A method for utilizing valuable elements in vanadium-titanium magnetite concentrate

Abstract

This invention provides a method for utilizing valuable elements in vanadium-titanium magnetite concentrate, comprising: mixing vanadium-titanium magnetite concentrate with a calcifying agent to prepare alkaline green pellets of vanadium-titanium magnetite concentrate; obtaining calcified oxide pellets through calcification and oxidative roasting; obtaining metallized reduced pellets through reduction at high temperature; crushing and grinding; and magnetic separation to obtain reduced iron powder and titanium-containing materials; acid leaching and solid-liquid separation of the titanium-containing materials to obtain a hydrochloric acid leaching solution and titanium middlings; alkaline leaching and solid-liquid separation of the titanium middlings to obtain titanium-rich material; spray-fiber vanadium-chromium powder and hydrochloric acid gas by incineration of the hydrochloric acid leaching solution; washing the vanadium-chromium-chromium powder with water to obtain vanadium-rich vanadium-chromium-chromium powder; selectively acid leaching vanadium with the vanadium-rich vanadium-chromium-chromium powder to obtain a vanadium-containing leaching solution and chromium-rich powder; and reducing and smelting the chromium-rich powder to produce ferrochrome. This invention can cleanly and efficiently extract iron, titanium, vanadium, and chromium elements from vanadium-titanium magnetite concentrate, and the prepared titanium-rich material meets the raw material requirements of the titanium dioxide industry, while also having advantages such as industrial operability and environmental friendliness.

Description

Technical Field

[0001] This invention relates to the fields of mineral processing and metallurgy, and more specifically, to a method for utilizing valuable elements in vanadium-titanium magnetite concentrate. Background Technology

[0002] Vanadium-titanium magnetite is a major and distinctive polymetallic mineral resource in my country, with enormous reserves. It is a polymetallic deposit containing vanadium, titanium, iron, chromium, and other metals, making it highly valuable for comprehensive utilization. my country's proven reserves exceed 10 billion tons, with potential reserves exceeding 30 billion tons, mainly distributed in the Panxi region of Sichuan, Chengde in Hebei, and Chaoyang in Liaoning. The Panxi region is the richest in vanadium and titanium resources in my country, with proven potential reserves exceeding 10 billion tons and existing reserves of 6.73 billion tons, including 14.75 million tons of vanadium (V₂O₅) and 593 million tons of titanium (TiO₂), accounting for 63% and 90.5% of the national vanadium and titanium reserves respectively, ranking third and first in the world. The Panxi region is rich in vanadium-titanium magnetite resources, consisting of four mining areas: Hongge, Taihe, Baima, and Panzhihua. The Hongge mining area, with reserves of approximately 3.6 billion tons, is the largest vanadium-titanium magnetite deposit in the Panxi region. The Honggenan mining area is a large-scale, comprehensive polymetallic deposit with iron as its main component. It boasts a high chromium content, with chromium trioxide reaching up to 1.8% in the concentrate, making it a highly valuable multimetallic deposit for comprehensive utilization. Due to the similar chemical properties of vanadium and chromium, the efficient, economical, and clean extraction and separation of these two minerals from vanadium- and chromium-endowed minerals has always been a global technical challenge and one of the key issues urgently needing to be addressed in the mining and utilization of high-chromium vanadium-titanium magnetite in the Honggenan mining area.

[0003] The core of the research on the comprehensive utilization of vanadium-titanium magnetite concentrate lies in the efficient and synergistic recovery of strategic metals such as iron, titanium, vanadium, and chromium.

[0004] Based on the different metal refining sequences, vanadium extraction processes from vanadium-titanium magnetite can be divided into two categories: vanadium-first, then iron-second, and iron-first, then vanadium-second. The blast furnace method is a typical example of the iron-first, then vanadium-second method. Blast furnace ironmaking remains the mainstream technology in modern ironmaking. Taking Panzhihua Iron and Steel Group as an example, blast furnace smelting of vanadium-titanium magnetite concentrate is mainly for steel production, with vanadium extraction as a secondary objective. The iron recovery rate is around 76%, while the vanadium recovery rate is only around 45%. However, the TiO2 content in the blast furnace slag is only around 22% to 25%, which cannot be effectively utilized and is instead stockpiled.

[0005] The vanadium-to-iron process, through roasting and leaching, dissolves vanadium in solution as a metal salt, achieving separation from other metals in vanadium-titanium magnetite. This method eliminates the need for iron smelting and slag blowing, resulting in high vanadium recovery rates. The tailings can still be used for steelmaking and titanium extraction, achieving efficient mineral utilization. However, the sodium roasting-water leaching vanadium extraction process requires high vanadium grades in the raw materials and presents numerous environmental pollution issues; therefore, this process has been largely phased out. The calcification / blank roasting-acid leaching vanadium extraction process also has drawbacks, such as high CO2 emissions during roasting, high roasting energy consumption, large sulfuric acid requirements, and difficulty in utilizing tailings. Further industrial optimization and improvement are needed.

[0006] Patent CN101906531A discloses a method for extracting iron, vanadium, and titanium from vanadium-titanium magnetite. This method employs a process combining vanadium chlorination roasting with medium-temperature solid-state reduction, achieving selective chlorination and volatilization of vanadium at temperatures between 1100 and 1300 K. Vanadium recovery rates can reach over 95%, iron recovery rates can reach 87%–92%, and the TiO2 content in the high-titanium slag can reach over 51 wt%, significantly improving titanium recovery efficiency. Simultaneously, this process can utilize industrial chlorination waste as a chlorinating agent, achieving resource utilization and harmless treatment of waste. However, its shortcomings are as follows: the chlorination roasting process needs to be carried out under argon protection, which requires high-end equipment and consumes a lot of energy; the chlorinating agent (such as FeCl3) is highly corrosive at high temperatures, which places high demands on the reactor material; the VOCl3 gas generated by chlorination needs to be further treated, and the tail gas purification and chlorine recovery system is complex; in addition, the composition fluctuates greatly when using industrial waste, which affects the stability of the process, and the energy consumption is high in the high-temperature melting stage (1800-1850K). Overall, the industrialization promotion still faces challenges such as equipment durability and economy.

[0007] Patent CN102690944A discloses a method for the comprehensive recovery of vanadium, titanium, and iron from high-vanadium-titanium magnetite. The process employs a flow of sodium / calcification roasting—acid leaching for vanadium extraction—rotary hearth furnace rapid reduction—melting—converter vanadium extraction—vanadium slag return. This process can achieve selective leaching of vanadium at 1100–1250℃, with vanadium recovery rates exceeding 93%, titanium recovery rates exceeding 95%, and iron recovery rates exceeding 93%, significantly improving the comprehensive utilization rate of vanadium and titanium resources. Simultaneously, this process uses a rotary hearth furnace for rapid reduction instead of a traditional rotary kiln, with a reduction time of only 20–50 minutes, effectively avoiding the ring formation problem in rotary kilns and greatly improving production efficiency. Furthermore, the vanadium slag after converter vanadium extraction is returned to the front end for further vanadium extraction, further increasing the overall vanadium recovery rate. However, its shortcomings are as follows: the process is relatively long and involves many steps, requiring significant equipment investment and is only suitable for high-vanadium-titanium magnetite; the acid leaching process uses sulfuric acid, generating acidic wastewater that necessitates comprehensive wastewater treatment facilities, resulting in high environmental costs; the amount of additives (sodium or calcium salts) used is 5%–12% of the ore weight, increasing auxiliary material costs; the TiO2 content in the titanium slag obtained from melting is only about 30%, indicating a low grade, leading to high acid consumption and large amounts of waste acid when used in the sulfuric acid process for titanium dioxide production; the rotary hearth furnace has a reduction temperature of 1250–1380℃ and a high melting temperature, resulting in high overall energy consumption and strict requirements for furnace lining materials. Further optimization of its economic and environmental aspects is needed for industrial-scale promotion.

[0008] Patent CN105112689 A discloses a method for extracting titanium from vanadium-titanium magnetite. This method involves adding sodium or potassium salt additives during the reduction smelting of iron concentrate in an electric furnace. These additives simultaneously reduce slag viscosity, improve slag-iron separation, and convert silicon and aluminum impurities into aluminosilicates soluble in dilute acid. Then, dilute acid leaching is used to remove impurities, yielding a titanium slag product containing >75% TiO2. The titanium recovery rate can reach over 96%, and the metallic iron recovery rate can reach over 98%, nearly three times higher than the traditional blast furnace process. Furthermore, the waste liquid after acid leaching can be used to prepare water purification agents, achieving wastewater resource utilization. The process is clean, with no waste discharge, and all steps utilize current mature industrial technologies and equipment. The process is short, easy to operate, and readily industrialized. However, its shortcomings are as follows: the amount of sodium or potassium salt additives needs to be adjusted according to the silicon and aluminum content of the raw materials, which requires a certain degree of adaptability to the raw materials; the electric furnace reduction smelting temperature is high, the power consumption is large, and the economics may be affected in areas sensitive to electricity prices; the obtained titanium slag has a TiO2 content of >75%, which meets the requirements of acid-soluble titanium slag, but it is still different from the high titanium slag required for chloride process titanium dioxide, and further enrichment is still required if it is used in the chloride process.

[0009] In summary, the following problems exist in the recovery of iron, titanium, and vanadium from vanadium-titanium magnetite concentrate using existing technologies: 1) Existing processes are complex, energy-intensive, and highly polluting, making industrialization difficult; 2) Raw material adaptability is narrow, resulting in low-grade titanium slag, making it difficult to prepare raw materials that meet the requirements of fluidized bed chlorination processes; 3) Other valuable elements cannot be utilized efficiently.

[0010] In view of this, the present invention is hereby proposed. Summary of the Invention

[0011] The purpose of this invention is to address the technical problems of low recovery rate, high energy consumption, and serious environmental pollution in existing processes for extracting iron, titanium, and vanadium from vanadium-titanium magnetite concentrate. This invention provides a method for utilizing valuable elements in vanadium-titanium magnetite concentrate, which can cleanly and efficiently extract iron, titanium, vanadium, and chromium from vanadium-titanium magnetite concentrate. The prepared titanium-rich material meets the raw material requirements of the titanium dioxide industry, and also has the advantages of industrial operability and environmental friendliness.

[0012] In order to achieve the above-mentioned objectives of the present invention, the following technical solution is adopted: A method for utilizing valuable elements in vanadium-titanium magnetite concentrate includes the following steps: Step 1: Mix vanadium-titanium magnetite concentrate with a calcifying agent to prepare alkaline green pellets of vanadium-titanium magnetite concentrate. The alkaline green pellets are then subjected to calcification oxidation roasting to obtain calcified oxide pellets. Step 2: Mix the calcified oxide pellets with the first reducing agent and reduce them at high temperature to obtain metallized reduced pellets. Crush and grind the metallized reduced pellets and then perform magnetic separation to obtain magnetic reduced iron powder and non-magnetic titanium-containing materials. Step 3: Acid leaching and solid-liquid separation of the titanium-containing material yields hydrochloric acid leaching solution and titanium ore. Step 4: The titanium ore is leached with alkali and separated into solid and liquid components to obtain titanium-rich material; Step 5: Spray-fire the hydrochloric acid leaching solution to generate vanadium-chromium-iron powder and hydrochloric acid gas; Step 6: Wash the vanadium-chromium-iron powder with water and separate the solid and liquid to obtain vanadium-chromium-iron powder; Step 7: Selectively acid-leach vanadium with the vanadium-rich ferrochrome powder, and separate the solid and liquid phases to obtain a vanadium-containing leaching solution and ferrochrome-rich powder; Step 8: Mix the chromium-rich iron powder with the second reducing agent and smelt it into ferrochrome.

[0013] Furthermore, the diameter of the alkaline green pellets described in step 1 is 9~16mm.

[0014] Further, the composition of the vanadium-titanium magnetite concentrate mentioned in step 1, by mass percentage, includes TFe 50~60%, TiO2 8~15%, V2O5 0.5~2.0%, SiO2 0.5~5.0%, Cr2O3 1.0~3.0%, and CaO 0.1~5%.

[0015] Further, the calcifying agent in step 1 is at least one of a calcium-containing compound or a calcium-containing mineral; the calcium-containing compound is at least one of CaO, CaSO4, and Ca(OH)2.

[0016] Furthermore, in step 1, the calcination oxidation calcination temperature is 1100~1250℃, and the calcination time is 20min~10h.

[0017] Furthermore, the compressive strength of the calcified oxide pellets after calcination in step 1 is ≥1500N / pellet.

[0018] Furthermore, during the alkaline pelletizing of vanadium-titanium magnetite concentrate in step 1, the amount of calcifying agent added relative to 100 parts of vanadium-titanium magnetite concentrate is calculated using the following formula: Wt Ca(OH)2 =K(123Wt SiO2 +41Wt V2O5 -132Wt CaO ), Among them Wt Ca(OH)2 The amount of calcifying agent added is specified, with an excess coefficient K of 1.2~1.5, Wt. SiO2 Wt V2O5 and Wt CaO The percentage contents of SiO2, V2O5 and CaO in vanadium-titanium magnetite concentrate are respectively. And Wt Ca(OH)2 The amount of calcifying agent added shall not be less than 1.2 parts.

[0019] Furthermore, in step 2, the mass ratio of the calcified oxide pellets to the first reducing agent is 1:0.1 to 1:2.

[0020] Furthermore, in step 2, the first reducing agent is at least one of carbon-based substances, hydrogen, and CO.

[0021] Furthermore, in step 2, the reduction temperature is 900~1300℃, and the reduction time is 30min~20h.

[0022] Furthermore, the metallization rate of the metallized reduced pellets after step 2 is ≥90%.

[0023] Furthermore, in step 2, the grinding method is at least one of ball milling, rod milling, and pebble milling.

[0024] Furthermore, in step 2, the grinding fine particles account for 50-90% of the total.

[0025] Furthermore, in step 2, the magnetic separation method is wet magnetic separation, and the magnetic field strength is 200~5000GS.

[0026] Furthermore, in step 3, the acid leaching is performed using a hydrochloric acid leaching solution, in which the titanium-containing material is leached with the hydrochloric acid leaching solution.

[0027] Furthermore, in step 3, acid leaching is carried out by high-temperature and high-pressure leaching to remove impurities and obtain leaching slurry. The leaching temperature is 110~180℃, the pressure corresponds to the vapor pressure of the hydrochloric acid leaching solution at the leaching temperature, the leaching liquid-solid ratio is 1:1~10:1ml / g, and the leaching time is 1~4h.

[0028] Furthermore, in step 3, the titanium yield is ≥99%, the vanadium leaching rate is ≥90%, and the chromium leaching rate is ≥50% during the acid leaching process.

[0029] Further, in step 4, the titanium ore is placed in an alkaline leaching solution and heated at normal pressure to remove silicon. The leaching temperature is 40~100℃, the leaching solution-to-solid ratio is 0.5:1~20:1ml / g, and the leaching time is 1~4h.

[0030] Further, the alkaline leaching solution in step 4 is at least one of sodium hydroxide, potassium hydroxide, or lithium hydroxide, and the concentration of the alkaline leaching solution is 5-80 wt%.

[0031] Further, the composition of the titanium-rich material obtained in step 4, in terms of mass percentage, is as follows: TFe ≤2.0%, TiO2 ≥85.00%, SiO2 ≤1.00%, MgO ≤1.35%, CaO ≤0.15%.

[0032] Furthermore, the hydrochloric acid leaching solution comprises hydrochloric acid, chloride salt, and oxidizing agent.

[0033] Furthermore, the concentration of the hydrochloric acid solution is 15-25 wt%.

[0034] Further, the chloride salt is at least one selected from FeCl3, FeCl2, AlCl3, MgCl2, and CaCl2, and the chloride salt content is 0.1~1.5 mol / L.

[0035] Further, the oxidant is at least one selected from FeCl3, O2, NaClO, H2O2, Cl2, and KMnO4, and the content of the oxidant is 0.1~3.0 mol / L, or the partial pressure of the oxidant is 0.1~1 MPa.

[0036] Furthermore, the hydrochloric acid leaching solution described in step 5 is spray-fired at a high temperature to produce solid oxides and hydrochloric acid gas, with the spray-fire temperature being 550~750℃.

[0037] Furthermore, the oxide is vanadium-chromium-iron powder, and its composition, by mass percentage, includes Fe2O3 ≥ 60%, V2O5 ≥ 2.5%, and Cr2O3 ≥ 2.0%.

[0038] Furthermore, the hydrochloric acid gas is subjected to multi-stage countercurrent absorption by water spraying to generate a regenerated hydrochloric acid solution with a concentration of 18~22wt%. The regenerated hydrochloric acid solution is returned to the acid leaching process in step 3.

[0039] Further, in step 6, the vanadium-chromium-iron powder is washed with water to remove CaCl2. Multi-stage countercurrent water flow is used, and the amount of water added is 2 to 5 times the weight of the vanadium-chromium-iron powder. After filtration, vanadium-chromium-iron powder is obtained.

[0040] Furthermore, in step 7, the acid used for selective acid leaching of vanadium-rich ferrochrome powder is at least one of sulfuric acid, hydrochloric acid, nitric acid, and oxalic acid, with an acid concentration of 0.1~2 mol / L.

[0041] Furthermore, in step 7, the selective acid leaching of vanadium is carried out under normal pressure, with a leaching temperature of 20~100℃ and a leaching liquid-to-solid ratio of 0.5:1~20:1 ml / g. After leaching, the leaching slurry is subjected to solid-liquid separation to obtain a vanadium-containing leaching solution and chromium-rich iron powder.

[0042] Furthermore, the composition of the chromium-rich iron powder, by mass percentage, includes Fe2O3 ≥ 70% and vanadium leaching ≥ 90%.

[0043] Further, the chromium-rich iron powder described in step 8 is mixed with the second reducing agent and smelted into ferrochrome in an electric furnace at a smelting temperature of 1550~1700℃.

[0044] Furthermore, in step 8, the ratio of chromium-rich iron powder to the second reducing agent is 1:0.15 to 1:0.30, and ferrochrome and smelting slag are obtained after smelting.

[0045] Furthermore, in step 8, the second reducing agent is at least one of coking coal, hydrogen, and CO.

[0046] Compared with the prior art, the beneficial effects of the present invention are as follows: 1. The present invention provides a method for utilizing valuable elements in vanadium-titanium magnetite concentrate. This method replaces high-temperature multiple roasting by preparing high-performance alkaline green pellets of vanadium-titanium magnetite. This increases the particle size of the raw material, improves its adaptability, allows for the addition of calcining agents to change the mineral phase, reduces energy consumption, generates less pollution from waste during roasting, and significantly increases the recovery rate of titanium.

[0047] 2. This invention provides the relationship between the amount of calcifying agent added and SiO2 and V2O5 in vanadium-titanium magnetite concentrate, as well as the optimal process conditions for the formation of CaSiO3 and CaV2O6, which is beneficial for the subsequent acid and alkaline leaching of elements such as aluminum, calcium, magnesium, vanadium, and silicon; and constructs a chloride-oxidant hydrochloric acid leaching system. This increases the solubility of vanadium and chromium, while simultaneously solving the problem of high titanium dissolution rate during the acid leaching process.

[0048] 3. This invention enables acid recycling, which greatly reduces production energy consumption and emissions of waste gas, wastewater, and solid waste, making it environmentally friendly.

[0049] 4. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate of the present invention is applicable not only to magnetite concentrates with low vanadium-titanium and high iron content, but also to vanadium-titanium magnetite concentrates with high vanadium-titanium and low iron content. Attached Figure Description

[0050] To more clearly illustrate the specific embodiments of the present invention or the technical solutions in the prior art, the drawings used in the description of the specific 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 from these drawings without creative effort.

[0051] Figure 1 Scanning electron microscope (SEM) images of vanadium-titanium magnetite metallized pellets of the present invention; Figure 2 XRD patterns of non-magnetic titanium-containing materials in this invention; Figure 3 Scanning electron microscope (SEM) morphology of the non-magnetic titanium-containing material of this invention; Figure 4 Scanning electron microscope (SEM) morphology of titanium minerals in this invention; Figure 5 Photograph of titanium-rich material after impregnation using Ca(OH)2 granulated pellets in this invention; Figure 6 XRD pattern of vanadium-chromium-iron powder obtained after incineration according to the present invention; Figure 7 Photograph of titanium-rich material after impregnation of bentonite granulated pellets in this invention; Figure 8 Scanning electron microscope image of the titanium-rich material after reaction without the addition of oxidant in this invention; Figure 9 The present invention provides a process flow diagram for the utilization of valuable elements in vanadium-titanium magnetite concentrate. Detailed Implementation

[0052] The technical solution of the present invention will be clearly and completely described below with reference to specific embodiments. However, those skilled in the art will understand that the embodiments described below are some embodiments of the present invention, but not all embodiments, and are only used to illustrate the present invention, and should not be regarded as limiting the scope of the present invention. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention. Where specific conditions are not specified in the embodiments, conventional conditions or conditions recommended by the manufacturer shall be followed. Where the manufacturers of reagents or instruments are not specified, they are all conventional products that can be purchased commercially.

[0053] A method for utilizing valuable elements in vanadium-titanium magnetite concentrate includes the following steps: Step 1: Using vanadium-titanium magnetite concentrate and calcifying agent as raw materials, the vanadium-titanium magnetite concentrate and calcifying agent are mixed to prepare alkaline green pellets of vanadium-titanium magnetite concentrate. The alkaline green pellets are then subjected to calcification, oxidation, and roasting to change the phase composition and improve the pellet strength, thereby obtaining calcified and oxidized pellets. Step 2: Mix the calcified oxide pellets with the first reducing agent and reduce them at high temperature to reduce Fe2O3 in the calcified oxide pellets to metallic iron and obtain metallized reduced pellets. Crush and grind the metallized reduced pellets to separate metallic iron particles and titanium-containing materials. Then, perform magnetic separation to obtain magnetic reduced iron powder and non-magnetic titanium-containing materials. Step 3: Add chloride salt and oxidant to hydrochloric acid solution to prepare hydrochloric acid leaching solution. Leach the titanium-containing material with the prepared hydrochloric acid leaching solution under high temperature and pressure to remove impurities and obtain leaching slurry. Perform solid-liquid separation and washing on the leaching slurry to obtain hydrochloric acid leaching solution and filter cake. Use the filter cake as titanium middlings or dry the filter cake to obtain titanium middlings. Step 4: Place the titanium ore in an alkaline leaching solution at normal pressure and temperature to remove silicon. After alkaline leaching, perform solid-liquid separation on the leaching slurry to obtain an alkaline leaching solution and a filter cake. Dry the filter cake to obtain titanium-rich material. Step 5: The hydrochloric acid leaching solution is sprayed and burned at high temperature to generate vanadium-chromium-iron powder and hydrochloric acid gas. The hydrochloric acid gas is absorbed by water spray to generate hydrochloric acid solution and returned to the acid leaching process in Step 3. Step 6: Wash the vanadium-chromium-iron powder with water to remove CaCl2, perform solid-liquid separation on the washed slurry to obtain CaCl2 solution and filter cake, and dry the filter cake to obtain vanadium-chromium-iron powder. Step 7: Selectively acid-leach vanadium with the vanadium-rich ferrochrome powder, and then perform solid-liquid separation on the leaching slurry to obtain a vanadium-containing leaching solution and a ferrochrome-rich powder filter cake. The filter cake is then dried to obtain ferrochrome-rich powder. Step 8: Mix the chromium-rich iron powder with the second reducing agent and smelt it into ferrochrome.

[0054] Preferably, the diameter of the alkaline green pellets in step 1 is 9~16mm, including but not limited to 9mm, 10mm, 12mm, 15mm, and 16mm.

[0055] Preferably, in step 1, the vanadium-titanium magnetite concentrate is dried and the moisture content is adjusted by a dryer, and then ground and mixed with a calcifying agent in a roller mill or a grinding mill. After adjusting the moisture content, alkaline green pellets of vanadium-titanium magnetite concentrate are prepared in a disc pelletizer. The alkaline green pellets are screened by 16mm and 9mm roller screens to obtain qualified alkaline green pellets of 9-16mm.

[0056] Preferably, the vanadium-titanium magnetite concentrate described in step 1 comprises, by mass percentage, 50-60% TFe, 8-15% TiO2, 0.5-2.0% V2O5, 0.5-5.0% SiO2, 1.0-3.0% Cr2O3, 0.1-5% CaO, and other impurities.

[0057] Preferably, the calcifying agent in step 1 is at least one of CaO, CaSO4, Ca(OH)2 calcium-containing compounds or calcium-containing minerals; preferably Ca(OH)2, which is both a calcifying agent and a ball-forming binder. Ca(OH)2 is dry slaked lime, which is slaked lime produced from CaO using a dry slaking process, wherein CaO ≥ 70% and the particle size is between 1 and 100 μm.

[0058] Preferably, in step 1, during the alkaline pelletizing of vanadium-titanium magnetite concentrate, the amount of calcifying agent such as Ca(OH)2 added relative to 100 parts of vanadium-titanium magnetite concentrate is calculated using the following formula (Wt). Ca(OH)2 =K(123Wt SiO2 +41Wt V2O5 -132Wt CaO ), where Wt Ca(OH)2 The amount of calcifying agent such as Ca(OH)2 added is specified, with an excess coefficient K of 1.2~1.5, Wt SiO2 Wt V2O5 and Wt CaO The percentage contents of SiO2, V2O5, and CaO in vanadium-titanium magnetite concentrate are respectively; and Wt Ca(OH)2 The amount of calcifying agent such as Ca(OH)2 added shall not be less than 1.2 parts.

[0059] One of the main functions of adding calcifying agents such as Ca(OH)2 is that during the calcification, oxidation, and roasting process, calcifying agents such as Ca(OH)2 react with SiO2 and V2O5 in vanadium-titanium magnetite concentrate to generate acid-soluble CaSiO3 and calcium vanadate (such as CaV2O6), which is beneficial for removing impurities such as calcium and vanadium from titanium-containing materials. At the same time, the alkali leaching of titanium ore removes silicon to produce high-grade titanium-rich materials.

[0060] The second main function of adding calcifying agents such as Ca(OH)2 is as a binder in pellet preparation. The amount of calcifying agent such as Ca(OH)2 added is related to the content of SiO2, V2O5 and CaO in vanadium-titanium magnetite concentrate. When the content of SiO2 and V2O5 is low or the content of CaO is high, the amount of calcifying agent such as Ca(OH)2 added according to the above calculation formula may be low and may not meet the requirements of the binder for pellet preparation. In order to ensure the preparation of alkaline green pellets, the minimum amount of calcifying agent such as Ca(OH)2 added is 1.2 parts.

[0061] In the preparation of vanadium-titanium magnetite pellets, a calcifying agent such as Ca(OH)2 is added as a binder to prepare alkaline green pellets. Ca(OH)2 replaces the bentonite binder used in conventional pelletizing, avoiding the introduction of impurities SiO2 and Al2O3 into the pellets and reducing the pressure of acid leaching for aluminum removal and alkaline leaching for silicon removal. By adding calcifying agent such as Ca(OH)2 to change the basicity of the pellets, during the calcification roasting process, elements Si and V combine with CaO, altering the mineral phase composition of Si and V in the pellets, presenting them as Ca-Si silicate phases and CaV2O6, which is beneficial for the subsequent acid and alkaline leaching of elements such as aluminum, calcium, magnesium, vanadium, and silicon.

[0062] Preferably, the calcination temperature in step 1 is 1100~1250℃ (including but not limited to 1100℃, 1150℃, 1200℃, 1250℃), more preferably 1180~1230℃, and the calcination time is 20min~10h (including but not limited to 20min, 30min, 1h, 2h, 3h, 5h, 6h, 8h, 10h), more preferably 2~6h, and the compressive strength of the calcined oxidized pellets after calcination is ≥1500N / piece.

[0063] Preferably, the mass ratio of the calcified oxide pellets to the first reducing agent in step 2 is 1:0.1 to 1:2 (including but not limited to 1:0.1, 1:0.5, 1:1, 1:1.5, and 1:2).

[0064] Preferably, in step 2, the first reducing agent is at least one of carbon-based substances (such as coal, coke, etc.), hydrogen, and CO, and the reduction method includes, but is not limited to, CO reduction, H2 reduction, and carbon-based (coal, coke) reduction.

[0065] Preferably, the reduction temperature in step 2 is 900~1300℃ (including but not limited to 900℃, 1000℃, 1100℃, 1200℃, 1300℃), more preferably 1200~1270℃, and the reduction time is 30min~20h (including but not limited to 30min, 1h, 2h, 3h, 5h, 6h, 8h, 10h, 15h, 20h), and the metallization rate of the metallized reduced pellets after reduction is ≥90%.

[0066] After reduction, the main phases of the pellets are metallic iron and titanium minerals and other silicate gangue phases. The evolution of iron grain diameter and embedding behavior is closely related to the changes in the enrichment and separation effects of iron, vanadium, and titanium during grinding and beneficiation. As the reduction temperature increases, the iron grains gradually grow, and the mineral phase interfaces become more defined, providing a structural basis for the effective liberation of titanium-containing materials with high vanadium and titanium content. This facilitates the grinding liberation and magnetic separation of iron with titanium and vanadium minerals during subsequent grinding and beneficiation processes.

[0067] Preferably, the grinding method in step 2 is at least one of ball milling, rod milling, and pebble milling.

[0068] Preferably, in step 2, the grinding fine particles have a proportion of -200 mesh (74μm) of 50-90% (including but not limited to 50%, 60%, 70%, 80%, 90%), with 70-90% being more preferred.

[0069] Reduction-grinding separation is a process that selectively reduces iron in vanadium-titanium magnetite concentrate under solid-state conditions, while vanadium and titanium remain in oxide form. Further fine grinding and magnetic separation then effectively separate iron from other elements. Compared to traditional processes such as the blast furnace method, the reduction-grinding separation process has a shorter flow rate, lower cost, and higher ferrotitanium recovery rate and grade, making it significantly advantageous for separating ferrotitanium from vanadium-titanium magnetite concentrate. After reduction-grinding separation, non-ferrous elements such as V, Ti, Si, Cr, and Al are enriched in the titanium-containing material, while the metallic iron particles do not contain V or other elements. The impurities such as V, Ti, Si, Al, and Cr in the reduced iron powder, the magnetic product, mainly originate from entrainment in the intergrowth of metallic iron particles.

[0070] Preferably, the magnetic separation method in step 2 is wet magnetic separation, and the magnetic field strength is 200~5000GS (including but not limited to 200GS, 500GS, 700GS, 1000GS, 2000GS, 3000GS, 4000GS, 5000GS), preferably 500~700GS.

[0071] Preferably, the hydrochloric acid leaching solution in step 3 comprises hydrochloric acid, a chloride salt, and an oxidizing agent, prepared by adding a chloride salt and an oxidizing agent to a hydrochloric acid solution. The concentration of the hydrochloric acid solution is 15-25 wt%; the chloride salt is at least one selected from FeCl3, FeCl2, AlCl3, MgCl2, and CaCl2, and the chloride salt content is 0.1-1.5 mol / L (including but not limited to 0.1 mol / L, 0.3 mol / L, 0.5 mol / L, 1 mol / L, 1.2 mol / L, and 1.5 mol / L); the oxidizing agent is at least one selected from FeCl3, O2, NaClO, H2O2, Cl2, and KMnO4, and the oxidizing agent (selected from FeCl3, NaClO, H2O2, and KMnO4) is... The content of the oxidant (when at least one of O4 is selected) is 0.1~3.0 mol / L (including but not limited to 0.1 mol / L, 0.3 mol / L, 0.5 mol / L, 1 mol / L, 1.2 mol / L, 1.5 mol / L, 2 mol / L, 2.5 mol / L, 3 mol / L), or the partial pressure of the oxidant (when at least one of O2 and Cl2 is selected) is 0.1~1 MPa (0.1 MPa, 0.2 MPa, 0.3 MPa, 0.4 MPa, 0.5 MPa, 0.6 MPa, 0.7 MPa, 0.8 MPa, 0.9 MPa, 1 MPa).

[0072] The addition of chloride salts enhances the activity of hydrogen ions in hydrochloric acid, which is beneficial for the leaching of impurities such as Fe, Al, Ca, Mg, Mn, V, and Cr from titanium-containing materials. Simultaneously, the addition of chloride salts increases the ionic strength of the leaching solution, which is conducive to the hydrolysis of metatitanic acid, thus increasing titanium recovery and improving the filtration performance of the leaching slurry. Oxidants are used to oxidize trivalent titanium to tetravalent titanium and trivalent vanadium to tetravalent or pentavalent vanadium during the leaching process, which is beneficial for increasing titanium recovery and improving vanadium leaching rate.

[0073] A chloride-oxidant hydrochloric acid leaching system is used. Adding chloride increases the chloride ion activity in the hydrochloric acid leaching system, which is beneficial for impurity removal. Since titanium-containing materials contain trivalent titanium, trivalent vanadium, and metallic iron, adding an oxidant to the chloride-based system further enhances the removal of dissolved Ti. 3+ In-situ oxidation to Ti 4+ The process involves hydrolysis to form metatitanic acid, which is adsorbed within the particles, thus resolving the issues of high titanium dissolution and low recovery rates during acid leaching. Simultaneously, the addition of an oxidant effectively improves the leaching rate of titanium (V). Tetravalent vanadium exhibits high solubility in hydrochloric acid solution, thereby enhancing the leaching rate of V from titanium-containing minerals. Furthermore, under high acidity conditions, as leaching progresses, the oxidant concentration decreases while the tetravalent vanadium concentration increases. Since tetravalent vanadium also acts as an oxidant, it further oxidizes trivalent titanium, thereby inhibiting titanium leaching.

[0074] Preferably, in step 3, acid leaching is carried out by high-temperature and high-pressure leaching to remove impurities and obtain leaching slurry. The leaching temperature is 110~180℃ (including but not limited to 110℃, 120℃, 130℃, 140℃, 150℃, 160℃, 170℃, 180℃), the pressure corresponds to the vapor pressure of the hydrochloric acid leaching solution at the leaching temperature, the leaching liquid-solid ratio is 1:1~10:1ml / g (including but not limited to 1:1ml / g, 3:1ml / g, 5:1ml / g, 6:1ml / g, 8:1ml / g, 10:1ml / g), and the leaching time is 1~4h (including but not limited to 1h, 2h, 3h, 4h).

[0075] Preferably, the leachate slurry in step 3 is subjected to solid-liquid separation using conventional solid-liquid separation methods such as pressure filtration and centrifugation. Diaphragm pressure filtration is preferred. After pressure filtration, the filter cake is washed with water and used as titanium middlings, or the filter cake is dried to obtain titanium middlings at a drying temperature ≤350℃. Preferably, the filter cake is used as titanium middlings.

[0076] In the calcification and oxidation roasting process, calcifying agents and binders such as Ca(OH)2 are added. The calcifying agent, such as Ca(OH)2, reacts with SiO2 and V2O5 in the vanadium-titanium magnetite concentrate at high temperature to generate CaSiO3 and calcium vanadate (such as CaV2O6). The CaV3O7 and CaSiO3 generated after reduction react with HCl according to the following reaction formula: CaSiO3 + 2HCl → Ca 2+ + 2Cl - + H2SiO3 CaV3O7 + 2HCl → Ca 2+ + 2Cl - + 2H + + V3O7 2- After the reaction, Ca 2+ V3O7 2- It dissolves in the leachate in ionic form, while H2SiO3 is insoluble in acidic solutions and exists in the filter cake of titanium ore in the form of silica gel.

[0077] Preferably, in step 3, the acid leaching process yields titanium ≥99%, vanadium leaching ≥90%, and chromium leaching ≥50%.

[0078] Preferably, in step 4, the titanium ore is placed in an alkaline leaching solution and subjected to alkaline leaching at normal pressure and temperature to remove silicon. The leaching temperature is 40~100℃ (including but not limited to 40℃, 50℃, 60℃, 70℃, 80℃, 90℃, 100℃), the leaching liquid-solid ratio is 0.5:1~20:1ml / g (including but not limited to 0.5:1ml / g, 1:1ml / g, 2:1ml / g, 5:1ml / g, 6:1ml / g, 8:1ml / g, 10:1ml / g, 15:1ml / g, 20:1ml / g), and the leaching time is 1~4h (including but not limited to 1h, 2h, 3h, 4h).

[0079] Preferably, the alkaline leaching solution in step 4 is at least one of sodium hydroxide, potassium hydroxide, or lithium hydroxide, and the concentration of the alkaline leaching solution is 5-80 wt%.

[0080] Preferably, after alkali leaching in step 4, the leachate is subjected to solid-liquid separation using conventional solid-liquid separation methods such as pressure filtration and centrifugation, with diaphragm pressure filtration being preferred. After pressure filtration, the filter cake is washed with water and dried to obtain titanium-rich material.

[0081] Preferably, the composition of the titanium-rich material obtained in step 4, by mass percentage, is TFe≤2.0%, TiO2≥85.00%, SiO2≤1.00%, MgO≤1.35%, and CaO≤0.15%.

[0082] Preferably, the hydrochloric acid leaching solution in step 5 is spray-fired at a high temperature to generate solid oxides and hydrochloric acid gas, with the spray-fire temperature being 550~750℃, thereby regenerating hydrochloric acid through spray-fire. The hydrochloric acid leaching solution is atomized in an incinerator at 550~750℃ (including but not limited to 550℃, 600℃, 650℃, 700℃, and 750℃). Other chloride salts besides CaCl2 (such as FeCl2, FeCl3, MgCl2, VOCl3, CrCl3, AlCl3, etc.) undergo hydrolysis and oxidation at high temperature to generate their respective high-valence oxides and HCl gas, while CaCl2 remains unhydrolyzed and exists as CaCl2 in the oxides. The oxides are vanadium-chromium-iron powder, and their composition, by mass percentage, includes Fe2O3 ≥ 60%, V2O5 ≥ 2.5%, and Cr2O3 ≥ 2.0%.

[0083] During incineration, metal chlorides such as FeCl2, MgCl2, Fe2Cl3, VOCl2, VOCl, CrCl3, and AlCl3 in the hydrochloric acid leaching solution are converted into metal oxides MgO, Fe2O3, V2O5, Cr2O3, and Al2O3. However, CaCl2 remains unchanged during incineration due to its high hydrolysis temperature and continues to exist as CaCl2. After high-temperature incineration, the hydrochloric acid leaching solution yields vanadium-chromium-iron powder and hydrochloric acid gas. The composition of the vanadium-chromium-iron powder, by mass percentage, is Fe2O3 ≥ 60%, V2O5 ≥ 2.5%, and Cr2O3 ≥ 2.0%. The hydrochloric acid gas is subjected to multi-stage countercurrent absorption via water spraying to generate a regenerated hydrochloric acid solution with a concentration of 18-22 wt%. This regenerated hydrochloric acid solution is returned to the acid leaching process in step 3.

[0084] Preferably, in step 6, the vanadium-chromium-iron powder is washed with water to remove CaCl2. The vanadium-chromium-iron powder discharged from the incinerator at 350~550℃ (including but not limited to 350℃, 400℃, 450℃, 500℃, and 550℃) is directly introduced into the water to raise the water temperature, which is conducive to the removal of CaCl2.

[0085] Preferably, in step 6, the vanadium-chromium-iron powder is washed with water using a multi-stage countercurrent water flow, and the amount of water added is 2 to 5 times (including but not limited to 2 times, 3 times, 4 times, and 5 times) the weight of the vanadium-chromium-iron powder. The filtration is carried out using conventional powder filtration methods such as belt filters and diaphragm filter presses. The filtrate is a CaCl2 solution, and the filter cake is vanadium-chromium-iron powder.

[0086] Preferably, in step 7, the acid used for selective acid leaching of vanadium-rich ferrochrome powder is at least one of sulfuric acid, hydrochloric acid, nitric acid, and oxalic acid, and the acid concentration is 0.1~2 mol / L (including but not limited to 0.1 mol / L, 0.5 mol / L, 1 mol / L, 1.5 mol / L, and 2 mol / L).

[0087] After hydrochloric acid regeneration, vanadium-chromium-iron powder is produced, in which vanadium exists in the form of vanadate, which can be dissolved into the solution through acid leaching. Other elements remain in the iron powder, and the vanadium element in the solution can be recovered.

[0088] Preferably, in step 7, the selective acid leaching of vanadium is carried out under normal pressure, with a leaching temperature of 20~100℃ (including but not limited to 20℃, 40℃, 50℃, 60℃, 80℃, 100℃), and a leaching liquid-to-solid ratio of 0.5:1~20:1ml / g (including but not limited to 0.5:1ml / g, 1:1ml / g, 2:1ml / g, 5:1ml / g, 6:1ml / g, 8:1ml / g, 10:1ml / g, 15:1ml / g, 20:1ml / g). After leaching, the leaching slurry is subjected to solid-liquid separation to obtain a vanadium-containing leaching solution and a chromium-rich iron powder filter cake, and the filter cake is dried to obtain chromium-rich iron powder.

[0089] Preferably, the chromium-rich iron powder comprises, by mass percentage, ≥70% Fe2O3 and ≥90% vanadium leaching.

[0090] Preferably, the chromium-rich iron powder in step 8 is mixed with the second reducing agent and smelted into ferrochrome in an electric furnace at a smelting temperature of 1550~1700℃ (including but not limited to 1550℃, 1600℃, 1650℃, and 1700℃).

[0091] Preferably, in step 8, the ratio of chromium-rich iron powder to the second reducing agent is 1:0.15 to 1:0.30 (including but not limited to 1:0.15, 1:0.20, 1:0.25, and 1:0.30), and ferrochrome and smelting slag are obtained after smelting.

[0092] Preferably, in step 8, the second reducing agent is at least one of coking coal, hydrogen, and CO.

[0093] Example 1 In this embodiment, calcified oxide pellets were prepared. The composition of the vanadium-titanium magnetite concentrate used is shown in Table 1-1, and the composition of the Ca(OH)2 used is shown in Table 1-2. The Ca(OH)2 particles are relatively fine, with most particles having a diameter of about 10 μm and a low proportion of large particles, with -200 mesh accounting for 95%. Alkaline green pellets were prepared by adding 1.9 parts of Ca(OH)2 to 100 parts of vanadium-titanium magnetite concentrate and then grinding them. The alkaline green pellets were then calcified and oxidized at 1200℃ for 2 hours to prepare calcified oxide pellets. The physical properties of the calcified oxide pellets are shown in Table 1-3, and the chemical composition of the vanadium-titanium magnetite concentrate calcified oxide pellets is shown in Table 1-4.

[0094] Table 1-1 Chemical composition of vanadium-titanium magnetite concentrate

[0095] Table 1-2 Chemical composition of Ca(OH)2

[0096] Table 1-3 Physical properties of vanadium-titanium magnetite concentrate calcified oxide pellets

[0097] Table 1-4 Chemical composition of vanadium-titanium magnetite concentrate calcified oxide pellets

[0098] Example 2 In Example 1, calcified oxide pellets were mixed with anthracite at a ratio of 1:0.5. Metallized pellets were prepared by reducing the mixture at 1250℃ for 2 hours. The chemical composition of the metallized pellets is shown in Table 2-1, and the metallization η of the pellets was 91.59%. Scanning electron microscopy images of the reduced pellets (reduction temperature 1250℃, 2 hours) are shown below. Figure 1 As shown. By Figure 1 It can be seen that the metallic iron phase inside the pellet is connected in sheets, resulting in a good reduction effect.

[0099] Table 2-1 Chemical composition of vanadium-titanium magnetite concentrate metallized pellets

[0100] Example 3 In Example 1, calcified oxide pellets were mixed with anthracite, and the surface of the roasted pellets was covered with coal for reduction in a muffle furnace. After reduction, the pellets were cooled in water to prevent iron oxidation. The reduction temperatures were 1200℃, 1250℃, and 1270℃, and the reduction time was 2 hours. After reduction, the pellets were ball-milled to particle sizes of -200 mesh (80%) and 90%, respectively. Wet magnetic separation was then performed after ball milling at a magnetic field strength of 700 GS. The magnetic separation results are shown in Tables 3-1 to 3-6. The XRD patterns of the non-magnetic titanium-containing materials separated by magnetic separation are shown in Tables 3-1 to 3-6. Figure 2 (Reduction temperature 1250℃, 2h; ball milling particle size -200 mesh 80%), as shown in the scanning electron microscope image. Figure 3 (Reduction temperature 1250℃, 2h; ball milling particle size -200 mesh 80%) As shown, from Figure 2 It can be seen that the main phase of the titanium-containing material after reduction and grinding at 1250℃ is titanomagnetic (Fe). 2.5 Ti 0.5 O4), magnesium-containing titanium dioxide (Mg 1.05 Ti 1.95 O5), complex calcium aluminate crystals (Na6O) 72 Al 24 Ca 33 ).

[0101] After magnetic separation, the recovery rates of V and Cr in titanium-containing minerals were correlated with those of TiO2, indicating that V, Cr, and Ti minerals were enriched in non-magnetic minerals, while metallic iron did not contain other elements such as V, Cr, and Ti. The presence of residual vanadium and chromium in the reduced iron powder was due to incomplete dissociation and separation of the titanium-containing material from the metallic iron during the grinding and separation process. This was caused by the intergrowth of some metallic iron and titanium-containing material within the pellets.

[0102] With increasing ball milling time, the Fe, Ti, V, and Cr content and yield of reduced iron powder first decreased and then increased, while the Fe, Ti, V, and Cr content and yield of titanium-containing minerals first increased and then decreased. Figure 3It is known that this is because as the grinding fineness increases, iron grains and non-ferrous phases gradually dissociate, reaching the optimal magnetic separation effect at a grinding fineness of -200 mesh (80%). Further increasing the grinding fineness leads to excessively fine non-ferrous mineral particles, resulting in magnetic agglomeration during magnetic separation. Over-ground non-ferrous minerals are trapped within metallic iron, causing an increase in the recovery rate of Fe, Ti, V, and Cr in reduced iron powder, and a decrease in the recovery rate of Fe, Ti, V, and Cr in titanium-containing minerals.

[0103] Meanwhile, the grinding and separation effect increased significantly with increasing reduction temperature. There was no significant difference in grinding and separation between reduction at 1250℃ and 1270℃. After reduction and grinding at 1200℃, the recovery rates of iron, titanium, and vanadium in the titanium material were the lowest, indicating the worst enrichment effect for vanadium and titanium. The experimental groups with the best vanadium and titanium enrichment effects were those with a grinding particle size of -200 mesh (90%) after reduction at 1250℃ and -200 mesh (80%) after reduction at 1270℃.

[0104] Table 3-1 Magnetic separation results (wt%) (reduction temperature 1200℃, 2h; ball milling particle size -200 mesh 80%)

[0105] Table 3-2 Magnetic separation results (wt%) (reduction temperature 1200℃, 2h; ball milling particle size -200 mesh 90%)

[0106] Table 3-3 Magnetic separation results (wt%) (reduction temperature 1250℃, 2h; ball milling particle size -200 mesh 80%)

[0107] Table 3-4 Magnetic separation results (wt%) (reduction temperature 1250℃, 2h; ball milling particle size -200 mesh 90%)

[0108] Table 3-5 Magnetic separation results (wt%) (reduction temperature 1270℃, 2h; ball milling particle size -200 mesh 80%)

[0109] Table 3-6 Magnetic separation results (wt%) (reduction temperature 1270℃, 2h; ball milling particle size -200 mesh 90%)

[0110] Example 4 The reduction conditions used in Example 3 were: reduction temperature 1250℃, 2h, ball milling particle size -200 mesh 90%, followed by wet magnetic separation with a magnetic field strength of 700GS to separate metallic iron from titanium-containing materials in the pellets. Non-magnetic titanium-containing material with a TiO2 content of 37.23% was obtained through magnetic separation. The titanium-containing material was then further purified using a chloride-oxidant hydrochloric acid leaching system. The leaching conditions were: hydrochloric acid concentration 20%wt, FeCl3 as both oxidant and chloride, with a total FeCl3 content of 0.3mol / L, a leaching liquid-to-solid ratio of 3:1ml / g, a leaching temperature of 145℃, and a leaching time of 2h. After the reaction, the slurry was filtered, washed, and dried to obtain titanium middlings. The filtrate was a hydrochloric acid leaching solution. The composition and leaching rate of the titanium middlings are shown in Table 4-1, and the scanning electron microscope images are shown below. Figure 4 As shown.

[0111] Table 4-1 Composition of titanium minerals and their leaching rates (wt%)

[0112] After acid leaching, the ωΣ(CaO+MgO) content in the titanium middlings was less than 1.5%, the TiO2 recovery rate was 99.64%, the V2O5 leaching rate was 94.04%, and the Cr2O3 leaching rate was 62.84%. The morphology of the leached titanium middlings under scanning electron microscopy showed that it was needle-shaped with crystals embedded in the particles.

[0113] Example 5 Because the SiO2 content in the titanium slag (titanium ore) of Example 4 is as high as 11.88%, the SiO2 content is too high when used as a raw material for titanium dioxide production, and further removal of SiO2 from the titanium slag is necessary. The titanium slag from Example 4 was selected as the raw material for alkaline leaching. The alkaline leaching conditions were: sodium hydroxide concentration 10wt%, leaching liquid-to-solid ratio 5:1ml / g, leaching temperature 95℃, and leaching time 3h. After alkaline leaching, the slurry was filtered, washed, and dried to obtain titanium-rich material, the composition of which is shown in Table 5-1. A photograph of the appearance of the titanium-rich material is shown below. Figure 5 As shown.

[0114] Table 5-1 Composition of titanium-rich material after alkali leaching (wt%)

[0115] After alkali leaching, the titanium-rich material has a MgO content of 1.19%, a CaO content of 0.10%, and a SiO2 content of 0.98%. The ωΣ(CaO+MgO) in the titanium-rich material is 1.29%, which meets the requirements of the fluidized bed chlorination process for raw material composition.

[0116] Example 6 The non-magnetic titanium-containing material with a TiO2 content of 37.23% from Table 3-4 of Example 3 was selected as the raw material for Example 6. The titanium-containing material was upgraded and impurities removed using a chloride-oxidant hydrochloric acid leaching system. The reaction conditions were: hydrochloric acid concentration 20 wt%, MgCl2 content 0.3 mol / L, O2 partial pressure 0.1 MPa, leaching liquid-to-solid ratio 3:1 ml / g, leaching temperature 145℃, and leaching time 2 h. After the reaction, the slurry was filtered, washed, and dried to obtain titanium middlings. The filtrate was a hydrochloric acid leaching solution. The composition of the titanium middlings is shown in Table 6-1.

[0117] Table 6-1 Mineral composition and leaching rate (wt%) in titanium

[0118] After acid leaching, the ωΣ(CaO+MgO) content in the titanium middlings ore was less than 1.5%, the TiO2 recovery rate was 99.42%, the V2O5 leaching rate was 94.73%, and the Cr2O3 leaching rate was 60.55%. With the same SiO2 content of 11.56%, the titanium middlings ore was further subjected to alkaline leaching under the following conditions: sodium hydroxide concentration 10 wt%, liquid-to-solid ratio 5:1 ml / g, leaching temperature 95℃, and leaching time 3 h. The slurry after alkaline leaching was filtered, washed, and dried to obtain titanium-rich material, the composition of which is shown in Table 6-2.

[0119] Table 6-2 Titanium-rich material composition (wt%)

[0120] After alkali leaching, the titanium-rich material has a MgO content of 1.31%, a CaO content of 0.12%, and a SiO2 content of 0.95%. The ωΣ(CaO+MgO) in the titanium-rich material is 1.43%, which meets the requirements of the fluidized bed chlorination process for raw material composition.

[0121] Example 7 The composition of the hydrochloric acid leaching solution for the non-magnetic titanium-containing material in Example 4 is shown in Table 7-1. Table 7-1 Composition of hydrochloric acid leaching solution for non-magnetic titanium-containing materials in Example 4

[0122] The treatment method for hydrochloric acid leaching solutions is widely used in hydrochloric acid regeneration processes in the iron and steel smelting industry. During the hydrochloric acid regeneration process, the Fe in the leaching solution... 2+ / Fe 3+ V2O4 2- Mg 2+ Al 3+ Mn 2+ Cr 3+At high temperatures, it hydrolyzes and oxidizes to Fe2O3, V2O5, MgO, Al2O3, MnO, and Cr2O3, respectively. CaCl2, due to its higher hydrolysis temperature, remains unhydrolyzed and exists as CaCl2 in the oxide solid phase. - The HCl gas is converted into HCl gas, which is then washed and adsorbed with water to form a 18-20 wt% regenerated hydrochloric acid solution. The hydrochloric acid regenerated oxide solid is vanadium-chromium-iron powder, the composition of which is shown in Table 7-2, and its XRD pattern is shown in [Table 7-2]. Figure 6 As shown, by Figure 6 It is known that the solid phase mainly consists of magnesium-iron spinel (MgFe2O4), iron oxide powder, and vanadates. Its 5% vanadium content is considered highly enriched, and iron can also be recovered and utilized. Furthermore, chromium, a valuable alloying element, can be recycled into the molten iron to enhance the value of the steel, resulting in significant overall economic value.

[0123] Table 7-2 Composition (wt%) of hydrochloric acid regenerated oxide solids (vanadium chromium iron powder)

[0124] The CaCl2 in vanadium-chromium-iron powder can be washed away with water. Using twice the weight of water in a three-stage countercurrent washing process, the CaCl2 removal rate can reach over 98%. The solid remaining after water washing is vanadium-chromium-iron powder, the composition of which is shown in Table 7-3. Vanadium-chromium-iron powder can be used as a raw material for extracting vanadium and chromium resources.

[0125] Table 7-3 Composition of Vanadium-Rich Chromium Iron Powder (wt%)

[0126] Example 8 The vanadium-chromium iron powder from Table 7-3 of Example 7 was selected as the raw material, and vanadium-chromium iron powder was leached with hydrochloric acid solution. The hydrochloric acid concentration was 0.7 mol / L, the leaching temperature was 25℃, the leaching solution-to-solid ratio was 2:1 ml / g, and the leaching time was 2 hours. After leaching, the slurry was filtered, washed, and dried. The filtrate was a vanadium-containing leaching solution, and the solid was vanadium-chromium iron powder. The components and leaching rates are shown in Table 8-1.

[0127] Table 8-1 Composition of vanadium-containing leaching solution and chromium-rich iron powder and their leaching rates (%)

[0128] The vanadium-containing leaching solution can be used to produce vanadium pentoxide using conventional ammonium precipitation technology, while chromium-rich iron powder can be used as a raw material for electric furnace reduction smelting to produce ferrochrome, thus realizing the stepwise comprehensive recovery of iron, titanium, vanadium and chromium in vanadium-titanium magnetite concentrate.

[0129] Comparative Example 1 Bentonite was used as a substitute for Ca(OH)2 as a calcifying agent and binder in the preparation of vanadium-titanium magnetite concentrate pellets. The chemical elemental composition and basic physical properties of the bentonite used are shown in Tables 9-1 and 9-2.

[0130] Table 9-1 Chemical elemental composition of bentonite

[0131] Table 9-2 Basic Physical Properties of Bentonite

[0132] Vanadium-titanium magnetite concentrate and bentonite (Tables 9-1 to 9-2) were added to 100 parts of vanadium-titanium magnetite concentrate and then milled to prepare green pellets. The green pellets were then oxidized and roasted at 1200℃ for 1 hour to prepare roasted pellets, the composition of which is shown in Table 9-3.

[0133] Table 9-3 Typical Chemical Compositions of Calcined Bentonite Pelletizes

[0134] Calcined bentonite pellets were mixed with anthracite coal, and the pellets were covered with coal before reduction in a muffle furnace. After reduction, the mixture was cooled in water to prevent iron oxidation. The reduction temperature was 1250℃, and the reduction time was 2 hours. The pellets were then ball-milled to a particle size of -200 mesh (90%). After ball milling, wet magnetic separation was performed at a magnetic field strength of 700 GS. The non-magnetic titanium-containing material was separated by magnetic separation, and its composition is shown in Table 9-4.

[0135] Table 9-4 Magnetic separation results (wt%) (reduction temperature 1250℃, 2h; ball milling particle size -200 mesh 90%)

[0136] Titanium-containing materials were upgraded and purified using a chloride-oxidant hydrochloric acid leaching system. The leaching conditions were as follows: hydrochloric acid concentration 20wt%, FeCl3 was selected as both oxidant and chloride, with a total FeCl3 content of 0.3mol / L, a leaching liquid-to-solid ratio of 3:1ml / g, a leaching temperature of 145℃, and a leaching time of 2h. After the reaction, the slurry was filtered, washed, and dried to obtain titanium middlings. The filtrate was a hydrochloric acid leaching solution. The composition of the titanium middlings is shown in Table 9-5.

[0137] Table 9-5 Composition of titanium minerals and their leaching rates (wt%)

[0138] Titanium middlings ore from Table 9-5 were selected as raw materials for alkaline leaching. The alkaline leaching conditions were: sodium hydroxide concentration 10 wt%, leaching liquid-to-solid ratio 5:1 ml / g, leaching temperature 95℃, and leaching time 3 h. After alkaline leaching, the slurry was filtered, washed, and dried to obtain titanium-rich material, the composition of which is shown in Table 9-6. A photograph of the appearance of the titanium-rich material is shown below. Figure 7 .

[0139] Table 9-6 Composition of titanium-rich material after alkaline leaching (wt%)

[0140] Compare the results in Table 9-5 with the results in Table 4-1 under the corresponding conditions. Using bentonite pellets resulted in poor leaching of elements such as V, Ca, and Mg from titanium-containing materials, and the impurity removal effect was significantly lower than when using Ca(OH)2 as a binder and calcifying agent. Compare the color of the titanium-rich material powder produced by using Ca(OH)2 pellets. Figure 5 ) and the color of titanium-rich powder using bentonite pellets ( Figure 7 The titanium-rich material produced by immersion in Ca(OH)2 pellets is white, while the titanium-rich material produced by immersion in bentonite pellets is black overall, showing a clear difference between the two.

[0141] Comparative Example 2 The pelletizing, reduction-grinding-leaching experimental conditions for the raw material vanadium-titanium magnetite concentrate were the same as in Example 4, except that: the acid leaching system used 20%wt hydrochloric acid, 0.3mol / L MgCl2, a leaching liquid-to-solid ratio of 3:1ml / g, a leaching temperature of 145℃, a leaching time of 2h, and no oxidant was added. The mineral composition of the titanium after the reaction is shown in Table 10-1, and the morphology under scanning electron microscopy is shown in... Figure 8 As shown.

[0142] Table 10-1 Titanium-containing materials, mineral composition of titanium after leaching, and leaching rate (wt%)

[0143] Without the addition of an oxidant, the TiO2 leaching rate increased to 14.84%, resulting in significant titanium loss. The V leaching rate decreased to 73.62%, indicating a decline in impurity removal efficiency. Scanning electron microscopy revealed that the particles were roughly rounded with a fine overall particle size, and no crystals formed within the particles.

Claims

1. A method for utilizing valuable elements in vanadium-titanium magnetite concentrate, characterized in that, Includes the following steps: Step 1: Mix vanadium-titanium magnetite concentrate with a calcifying agent to prepare alkaline green pellets of vanadium-titanium magnetite concentrate. The alkaline green pellets are then subjected to calcification oxidation roasting to obtain calcified oxide pellets. Step 2: Mix the calcified oxide pellets with the first reducing agent and reduce them at high temperature to obtain metallized reduced pellets. Crush and grind the metallized reduced pellets and then perform magnetic separation to obtain magnetic reduced iron powder and non-magnetic titanium-containing materials. Step 3: Acid leaching and solid-liquid separation of the titanium-containing material yields hydrochloric acid leaching solution and titanium ore. Step 4: The titanium ore is leached with alkali and separated into solid and liquid components to obtain titanium-rich material; Step 5: Spray-fire the hydrochloric acid leaching solution to generate vanadium-chromium-iron powder and hydrochloric acid gas; Step 6: Wash the vanadium-chromium-iron powder with water and separate the solid and liquid to obtain vanadium-chromium-iron powder; Step 7: Selectively acid-leach vanadium with the vanadium-rich ferrochrome powder, and separate the solid and liquid phases to obtain a vanadium-containing leaching solution and ferrochrome-rich powder; Step 8: Mix the chromium-rich iron powder with the second reducing agent and smelt it into ferrochrome.

2. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 1, characterized in that, Includes at least one of the following technical features: (1) The diameter of the alkaline green pellets mentioned in step 1 is 9~16mm; (2) The composition of the vanadium-titanium magnetite concentrate mentioned in step 1, by mass percentage, includes TFe 50~60%, TiO2 8~15%, V2O5 0.5~2.0%, SiO2 0.5~5.0%, Cr2O3 1.0~3.0%, and CaO 0.1~5%; (3) The calcifying agent mentioned in step 1 is at least one of calcium-containing compounds or calcium-containing minerals; the calcium-containing compound is at least one of CaO, CaSO4, and Ca(OH)2; (4) In step 1, the calcification oxidation calcination temperature is 1100~1250℃ and the calcination time is 20min~10h; (5) The compressive strength of the calcified oxide pellets after calcination in step 1 is ≥1500N / piece.

3. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 2, characterized in that, In step 1, during the alkaline pelletizing of vanadium-titanium magnetite concentrate, the amount of calcifying agent added relative to 100 parts of vanadium-titanium magnetite concentrate is calculated using the following formula: Wt Ca(OH)2 =K(123Wt SiO2 +41Wt V2O5 -132Wt CaO ), Among them Wt Ca(OH)2 The amount of calcifying agent added is specified, with an excess coefficient K of 1.2~1.5, Wt. SiO2 Wt V2O5 and Wt CaO The percentage contents of SiO2, V2O5 and CaO in vanadium-titanium magnetite concentrate are respectively. And Wt Ca(OH)2 The amount of calcifying agent added shall not be less than 1.2 parts.

4. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 1, characterized in that, Includes at least one of the following technical features: (1) The mass ratio of the calcified oxide pellets to the first reducing agent in step 2 is 1:0.1~1:2; (2) In step 2, the first reducing agent is at least one of carbon-based substances, hydrogen, and CO; (3) The reduction temperature in step 2 is 900~1300℃ and the reduction time is 30min~20h; (4) The metallization rate of the metallized reduced pellets after step 2 reduction is ≥90%; (5) The grinding method in step 2 is at least one of ball milling, rod milling, and pebble milling; (6) In step 2, the grinding fine particles account for 50-90% of the total. (7) In step 2, the magnetic separation method is wet magnetic separation, and the magnetic field strength is 200~5000GS.

5. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 1, characterized in that, Includes at least one of the following technical features: (1) In step 3, the acid leaching is carried out using hydrochloric acid leaching solution, and the titanium-containing material is acid-leached with hydrochloric acid leaching solution; (2) In step 3, acid leaching is carried out by high temperature and pressure leaching to remove impurities and obtain leaching slurry. The leaching temperature is 110~180℃, the pressure corresponds to the vapor pressure of hydrochloric acid leaching solution at the leaching temperature, the leaching liquid-solid ratio is 1:1~10:1ml / g, and the leaching time is 1~4h. (3) In step 3, the titanium yield is ≥99%, the vanadium leaching rate is ≥90%, and the chromium leaching rate is ≥50% during the acid leaching process.

6. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 1, characterized in that, Includes at least one of the following technical features: (1) In step 4, titanium ore is placed in an alkaline leaching solution and heated under normal pressure to remove silicon. The leaching temperature is 40~100℃, the leaching liquid-solid ratio is 0.5:1~20:1ml / g, and the leaching time is 1~4h. (2) The alkaline leaching solution in step 4 is at least one of sodium hydroxide, potassium hydroxide, or lithium hydroxide, and the concentration of the alkaline leaching solution is 5-80 wt%. (3) The composition of the titanium-rich material obtained in step 4, in terms of mass percentage, is as follows: TFe ≤2.0%, TiO2≥85.00%, SiO2≤1.00%, MgO≤1.35%, CaO≤0.15%.

7. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 5, characterized in that, Includes at least one of the following technical features: (1) The hydrochloric acid leaching solution includes hydrochloric acid, chloride salt and oxidant; (2) The concentration of the hydrochloric acid solution is 15~25wt%; (3) The chloride salt is at least one of FeCl3, FeCl2, AlCl3, MgCl2, and CaCl2, and the chloride salt content is 0.1~1.5 mol / L; (4) The oxidant is at least one of FeCl3, O2, NaClO, H2O2, Cl2, and KMnO4, and the content of the oxidant is 0.1~3.0 mol / L, or the partial pressure of the oxidant is 0.1~1 MPa; (5) The hydrochloric acid leaching solution described in step 5 is sprayed and incinerated at high temperature to produce solid oxides and hydrochloric acid gas. The spray incineration temperature is 550~750℃. (6) The oxide is vanadium-chromium-iron powder, and its composition, by mass percentage, includes Fe2O3 ≥ 60%, V2O5 ≥ 2.5%, and Cr2O3 ≥ 2.0%; (7) The hydrochloric acid gas is absorbed by water spraying in a multi-stage countercurrent to generate a regenerated hydrochloric acid solution with a concentration of 18~22wt%. The regenerated hydrochloric acid solution is returned to the acid leaching process in step 3.

8. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 1, characterized in that, Step 6 involves washing the vanadium-chromium-iron powder with water to remove CaCl2. A multi-stage countercurrent water flow is used, and the amount of water added is 2 to 5 times the weight of the vanadium-chromium-iron powder. The powder is then filtered to obtain vanadium-chromium-iron powder rich in vanadium-chromium.

9. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 1, characterized in that, Includes at least one of the following technical features: (1) In step 7, the acid used for selective acid leaching of vanadium-rich ferrochrome powder is at least one of sulfuric acid, hydrochloric acid, nitric acid, and oxalic acid, and the acid concentration is 0.1~2 mol / L; (2) In step 7, selective acid leaching of vanadium is carried out under normal pressure, with a leaching temperature of 20~100℃ and a leaching liquid-to-solid ratio of 0.5:1~20:1 ml / g. After leaching, the leaching slurry is subjected to solid-liquid separation to obtain a vanadium-containing leaching solution and chromium-rich iron powder. (3) The composition of the chromium-rich iron powder, by mass percentage, includes Fe2O3 ≥ 70% and vanadium leaching ≥ 90%.

10. The method for utilizing valuable elements in vanadium-titanium magnetite concentrate according to claim 1, characterized in that, Includes at least one of the following technical features: (1) The chromium-rich iron powder described in step 8 is mixed with the second reducing agent and smelted into ferrochrome in an electric furnace at a smelting temperature of 1550~1700℃. (2) The ratio of the chromium-rich iron powder to the second reducing agent in step 8 is 1:0.15~1:0.30, and ferrochrome and smelting slag are obtained after smelting; (3) In step 8, the second reducing agent is at least one of coking coal, hydrogen, and CO.

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