Cooperative treatment method for lithium-containing waste

Abstract

The method comprises the following steps that the silicon-aluminum-rich lithium-containing waste and an acid solution are mixed, solid-liquid separation is conducted after reaction, a leaching solution A and silicon-containing leaching residues are obtained, and the molar weight of hydrogen ions in the acid solution is 1.5 times or above of the theoretical consumption of lithium and aluminum in the silicon-aluminum-rich lithium-containing waste; and mixing the obtained leachate A and the fluorine-aluminum-rich lithium-containing waste, regulating the pH value to be 1-3.5, reacting at the temperature of 80 DEG C or above, and then carrying out solid-liquid separation to obtain lithium-containing leachate and aluminum-containing leaching residues. According to the treatment method, synergistic separation and recovery of the lithium-containing waste can be achieved, valuable lithium in the lithium-containing waste can be effectively leached out, meanwhile, leaching of impurities difficult to separate is inhibited, lithium entrainment loss caused by subsequent impurity removal is avoided, and battery-grade lithium carbonate can be efficiently prepared through simple impurity removal of the leachate.

Description

TECHNICAL FIELD

[0001] The application belongs to the field of lithium-containing waste treatment, and more particularly relates to a method for synergistically treating lithium-containing waste. BACKGROUND

[0002] Lithium batteries are widely used in energy storage base stations, new energy vehicles, electronic products, aerospace and military communication fields. In the context of global energy transformation, the demand for lithium resources as a key raw material is growing rapidly.

[0003] At present, lithium-containing compounds used in lithium batteries in China mainly come from two paths of salt lake lithium extraction and ore lithium extraction. However, the lithium compounds prepared by salt lake lithium extraction have many impurity ions, and the purity is limited, which makes it difficult to be used in high-end material field. The preparation of lithium compounds from ores is limited by resource exploitation, cost and regional characteristics, and cannot be effectively released at present. Therefore, developing new lithium extraction paths and reasonably developing and recycling existing lithium resources to reduce the exploitation of natural lithium-containing resources is an important issue that the lithium industry will face in the long term in the future.

[0004] Waste residues produced from the extraction of lithium products from lithium-aluminum-silicate minerals or various types of lithium-containing waste generated in the process of glass, ceramic, aluminum material production. Due to the different forms, contents of lithium and the forms, contents of non-lithium elements, and most of the lithium-containing waste does not have recycling value due to the limitation of process cost, most of such lithium-containing materials are in the form of waste residues for low utilization non-extraction treatment (such as used in the construction industry) or stacking, not only causing waste of valuable resources such as lithium, but also causing great pressure on environmental protection, and such waste residues are generally difficult to be fixed by vegetation coverage, so they are vulnerable to extreme weather such as floods, heavy rains, etc., which can easily cause secondary disasters such as mudslides and landslides. Therefore, how to solve the resource waste caused by the low-value treatment of various different lithium-containing waste or the pollution or potential disaster hidden danger caused by direct stacking to the environment, and at the same time alleviate the pressure of lithium resources brought by the development of new energy industry, is an urgent problem to be solved at present.

[0005] For lithium-containing waste residues produced from the extraction of lithium products from lithium-aluminum-silicate minerals or lithium-containing waste generated in the process of glass, ceramic, aluminum material production, based on the main element types, they can be divided into silicon-aluminum-rich lithium-containing waste and fluorine-aluminum-rich lithium-containing waste.

[0006] For the separation and recovery of valuable elements such as lithium in silicon-aluminum-rich lithium-containing waste or fluorine-aluminum-rich lithium-containing waste, some existing technologies are involved.

[0007] For silicon-aluminum-rich lithium-containing waste: CN202211111080.9 discloses a comprehensive recycling method for extracting silicon and aluminum from lithium slag. The lithium slag is the lithium-containing leaching residue after lithium extraction from lithium-containing ores. The main extraction process is as follows: the lithium-containing slag slurry is treated with CO2 at normal temperature and pressure to remove alkali and obtain alkali-removed lithium slag. The alkali-removed lithium slag is dissolved in dilute acid to obtain a fluorine-lithium enriched solution and a dilute acid leaching residue. The fluorine-lithium enriched solution is treated with sodium carbonate in three steps to remove impurities and precipitate lithium to obtain lithium carbonate. The filtrate is added with calcium chloride to obtain calcium fluoride. The dilute acid leaching residue is dissolved with concentrated sulfuric acid with a mass fraction of 98% to obtain a pure aluminum-containing solution and an acid leaching residue. The aluminum-containing solution is adjusted to a certain pH by adding alkali to further precipitate aluminum to obtain high-quality aluminum hydroxide product. The filtrate after aluminum precipitation is evaporated to obtain a product containing rubidium sulfate. The acid leaching residue is further treated with alkali to obtain a silicate solution, which is reacted with CO2 to produce white carbon black product. For the main elements of lithium, aluminum, silicon, and fluorine in waste materials, lithium and fluorine are leached by dilute acid, aluminum is leached by concentrated acid, and silicon is leached by alkali. The leaching operation is complex, and after leaching, each leaching solution is treated separately. The lithium-containing leaching solution needs to be adjusted to a certain pH three times to separate and remove impurities. The overall process has a long and complex impurity removal process, and the purity of the obtained lithium carbonate is only above 90%. Further treatment is required for it to be used in lithium battery material production. In addition, it requires pressurized treatment and the use of strong acids such as concentrated sulfuric acid, which has high requirements for equipment.

[0008] CN202410536576.3 discloses a method for extracting lithium from lithium-containing silicon-aluminum glass. To address the problems of high aluminum impurity content in the leaching solution and poor filtration caused by aluminum hydroxide precipitation, the patent proposes mixing lithium-containing silicon-aluminum glass, soluble M salt, and sulfuric acid solution, pressurized leaching, and separating to obtain filtrate and filter residue. The filter residue contains aluminum alum precipitate, and the aluminum alum precipitate contains M, Al, and sulfate. The pressure of pressurized leaching is 2 MPa to 5 MPa, the temperature is 170°C to 270°C, and the time is 0.5 hours to 4 hours. Although this method can achieve efficient leaching of lithium, it requires high temperature and high pressure for leaching, and additional M salt reagent is added. The molar ratio of M element to Al element in the lithium-containing silicon-aluminum glass is (0.3-1):1. M is at least one of lithium, sodium, potassium, rubidium, cesium, silver, thallium, and ammonium. The cost of reagent is high, and the economic value of aluminum alum residue, a byproduct, is low.

[0009] CN202410368977.2 discloses a method for extracting lithium from lithium-containing glass powder. To address the defects of lithium recovery rate and purity in traditional lithium extraction methods, the patent proposes mixing lithium-containing glass powder with alkaline substances and high-temperature roasting. The mass ratio of lithium-containing glass powder to alkaline substances is 5:(3-5), the high-temperature roasting temperature is 600-700°C, the roasting time is 1.5-2 hours, and the roasting material is then acid leached to extract lithium. The leaching rate is between 52.8% and 99.5%. This patent requires the addition of a large amount of alkaline substances and high-temperature roasting to achieve an increase in lithium leaching rate. This not only consumes a lot of energy and has high reagent costs, but also fails to solve the problem of removing aluminum impurities from lithium-containing leaching solution.

[0010] For fluorine-rich aluminum-containing lithium-containing waste materials: CN202210969842.2 discloses a method for preparing lithium carbonate by recycling fluorine-containing lithium tailings. The fluorine-containing lithium tailings are mixed with sulfuric acid in an autoclave and heated to react, and the hydrogen fluoride aqueous solution is collected by condensation. The lithium-containing and aluminum-containing leaching solution is adjusted in pH and then barium hydrogen carbonate is added to react, obtaining a mixed precipitate of barium sulfate and aluminum hydroxide and a lithium-containing solution. The mixed precipitate is dissolved with alkali to obtain barium sulfate and an aluminum-containing solution. The lithium-containing solution is reacted with sodium hydroxide or pyrolyzed to obtain lithium carbonate and soda solution. The obtained hydrogen fluoride and insoluble matter, aluminum-containing solution, soda solution, and a certain amount of aluminum hydroxide slurry are mixed and reacted in an autoclave to obtain cryolite product. The whole process of the process method produces harmful other hydrogen fluoride gas, which has high requirements for the sealing and safety of the equipment; the cost of barium salt as an aluminum and lithium separation precipitant is high; the aluminum in the leaching solution is first removed in the form of aluminum hydroxide, and then the alkali solution is obtained to obtain an aluminum-containing solution, which is complicated; finally, the fluorine and aluminum products are added with aluminum hydroxide to synthesize cryolite, which requires a high-pressure reaction kettle, and the synthesis equipment requirements and investment cost are high.

[0011] CN201910609881.X discloses that the lithium-rich aluminum electrolyte is subjected to heat treatment, and then a water-soluble inorganic salt is used as a leaching agent to leach the heat-treated product in water. After the leaching solution is adjusted in pH to remove aluminum, lithium carbonate is obtained by adding a carbonate salt to precipitate lithium. The process requires high-temperature treatment at 750-1100°C for 1.5-3.5h, which has high energy consumption and long time; the lithium-containing leaching solution cannot avoid the loss of entrained lithium by adjusting the pH to remove aluminum; the leaching rate and recovery rate of lithium are not disclosed.

[0012] CN202410630489.4 discloses a method for treating and utilizing aluminum electrolysis cell overhaul slag. The overhaul slag raw material is crushed and then mixed uniformly with a calcium-containing substance, and then sent into a rotary kiln for calcination at 600-900°C for 1-4h. The cooled clinker is reacted with an acidic filtrate to obtain a filtrate and a filter residue. The filter residue is subjected to high-temperature leaching with a strong acid, and then filtered. The acidic filtrate is returned to leach the clinker, and the solid is washed with water to produce calcium fluoride product. Sodium hydroxide and aluminum oxide are added to the filtrate obtained by leaching the clinker to remove aluminum in the form of sodium fluoroaluminate. After removing aluminum, the solution is reacted with sodium hydroxide and sodium carbonate in a removal impurity kettle, filtered, and then the filtrate is concentrated and filtered to obtain sodium sulfate solid. The filtrate enters a lithium precipitation process. The method requires 60-90% of the calcium-containing substance based on the mass of the raw material for high-temperature calcination, and the calcined material is subjected to two-step acid leaching and multi-step separation and impurity removal, which has a long process flow and complex operation. The aluminum ion content in the solution after removing aluminum is still about 1g / L, and the lithium recovery rate is about 83%, which needs to be further improved.

[0013] The above existing technologies are all aimed at treating single type of lithium-containing waste materials. There are occasional references to the common treatment of waste materials, such as: CN202410164609.6 proposes a method for the co-processing and utilization of overhaul slag, fly ash, and desulfurized gypsum. The desulfurized gypsum is calcined with calcium carbonate at 1000-1200℃ to obtain calcium oxide and SO2. The SO2 is then oxidized at high temperature and reacted with water and fly ash via pressure cooking to obtain aluminum sulfate solution and silicon dioxide. The overhaul slag is calcined with peroxide at high temperature, followed by water leaching to separate sodium fluoride. The water-leached slag, aluminum sulfate solution, and calcium-containing substances are then calcined again and water-leached to obtain alumina-containing slag and a lithium-containing solution. The alumina-containing slag is then calcined with calcium oxide at high temperature to obtain calcium aluminate. This method is based on the co-processing of desulfurized gypsum and fly ash to prepare aluminum sulfate and calcium oxide. Aluminum sulfate reacts with overhaul slag to extract fluorine and lithium. The calcium oxide obtained from the desulfurized gypsum calcination reacts with the fluorine and aluminum elements in the overhaul slag to produce calcium aluminate. The process employs a combination of pyrometallurgical and hydrometallurgical methods to treat various waste materials, but the materials require multiple high-temperature roasting treatments, resulting in high overall energy consumption; it involves multiple liquid-solid separation operations, leading to low process efficiency; and it involves high-temperature treatment, dust removal, and harmless treatment of highly acidic and harmful gases, requiring high-end equipment and significant investment.

[0014] CN202210382816.X discloses a system process for high-value utilization of lithium-containing waste. This process involves mixing various lithium-containing wastes, including lithium-containing glass, various lithium-containing molecular sieves, and other lithium-containing silicates, aluminosilicates, or phosphoaluminates, with one or more auxiliary materials 1 (such as lepidolite, spodumene, and lithium clay) and one or more auxiliary materials 2 (such as sodium sulfate, potassium sulfate, calcium sulfate, calcium carbonate, calcium hydroxide, calcium oxide, ferric sulfate, aluminum sulfate, ammonium sulfate, and sulfuric acid). The lithium-containing waste : auxiliary material 1 : auxiliary material 2 = 0.5%~80% : 0%~70% : 15%~60%. After high-temperature roasting at 800~1300℃, the mixture is leached in water. The leachate is then pH-adjusted, purified by adding soda ash, and concentrated to precipitate lithium. Although this patent can process multiple lithium-containing wastes simultaneously, it requires a variety of auxiliary materials in large quantities, resulting in high costs; high-temperature roasting consumes a lot of energy and is not environmentally friendly; the lithium-containing solution does not effectively solve the problem of removing aluminum impurities; in addition, it only focuses on the preparation of lithium salt products and does not reflect the recycling of other valuable metals, resulting in low overall process value.

[0015] Acid treatment of lithium-containing silicon-aluminum waste can efficiently leach valuable elements such as lithium and aluminum. However, excessively high aluminum concentrations in the leachate make subsequent lithium-aluminum separation difficult, with a large amount of lithium being lost in the aluminum hydroxide precipitate, failing to achieve a simple and effective lithium-aluminum separation. Direct acid leaching of lithium-containing fluorine-aluminum waste can dissolve some lithium salts, such as LiF, but insoluble lithium salts like Na₂LiAlF₆ and NaLi₂AlF₆ are poorly soluble in acid, resulting in a low final lithium leaching rate. Separate wet leaching treatments of lithium-containing silicon-aluminum and lithium-containing fluorine-aluminum waste each have their drawbacks. Existing co-processing methods for lithium-containing waste involve varying degrees of high-temperature roasting and the use of large amounts of additives, failing to achieve low-consumption, simple, and efficient recycling and regeneration of lithium-containing waste.

[0016] Therefore, it is necessary to improve the treatment methods for lithium-containing silicon-rich aluminum and lithium-containing fluorine-rich aluminum to achieve simple, efficient, low-cost, and high-value comprehensive recycling of lithium-containing waste. Summary of the Invention

[0017] The purpose of this invention is to provide a method for the co-processing of lithium-containing waste, so as to solve at least one technical problem existing in the above-mentioned background art.

[0018] This invention provides a method for the co-processing of lithium-containing waste, comprising the following steps: (1) Leaching silicon-rich aluminum lithium-containing waste with acidic solution to obtain leachate A and silicon-containing leaching residue, wherein the molar amount of hydrogen ions in the acidic solution is more than 1.5 times the theoretical consumption of lithium and aluminum in the silicon-rich aluminum lithium-containing waste; The theoretical consumption is determined based on the theoretical amount required for the following reaction: Li+H + →Li + Al+3H + →Al 3+ (2) Mix the leachate A obtained in step (1) with the fluorine-rich aluminum lithium-containing waste, adjust and control the pH of the slurry to 1~3.5, react at above 80℃ and then separate the solid and liquid to obtain lithium-containing leachate and aluminum-containing leachate residue.

[0019] Furthermore, in step (1), the lithium-rich silicon aluminum waste includes one or more combinations of lithium-containing aluminosilicate tailings, fly ash, glass, and ceramics.

[0020] Furthermore, in step (1), the acidic solution contains at least one of hydrochloric acid, sulfuric acid, and nitric acid, and the amount used is 1.5 to 4 times the theoretical consumption of lithium and aluminum in the silicon-aluminum lithium-containing waste, preferably 1.5 to 2.5 times, and more preferably 1.75 to 2 times.

[0021] Furthermore, in step (1), the leaching temperature is 50~99℃, preferably 60~95℃, and even more preferably 70~95℃.

[0022] And / or, the leaching time is 0.5 to 3 hours, preferably 1 to 2.5 hours, more preferably 1.5 to 2 hours.

[0023] And / or, the solid-liquid mass ratio of the leaching is 1:4 to 10, preferably 1:4 to 9, more preferably 1:5 to 8.

[0024] In step (1), the lithium ion concentration is 1.0 g / L to 10.0 g / L, preferably 2.0 to 7.5 g / L, and more preferably 2.5 to 6.0 g / L.

[0025] The aluminum ion concentration is 6.0~40.0 g / L, preferably 10.0~30.0 g / L, and more preferably 15.0 g / L~25.0 g / L.

[0026] In step (1), the lithium content in the silicon-rich aluminum lithium-containing waste is above 0.5%, for example above 1%, above 2%, above 5%, above 7.5%, etc.

[0027] The silicon-rich aluminum-containing lithium waste contains aluminum and silicon with a content of 5% or more, for example, 7% or more, 9% or more, 15% or more, 20% or more, etc.

[0028] Furthermore, in step (1), the silicon-rich aluminum lithium-containing waste is powder; preferably, it is powder passing through 50-200 mesh; more preferably, it is powder passing through 80-150 mesh; and even more preferably, it is powder passing through 100 mesh.

[0029] Furthermore, in step (2), the fluorine-rich aluminum lithium-containing waste contains one or more of the following: lithium aluminum electrolyte waste, electrolytic aluminum waste, or aluminum smelting slag.

[0030] Furthermore, the lithium content in the fluorine-rich aluminum lithium-containing waste is above 0.5%, for example above 1%, above 2%, above 5%, above 7.5%, etc.

[0031] The aluminum content in the fluorine-rich aluminum lithium waste is above 5%, for example above 7%, above 12%, above 15%, above 20%, above 25%, etc.

[0032] The fluorine-rich aluminum lithium waste contains more than 5% fluorine, for example, more than 7.5%, more than 10%, more than 15%, etc.

[0033] Furthermore, in step (2), the solid-liquid mass ratio of the reaction is 1:3~15, preferably 1:6~12, and more preferably 1:8~11.

[0034] Furthermore, calcium-containing substances are added in step (2).

[0035] Furthermore, in step (2), the amount of calcium-containing material added is 5-50% of the total molar amount of lithium and aluminum in the fluorine-rich aluminum lithium-containing waste, preferably 10-45%, and more preferably 20-35%.

[0036] And / or, the calcium-containing substance is one or more of calcium chloride, calcium sulfate, calcium carbonate, calcium hydroxide, and calcium oxide, preferably calcium chloride.

[0037] Furthermore, in step (2), the pH is controlled to be 1.5~3.5, preferably 2~3.

[0038] Furthermore, in step (2), the reaction temperature is 85~99℃, more preferably 90~95℃.

[0039] And / or, the reaction time is 0.5 to 6 hours, preferably 3 to 5 hours, and more preferably 4 to 5 hours.

[0040] Furthermore, it also includes step (3), which removes impurities and precipitates lithium in the lithium-rich leachate obtained in step (2) to obtain lithium salt.

[0041] Further, in step (3), the impurity removal operation is to adjust the pH of the lithium-rich leachate obtained in step (2) to remove aluminum, fluorine, and magnesium in sequence, and then add carbonate or oxalate to remove calcium; preferably, the pH of the lithium-rich leachate obtained in step (2) is adjusted to 7±0.5 to remove aluminum and fluorine, then adjusted to 11±0.5 to remove magnesium, and then 1 to 1.2 times the theoretical amount of calcium ions of carbonate or oxalate is added to remove calcium.

[0042] In silicon-rich aluminum-containing lithium waste, aluminum, silicon, and lithium exist in a variety of mixed phases, such as complex silicates, aluminates, aluminosilicates, and oxides. In fluorine-rich aluminum-containing lithium waste, fluorine, aluminum, and lithium exist in a variety of mixed phases, such as various fluoroaluminates, fluorides, and oxides. Research has found that after acid leaching of silicon-rich aluminum-containing lithium waste, the resulting leachate reacts synergistically with the fluorine-rich aluminum-containing lithium waste. The active components in the leachate effectively leach lithium from the fluorine-rich aluminum-containing lithium waste. At the same time, impurities such as aluminum, calcium, magnesium, and fluorine in the leachate obtained in step (1) are removed in the form of multiphase mixed precipitation. The leaching of fluorine and aluminum compounds in the fluorine-rich aluminum waste can be inhibited, allowing them to remain in the slag. This achieves selective leaching of lithium from the fluorine-rich aluminum-containing lithium waste, resulting in a low-aluminum, fluorine-rich lithium leachate, which solves the shortcomings of the existing technology in recycling single-category lithium waste. This invention addresses, to varying degrees, the challenges of lithium-aluminum separation in single wet leaching processes and the low lithium leaching rate of direct acid leaching of lithium-containing fluorine-rich aluminum waste. It also avoids the use of high-pollution, high-energy-consumption, and high-cost equipment / conditions such as high-temperature roasting and high-temperature, high-pressure reactions, effectively reducing the cost of auxiliary material addition and purification. During the lithium extraction process, major elements such as aluminum, silicon, calcium, and fluorine can be recycled and reused. This invention achieves efficient and high-value-added separation and recycling of lithium-containing waste through a pure wet process and simple operation, thereby effectively alleviating the waste of resources and environmental pollution caused by waste dumping, and realizing low-cost, high-value-added resource recycling.

[0043] The present invention has at least the following advantages over the prior art: (1) The pure wet process is adopted, and the leaching conditions are easy to achieve, thus effectively avoiding the high energy consumption and high pollution problems caused by high-temperature fire roasting. At the same time, it reduces the safety hazards in the high-temperature and high-pressure leaching process and the demand for high equipment and operation requirements.

[0044] (2) It achieves the synergistic recycling of waste materials, with a short and easy-to-operate process. Through the combination of staged leaching, the effective components in the two types of waste materials are fully utilized, achieving efficient extraction of lithium while promoting the precipitation and separation of non-lithium elements. This process significantly reduces the amount of leaching reagents and industrial additives used, effectively reducing production costs and solving the environmental pollution problem caused by waste accumulation.

[0045] (3) It achieves efficient separation of lithium and aluminum in hydrometallurgical process, avoids the loss of lithium and filtration difficulties caused by the formation of a large amount of aluminum hydroxide colloid with adsorption capacity during conventional pH adjustment precipitation aluminum removal process, and the high-purity lithium-containing solution can be obtained by simple impurity removal of the synergistic leachate, and high-purity lithium salt product can be obtained by lithium precipitation.

[0046] (4) The lithium carbonate prepared by this process can be used as a raw material for the production of lithium-ion battery materials. The by-products such as insoluble silicon slag can be recycled for the preparation of cement, ceramics, glass and other products. The aluminum slag can be used as the main raw material for electrolytic aluminum and returned to the electrolytic cell for continued use. This process fully realizes the recycling of valuable elements in silicon-rich aluminum lithium-containing waste and fluorine-rich aluminum lithium-containing waste slag, and effectively improves the recycling value of process waste. Attached Figure Description

[0047] The accompanying drawings are provided to further illustrate the invention and form part of the specification. They are used in conjunction with embodiments of the invention to explain the invention and do not constitute a limitation thereof. In the drawings: Figure 1 This is an XRD pattern of lithium-containing silicon-aluminum waste from an embodiment of the present invention.

[0048] Figure 2 This is an XRD pattern of fluorine-rich aluminum lithium-containing waste from an embodiment of the present invention.

[0049] Figure 3 This is the XRD pattern of silicon slag byproduct 1 in Example 3 of the present invention.

[0050] Figure 4 This is the XRD pattern of aluminum slag byproduct 2 in Example 3 of the present invention.

[0051] Figure 5 The image shows the XRD pattern of the leaching residue from Comparative Example 1.

[0052] Figure 6 This is the XRD pattern of aluminum slag byproduct 2 in Example 7 of the present invention.

[0053] Figure 7 The image shows the XRD pattern of aluminum slag byproduct 2 in Comparative Example 2.

[0054] Figure 8 The image shows the XRD pattern of aluminum slag byproduct 2 from Comparative Example 4.

[0055] Figure 9 The image shows the XRD pattern of aluminum slag byproduct 2 from Comparative Example 5. Detailed Implementation

[0056] To facilitate understanding of the present invention, it will be described more fully and in detail below with reference to the accompanying drawings and specific embodiments. However, the scope of protection of the present invention is not limited to the specific embodiments described below. Unless otherwise defined, all technical terms used herein have the same meaning as commonly understood by those skilled in the art. The technical terms used herein are for the purpose of describing specific embodiments only and are not intended to limit the scope of protection of the present invention.

[0057] The silicon-rich aluminum lithium-containing waste used in the following examples is lithium-containing glass waste, such as... Figure 1 As shown, silicon and aluminum mainly exist in the form of quartz and corundum, and their main components (wt.%) are: 1.994% Li, 10.97% Al, 10.878% Si, 1.65% Na, 1.17% Ca, and 1.25% Mg.

[0058] Fluorine-rich aluminum lithium-containing waste is waste aluminum electrolyte, such as... Figure 2 As shown, aluminum, fluorine and lithium mainly exist in the form of cryolite, corundum, fluorite, Na2LiAlF6, etc., and their main components (wt.%) are: 1.575%Li, 14.344%Al, 9.556%F, 16.995%Na, 4.307%Ca, and 1.219%Mg.

[0059] Example 1 Lithium-containing silicon-aluminum waste was mixed with hydrochloric acid solution at a solid-liquid mass ratio of 1:4, with the amount of hydrochloric acid being 1.6 times the theoretical amount. The mixture was stirred at 85°C for 2 hours, followed by solid-liquid separation to obtain filtrate A and leaching residue. The leaching residue was washed and dried to obtain silicon slag byproduct 1. In this embodiment, the lithium leaching rate of the lithium-containing silicon-aluminum waste was 99.29%, and the aluminum leaching rate was 99.45%. The lithium-aluminum mass ratio of filtrate A was 0.18. Fluorine-containing aluminum-aluminum waste and filtrate A were mixed at a solid-liquid mass ratio of 1:7. Calcium chloride was added, with the amount of calcium added being 0.05 times the total molar amount of lithium and aluminum in the fluorine-containing aluminum-aluminum waste. After stirring evenly, the pH of the system was adjusted to 2.6, and the mixture was reacted at 90°C for 4 hours. After filtration, the filter cake was washed and dried to obtain aluminum slag byproduct 2. Filtrate B was the lithium-rich leaching solution. In this embodiment, the lithium co-leaching rate (based on the total lithium of the two wastes, the same below) is 92.41%, the co-aluminum fixation rate is 99.36% (based on the total aluminum of the two wastes, the same below), the fluorine fixation rate is 96.20%, the lithium-aluminum mass ratio of filtrate B is 19.22, and the lithium-fluorine mass ratio is 11.55.

[0060] Example 2 Lithium-containing silicon-aluminum waste was mixed with sulfuric acid solution at a solid-liquid mass ratio of 1:5, with the amount of sulfuric acid being twice the theoretical amount. The mixture was stirred at 95°C for 1.5 hours, followed by solid-liquid separation to obtain filtrate A and leaching residue. The leaching residue was washed and dried to obtain silicon slag byproduct 1. In this embodiment, the lithium leaching rate of the lithium-containing silicon-aluminum waste was 99.54%, and the aluminum leaching rate was 99.03%. The lithium-aluminum mass ratio of filtrate A was 0.18. Fluorine-containing aluminum-aluminum waste and filtrate A were mixed at a solid-liquid mass ratio of 1:7.5, and calcium chloride was added. The amount of calcium added was 0.3 times the total molar amount of lithium and aluminum in the fluorine-containing aluminum-aluminum waste. After stirring evenly, the pH of the system was adjusted to 2.3, and the mixture was reacted at 90°C for 4 hours. After filtration, the filter cake was washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B was the lithium-rich leaching solution filtrate. In this embodiment, the lithium co-leaching rate is 92.36%, the co-aluminum fixation rate is 99.28%, and the fluorine fixation rate is 98.88%. The lithium-aluminum mass ratio of filtrate B is 14.64, and the lithium-fluorine mass ratio is 106.90. The pH of filtrate B is adjusted to 7, and after sufficient reaction, solid-liquid separation is performed. The pH of the filtrate is then adjusted to 11, and after sufficient reaction, solid-liquid separation is performed again. Sodium oxalate with a calcium molar ratio of 1.1 times is added to the filtrate, and the purified liquid is obtained. After lithium precipitation, washing, and drying, battery-grade lithium carbonate product is obtained. In this embodiment, the lithium recovery rate of the leachate is 95.78%, and the composition of the battery-grade lithium carbonate product is shown in Table 1.

[0061] Table 1 Example 3 Lithium-containing silicon-aluminum waste was mixed with hydrochloric acid solution at a solid-liquid mass ratio of 1:5, with the amount of hydrochloric acid being twice the theoretical amount. The mixture was stirred at 90°C for 1 hour, followed by solid-liquid separation to obtain filtrate A and leaching residue. The leaching residue was washed and dried to obtain silicon slag byproduct 1. In this embodiment, the lithium leaching rate of the lithium-containing silicon-aluminum waste was 99.52%, and the aluminum leaching rate was 98.79%. The lithium-aluminum mass ratio of filtrate A was 0.18. Fluorine-containing aluminum-aluminum waste and filtrate A were mixed at a solid-liquid mass ratio of 1:10, and calcium chloride was added. The amount of calcium added was 0.42 times the total molar amount of lithium and aluminum in the fluorine-containing aluminum-aluminum waste. After stirring evenly, the pH of the system was adjusted to 2.3, and the mixture was reacted at 95°C for 4.5 hours. After filtration, the filter cake was washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B was the lithium-rich leachate. In this embodiment, the lithium synergistic leaching rate is 96.30%, the synergistic aluminum fixation rate is 94.31%, the fluorine fixation rate is 95.92%, the lithium-aluminum mass ratio of filtrate B is 2.33, and the lithium-fluorine mass ratio is 305.22.

[0062] Example 4 Lithium-containing silicon-aluminum waste was mixed with nitric acid solution at a solid-liquid mass ratio of 1:5, with the amount of nitric acid being 1.75 times the theoretical amount. The mixture was stirred at 70°C for 2 hours, followed by solid-liquid separation to obtain filtrate A and leaching residue. The leaching residue was washed and dried to obtain silicon slag byproduct 1. In this embodiment, the lithium leaching rate of the lithium-containing silicon-aluminum waste was 96.67%, and the aluminum leaching rate was 96.24%. The lithium-aluminum mass ratio of filtrate A was 0.17. Fluorine-containing aluminum waste and filtrate A were mixed at a solid-liquid mass ratio of 1:3, and calcium sulfate was added. The amount of calcium added was 0.6 times the total molar amount of lithium and aluminum in the fluorine-containing aluminum waste. After stirring evenly, the pH of the system was adjusted to 1.17, and the mixture was reacted at 90°C for 2.0 hours. After filtration, the filter cake was washed and dried to obtain fluorinated slag byproduct 2, including aluminum fluoride. Filtrate B was the lithium-rich leachate. In this embodiment, the lithium co-leaching rate was 80.52%, the co-aluminum fixation rate was 94.10%, the fluorine fixation rate was 90.86%, the lithium-aluminum mass ratio of filtrate B was 1.48, and the lithium-fluorine mass ratio was 15.52.

[0063] Example 5 Lithium-containing silicon-aluminum waste was mixed with hydrochloric acid solution at a solid-liquid mass ratio of 1:10, with the acid amount being 4 times the theoretical amount. The mixture was stirred at 50°C for 0.5 hours, followed by solid-liquid separation to obtain filtrate A and leaching residue. The leaching residue was washed and dried to obtain silicon slag byproduct 1. In this embodiment, the lithium leaching rate of the lithium-containing silicon-aluminum waste was 78.33%, and the aluminum leaching rate was 73.11%. The lithium-aluminum mass ratio of filtrate A was 0.16. Fluorine-containing aluminum-aluminum waste and filtrate A were mixed at a solid-liquid mass ratio of 1:10, and calcium carbonate was added. The amount of calcium added was 0.4 times the total molar amount of lithium and aluminum in the fluorine-containing aluminum-aluminum waste. After stirring evenly, the pH of the system was adjusted to 1.5, and the mixture was reacted at 90°C for 2.0 hours. After filtration, the filter cake was washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B was the lithium-rich leachate. In this embodiment, the lithium co-leaching rate was 70.31%, the co-aluminum fixation rate was 98.06%, the fluorine fixation rate was 84.59%, the lithium-aluminum mass ratio of filtrate B was 4.83, and the lithium-fluorine mass ratio was 1.76.

[0064] Example 6 The leaching steps for the silicon-rich aluminum-containing lithium waste in this embodiment are the same as in Example 5. The specific co-treatment steps are as follows: Fluorine-rich aluminum-containing lithium waste and filtrate A are mixed at a solid-liquid mass ratio of 1:10, and calcium hydroxide is added. The amount of calcium added is 0.36 times the total molar amount of lithium and aluminum in the fluorine-rich aluminum-containing lithium waste. After stirring evenly, the pH of the system is adjusted to 3.5, and the reaction is carried out at 85°C for 2.5 hours. After filtration, the filter cake is washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B is the lithium-rich leachate. In this embodiment, the lithium co-leaching rate is 71.02%, the co-aluminum fixation rate is 99.16%, the fluorine fixation rate is 93.04%, the lithium-aluminum mass ratio of filtrate B is 18.21, and the lithium-fluorine mass ratio is 5.38.

[0065] Example 7 The leaching steps for the lithium-containing silicon-aluminum waste in this embodiment are the same as in Embodiment 3. The specific co-treatment steps are as follows: The lithium-containing fluorine-rich aluminum waste and filtrate A are mixed at a solid-liquid mass ratio of 1:10. After stirring evenly, the pH of the system is adjusted to 2.3. The mixture is reacted at 95°C for 4.5 hours and then filtered. The filter cake is washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B is the lithium-rich leachate. In this embodiment, the lithium co-leaching rate is 84.60%, the co-aluminum fixation rate is 99.92%, the fluorine fixation rate is 98.20%, the lithium-aluminum mass ratio of filtrate B is 165.90, and the lithium-fluorine mass ratio is 33.18.

[0066] Comparative Example 1 Fluorine-rich aluminum lithium-containing waste was mixed with hydrochloric acid solution at a solid-liquid mass ratio of 1:10, with the amount of hydrochloric acid being 4 times the theoretical amount. Calcium chloride was added, with the amount of calcium added being 0.42 times the total molar amount of lithium and aluminum in the fluorine-rich aluminum lithium-containing waste. The mixture was stirred at 95℃ for 4 hours, followed by solid-liquid separation to obtain filtrate and leaching residue. In this comparative example, the lithium leaching rate was 56.43%, the aluminum leaching rate was 0.30%, the fluorine fixation rate was 92.19%, and the lithium-aluminum mass ratio of filtrate B was 14.81, and the lithium-fluorine mass ratio was 1.75. Direct acid leaching of fluorine-rich aluminum lithium-containing waste resulted in a low lithium leaching rate and could not effectively leach the valuable element lithium.

[0067] Comparative Example 2 The leaching steps for the silicon-rich aluminum-containing lithium waste in this comparative example are the same as in Example 3. The specific co-treatment steps are as follows: Fluorine-rich aluminum-containing lithium waste and filtrate A are mixed at a solid-liquid mass ratio of 1:10. After stirring evenly, the pH of the system is adjusted to 4.53, and the reaction is carried out at 95℃ for 4.5 hours. After filtration, the filter cake is washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B is the lithium-rich leachate. In this comparative example, the lithium leaching rate is -43.32%, the lithium co-leaching rate is 63.25%, the co-aluminum fixation rate is 99.92%, the fluorine fixation rate is 98.40%, the lithium-aluminum mass ratio of filtrate B is 686.84, and the lithium-fluorine mass ratio is 25.59. When the pH of the co-treatment was 4.53, although the fluorine fixation rate and aluminum fixation rate were further improved, the valuable lithium in the fluorine-rich aluminum lithium-containing waste could not be leached. At higher pH, the lithium in the leachate of the silicon-rich aluminum lithium-containing waste formed compounds with aluminum and fluorine, and the precipitation caused a large amount of entrainment loss.

[0068] Comparative Example 3 The leaching steps for the silicon-rich aluminum-containing lithium waste in this comparative example are the same as in Example 3. The specific co-treatment steps are as follows: Fluorine-rich aluminum-containing lithium waste and filtrate A are mixed at a solid-liquid mass ratio of 1:10, and calcium chloride is added. The amount of calcium added is 0.42 times the total molar amount of lithium and aluminum in the fluorine-rich aluminum-containing lithium waste. After stirring evenly, the pH of the system is adjusted to 2.3, and the reaction is carried out at 60℃ for 4.5 hours. After filtration, the filter cake is washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B is the lithium-rich leachate. In this comparative example, the lithium co-leaching rate is 91.23%, the co-aluminum fixation rate is 41.36%, and the fluorine fixation rate is 35.16%. The lithium-aluminum mass ratio of filtrate B is 0.18, and the lithium-fluorine mass ratio is 1.65. Although a higher co-leaching rate can be achieved at excessively low temperatures, selective separation of aluminum and fluorine cannot be achieved.

[0069] Comparative Example 4 In Example 3, lithium carbonate and alumina, containing equal amounts of lithium and aluminum, were used to replace the lithium-rich silicon-aluminum waste, with other conditions remaining the same as in Example 3. In this comparative example, in step (1), all solid lithium carbonate and alumina were dissolved, resulting in a lithium co-leaching rate of 89.68%, a co-aluminum fixation rate of 98.66%, and a co-fluorine fixation rate of 95.65%. The lithium-aluminum mass ratio in filtrate B was 24.52, and the lithium-fluorine mass ratio was 10.02. The lithium-rich silicon-aluminum waste had a positive effect on promoting the leaching of lithium from the lithium-rich fluorine-aluminum waste.

[0070] Comparative Example 5 The leaching steps for the silicon-rich aluminum-containing lithium waste in this comparative example are the same as in Example 3. The specific co-treatment steps are as follows: Sodium fluoride with an equivalent molar fluorine content in the fluorine-rich aluminum-containing lithium waste from Example 3 and filtrate A are added at a solid-liquid mass ratio of 1:10. An equal amount of calcium chloride as in Example 3 is added, and the mixture is stirred and mixed thoroughly. The pH of the system is adjusted to 2.3, and the reaction is carried out at 95°C for 4.5 hours. After filtration, the filter cake is washed and dried to obtain fluorinated slag byproduct 2, such as aluminum fluoride. Filtrate B is the lithium-rich leachate. In this comparative example, the lithium leaching rate is -25.87%, the lithium co-leaching rate is 77.08%, the co-aluminum fixation rate is 99.59%, the co-fluorine fixation rate is 95.19%, and the lithium-aluminum mass ratio of filtrate B is 21.47, and the lithium-fluorine mass ratio is 5.44. It can be seen that although adding fluoride alone can remove aluminum from the leachate, it not only fails to leach lithium from the fluorine-rich aluminum-containing lithium waste but also causes a significant loss of lithium due to entrainment in leachate A.

[0071] Comparative Example 6 Lithium-containing silicon-aluminum waste was mixed with hydrochloric acid solution at a solid-liquid mass ratio of 1:3, with the amount of hydrochloric acid being 1.2 times the theoretical amount. The mixture was stirred at 60℃ for 3 hours, followed by solid-liquid separation to obtain filtrate A and leaching residue. The leaching residue was washed and dried to obtain silicon slag byproduct 1. In this comparative example, the lithium leaching rate of the lithium-containing silicon-aluminum waste was 57.52%, and the aluminum leaching rate was 52.43%, with a lithium-aluminum mass ratio of 0.16 in filtrate A. Fluorine-containing aluminum waste and filtrate A were mixed at a solid-liquid mass ratio of 1:10, and calcium sulfate was added, with the amount of calcium added being 0.33 times the total molar amount of lithium and aluminum in the fluorine-containing aluminum waste. After stirring evenly, the pH of the system was adjusted to 2, and the mixture was reacted at 80℃ for 3.0 hours. After filtration, the filter cake was washed and dried to obtain fluorinated slag byproduct 2, including aluminum fluoride. Filtrate B was the lithium-rich leachate. In this comparative example, the lithium co-leaching rate was 54.06%, the aluminum co-fixing rate was 96.76%, the fluorine co-fixing rate was 60.65%, the lithium-aluminum mass ratio of filtrate B was 2.03, and the lithium-fluorine mass ratio was 0.81.

[0072] Comparative Example 7 Fluorine-rich aluminum lithium-containing waste was mixed with calcium chloride and water at a lithium to calcium molar ratio of 1:0.5, with a solid-liquid mass ratio of 1:0.5. Nitric acid was added and the pH of the slurry was controlled at 2.5. The mixture was leached at 95℃ for 3 hours, and the solid and liquid phases were separated to obtain leachate 1 and leaching residue 1. Ammonia and calcium carbonate were added to leachate 1 to adjust and control the pH of the slurry to 10. The mixture was reacted at 50℃ for 0.1 hours, and the solid and liquid phases were separated to obtain leachate 2 and residue 2. The composition of residue phase 1, leachate 2, and residue phase 2 is shown in Table 2.

[0073] Table 2 It is evident that, although the leaching of aluminum can be effectively suppressed by using calcium salts to assist the acidic leaching of fluorine-rich aluminum lithium-containing waste, the single-pass leaching rate of the entire process is only 34.67%, meaning that most of the lithium enters the slag phase with the aluminum and cannot be effectively extracted by simple leaching.

[0074] The above embodiments should be understood as being used only to illustrate the present invention more clearly, and not to limit the scope of the present invention. After reading the present invention, any modifications of the present invention in various equivalent forms by those skilled in the art fall within the scope defined by the appended claims.

Claims

1. A method for the synergistic treatment of lithium-containing waste material, characterized in that, The method comprises the following steps: (1) leaching the silicon-rich aluminum lithium-containing waste material with an acidic solution to obtain leaching solution A and a silicon-containing leaching residue; (2) mixing the leaching solution A obtained in step (1) and the fluorine-rich aluminum lithium-containing waste material, adjusting the pH to 1-3.5, and then performing a reaction at a temperature of 80°C or higher, followed by solid-liquid separation to obtain a lithium-containing leaching solution and an aluminum-containing leaching residue.

2. The method of co-processing lithium-containing waste material according to claim 1, wherein, The molar amount of hydrogen ions in the acidic solution is 1.5 times or more than the theoretical consumption amount of lithium and aluminum in the silicon-rich aluminum lithium-containing waste material.

3. The method of co-processing lithium-containing waste material of claim 1, wherein, In step (1), the acidic solution contains at least one of hydrochloric acid, sulfuric acid and nitric acid, and the amount of the acidic solution is 1.5-4 times, preferably 1.5-2.5 times, and more preferably 1.75-2 times the theoretical consumption amount of lithium and aluminum in the silicon-rich aluminum lithium-containing waste material.

4. The method of co-processing lithium-containing waste material of claim 1, wherein, In step (1), the leaching temperature is 50-99°C, preferably 60-95°C, and further preferably 70-95°C. And / or, the leaching time is 0.5-3h, preferably 1-2.5h, and more preferably 1.5-2h. And / or, the solid-liquid mass ratio of the leaching is 1:4-10, preferably 1:4-9, and more preferably 1:5-8.

5. The method of co-processing lithium-containing waste material of claim 1, wherein, In step (1), the silicon-rich aluminum lithium-containing waste material is a powder; preferably a powder with a mesh size of 50-200; further preferably a powder with a mesh size of 80-150; and more preferably a powder with a mesh size of 100.

6. The method of co-processing lithium-containing waste material of claim 1, wherein, In step (2), the solid-liquid mass ratio of the reaction is 1:3-15, preferably 1:6-12, and more preferably 1:8-11. And / or, a calcium-containing substance is further added in step (2).

7. The method of synergistic processing of lithium-containing waste material according to claim 6, characterized in that, In step (2), the amount of the calcium-containing substance added is 5-50% of the total molar amount of lithium and aluminum in the fluorine-rich aluminum lithium-containing waste material, preferably 10-45%, and more preferably 20-35%. And / or, in step (2), the calcium-containing substance is one or more of calcium chloride, calcium sulfate, calcium carbonate, calcium hydroxide and calcium oxide, and is preferably calcium chloride.

8. The method of co-processing lithium-containing waste material of claim 1, wherein, In step (2), the reaction temperature is 85-99°C, and further preferably 90-95°C. And / or, the reaction time is 0.5-6h, preferably 3-5h, and further preferably 4-5h. And / or, the pH is adjusted to 1.5-3.5, and preferably 2-3.

9. The method of co-processing lithium-containing waste material of claim 1, wherein, Step (3) is further included, in which the lithium-rich leaching solution obtained in step (2) is subjected to impurity removal treatment, and then lithium is precipitated to obtain a lithium salt.

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