Method for efficiently extracting aluminum oxide from low-grade aluminum ore

Abstract

The invention provides a method for efficiently extracting aluminum oxide from low-grade aluminum ore, and belongs to the technical field of metallurgical engineering. The method provided by the invention comprises the following steps: crushing and grinding the low-grade aluminum ore in sequence to obtain pretreated aluminum ore; the pretreated aluminum ore is mixed with water, first alkali and a collecting agent, reverse flotation is conducted, and aluminum concentrate is obtained; the aluminum concentrate is subjected to pressurized acid leaching, and a leaching solution is obtained; mixing the leachate with alkali, and carrying out first iron precipitation and second aluminum precipitation step by step to obtain aluminum hydroxide; mixing aluminum hydroxide with a second acid solution to obtain an aluminum salt solution; concentrating the aluminum salt solution, adding a seed crystal, and crystallizing to obtain aluminum salt; and calcining the aluminum salt to obtain the aluminum oxide. According to the method, the aluminum-silicon ratio of the aluminum concentrate is increased through reverse flotation pre-enrichment, the acid consumption in the subsequent pressurized acid leaching process is reduced, and the aluminum oxide recovery cost is reduced; through fractional precipitation, efficient separation of aluminum and iron is achieved, and the purity of final aluminum oxide is improved.

Description

Technical Field

[0001] This invention belongs to the field of metallurgical engineering technology, specifically relating to a method for efficiently extracting alumina from low-grade bauxite. Background Technology

[0002] Currently, the main industrial methods for producing alumina include alkaline processes (such as the Bayer process, sintering process, Bayer-sintering combined process, high-pressure water chemistry process, etc.), acid processes (such as hydrochloric acid process, sulfuric acid process, nitric acid process, etc.), and combined acid-alkali processes. The Bayer process is suitable for processing high-quality bauxite and has a relatively simple process. However, for low-grade bauxite, especially high-silicon low-grade bauxite, this method suffers from high alkali consumption and large amounts of waste due to the high silicon content in the ore. While the sintering process is suitable for processing high-silicon aluminum-containing materials, achieving high alumina recovery rates and high product whiteness, the process is relatively complex. Acid processes face challenges in separating aluminum and iron when processing aluminum-containing materials. my country suffers from a severe shortage of aluminum resources, and a large amount of low-grade bauxite remains unutilized. Therefore, developing a method for efficiently extracting alumina from low-grade bauxite is of significant practical importance. Summary of the Invention

[0003] The purpose of this invention is to provide a method for efficiently extracting alumina from low-grade bauxite. The method provided by this invention efficiently extracts high-quality alumina from low-grade bauxite.

[0004] To achieve the above-mentioned objectives, the present invention provides the following technical solution: This invention provides a method for efficiently extracting alumina from low-grade bauxite, comprising the following steps: (1) The low-grade aluminum ore is crushed and ground in sequence to obtain pretreated aluminum ore; (2) The pretreated aluminum ore obtained in step (1) is mixed with water and a first alkali to obtain a slurry; the slurry is mixed with a collector and subjected to reverse flotation to obtain aluminum concentrate; (3) The aluminum concentrate obtained in step (2) is mixed with the first acid solution and subjected to pressure acid leaching to obtain a leachate; (4) The leachate obtained in step (3) is mixed with the second alkali to precipitate iron for the first time, and the filtrate is obtained; the filtrate is mixed with the third alkali to precipitate aluminum for the second time, and aluminum hydroxide is obtained. (5) The aluminum hydroxide obtained in step (4) is mixed with the second acid solution to obtain an aluminum salt solution; the aluminum salt solution is concentrated and seed crystals are added to crystallize and obtain aluminum salt; the aluminum salt is calcined to obtain aluminum oxide.

[0005] Preferably, the particle size of the pretreated aluminum ore in step (1) is ≥200 mesh.

[0006] Preferably, the mass concentration of the slurry in step (2) is 20-30%; and the pH value of the slurry is 8-9.

[0007] Preferably, the collector in step (2) includes fatty acids and sulfonates; the mass ratio of the fatty acids and sulfonates is (2~4):1; the mass ratio of the collector to the slurry is (0.05~0.2):1000.

[0008] Preferably, the reverse flotation in step (2) is performed 2 to 4 times.

[0009] Preferably, the first acid solution in step (3) is hydrochloric acid; the mass concentration of the first acid solution is 15~25%; and the volume ratio of the first acid solution to the mass of aluminum concentrate is (6~8) mL:1g.

[0010] Preferably, in step (3), the pressure of the acid leaching is 0.35~0.4MPa, the temperature of the acid leaching is 100~120℃, and the time of the acid leaching is 1.5~2h.

[0011] Preferably, in step (4), the pH value of the first precipitated iron is 1~2; and the pH value of the second precipitated aluminum is 4~5.

[0012] Preferably, the second acid solution in step (5) is hydrochloric acid; the mass concentration of the second acid solution is 10~20%; and the mass ratio of the aluminum hydroxide to the second acid solution is 1:(8~10).

[0013] Preferably, the calcination in step (5) includes a first calcination, a second calcination, and a third calcination performed sequentially; the temperature of the first calcination is 200~390℃, and the time of the first calcination is 10~15min; the temperature of the second calcination is 400~600℃, and the time of the second calcination is 30~50min; the temperature of the third calcination is 800~1000℃, and the time of the third calcination is 30~70min.

[0014] This invention provides a method for efficiently extracting alumina from low-grade bauxite, comprising the following steps: (1) crushing and grinding the low-grade bauxite sequentially to obtain pretreated bauxite; (2) mixing the pretreated bauxite obtained in step (1) with water and a first alkali to obtain a slurry; mixing the slurry with a collector and performing reverse flotation to obtain alumina concentrate; (3) mixing the alumina concentrate obtained in step (2) with a first acid solution and performing pressurized acid leaching to obtain a leachate; (4) mixing the leachate obtained in step (3) with a second alkali and performing a first precipitation of iron to obtain a filtrate; mixing the filtrate with a third alkali and performing a second precipitation of aluminum to obtain aluminum hydroxide; (5) mixing the aluminum hydroxide obtained in step (4) with a second acid solution to obtain an aluminum salt solution; concentrating the aluminum salt solution and adding seed crystals to perform crystallization to obtain an aluminum salt; calcining the aluminum salt to obtain alumina. This invention improves the aluminum-silicon ratio of aluminum concentrate through reverse flotation pre-enrichment, reduces acid consumption in the subsequent pressurized acid leaching process, and lowers alumina recovery costs; through stepwise precipitation, it achieves efficient separation of aluminum and iron, and improves the purity of the final alumina. Detailed Implementation

[0015] This invention provides a method for efficiently extracting alumina from low-grade bauxite, comprising the following steps: (1) The low-grade aluminum ore is crushed and ground in sequence to obtain pretreated aluminum ore; (2) The pretreated aluminum ore obtained in step (1) is mixed with water and a first alkali to obtain a slurry; the slurry is mixed with a collector and subjected to reverse flotation to obtain aluminum concentrate; (3) The aluminum concentrate obtained in step (2) is mixed with the first acid solution and subjected to pressure acid leaching to obtain a leachate; (4) The leachate obtained in step (3) is mixed with the second alkali to precipitate iron for the first time, and the filtrate is obtained; the filtrate is mixed with the third alkali to precipitate aluminum for the second time, and aluminum hydroxide is obtained. (5) The aluminum hydroxide obtained in step (4) is mixed with the second acid solution to obtain an aluminum salt solution; the aluminum salt solution is concentrated and seed crystals are added to crystallize and obtain aluminum salt; the aluminum salt is calcined to obtain aluminum oxide.

[0016] Unless otherwise specified, the present invention does not impose any special restrictions on the source of the raw materials, and commercially available products well known to those skilled in the art can be used.

[0017] This invention involves sequentially crushing and grinding low-grade aluminum ore to obtain pretreated aluminum ore.

[0018] The present invention does not have any special limitations on the source and specific composition of the low-grade aluminum ore, and any low-grade aluminum ore well known to those skilled in the art can be used.

[0019] In this invention, the particle size of the pretreated aluminum ore is preferably ≥200 mesh.

[0020] The present invention does not impose any special limitations on the crushing and grinding operations. As long as the pre-treated aluminum ore particle size is within the required range, a technical solution known to those skilled in the art can be used.

[0021] After obtaining the pretreated aluminum ore, the present invention mixes the pretreated aluminum ore with water and a first alkali to obtain a slurry; the slurry is then mixed with a collector and subjected to reverse flotation to obtain aluminum concentrate.

[0022] In this invention, the first alkali is preferably sodium carbonate or sodium hydroxide.

[0023] In this invention, the mass concentration of the slurry is preferably 20-30%, more preferably 25%; the pH value of the slurry is preferably 8-9. This invention does not impose any special limitations on the amount of water and the first alkali used, as long as the mass concentration and pH value of the slurry are within the required range.

[0024] In this invention, the collector preferably comprises fatty acids and sulfonates; the fatty acids are preferably oleic acid; the sulfonates are preferably sodium petroleum sulfonate; the mass ratio of the fatty acids to the sulfonates is preferably (2~4):1, more preferably 3:1; the mass ratio of the collector to the slurry is preferably (0.05~0.2):1000.

[0025] In this invention, the reverse flotation is preferably performed 2 to 4 times, more preferably 3 times; the temperature of each reverse flotation is preferably 10 to 20°C; and the time of each reverse flotation is preferably 15 to 20 minutes.

[0026] In one implementation, when the reverse flotation is performed three times, the present invention adjusts the pulp concentration after the first reverse flotation to 20-30%, the pH value to 8-9, and adds 90-150g / ton of collector before performing the second reverse flotation; after the second reverse flotation, the pulp concentration is adjusted to 20-30%, the pH value to 8-9, and adds 50-100g / ton of collector before performing the third reverse flotation.

[0027] In this invention, air is preferably introduced from the bottom of the slurry during the reverse flotation process. As one embodiment, the air pressure can be 0.25 MPa and the flow rate can be 0.25 m³ / s. 3 / h.

[0028] By performing reverse flotation and controlling the parameters of reverse flotation within the above-mentioned range, the present invention can further enrich and increase the aluminum-silicon ratio of aluminum concentrate, and further reduce the acid consumption of subsequent pressurized acid leaching.

[0029] After obtaining aluminum concentrate, the present invention mixes the aluminum concentrate with a first acid solution and performs pressurized acid leaching to obtain a leachate.

[0030] In this invention, the first acid solution is preferably hydrochloric acid; the mass concentration of the first acid solution is preferably 15-25%, more preferably 20%; the volume ratio of the first acid solution to the mass ratio of aluminum concentrate is preferably (6-8) mL:1g, more preferably 7 mL:1g.

[0031] In this invention, the pressure of the pressurized acid leaching is preferably 0.35~0.4MPa; the temperature of the pressurized acid leaching is preferably 100~120℃, more preferably 110℃; and the time of the pressurized acid leaching is preferably 1.5~2h.

[0032] This invention controls the parameters of pressurized acid leaching within the above-mentioned range, which enables the elements in the aluminum concentrate to be fully dissolved.

[0033] After pressurized acid leaching, the present invention preferably filters the system to obtain a leachate and a leachate residue. In the present invention, the pH value of the leachate is preferably 0.1 to 0.4.

[0034] The present invention does not impose any special limitations on the filtering operation; any technical solution known to those skilled in the art can be used.

[0035] After obtaining the leachate, the present invention mixes the leachate with a second alkali to precipitate iron in the first stage, and obtains a filtrate; the filtrate is then mixed with a third alkali to precipitate aluminum in the second stage, and obtains aluminum hydroxide.

[0036] In this invention, the second alkali is preferably sodium carbonate.

[0037] In this invention, the pH value of the first precipitated iron is preferably 1 to 2, more preferably 1.5. This invention does not have a specific limitation on the amount of the second alkali used, as long as the pH value of the first precipitated iron is within the above range.

[0038] In this invention, the temperature of the first precipitated iron is preferably 15~25℃, more preferably 20℃; the precipitation time of the first precipitated iron is preferably 40~50min, more preferably 45min. By controlling the pH value, temperature, and time of the first precipitated iron within the above ranges, this invention enables iron impurities to precipitate sufficiently, which is more conducive to obtaining high-purity alumina subsequently.

[0039] After the first precipitation of iron is completed, the present invention preferably filters the system after the first precipitation of iron to obtain filtrate and iron slag.

[0040] After obtaining the iron slag, the present invention preferably filters and washes the iron slag twice with distilled water, and then heats and decomposes it at 280~320℃ until constant weight is achieved to obtain Fe2O3. Fe2O3 is a raw material for producing high-purity ultrafine (600~800 mesh) iron powder. The iron powder obtained can be used to manufacture high-strength, high-precision mechanical parts, such as connecting rods, crankshafts, clutch plates, gears and other complex-shaped parts.

[0041] In this invention, the third alkali is preferably sodium carbonate.

[0042] In this invention, the pH value of the second precipitated aluminum is preferably 4-5, more preferably 4.5. This invention does not have a specific limitation on the amount of the third alkali used, as long as the pH value of the second precipitated aluminum is within the required range.

[0043] In this invention, the temperature of the second precipitated aluminum is preferably 38~42℃, more preferably 40℃; the precipitation time is preferably 80~100min, more preferably 90min. By controlling the pH value, temperature, and time of the second precipitated aluminum within the above ranges, this invention ensures sufficient aluminum precipitation.

[0044] After the second aluminum precipitation is completed, the present invention preferably filters the system after the second aluminum precipitation to obtain aluminum hydroxide.

[0045] The present invention does not impose any special limitations on the filtering operation; any technical solution known to those skilled in the art can be used.

[0046] After obtaining aluminum hydroxide, the present invention mixes the aluminum hydroxide with a second acid solution to obtain an aluminum salt solution; after concentrating the aluminum salt solution, seed crystals are added to crystallize and obtain an aluminum salt; the aluminum salt is calcined to obtain aluminum oxide.

[0047] In this invention, the second acid solution is preferably hydrochloric acid; the mass concentration of the second acid solution is preferably 10-20%, more preferably 15%; the mass ratio of the aluminum hydroxide to the second acid solution is preferably 1:(8-10), more preferably 1:9.

[0048] In this invention, the mixing temperature of the aluminum hydroxide and the second acid solution is preferably 80-85°C; the mixing time is preferably 45-60 minutes. By controlling the mixing temperature and time within the above ranges, this invention enables the aluminum hydroxide to fully dissolve and form an aluminum chloride solution.

[0049] In this invention, the concentration temperature is preferably 35~45℃, more preferably 40℃. This invention does not have a specific limitation on the concentration time; concentration can be achieved until supersaturation.

[0050] In this invention, the seed crystal is preferably an aluminum chloride hexahydrate seed crystal; the mass of the seed crystal is preferably 0.4~0.6% of the mass of the aluminum salt solution, more preferably 0.5%.

[0051] The present invention does not impose any special limitations on the crystallization operation; any technical solution known to those skilled in the art can be used to fully crystallize and precipitate the aluminum salt.

[0052] After crystallization is completed, the present invention preferably filters and washes the crystallized product sequentially to obtain aluminum salt.

[0053] The present invention does not impose any special limitations on the filtering operation; any technical solution known to those skilled in the art can be used.

[0054] In this invention, the water used for washing is preferably at room temperature. This invention does not impose any special limitations on other operations of the washing process; the goal is simply to remove impurities from the aluminum salt surface.

[0055] In this invention, the calcination preferably includes a first calcination, a second calcination, and a third calcination performed sequentially; the temperature of the first calcination is preferably 200~390℃; the time of the first calcination is preferably 10~15 min; the temperature of the second calcination is preferably 400~600℃; the time of the second calcination is preferably 30~50 min; the temperature of the third calcination is preferably 800~1000℃; and the time of the third calcination is preferably 30~70 min. As one embodiment, the temperature of the first calcination can specifically be 200℃, 220℃, 250℃, 280℃, 300℃, 320℃, 350℃, 380℃, or 390℃; the time of the first calcination can specifically be 10 min, 11 min, 12 min, 13 min, 14 min, or 15 min; and the temperature of the second calcination can specifically be 400℃, 420℃, 450℃, 480℃, 500℃, 520℃, 550℃, or 5... The first calcination temperature can be 80℃ or 600℃; the second calcination time can be specifically 30 min, 35 min, 40 min, 45 min, or 50 min; the third calcination temperature can be specifically 800℃, 820℃, 850℃, 880℃, 900℃, 920℃, 950℃, 980℃, or 1000℃; the third calcination time can be specifically 30 min, 40 min, 50 min, 55 min, 60 min, 65 min, or 70 min. By controlling the calcination temperature and time within the above ranges, this invention can further improve the purity of alumina.

[0056] In this invention, the hydrogen chloride gas generated during the calcination process is preferably absorbed and recovered using water through a packed tower or a spray tower. This invention does not impose any particular limitations on the operation of absorbing the gas using water through a packed tower or a spray tower; any technical solution well-known to those skilled in the art can be used.

[0057] This invention offers the following advantages: 1. Significant economic benefits: By increasing the aluminum-silicon ratio of the aluminum concentrate through reverse flotation pre-enrichment, the acid consumption in the subsequent acid leaching process is reduced, resulting in a 32% reduction in alumina recovery costs (compared to the traditional Bayer process for the same grade ore). Resource utilization of iron slag and leaching residue allows for product value-added revenue covering 40% of the processing costs. 2. Good environmental benefits: The comprehensive utilization rate of waste residue is >98%, reducing solid waste emissions. The hydrochloric acid recycling rate is ≥85%, reducing acid consumption and environmental pollution. Energy consumption is 50% lower than the alkaline process, meeting energy conservation and emission reduction requirements. 3. Obvious process advantages: The use of collectors and a three-stage reverse flotation process effectively improves the quality of the aluminum concentrate. The acid leaching-separation coupling technology achieves efficient separation of aluminum and iron, improving the purity of alumina. The solid waste resource utilization technology maximizes resource utilization and reduces environmental impact.

[0058] The technical solutions of this invention will be clearly and completely described below with reference to the embodiments thereof. Obviously, the described embodiments are only a part of the embodiments of this invention, and not all of them. All other embodiments obtained by those skilled in the art based on the embodiments of this invention without creative effort are within the scope of protection of this invention.

[0059] Example 1 A method for efficiently extracting alumina from low-grade bauxite is as follows: (1) Take 5 tons of low-grade bauxite (32wt% Al2O3, 8.5wt% SiO2, 8.2wt% Fe2O3, 1.85wt% TiO2, 5.3wt% CaO, 0.6wt% MgO, 0.65wt% K2O, 0.3wt% Na2O) and feed it into a crusher for crushing. Then, grind the ore to -200 particle size using a grinding mill to obtain pretreated bauxite. (2) Add water to the pretreated aluminum ore in a mixing tank to prepare the slurry (slurry mass concentration 25%). Add Na2CO3 to adjust the pH of the slurry to 8. Pump the prepared slurry into the flotation tank. At the same time, add the prepared collector (the collector is made by mixing oleic acid and sodium petroleum sulfonate at a mass ratio of 3:1) to the tank along with the slurry using a dosing pump. Add 150g of collector per ton of slurry (the mass ratio of collector to slurry is 0.15:1000). Perform the first reverse flotation at 20℃ for 20min (air is introduced from the bottom during the reverse flotation process, with an air pressure of 0.25MPa and a flow rate of 0.25m³). 3 / h), skim off the foam; adjust the pulp concentration after the first reverse flotation to 30wt%, add sodium hydroxide to adjust the pH to 8.5, add 100g / t collector, and perform a second aerated reverse flotation at 20℃ for 20min, skim off the foam; adjust the pulp concentration after the second reverse flotation to 30wt%, add sodium hydroxide to adjust the pH to 9, add 60g / t collector, and perform a third aerated reverse flotation at 20℃ for 15min to obtain aluminum concentrate; (3) Mix the flotation-fled aluminum concentrate with hydrochloric acid solution (mass concentration 20%) (the volume ratio of hydrochloric acid solution to aluminum concentrate is 7 mL: 1 g), and perform pressurized acid leaching for 2 h at a temperature of 110℃ and a pressure of 0.4 MPa to fully dissolve the aluminum oxide and iron oxide in the aluminum concentrate. After leaching, perform solid-liquid separation to obtain aluminum-containing leachate and leaching residue. (4) Add Na2CO3 to the leachate to adjust the pH of the leachate to 1.5. Precipitate iron impurities at 20℃ for 45 min, filter to separate the iron slag, add Na2CO3 again to adjust the pH of the filtrate to 4.5, precipitate aluminum at 40℃ for 90 min, filter to obtain aluminum hydroxide; rinse the iron slag twice with distilled water, and then heat at 300℃ to decompose to constant weight to obtain Fe2O3; (5) Add aluminum hydroxide and 15% hydrochloric acid at a mass ratio of 1:9 to the corrosion-resistant reactor and react at 80°C for 50 min to generate aluminum chloride solution. The aluminum chloride solution was transferred to a vacuum evaporator and concentrated under stirring at 40°C. When the solution reached supersaturation, 0.5% (by mass) of aluminum chloride hexahydrate seed crystals were added, and stirring continued to allow aluminum chloride to crystallize and precipitate around the seed crystals. The solution was then filtered through a vacuum filter and the crystals were washed with water to obtain crystalline aluminum chloride. The crystalline aluminum chloride was then pyrolyzed in a rotary kiln, calcined sequentially at 300°C for 10 min, 500°C for 40 min, and 800°C for 60 min to obtain high-purity alumina. The hydrogen chloride gas generated during pyrolysis was absorbed and recovered using water in a spray tower.

[0060] Testing revealed that the alumina obtained in Example 1 contained 93.42 wt% Al2O3, 2.19 wt% Fe2O3, 1.72 wt% SiO2, and 1.21 wt% CaO. With 5 tons of raw materials, the recovered product 1 was 1.19 tons of Al2O3, achieving a leaching rate of 87%. Product 2 was 0.386 tons of Fe2O3.

[0061] Example 2 A method for efficiently extracting alumina from low-grade bauxite is as follows: (1) Take 5 tons of low-grade bauxite (Al2O3 41wt%, SiO2 9.76wt%, Fe2O3 6wt%, TiO2 2.5wt%, CaO 1.5wt%, MgO 0.8wt%, K2O 0.4wt%, Na2O 0.2wt%) and feed it into a crusher for crushing. Then, grind the ore to -200 particle size using a grinding mill to obtain pretreated bauxite. (2) Add water to the pretreated aluminum ore in a mixing tank to prepare the slurry (slurry mass concentration 25%). Add Na2CO3 to adjust the pH of the slurry to 8. Pump the prepared slurry into the flotation tank. At the same time, add the prepared collector (the collector is made by mixing oleic acid and sodium petroleum sulfonate at a mass ratio of 3:1) to the flotation tank along with the slurry using a dosing pump. Add 150g of collector per ton of slurry. Perform the first reverse flotation at 20℃ for 20min (air is introduced from the bottom during the reverse flotation process, with an air pressure of 0.25MPa and a flow rate of 0.25m³ / h). 3 / h), skim off the foam; adjust the pulp concentration after the first reverse flotation to 30wt%, add sodium hydroxide to adjust the pH to 8.5, add 100g / t collector, and perform a second aerated reverse flotation at 20℃ for 20min, skim off the foam; adjust the pulp concentration after the second reverse flotation to 30wt%, add sodium hydroxide to adjust the pH to 9, add 60g / t collector, and perform a third aerated reverse flotation at 20℃ for 15min to obtain aluminum concentrate; (3) Mix the flotation-fled aluminum concentrate with hydrochloric acid solution (mass concentration 20%) (the volume ratio of hydrochloric acid solution to aluminum concentrate is 7 mL: 1 g), and perform pressurized acid leaching for 2 h at a temperature of 110℃ and a pressure of 0.4 MPa to fully dissolve the aluminum oxide and iron oxide in the aluminum concentrate. After leaching, perform solid-liquid separation to obtain aluminum-containing leachate and leaching residue. (4) Add Na2CO3 to the leachate to adjust the pH of the leachate to 1.5. Precipitate iron impurities at 20℃ for 45 min, filter to separate the iron slag, add Na2CO3 again to adjust the pH of the filtrate to 4.5, precipitate aluminum at 40℃ for 90 min, filter to obtain aluminum hydroxide; rinse the iron slag twice with distilled water, and then heat at 300℃ to decompose to constant weight to obtain Fe2O3; (5) Add aluminum hydroxide and 15% hydrochloric acid at a mass ratio of 1:9 to the corrosion-resistant reactor and react at 80°C for 50 min to generate aluminum chloride solution. The aluminum chloride solution was transferred to a vacuum evaporator and concentrated under stirring at 40°C. When the solution reached supersaturation, 0.5% (by mass) of aluminum chloride hexahydrate seed crystals were added, and stirring continued to allow aluminum chloride to crystallize and precipitate around the seed crystals. The solution was then filtered through a vacuum filter and the crystals were washed with water to obtain crystalline aluminum chloride. The crystalline aluminum chloride was then pyrolyzed in a rotary kiln, calcined sequentially at 300°C for 10 min, 500°C for 30 min, and 800°C for 30 min to obtain high-purity alumina. The hydrogen chloride gas generated during pyrolysis was absorbed and recovered by water in a spray tower. Testing revealed that the alumina obtained in Example 2 contained 96.4 wt% Al2O3, 1.73 wt% Fe2O3, and 1.58 wt% SiO2. With 5 tons of raw materials, the recovered product 1 consisted of 1.46 tons of Al2O3, achieving a leaching rate of 86%; the product 2 consisted of 0.29 tons of Fe2O3.

[0062] Example 3 A method for efficiently extracting alumina from low-grade bauxite is as follows: (1) Take 5 tons of low-grade bauxite (Al2O3 41wt%, SiO2 9.76wt%, Fe2O3 6wt%, TiO2 2.5wt%, CaO 1.5wt%, MgO 0.8wt%, K2O 0.4wt%, Na2O 0.2wt%) and feed it into a crusher for crushing. Then, grind the ore to -200 particle size using a grinding mill to obtain pretreated bauxite. (2) Add water to the pretreated aluminum ore in a mixing tank to prepare the slurry (slurry concentration 30%). Add Na2CO3 to adjust the pH of the slurry to 8. Pump the prepared slurry into the flotation tank. At the same time, add the prepared collector (the collector is made by mixing oleic acid and sodium petroleum sulfonate at a mass ratio of 3:1) to the tank along with the slurry using a dosing pump. Add 180g of collector per ton of slurry (the mass ratio of collector to slurry is 0.18:1000). Perform the first reverse flotation at 20℃ for 20min (air is introduced from the bottom during the reverse flotation process, with an air pressure of 0.25MPa and a flow rate of 0.2m³). 3 / h), skim off the foam; adjust the pulp concentration after the first reverse flotation to 30wt%, add sodium hydroxide to adjust the pH to 8.5, add 150g / t collector, and perform a second aerated reverse flotation at 20℃ for 20min, skim off the foam; adjust the pulp concentration after the second reverse flotation to 30wt%, add sodium hydroxide to adjust the pH to 9, add 100g / t collector, and perform a third aerated reverse flotation at 20℃ for 15min to obtain aluminum concentrate; (3) Mix the flotation-fled aluminum concentrate with hydrochloric acid solution (mass concentration 20%) (the volume ratio of hydrochloric acid solution to aluminum concentrate is 7 mL: 1 g), and perform pressurized acid leaching for 2 h at a temperature of 110℃ and a pressure of 0.4 MPa to fully dissolve the aluminum oxide and iron oxide in the aluminum concentrate. After leaching, perform solid-liquid separation to obtain aluminum-containing leachate and leaching residue. (4) Add Na2CO3 to the leachate to adjust the pH of the leachate to 1.5. Precipitate iron impurities at 20℃ for 45 min, filter to separate the iron slag, add Na2CO3 again to adjust the pH of the filtrate to 4.5, precipitate aluminum at 40℃ for 90 min, filter to obtain aluminum hydroxide; rinse the iron slag twice with distilled water, and then heat at 300℃ to decompose to constant weight to obtain Fe2O3; (5) Add aluminum hydroxide and 15% hydrochloric acid at a mass ratio of 1:9 to the corrosion-resistant reactor and react at 80°C for 50 min to generate aluminum chloride solution. The aluminum chloride solution was transferred to a vacuum evaporator and concentrated under stirring at 40°C. When the solution reached supersaturation, 0.5% (by mass) of aluminum chloride hexahydrate seed crystals were added, and stirring continued to allow aluminum chloride to crystallize and precipitate around the seed crystals. The solution was then filtered through a vacuum filter and the crystals were washed with water to obtain crystalline aluminum chloride. The crystalline aluminum chloride was then pyrolyzed in a rotary kiln, calcined sequentially at 300°C for 10 min, 500°C for 30 min, and 800°C for 30 min to obtain high-purity alumina. The hydrogen chloride gas generated during pyrolysis was absorbed and recovered by water in a spray tower. Testing revealed that the alumina obtained in Example 3 contained 97.4 wt% Al2O3, 0.26 wt% Fe2O3, and 1.03 wt% SiO2. With 5 tons of raw materials, the recovered product 1 was 1.54 tons of Al2O3, with a leaching rate of 88%, and product 2 was 0.331 tons of Fe2O3.

[0063] The above description is only a preferred embodiment of the present invention. It should be noted that for those skilled in the art, several improvements and modifications can be made without departing from the principle of the present invention, and these improvements and modifications should also be considered within the scope of protection of the present invention.

Claims

1. A method for efficiently extracting alumina from low-grade bauxite, comprising the following steps: (1) The low-grade aluminum ore is crushed and ground in sequence to obtain pretreated aluminum ore; (2) The pretreated aluminum ore obtained in step (1) is mixed with water and a first alkali to obtain a slurry; the slurry is mixed with a collector and subjected to reverse flotation to obtain aluminum concentrate; (3) The aluminum concentrate obtained in step (2) is mixed with the first acid solution and subjected to pressure acid leaching to obtain a leachate; (4) The leachate obtained in step (3) is mixed with the second alkali to precipitate iron for the first time, and the filtrate is obtained; the filtrate is mixed with the third alkali to precipitate aluminum for the second time, and aluminum hydroxide is obtained. (5) Mix the aluminum hydroxide obtained in step (4) with the second acid solution to obtain an aluminum salt solution; The aluminum salt solution is concentrated and seed crystals are added to crystallize and obtain aluminum salt; the aluminum salt is then calcined to obtain aluminum oxide.

2. The method according to claim 1, characterized in that, The pretreated aluminum ore in step (1) has a particle size ≥200 mesh.

3. The method according to claim 1, characterized in that, In step (2), the mass concentration of the slurry is 20-30%; the pH value of the slurry is 8-9.

4. The method according to claim 1, characterized in that, The collector in step (2) includes fatty acids and sulfonates; the mass ratio of the fatty acids and sulfonates is (2~4):1; the mass ratio of the collector to the slurry is (0.05~0.2):1000.

5. The method according to claim 1, characterized in that, The number of reverse flotation steps in step (2) is 2 to 4.

6. The method according to claim 1, characterized in that, The first acid in step (3) is hydrochloric acid; the mass concentration of the first acid is 15~25%; the volume ratio of the first acid to the mass of aluminum concentrate is (6~8) mL:1g.

7. The method according to claim 1, characterized in that, In step (3), the pressure of the acid leaching is 0.35~0.4MPa, the temperature of the acid leaching is 100~120℃, and the time of the acid leaching is 1.5~2h.

8. The method according to claim 1, characterized in that, In step (4), the pH value of the first precipitated iron is 1~2; the pH value of the second precipitated aluminum is 4~5.

9. The method according to claim 1, characterized in that, The second acid solution in step (5) is hydrochloric acid; the mass concentration of the second acid solution is 10~20%; the mass ratio of the aluminum hydroxide to the second acid solution is 1:(8~10).

10. The method according to claim 1, characterized in that, The calcination in step (5) includes a first calcination, a second calcination, and a third calcination performed sequentially; the temperature of the first calcination is 200~390℃ and the time of the first calcination is 10~15min; the temperature of the second calcination is 400~600℃ and the time of the second calcination is 30~50min; the temperature of the third calcination is 800~1000℃ and the time of the third calcination is 30~70min.