A process for preparing iron oxide red from high-iron red mud and extracting titanium dioxide.

Abstract

This invention belongs to the field of solid waste resource recycling technology in the metallurgical industry. It discloses a process for preparing iron oxide red and extracting titanium dioxide from high-iron red mud. The process involves mixing high-iron red mud with an alkaline agent, followed by water leaching to obtain a water-leached solution and a water-leached residue. The water-leached solution undergoes desilication and carbon removal treatment to obtain aluminum hydroxide and sodium carbonate solutions. The water-leached residue is then acid-leached with sulfuric acid; silicon is recovered as silica gel, while aluminum and iron enter the solution and undergo polymerization to prepare a water purification agent. The acid-leached residue is treated with a high-temperature heap leaching-water leaching process using concentrated sulfuric acid to achieve efficient dissolution and separation of titanium. The titanium-iron mixed solution is subjected to stepwise crystallization control, successively precipitating ferric sulfate and ferrous sulfate crystals, which are then pyrolyzed to generate high-purity iron oxide red. Finally, the mother liquor is subjected to titanium precipitation and pyrolysis to obtain titanium dioxide. This invention achieves efficient separation and high-value utilization of all components of iron, titanium, aluminum, and silicon in red mud. The process flow is rationally cyclical, resulting in high-value-added products with low energy consumption and no secondary pollution.

Description

Technical Field

[0001] This invention belongs to the field of solid waste resource recycling technology in the metallurgical industry, specifically involving a process for preparing iron oxide red from high-iron red mud and extracting titanium dioxide. Background Technology

[0002] Red mud is a solid waste generated during the smelting of alumina from bauxite. Large stockpiles of red mud not only occupy land resources, but their high alkalinity and heavy metal content also easily pollute the soil and groundwater, posing a serious threat to the ecological environment.

[0003] High-speed rail red mud is rich in valuable elements such as iron, titanium, aluminum, and silicon, and has significant comprehensive recycling value. Chinese patent CN114212809A discloses a novel low-temperature leaching process for recovering alumina and iron oxide from red mud, which boasts good red mud settling performance and significant emission reduction. However, the process is complex, and other valuable elements such as titanium are difficult to recover effectively, resulting in limited resource utilization. Chinese patent CN112095009A relates to a process for leaching calcium, sodium, iron, aluminum, and rare earth elements from red mud using hydrochloric acid, but it requires at least three leaching devices, placing high demands on equipment and leading to high production costs. Chinese patent CN108031546A discloses a method for recovering iron from red mud, with a large processing capacity, high iron concentrate grade, high recovery rate, low energy consumption, and low cost. However, using red mud for building material production cannot effectively extract valuable metal elements, only achieving low-value utilization of red mud with limited economic benefits. Chinese patent CN109250741A discloses a method for the comprehensive utilization of high-iron red mud. Through a process of pre-iron removal, low-temperature sintering, leaching, and dealkali removal, it simultaneously recovers iron and aluminum and prepares calcium silicate materials, achieving multi-component recovery. However, this process still requires a pre-iron removal step, making the process relatively long and failing to achieve efficient separation and enrichment of iron and titanium. Chinese patent CN119061267A discloses a method for separating titanium from red mud. It involves pre-treating the red mud with carbonization and dealkali removal to obtain carbonized red mud; then pre-treating the carbonized red mud with acid leaching and microbubble flotation to obtain titanium concentrate and red mud tailings. This method can selectively extract titanium from red mud, achieving the recovery of titanium resources and the preparation of titanium concentrate, but it fails to achieve precise separation of iron and titanium.

[0004] Red mud has a complex composition and stable phase structure, with elements intertwined and embedded, making separation difficult. Existing recovery technologies have many limitations in separating iron and titanium, and traditional separation methods struggle to achieve efficient separation. Wet leaching processes consume large amounts of acids and alkalis, increasing production costs and potentially causing secondary pollution. While some pyrometallurgical processes can enrich iron, they are energy-intensive and generate significant amounts of waste gas and residue, polluting the environment. Furthermore, processes that extract only iron or titanium cannot achieve full utilization of red mud components, resulting in low resource utilization and limited economic benefits.

[0005] Therefore, developing a resource utilization technology that can efficiently separate iron and titanium from red mud and simultaneously achieve high-value utilization of components such as aluminum and silicon, with a simple process flow, low energy consumption, and environmental friendliness, has become a key challenge that urgently needs to be overcome in the field of comprehensive utilization of red mud. This invention aims to provide an innovative solution to the core technical bottleneck of iron and titanium separation, based on previous research. Summary of the Invention

[0006] Based on the above-mentioned situation, this invention provides a process for preparing iron oxide red and extracting titanium dioxide from high-iron red mud. By employing a combination of sulfuric acid heap leaching and stepwise crystallization by iron salt reduction, the Al and Si elements, especially Fe and Ti, in the red mud can be efficiently separated and converted into high-value products such as iron oxide red, water purification agents, silica gel, and aluminum hydroxide. Compared with existing technologies, this invention can efficiently separate iron and titanium elements, and has significant advantages such as high resource utilization, high product added value, low energy consumption, and no secondary pollution, realizing the full-component recovery and high-value utilization of red mud.

[0007] To achieve the above objectives, the technical solution of the present invention is as follows: This invention provides a process for preparing iron oxide red from high-iron red mud and extracting titanium dioxide, comprising the following steps: (1) Preparation of raw material and clinker: The high-iron red mud and alkali agent are mixed in a mass ratio of 1:2-3 and ground into raw material of 150-300 mesh. The raw material is then calcined and activated at 800-1000℃ for 0.5-1h to obtain clinker; (2) Water extraction: The clinker and water are mixed and heated in a water bath to extract water, resulting in water extract and water residue; the water extract is circulated to extract clinker multiple times to obtain circulating water extract; (3) Desilication and aluminum extraction: Heat the circulating water leaching solution to 90℃-110℃, add desilication agent to carry out desilication reaction, filter to obtain desilication solution, heat the desilication solution to 60-90℃, add seed aluminum hydroxide and stir, and at the same time pass CO2 gas to carry out carbon separation reaction. After the reaction is completed, the solid and liquid are separated, and the precipitate is washed to obtain aluminum hydroxide and sodium carbonate solution. (4) Acid leaching: The water leaching residue obtained in step 2 is mixed with a sulfuric acid solution of 1-3 mol / L. The liquid-solid ratio of the sulfuric acid solution to the water leaching residue is 4-6 mL / g. The mixture is stirred and acid leached at room temperature for 5-15 min. After filtration and separation, the acid leaching solution and acid leaching residue are obtained by circulating the leaching solution. (5) Desiliconization and preparation of water purification agent: Heat the acid leaching solution to 80-100℃ and keep it warm. After aging for 2-5 hours, filter to obtain silica gel and leachate A containing aluminum sulfate and ferric sulfate. Add lime milk to leachate A to adjust the pH and control the polymerization reaction. After filtration, gypsum and liquid water purification agent are obtained.

[0008] (6) Heap leaching for iron and titanium: The acid leaching residue obtained in step 4 is mixed with a sulfuric acid solution with a concentration of 14-18 mol / L at a solid-liquid ratio of 1g:2-2.5mL, and then continuously heated at 200℃-320℃ for heap leaching for 1-5h; water is added to the mixed solution after heap leaching at a mass ratio of 1:1 for water leaching, and solid-liquid separation is performed to obtain leachate B and titanium concentrate.

[0009] (7) Iron oxide red is prepared by ferric sulfate crystals: the leachate B is evaporated and concentrated, cooled to room temperature and crystallized to obtain ferric sulfate crystals and filtrate A; the ferric sulfate crystals are pyrolyzed at 650-800℃ for 30-90 min to obtain iron oxide red.

[0010] (8) Preparation of iron red from ferrous sulfate heptahydrate crystals: Add excess reduced iron powder to the obtained filtrate A. Based on the concentration of ferric iron in filtrate A, the amount of reduced iron powder added is 1.1-1.4 times the amount of iron powder required to reduce ferric iron to ferrous iron. Filter to separate filtrate B and unreacted reduced iron powder. Evaporate B to concentrate and cool to room temperature to obtain filtrate C and ferrous sulfate heptahydrate crystals. Pyrolyze the ferrous sulfate heptahydrate crystals at 700-800℃ to obtain iron oxide red. (9) Titanium dioxide preparation by metatitanic acid: filtrate C and slightly boiling water with a volume ratio of 1:3-5 are mixed to precipitate titanium dioxide and obtain metatitanic acid precipitate. The precipitate is then pyrolyzed at 500-650℃ to obtain titanium dioxide.

[0011] Preferably, in step (1), the iron-to-aluminum ratio (Fe2O3 / Al2O3) in the high-iron red mud is ≥2, and the alkali is sodium carbonate or sodium hydroxide, more preferably sodium carbonate; the main reaction equations involved in the conversion of iron oxides and decomposition of aluminosilicates in the red mud are: Preferably, in step (2), the water leaching temperature is 60-90℃, the time is 10-30 min, and the solid-liquid ratio of clinker to water is 1g:2-3mL; the water leaching solution is circulated to leach the clinker 1-3 times, so that sodium is enriched in the circulating water leaching solution, and the concentration of sodium aluminate in the circulating water leaching solution is 80-180g / L based on alumina, before entering the next step of aluminum extraction; the separated water leaching residue is used as the raw material for acid leaching in step (4). Specifically, controlling the temperature to 60-90℃ can ensure efficient aluminum dissolution. By circulating the leaching solution 1-3 times, the concentration of sodium aluminate in the water leaching solution is increased to 80-180g / L, while reducing aluminum residue in the residue. During the water leaching process, aluminum and sodium in the clinker mainly enter the solution in the form of soluble sodium salts, while iron is insoluble in alkaline media and is enriched in the water leaching residue. The main reactions are as follows: Preferably, in step (3), the desilication agent is one or more of calcium oxide, lime milk, or Bayer red mud (solid waste generated after alumina extraction from bauxite using the Bayer process).

[0012] Specifically, the silicon concentration in the circulating water leaching solution is first determined. Based on the theoretical amount of silicon to be precipitated, calcium oxide, lime slurry, or red mud are added in proportion. Sodium silicate in the circulating water leaching solution reacts to form calcium silicate precipitate, thus removing silicon impurities. Alternatively, SiO2 in the sodium aluminate solution may precipitate as hydrated sodium aluminosilicate to obtain a desilication solution. The precipitate is primarily calcium silicate, and the main reaction equation is as follows: When Bayer red mud is added and stirred under pressure, the following reaction occurs: 1.7Na2SiO3+2NaAl(OH)4→Na2O·Al2O3·1.7SiO2·nH2O+3.4NaOH; Preferably, at the endpoint of the carbonation reaction, the aluminum ion concentration in the solution, calculated as alumina, is 1-10 g / L. The obtained aluminum hydroxide can be further pyrolyzed to prepare alumina. The sodium carbonate solution can be returned to step (1) as an alkali. The main reaction equation is: Preferably, in step (4), the acid leaching solution obtained after filtration is first used to extract silica to prepare silica gel, and then used to prepare a water purification agent. The acid leaching residue is then used for heap leaching to extract iron and titanium. The main chemical reactions involved in the above acid leaching process are as follows: Fe2O3+3H2SO4→Fe2(SO4)3+3H2O; Al2O3+3H2SO4→Al2(SO4)3+3H2O; SiO2 + 2H2O → H4SiO4; Preferably, in step (5), the main chemical reactions involved in the heat preservation and aging process are: The obtained silica gel can be further washed to obtain pure silica gel.

[0013] Preferably, in step (5), the sulfuric acid concentration of the desilication leachate A is adjusted to 1-3 mol / L, and the process can be repeated in step (4) to circulate the water leaching residue 1-3 times, so that aluminum sulfate and ferric sulfate are enriched in the acid leaching solution. More preferably, the number of cycles is 1-3, and the aluminum sulfate and ferric sulfate content, calculated as Al2O3, reaches 10-13%. Alternatively, lime milk can be added to leachate A for repolymerization, and the pH value can be controlled to 1.5-4 to obtain gypsum and liquid water purification agent. In this step, silicon and aluminum in the high-iron red mud are completely leached and extracted, and some iron elements are also dissolved and utilized.

[0014] Preferably, in step (6), titanium is dissolved by high-temperature sulfuric acid heap leaching: TiO2 + H2SO4 → Ti(SO4)2 + H2O; the remaining iron in the acid leaching residue is in the form of sulfate in the solution: Fe2O3 + 3H2SO4 → Fe2(SO4)3 + 3H2O, which is used for crystallization in the subsequent steps. Concentrated sulfuric acid can only dissolve TiO2 under high temperature conditions, and the reaction requires continuous heating to convert the titanium in the acid leaching residue (mainly containing TiO2 and Fe2O3) into soluble Ti(SO4)2. The titanium sulfate generated by the slow reaction is easily hydrolyzed and may reform TiO2 precipitate. Water leaching is used to extract unreacted TiO2, Fe2O3, etc., which are chemically inert and do not dissolve. They are separated by filtration to obtain titanium concentrate and leachate B containing Ti(SO4)2 and Fe2(SO4)3.

[0015] Preferably, the leachate B (containing Fe) is first concentrated by heating and evaporation. 3+ Evaporation and concentration to 50-70% of the volume increases the Fe content in the leachate. 3+ The concentration promotes the supersaturation precipitation of ferric sulfate (Fe2(SO4)3). Afterwards, the solution is cooled to room temperature or rapidly cooled in an ice-water bath to crystallize, resulting in pale yellow or white ferric sulfate crystals (Fe2(SO4)3·xH2O). Slow cooling yields larger crystals with fewer impurities. The pyrolysis reaction equation is: The resulting iron oxide red can be used in pigments, ceramics, and lithium battery cathode materials.

[0016] Preferably, excess reduced iron powder is added in batches to the filtrate A obtained in step (7) to remove Fe from the filtrate A. 3+ (Incompletely crystallized Fe2(SO4)3) is reduced to Fe 2+ Filter while hot to avoid crystallization of ferrous sulfate upon cooling. The main reaction equation is: 2Fe 3+ +Fe→3Fe 2+ .

[0017] Specifically, due to trivalent iron (Fe) 3+ Its weaker electronic structure and larger charge make it more susceptible to damage than divalent iron (Fe²⁺). 2 + It is more easily hydrolyzed. Reducing ferric iron to ferrous iron before stepwise crystallization before titanium precipitation allows for efficient separation of iron and titanium. 3 + Under heating conditions, it readily hydrolyzes to form ferric hydroxide (Fe(OH)3) precipitate, and reduction avoids the problem of iron-titanium co-precipitation. Furthermore, by fractional crystallization, ferrous iron is recovered separately as ferrous sulfate, and then pyrolyzed to prepare iron oxide red, achieving secondary iron recovery while ensuring the purity of the titanium product.

[0018] Preferably, filtrate B is evaporated and concentrated to 60-80%, then slowly cooled to room temperature to precipitate light green FeSO4·7H2O crystals. Ferrous sulfate crystals are then pyrolyzed at 700-800℃, with the main reaction equation being: Iron oxide red was obtained.

[0019] Preferably, the obtained filtrate C contains Ti(SO4)2. Adding slightly boiling water to precipitate titanium yields metatitanic acid. The main reaction equation is as follows: Titanium dioxide is produced by pyrolysis of metatitanic acid at 500-650℃. The main reaction equation is: H2TiO3→TiO2+H2O.

[0020] This invention successfully achieves the efficient separation and recovery of iron, titanium, and other elements from high-iron red mud through a staged selective leaching process. First, in the activation roasting-water leaching stage, iron oxides are selectively enriched in the leaching residue, achieving preliminary separation of iron from aluminum and silicon. Subsequently, through sulfuric acid heap leaching, titanium is dissolved to form soluble Ti(SO4)2, which enters the solution, while iron remains in the residue or crystallizes out as sulfate. The iron component is converted into ferric sulfate and ferrous sulfate through stepwise evaporation and crystallization, and finally pyrolyzed to obtain high-purity iron oxide red (Fe2O3, purity ≥98%). The titanium component is precipitated through a titanium precipitation process, causing Ti(SO4)2 in the solution to form metatitanic acid (H2TiO3) precipitate, which is then pyrolyzed to obtain titanium dioxide (TiO2, purity up to 99.6%). This process effectively solves the problems of iron-aluminum co-solubility and iron-titanium separation in traditional methods, and realizes the high-value utilization of all components of red mud. It also produces water purification agents, silica gel and aluminum hydroxide, etc., with advantages such as high resource utilization rate, high product added value and environmental friendliness. It provides an efficient and feasible technical path for the resource utilization of metallurgical solid waste such as red mud.

[0021] The advantages and beneficial effects of this invention are: (1) A synergistic process of alkaline activation for aluminum and silicon extraction and acidic heap leaching-reduction-stepwise crystallization for iron and titanium extraction is adopted. Aluminum / silicon is preferentially dissolved, while iron is retained in the slag. Acid leaching selectively dissolves titanium and iron, and iron is separated by stepwise crystallization to remove most of the Fe. 3+ Then reduce it to Fe 2+ Post-crystallization separation eliminates the co-precipitation of iron and titanium at the source, successfully solving the problem of deep separation of iron and titanium in traditional wet processes.

[0022] (2) The sulfuric acid solution is recycled, and all titanium concentrate and silica gel are fully utilized. There is no harmful waste discharge throughout the process. High value-added products such as iron oxide red, titanium dioxide and silica gel are produced in conjunction with the production of these products, which improves the overall income and reduces the environmental risks of red mud storage.

[0023] (3) The titanium dioxide obtained by this invention has a TiO2 purity of ≥99.6%, which meets the pigment grade standard; the Fe2O3 purity in iron oxide red is ≥98%, realizing the simultaneous and efficient separation of iron and titanium. Attached Figure Description

[0024] Figure 1 This is a flowchart of the process for preparing iron oxide red and extracting titanium dioxide from high-iron red mud according to the present invention. Detailed Implementation

[0025] To better explain and facilitate understanding of the present invention, the present invention will be described in detail below with reference to the accompanying drawings and specific embodiments.

[0026] Example 1 The high-iron red mud used in this embodiment all came from Shaanxi Longmen Iron & Steel (Group) Co., Ltd. The main chemical components of the red mud were determined by chemical analysis to be: Al2O3: 14.33%, SiO2: 4.43%, Fe2O3: 64.84%, TiO2: 3.58%, CaO: 1.51%, MgO: 0.11%, Na2O: 2.84%, K2O: 0.20%, P2O5: 0.23%, MnO2: 0.07%; the alkali agent was sodium carbonate.

[0027] Combination Figure 1 As shown, this embodiment provides a process for preparing iron oxide red and extracting titanium dioxide from the above-mentioned high-iron red mud, including the following steps: Step 1: Mix red mud and sodium carbonate in a mass ratio of 1:2.5 and grind them into raw material with a mesh size of 200. Then, calcine the raw material at 900℃ for 0.5 hours to obtain cooked material.

[0028] Step 2: Mix the clinker and water at a solid-liquid ratio of 1g:2mL, heat in a water bath to 60℃, stir and dissolve in water for 15min, separate the solid and liquid, and obtain water leaching solution and water leaching residue; circulate the water leaching solution to leach the clinker twice to obtain circulating water leaching solution, and the concentration of sodium aluminate in the circulating water leaching solution is 80g / L based on alumina.

[0029] Step 3: Heat the circulating water leaching solution obtained in Step 2 to 90°C. First, determine the silicon concentration in the circulating water leaching solution. Based on the theoretical silicon precipitation amount, add calcium oxide in proportion to carry out the desilication reaction. Filter to obtain the desilication solution. Heat the desilication solution to 90°C, add aluminum hydroxide seed crystals and stir, while simultaneously introducing CO2 gas to carry out the carbon separation reaction. After the reaction is complete, separate the solid and liquid, wash the precipitate to obtain aluminum hydroxide and sodium carbonate solutions. The sodium carbonate solution is returned to Step 1 as an alkali.

[0030] Al(OH)3 was pyrolyzed under certain temperature conditions to obtain Al2O3 product. The Al element in the red mud was recovered in the form of alumina product with a yield of 62.32%. The chemical composition of the obtained Al2O3 was analyzed, and the contents of the main components are shown in Table 1.

[0031] Table 1. Main chemical composition of alumina

[0032] The chemical composition of the product meets the industry standard GB / T24487-2022. Referring to the national standard for metallurgical grade alumina, the alumina product exceeds the requirements of AO-G grade.

[0033] Step 4: Mix the water leaching residue obtained in Step 2 with a 2 mol / L sulfuric acid solution at a liquid-to-solid ratio of 4 mL / g, stir and acid leaching at room temperature for 10 min, then circulate and filter to separate the acid leaching solution and acid leaching residue. Step 5: Heat the acid leaching solution obtained in Step 4 to 90°C and keep it at that temperature. After aging for 2 hours, filter to obtain silica gel and leachate A containing aluminum sulfate and ferric sulfate. Add lime milk to leachate A containing aluminum sulfate and ferric sulfate to adjust the pH to 2. Perform a temperature-controlled polymerization reaction. After filtration, obtain gypsum and liquid aluminum ferric sulfate water purification agent.

[0034] Analysis of the liquid aluminum ferric sulfate water purification agent showed that the recovery rates of Al2O3 and Fe2O3 were 33.65% and 31.58%, respectively. The main components of the water purification agent are shown in Table 2. The chemical composition of the product meets the requirements of the (HG / T 5565-2019) standard. Referring to Table 3, the performance indicators of the water purification agent, the test results indicate that the water purification agent meets the industry standard for water purification agents.

[0035] Table 2 Chemical Composition of Water Purification Agent (%)

[0036] Table 3 Performance Indicators of Water Purification Agents (%)

[0037] After silicon extraction, the solution was polymerized and separated to obtain gypsum products. The main components after removing loss on ignition are shown in Table 4.

[0038] Table 4 Chemical Composition of Gypsum (%)

[0039] Step 6: Mix the acid leaching residue obtained in Step 4 with a sulfuric acid solution with a concentration of 14 mol / L at a solid-liquid mass ratio of 1 g: 2 mL, and continue heating at 320℃ for heap leaching for 4 h; add water to the mixed solution after heap leaching at a mass ratio of 1:1 for water leaching, and separate the solid and liquid to obtain leachate B and titanium concentrate.

[0040] Step 7: Preparation of iron oxide red from ferric sulfate crystals: The leachate B after heap leaching is evaporated and concentrated to 50% of its original volume, then cooled to room temperature for crystallization to obtain ferric sulfate crystals and filtrate A; the ferric sulfate crystals are pyrolyzed at 650℃ for 90 min to obtain iron oxide red.

[0041] Step 8: Add excess reduced iron powder to filtrate A obtained in step 7. Based on the concentration of ferric iron in filtrate A, the amount of reduced iron powder added is 1.2 times the amount of iron powder required to reduce ferric iron to ferrous iron. Filter to separate filtrate B and unreacted reduced iron powder. Evaporate filtrate B to 70% and cool to room temperature to obtain filtrate C and ferrous sulfate heptahydrate crystals. Pyrolyze the ferrous sulfate heptahydrate crystals at 750℃ for 30 min to obtain iron oxide red. Step 9: Mix filtrate C (volume ratio 1:4) with slightly boiling water to precipitate titanium dioxide, obtaining metatitanic acid precipitate. Pyrolyze the precipitate at 650℃ to obtain titanium dioxide. The metatitanic acid obtained after titanium precipitation is then pyrolyzed to obtain titanium dioxide.

[0042] Example 1: Ti element in red mud was recovered in the form of titanium dioxide product, with a recovery rate of 70.43%.

[0043] Example 2 The high-speed iron red mud used in this embodiment all came from Weiqiao high-speed iron red mud in Shandong. The main chemical components of the red mud were determined by chemical analysis to be: Al2O3: 17.99%, SiO2: 8.66%, Fe2O3: 49.53%, TiO2: 6.58%, CaO: 2.19%, Na2O: 5.64%, and loss on ignition: 9.92%; the alkali agent was sodium carbonate.

[0044] Combination Figure 1 As shown, this embodiment provides a process for preparing iron oxide red and extracting titanium dioxide from the above-mentioned high-iron red mud, including the following steps: Step 1: Mix red mud and sodium carbonate in a mass ratio of 1:2 and grind them into raw material of 200 mesh. Then, roast the raw material at 800℃ for 1 hour to obtain cooked material.

[0045] Step 2: Mix the clinker and water at a solid-liquid ratio of 1g:3mL, heat in a water bath to 90℃, stir and dissolve in water for 30min, separate the solid and liquid, and obtain water leaching solution and water leaching residue; circulate the water leaching solution to leach the clinker 3 times to obtain circulating water leaching solution, and the concentration of sodium aluminate in the circulating water leaching solution is 54 g / L based on alumina.

[0046] Step 3: Heat the circulating water leaching solution obtained in Step 2 to 100℃, add calcium oxide in proportion to carry out a desilication reaction, and filter to obtain a desilication solution. Heat the desilication solution to 60℃, add aluminum hydroxide and stir, while simultaneously introducing CO2 gas to carry out a carbon separation reaction. After the reaction is complete, separate the solid and liquid, wash the precipitate to obtain aluminum hydroxide and sodium carbonate solutions. The sodium carbonate solution is returned to Step 1 as an alkali.

[0047] Al(OH)3 is pyrolyzed under certain temperature conditions to obtain Al2O3 product. The chemical composition of the obtained Al2O3 is analyzed, and the contents of the main components are shown in Table 5.

[0048] Table 5. Main chemical composition of alumina

[0049] Step 4: Mix the water leaching residue obtained in Step 2 with a 1.5 mol / L sulfuric acid solution at a liquid-to-solid ratio of 4 mL / g, stir at room temperature for 15 min, and then filter to separate the acid leaching solution and acid leaching residue.

[0050] Step 5: Heat the acid leaching solution obtained in Step 4 to 90°C and keep it at that temperature. After aging for 4 hours, filter to obtain silica gel and leachate A containing aluminum sulfate and ferric sulfate. Add lime milk to leachate A to adjust the pH to 4, control the temperature for polymerization reaction, and filter to obtain gypsum and liquid water purification agent.

[0051] The liquid aluminum ferric sulfate water purification agent was analyzed, and its main components are shown in Tables 6-7. The chemical composition of the product meets the requirements of the standard (HG / T 5565-2019). The performance indicators of the water purification agent are shown in Table 7. The test results show that the water purification agent meets the industry standard for water purification agents.

[0052] Table 6 Chemical Composition of Water Purification Agent (%)

[0053] Table 7 Performance Indicators of Water Purification Agents (%)

[0054] After silicon extraction, the solution was polymerized and separated to obtain gypsum products. The main components after removing loss on ignition are shown in Table 8.

[0055] Table 8 Chemical Composition of Gypsum (%)

[0056] Step 6: Mix the acid leaching residue obtained in Step 4 with sulfuric acid of concentration 18 mol / L at a solid-liquid ratio of 1 g: 2.5 mL, and heat to 250℃ for heap leaching for 3 h; add water to the mixed solution after heap leaching at a mass ratio of 1:1 for water leaching, and separate the solid and liquid to obtain leachate B and titanium concentrate.

[0057] Step 7: The leachate obtained in Step 6 is concentrated to 60% of its original volume by evaporation, then cooled to room temperature for crystallization to obtain ferric sulfate crystals and filtrate A. The ferric sulfate crystals are then pyrolyzed at 700℃ for 60 min to obtain iron oxide red.

[0058] Step 8: Add excess reduced iron powder to filtrate A obtained in Step 7 to reduce ferric iron to ferrous iron. Based on the ferric iron concentration in filtrate A, the amount of reduced iron powder added is 1.2 times the amount required to reduce ferric iron to ferrous iron. Reduce the acid concentration, filter to separate filtrate B and unreacted reduced iron powder. Evaporate and concentrate filtrate B to 70% of its volume and cool to room temperature. Filter to obtain filtrate C and ferrous sulfate heptahydrate crystals. Pyrolyze at 750℃ for 30 min to obtain iron oxide red.

[0059] Step 9: Mix the filtrate C after crystallization filtration at a volume ratio of 1:4 with slightly boiling water to precipitate titanium dioxide, and then pyrolyze the titanium dioxide at 600℃ to obtain titanium dioxide.

[0060] The metatitanic acid obtained after titanium precipitation was pyrolyzed to obtain titanium dioxide. Ti element in red mud was recovered in the form of titanium dioxide product, with a yield of 73.01%. Performance indicators are shown in Table 9.

[0061] Table 9 Performance Indicators of Titanium Dioxide

[0062] The chemical composition meets the requirements of the standard (GB / T1706-2006). Referring to the quality standard of titanium dioxide pigment, the test results show that the titanium dioxide can meet the requirements of titanium dioxide pigment type A1.

[0063] The performance indicators of the iron oxide red obtained in Examples 1 and 2 are shown in Table 10, and the chemical composition is shown in Table 11.

[0064] Table 10 Performance Indicators of Iron Oxide Red

[0065] The chemical composition of iron oxide red pigment meets the requirements of the standard (GB / T1863-2008). Referring to the quality standard of iron oxide red pigment, the test results show that the iron oxide can meet the requirements of Class A red iron oxide red pigment.

[0066] Table 11 Chemical composition of iron oxide red (%)

[0067] Comparative Example 1 The only difference from Example 2 is that step 8 skips the process of adding excess iron powder to the filtrate A obtained in step 7 to reduce ferric iron and prepare ferrous sulfate heptahydrate crystals. Instead, the filtrate A obtained after crystallization and filtration in step 7 is directly mixed with slightly boiling water at a volume ratio of 1:4 to precipitate titanium dioxide. The resulting metatitanic acid precipitate is then pyrolyzed at 600 °C to finally obtain the titanium dioxide product.

[0068] Comparative Example 2 The only difference from Example 2 is that in step 8, based on the concentration of ferric iron in filtrate A, the amount of reduced iron powder added is 80% of the amount of iron powder required to reduce ferric iron to ferrous iron.

[0069] Comparative Example 3 The difference from Example 2 is only in step 8: without stepwise crystallization, excess reduced iron powder is added to the filtrate A obtained in step 7 to reduce the ferric iron to ferrous iron, thereby lowering the acid concentration. The mixture is then filtered to separate filtrate B and excess iron. Filtrate B (volume ratio 1:4) is then mixed with slightly boiling water to precipitate titanium dioxide, resulting in metatitanic acid precipitate. The metatitanic acid is then pyrolyzed at 600°C to obtain titanium dioxide.

[0070] The chemical composition of the titanium dioxide products obtained in the examples and comparative examples was analyzed, and the results are shown in Table 12.

[0071] Table 12 Chemical composition of titanium dioxide (%)

[0072] The titanium dioxide obtained in Comparative Example 1 had a TiO2 content of 89.00%, while the main impurity, Fe2O3, accounted for 9.66%, indicating a severe excess of iron impurities, failing to meet the requirements for Type A1 titanium dioxide pigment. This significantly differs from the TiO2 purity in Example 2 of this invention. This is because residual Fe2O3 in filtrate A occurred during titanium precipitation. 3+ With Ti 4+ Ions undergo hydrolysis simultaneously under heating and dilution conditions. Fe 3 + It readily hydrolyzes to form Fe(OH)3 precipitate, which then co-precipitates with metatitanic acid (H2TiO3) generated from titanium hydrolysis. This co-precipitate cannot be separated by conventional washing, resulting in iron being completely encapsulated in the final titanium dioxide product. The "reduction-stepwise crystallization" method of this invention introduces iron powder to... 3+ Reduced to Fe 2+ This further separates iron and titanium, avoiding the co-precipitation of iron and titanium.

[0073] Comparative Example 2 showed insufficient reduced iron powder, with TiO2 purity of 96.30% and Fe2O3 content of 2.45%. It did not meet the requirements for Titanium Dioxide Pigment Type A1.

[0074] Compared to Comparative Example 1, the purity of titanium dioxide was improved, proving that partial reduction was effective, but not complete. The purity was far lower than the 99.60% of Example 2, mainly because 80% iron powder was added, meaning that filtrate A still contained 20% Fe. 3+ Unreduced. Irreversible chemical coprecipitation occurs in subsequent reactions, where it is encapsulated in metatitanic acid and cannot be removed by washing. This comparative example demonstrates that the reduction step must be thorough and complete; insufficient reducing agent will result in residual Fe. 3+The co-precipitation of these substances prevents the production of high-purity titanium dioxide.

[0075] Comparative Example 3 did not crystallize in steps, and its purity was higher than that of Comparative Examples 1 and 2, thus avoiding Fe... 3+ The problem of chemical coprecipitation exists. However, its purity (98.09%) and iron content (1.20%) are still significantly worse than those of Example 2 (99.60%, 0.16%). Although the reduction step will reduce Fe... 3+ All converted to Fe 2+ However, skipping "stepwise crystallization" means that the concentration of ferrous ions in the pre-precipitation filtrate B is relatively high. In Ti 4+ During the hydrolysis and precipitation of titanium, these Fe 2+ It will be physically adsorbed and encapsulated by the newly generated metatitanic acid, and is difficult to completely remove by washing.

[0076] In summary, neither reduction alone nor stepwise crystallization alone can produce the high-purity titanium dioxide of this invention. The core function of the "stepwise crystallization" step is to significantly reduce the total iron concentration in the solution before titanium precipitation, thereby minimizing Fe emissions at the source. 2+ The potential for product contamination through adsorption and encapsulation pathways. The synergistic technology of reduction-step crystallization produced unexpected synergistic effects, significantly improving the purity of titanium dioxide and iron recovery rate.

[0077] The above description is only a preferred embodiment of the present invention. It should be noted that for those skilled in the art, several modifications and improvements can be made without departing from the inventive concept, and these all fall within the protection scope of the present invention.

Claims

1. A process for preparing iron oxide red from high-iron red mud and extracting titanium dioxide, characterized in that, Includes the following steps: (1) Mix high-iron red mud with an alkaline agent at a mass ratio of 1:2-3 and grind it. Then, calcine and activate it at 800-1000℃ for 0.5-1h to obtain clinker; (2) Mix the clinker with water and perform water extraction to obtain water extract and water residue; circulate the water extract to extract the clinker multiple times to obtain circulating water extract; (3) Heat the circulating water immersion liquid to 90℃-110℃ and add desilication agent to carry out desilication reaction. Filter to obtain desilication liquid. Heat the desilication liquid to 60-90℃, add aluminum hydroxide seed crystals and stir. At the same time, introduce CO2 gas to carry out carbon separation reaction. Solid-liquid separation yields aluminum hydroxide and sodium carbonate solution. (4) Mix the obtained water leaching residue with a 1-3 mol / L sulfuric acid solution, stir and leach at room temperature, and circulate the leaching solution and filter to separate the acid leaching residue; (5) Heat the acid leaching solution to 80-100℃ and keep it at that temperature. After aging for 2-5 hours, filter to obtain silica gel and leachate A. (6) The acid leaching residue is mixed with a sulfuric acid solution with a concentration of 14-18 mol / L at a solid-liquid ratio of 1g:2-2.5mL, and then heated and piled at 200℃-320℃ for 1-5h. After pile leaching, the mixed solution is leached with water, and the solid and liquid are separated to obtain leachate B and titanium concentrate. (7) After evaporation, the leachate B is cooled to room temperature to crystallize, resulting in ferric sulfate crystals and filtrate A. The ferric sulfate crystals are then pyrolyzed to obtain iron oxide red. (8) Add excess reduced iron powder to filtrate A, filter to separate filtrate B and unreacted reduced iron powder, evaporate and concentrate filtrate B and cool to room temperature to obtain filtrate C and ferrous sulfate heptahydrate crystals, and pyrolyze ferrous sulfate heptahydrate crystals to obtain iron oxide red. (9) Mix filtrate C with slightly boiling water at a volume ratio of 1:3-5 to precipitate titanium dioxide, and then pyrolyze the precipitate to obtain titanium dioxide.

2. The process method according to claim 1, characterized in that, The iron-aluminum ratio in the high-iron red mud is ≥2; the alkaline agent is sodium carbonate or sodium hydroxide.

3. The process method according to claim 1, characterized in that, The solid-liquid ratio of the clinker to water is 1g:2-3mL; the water extraction temperature is 60-90℃, the time is 10-30min, and the concentration of sodium aluminate in the circulating water extraction solution is 80-180g / L (calculated as alumina).

4. The process method according to claim 1, characterized in that, The desilication agent is one or more of calcium oxide, lime milk, or red mud.

5. The process method according to claim 1, characterized in that, Step (4) The acid immersion time is 5-15 minutes.

6. The process method according to claim 1, characterized in that, The sulfuric acid concentration of the leachate A is adjusted to 1-3 mol / L, and the water leaching residue is recycled to step (4) 1-3 times. The aluminum ferric sulfate content in the leachate A is 10-13% based on Al2O3. Alternatively, lime milk is added to the leachate A and polymerized again, and the pH value is controlled to 1.5-4 to obtain gypsum and liquid water purification agent.

7. The process method according to claim 1, characterized in that, The pyrolysis temperature of the ferric sulfate crystals is 650-800℃, and the pyrolysis time is 30-90 min.

8. The process method according to claim 1, characterized in that, The pyrolysis temperature for the crystallization of the ferrous sulfate heptahydrate is 700-800℃.

9. The process method according to claim 1, characterized in that, The pyrolysis temperature of the metatitanic acid precipitate is 500-650℃.

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