A method for extracting alumina from an industrial solid waste red mud

Abstract

The application discloses a method for extracting alumina from industrial solid waste red mud. The method comprises the following steps: mixing the red mud with a composite additive composed of sodium carbonate, potassium sulfate and limestone, wet ball milling the mixture until the particle size of the mixture reaches D90≤74 μm, drying the obtained slurry until the water content is less than 5%, and sintering at 950-1050 ℃ to obtain sintered clinker; leaching and separating the sintered clinker in hot water at a liquid-solid ratio of 10:1-25:1 and at 65-95 ℃ to obtain an aluminum-containing filtrate; introducing CO2 into the filtrate to perform carbonization, and performing aging and solid-liquid separation to obtain aluminum hydroxide; and then performing washing, drying and calcination to obtain an alumina product. By introducing potassium sulfate as a fluxing agent, the transformation of inert aluminum mineral phases is promoted, the sintering temperature can be reduced by 50-100 ℃, the time can be shortened by 30-50%, the process energy consumption is low, the aluminum extraction efficiency is as high as 88% or more, and a brand-new path is provided for large-scale resource utilization of the red mud.

Description

Technical Field

[0001] This invention belongs to the field of industrial solid waste resource utilization and metallurgical chemical technology, specifically relating to a method for efficiently extracting alumina from industrial solid waste red mud using a composite additive sintering-water leaching-carbonization process. Background Technology

[0002] Red mud is a highly alkaline solid waste generated during the Bayer process of alumina production from bauxite. It is characterized by a high pH (typically 10–13), high aluminum content (10%–20% Al2O3), and trace amounts of heavy metals (such as Cr, V, and As). Large-scale stockpiling of red mud not only occupies significant land resources but also easily leads to environmental problems such as alkali leakage, dust dispersion, and heavy metal migration, posing potential risks to the ecological environment and safe production. As the world's largest alumina producer, China's annual red mud emissions reached approximately 115 million tons as of 2024, with historical stockpiles exceeding 1.6 billion tons. However, the current annual comprehensive utilization of red mud is less than 15 million tons, with a utilization rate of less than 15%. The long-term stockpiling of large amounts of red mud not only causes severe environmental pressure but also results in a huge waste of recoverable aluminum resources. The aluminum in red mud mainly exists in the form of chemically stable mineral phases such as aluminum-iron solid solutions and aluminosilicates, exhibiting low reactivity, which is one of the key factors restricting its efficient recovery and utilization.

[0003] Currently, the main technical routes for recovering aluminum from red mud include acid leaching, reduction roasting-acid leaching, and alkali sintering. While acid leaching can dissolve aluminum components from red mud to a certain extent, it introduces large amounts of impurities such as iron and silicon into the solution during the leaching process, resulting in low purity of the leachate and difficulties in subsequent separation. It also presents problems such as large volumes of waste acid and severe equipment corrosion. Alkali sintering has attracted attention due to its relatively simple process and the ability to recycle some reagents. However, existing alkali sintering processes still generally suffer from uneven mixing, incomplete sintering reactions, low aluminum leaching rates, and high sintering temperatures.

[0004] To address these issues, researchers have made numerous attempts. For example, Chinese patent CN111825113B discloses a process for recovering sodium oxide and alumina from Bayer process red mud and returning them to the Bayer cycle. However, this process involves multiple steps, including red mud slurry preparation, leaching, and membrane separation, resulting in a complex process, high energy consumption, and limited control over silicon impurities. In contrast, alkali sintering, which promotes the transformation of inert bauxite phases in red mud into soluble aluminate phases under high-temperature conditions, is considered an effective way to improve aluminum recovery. However, existing alkali sintering processes still have significant shortcomings. For instance, Chinese patent CN109250737B discloses a method for extracting alumina from Bayer process red mud using a high-alkali concentration solution. This method has stringent reaction conditions, is prone to side reactions, and has low solid-liquid separation efficiency. Chinese patent CN112410559B discloses a high-temperature alkaline roasting process that improves the extractability of aluminum by altering the mineral phase structure of red mud, but its aluminum leaching rate remains low, limiting the practical application of this process.

[0005] Existing research and engineering practice indicate that the uniformity of raw material mixing and the liquid phase formation behavior in the sintering system have a significant impact on the degree of reaction during alkali sintering. To improve sintering performance, existing technologies have attempted to introduce fluxes to lower the sintering temperature or promote mineral phase transformation. However, most methods still employ dry mixing, making it difficult to achieve uniform dispersion of red mud, alkali source, and flux at the microscale, resulting in incomplete sintering reactions and limited improvement in aluminum activation efficiency and selective dissolution. In particular, existing technologies mainly focus on optimizing the ratio of sodium or calcium salts, with few reports on the application of potassium salts (especially potassium sulfate) in this system. A long-standing technological preference in this field has led to the use of sodium salts as the default alkali metal additive, failing to recognize the potential changes in sintering behavior that potassium salt additives may induce. Therefore, there is an urgent need to develop a new additive system with good mixing uniformity and sufficient sintering reaction to overcome the technical bottlenecks of existing sodium-calcium sintering processes.

[0006] This invention is the first to systematically investigate how potassium sulfate can significantly improve the aluminum dissolution rate (peak value of 88.4%) in a red mud alkaline sintering system within a specific dosage range (3%~10%), indicating that the mechanism of action of potassium salts is unique and different from that of conventional sodium / calcium salt systems. Summary of the Invention

[0007] In view of the problems existing in the red mud alkaline sintering aluminum extraction process, such as insufficient sintering reaction, low aluminum activation degree, and low leaching rate, this invention aims to provide a method for extracting alumina from industrial solid waste red mud by sintering using a composite flux.

[0008] To achieve the above objectives, the present invention adopts the following technical solution: A method for extracting alumina from industrial solid waste red mud includes the following steps: (1) Mix red mud with composite additives, wet ball mill the resulting mixture to obtain a mixed slurry, and then dry it to obtain dry powder; (2) The obtained dry powder is sintered at high temperature to obtain sintered clinker; (3) The sintered clinker was treated with water immersion and filtered to obtain an aluminum-containing filtrate; (4) CO2 is introduced into the aluminum-containing filtrate to carry out a carbonization reaction, followed by aging treatment, and solid-liquid separation to obtain aluminum hydroxide precipitate; (5) The obtained aluminum hydroxide precipitate was washed, dried, calcined, cooled to room temperature and ball-milled to obtain α-Al2O3 product.

[0009] Further, the composite additive in step (1) includes sodium carbonate, potassium sulfate, and limestone; and requires that: Al2O3 in the red mud ≥ 27.0%; sodium carbonate purity ≥ 98.0% (industrial grade); potassium sulfate purity ≥ 98.0% (industrial grade); and limestone purity ≥ 90.0% (general industrial grade, applicable to chemical and building materials). The molar ratio of Na in sodium carbonate to Al in red mud is (1~3):1, the amount of limestone added is 3%~10% of the mass of red mud, and the amount of potassium sulfate added is 3%~10% of the mass of red mud.

[0010] Further, in step (1), the liquid-solid ratio of wet ball milling is preferably 1:1 to 5:1, and the ball milling time is 10 to 30 min, so that the red mud and composite additives are uniformly dispersed at the microscale to obtain a slurry with a particle size D90≤74 μm. The mixed slurry is dried at 60 to 80°C until the moisture content is less than 5% to obtain dry powder.

[0011] Further, in step (2), the dry powder is sintered at 950~1050℃ for 0.5~2 h, preferably 0.8~2 h, so that the inert aluminum phase in the red mud is converted into a soluble aluminate phase, and sintered clinker is obtained.

[0012] Further, in step (3), the sintered clinker is treated with water at 65-95℃ for 20-60 min at a liquid-solid ratio of 10:1~25:1 (preferably 15:1~25:1), and then filtered to obtain an aluminum-containing filtrate.

[0013] Further, in step (4), CO2 is introduced into the aluminum-containing filtrate to carry out a carbonization reaction, the endpoint pH is controlled to be 6.5~10 (preferably 8~10), and the filtrate is aged in hot water at 60~80℃ for 1~3h, and aluminum hydroxide precipitate is obtained by solid-liquid separation.

[0014] Further, in step (5), the obtained aluminum hydroxide precipitate is washed 2-4 times at a washing liquid-to-solid ratio of 2:1 to 5:1, a stirring rate of 150-300 r / min, until the washing liquid is neutral. The washed filter cake is dried to constant weight at 100-120℃ and then calcined at 1050-1200℃ (preferably 1080-1150℃) for 1-2 hours. The calcined product is cooled to room temperature and then ball-milled to achieve a particle size of D50≤45 μm, finally obtaining an α-Al2O3 product with a purity ≥98.0%.

[0015] Furthermore, the carbonized tail liquid generated after solid-liquid separation in step (4) and the washing liquid generated in step (5) can be combined and returned to the preceding system for recycling. During the recycling process, the sodium and potassium ion concentrations of the returned mixed solution are detected. Based on the actual detection results and the initial set values ​​of the system, the consumed sodium carbonate and potassium sulfate are dynamically replenished, and limestone is added simultaneously to maintain the reaction system within the set molar ratio and mass ratio range of step (1). Preferably, in a single cycle, the amount of sodium carbonate replenished is preferably 40% to 60% of the initial addition amount, the amount of potassium sulfate replenished is 85% to 100% of the initial addition amount, and the amount of limestone replenished is 95% to 100% of the initial addition amount.

[0016] The present invention has the following beneficial effects: 1. This invention introduces a composite flux (especially potassium sulfate) into the red mud alkaline sintering system, which is beneficial to improving the formation of the liquid phase in the sintering system and promoting the transformation of the inert bauxite phase in red mud into the soluble aluminate phase. Experiments show that, with equal mass addition, the leaching effect of potassium sulfate (approximately 88.4%) is significantly better than that of traditional sodium sulfate (approximately 80.9%); at the same time, the introduction of potassium sulfate does not change the main nepheline crystal phase, but through K... + Interference with the ordered growth of the crystal lattice reduces the crystallinity of the product, making it more prone to decomposition during water immersion.

[0017] 2. This invention employs a wet ball milling process, which enables the composite additive and red mud to achieve highly uniform dispersion at the microscale. Combined with the phase characteristics formed by potassium sulfate, the system can achieve excellent activation effects at a relatively low sintering temperature (950-1050℃) and in a short time, effectively reducing energy consumption in industrial applications.

[0018] 3. The α-Al₂O₃ product obtained by the method of this invention can achieve a purity of over 98.0%, fully meeting the quality standards for metallurgical-grade alumina. Simultaneously, the alkaline tailings generated during the carbonization and washing processes are directly returned to the preceding processes for recycling, achieving zero secondary pollution emissions and excellent separation effect. This provides a practical and feasible technical solution for the reduction and high-value utilization of red mud. Attached Figure Description

[0019] Figure 1 The present invention provides a process flow diagram for the preparation of a method for extracting alumina from industrial solid waste red mud.

[0020] Figure 2 A comparative graph showing the effects of different additives on the aluminum leaching rate of red mud sintered clinker.

[0021] Figure 3 A comparative graph showing the effect of potassium sulfate addition on the aluminum leaching rate of red mud sintered clinker.

[0022] Figure 4 The XRD patterns are comparisons of the sintered products from Example 1 (5% potassium sulfate) and Example 2 (no potassium sulfate). Detailed Implementation

[0023] The technical solution of the present invention will be further described in detail below with reference to specific implementation cases and accompanying drawings, but the scope of protection and implementation of the present invention are not limited thereto.

[0024] The red mud used in the embodiments and comparative examples of this invention was pretreated (dried, pulverized, and sieved to a particle size of less than 0.075 mm). Its main components by mass percentage were: Al₂O₃ 27.8%, SiO₂ 21.6%, and Fe₂O₃ 14.9%. Reagents used included: sodium carbonate (industrial grade, purity ≥98.0%), potassium sulfate (industrial grade, purity ≥98.0%), sodium sulfate (industrial grade, purity ≥98.0%), and limestone (ordinary industrial grade, purity ≥90.0%). The aluminum leaching rate described in this invention refers to the percentage of aluminum (converted to alumina) detected in the filtrate after water leaching, relative to the total mass of Al₂O₃ in the original red mud sample. The aluminum precipitation recovery rate described in this invention refers to the percentage of aluminum mass in the carbonized precipitated aluminum hydroxide relative to the total aluminum mass in the water leaching filtrate.

[0025] Figure 1 This invention provides a process flow diagram for extracting alumina from industrial solid waste red mud, illustrating the main materials and processes. The process includes: mixing red mud with a composite additive composed of sodium carbonate, potassium sulfate, and limestone; and then performing a series of processes such as ball milling, drying, sintering, sieving, water leaching and filtration, CO2 carbonization, aging, filtration, washing, drying, and calcination to obtain α-Al₂O₃ with a purity exceeding 98%. Furthermore, the carbonization tail liquid and washing liquid produced can be combined and recycled back to the preceding system.

[0026] The following examples and comparative examples are in accordance with Figure 1 Follow the steps shown: Example 1 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and 5% of the red mud mass were added to the red mud. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid components. Analysis of the filtrate showed an aluminum leaching rate of 88.4%.

[0027] CO2 gas was uniformly introduced into the filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed with deionized water at a solid-liquid ratio of 3:1 and a stirring rate of 200 r / min, for a total of three washes until the washing liquid was neutral. The washed filter residue was dried in a 105℃ oven to constant weight, then calcined again in a muffle furnace at 1100℃ for 1.5 hours. After cooling, it was ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. Testing showed that the product had a purity of 98.6%, meeting the quality standards for metallurgical-grade alumina.

[0028] In this embodiment, the aluminum precipitation recovery rate was 92.8%, and the total recovery rate reached 82.1%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and sodium carbonate (approximately 45% of the initial addition), potassium sulfate (approximately 92% of the initial addition), and limestone (approximately 98% of the initial addition) were added to maintain the equilibrium of the reaction system.

[0029] Example 2 (without potassium sulfate) 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate and 2.4 g of limestone were added to it, without adding potassium sulfate. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid. Analysis of the filtrate showed an aluminum leaching rate of 73.7%. A comparison of Example 1 and Example 2 shows that the introduction of potassium sulfate increased the leaching rate by approximately 14.7 percentage points, demonstrating a significant technical effect.

[0030] CO2 gas was uniformly introduced into the obtained filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed with deionized water at a solid-liquid ratio of 3:1 and a stirring rate of 200 r / min, for a total of three washes until the washing liquid was neutral. The washed filter residue was dried in a 105℃ oven to constant weight, then calcined again in a muffle furnace at 1100℃ for 1.5 hours. After cooling, it was ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. The product purity was tested to be 98.5%. In this embodiment, the aluminum precipitation recovery rate was 87.2%, and the total recovery rate was 64.3%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and sodium carbonate (approximately 48% of the initial addition) and limestone (approximately 98% of the initial addition) were added to maintain the equilibrium of the reaction system.

[0031] Example 3 (replaced with sodium sulfate) 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and sodium sulfate (5% of the red mud mass) were added to the red mud. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid phases. Analysis of the filtrate showed an aluminum leaching rate of 80.9%. Comparison of Example 3 and Example 1 shows that, with equal mass additions, the leaching effect of potassium sulfate (approximately 88.4%) was significantly better than that of sodium sulfate (approximately 80.9%).

[0032] CO2 gas was uniformly introduced into the obtained filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed with deionized water at a solid-liquid ratio of 3:1 and a stirring rate of 200 r / min, for a total of three washes until the washing liquid was neutral. The washed filter residue was dried in a 105℃ oven to constant weight, then calcined a second time in a muffle furnace at 1100℃ for 1.5 hours. After cooling, it was ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. The product purity was tested to be 98.0%.

[0033] In this embodiment, the aluminum precipitation recovery rate was 83.4%, and the total recovery rate was 67.5%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and the consumed sodium carbonate (approximately 40% of the initial addition) and sodium sulfate (approximately 88% of the initial addition) were replenished, and limestone was added in full.

[0034] Example 4 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and 5% of the red mud mass were added to the red mud. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 950°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-volume, multiple-stage filtration method to separate the solid and liquid components. Analysis of the filtrate showed an aluminum leaching rate of 85.4%.

[0035] CO2 gas was uniformly introduced into the obtained filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed with deionized water at a liquid-to-solid ratio of 4:1 and a stirring rate of 250 r / min, for a total of two washes until the washing liquid was neutral. The filter residue was dried at 105℃ to constant weight, calcined twice at 1100℃ for 1.5 hours, cooled, and then ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. The purity was tested and found to be 98.5%.

[0036] In this embodiment, the aluminum precipitation recovery rate was 90.5%, and the total recovery rate was 77.3%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and sodium carbonate (approximately 40% of the initial addition) and potassium sulfate (approximately 85% of the initial addition) were added, while limestone (approximately 98% of the initial addition) was added in full.

[0037] Example 5 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and 5% of the red mud mass were added to the red mud. The mixture was ball-milled at a liquid-to-solid ratio of 3:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid phases. Analysis of the filtrate showed an aluminum leaching rate of 77.1%.

[0038] CO2 gas was uniformly introduced into the obtained filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed with deionized water at a liquid-to-solid ratio of 3:1 and a stirring rate of 200 r / min, for a total of three washes until the washing liquid was neutral. The filter residue was dried to constant weight at 105℃, calcined twice at 1100℃ for 1.5 hours, cooled, and then ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. The purity was tested and found to be 98.5%.

[0039] In this embodiment, the aluminum precipitation recovery rate was 89.9%, and the total recovery rate was 69.3%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and sodium carbonate (approximately 55% of the initial addition) and potassium sulfate (approximately 96% of the initial addition) were added, and limestone was added in full.

[0040] Example 6 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and 5% of the red mud mass were added to the red mud. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1050°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid phases. Analysis of the filtrate showed an aluminum leaching rate of 85.2%.

[0041] CO2 gas was uniformly introduced into the obtained filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed multiple times with deionized water, maintaining a washing liquid-to-solid ratio of 3:1 and a stirring rate of 200 r / min, for a total of three washes until the washing liquid was neutral. The washed filter residue was dried in a 105℃ oven to constant weight, then calcined again in a muffle furnace at 1100℃ for 1.5 hours. After cooling, it was ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. The purity was tested and found to be 98.0%.

[0042] In this embodiment, the aluminum precipitation recovery rate was 92.7%, and the total recovery rate was 79.0%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and sodium carbonate (approximately 42% of the initial addition) and potassium sulfate (approximately 88% of the initial addition) were added, and limestone was added in full.

[0043] Example 7 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and potassium sulfate (3% of the red mud mass) were added to it. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid phases. Analysis of the filtrate showed an aluminum leaching rate of 78.9%.

[0044] CO2 gas was uniformly introduced into the obtained filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed multiple times with deionized water at a liquid-to-solid ratio of 3:1 and a stirring rate of 200 r / min until neutral. The filter residue was dried in a 105℃ oven and then calcined again at 1100℃ for 1.5 hours. After cooling, it was ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. The purity was tested and found to be 98.5%.

[0045] In this embodiment, the aluminum precipitation recovery rate was 90.2%, and the total recovery rate was 71.2%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and sodium carbonate (approximately 46% of the initial addition) and potassium sulfate (approximately 90% of the initial addition) were added, and limestone was added in full.

[0046] Example 8 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and potassium sulfate (7.5% of the red mud mass) were added to the red mud. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid phases. Analysis of the filtrate showed an aluminum leaching rate of 85.2%.

[0047] CO2 gas was uniformly introduced into the obtained filtrate to initiate a carbonization reaction. The solution was aged at 60℃ for 2 hours after reaching a final pH of 9.0, followed by filtration. The filter cake was washed multiple times with deionized water at a liquid-to-solid ratio of 3:1 and a stirring rate of 200 r / min until neutral. The filter residue was dried in a 105℃ oven and then calcined again at 1100℃ for 1.5 hours. After cooling, it was ball-milled to a particle size D50 ≤ 45 μm to obtain the α-Al2O3 product. The purity was tested and found to be 98.6%.

[0048] In this embodiment, the aluminum precipitation recovery rate was 89.8%, and the total recovery rate was 76.5%. The aluminum extraction tail liquid and washing liquid were mixed and returned to the system for circulation. Before circulation, the loss was detected and calculated, and sodium carbonate (approximately 48%) and potassium sulfate (approximately 90%) were added, with limestone added in full.

[0049] Comparative Example 9 (the amount of potassium sulfate used is outside the scope of this invention) 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and potassium sulfate (1.5% of the red mud mass) were added to the red mud. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid phases. Analysis of the filtrate showed an aluminum leaching rate of 74.1%. Subsequent carbonization, washing, and calcination steps were the same as in Example 1. The product purity was tested to be 97.7%. In this example, the aluminum precipitation recovery rate was 86.4%, and the total recovery rate was 59.2%.

[0050] Comparative Example 10 (the amount of potassium sulfate used is outside the scope of this invention) 20.0 g of pretreated red mud was taken, and 13.6 g of sodium carbonate, 2.4 g of limestone, and potassium sulfate were added to the red mud, with the amount added being 12.0% of the red mud mass. The mixture was ball-milled at a liquid-to-solid ratio of 5:1 to obtain a slurry. The slurry was dried at 65°C for 12 h and then sintered in a muffle furnace at 1000°C for 1 h to obtain sintered clinker. The sintered clinker was leached in 80°C hot water at a liquid-to-solid ratio of 25:1 for 30 min. The filter cake was then washed using a small-batch, multiple-stage filtration method to separate the solid and liquid phases. Analysis of the filtrate showed an aluminum leaching rate of 75.1%. Subsequent carbonization, washing, and calcination steps were the same as in Example 1. The product purity was tested to be 98.0%. In this example, the aluminum precipitation recovery rate was 91.0%, and the total recovery rate was 68.3%.

[0051] A comparative analysis of the above embodiments and comparative examples shows that, Figure 2 The invention clearly demonstrates that, with equal mass addition, the leaching effect of potassium sulfate (approximately 88.4%) is significantly better than that of traditional sodium sulfate (approximately 80.9%), which is the core technological breakthrough brought about by the present invention.

[0052] like Figure 3Based on Examples 7 (3%, 78.9%), 1 (5%, 88.4%), 8 (7.5%, 85.2%), 9 (1.5%, 74.1%), and 10 (12%, 75.1%), it can be seen that the aluminum leaching rate exhibits a non-monotonic trend of first increasing and then decreasing with the amount of potassium sulfate added, reaching a peak around 5%. When the addition amount is below 3%, the leaching rate is basically the same as the system without additives, failing to reflect the beneficial effect of potassium sulfate; above 10%, the leaching rate actually decreases, further indicating that the mechanism of action of potassium sulfate is not a simple fluxing effect, but involves specific eutectic control.

[0053] To further investigate the role of potassium sulfate in the sintering process, X-ray diffraction (XRD) analysis was performed on the sintering products of Example 2 (no additives) and Example 1 (5% potassium sulfate). The results are as follows: Figure 4 As shown, the main diffraction peaks all belong to nepheline aluminosilicates (NaAlSiO4 type structure), with a small amount of hematite (Fe2O3) characteristic peaks also present. The diffraction peak positions of the two samples are basically overlapping (both main peaks are located at 2θ=33.71°), indicating that the introduction of potassium sulfate did not change the phase composition of the sintered product. Taking the strongest diffraction peak (2θ=33.71°) as an example, the intensity of Example 2 (without potassium sulfate) is 990 counts, while the intensity of Example 1 (potassium sulfate) is 942 counts, a decrease of about 4.8%. The decrease in diffraction peak intensity may be due to a decrease in the content of the main crystalline phase or a decrease in crystallinity. Combined with the leaching rate data analysis, if the decrease in peak intensity is due to a decrease in the content of the main crystalline phase, it means that the aluminum component transformed into the nepheline phase is reduced, and the leaching rate should decrease accordingly; however, the aluminum leaching rate of Example 1 (88.4%) is significantly higher than that of Example 2 (73.7%), thus ruling out the influence of content factors. The above results indicate that although the eutectic liquid phase formed by potassium sulfate at high temperatures promotes mass transfer and the formation of the aluminate phase, the K in the liquid phase... + The presence of potassium sulfate interferes with the orderly growth of nepheline crystals, leading to decreased crystallinity and increased lattice defects in the product. This reduced crystallinity makes the nepheline phase more easily decomposed during water leaching, resulting in a more complete release of aluminate ions and a significant increase in leaching rate. This mechanism of action, which "does not change the phase composition but only regulates the crystallization state," reflects the unique nature of potassium sulfate's mechanism of action.

[0054] The above description is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any way. Any equivalent changes, modifications, or variations made by those skilled in the art to the above embodiments using the technical solutions of the present invention shall still fall within the scope of the technical solutions of the present invention.

Claims

1. A process for the extraction of alumina from an industrial solid waste red mud, characterised in that, Includes the following steps: (1) Mix red mud with composite additives, wet ball mill the resulting mixture to obtain a mixed slurry, and dry it to obtain dry powder; (2) The obtained dry powder is sintered at high temperature to obtain sintered clinker; (3) The sintered clinker was treated with water immersion and filtered to obtain an aluminum-containing filtrate; (4) CO2 is introduced into the aluminum-containing filtrate to carry out a carbonization reaction, followed by aging treatment, and solid-liquid separation to obtain aluminum hydroxide precipitate; (5) The obtained aluminum hydroxide precipitate was washed, dried, calcined, cooled to room temperature and ball-milled to obtain α-Al2O3 product.

2. The method of claim 1, wherein, The composite additives mentioned in step (1) include sodium carbonate, potassium sulfate, and limestone; and require that: Al2O3 in the red mud ≥ 27.0%; sodium carbonate purity ≥ 98.0%; potassium sulfate purity ≥ 98.0%; and limestone purity ≥ 90.0%. The molar ratio of Na in sodium carbonate to Al in red mud is (1~3):1, the amount of limestone added is 3%~10% of the mass of red mud, and the amount of potassium sulfate added is 3%~10% of the mass of red mud.

3. The method of claim 1, wherein, In step (1), the liquid-solid ratio of wet ball milling is 1:1 to 5:1, and the ball milling time is 10 to 30 min, so that the red mud and composite additives are uniformly dispersed at the microscale to obtain a slurry with a particle size D90≤74 μm. The mixed slurry is dried at 60 to 80°C until the moisture content is less than 5% to obtain dry powder.

4. The method of claim 1, wherein, In step (2), the dry powder is sintered at 950~1050℃ for 0.5~2 h to obtain sintered clinker.

5. The method of claim 1, wherein, In step (3), the sintered clinker is treated with water at 65-95℃ for 20-60 min at a liquid-solid ratio of 10:1 to 25:1, and then filtered to obtain an aluminum-containing filtrate.

6. The method of claim 1, wherein, In step (4), CO2 is introduced into the aluminum-containing filtrate to carry out a carbonization reaction, the final pH is controlled to be 6.5~10, and the filtrate is aged at 60~80℃ for 1~3h, and aluminum hydroxide precipitate is obtained by solid-liquid separation.

7. The method of claim 1, wherein, In step (5), the obtained aluminum hydroxide precipitate is washed 2 to 4 times at a washing liquid-solid ratio of 2:1 to 5:1, a stirring rate of 150 to 300 r / min, until the washing liquid is neutral.

8. The method of claim 1, wherein, In step (5), the washed filter cake is dried and then calcined at 1050~1200℃ for 1~2 h.

9. The method of claim 1, wherein, In step (5), the calcined product is cooled to room temperature and then ball-milled to achieve a particle size of D50≤45 μm, ultimately obtaining an α-Al2O3 product with a purity ≥98.0%.

10. The method according to any one of claims 1-9, characterized in that, The carbonized tail liquid generated after solid-liquid separation in step (4) and the washing liquid generated in step (5) can be combined and returned to the preceding system for recycling.

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