Method for extracting lithium and aluminum from red mud and lithium chinastone

Abstract

The present disclosure belongs to the field of lithium ore metallurgy. Provided in the present disclosure is a method for extracting lithium and aluminum from red mud and lithium chinastone. The method comprises: mixing lithium chinastone and red mud, subjecting the mixture to primary roasting, soaking same in water to obtain a soaked material, subjecting the soaked material to secondary roasting, subjecting the soaked material from the secondary roasting to primary leaching to obtain a leachate, and removing impurities from the resulting leachate to obtain an accepted lithium liquid product, and subjecting the resulting leaching residue to secondary leaching to obtain an aluminum solution with a high aluminum content, thereby achieving the extraction of both lithium and aluminum, and the effective separation thereof. In the method, the red mud, which is also industrial solid waste, is used as a raw material, and lithium is directly extracted from the lithium chinastone ore raw material; the method is simple and reliable, and has low production costs and a high lithium recovery rate reaching 98.7% or more, thereby facilitating large-scale industrialization; and the method uses a large amount of industrial solid waste, and is more environmentally friendly with considerable economic benefits.

Description

A method for extracting lithium and aluminum using red mud and lithium porcelain stone Technical Field

[0001] The present invention belongs to the field of metallurgical production and relates to a lithium ore extraction technology, and in particular to a method for extracting lithium and aluminum by using red mud in conjunction with lithium porcelain stone. Background Art

[0002] Currently, there are two main methods for producing high-purity lithium products (lithium extraction technology): lithium extraction from lithium ore and lithium extraction from salt lakes. Salt lake extraction has consistently accounted for a relatively low share of lithium production, primarily due to two factors: 1) Salt lake resources are generally concentrated in plateaus, where high altitudes, harsh climate conditions, and inconvenient transportation are common; and 2) the extraction technology for high-lithium-to-magnesium ratio brines is still in its early stages of exploration.

[0003] Therefore, the extraction of lithium from ores still occupies a dominant position. The current mainstream raw materials for extracting lithium from ores are mainly spodumene and lithium clay. However, lithium ore resources are limited, and in the future they will inevitably face the unfavorable situation of reduced grade and resource depletion.

[0004] Lithium chinastone is a type of lithium ore, but its chemical composition and structure differ from those of spodumene and lithium clay. Its lithium content is lower than spodumene, and its structure is more complex. Currently, commonly used lithium extraction methods are not suitable for lithium chinastone. Lithium extraction from lithium chinastone remains difficult, often requiring specialized protocols.

[0005] Currently, lithium extraction from lithium-based porcelain stone is generally divided into two steps: first, extracting the lepidolite from the ore through a mineral processing process; then, extracting the lithium from the lepidolite. CN108014901A discloses a process for extracting lepidolite from lithium-based porcelain stone ore. This method uses impact treatment of the ore particles, gravity flotation, and continuous high-gradient strong magnetic purification to obtain feldspar and lepidolite. However, this method is complex and lengthy, and only produces lepidolite concentrate. CN113957268A discloses a method for extracting lithium from lithium-based porcelain stone raw materials. This method uses a composite salt composed of sodium sulfate, calcium sulfate, and calcium carbonate as a roasting agent. However, this method produces a large amount of calcium-silicon waste residue, which is difficult to process.

[0006] It can be seen that the existing technology will produce a large amount of tailings in the process of obtaining lithium mica from lithium porcelain stone, which will consume a lot of energy and manpower, resulting in high costs; moreover, in the complete lithium extraction process from lithium porcelain stone, the lithium recovery rate is only about 65%, which will cause a large amount of waste of resources.

[0007] Therefore, there is still a need to develop a new solution for extracting lithium from lithium porcelain stone ore to provide a new way to extract and produce lithium, thereby alleviating the tightening and tense trend of lithium extraction from ore and ensuring the low cost and stable supply of lithium-containing products.

[0008] Summary of the Invention

[0009] The following is a summary of the subject matter described in detail herein. This summary is not intended to limit the scope of the claims.

[0010] The present invention provides a method for extracting lithium and aluminum from red mud in combination with lithium porcelain stone. The method involves first mixing the lithium porcelain stone with red mud and calcining it once. The mixture is then soaked in water to obtain a soaked material, which is then calcined a second time. The resulting leached solution is then leached once. The resulting leachate is then cleaned to produce a qualified lithium liquid product. The resulting leached residue is then leached a second time to produce an aluminum solution with a high aluminum content, thereby achieving simultaneous lithium and aluminum extraction and effective separation. The method utilizes red mud, also an industrial solid waste, as a raw material, directly extracting lithium from the lithium porcelain stone ore. The process is simple and reliable, with low production costs and a high lithium recovery rate of over 98.7%. This method is suitable for large-scale industrialization and utilizes industrial solid waste in a green and environmentally friendly manner, resulting in considerable economic benefits.

[0011] To achieve this goal, the present disclosure adopts the following technical solutions:

[0012] The present disclosure provides a method for extracting lithium and aluminum from red mud in combination with lithium porcelain stone, comprising the following steps:

[0013] S1: lithium porcelain stone and red mud are mixed and calcined once to obtain a primary calcined material;

[0014] S2: mixing the primary roasted material with red mud and soaking it to obtain a soaked material;

[0015] S3: secondary roasting the soaked material to obtain a secondary roasted material;

[0016] S4: Leaching the secondary roasted material once to obtain a leachate and leach residue;

[0017] S5: removing impurities from the leaching solution to obtain metallurgical slag and lithium liquid products;

[0018] S6: performing secondary leaching on the leached residue to obtain silicon residue and aluminum solution;

[0019] There is no particular order for steps S5 and S6.

[0020] Red mud is an industrial solid waste produced during the extraction of alumina in the aluminum industry. It is called red mud because of its high iron oxide content and its resemblance to reddish soil. However, some mud is brown or even off-white due to its lower iron oxide content. Red mud is an insoluble residue that can be divided into sintered, Bayer, and combined processes. Its main components are SiO2, Al2O3, CaO, and Fe2O3. With the increasing stockpiles of red mud and the increasing pollution it causes to the environment, maximizing its resource utilization is imperative.

[0021] This paper proposes a method for efficiently extracting lithium aluminum by synergizing red mud and lithium porcelain stone ore. The method uses a pyrometallurgical process to extract lithium and realizes efficient lithium extraction based on the mutual replacement of ions at high temperature. First, the lithium porcelain stone and red mud are mixed to fully contact each other, and then the mixture is roasted once. At high temperature, the crystal structure of the lithium porcelain stone ore changes, the connecting bonds of the ions are broken, and the lithium + It is easy to replace metal ions such as Fe, Na and K in red mud; then, the primary roasted material is added to the red mud mixture and soaked in water. The red mud itself is strongly alkaline. After the soaked material is roasted for the second time, the ionic bond is broken more thoroughly, and the lithium extraction effect is better. Finally, under the action of the leaching agent, the lithium is dissolved into the solution for leaching (i.e., primary leaching). The obtained leachate is a lithium liquid with a high lithium content. The lithium liquid is further impurity-removed to obtain a lithium liquid product. The lithium liquid product at this time can be directly used for lithium precipitation. Of course, technical personnel in this field can also perform other treatments or use the leachate with a high lithium content according to actual needs; the present invention discloses that the leached material obtained by the primary leaching can selectively extract aluminum through secondary leaching. The method has a high lithium extraction rate, which can reach up to more than 98.7%. The obtained aluminum solution has a high aluminum content and can also be further used in subsequent applications.

[0022] It should be noted that the present disclosure adds red mud in two steps to avoid the problem of excessive red mud addition in a single step, which can result in hardened slag and inconvenient subsequent handling. Therefore, a two-step addition is more reasonable, with the second step primarily intended to ensure a more thorough reaction. Similarly, when red mud is added in two steps but the total amount added is excessive, it not only results in waste but also leads to severe material agglomeration, cumbersome subsequent handling, increased processing costs, and a negative impact on the leaching rate.

[0023] This disclosed method aims to address the complex and lengthy lithium extraction process from traditional lithium-based porcelain stoneware, characterized by low lithium recovery rates and the difficulty in resource utilization of waste residues. The method not only enables the extensive utilization of industrial solid waste red mud, but also allows for direct lithium extraction from lithium-based porcelain stone ore. This simple and reliable process offers low production costs, a high lithium recovery rate, and ease of industrialization. It possesses broad applicability, is in line with green development, and demonstrates cutting-edge innovation.

[0024] The following are optional technical solutions of the present disclosure, but are not intended to limit the technical solutions provided by the present disclosure. Through the following technical solutions, the technical objectives and beneficial effects of the present disclosure can be better achieved and realized.

[0025] As an optional technical solution of the present disclosure, the lithium porcelain stone and red mud in step S1 are first mixed and crushed, and then roasted once.

[0026] In one embodiment, after the mixing and crushing, the particle size of the lithium porcelain stone and the red mud is 150-300 mesh, for example, 150 mesh, 160 mesh, 170 mesh, 180 mesh, 190 mesh, 200 mesh, 210 mesh, 220 mesh, 230 mesh, 240 mesh, 250 mesh, 260 mesh, 270 mesh, 280 mesh, 290 mesh or 300 mesh, etc., and can be selected from 200-250 mesh, but is not limited to the listed values. Other values ​​not listed within the above numerical range are also applicable.

[0027] As an optional technical solution of the present disclosure, the mass ratio of the lithium porcelain stone to the red mud in step S1 is 1:(0.2-0.8), for example, 1:0.2, 1:0.25, 1:0.3, 1:0.35, 1:0.4, 1:0.45, 1:0.5, 1:0.55, 1:0.6, 1:0.65, 1:0.7, 1:0.75 or 1:0.8, etc., and can be optionally 1:(0.4-0.6), but is not limited to the listed values, and other values ​​not listed within the above numerical range are also applicable.

[0028] As an optional technical solution disclosed in the present invention, the temperature of the primary roasting in step S1 is 750-950°C, for example, 750°C, 760°C, 770°C, 780°C, 790°C, 800°C, 810°C, 820°C, 830°C, 840°C, 850°C, 860°C, 870°C, 880°C, 890°C, 900°C, 910°C, 920°C, 930°C, 940°C or 950°C, etc., and can be selected from 800-850°C, but is not limited to the listed values, and other values ​​not listed within the above numerical range are also applicable.

[0029] If the primary roasting temperature is too low, the roasting effect is poor, and if the temperature is too high, the leaching rate will not be significantly improved.

[0030] In one embodiment, the time for the first calcination in step S1 is 1 to 3 hours, for example, 1 hour, 1.2 hours, 1.4 hours, 1.6 hours, 1.8 hours, 2 hours, 2.2 hours, 2.4 hours, 2.6 hours, 2.8 hours or 3 hours, etc., and can be optionally 1.5 to 2.5 hours, but is not limited to the listed values. Other values ​​not listed within the above numerical range are also applicable.

[0031] As an optional technical solution of the present disclosure, the red mud described in step S2 is red mud with a particle size of 150 to 300 mesh after crushing, such as 150 mesh, 160 mesh, 170 mesh, 180 mesh, 190 mesh, 200 mesh, 210 mesh, 220 mesh, 230 mesh, 240 mesh, 250 mesh, 260 mesh, 270 mesh, 280 mesh, 290 mesh or 300 mesh, etc., and can be selected as 200 to 250 mesh, but is not limited to the listed values, and other values ​​not listed within the above numerical range are also applicable.

[0032] As an optional technical solution of the present disclosure, the mass ratio of the primary roasting material to the red mud in step S2 is 1:(0.2~0.8), for example, 1:0.2, 1:0.25, 1:0.3, 1:0.35, 1:0.4, 1:0.45, 1:0.5, 1:0.55, 1:0.6, 1:0.65, 1:0.7, 1:0.75 or 1:0.8, etc., and can be selected as 1:(0.4~0.6), but is not limited to the listed values, and other values ​​not listed within the above numerical range are also applicable.

[0033] As an optional technical solution of the present disclosure, the soaking liquid in step S2 includes water.

[0034] In one embodiment, the mass ratio of the total mass of the primary roasted material and the red mud after mixing to water is (0.8-1.5):1, for example, 0.8:1, 0.85:1, 0.9:1, 0.95:1, 1:1, 1.05:1, 1.1:1, 1.15:1, 1.2:1, 1.25:1, 1.3:1, 1.35:1, 1.4:1, 1.45:1 or 1.5:1, etc., and can be selected as (0.8-1):1, but is not limited to the listed values, and other values ​​not listed within the above numerical range are also applicable.

[0035] The soaking material obtained by the present invention is in a state similar to a soft dough or a thin dough formed by mixing flour and water, and therefore does not require solid-liquid separation and can be directly subjected to secondary roasting.

[0036] As an optional technical solution of the present disclosure, the secondary calcination in step S3 is performed under vacuum.

[0037] The secondary roasting is carried out under vacuum to apply high pressure to the material to make the ion exchange more thorough, which is beneficial to improving the leaching rate; at this time, the vacuum degree is adjusted as needed, and a conventional vacuum furnace can be used for sintering.

[0038] In one embodiment, the temperature of the secondary calcination is 850-1000°C, for example, 850°C, 860°C, 870°C, 880°C, 890°C, 900°C, 910°C, 920°C, 930°C, 940°C, 950°C, 960°C, 970°C, 980°C, 990°C or 1000°C, etc., and can be selected from 850-950°C, but is not limited to the listed values. Other values ​​not listed within the above numerical range are also applicable.

[0039] In one embodiment, the secondary calcination time is 2 to 5 hours, for example, 2 hours, 2.2 hours, 2.4 hours, 2.6 hours, 2.8 hours, 3 hours, 3.2 hours, 3.4 hours, 3.6 hours, 3.8 hours, 4 hours, 4.2 hours, 4.4 hours, 4.6 hours, 4.8 hours or 5 hours, etc., and can be selected from 3 to 4.5 hours, but is not limited to the listed values. Other values ​​not listed within the above numerical range are also applicable.

[0040] As an optional technical solution of the present disclosure, the primary leaching in step S4 and the secondary leaching in step S6 are to mix the secondary roasted material and the leaching residue with the leaching agent respectively, and perform solid-liquid separation after the leaching is completed.

[0041] In one embodiment, the leaching agent includes any one of water, hydrochloric acid solution or sulfuric acid solution, or a combination of at least two of them. Typical but non-limiting examples of the combination include a combination of water and hydrochloric acid solution, a combination of water and sulfuric acid solution, or a combination of hydrochloric acid solution and sulfuric acid solution, and sulfuric acid solution can be selected.

[0042] In one embodiment, when the leaching agent includes a hydrochloric acid solution and / or a sulfuric acid solution, the concentration of the leaching agent is 6 to 12 mol / L, for example, 6 mol / L, 6.5 mol / L, 7 mol / L, 7.5 mol / L, 8 mol / L, 8.5 mol / L, 9 mol / L, 9.5 mol / L, 10 mol / L, 10.5 mol / L, 11 mol / L, 11.5 mol / L or 12 mol / L, etc., and can be optionally 8 to 10 mol / L, but is not limited to the listed values, and other values ​​not listed within the above numerical range are also applicable.

[0043] In one embodiment, in the primary leaching and the secondary leaching, the liquid-solid ratio of the leaching agent to the secondary roasting material and the leaching residue is (3-5) mL: 1g, for example, 3 mL: 1g, 3.2 mL: 1g, 3.4 mL: 1g, 3.6 mL: 1g, 3.8 mL: 1g, 4 mL: 1g, 4.2 mL: 1g, 4.4 mL: 1g, 4.6 mL: 1g, 4.8 mL: 1g or 5 mL: 1g, etc., and can be selected as 3 mL: 1g, but is not limited to the listed values, and other values ​​not listed within the above numerical range are equally applicable.

[0044] As an optional technical solution of the present disclosure, the impurity removal method described in step S5 includes using a chemical method to add liquid alkali to remove nickel, cobalt and manganese, and then using resin to deeply remove calcium and magnesium.

[0045] The purpose of the impurity removal is to convert the leachate into a lithium liquid product that can meet the needs of lithium precipitation, mainly removing impurity elements such as nickel, cobalt, manganese, calcium, magnesium, phosphorus, and fluorine. The resulting metallurgical slag is mainly solid slag produced during the impurity removal process, and its components include nickel hydroxide, cobalt, manganese, calcium fluoride, etc.

[0046] As an optional technical solution of the present disclosure, the method includes the following steps:

[0047] (1) mixing and crushing lithium porcelain stone and red mud in a mass ratio of 1:(0.4-0.6) to obtain a mixed crushed material, wherein the particle size of the lithium porcelain stone and the red mud is 200-250 mesh, and calcining the mixed crushed material at 800-850° C. for 1.5-2.5 hours to obtain a primary calcined material;

[0048] (2) taking red mud with a crushed particle size of 200-250 mesh, mixing the primary roasted material with the red mud in a mass ratio of 1:(0.4-0.6), and soaking them in water, controlling the mass ratio of the total mass of the primary roasted material and the red mud to water to be (0.8-1):1, to obtain a soaked material;

[0049] (3) secondary calcining the soaked material at 850-950° C. under vacuum for 3-4.5 hours to obtain a secondary calcined material;

[0050] (4) mixing the secondary roasted material with a leaching agent, wherein the leaching agent includes water, or any one of a hydrochloric acid solution or a sulfuric acid solution with a concentration of 8 to 10 mol / L, or a combination of at least two of the two, controlling the liquid-solid ratio of the leaching agent to the secondary roasted material to be (3 to 5) mL:1 g, performing a primary leaching on the secondary roasted material, and obtaining a leachate and a leaching residue after solid-liquid separation;

[0051] (5) removing impurities from the leachate to obtain metallurgical slag and lithium liquid products;

[0052] (6) mixing the leached residue with a leaching agent, wherein the leaching agent includes water, or any one of a hydrochloric acid solution or a sulfuric acid solution with a concentration of 8 to 10 mol / L, or a combination of at least two of the two, controlling the liquid-solid ratio of the leaching agent to the leached residue to be (3 to 5) mL:1 g, performing secondary leaching on the leached residue, and obtaining silicon slag and aluminum solution after solid-liquid separation;

[0053] There is no particular order for steps S5 and S6.

[0054] Compared with the existing technical solutions, the present disclosure has at least the following beneficial effects:

[0055] The disclosed method for extracting lithium and aluminum from red mud in conjunction with lithium porcelain stone can address the complex and lengthy processes, low lithium recovery rates, and difficulty in resource utilization of waste residues associated with traditional lithium porcelain stone extraction. This method not only enables the extensive utilization of industrial solid waste red mud, but also allows for the direct extraction of lithium from lithium porcelain stone ore. The process is simple and reliable, with low production costs and a high lithium recovery rate of over 98.7%. The resulting aluminum solution is a high-aluminum solution, and the resulting silicon slag is silicon slag. This method is easily scalable for industrial production, has broad applicability, is in line with green development, and offers considerable economic benefits.

[0056] Still other aspects will become apparent upon reading and understanding the accompanying drawings and detailed description. BRIEF DESCRIPTION OF THE DRAWINGS

[0057] The accompanying drawings are used to provide a further understanding of the technical solution of this article and constitute a part of the specification. Together with the embodiments of this application, they are used to explain the technical solution of this article and do not constitute a limitation on the technical solution of this article.

[0058] FIG1 is a schematic flow chart of the method for extracting lithium aluminum from red mud in conjunction with lithium porcelain stone according to Example 1. DETAILED DESCRIPTION

[0059] The technical solution of the present disclosure is further described below through specific implementation methods. Those skilled in the art should understand that the embodiments are only used to help understand the present disclosure and should not be regarded as specific limitations of the present disclosure.

[0060] Example 1

[0061] This embodiment provides a method for extracting lithium aluminum from red mud in conjunction with lithium porcelain stone, as shown in FIG1 , the method comprising the following steps:

[0062] S1: 500g of lithium porcelain stone ore and 250g of red mud are mixed and crushed to 200 mesh using a crusher, with the mass ratio of lithium porcelain stone to red mud being 1:0.5, to obtain a mixed crushed material; the mixed crushed material is added to a roasting furnace for primary roasting at a roasting temperature of 850°C for 2 hours to obtain a primary roasted material;

[0063] S2: The primary calcined material obtained in S1 is added to a 200-mesh red mud mixture, and then soaked in deionized water. The mass ratio of the primary calcined material to the red mud is 1:0.5, and the mass ratio of the mixture to deionized water is 1:1, to obtain a soaked material;

[0064] S3: adding the soaked material of S2 into a vacuum furnace for secondary calcination at a temperature of 950°C for 4 hours to obtain a secondary calcined material;

[0065] S4: Mix the leaching agent with the secondary roasted material of S3 at a liquid-solid ratio of 3 mL:1 g. Use sulfuric acid as the leaching agent at a concentration of 8 mol / L. Perform a primary leaching. After solid-liquid separation, obtain the leachate and leach residue.

[0066] S5: removing impurities from the leaching solution obtained in S4 to obtain a qualified lithium liquid product and metallurgical slag;

[0067] S6: The leaching agent and the leaching residue obtained in S4 are mixed at a liquid-solid ratio of 3 mL:1 g. Sulfuric acid is used as the leaching agent at a concentration of 8 mol / L. A secondary leaching is performed to obtain an aluminum solution and silicon slag after solid-liquid separation.

[0068] Example 2

[0069] This embodiment provides a method for extracting lithium and aluminum from red mud in combination with lithium porcelain stone, the method comprising the following steps:

[0070] S1: 500g of lithium porcelain stone ore and 100g of red mud are mixed and crushed to 200 mesh using a crusher, with the mass ratio of lithium porcelain stone to red mud being 1:0.2, to obtain a mixed crushed material; the mixed crushed material is added to a roasting furnace for primary roasting at a roasting temperature of 950°C for 3 hours to obtain a primary roasted material;

[0071] S2: The primary calcined material obtained in S1 is added to a 200-mesh red mud mixture, and then soaked in deionized water. The mass ratio of the primary calcined material to the red mud is 1:0.6, and the mass ratio of the mixture to deionized water is 1:0.8, to obtain a soaked material;

[0072] S3: adding the soaked material of S2 into a vacuum furnace for secondary calcination at a temperature of 1000°C for 5 hours to obtain a secondary calcined material;

[0073] S4: The leaching agent is mixed with the secondary roasted material of S3 at a liquid-solid ratio of 3mL:1g. Sulfuric acid is used as the leaching agent at a concentration of 12mol / L. A leaching is performed once. After solid-liquid separation, a leachate and a leach residue are obtained;

[0074] S5: removing impurities from the leaching solution obtained in S4 to obtain a qualified lithium liquid product and metallurgical slag;

[0075] S6: The leaching agent and the leaching residue obtained in S4 are mixed at a liquid-solid ratio of 3 mL:1 g. Sulfuric acid is used as the leaching agent at a concentration of 12 mol / L. A secondary leaching is performed to obtain an aluminum solution and silicon slag after solid-liquid separation.

[0076] Example 3

[0077] This embodiment provides a method for extracting lithium and aluminum from red mud in combination with lithium porcelain stone, the method comprising the following steps:

[0078] S1: 500 g of lithium porcelain stone ore and 400 g of red mud are mixed and crushed to 200 mesh using a crusher, with the mass ratio of lithium porcelain stone to red mud being 1:0.8, to obtain a mixed crushed material; the mixed crushed material is added to a roasting furnace for primary roasting at a roasting temperature of 750° C. for 1 hour to obtain a primary roasted material;

[0079] S2: The primary calcined material obtained in S1 is added to a 200-mesh red mud mixture, and then soaked in deionized water. The mass ratio of the primary calcined material to the red mud is 1:0.5, and the mass ratio of the mixture to deionized water is 1:0.8, to obtain a soaked material;

[0080] S3: adding the soaked material of S2 into a vacuum furnace for secondary calcination at a temperature of 850°C for 2 hours to obtain a secondary calcined material;

[0081] S4: The leaching agent is mixed with the secondary roasted material of S3 at a liquid-solid ratio of 3mL:1g. Sulfuric acid is used as the leaching agent at a concentration of 6mol / L. A leaching is performed once. After solid-liquid separation, a leachate and a leach residue are obtained;

[0082] S5: removing impurities from the leaching solution obtained in S4 to obtain a qualified lithium liquid product and metallurgical slag;

[0083] S6: The leaching agent and the leaching residue obtained in S4 are mixed at a liquid-solid ratio of 3 mL:1 g. Sulfuric acid is used as the leaching agent at a concentration of 6 mol / L. A secondary leaching is performed. After solid-liquid separation, an aluminum solution and silicon slag are obtained.

[0084] Example 4

[0085] This embodiment provides a method for extracting lithium and aluminum from red mud in combination with lithium porcelain stone, the method comprising the following steps:

[0086] S1: 500g of lithium porcelain stone ore and 250g of red mud are mixed and crushed to 200 mesh using a crusher, with the mass ratio of lithium porcelain stone to red mud being 1:0.5, to obtain a mixed crushed material; the mixed crushed material is added to a roasting furnace for primary roasting at a roasting temperature of 850°C for 1 hour to obtain a primary roasted material;

[0087] S2: The primary calcined material obtained in S1 is added to a 200-mesh red mud mixture, and then soaked in deionized water. The mass ratio of the primary calcined material to the red mud is 1:0.5, and the mass ratio of the mixture to deionized water is 1:0.8, to obtain a soaked material;

[0088] S3: adding the soaked material of S2 into a vacuum furnace for secondary calcination at a temperature of 850°C for 2 hours to obtain a secondary calcined material;

[0089] S4: The leaching agent is mixed with the secondary roasted material of S3 at a liquid-solid ratio of 3mL:1g. The leaching agent is hydrochloric acid with a concentration of 6mol / L. A primary leaching is performed. After solid-liquid separation, a leachate and a leach residue are obtained;

[0090] S5: removing impurities from the leaching solution obtained in S4 to obtain a qualified lithium liquid product and metallurgical slag;

[0091] S6: The leaching agent and the leaching residue obtained in S4 are mixed at a liquid-solid ratio of 3 mL:1 g. The leaching agent is hydrochloric acid with a concentration of 6 mol / L. Secondary leaching is performed. After solid-liquid separation, aluminum solution and silicon slag are obtained.

[0092] Example 5

[0093] This embodiment provides a method for extracting lithium and aluminum from red mud in combination with lithium porcelain stone, the method comprising the following steps:

[0094] S1: 500g of lithium porcelain stone ore and 250g of red mud are mixed and crushed to 200 mesh using a crusher, with the mass ratio of lithium porcelain stone to red mud being 1:0.5, to obtain a mixed crushed material; the mixed crushed material is added to a roasting furnace for primary roasting at a roasting temperature of 850°C for 1 hour to obtain a primary roasted material;

[0095] S2: The primary calcined material obtained in S1 is added to a 200-mesh red mud mixture, and then soaked in deionized water. The mass ratio of the primary calcined material to the red mud is 1:0.5, and the mass ratio of the mixture to deionized water is 1:0.8, to obtain a soaked material;

[0096] S3: adding the soaked material of S2 into a vacuum furnace for secondary calcination at a temperature of 850°C for 2 hours to obtain a secondary calcined material;

[0097] S4: The leaching agent is mixed with the secondary roasted material of S3 at a liquid-solid ratio of 3mL:1g. Pure water is used as the leaching agent for a primary leaching. After solid-liquid separation, the leachate and leach residue are obtained;

[0098] S5: removing impurities from the leaching solution obtained in S4 to obtain a qualified lithium liquid product and metallurgical slag;

[0099] S6: The leaching agent and the leaching residue obtained in S4 are mixed at a liquid-solid ratio of 3 mL:1 g. The leaching agent is hydrochloric acid with a concentration of 6 mol / L. Secondary leaching is performed. After solid-liquid separation, aluminum solution and silicon slag are obtained.

[0100] Example 6

[0101] This example provides a method for extracting lithium aluminum using red mud in combination with lithium porcelain stone. The method is identical to Example 1 except that in step S1, the amount of red mud is adjusted from 250 g to 150 g, thereby changing the mass ratio of lithium porcelain stone to red mud from 1:0.5 to 1:0.3.

[0102] Example 7

[0103] This example provides a method for extracting lithium aluminum using red mud in combination with lithium porcelain stone. The method is identical to Example 1 except that in step S1, the amount of red mud is adjusted from 250 g to 425 g, thereby changing the mass ratio of lithium porcelain stone to red mud from 1:0.5 to 1:0.85.

[0104] Example 8

[0105] This embodiment provides a method for extracting lithium aluminum from red mud in combination with lithium porcelain stone. The method is identical to that of Example 1 except that the primary calcination temperature is adjusted from 850°C to 700°C in step S1.

[0106] Example 9

[0107] This embodiment provides a method for extracting lithium aluminum from red mud in combination with lithium porcelain stone. The method is identical to that of Example 1 except that the primary calcination temperature is adjusted from 850°C to 1000°C in step S1.

[0108] Example 10

[0109] This embodiment provides a method for extracting lithium aluminum using red mud in conjunction with lithium porcelain stone. The method is identical to that of Example 1, except that the amount of red mud is adjusted in step S2 so that the mass ratio of the primary roasted material to the red mud is changed from 1:0.5 to 1:0.15.

[0110] Example 11

[0111] This embodiment provides a method for extracting lithium aluminum using red mud in conjunction with lithium porcelain stone. The method is identical to that of Example 1, except that the amount of red mud is adjusted in step S2 so that the mass ratio of the primary roasted material to the red mud changes from 1:0.5 to 1:0.85.

[0112] Example 12

[0113] This embodiment provides a method for extracting lithium aluminum from red mud in combination with lithium porcelain stone. The method is identical to that of Example 1 except that the secondary calcination temperature is adjusted from 950°C to 800°C in step S3.

[0114] Example 13

[0115] This embodiment provides a method for extracting lithium aluminum from red mud in combination with lithium porcelain stone. The method is identical to that of embodiment 1 except that the secondary calcination temperature is adjusted from 950°C to 1050°C in step S3.

[0116] Example 14

[0117] This embodiment provides a method for extracting lithium aluminum from red mud in conjunction with lithium porcelain stone. The secondary roasting in step S3 of the method is performed in a roasting furnace, not in a vacuum furnace, that is, no vacuum environment is applied. Other than this, the other conditions are exactly the same as in Example 1.

[0118] Comparative Example 1

[0119] This comparative example provides a method for extracting lithium aluminum using lithium porcelain stone. The method does not use red mud in step S1, and uses pure water for leaching in step S6. The method comprises the following steps:

[0120] S1: crushing 500g of lithium porcelain stone ore into 200 mesh using a crusher to obtain crushed material; adding the crushed material to a roasting furnace for primary roasting at a roasting temperature of 850°C for 2 hours to obtain a primary roasted material;

[0121] S2: The primary calcined material obtained in S1 is added to a 200-mesh red mud mixture, and then soaked in deionized water. The mass ratio of the primary calcined material to the red mud is 1:0.5, and the mass ratio of the mixture to deionized water is 1:0.8, to obtain a soaked material;

[0122] S3: adding the soaked material of S2 into a vacuum furnace for secondary calcination at a temperature of 950°C for 4 hours to obtain a secondary calcined material;

[0123] S4: The leaching agent is mixed with the secondary roasted material of S3 at a liquid-solid ratio of 3mL:1g. The leaching agent is hydrochloric acid with a concentration of 8mol / L. A leaching is performed once. After solid-liquid separation, a leachate and a leach residue are obtained;

[0124] S5: removing impurities from the leachate obtained in S4 to obtain a lithium liquid product and metallurgical slag;

[0125] S6: The leaching agent and the leaching residue obtained in S4 are mixed at a liquid-solid ratio of 3 mL:1 g. Pure water is used as the leaching agent for secondary leaching. After solid-liquid separation, aluminum solution and silicon slag are obtained.

[0126] Comparative Example 2

[0127] This comparative example provides a method for extracting lithium aluminum using lithium porcelain stone. The method does not use red mud in step S2, that is, step S2 is:

[0128] The primary calcined material obtained in S1 is soaked in deionized water, with the mass ratio of the primary calcined material to the deionized water being 1:1, to obtain a soaked material;

[0129] Apart from this, other conditions are exactly the same as those in Example 1.

[0130] Comparative Example 3

[0131] This comparative example provides a method for extracting lithium aluminum using lithium porcelain stone. The method does not use red mud in step S1, that is, step S1 is:

[0132] 500g of lithium porcelain stone ore was crushed into 200 meshes by a crusher to obtain crushed material; the crushed material was added to a roasting furnace for primary roasting at a roasting temperature of 850°C and a roasting time of 2h to obtain a primary roasted material;

[0133] Apart from this, other conditions are exactly the same as those in Example 1.

[0134] Comparative Example 4

[0135] This comparative example provides a method for extracting lithium aluminum using lithium porcelain stone. In the method, deionized water is not used for soaking in step S2. The primary roasted material obtained in step S1 is directly added to 200-mesh red mud in a mass ratio of 1:0.5, and the resulting mixture is directly subjected to secondary roasting in step S3.

[0136] Apart from this, other conditions are exactly the same as those in Example 1.

[0137] Comparative Example 5

[0138] This comparative example provides a method for extracting lithium aluminum using lithium porcelain stone. In the method, the first roasting is not performed in step S1. The obtained mixed crushed material is directly added to a 200-mesh red mud mixture in step S2, and the mass ratio of the mixed crushed material to the red mud is controlled to be 1:0.5, and then the mixed crushed material is added to deionized water for soaking.

[0139] Apart from this, other conditions are exactly the same as those in Example 1.

[0140] Comparative Example 6

[0141] This comparative example provides a method for extracting lithium aluminum using lithium porcelain stone. In the method, no secondary roasting is performed in step S3. Instead, the soaked material in step S2 is dried at 100-120°C, and the obtained dried material is directly subjected to primary leaching in step S4.

[0142] Apart from this, other conditions are exactly the same as those in Example 1.

[0143] The lithium porcelain stone, red mud, leached residue, leachate, lithium solution product, and aluminum solution disclosed herein were all tested using an atomic absorption spectrophotometer in accordance with the national standard GB / T 15337-2008. The lithium content of the lithium porcelain stone used in the examples and comparative examples was measured to be 0.56 wt %, and the aluminum content of the red mud was 15 wt %. The lithium-aluminum leaching rate was calculated using the following formula:

[0144] Lithium leaching rate = (volume of lithium-containing leachate * lithium concentration in leachate) / (mass of lithium porcelain stone * percentage of lithium in lithium porcelain stone); lithium-containing leachate is the lithium liquid product;

[0145] Aluminum leaching rate = (volume of aluminum-containing leachate * aluminum concentration in leachate) / ((mass of red mud in mixed crushed material + mass of red mud added to the mixed material before immersion) * percentage of aluminum in red mud); the aluminum-containing leachate is the aluminum solution;

[0146] The results are recorded in Tables 1 and 2.

[0147] Table 1 Primary leaching results

[0148] Table 2 Secondary leaching results

[0149] It can be seen from Table 1 and Table 2 that:

[0150] The method provided by the present disclosure can change the crystal structure of lithium porcelain stone ore at high temperature during the first and second roasting, break the ionic connection bond, and + It easily replaces metal ions such as Fe, Na, and K, thereby increasing the final leaching rate of lithium. A comparison of Example 1 with Examples 2 and 3 shows that the best results were achieved when the mass ratio of lithium porcelain stone to red mud was 1:0.5, and the mass ratio of primary calcined material to red mud was 1:0.5. The use of 8 mol / L sulfuric acid achieved the best results, with lithium leaching rates reaching a maximum of 98.74% and aluminum leaching rates reaching a maximum of 99.16%. Example 4 used hydrochloric acid as the leaching agent, and found that its lithium leaching rates reached 87.60% and aluminum leaching rates reached 86.93%.

[0151] In Comparative Example 1, no red mud was added during crushing, and pure water was used for secondary leaching. Combined with Example 5, adding pure water for leaching aluminum was difficult, while lithium could be leached with pure water and achieved good leaching results. In Comparative Example 2, no red mud was added to the primary calcined material for leaching. The results showed that the leaching effect was somewhat worse, with a lithium leaching rate of only 77.38% and an aluminum leaching rate of 81.81%. This indicates that the primary calcination resulted in incomplete cleavage of the ionic bonds, resulting in a low leaching rate.

[0152] Comparing Example 1 with Examples 6 and 7, it was found that: if the amount of red mud added in the first roasting was too little, the leaching rate would be low and the reaction would be incomplete; if the amount of red mud added was too much, the material would be severely agglomerated, and subsequent leaching would be difficult, affecting the leaching rate;

[0153] Comparing Example 1 with Examples 8 and 9, it was found that the primary roasting temperature had a significant effect on the leaching rate of the material. When the temperature was low, the leaching rate of Li and Al was only about 87%, while the effect of high temperature on the leaching rate was not very obvious.

[0154] Comparing Example 1 with Examples 10 and 11, it was found that: the addition of too little red mud in the vacuum furnace resulted in incomplete reaction, with the Li and Al leaching rates being only about 85%; the addition of too much red mud resulted in severe material agglomeration, making subsequent leaching difficult and affecting the leaching rate;

[0155] Comparing Example 1 with Examples 12 and 13, it was found that high vacuum roasting temperature had little effect on the leaching rates of Li and Al, while low temperature would lead to a decrease in the leaching rate.

[0156] Comparing Example 1 with Example 14, it was found that the leaching rate dropped sharply, and the Li leaching rate and Al leaching rate were about 85%, indicating that the vacuum environment promoted ion exchange;

[0157] Comparing Example 1 with Comparative Examples 3-6, it was found that not adding the primary roasted red mud, not performing roasting in a roasting furnace, and not performing roasting in a vacuum furnace would directly lead to a decrease in the leaching rate. The Li and Al leaching rates were only about 60%. The effect of not immersing was not very obvious, with the Li leaching rate being 88.74% and the Al leaching rate being 92.37%.

Claims

1. A method for extracting lithium aluminum by using red mud in combination with lithium porcelain stone, comprising the following steps: S1: mixing lithium porcelain stone and red mud and then calcining them once to obtain a primary calcined material; S2: mixing the primary roasted material with red mud and soaking it to obtain a soaked material; S3: secondary roasting the soaked material to obtain a secondary roasted material; S4: Leaching the secondary roasted material once to obtain a leachate and leach residue; S5: removing impurities from the leaching solution to obtain metallurgical slag and lithium liquid products; S6: performing secondary leaching on the leached residue to obtain silicon residue and aluminum solution; in, Steps S5 and S6 are performed in no particular order.

2. The method according to claim 1, wherein: In step S1, the lithium porcelain stone and red mud are first mixed and crushed, and then roasted.

3. The method according to claim 1 or 2, wherein: After the mixing and crushing, the particle sizes of the lithium porcelain stone and the red mud are both 150-300 meshes.

4. [Corrected 20.12.2023 in accordance with Rule 91] A method according to any one of claims 1 to 3, wherein: After the mixing and crushing, the particle sizes of the lithium porcelain stone and the red mud are both 200-250 meshes.

5. The method according to any one of claim 14, wherein: In step S1, the mass ratio of the lithium porcelain stone to the red mud is 1:(0.2-0.8); Optionally, in step S1, the mass ratio of the lithium porcelain stone to the red mud is 1:(0.4-0.6).

6. The method according to any one of claims 1 to 5, wherein: The temperature of the primary calcination in step S1 is 750-950°C; Optionally, the primary calcination temperature in step S1 is 800-850° C.; Optionally, the time of the first calcination in step S1 is 1 to 3 hours; Optionally, the time of the first calcination in step S1 is 1.5 to 2.5 hours.

7. The method according to any one of claims 1 to 6, wherein: The red mud in step S2 is red mud with a particle size of 150 to 300 mesh after crushing; Optionally, the red mud in step S2 has a particle size of 200-250 mesh after crushing.

8. The method according to any one of claims 1 to 7, wherein: The mass ratio of the primary roasting material to the red mud in step S2 is 1:(0.2-0.8); Optionally, in step S2, the mass ratio of the primary roasting material to the red mud is 1:(0.4-0.6).

9. The method according to any one of claims 1 to 8, wherein: The soaking liquid used in the soaking in step S2 includes water; Optionally, the mass ratio of the total mass of the primary roasted material and the red mud after mixing to water is (0.8-1.5):1; Optionally, the mass ratio of the total mass of the primary roasted material and the red mud after mixing to water is (0.8-1):

1.

10. The method according to any one of claims 1 to 9, wherein: The secondary calcination in step S3 is carried out under vacuum; Optionally, the secondary calcination temperature is 850-1000°C; Optionally, the secondary calcination temperature is 850-950°C; Optionally, the secondary roasting time is 2 to 5 hours; Optionally, the secondary calcination time is 3 to 4.5 hours.

11. The method according to any one of claims 1 to 10, wherein: The primary leaching in step S4 and the secondary leaching in step S6 are to mix the secondary roasted material and the leached residue with the leaching agent respectively, and after the leaching is completed, solid-liquid separation is performed; Optionally, the leaching agent includes any one of water, hydrochloric acid solution or sulfuric acid solution, or a combination of at least two thereof; Optionally, the leaching agent is a sulfuric acid solution; Optionally, when the leaching agent includes hydrochloric acid solution and / or sulfuric acid solution, the concentration of the leaching agent is 6 to 12 mol / L; Optionally, when the leaching agent includes hydrochloric acid solution and / or sulfuric acid solution, the concentration of the leaching agent is 8 to 10 mol / L; Optionally, in the primary leaching and the secondary leaching, the liquid-to-solid ratio of the leaching agent to the secondary roasted material and the leached residue is (3-5) mL:1 g respectively; Optionally, in the primary leaching and the secondary leaching, the liquid-to-solid ratio of the leaching agent to the secondary roasted material and the leaching residue is 3 mL:1 g respectively.

12. The method according to any one of claims 1 to 11, wherein: The method comprises the following steps: S1: mixing and crushing lithium porcelain stone and red mud in a mass ratio of 1:(0.4-0.6) to obtain a mixed crushed material, wherein the particle size of the lithium porcelain stone and the red mud is 200-250 mesh, and calcining the mixed crushed material at 800-850° C. for 1.5-2.5 hours to obtain a primary calcined material; S2: taking red mud with a particle size of 200-250 mesh after crushing, mixing the primary roasted material with the red mud in a mass ratio of 1:(0.4-0.6), and soaking them in water, controlling the mass ratio of the total mass of the primary roasted material and the red mud to water to be (0.8-1):1, to obtain a soaked material; S3: The soaked material is subjected to secondary roasting at 850-950° C. under vacuum for 3-4.5 hours to obtain a secondary roasted material; S4: mixing the secondary roasted material with a leaching agent, wherein the leaching agent includes water, or any one of a hydrochloric acid solution or a sulfuric acid solution with a concentration of 8 to 10 mol / L, or a combination of at least two of them, controlling the liquid-solid ratio of the leaching agent to the secondary roasted material to be (3 to 5) mL: 1 g, performing a primary leaching on the secondary roasted material, and obtaining a leaching solution and leaching residue after solid-liquid separation; S5: removing impurities from the leaching solution to obtain metallurgical slag and lithium liquid products; S6: mixing the leached residue with a leaching agent, wherein the leaching agent includes water, or any one of a hydrochloric acid solution or a sulfuric acid solution with a concentration of 8 to 10 mol / L, or a combination of at least two of them, controlling the liquid-solid ratio of the leaching agent to the leached residue to be (3 to 5) mL: 1 g, performing secondary leaching on the leached residue, and obtaining silicon slag and aluminum solution after solid-liquid separation; There is no particular order for steps S5 and S6.

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