Method for high-valued recycling of electrolytic aluminum waste residues

Abstract

The invention provides a method for high-valued recycling of electrolytic aluminum waste residues. The method mainly comprises the following steps: adding a roasting agent for low-temperature roasting, performing water leaching, and adding ammonium salt for aluminum precipitation; ammonium alum obtained by aluminum precipitation is used for obtaining an aluminum oxide product; the aluminum precipitation post-liquid obtained through aluminum precipitation is sequentially subjected to impurity removal, solid-liquid separation and lithium precipitation by adding a lithium precipitation agent, a lithium carbonate product is obtained, and in addition, a villiaumite product and high-carbon residues can be obtained through the method. According to the method, high-valued recycling of the F resource, the C resource, the Al resource and the Li resource in the electrolytic aluminum waste residues can be achieved, especially the Li recovery rate can reach 90% or above, the lithium carbonate product contains few impurities and is high in purity, and the method has the advantages of being wide in raw material adaptability, low in raw material requirement, free of waste water and waste gas emission, environmentally friendly and the like.

Description

Technical Field

[0001] The present invention relates to the technical field of resource recovery and utilization of hazardous solid wastes in the electrolytic aluminum industry, and particularly to a method for high-value recovery and utilization of electrolytic aluminum waste slag. Background Art

[0002] During the aluminum smelting process, a large amount of hazardous solid wastes such as spent electrolyte, overhaul slag, aluminum ash, and carbon slag are inevitably generated, and the environmental pollution problem is becoming increasingly prominent. These waste slags are rich in valuable components such as a large amount of fluorides, aluminum compounds, and carbon. How to turn waste into treasure and environmentally friendly, economically and efficiently recover high-value components such as fluorine, aluminum, lithium, and carbon in electrolytic aluminum waste slag has become a difficult problem to be solved urgently for the sustainable development of the electrolytic aluminum industry.

[0003] CN116715263A discloses a method for pollution-free treatment and recycling of electrolytic aluminum overhaul slag. In this method, the overhaul slag is mixed with concentrated sulfuric acid, and then low-temperature ripening is carried out to recover hydrogen fluoride for preparing cryolite, high-temperature roasting is carried out to recycle sulfuric acid to recover corundum sand, the water leaching solution is recovered to prepare lithium carbonate products, and the solution is crystallized to obtain sodium sulfate products. Although this method realizes the recycling of overhaul slag, the raw material applicability is poor. If the overhaul slag contains a large amount of carbon or other impurities such as calcium, magnesium, and iron, the corundum sand does not have the recycling value, and additional NaAlO2 needs to be supplemented to recover cryolite, increasing the cost of auxiliary materials.

[0004] CN117735585A discloses a method for enriching lithium and preparing aluminum fluoride from low-lithium-containing electrolytic aluminum overhaul slag. In this method, concentrated sulfuric acid and aluminum sulfate are mixed and reacted with the overhaul slag, the generated HF gas is volatilized, dust-removed and decontaminated, and then reacted with alumina to prepare anhydrous AlF3. The solution is adjusted with alkali to precipitate aluminum hydroxide and calcined to recover alumina, and sodium phosphate is used to precipitate lithium to recover lithium phosphate, achieving the purpose of recovering lithium, fluorine, and aluminum. However, the products generated during the iron precipitation and aluminum precipitation processes are all hydroxides, resulting in a large loss of lithium and a low overall lithium recovery rate.

[0005] The existing treatment and recycling processes for electrolytic aluminum waste slag are difficult to achieve the full recovery of valuable components and have a low recovery rate. Most processes can only treat and recycle a certain type of electrolytic aluminum waste slag, with low raw material applicability. Therefore, it is necessary to develop a method for high-value recovery and utilization of electrolytic aluminum waste slag. Summary of the Invention

[0006] In view of the problems existing in the prior art, the present invention provides a method for high-value recovery and utilization of electrolytic aluminum waste slag. The method of the present invention can achieve the high-value recovery and utilization of F resources, C resources, Al resources, and Li resources in electrolytic aluminum waste slag. In particular, the Li recovery rate can reach more than 90%. The lithium carbonate product has few impurities and high purity. Moreover, the method of the present invention has the advantages of wide raw material adaptability, low requirements for raw materials, no waste water and waste gas emissions, and environmental friendliness.

[0007] To achieve this purpose, the present invention adopts the following technical solutions:

[0008] The purpose of the present invention is to provide a method for the high-value recycling and utilization of electrolytic aluminum waste residue. The method comprises the following steps:

[0009] (1) Refine the electrolytic aluminum waste residue to obtain electrolytic aluminum waste residue powder;

[0010] (2) Mix the roasting agent evenly with the electrolytic aluminum waste residue powder obtained in step (1) to obtain a mixture;

[0011] (3) Carry out low-temperature roasting on the mixture obtained in step (2), and use the generated HF-containing tail gas to prepare fluoride products;

[0012] (4) Refine, water-leach, and carry out solid-liquid separation on the roasted material obtained in step (3) in sequence to obtain a water-leachate and a water-leached residue; wherein, after the water-leached residue is washed, a high-carbon residue is obtained;

[0013] (5) Add an ammonium salt to the water-leachate obtained in step (4), and carry out heating, stirring, dissolution, cooling, aging, crystallization, and solid-liquid separation in sequence to obtain a post-aluminum-precipitation solution and ammonium alum;

[0014] (6) Purify and roast the ammonium alum described in step (5) in sequence to obtain an alumina product;

[0015] (7) Carry out impurity removal and solid-liquid separation on the post-aluminum-precipitation solution obtained in step (5) in sequence, add a lithium-precipitating agent to the obtained lithium-containing solution for lithium precipitation, and carry out solid-liquid separation to obtain a post-lithium-precipitation solution and a lithium carbonate product;

[0016] Among them, there is no sequence priority between step (6) and step (7).

[0017] In the method of the present invention, the acidification low-temperature roasting process is adopted, which can realize the flue gas recovery of F resources and the water-leach recovery of Al and Li resources. The process is simple and reliable, and the recovery rate of valuable resources is high. The control of the roasting process is the key to achieving high yield; adding an ammonium salt to the water-leachate, adopting the ammonium alum aluminum precipitation process to separate Al elements from the water-leachate, with small Li loss and high recovery rate, adopting the ammonium alum recrystallization roasting process to produce high-purity alumina products, with high added value and significant economic benefits. The ammonium alum crystallization purification process is the key process for producing high-purity alumina. Producing high-purity alumina products can greatly improve the product added value, and the tail gas generated by the roasting of ammonium alum can be recycled through absorption treatment, without generating any ammonia-nitrogen waste gas; washing the water-leached residue can make the high-carbon residue completely harmless, with high calorific value, and can be sold as a general energy material.

[0018] The method of the present invention can achieve the high-value recycling and utilization of F resources, C resources, Al resources, and Li resources in electrolytic aluminum waste residues. In particular, the Li recovery rate can reach more than 90%. The lithium carbonate product has few impurities and high purity. Moreover, the method of the present invention has the advantages of wide raw material adaptability, low requirements for raw materials, no wastewater and waste gas emissions, and environmental friendliness.

[0019] As a preferred technical solution of the present invention, the electrolytic aluminum waste residue in step (1) includes a mixture of one or more of waste electrolytes, overhaul slag, aluminum ash, and carbon slag generated from electrolytic aluminum.

[0020] The method of the present invention has the advantages of wide raw material adaptability and low requirements for raw materials.

[0021] Preferably, the refinement in step (1) includes crushing and ball milling carried out in sequence.

[0022] Preferably, the particle size of the powder of the electrolytic aluminum waste residue in step (1) is not greater than 100 mesh.

[0023] As a preferred technical solution of the present invention, the roasting agent in step (2) includes sulfuric acid and / or sodium bisulfate, and the added mass of the roasting agent is 0.8 - 1.5 times the mass of the electrolytic aluminum waste residue powder, such as 0.8 times, 0.9 times, 1.0 times, 1.1 times, 1.2 times, 1.3 times, 1.4 times, or 1.5 times, etc.

[0024] As a preferred technical solution of the present invention, the low-temperature roasting in step (3) is carried out in a tubular furnace.

[0025] Preferably, in step (3), the roasting temperature of the low-temperature roasting is 100 - 500 °C, such as 100 °C, 150 °C, 200 °C, 250 °C, 300 °C, 350 °C, 400 °C, 450 °C, or 500 °C, etc., and the roasting time is 1 - 5 h, such as 1 h, 1.5 h, 2 h, 2.5 h, 3 h, 3.5 h, 4 h, 4.5 h, or 5 h, etc.

[0026] Preferably, in step (3), the HF-containing tail gas is absorbed by an alkaline solution, and the absorption solution is purified and concentrated in sequence to obtain a fluoride salt product; wherein, the alkaline solution includes sodium hydroxide solution and / or sodium carbonate solution.

[0027] As a preferred technical solution of the present invention, the refinement in step (4) includes crushing and ball milling carried out in sequence, and the particle size of the refined powder is not greater than 100 mesh.

[0028] Preferably, the water immersion temperature in step (4) is 20 - 90 °C, such as 20 °C, 30 °C, 40 °C, 50 °C, 60 °C, 70 °C, 80 °C or 90 °C, etc., the leaching time is 0.5 - 5 h, such as 0.5 h, 1 h, 1.5 h, 2 h, 2.5 h, 3 h, 3.5 h, 4 h, 4.5 h or 5 h, etc., and the liquid-solid ratio of the water immersion is (1 - 10) mL:1 g, that is, the ratio of the water volume to the mass of the calcined material is (1 - 10) mL:1 g, such as 1 mL:1 g, 2 mL:1 g, 3 mL:1 g, 4 mL:1 g, 5 mL:1 g, 6 mL:1 g, 7 mL:1 g, 8 mL:1 g, 9 mL:1 g or 10 mL:1 g, etc.

[0029] Preferably, the washing in step (4) includes three times of water washing; the water washing temperature for each time is 20 - 90 °C, such as 20 °C, 30 °C, 40 °C, 50 °C, 60 °C, 70 °C, 80 °C or 90 °C, etc., the water washing time is 0.5 - 5 h, such as 0.5 h, 1 h, 1.5 h, 2 h, 2.5 h, 3 h, 3.5 h, 4 h, 4.5 h or 5 h, etc., and the liquid-solid ratio of the water washing is (1 - 10) mL:1 g, that is, the ratio of the water volume to the mass of the water immersion residue is (1 - 10) mL:1 g, such as 1 mL:1 g, 2 mL:1 g, 3 mL:1 g, 4 mL:1 g, 5 mL:1 g, 6 mL:1 g, 7 mL:1 g, 8 mL:1 g, 9 mL:1 g or 10 mL:1 g, etc.

[0030] As a preferred technical solution of the present invention, the ammonium salt in step (5) includes ammonium sulfate, and the ratio of the molar amount of aluminum element to the molar amount of ammonium sulfate in the water immersion liquid is controlled to be 1:(0.5 - 2), such as 1:0.5, 1:0.7, 1:1, 1:1.1, 1:1.3, 1:1.5, 1:1.7 or 1:2, etc.

[0031] Preferably, the temperature for heating and stirring to dissolve in step (5) is 50 - 90 °C, such as 50 °C, 55 °C, 60 °C, 65 °C, 70 °C, 75 °C, 80 °C, 85 °C or 90 °C, etc.

[0032] Preferably, the temperature for cooling, aging and crystallizing in step (5) is 5 - 25 °C, such as 5 °C, 10 °C, 15 °C, 20 °C or 25 °C, etc., and the aging time is 1 - 8 h, such as 1 h, 2 h, 3 h, 4 h, 5 h, 6 h, 7 h or 8 h, etc.

[0033] As a preferred technical solution of the present invention, the purification in step (6) includes recrystallization of ammonium alum, and the number of recrystallization times is 1 - 3 times.

[0034] Preferably, the calcination temperature in step (6) is 800 - 1100 °C, such as 800 °C, 850 °C, 900 °C, 950 °C, 1000 °C, 1050 °C or 1100 °C, etc., and the calcination time is 1 - 5 h, such as 1 h, 1.5 h, 2 h, 2.5 h, 3 h, 3.5 h, 4 h, 4.5 h or 5 h, etc.

[0035] It should be noted that the purified ammonium alum is placed in a muffle furnace for calcination, and the calcination tail gas is introduced into an absorption device to recover ammonium salts, for example, it is absorbed by dilute sulfuric acid to obtain ammonium sulfate.

[0036] As a preferred technical solution of the present invention, the impurity removal in step (7) includes aluminum removal and calcium removal in sequence;

[0037] The aluminum removal includes: adding lime to the solution after aluminum precipitation to remove aluminum, controlling the aluminum removal temperature at 60 - 90 °C, such as 60 °C, 65 °C, 70 °C, 75 °C, 80 °C, 85 °C or 90 °C, etc., and the reaction time is 1 - 5 h, such as 1 h, 1.5 h, 2 h, 2.5 h, 3 h, 3.5 h, 4 h, 4.5 h or 5 h, etc., and after solid-liquid separation, an aluminum-removed solution and aluminum-removed slag are obtained;

[0038] The calcium removal includes: adding a saturated sodium carbonate solution to the aluminum-removed solution to remove calcium, controlling the calcium removal temperature at 25 - 90 °C, such as 25 °C, 30 °C, 35 °C, 40 °C, 45 °C, 50 °C, 55 °C, 60 °C, 65 °C, 70 °C, 75 °C, 80 °C, 85 °C or 90 °C, etc., and the reaction time is 1 - 5 h, such as 1 h, 1.5 h, 2 h, 2.5 h, 3 h, 3.5 h, 4 h, 4.5 h or 5 h, etc., and after solid-liquid separation, a lithium-containing solution and calcium-removed slag are obtained.

[0039] As a preferred technical solution of the present invention, the lithium precipitation agent in step (7) includes a saturated sodium carbonate solution, and the ratio of the molar amount of lithium element in the lithium-containing solution to the molar amount of carbonate radical in the saturated sodium carbonate solution is controlled at 1:(0.5 - 2), such as 1:0.5, 1:0.7, 1:1, 1:1.1, 1:1.3, 1:1.5, 1:1.7 or 1:2, etc.

[0040] Preferably, the reaction temperature for lithium precipitation in step (7) is 60 - 95 °C, such as 60 °C, 65 °C, 70 °C, 75 °C, 80 °C, 85 °C, 90 °C or 95 °C, etc., and the lithium precipitation time is 1 - 5 h, such as 1 h, 1.5 h, 2 h, 2.5 h, 3 h, 3.5 h, 4 h, 4.5 h or 5 h, etc.

[0041] As a preferred technical solution of the present invention, the solution after lithium precipitation in step (7) is subjected to evaporation concentration and cooling crystallization in sequence to obtain a sulfate product.

[0042] Preferably, the crystallization mother liquor from cooling crystallization is returned to step (4) and used as the water required for water leaching.

[0043] Compared with the prior art solutions, the present invention has at least the following beneficial effects:

[0044] (1) In the method of the present invention, by adopting the acidification low-temperature roasting process, the flue gas recovery of F resources and the water leaching recovery of Al and Li resources can be realized. The process is simple and reliable, and the recovery rate of valuable resources is high. The control of the roasting process is the key to achieving high yield;

[0045] (2) In the method of the present invention, an ammonium salt is added to the water leaching solution, and the ammonium alum precipitation process is used to separate Al elements from the water leaching solution. The loss of Li is small and the recovery rate is high. The ammonium alum recrystallization roasting process is used to produce high-purity alumina products with high added value and significant economic benefits. The ammonium alum crystallization purification process is the key process for producing high-purity alumina products. Producing high-purity alumina products can greatly improve the added value of products. The tail gas generated by the roasting of ammonium alum can be absorbed and treated to realize the closed-loop reuse of ammonia without generating any ammonia-nitrogen waste gas;

[0046] (3) In the method of the present invention, by washing the water leaching residue, the high-carbon residue can be completely rendered harmless, with a high calorific value, and can be sold as a general energy material;

[0047] (4) The method of the present invention can realize the high-value recycling and utilization of F resources, C resources, Al resources, and Li resources in electrolytic aluminum waste residue. In particular, the recovery rate of Li can reach more than 90%, and the impurities in the lithium carbonate product are few and the purity is high;

[0048] (5) The method of the present invention has the advantages of wide raw material adaptability, low requirements for raw materials, no wastewater and waste gas emissions, and environmental friendliness. Description of the Drawings

[0049] Figure 1 is the process flow chart of a method for the high-value recycling and utilization of electrolytic aluminum waste residue in a specific embodiment of the present invention. Detailed Embodiments

[0050] The technical solution of the present invention will be further described below in conjunction with the drawings and through specific embodiments.

[0051] The present invention provides a specific embodiment, that is, a method for the high-value recycling and utilization of electrolytic aluminum waste residue, as Figure 1 shown, the method comprises the following steps:

[0052] (1) Refine the electrolytic aluminum waste residue to obtain electrolytic aluminum waste residue powder;

[0053] (2) Mix concentrated sulfuric acid evenly with the electrolytic aluminum waste residue powder obtained in step (1) to obtain a mixture;

[0054] (3) Calcinate the mixture obtained in step (2) at low temperature, and use the generated HF-containing tail gas to prepare fluorine salt products; specifically, the HF-containing tail gas is absorbed by an alkaline solution, and the absorption liquid is purified and concentrated in sequence to obtain fluorine salt products;

[0055] (4) Refine, leach with water, and filter the calcined material obtained in step (3) in sequence to obtain a water leaching solution and a water leaching residue; among them, after the water leaching residue is washed, a high-carbon residue is obtained;

[0056] (5) Add ammonium sulfate to the water leaching solution obtained in step (4) to precipitate aluminum, specifically, heat, stir and dissolve, cool and age for crystallization, and filter in sequence to obtain the solution after aluminum precipitation and ammonium alum;

[0057] (6) Purify and calcine the ammonium alum described in step (5) in sequence to obtain high-purity alumina products; among them, the calcination tail gas is introduced into an absorption device to recover ammonium sulfate, for example, absorb it with dilute sulfuric acid to recover ammonium sulfate, realizing the recycling of ammonia and sulfur;

[0058] (7) Add a impurity remover to the solution after aluminum precipitation obtained in step (5) for impurity removal and filtration in sequence to obtain an impurity removal residue and a lithium-containing solution, add saturated sodium carbonate solution to the lithium-containing solution to precipitate lithium, and filter to obtain the solution after lithium precipitation and lithium carbonate products; evaporate and concentrate the solution after lithium precipitation in sequence, and cool and crystallize to obtain sodium sulfate products;

[0059] Among them, there is no sequence between step (6) and step (7).

[0060] To better illustrate the present invention and facilitate understanding of the technical solution of the present invention, the typical but non-limiting embodiments of the present invention are as follows:

[0061] Example 1

[0062] This example provides a method for the high-value recycling of electrolytic aluminum waste residue, and the method includes the following steps:

[0063] (1) Take 200 g of electrolytic aluminum overhaul slag, successively crush, ball mill, and pass through a 150-mesh sieve to obtain electrolytic aluminum waste residue powder; among them, calculated by mass percentage, the elemental composition of the electrolytic aluminum overhaul slag is shown in Table 1, and the remaining components are oxygen and trace impurities;

[0064] Table 1

[0065] Element F Al Na Li C Si Fe Content, wt% 35 27 18 1.5 9 1 2

[0066] (2) Slowly add 200 g of concentrated sulfuric acid to the electrolytic aluminum waste residue powder and mix evenly to obtain a mixture;

[0067] (3) Place the mixture in a tube furnace for low-temperature roasting. The roasting temperature is 300 °C, and the roasting time is 2 h. The roasting tail gas (including HF tail gas) is absorbed by a sodium hydroxide solution with a concentration of 2 mol / L for the preparation of sodium fluoride products. After roasting, the sodium fluoride content in the tail gas absorption liquid is measured, and the fluorine recovery rate is 85%;

[0068] (4) The roasted material is successively crushed, ball-milled, passed through a 150-mesh sieve, and added to 800 mL of water for water leaching. The water leaching temperature is 70 °C, and the leaching time is 2 h to obtain a water leaching solution and water leaching residue. Among them, the water leaching residue is washed three times with 800 mL of water each time. The washing temperature is 60 °C, and the single washing time is 1 h. After the water leaching residue is washed three times, the carbon content of the high-carbon residue is measured to be 52.90%, and the fluorine content is 0.05%;

[0069] (5) Add ammonium sulfate solid to the water leaching solution, control the molar ratio of aluminum element in the water leaching solution to ammonium sulfate to be 1:0.6, heat to 80 °C until completely dissolved, then naturally cool down to 20 °C, and age for 5 h. After filtration, ammonium alum crystals and aluminum-depleted liquid are obtained;

[0070] (6) The ammonium alum is subjected to 2 recrystallizations and then roasted. The roasting temperature is 1000 °C, and the roasting time is 2 h to obtain high-purity alumina. The aluminum recovery rate is calculated to be 72%, and the alumina purity is 99.95%. Among them, the roasting tail gas is absorbed by dilute sulfuric acid to recover ammonium sulfate, realizing the recycling of ammonia and sulfur;

[0071] (7) Add lime to the aluminum-depleted liquid to remove aluminum, adjust the pH to 13, react at 80 °C for 2 h, filter to obtain aluminum-removed liquid and aluminum-removed slag. Then add saturated sodium carbonate solution to the aluminum-removed liquid to remove calcium. The molar ratio of calcium ions in the aluminum-removed liquid to sodium carbonate is 1:1.2, react at 30 °C for 1 h, filter to obtain a lithium-containing solution and calcium-removed slag. Add saturated sodium carbonate solution to the lithium-containing solution. The molar ratio of lithium element in the lithium-containing solution to the carbonate in the saturated sodium carbonate solution is 1:0.6, react at 90 °C for 2 h, filter to obtain lithium carbonate products and lithium-depleted liquid. The Li recovery rate is measured to be 91%. The lithium-depleted liquid is successively subjected to evaporation concentration and cooling crystallization to obtain sodium sulfate products;

[0072] Among them, there is no sequential order between step (6) and step (7).

[0073] Example 2

[0074] This example provides a method for the high-value recycling and utilization of electrolytic aluminum waste residue. The method includes the following steps:

[0075] (1) Take 300 g of electrolytic aluminum overhaul slag, successively crush, ball-mill, and pass through a 150-mesh sieve to obtain electrolytic aluminum waste residue powder. Among them, the electrolytic aluminum overhaul slag treated in this example is the same as that in Example 1;

[0076] (2) Slowly add 320 g of concentrated sulfuric acid to the electrolytic aluminum waste residue powder and mix well to obtain a mixture.

[0077] (3) Place the mixture in a tubular furnace for low-temperature roasting. The roasting temperature is 260 °C and the roasting time is 3 h. The roasting tail gas (including HF tail gas) is absorbed by a sodium hydroxide solution with a concentration of 2 mol / L for the preparation of sodium fluoride products. After roasting, measure the sodium fluoride content in the tail gas absorption solution, and the fluorine recovery rate is measured to be 88%.

[0078] (4) The roasted material is successively crushed, ball-milled, passed through a 150-mesh sieve, and added to 1000 mL of water for water leaching. The water leaching temperature is 80 °C and the leaching time is 2 h to obtain a water leaching solution and water leaching residue. Among them, the water leaching residue is washed three times with 1000 mL of water each time. The washing temperature is 60 °C and the single washing time is 1 h. After the water leaching residue is washed three times, the carbon content of the high-carbon residue is measured to be 48.85% and the fluorine content is 0.04%.

[0079] (5) Add ammonium sulfate solid to the water leaching solution, control the molar ratio of aluminum element in the water leaching solution to ammonium sulfate to be 1:0.7, heat to 80 °C to dissolve completely, then cool naturally to 20 °C, age for 5 h, and filter to obtain ammonium alum crystals and the solution after aluminum precipitation.

[0080] (6) The ammonium alum is subjected to 2 times of recrystallization and then roasted. The roasting temperature is 1100 °C and the roasting time is 1 h to obtain high-purity alumina. The aluminum recovery rate is calculated to be 76% and the alumina purity is 99.96%. Among them, the roasting tail gas is absorbed by dilute sulfuric acid to recover ammonium sulfate, realizing the recycling of ammonia and sulfur.

[0081] (7) Add lime to the solution after aluminum precipitation for aluminum removal, adjust the pH = 13, react at 80 °C for 2 h, filter to obtain the solution after aluminum removal and the aluminum removal residue, then add saturated sodium carbonate solution to the solution after aluminum removal for calcium removal. The molar ratio of calcium ion in the solution after aluminum removal to sodium carbonate is 1:1.1, react at 30 °C for 1 h, filter to obtain a lithium-containing solution and the calcium removal residue, add saturated sodium carbonate solution to the lithium-containing solution. The molar ratio of lithium element in the lithium-containing solution to carbonate in the saturated sodium carbonate solution is 1:0.6, react at 90 °C for 2 h, filter to obtain lithium carbonate products and the solution after lithium precipitation. The Li recovery rate is measured to be 92%. The solution after lithium precipitation is successively subjected to evaporation concentration and cooling crystallization to obtain sodium sulfate products.

[0082] Among them, there is no sequence between step (6) and step (7).

[0083] Example 3

[0084] This example provides a method for the high-value recycling of electrolytic aluminum waste residue. The method includes the following steps:

[0085] (1) Take 500 g of the mixture of electrolytic aluminum carbon slag and waste electrolyte, successively crush, ball-mill it, and pass through a 150-mesh sieve to obtain electrolytic aluminum waste residue powder; among them, according to the mass percentage, the elemental composition of the mixture in this example is shown in Table 2, and the remaining components are oxygen and trace impurities;

[0086] Table 2

[0087] Element F Al Na Li C Si Fe Content, wt% 26 18 12 1.0 38 1 1

[0088] (2) Slowly add 450 g of concentrated sulfuric acid to the electrolytic aluminum waste residue powder and mix well to obtain a mixture;

[0089] (3) Place the mixture in a tube furnace for low-temperature roasting. The roasting temperature is 250 °C, and the roasting time is 3 h. The roasting tail gas (containing HF tail gas) is absorbed by a sodium hydroxide solution with a concentration of 2.5 mol / L for the preparation of sodium fluoride products; after roasting, the sodium fluoride content in the tail gas absorption solution is measured, and the fluorine recovery rate is measured to be 88%;

[0090] (4) Crush and ball-mill the roasted material successively, pass through a 150-mesh sieve, add it to 2000 mL of water for water leaching. The water leaching temperature is 75 °C, and the leaching time is 4 h to obtain a water leaching solution and a water leaching residue; among them, the water leaching residue is washed three times, with 2000 mL of washing water each time, the washing temperature is 60 °C, and the single washing time is 0.5 h; after the water leaching residue is washed three times, the carbon content of the high-carbon slag is measured to be 45.00%, and the fluorine content is 0.06%;

[0091] (5) Add ammonium sulfate solid to the water leaching solution, control the molar ratio of aluminum element in the water leaching solution to ammonium sulfate to be 1:0.6, heat to 80 °C until completely dissolved, then cool naturally to 25 °C, age for 4 h, and filter to obtain ammonium alum crystals and the solution after aluminum precipitation;

[0092] (6) After ammonium alum is recrystallized twice, it is roasted. The roasting temperature is 1000 °C, and the roasting time is 2 h to obtain high-purity alumina. The aluminum recovery rate is calculated to be 78%, and the alumina purity is 99.95%; among them, the roasting tail gas is absorbed by dilute sulfuric acid to recover ammonium sulfate, realizing the recycling of ammonia and sulfur;

[0093] (7) Add lime to the solution after aluminum precipitation to remove aluminum. Adjust the pH to 13 and react at 85 °C for 2 h. Filter to obtain the aluminum-removed solution and aluminum-removal slag. Then, add saturated sodium carbonate solution to the aluminum-removed solution to remove calcium. The molar ratio of calcium ions to sodium carbonate in the aluminum-removed solution is 1:1.2. React at 50 °C for 1 h. Filter to obtain the lithium-containing solution and calcium-removal slag. Add saturated sodium carbonate solution to the lithium-containing solution. The molar ratio of lithium element to carbonate in the saturated sodium carbonate solution in the lithium-containing solution is 1:0.7. React at 90 °C for 2 h. Filter to obtain lithium carbonate product and the solution after lithium precipitation. The Li recovery rate is measured to be 90%. The solution after lithium precipitation is successively subjected to evaporation concentration and cooling crystallization to obtain sodium sulfate product;

[0094] Among them, there is no sequence between step (6) and step (7).

[0095] Organize the product information obtained from the above examples and comparative examples and summarize them in Table 3.

[0096] Table 3

[0097]

[0098] The present invention provides a method for the high-value recycling and utilization of electrolytic aluminum waste residue. In the method, an acidification low-temperature roasting process is adopted, which can realize the flue gas recovery of F resources and the water leaching recovery of Al and Li resources. The process is simple and reliable, and the recovery rate of valuable resources is high. The control of the roasting process is the key to achieving high yield; add ammonium salt to the water leaching solution, and adopt the ammonium alum aluminum precipitation process to separate Al element from the water leaching solution. The Li loss is small and the recovery rate is high. Adopt the ammonium alum recrystallization roasting process to produce high-purity alumina products with high added value and significant economic benefits. The ammonium alum crystallization purification process is the key process for producing high-purity alumina products. Producing high-purity alumina products can greatly improve the product added value. The tail gas generated by the ammonium alum roasting can be absorbed and treated to realize the closed-loop reuse of ammonia without generating any ammonia-nitrogen waste gas; wash the water leaching residue, which can make the high-carbon slag completely harmless, with high calorific value, and can be sold as general energy materials.

[0099] The method of the present invention can realize the high-value recycling and utilization of F resources, C resources, Al resources, and Li resources in electrolytic aluminum waste residue. In particular, the Li recovery rate can reach more than 90%. The lithium carbonate product has few impurities and high purity; moreover, the method of the present invention has the advantages of wide raw material adaptability, low requirements for raw materials, no wastewater and waste gas emissions, and environmental friendliness.

[0100] The present invention uses the above embodiments to illustrate the detailed structural features of the present invention. However, the present invention is not limited to the above detailed structural features, that is, it does not mean that the present invention must rely on the above detailed structural features to be implemented. Those skilled in the art should understand that any improvement to the present invention, the equivalent replacement of the components selected for the present invention, the addition of auxiliary components, and the selection of specific methods, etc., all fall within the protection scope and the disclosure scope of the present invention.

[0101] The preferred embodiments of the present invention have been described in detail above. However, the present invention is not limited to the specific details in the above embodiments. Within the scope of the technical concept of the present invention, various simple modifications can be made to the technical solutions of the present invention, and these simple modifications all fall within the protection scope of the present invention.

[0102] In addition, it should be noted that, among the various specific technical features described in the above specific embodiments, without conflict, they can be combined in any appropriate manner. To avoid unnecessary repetition, the present invention will not separately describe various possible combination methods.

[0103] Furthermore, any combination can be made between various different embodiments of the present invention, as long as it does not violate the idea of the present invention, it should also be regarded as the content disclosed by the present invention.

Claims

1. A method for high-value recycling of electrolytic aluminum waste residue, characterized in that, The method includes the following steps: (1) Refine the electrolytic aluminum waste residue to obtain electrolytic aluminum waste residue powder; (2) Mix the roasting agent evenly with the electrolytic aluminum waste residue powder obtained in step (1) to obtain a mixture; (3) Roast the mixture obtained in step (2) at a low temperature, and use the generated HF-containing tail gas to prepare fluorine salt products; (4) Refine, leach with water, and separate solid from liquid the roasted material obtained in step (3) in sequence to obtain a water leaching solution and a water leaching residue; wherein, after the water leaching residue is washed, a high-carbon residue is obtained; (5) Add an ammonium salt to the water leaching solution obtained in step (4), and carry out heating, stirring and dissolving, cooling and aging crystallization, and solid-liquid separation in sequence to obtain a solution after aluminum precipitation and ammonium alum; (6) Purify and roast the ammonium alum described in step (5) in sequence to obtain alumina products; (7) Carry out impurity removal and solid-liquid separation on the solution after aluminum precipitation obtained in step (5) in sequence, add a lithium precipitation agent to the obtained lithium-containing solution for lithium precipitation, and obtain a solution after lithium precipitation and lithium carbonate products through solid-liquid separation; Among them, there is no sequence between step (6) and step (7).

2. The method according to claim 1, characterized in that, The electrolytic aluminum waste residue described in step (1) includes a mixture of one or more of waste electrolytes, overhaul slag, aluminum ash, and carbon slag generated during electrolytic aluminum production; Preferably, the refinement described in step (1) includes crushing and ball milling carried out in sequence; Preferably, the particle size of the powder of the electrolytic aluminum waste residue described in step (1) is not greater than 100 mesh.

3. The method according to claim 1 or 2, characterized in that, The roasting agent described in step (2) includes sulfuric acid and / or sodium bisulfate, and the added mass of the roasting agent is 0.8 - 1.5 times the mass of the electrolytic aluminum waste residue powder.

4. The method according to any one of claims 1 to 3, characterized in that, The roasting temperature of the low-temperature roasting described in step (3) is 100 - 500 °C, and the roasting time is 1 - 5 h; Preferably, in step (3), the HF-containing tail gas is absorbed by an alkaline solution, and the absorption solution is purified and concentrated in sequence to obtain fluorine salt products; wherein, the alkaline solution includes sodium hydroxide solution and / or sodium carbonate solution.

5. The method according to any one of claims 1-4, characterized in that, The refinement described in step (4) includes crushing and ball milling carried out in sequence, and the particle size of the refined powder is not greater than 100 mesh; Preferably, the water leaching temperature of the water leaching described in step (4) is 20 - 90 °C, the leaching time is 0.5 - 5 h, and the liquid-solid ratio of the water leaching solution is (1 - 10) mL:1 g; Preferably, the washing described in step (4) includes three times of water washing; the water washing temperature for each water washing is 20 - 90 °C, the water washing time is 0.5 - 5 h, and the liquid-solid ratio of the water washing solution is (1 - 10) mL:1 g.

6. The method according to any one of claims 1 to 5, characterized in that, The ammonium salt described in step (5) includes ammonium sulfate, and the molar ratio of aluminum element in the water leaching solution to ammonium sulfate is controlled to be 1:(0.5 - 2); Preferably, the temperature of the heating, stirring and dissolving described in step (5) is 50 - 90 °C; Preferably, the temperature of the cooling and aging crystallization described in step (5) is 5 - 25 °C, and the aging time is 1 - 8 h.

7. The method according to any one of claims 1-6, characterized in that, The purification described in step (6) includes recrystallization of ammonium alum, and the number of recrystallization times is 1 - 3 times; Preferably, the roasting temperature of the roasting described in step (6) is 800 - 1100 °C, and the roasting time is 1 - 5 h.

8. The method according to any one of claims 1-7, characterized in that The impurity removal described in step (7) includes aluminum removal and calcium removal carried out in sequence; The aluminum removal includes: adding lime to the obtained solution after aluminum precipitation for aluminum removal, controlling the aluminum removal temperature at 60 - 90 °C, the reaction time at 1 - 5 h, and obtaining an aluminum-removed solution and aluminum-removed slag through solid-liquid separation; The calcium removal includes: adding a saturated sodium carbonate solution to the aluminum-removed solution for calcium removal, controlling the calcium removal temperature at 25 - 90 °C, the reaction time at 1 - 5 h, and obtaining a lithium-containing solution and calcium-removed slag through solid-liquid separation.

9. The method according to any one of claims 1 to 8, characterized in that, The lithium precipitation agent in step (7) includes a saturated sodium carbonate solution, and the molar ratio of lithium element in the lithium-containing solution to carbonate radical in the saturated sodium carbonate solution is controlled to be 1:(0.5 - 2); Preferably, the reaction temperature for lithium precipitation in step (7) is 60 - 95 °C, and the lithium precipitation time is 1 - 5 h.

10. The method according to any one of claims 1-9, characterized in that, The solution after lithium precipitation in step (7) is successively subjected to evaporation concentration and cooling crystallization to obtain a sulfate product; Preferably, the crystallization mother liquor from cooling crystallization is returned to step (4) and used as the water required for water leaching.

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