Method for lithium-aluminum efficient separation and lithium enrichment of aluminum electrolyte waste residue

Abstract

The present application belongs to the technical field of solid waste treatment, and provides a method for efficient separation of lithium and aluminum and enrichment of lithium from aluminum electrolyte waste residue, comprising the following steps: mixing the aluminum electrolyte waste residue and aluminum sulfate, performing one-stage low-temperature calcination and two-stage high-temperature thermal decomposition to obtain calcined material and sulfur-containing flue gas; passing the sulfur-containing flue gas into an alkali solution to obtain a sulfate solution; filtering the calcined material after leaching with water to obtain water leaching residue and sodium-potassium-containing lithium sulfate solution; repeatedly performing the water leaching process on the sodium-potassium-containing lithium sulfate solution to obtain a lithium-enriched sulfate solution. The present application realizes efficient separation of lithium and aluminum in the aluminum electrolyte waste residue while recycling and utilizing aluminum fluoride resources, and obtains a lithium-enriched sulfate solution, which provides convenience for subsequent efficient recovery of high-value lithium salt products. The method of the present application has a simple process flow, low equipment requirements and strong practicability.

Description

Technical Field

[0001] This invention relates to the field of solid waste treatment technology, and in particular to a method for efficient separation and enrichment of lithium from aluminum electrolyte waste residue. Background Technology

[0002] my country's monohydrate gibbsite bauxite contains alkali metal impurities such as lithium, which accumulate in the aluminum electrolyte during the "alumina-aluminum electrolysis" production process, severely impacting aluminum electrolytic production. Lithium-rich aluminum electrolyte waste is a unique, non-traditional lithium resource generated from the development and utilization of lithium-containing monohydrate gibbsite bauxite. The forms and mineral structures of lithium are still unclear, and existing lithium extraction technologies are not applicable. Currently, aluminum electrolyte waste is mainly disposed of through stockpiling, which not only poses a risk of fluoride pollution but also prevents the effective utilization of the strategic lithium resource.

[0003] In existing technologies, CN113981232A discloses a method for direct leaching and recovery of lithium from aluminum electrolyte waste residue using aluminum sulfate. This method involves leaching the lithium-containing aluminum electrolyte waste residue with aluminum sulfate solution, followed by a series of alkaline reactions, carbonation reactions, and pyrolysis to obtain lithium carbonate. However, this method suffers from low phase conversion rates when using aluminum sulfate solution to leach lithium-containing fluorides, a lengthy process, and the generation of difficult-to-separate Al(OH)3 colloids that affect subsequent lithium extraction. CN105293536A discloses a method for extracting lithium from electrolytic aluminum waste residue. This method involves roasting the electrolytic aluminum waste residue with concentrated sulfuric acid, alkaline hydrolysis with sodium carbonate, causticization with lime, and a carbonation reaction to obtain lithium carbonate. This method uses concentrated sulfuric acid for roasting, requires sophisticated equipment, has a complex process, and generates difficult-to-separate Al(OH)3 colloids that affect subsequent lithium extraction. CN105349786A discloses a method for comprehensive recovery and utilization of lithium-containing aluminum electrolyte. The method involves mixing lithium-containing aluminum electrolyte with water and adjusting the pH to <2 using an inorganic acid. Aluminum salts are then added to prepare cryolite, followed by alkaline hydrolysis with sodium carbonate, CO2 carbonation, impurity removal using cation exchange resin, and CO2 removal treatment to obtain high-purity lithium carbonate. However, this method uses cation exchange resin for impurity removal, which is too costly and produces difficult-to-separate Al(OH)3 colloids that affect subsequent lithium extraction.

[0004] Currently, the technology for resource utilization of fluoride electrolyte waste residue from the aluminum electrolysis industry is not yet mature, and there is no method for purifying lithium to separate lithium from aluminum. This results in existing lithium extraction processes being lengthy and complex. Therefore, research on clean recycling and efficient lithium enrichment technology for lithium-rich aluminum electrolyte waste residue can avoid the risks of large-scale electrolyte waste residue accumulation and fluoride pollution leaks caused by the accumulation of lithium and other alkali metals in the aluminum electrolysis industry. It can also alleviate the lithium resource security issues brought about by the rapid development of new energy sources. This is of great significance for strengthening clean production in the aluminum electrolysis industry, promoting the comprehensive utilization of bulk hazardous waste, and supporting the development of the new energy vehicle industry. Summary of the Invention

[0005] The purpose of this invention is to address the shortcomings of existing technologies by providing a method for efficient separation and enrichment of lithium from aluminum electrolyte waste residue.

[0006] To achieve the above-mentioned objectives, the present invention provides the following technical solution:

[0007] This invention provides a method for efficient separation and enrichment of lithium from aluminum electrolyte waste residue, comprising the following steps:

[0008] 1) Mix aluminum electrolyte waste residue and aluminum sulfate, and carry out a first stage of low-temperature roasting and a second stage of high-temperature thermal decomposition to obtain roasted material and sulfur-containing flue gas;

[0009] 2) Sulfur-containing flue gas is passed into an alkaline solution to obtain a sulfate solution;

[0010] 3) After leaching the roasted material with water, filter it to obtain water-leached residue and a lithium sulfate solution containing sodium and potassium.

[0011] 4) Return the sodium and potassium-containing lithium sulfate solution to the water leaching process in step 3), and after multiple cycles of leaching, obtain a lithium-rich sulfate solution.

[0012] Preferably, the mass ratio of aluminum electrolyte waste residue to aluminum sulfate in step 1) is 1:0.5 to 1.5.

[0013] Preferably, the temperature of the first low-temperature roasting in step 1) is 200-600℃, and the roasting time is 0.5-5h.

[0014] Preferably, the temperature of the two-stage high-temperature thermal decomposition in step 1) is 700-1000℃, and the time of the two-stage high-temperature thermal decomposition is 0.5-3h.

[0015] Preferably, the concentration of the alkaline solution in step 2) is 50–300 g / L.

[0016] Preferably, in step 3), the mass-to-volume ratio of the calcining material to water is 1g:3-15mL.

[0017] Preferably, the leaching temperature in step 3) is 30–95°C, and the leaching time is 0.5–5 h.

[0018] Preferably, the number of leaching cycles in step 4) is 5 to 20.

[0019] The beneficial effects of this invention include the following:

[0020] 1) This invention provides a method for the efficient separation and enrichment of lithium from aluminum electrolyte waste residue and aluminum sulfate through a two-stage roasting-decomposition-water leaching process. This invention promotes the conversion of lithium sodium, potassium, and calcium fluorides in the electrolyte waste residue into sulfate phases and aluminum fluoride through a two-stage roasting-decomposition-water leaching process, and selectively decomposes unreacted aluminum sulfate at high temperatures. This achieves efficient leaching and separation of soluble lithium sodium and potassium sulfates from the insoluble aluminum-containing phases (Al2O3 and AlF3) in the water-leached residue, avoiding the impact of aluminum sulfate hydrolysis on lithium recovery. The enriched lithium-rich sulfate solution facilitates subsequent efficient recovery of metallic lithium.

[0021] 2) This invention achieves the resource recycling of aluminum fluoride by returning the water-leached residue (Al2O3 and AlF3) to the aluminum electrolysis system, thus avoiding the risk of fluoride pollution spread. While ensuring efficient separation of lithium and aluminum in the aluminum electrolyte waste residue, this invention also achieves the resource recycling of aluminum fluoride, obtaining a lithium-rich sulfate solution, which facilitates the subsequent efficient recovery of high-value-added metallic lithium.

[0022] 3) The method of the present invention has a simple process flow, low equipment requirements, and strong practicality. Attached Figure Description

[0023] Figure 1 This is a schematic diagram of the process flow for efficient lithium-aluminum separation and lithium enrichment from aluminum electrolyte waste residue of the present invention. Detailed Implementation

[0024] This invention provides a method for efficient separation and enrichment of lithium from aluminum electrolyte waste residue, comprising the following steps:

[0025] 1) Mix aluminum electrolyte waste residue and aluminum sulfate, and carry out a first stage of low-temperature roasting and a second stage of high-temperature thermal decomposition to obtain roasted material and sulfur-containing flue gas;

[0026] 2) Sulfur-containing flue gas is passed into an alkaline solution to obtain a sulfate solution;

[0027] 3) After leaching the roasted material with water, filter it to obtain water-leached residue and a lithium sulfate solution containing sodium and potassium.

[0028] 4) Return the sodium and potassium-containing lithium sulfate solution to the water leaching process in step 3), and after multiple cycles of leaching, obtain a lithium-rich sulfate solution.

[0029] In this invention, the aluminum electrolyte waste residue in step 1) is preferably aluminum electrolyte waste residue obtained by crushing treatment, and the mesh size of the aluminum electrolyte waste residue is preferably 100-400 mesh, more preferably 150-350 mesh, and even more preferably 200-300 mesh; the aluminum electrolyte waste residue is lithium-rich aluminum electrolyte waste residue.

[0030] In this invention, the mass ratio of aluminum electrolyte waste residue and aluminum sulfate in step 1) is preferably 1:0.5 to 1.5, more preferably 1:0.7 to 1.3, and even more preferably 1:0.8 to 1.2.

[0031] In this invention, the temperature of the first stage of low-temperature roasting in step 1) is preferably 200-600℃, more preferably 250-550℃, and even more preferably 300-500℃; the time of the first stage of low-temperature roasting is preferably 0.5-5h, more preferably 1-4h, and even more preferably 2-3h.

[0032] In this invention, the temperature of the two-stage high-temperature thermal decomposition in step 1) is preferably 700-1000℃, more preferably 750-950℃, and even more preferably 800-900℃. The time of the two-stage high-temperature thermal decomposition is preferably 0.5-3h, more preferably 1-2.5h, and even more preferably 1.5-2h.

[0033] In this invention, the concentration of the alkaline solution in step 2) is preferably 50-300 g / L, more preferably 100-250 g / L, and even more preferably 150-200 g / L.

[0034] In this invention, the alkaline solution in step 2) is preferably an aqueous solution of sodium hydroxide.

[0035] In step 3) of the present invention, the mass-volume ratio of the calcining material to water is preferably 1g:3-15mL, more preferably 1g:5-12mL, and even more preferably 1g:8-10mL.

[0036] In this invention, the leaching temperature in step 3) is preferably 30-95°C, more preferably 40-90°C, and even more preferably 50-85°C; the leaching time is preferably 0.5-5h, more preferably 1-4h, and even more preferably 2-3h.

[0037] In this invention, the water-leached residue in step 3) contains AlF3 and Al2O3, and the water-leached residue is returned to the aluminum electrolysis system.

[0038] In this invention, the number of leaching cycles in step 4) is preferably 5 to 20 times, more preferably 6 to 18 times, and even more preferably 8 to 15 times; the lithium-rich sulfate solution is used to purify lithium precipitation.

[0039] A schematic diagram of the process flow for efficient lithium-aluminum separation and lithium enrichment from aluminum electrolyte waste residue of the present invention is shown below. Figure 1 As shown, the lithium-rich aluminum electrolyte waste residue is the aluminum electrolyte waste residue of this invention.

[0040] The technical solutions provided by the present invention will be described in detail below with reference to the embodiments, but they should not be construed as limiting the scope of protection of the present invention.

[0041] Example 1

[0042] 100g of aluminum electrolyte waste residue from an electrolytic aluminum plant, crushed to 400 mesh, and 70g of aluminum sulfate were mixed evenly. The mixture was subjected to a first-stage low-temperature roasting at 300℃ (roasting time: 4h), followed by a second-stage high-temperature thermal decomposition at 750℃ (high-temperature thermal decomposition time: 3h), yielding roasted material and sulfur-containing flue gas. The sulfur-containing flue gas was passed into a 100g / L sodium hydroxide solution, and after absorption, a sodium sulfate solution was obtained. The roasted material was mixed with water (roasted material to water ratio: 1g:5mL) and leached at 75℃ and 500r / min for 4h. The leached residue and a sodium-potassium lithium sulfate solution were obtained by filtration.

[0043] In this embodiment, no Al was detected in the sodium-potassium lithium sulfate solution. 3+ The leaching rates of Na and K were 95.45% and 96.12%, respectively, and the leaching rate of Li was 96.48%. After 12 cycles of water leaching under the same conditions, the lithium-rich sulfate solution containing sodium and potassium sulfate contained 15 g / L.

[0044] Example 2

[0045] 100g of aluminum electrolyte waste residue from an electrolytic aluminum plant, crushed to 300 mesh, and 90g of aluminum sulfate were mixed evenly. The mixture was subjected to a first-stage low-temperature roasting at 350℃ (roasting time: 3h), followed by a second-stage high-temperature thermal decomposition at 800℃ (high-temperature thermal decomposition time: 2.5h), yielding roasted material and sulfur-containing flue gas. The sulfur-containing flue gas was passed into a 150g / L sodium hydroxide solution, and after absorption, a sodium sulfate solution was obtained. The roasted material was mixed with water (roasted material to water ratio: 1g:7mL), and then leached by stirring at 80℃ and 600r / min for 2.5h. The leached residue and a sodium-potassium lithium sulfate solution were obtained by filtration.

[0046] In this embodiment, no Al was detected in the sodium-potassium lithium sulfate solution. 3+ The leaching rates of Na and K were 96.43% and 96.97%, respectively, and the leaching rate of Li was 98.21%. After 10 cycles of water leaching under the same conditions, the lithium-rich sulfate solution containing sodium and potassium sulfate contained 12 g / L.

[0047] Example 3

[0048] 100g of aluminum electrolyte waste residue from an electrolytic aluminum plant, crushed to 200 mesh, and 120g of aluminum sulfate were mixed evenly. The mixture was subjected to a first-stage low-temperature roasting at 400℃ (roasting time: 2h), followed by a second-stage high-temperature thermal decomposition at 850℃ (high-temperature thermal decomposition time: 1h), yielding roasted material and sulfur-containing flue gas. The sulfur-containing flue gas was passed into a sodium hydroxide solution with a concentration of 175g / L, and after absorption, a sodium sulfate solution was obtained. The roasted material was mixed with water (roasted material to water ratio: 1g:9mL) and leached by stirring at 85℃ and a rotation speed of 700r / min for 2h. The leached residue and a sodium-potassium lithium sulfate solution were obtained by filtration.

[0049] In this embodiment, no Al was detected in the sodium-potassium lithium sulfate solution. 3+ The leaching rates of Na and K were 97.33% and 97.67%, respectively, and the leaching rate of Li was 96.59%. After the sodium and potassium-containing lithium sulfate solution was subjected to 8 cycles of water leaching under the same conditions, the lithium content in the resulting lithium-rich sulfate solution was 10 g / L.

[0050] Example 4

[0051] 100g of aluminum electrolyte waste residue from an electrolytic aluminum plant, crushed to 150 mesh, and 140g of aluminum sulfate were mixed evenly. The mixture was subjected to a first-stage low-temperature roasting at 500℃ (roasting time: 1 h), followed by a second-stage high-temperature thermal decomposition at 900℃ (high-temperature thermal decomposition time: 1.5 h), yielding roasted material and sulfur-containing flue gas. The sulfur-containing flue gas was passed into a 225g / L sodium hydroxide solution, and after absorption, a sodium sulfate solution was obtained. The roasted material was mixed with water (roasted material to water ratio: 1g:12mL) and leached by stirring at 50℃ and 700r / min for 3 h. The leached residue and a sodium-potassium lithium sulfate solution were obtained by filtration.

[0052] In this embodiment, no Al was detected in the sodium-potassium lithium sulfate solution. 3+ The leaching rates of Na and K were 96.89% and 97.32%, respectively, and the leaching rate of Li was 97.89%. After 15 cycles of water leaching under the same conditions, the lithium-rich sulfate solution containing sodium and potassium sulfate contained 14 g / L.

[0053] This invention provides a method for the efficient separation and enrichment of lithium from aluminum electrolyte waste residue and aluminum sulfate through a two-stage roasting-decomposition-water leaching process. The invention promotes the conversion of lithium, sodium, potassium, and calcium fluorides in the electrolyte waste residue into sulfate phases and aluminum fluoride through two-stage roasting-decomposition-water leaching, and selectively decomposes unreacted aluminum sulfate at high temperatures. This achieves efficient leaching and separation of soluble lithium-containing sodium and potassium sulfates from insoluble aluminum-containing phases (Al₂O₃ and AlF₃), avoiding the impact of aluminum sulfate hydrolysis on lithium recovery. The enriched lithium-rich sulfate solution facilitates subsequent efficient recovery of metallic lithium. The invention also achieves resource recycling of aluminum fluoride by returning the water-leached residue (Al₂O₃ and AlF₃) to the aluminum electrolysis system, avoiding the risk of fluoride pollution. While ensuring efficient separation of lithium and aluminum from the aluminum electrolyte waste residue, this invention also achieves resource recycling of aluminum fluoride, obtaining a lithium-rich sulfate solution, which facilitates subsequent efficient recovery of high-value-added metallic lithium. The method of this invention has a simple process flow, low equipment requirements, and strong practicality.

[0054] The above description is only a preferred embodiment of the present invention. It should be noted that for those skilled in the art, several improvements and modifications can be made without departing from the principle of the present invention, and these improvements and modifications should also be considered within the scope of protection of the present invention.

Claims

1. A method for efficient separation and enrichment of lithium from aluminum electrolyte waste residue, characterized in that, Includes the following steps: 1) Mix aluminum electrolyte waste residue and aluminum sulfate, and carry out a first stage of low-temperature roasting and a second stage of high-temperature thermal decomposition to obtain roasted material and sulfur-containing flue gas; 2) Sulfur-containing flue gas is passed into an alkaline solution to obtain a sulfate solution; 3) After leaching the roasted material with water, filter it to obtain water-leached residue and a lithium sulfate solution containing sodium and potassium. 4) Return the sodium and potassium-containing lithium sulfate solution to the water leaching process in step 3), and after multiple cycles of leaching, obtain a lithium-rich sulfate solution; Step 1) The temperature of the first stage of low-temperature roasting is 200-400℃, and the time of the first stage of low-temperature roasting is 0.5-5h; Step 1) The temperature of the two-stage high-temperature thermal decomposition is 750-1000℃, and the time of the two-stage high-temperature thermal decomposition is 0.5-3h; Step 3) The water-leached residue contains AlF3 and Al2O3.

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

5.

3. The method according to claim 2, characterized in that, The concentration of the alkaline solution in step 2) is 50–300 g / L.

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

5. The method according to claim 4, characterized in that, Step 3) The leaching temperature is 30-95℃, and the leaching time is 0.5-5h.

6. The method according to claim 5, characterized in that, Step 4) involves 5 to 20 cycles of leaching.

Images