A processing method for red mud recycling

Abstract

The application provides a treatment method for recycling red mud, and belongs to the technical field of red mud resource utilization; a short-process closed-circuit circulation process is used for removing alkali through water washing, hydrochloric acid leaching, solid-liquid separation, solvent extraction separation of iron, selective precipitation and separation of rare earth, preparation of polyaluminum chloride, preparation of iron oxide and regeneration of hydrochloric acid; the selective precipitation method is used for preferentially separating and recycling rare earth, and the polyaluminum chloride is prepared again; the step-by-step separation and high-value utilization of aluminum, iron and rare earth are realized; the Al / Fe element leaching rate is greater than 85%, and the comprehensive utilization rate of acid-iron resources is improved; in addition, the content of rare earth in the red mud is extremely low, the rare earth precipitation in the application is separated through solid-liquid separation, trace rare earth elements are captured from a large amount of solution, an independent rare earth product is formed, the silicon-titanium slag is used for building material, the recovery rate is significantly improved, and there is no secondary pollution; the application has the advantages of simplified process, low energy consumption, high resource recovery rate and superior economy, effectively solves the problems of red mud storage pollution and resource waste, and has the industrialization popularization value.

Description

Technical Field

[0001] This invention belongs to the field of red mud resource utilization technology, specifically relating to a method for recycling and utilizing red mud. Background Technology

[0002] Red mud is an industrial solid waste generated during the alumina smelting process from bauxite. Approximately 1.0-1.8 tons of red mud are produced for every ton of alumina produced. Red mud contains valuable metallic elements such as aluminum, iron, and rare earth elements, and is also rich in oxides such as silicon and titanium. However, due to its high alkalinity, fine particles, and easy leaching, it is currently mainly disposed of through stockpiling. This not only occupies a large amount of land resources but may also cause environmental problems such as soil and water pollution, resulting in serious waste of resources.

[0003] Existing red mud resource utilization technologies suffer from the following shortcomings: First, current red mud recycling processes are lengthy, often employing multi-stage leaching and multiple separation processes, resulting in high equipment investment and energy consumption. Second, the recovery rate of valuable components in red mud is low, especially rare earth elements, which are highly dispersed and difficult to recover on a large scale. Third, existing red mud recycling systems do not form a closed-loop cycle, leading to high consumption of resources such as acid and water, high production costs, and the generation of secondary pollutants. Finally, existing red mud resource utilization technologies produce products with low added value, primarily primary chemical raw materials, resulting in weak market competitiveness. Therefore, developing a red mud recycling technology that features a short process, high resource recovery rate, and is environmentally friendly and economical is of significant practical importance. Summary of the Invention

[0004] This invention addresses the problems of existing technologies by providing a method for the recycling and utilization of red mud.

[0005] To achieve the above objectives, the technical solution adopted in this application is as follows: A method for recycling and utilizing red mud, the specific steps of which are as follows: Step 1: Red mud pretreatment: The red mud is sent into a three-stage countercurrent washing tank and washed in a three-stage countercurrent condition at 60-80℃ and a liquid-solid ratio of 4:1. After being treated by a thickener, the pH of the red mud at the outlet is controlled to be less than 10.0 in order to remove free alkali and reduce the subsequent hydrochloric acid consumption, thus obtaining pretreated red mud. Step 2: Hydrochloric acid leaching: The pretreated red mud is fed into an enamel-lined reactor, and leached for 15 minutes in a hydrochloric acid solution with a concentration of 4-5 mol / L. Then, a polyquaternary ammonium salt silicate inhibitor with a mass ratio of 0.08% is added, and leaching continues until the total leaching time reaches 100 minutes. The final pH is controlled at 0.8 to obtain the leachate. The hydrochloric acid solution includes hydrochloric acid and recycled hydrochloric acid; the mass fraction of hydrochloric acid and recycled hydrochloric acid is 14%-17%. Step 3: Solid-liquid separation: The leachate is sent to a plate and frame filter press for filtration; a three-stage countercurrent washing is carried out under a pressure of 0.7 MPa to obtain a filtrate rich in aluminum, iron, and rare earth elements and a silicon-titanium slag filter residue; the filtrate includes AlCl3, FeCl3, and REECl3; the silicon-titanium slag filter residue is an inert material and can be directly used for building materials or stored in a red mud pond without environmental risks; the filtrate is reserved for later use. Step 4: Solvent extraction to separate iron: The filtrate is sent to a three-stage countercurrent mixing and clarifying extraction tank. An N503 / sulfonated kerosene mixture with a volume ratio of 3:7 is used as the extraction system, and the ratio of organic phase to aqueous phase is controlled at 1:2.5, with a mixing time of 3 min. During extraction, the acidity of the aqueous phase feed is maintained at 4.0~4.5 mol / L HCl. Iron-loaded organic phase and aluminum-containing rare earth raffinate are obtained. Pure water is used as the back-extraction agent, and the ratio of organic phase to aqueous phase is controlled at 3:1 to back-extract the iron-loaded organic phase, yielding a pure FeCl3 solution. Step 5: Selective precipitation and separation of rare earth elements: The aluminum-containing rare earth raffinate obtained in Step 4 is pumped into a multi-effect evaporator for concentration to ensure that the aluminum chloride content is ≥8%. If the initial aluminum chloride content is ≥8%, subsequent processing can be carried out directly. Then, oxalic acid or carbonate is added as a precipitant to adjust the pH to 4.0~4.5, and the mixture is heated and stirred to allow the rare earth elements to precipitate selectively. After filtration, the (oxalic acid) rare earth precipitate and the purified aluminum acidic mother liquor are obtained. Step 6: Preparation of polyaluminum chloride (PAC): Pump the purified aluminic mother liquor into the polymerization reactor, heat it to 80~90℃, add aluminate or alkaline solution to carry out the polymerization reaction, control the alkalinity (B value) at 50%~65%, and obtain liquid polyaluminum chloride product. Step 7: Iron oxide preparation and hydrochloric acid recovery: The back-extraction liquid of the pure FeCl3 solution obtained in Step 4 is sent to a spray-type roasting furnace for roasting; the furnace temperature is controlled at 410~450℃ and the atomization pressure is 0.5MPa; during this process, the reaction 2FeCl3+3H2O→Fe2O3+6HCl↑ occurs, yielding pigment-grade Fe2O3 solid powder and HCl gas. The HCl gas is absorbed by a graphite falling film absorption tower to generate regenerated hydrochloric acid with a mass fraction of 20%~22%, which is returned to Step 2 for recycling.

[0006] Preferably, in step two, the leaching rate of Al element is >85%, the leaching rate of Fe element is >85%, and the leaching rate of rare earth elements is >75%.

[0007] Preferably, in step two, the solid-liquid ratio of hydrochloric acid to pretreated red mud is 5:1, and the leaching temperature is 70~80℃.

[0008] Preferably, in step four, the extraction system further includes one or more of the following: tributyl phosphate (TBP), dimethylheptyl methylphosphonate (P350), di(2-ethylhexyl)phosphonic acid (P204), mono-2-ethylhexyl 2-ethylhexyl phosphate (P507), and methyl isobutyl ketone (MIBK).

[0009] Preferably, in step six, the alkaline solution is either the alkaline wastewater generated during the rare earth precipitation in step five, or an added sodium carbonate solution or sodium hydroxide solution with a mass fraction of 10%.

[0010] Compared with the prior art, the advantages and positive effects of the present invention are as follows: This application discloses a method for recycling red mud, employing a short-process closed-loop cycle involving water washing and alkali removal, hydrochloric acid leaching, solid-liquid separation, solvent extraction for iron separation, selective precipitation and separation of rare earth elements, preparation of polyaluminum chloride, preparation of iron oxide, and hydrochloric acid regeneration. This process simultaneously recovers three valuable components from red mud: aluminum, iron, and rare earth elements. The Al / Fe leaching rate is >85%, significantly improving the comprehensive utilization rate of acid iron resources. Furthermore, given the extremely low rare earth content in red mud, this method uses solid-liquid separation to extract trace amounts of rare earth elements from a large volume of solution, forming independent rare earth products. The silicon-titanium slag can be used in building materials, and the sodium salt can be recycled into industrial salt. There are no secondary pollutant emissions, and the recovery rate is significantly improved, solving the pollution problem caused by red mud stockpiling. The produced polyaluminum chloride and pigment-grade iron oxide red are both in-demand commodities with large market capacity and stable demand. The mixed rare earth oxides are strategic resources, and the products have high market acceptance and superior economic performance. The process parameters are clear, the equipment is mature, and the operation is simple, making it suitable for large-scale promotion and application. Attached Figure Description

[0011] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0012] Fig. 1 This is a picture of a pigment-grade Fe2O3 solid powder product. Fig. 2 Here is a picture of a polyaluminum chloride product. Fig. 3 This is a flowchart of a red mud recycling and treatment method. Detailed Implementation

[0013] To better understand the above-mentioned objectives, features, and advantages of the present invention, the present invention will be further described below in conjunction with the accompanying drawings and embodiments. It should be noted that, unless otherwise specified, the embodiments and features described in these embodiments can be combined with each other.

[0014] Numerous specific details are set forth in the following description in order to provide a full understanding of the invention. However, the invention may also be practiced in other ways than those described herein, and therefore the invention is not limited to the specific embodiments disclosed in the following specification.

[0015] Example 1, such as Figs. 1-3 As shown, the specific steps of the red mud recycling method of this application are as follows: Step 1: Red mud pretreatment: The red mud is sent into a three-stage countercurrent washing tank and washed in a three-stage countercurrent condition at 70℃ and a liquid-solid ratio of 4:1. After being treated by a thickener, the pH of the red mud at the outlet is controlled to be 9.5 to obtain pretreated red mud. Step 2: Hydrochloric acid leaching: The pretreated red mud is fed into an enamel-lined reactor and leached for 15 minutes in a 4 mol / L hydrochloric acid solution. The solid-liquid ratio of hydrochloric acid to pretreated red mud is 5:1, and the leaching temperature is 75℃. Subsequently, 0.08% (by mass) of polyquaternary ammonium silicate inhibitor is added, and leaching continues until the total leaching time reaches 100 minutes. The final pH is controlled at 0.8 to obtain the leachate. The hydrochloric acid solution includes hydrochloric acid and recovered hydrochloric acid. The concentration of the hydrochloric acid solution is 15% by mass, and the mass percentage of recovered hydrochloric acid is 90%. Step 3: Solid-liquid separation: The leachate is sent to a plate and frame filter press for filtration; a three-stage countercurrent washing is carried out under a pressure of 0.7 MPa to obtain a filtrate containing AlCl3, FeCl3, HCl, and REECl3 and a silicon-titanium slag filter residue. Step 4: Solvent extraction to separate iron: The filtrate is sent to a three-stage countercurrent mixing and clarifying extraction tank. N503 / sulfonated kerosene with a volume ratio of 3:7 is used as the extraction system, and the ratio of organic phase to aqueous phase (O / A) is controlled at 1:2.5, with a mixing time of 3 min. During the extraction process, the acidity of the aqueous phase feed is maintained at 4.2 mol / L HCl. Iron-loaded organic phase and aluminum-containing rare earth raffinate are obtained. Pure water is used as the back-extraction agent, and the ratio of organic phase to aqueous phase is controlled at 3:1 to back-extract the iron-loaded organic phase, resulting in a pure FeCl3 solution. Step 5: Selective Precipitation and Separation of Rare Earths: The aluminum-containing rare earth raffinate from Step 4 is concentrated in a multi-effect evaporator after impurity removal from the acidic mother liquor to achieve an aluminum chloride content ≥8%. The concentrate is then transferred to a precipitation tank and heated to 75°C. A 10% sodium carbonate solution is slowly and evenly added while stirring to adjust the pH to 4.2. At this point, rare earth ions rapidly hydrolyze to form rare earth hydroxide precipitate, while aluminum ions remain stably dissolved in the acidic solution. After filtration, rare earth precipitate (crude rare earth hydroxide) and the impurity-removed acidic mother liquor are obtained. The rare earth precipitate is converted using oxalic acid precipitation, washed, dried, and then calcined at 850°C for 2 hours to obtain a mixed rare earth oxide product. Step Six: Preparation of Polyaluminum Chloride (PAC); Subsequently, the aluminic mother liquor after impurity removal in Step Five is transferred into a polymerization reactor, heated to 85°C, calcium aluminate powder is added to adjust the basicity, the pH at the reaction endpoint is controlled to be 4.0, and the reaction is carried out for 90 minutes to obtain liquid polyaluminum chloride product; Step 7: Iron oxide preparation and hydrochloric acid recovery: The back-extraction liquid of the pure FeCl3 solution is sent to a spray-type roasting furnace for roasting; the furnace temperature is controlled at 435℃ and the atomization pressure is 0.5MPa to obtain pigment-grade Fe2O3 solid powder and HCl gas. The HCl gas is absorbed by a graphite falling film absorption tower to generate regenerated hydrochloric acid with a concentration of 6.4mol / L. The regenerated hydrochloric acid is used for leaching in step two of the subsequent red mud recycling process.

[0016] Example 2 describes a method for recycling red mud, with the following specific steps: Step 1: Red mud pretreatment: The red mud is sent into a three-stage countercurrent washing tank and washed in a three-stage countercurrent condition at 70℃ and a liquid-solid ratio of 4:1. After being treated by a thickener, the pH of the red mud at the outlet is controlled to be 9.5 to obtain pretreated red mud. Step 2: Hydrochloric acid leaching: The pretreated red mud was fed into an enamel-lined reactor and leached for 15 minutes in a 4.2 mol / L hydrochloric acid solution. The solid-liquid ratio of hydrochloric acid to pretreated red mud was 5:1, and the leaching temperature was 75°C. Subsequently, 0.08% by mass of polyquaternary ammonium silicate inhibitor was added, and leaching continued until the total leaching time reached 100 minutes. The final pH was controlled at 0.8 to obtain the leachate. 95% of the hydrochloric acid solution was 6.4 mol / L regenerated hydrochloric acid produced in Step 6 of Example 1, with only 5% fresh hydrochloric acid added to compensate for chloride ions carried away by polyaluminum chloride, etc. The solution was diluted to 4.2 mol / L with process washing water within the system. Step 3: Solid-liquid separation: The leachate is sent to a plate and frame filter press for filtration; a three-stage countercurrent washing is carried out under a pressure of 0.7 MPa to obtain a filtrate containing AlCl3, FeCl3, HCl, and REECl3 and a silicon-titanium slag filter residue. Step 4: The filtrate is sent to a three-stage countercurrent mixing and clarifying extraction tank. The N503 / sulfonated kerosene volume ratio of 3:7 is used as the extraction system, and the ratio of organic phase to aqueous phase is controlled at 1:2.5, with a mixing time of 3 min. During the extraction process, the acidity of the aqueous phase feed is maintained at 4.2 mol / L HCl. Pure water is used as the back-extraction agent, and the ratio of organic phase to aqueous phase is controlled at 3:1 for back-extraction to obtain an organic phase containing AlCl3, unreacted hydrochloric acid, and rare earth elements, as well as a raffinate containing FeCl3. Step 5: The aluminum-containing rare earth raffinate from Step 4 is concentrated in a multi-effect evaporator after impurity removal from the aluminic acid mother liquor to ensure an aluminum chloride content of ≥8%. The concentrate is then transferred to a precipitation tank and heated to 75°C. A 5% oxalic acid solution is then slowly and evenly added while stirring to adjust the pH to 4.5. At this point, rare earth ions rapidly hydrolyze to form rare earth hydroxide precipitate, while aluminum ions remain stably dissolved in the acidic solution. After filtration, rare earth oxalate precipitate (crude rare earth hydroxide) and the aluminic acid mother liquor after impurity removal are obtained. The rare earth oxalate precipitate is then washed, dried, and calcined at 850°C for 2 hours to obtain a mixed rare earth oxide product.

[0017] Step 6: Transfer the purified aluminic mother liquor into the polymerization reactor, heat to 85°C, react for 75 minutes, add calcium aluminate powder at a stirring frequency of 40 rpm to adjust the basicity, control the pH at the reaction endpoint to be around 3.8 and the basicity to 50%, and obtain liquid polyaluminum chloride product. Step 7: The back-extraction solution of the pure FeCl3 solution is fed into a spray-type roasting furnace for roasting; the furnace temperature is controlled at 435℃ and the atomization pressure is 0.5MPa to obtain pigment-grade Fe2O3 solid powder and HCl gas. The HCl gas is absorbed by a graphite falling film absorption tower to generate regenerated hydrochloric acid with a concentration of 6.2mol / L. The regenerated hydrochloric acid is used for leaching in step 2 of the subsequent red mud recycling process.

[0018] Example 3 describes a method for recycling red mud, with the following specific steps: Step 1: Red mud pretreatment: The red mud is sent into a three-stage countercurrent washing tank and washed in a three-stage countercurrent condition at 70℃ and a liquid-solid ratio of 4:1. After being treated by a thickener, the pH of the red mud at the outlet is controlled to be 9.5 to obtain pretreated red mud. Step 2: Hydrochloric acid leaching: The pretreated red mud was fed into an enamel-lined reactor and leached for 15 minutes in a 4.2 mol / L hydrochloric acid solution. The solid-liquid ratio of hydrochloric acid to pretreated red mud was 5:1, and the leaching temperature was 75°C. Subsequently, 0.08% (w / w) of polyquaternary ammonium salt silicon inhibitor was added, and leaching continued until the total leaching time reached 100 minutes. The final pH was controlled at 0.8 to obtain the leachate. The hydrochloric acid solution was 100% the 6.2 mol / L solution produced in Step 6 of Example 2. L of regenerated hydrochloric acid was diluted with process washing water and supplemented with a very small amount of fresh hydrochloric acid to make up for the trace losses in the two cycles, and adjusted to 4.2 mol / L. The specific preparation method is as follows: for every 1000L of leaching hydrochloric acid solution prepared, take 645L of 6.2 mol / L regenerated hydrochloric acid, add 340L of process washing water, and then add 15L of 4.2 mol / L fresh industrial hydrochloric acid to make up for the process losses. The hydrochloric acid regeneration utilization rate reaches 92.5%, which fully meets the technical effect of the present invention of hydrochloric acid regeneration rate >90%. Step 3: Solid-liquid separation: The leachate is sent to a plate and frame filter press for filtration; a three-stage countercurrent washing is carried out under a pressure of 0.7 MPa to obtain a filtrate containing AlCl3, FeCl3, HCl, and REECl3 and a silicon-titanium slag filter residue. Step 4: The filtrate is sent to a three-stage countercurrent mixing and clarifying extraction tank. The N503 / sulfonated kerosene volume ratio of 3:7 is used as the extraction system, and the ratio of organic phase to aqueous phase is controlled at 1:2.5, with a mixing time of 3 min. During the extraction process, the acidity of the aqueous phase feed is maintained at 4.2 mol / L HCl. Pure water is used as the back-extraction agent, and the ratio of organic phase to aqueous phase is controlled at 3:1 for back-extraction to obtain an organic phase containing AlCl3, unreacted hydrochloric acid, and rare earth elements, as well as a raffinate containing FeCl3. Step 5: The aluminum-containing rare earth raffinate from Step 4 is concentrated in a multi-effect evaporator after impurity removal from the aluminic acid mother liquor to ensure an aluminum chloride content of ≥8%. The concentrate is then transferred to a precipitation tank and heated to 75°C. A 12% (w / w) oxalic acid solution is then slowly and evenly added while stirring to adjust the pH to 4.5. At this point, rare earth ions rapidly hydrolyze to form rare earth hydroxide precipitate, while aluminum ions remain stably dissolved in the acidic solution. After filtration, rare earth oxalate precipitate (crude rare earth hydroxide) and the purified aluminic acid mother liquor are obtained. The rare earth oxalate precipitate is then washed, dried, and calcined at 850°C for 2 hours to obtain a mixed rare earth oxide product.

[0019] Step 6: Transfer the purified aluminic mother liquor into the polymerization reactor, heat to 90°C, react for 90 minutes, and stir at 60 rpm; add 10% sodium carbonate solution to adjust the basicity, control the final pH of the reaction to be around 4.2 and the basicity to 65%, and obtain liquid polyaluminum chloride product. Step 7: The back-extraction solution of the pure FeCl3 solution is fed into a spray-type roasting furnace for roasting; the furnace temperature is controlled at 440℃ and the atomization pressure is 0.5MPa to obtain pigment-grade Fe2O3 solid powder and HCl gas. The HCl gas is absorbed by a graphite falling film absorption tower to generate regenerated hydrochloric acid with a concentration of 6.2mol / L. The regenerated hydrochloric acid is used for leaching in step 2 of the subsequent red mud recycling process.

[0020] I. Purity and performance verification of polyaluminum chloride products, pigment-grade Fe2O3 solid powder, and mixed rare earth oxide products: (A1) Verification of the purity and water treatment performance of polyaluminum chloride products Polyaluminum chloride (PAC) yield (based on Al2O3) = [actual total mass of Al2O3 in PAC product / initial total mass of Al2O3 in red mud] × 100%; The actual total mass of Al2O3 in the PAC product = liquid PAC volume (L) × density (g / cm³) 3 ) × Al2O3 mass fraction (%) / 100 (For solid PAC, calculate directly using [product dry weight × Al2O3 mass fraction]) The initial total mass of Al2O3 in red mud = dry mass of treated red mud (kg) × initial Al2O3 content in red mud (%) / 100; The specific experimental steps are as follows: Take 50g of representative dry red mud sample, crush it through a 200-mesh sieve, and determine the Al content using ICP-OES according to GB / T14506.28-2010 Chemical Analysis Methods for Silicate Rocks, and convert it to the initial Al2O3 content (denoted as ω1).

[0021] PAC Product Measurement and Testing: Record the total volume (V, L) of liquid PAC produced in each batch of Examples 1-3, and measure the product density (ρ, g / cm³) using a densitometer. 3 The mass fraction of Al2O3 was determined by the gravimetric method according to GB / T22627-2014 (denoted as ω2).

[0022] Yield calculation and parallel verification: Substitute into the formula to calculate the yield, and take the average of the results of 3 sets of parallel experiments; if the yield is greater than 75%, the corresponding Al leaching rate is greater than 85%.

[0023] In Examples 1-3, the products were tested according to GB / T22627-2022 Water Treatment Agents Polyaluminum Chloride: the initial turbidity of the wastewater was 200 NTU and the initial COD concentration was 300 mg / L. The PAC sample was added at a dosage of 50 mg / L, and after being stirred rapidly for 1 min (300 r / min) and slowly for 10 min (60 r / min) with a six-unit stirrer, it was allowed to settle for 30 min. The turbidity and COD concentration of the supernatant were measured, and the corresponding removal rate was calculated. Example 1: Turbidity removal rate of the product is 98.2% and COD removal rate is 85.6%; Example 2: Turbidity removal rate of the product is 96.5% and COD removal rate is 82.1%; Example 3: Turbidity removal rate of the product is 97.8% and COD removal rate is 84.3%. The purity and basicity of the products in Examples 1-3 all meet the national standards, and they have excellent flocculation performance, so they can be directly used in industrial water treatment, municipal sewage treatment and other scenarios.

[0024] (A2) Verification of the purity and pigment performance of pigment-grade Fe2O3 solid powder The products in Examples 1-3 were all tested according to GB / T1863-2008 Iron Oxide Pigments. The specific method for determining the purity of Fe2O3 is as follows: the test is carried out according to the potassium dichromate titration method specified in the national standard. The sample is dissolved by heating with hydrochloric acid, ferric iron is reduced with stannous chloride, excess stannous chloride is oxidized with mercuric chloride, and the purity of Fe2O3 is calculated by titrating with potassium dichromate standard solution using sodium diphenylamine sulfonate as an indicator. The specific testing methods for the core performance of pigments are as follows: According to national standards, the tinting strength of the product is determined using the relative tinting strength method, and the hiding power is determined using the black and white grid method. The pass / fail criteria are: Fe2O3 purity ≥ 95%, tinting strength ≥ 95%, and hiding power ≤ 80 g / m³. 2 ; Example 1: The product has a ferric oxide purity of 98.5%, a tinting strength of 102%, and a hiding power of 75 g / m³. 2 Example 2: The product has a ferric oxide purity of 98.1%, a tinting strength of 100%, and a hiding power of 78 g / m³. 2 Example 3: The product has a ferric oxide purity of 98.3%, a tinting strength of 101%, and a hiding power of 76 g / m³. 2 All of them meet the national standard requirements for pigment-grade iron oxide red.

[0025] (A3) Verification of the purity and enrichment factor of mixed rare earth oxide products ICP-MS analysis accurately determined the total amount of rare earth elements such as La, Ce, and Nd, and calculated the total purity of rare earth oxides (REO) in the product: Example 1: Total purity of mixed rare earth oxides 92.5%, total rare earth recovery rate 76.8%, perfectly matching the process design value, with an enrichment factor of 48 times; Example 2: Total purity of mixed rare earth oxides 91.2%, total rare earth recovery rate 75.3%, perfectly matching the process design value, with an enrichment factor of 45 times; Example 3: Total purity of mixed rare earth oxides 92.1%, total rare earth recovery rate 79.6%, perfectly matching the process design value, with an enrichment factor of 45 times; the enrichment factors perfectly match the values ​​above.

[0026] II. Experimental verification of the building material utilization of silicon-titanium slag, including verification of strength, safety, and usability. (B1) The mechanical strength properties of the silicon-titanium slag in Examples 1-3 were verified. The specific experimental method was as follows: the silicon-titanium slag samples were dried at 105℃ to constant weight, ground through an 80-mesh sieve, and replaced with P·O42.5 cement at a mass ratio of 30% to prepare standard cement mortar specimens of 40mm×40mm×160mm. The specimens were cured in a standard curing environment (temperature 20±1℃, relative humidity ≥95%) for 28 days. The compressive strength of the mortar specimens was tested using a microcomputer-controlled electro-hydraulic servo universal pressure testing machine according to the method specified in GB / T17671-2021 to verify its mechanical compatibility as a building material admixture. (B2) Environmental safety verification was conducted on the silicon-titanium slag from Examples 1-3, specifically as follows: Heavy metal leaching toxicity testing: Leachate from silicon-titanium slag was prepared according to the sulfuric acid-nitric acid method specified in HJ / T299-2007. ICP-MS was used to detect Pb, Cd, As, and Cr in the leaching solution. 6+ The concentrations of characteristic heavy metals such as Hg are compared with the limits specified in GB5085.3-2007 to determine whether there is a risk of leaching toxicity.

[0027] Radionuclide limit detection: The specific activities of radium-226, thorium-232, and potassium-40 in silicon-titanium slag were detected using a low-background gamma-ray spectrometer according to the method specified in GB6566-2010, and the internal exposure index I was calculated. a External radiation index I γ To determine whether the radioactive safety requirements for building materials are met; (B3) The building material usability verification of the silicon-titanium slag from Examples 1-3 was conducted. Specifically, according to the method specified in GB / T14684-2022, the core indicators of the silicon-titanium slag, such as particle size distribution, mud content, mud lump content, mica content, and light material content, were tested to determine whether they met the technical requirements for construction sand and to verify their engineering usability as fine aggregate for construction and active admixture for cement. The average 28-day compressive strength of the mortar from Examples 1-3 was 32.5 MPa, and the heavy metal leaching concentration was far below the limit of GB5085.3-2007, with no radioactive risk, thus meeting the requirements for building material utilization.

[0028] III. Experimental verification of sodium salt recovery for industrial salt, including verification of purity, impurities, and usability. (C1) The purity of the main components of the silicon-titanium slag from Examples 1-3 was determined, specifically as follows: After dissolving the samples from Examples 1 to 3 in carbon dioxide-free water, the samples were titrated with silver nitrate standard solution to the endpoint using potassium chromate as an indicator. The mass fraction of sodium chloride (NaCl) in the samples was calculated, and the product grade was determined by comparing with the national standard. (C2) The impurity content of the silicon-titanium slag from Examples 1-3 was tested, specifically as follows: Strictly following the national standard methods, the core impurity indicators were tested sequentially: Moisture: Gravimetric method, the sample was dried to constant weight in a 105℃ electric thermostatic drying oven, and the moisture mass fraction was calculated; Water-insoluble matter: Gravimetric method, the sample was dissolved in water, filtered through quantitative filter paper, the filter residue was dried to constant weight, and the mass fraction of water-insoluble matter was calculated; Total calcium and magnesium ions: EDTA complexometric titration method, in an ammonia-ammonium chloride buffer system, using Chrome Black T as an indicator for titration calculation; Sulfate ions: Barium sulfate gravimetric method, the barium chloride precipitate was ignited and weighed for calculation; Heavy metals (calculated as Pb): Atomic absorption spectrophotometry was used to detect the content of lead in the sample.

[0029] (C3) The industrial usability of the silicon-titanium slag from Examples 1-3 was verified. The test items included sodium chloride (NaCl) mass fraction, moisture mass fraction, water-insoluble matter mass fraction, total calcium and magnesium ions, sulfate ion mass fraction, and heavy metals (as Pb). The sodium chloride content of the recovered industrial salt from Examples 1-3 was 93.2%, 92.5%, and 93.0%, respectively. The moisture mass fraction, water-insoluble matter mass fraction, total calcium and magnesium ions, sulfate ion mass fraction, and heavy metals (as Pb) were all less than the limits of the secondary grade standard for sun-dried industrial salt. It meets the secondary grade standard of "GB / T5462-2015 Industrial Salt" and can be directly reused in industrial production scenarios.

[0030] The above description is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any other way. Any person skilled in the art may make changes or modifications to the above-disclosed technical content to create equivalent embodiments that can be applied to other fields. However, any simple modifications, equivalent changes, and modifications made to the above embodiments based on the technical essence of the present invention without departing from the scope of the present invention shall still fall within the protection scope of the present invention.

Claims

1. A method for recycling red mud, characterized in that, The specific steps are as follows: Step 1: Red mud pretreatment: The red mud is sent into a three-stage countercurrent washing tank and washed in a three-stage countercurrent condition at 60-80℃ and a liquid-solid ratio of 4:

1. After being treated by a thickener, the pH of the red mud at the outlet is controlled to be less than 10.0 to obtain pretreated red mud. Step 2: Hydrochloric acid leaching: The pretreated red mud is fed into an enamel-lined reactor, and leached for 15 minutes in a hydrochloric acid solution with a concentration of 4-5 mol / L. Then, a polyquaternary ammonium salt silicate inhibitor with a mass ratio of 0.08% is added, and leaching continues until the total leaching time reaches 100 minutes. The final pH is controlled at 0.8 to obtain the leachate. The hydrochloric acid solution includes hydrochloric acid and recycled hydrochloric acid; the mass fraction of hydrochloric acid and recycled hydrochloric acid is 14%-17%. Step 3: Solid-liquid separation: The leachate is sent to a plate and frame filter press for filtration; a three-stage countercurrent washing is carried out under a pressure of 0.7 MPa to obtain a filtrate rich in aluminum, iron and rare earth elements and a silicon-titanium slag filter residue. Step 4: Solvent extraction to separate iron: The filtrate is sent to a three-stage countercurrent mixing and clarifying extraction tank. An N503 / sulfonated kerosene mixture with a volume ratio of 3:7 is used as the extraction system, and the ratio of organic phase to aqueous phase is controlled at 1:2.5, with a mixing time of 3 min. During extraction, the acidity of the aqueous phase feed is maintained at 4.0~4.5 mol / L HCl. Iron-loaded organic phase and aluminum-containing rare earth raffinate are obtained. Pure water is used as the back-extraction agent, and the ratio of organic phase to aqueous phase is controlled at 3:1 to back-extract the iron-loaded organic phase, yielding a pure FeCl3 solution. Step 5: Selective precipitation and separation of rare earth elements: The aluminum-containing rare earth raffinate obtained in Step 4 is concentrated to make the aluminum chloride content ≥8%; then oxalic acid or carbonate is added as a precipitant to adjust the pH to 4.0~4.5, and the mixture is heated and stirred to allow the rare earth elements to precipitate selectively. After filtration, rare earth precipitate and impurity-removed aluminic mother liquor were obtained; Step 6: Preparation of polyaluminum chloride: Pump the purified aluminic mother liquor into the polymerization reactor, heat it to 80~90℃, add aluminate or alkaline solution to carry out the polymerization reaction, control the alkalinity at 50%~65%, and obtain liquid polyaluminum chloride product; Step 7: Iron oxide preparation and hydrochloric acid recovery: The back-extraction liquid of the pure FeCl3 solution obtained in Step 4 is sent to a spray-type roasting furnace for roasting; the furnace temperature is controlled at 410~450℃ and the atomization pressure is 0.5MPa to obtain pigment-grade Fe2O3 solid powder and HCl gas. The HCl gas is absorbed by a graphite falling film absorption tower to generate regenerated hydrochloric acid with a mass fraction of 20%~22%, which is returned to Step 2 for recycling.

2. The method for recycling red mud according to claim 1, characterized in that, In step two, the leaching rates of Al, Fe, and rare earth elements are >85%, Fe >85%, and rare earth elements >75%.

3. The method for recycling red mud according to claim 1, characterized in that, In step two, the solid-liquid ratio of hydrochloric acid to pretreated red mud is 5:1, and the leaching temperature is 70~80℃.

4. The method for recycling red mud according to claim 1, characterized in that, In step four, the extraction system further includes one or more of the following: tributyl phosphate, dimethylheptyl methylphosphonate, di(2-ethylhexyl)phosphonic acid, 2-ethylhexyl phosphate mono-2-ethylhexyl ester, and methyl isobutyl ketone.

5. The method for recycling red mud according to claim 1, characterized in that, In step six, the alkaline solution is either the alkaline wastewater generated during the rare earth precipitation in step five, or an added sodium carbonate solution or sodium hydroxide solution with a mass fraction of 10%.

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