline hydrolysis reaction, carbonization reaction, and pyrolysis in sequence; this method uses aluminum sulfate solution to leach lithium-containing fluoride with a low phase conversion rate and a lengthy process. And it produces Al(OH) ₃ colloid that is difficult to separate and affects subsequent lithium extraction. CN105293536A discloses a method for extracting lithium from electrolytic aluminum waste residue. The method is to roast electrolytic aluminum waste residue through concentrated sulfuric acid roasting, sodium carbonate alkaline hydrolysis, lime causticization and carbonization reaction to obtain lithium carbonate; this method uses concentrated sulfuric acid for roasting, which has high equipment requirements, complicated processes, and produces Al(OH which is difficult to separate). ) ₃ colloid and affect subsequent lithium extraction. CN105349786A discloses a method for comprehensive recycling of lithium-containing aluminum electrolytes. The method is to mix the lithium-containing aluminum electrolyte with water and use inorganic acid to adjust the pH <2, add aluminum salt to prepare cryolite, and then proceed through sodium carbonate alkaline hydrolysis, CO ₂ carbonization, cation exchange resin impurity removal, and CO ₂ removal to obtain a high-quality electrolyte. Pure lithium carbonate. This method uses cation exchange resin to remove impurities, which is too costly and produces Al(OH) ₃ colloid that is difficult to separate and affects subsequent lithium extraction.

At present, the resource utilization technology for fluoride electrolyte waste residue in the aluminum electrolysis industry is not yet mature, and there is no method to purify lithium for lithium and aluminum separation, causing the existing lithium extraction process to be lengthy and complex. Therefore, carrying out research on the clean recovery of lithium-rich aluminum electrolyte waste residue and efficient lithium enrichment technology can not only avoid the risk of large amounts of electrolyte waste residue accumulation and fluorine pollution leakage due to the accumulation of alkali metals such as lithium in the aluminum electrolysis industry, but also alleviate the risk of new energy sources. The security issues of lithium resources brought about by the rapid development are of great significance to strengthening the clean production of the aluminum electrolysis industry, promoting the comprehensive utilization of bulk hazardous waste, and supporting the development of the new energy automobile industry.

Contents of the invention

The object of the present invention is to provide a method for efficiently separating lithium and aluminum from aluminum electrolyte waste residue and enriching lithium in view of the shortcomings of the existing technology.

In order to achieve the above-mentioned object of the invention, the present invention provides the following technical solutions:

The invention provides a method for efficiently separating lithium and aluminum from aluminum electrolyte waste residue and enriching lithium, which includes the following steps:

1) Mix aluminum electrolyte waste residue and aluminum sulfate, perform one stage of low-temperature roasting and two stages of high-temperature thermal decomposition, and obtain roasting materials and sulfur-containing flue gas;

2) Pass the sulfur-containing flue gas into the alkali solution to obtain a sulfate solution;

3) Leach the roasted material with water and then filter it to obtain water leaching residue and a lithium sulfate solution containing sodium and potassium;

4) Return the lithium sulfate solution containing sodium and potassium to the water leaching process in step 3), and obtain a lithium-rich sulfate solution after multiple cycles of leaching.

Preferably, the mass ratio of the aluminum electrolyte waste residue and aluminum sulfate described in step 1) is 1:0.5-1.5.

Preferably, the temperature of one stage of low-temperature roasting in step 1) is 200-600°C, and the time of one stage of low-temperature roasting is 0.5-5 hours.

Preferably, the temperature of the second-stage high-temperature thermal decomposition in step 1) is 700-1000°C, and the time of the second-stage high-temperature thermal decomposition is 0.5-3 hours.

Preferably, the concentration of the alkali solution in step 2) is 50-300g/L.

Preferably, in step 3), the mass volume ratio of the roasted material and water is 1g:3-15mL.

Preferably, the leaching temperature in step 3) is 30 to 95°C, and the leaching time is 0.5 to 5 hours.

Preferably, the number of multiple cycles of leaching in step 4) is 5 to 20 times.

The beneficial effects of the present invention include the following points:

1) The present invention provides a method for efficiently separating lithium and aluminum and enriching lithium through two-stage roasting and decomposition of aluminum electrolyte waste residue and aluminum sulfate - water leaching. The invention promotes the conversion of lithium, sodium, potassium and calcium fluoride in electrolyte waste residue into sulfate phase and aluminum fluoride through two stages of roasting and decomposition-water immersion, and selectively high-temperature thermal decomposition of unreacted aluminum sulfate to achieve lithium-containing sodium and potassium. The efficient leaching and separation of soluble sulfate and the insoluble aluminum-containing phase (Al ₂ O ₃ and AlF ₃ ) in the water leaching residue avoids the impact of aluminum sulfate hydrolysis on lithium recovery. The enriched lithium-rich sulfate solution is It provides convenience for subsequent efficient recovery of metallic lithium.

2) The present invention realizes resource recycling of aluminum fluoride by returning water-leached slag (Al ₂ O ₃ and AlF ₃ ) to the aluminum electrolysis system, thereby avoiding the risk of the spread of fluorine pollution. While ensuring efficient separation of lithium and aluminum from aluminum electrolyte waste residue, the present invention also realizes resource recycling of aluminum fluoride and obtains a lithium-rich sulfate solution, which facilitates the subsequent efficient recovery of high value-added metal lithium.

3) The method of the present invention has a simple process flow, low equipment requirements and strong practicability.

Description of drawings

Figure 1 is a schematic process flow diagram of the efficient separation of lithium and aluminum from aluminum electrolyte waste residue and the enrichment of lithium according to the present invention.

Detailed ways

The invention provides a method for efficiently separating lithium and aluminum from aluminum electrolyte waste residue and enriching lithium, which includes the following steps:

1) Mix aluminum electrolyte waste residue and aluminum sulfate, perform one stage of low-temperature roasting and two stages of high-temperature thermal decomposition, and obtain roasting materials and sulfur-containing flue gas;

2) Pass the sulfur-containing flue gas into the alkali solution to obtain a sulfate solution;

3) Leach the roasted material with water and then filter it to obtain water leaching residue and a lithium sulfate solution containing sodium and potassium;

4) Return the lithium sulfate solution containing sodium and potassium to the water leaching process in step 3), and obtain a lithium-rich sulfate solution after multiple cycles of leaching.

In the present invention, the aluminum electrolyte waste residue in step 1) is preferably the aluminum electrolyte waste residue obtained by crushing treatment. The mesh number of the aluminum electrolyte waste residue is preferably 100 to 400 mesh, more preferably 150 to 350 mesh, and more preferably 200 to 300 mesh. ; Aluminum electrolyte waste residue is lithium-rich aluminum electrolyte waste residue.

In the present invention, the mass ratio of the aluminum electrolyte waste residue and aluminum sulfate described in step 1) is preferably 1:0.5-1.5, more preferably 1:0.7-1.3, and more preferably 1:0.8-1.2.

In the present invention, the temperature of one stage of low-temperature roasting in step 1) is preferably 200~600°C, more preferably 250~550°C, more preferably 300~500°C; the time of one stage of low-temperature roasting is preferably 0.5~5h, further preferably It is 1-4h, more preferably, it is 2-3h.

In the present invention, the temperature of the second-stage high-temperature thermal decomposition in step 1) is preferably 700-1000°C, further preferably 750-950°C, more preferably 800-900°C, and the time of the second-stage high-temperature thermal decomposition is preferably 0.5-1000°C. 3h, more preferably 1 to 2.5h, more preferably 1.5 to 2h.

In the present invention, the concentration of the alkali solution in step 2) is preferably 50-300g/L, more preferably 100-250g/L, and more preferably 150-200g/L.

In the present invention, the alkali solution in step 2) is preferably an aqueous sodium hydroxide solution.

In step 3) of the present invention, the mass-volume ratio of the roasting material and water is preferably 1g:3-15mL, more preferably 1g:5-12mL, and more preferably 1g:8-10mL.

In the present invention, the leaching temperature in step 3) is preferably 30-95°C, more preferably 40-90°C, more preferably 50-85°C; the leaching time is preferably 0.5-5h, further preferably 1-4h , more preferably 2 to 3 hours.

In the present invention, the water leaching slag in step 3) contains AlF ₃ and Al ₂ O ₃ , and the water leaching slag is returned to the aluminum electrolysis system.

In the present invention, the number of multiple cycles of leaching in step 4) is preferably 5 to 20 times, more preferably 6 to 18 times, and more preferably 8 to 15 times; the lithium-rich sulfate solution is used to purify lithium precipitation. .

The schematic process flow diagram of the efficient separation of lithium and aluminum from the aluminum electrolyte waste residue of the present invention and the enrichment of lithium is shown in Figure 1, in which the lithium-rich aluminum electrolyte waste residue is the aluminum electrolyte waste residue of the present invention.

The technical solutions provided by the present invention will be described in detail below with reference to the examples, but they should not be understood as limiting the protection scope of the present invention.

Example 1

100g of aluminum electrolyte waste residue from an electrolytic aluminum plant and 70g of aluminum sulfate were mixed evenly after being crushed to 400 mesh. The mixture was subjected to a low-temperature roasting at 300°C (low-temperature roasting time is 4 hours) and then was subjected to a second-stage high-temperature thermal decomposition (high-temperature thermal decomposition) at 750°C. The time is 3h), and the roasted material and sulfur-containing flue gas are obtained. Pass the sulfur-containing flue gas into a sodium hydroxide solution with a concentration of 100g/L, and obtain a sodium sulfate solution after absorption. Mix the roasted material and water (the dosage ratio of roasted material and water is 1g:5mL), stir and leach for 4 hours at a temperature of 75°C and a rotation speed of 500r/min, then filter to obtain the water leaching residue and lithium sulfate containing sodium and potassium. solution.

No Al ³⁺ was detected in the lithium sulfate solution containing sodium and potassium in this example. The leaching rates of Na and K were 95.45% and 96.12% respectively. The leaching rate of Li was 96.48%. The lithium sulfate solution containing sodium and potassium was in the same After 12 cycles of water immersion under the conditions, the lithium content in the obtained lithium-rich sulfate solution was 15g/L.

Example 2

100g of aluminum electrolyte waste residue from an electrolytic aluminum plant and 90g of aluminum sulfate were mixed evenly after crushing and processing to 300 mesh. The mixture was subjected to a low-temperature roasting at 350°C (low-temperature roasting time is 3 hours) and then was subjected to a second-stage high-temperature thermal decomposition (high-temperature thermal decomposition) at 800°C. The time is 2.5h) to obtain roasted materials and sulfur-containing flue gas. Pass the sulfur-containing flue gas into a sodium hydroxide solution with a concentration of 150g/L, and obtain a sodium sulfate solution after absorption. Mix the roasted material and water (the dosage ratio of roasted material and water is 1g:7mL), stir and leach for 2.5 hours at a temperature of 80°C and a rotation speed of 600r/min, then filter to obtain the water leaching residue and sulfuric acid containing sodium and potassium. Lithium solution.

No Al ³⁺ was detected in the lithium sulfate solution containing sodium and potassium in this example. The leaching rates of Na and K were 96.43% and 96.97% respectively. The leaching rate of Li was 98.21%. The lithium sulfate solution containing sodium and potassium was in the same After 10 cycles of water immersion under the conditions, the lithium content in the obtained lithium-rich sulfate solution was 12g/L.

Example 3

100g of aluminum electrolyte waste residue from an electrolytic aluminum plant and 120g of aluminum sulfate were mixed evenly after crushing and processing to 200 mesh. The mixture was subjected to a low-temperature roasting at 400°C (low-temperature roasting time is 2 hours) and then was subjected to a second-stage high-temperature thermal decomposition (high-temperature thermal decomposition) at 850°C. The time is 1h), and the roasted material and sulfur-containing flue gas are obtained. Pass the sulfur-containing flue gas into a sodium hydroxide solution with a concentration of 175g/L, and obtain a sodium sulfate solution after absorption. Mix the roasted material and water (the dosage ratio of roasted material and water is 1g:9mL), stir and leach for 2 hours at a temperature of 85°C and a rotation speed of 700r/min, then filter to obtain the water leaching residue and lithium sulfate containing sodium and potassium. solution.

No Al ³⁺ was detected in the lithium sulfate solution containing sodium and potassium in this example. The leaching rates of Na and K were 97.33% and 97.67% respectively. The leaching rate of Li was 96.59%. The lithium sulfate solution containing sodium and potassium was in the same After 8 cycles of water immersion under the conditions, the lithium content in the obtained lithium-rich sulfate solution was 10g/L.

Example 4

Mix 100g of aluminum electrolyte waste residue from an electrolytic aluminum plant to 150 mesh and 140g of aluminum sulfate. The mixture is subjected to a low-temperature roasting at 500°C (low-temperature roasting time is 1 hour) and then a second-stage high-temperature thermal decomposition (high-temperature thermal decomposition) at 900°C. The time is 1.5h) to obtain roasted material and sulfur-containing flue gas. Pass the sulfur-containing flue gas into a sodium hydroxide solution with a concentration of 225g/L, and obtain a sodium sulfate solution after absorption. Mix the roasted material and water (the dosage ratio of roasted material and water is 1g:12mL), stir and leach for 3 hours at a temperature of 50°C and a rotation speed of 700r/min, then filter to obtain the water leaching residue and lithium sulfate containing sodium and potassium. solution.

No Al ³⁺ was detected in the lithium sulfate solution containing sodium and potassium in this example. The leaching rates of Na and K were 96.89% and 97.32% respectively. The leaching rate of Li was 97.89%. The lithium sulfate solution containing sodium and potassium was in the same After 15 cycles of water immersion under the conditions, the lithium content in the obtained lithium-rich sulfate solution was 14g/L.

The invention provides a method for efficiently separating lithium and aluminum and enriching lithium through two-stage roasting and decomposition of aluminum electrolyte waste residue and aluminum sulfate-water leaching. The present invention promotes the conversion of lithium, sodium, potassium and calcium fluoride in electrolyte waste residue into sulfate phase and aluminum fluoride through two stages of roasting and decomposition-water leaching, and selectively high-temperature thermal decomposition of unreacted aluminum sulfate to achieve lithium-containing sodium and potassium. Efficient leaching and separation of soluble sulfate and insoluble aluminum-containing phases (Al ₂ O ₃ and AlF ₃ ) avoids the impact of aluminum sulfate hydrolysis on lithium recovery. The enriched lithium-rich sulfate solution provides the subsequent efficient recovery of metals. Lithium provides convenience. The present invention realizes resource recycling of aluminum fluoride by returning water-leached slag (Al ₂ O ₃ and AlF ₃ ) to the aluminum electrolysis system, thereby avoiding the risk of fluorine pollution diffusion. While ensuring efficient separation of lithium and aluminum from aluminum electrolyte waste residue, the present invention also realizes resource recycling of aluminum fluoride and obtains a lithium-rich sulfate solution, which facilitates the subsequent efficient recovery of high value-added metal lithium. The method of the invention has simple process flow, low equipment requirements and strong practicability.

The above are only preferred embodiments of the present invention. It should be noted that those skilled in the art can make several improvements and modifications without departing from the principles of the present invention. These improvements and modifications can also be made. should be regarded as the protection scope of the present invention.

Claims (8)

1. The method for efficiently separating and enriching lithium from lithium and aluminum in the aluminum electrolyte waste residue is characterized by comprising the following steps of:

1) Mixing aluminum electrolyte waste residue and aluminum sulfate, and performing primary low-temperature roasting and secondary high-temperature thermal decomposition to obtain a roasting material and sulfur-containing flue gas;

2) Introducing sulfur-containing flue gas into an alkali solution to obtain a sulfate solution;

3) Leaching the roasting material with water, and filtering to obtain water leaching slag and sodium-potassium-containing lithium sulfate solution;

4) And returning the lithium sulfate solution containing sodium and potassium to the water leaching process of the step 3), and circularly leaching for a plurality of times to obtain the lithium-rich sulfate solution.

2. The method according to claim 1, wherein the mass ratio of the aluminum electrolyte waste residue to aluminum sulfate in step 1) is 1:0.5 to 1.5.

3. The method according to claim 2, wherein the temperature of the one-stage low-temperature calcination in step 1) is 200 to 600 ℃ and the time of the one-stage low-temperature calcination is 0.5 to 5 hours.

4. A method according to claim 2 or 3, wherein the temperature of the two-stage pyrolysis in step 1) is 700 to 1000 ℃ and the time of the two-stage pyrolysis is 0.5 to 3 hours.

5. The method according to claim 4, wherein the concentration of the alkaline solution in step 2) is 50 to 300g/L.

6. The method according to claim 4, wherein in the step 3), the mass-to-volume ratio of the roasting material to the water is 1g: 3-15 mL.

7. The method according to claim 5 or 6, wherein the leaching in step 3) is carried out at a temperature of 30-95 ℃ for a period of 0.5-5 hours.

8. The method of claim 7, wherein the number of times of the multiple cycle leaching of step 4) is 5 to 20.