Method for comprehensively recovering iron, vanadium, titanium, scandium and gallium through cooperation of vanadium titano-magnetite and red mud

Abstract

The invention belongs to the field of metallurgy, and particularly relates to a method for comprehensively recovering iron, vanadium, titanium, scandium and gallium through cooperation of vanadium titano-magnetite and red mud, which comprises the following steps: a) mixing vanadium titano-magnetite concentrate powder and red mud powder, roasting, leaching in water, and carrying out solid-liquid separation to obtain leached residues and leached liquid; b) mixing the leachate with a precipitant to precipitate vanadium, and carrying out solid-liquid separation to obtain vanadium-rich slag and vanadium-precipitated clear liquid; dissolving the vanadium-rich slag in a sulfuric acid solution, mixing the vanadium-rich slag with ammonium salt to react, carrying out solid-liquid separation, and roasting the obtained solid product ammonium polyvanadate to obtain V2O5; resin is used for adsorbing and desorbing gallium ions in the vanadium precipitation clear liquid, the pH of desorption liquid is adjusted to be alkaline, then electrolysis is conducted, and crude gallium is obtained; c) washing and drying the leaching residues to prepare pellets; after the pellets are pre-reduced, a carbonaceous reducing agent is mixed, melt separation treatment is carried out, and scandium-containing high-titanium slag and molten iron are obtained; and Ti and Sc in the scandium-containing high titanium slag are recycled. The method has the advantages of simple process, cleanness, high efficiency and the like.

Description

Technical Field

[0001] The invention belongs to the field of metallurgy, and in particular relates to a method for comprehensively recovering iron, vanadium, titanium, scandium and gallium by using vanadium-titanium magnetite in conjunction with red mud. Background Art

[0002] Red mud is a highly alkaline solid waste produced when alumina is extracted from bauxite through processes such as the Bayer process and sintering. Every ton of alumina produced generates 1 to 2 tons of red mud. my country, the world's largest alumina producer, produced 107 million tons of red mud annually in 2023, with cumulative stockpiles exceeding 1.1 billion tons, covering an area of ​​over 120,000 mu (approximately 16,000 acres). This waste residue is rich in iron oxide (15-40%), aluminum oxide (15-20%), sodium oxide (4-10%), calcium oxide (5-12%), titanium dioxide (3-8%), and rare metals such as scandium and gallium. Rare earth element content reaches 9 to 12 parts per million (ppm). Comprehensive recycling of red mud is urgently needed to address environmental risks such as soil salinization and groundwater contamination caused by the storage of highly alkaline red mud, and to promote the green transformation of the aluminum industry.

[0003] Vanadium-titanium magnetite is a polymetallic symbiotic resource mainly composed of iron, vanadium and titanium. Its mineral composition is complex. Magnetite and ilmenite are closely symbiotic. Vanadium is present in magnetite in an isomorphous manner, and is accompanied by rare elements such as cobalt, nickel, chromium and scandium. The efficient utilization of vanadium-titanium magnetite resources supports the development of aerospace, new energy and other industries. Summary of the Invention

[0004] In view of this, the purpose of the present invention is to provide a method for comprehensive recovery of iron, vanadium, titanium, scandium and gallium from vanadium-titanium magnetite and red mud. This method can realize the selective separation and extraction of iron, vanadium, titanium, scandium and gallium from vanadium-titanium magnetite and red mud, and has the advantages of simple process, clean and high efficiency.

[0005] The present invention provides a method for comprehensively recovering iron, vanadium, titanium, scandium and gallium from vanadium-titanium magnetite in collaboration with red mud, comprising the following steps:

[0006] a) mixing vanadium-titanium magnetite concentrate powder with red mud powder, roasting, leaching, and solid-liquid separation to obtain leaching residue and leachate;

[0007] b) mixing the leachate with a precipitant to precipitate vanadium, and performing solid-liquid separation to obtain vanadium-rich slag and a vanadium-precipitated clear solution; dissolving the vanadium-rich slag in a sulfuric acid solution, reacting the mixture with an ammonium salt, and performing solid-liquid separation to obtain a solid product, ammonium polyvanadate, and roasting to obtain V2O5; adsorbing and desorbing gallium ions in the vanadium-precipitated clear solution using a resin, and adjusting the pH of the desorption solution to alkaline before electrolysis to obtain crude gallium;

[0008] c) washing and drying the leached residue to prepare pellets; pre-reducing the pellets, mixing them with a carbonaceous reducing agent, and performing a smelting treatment to obtain scandium-containing high-titanium slag and molten iron; and recovering Ti and Sc in the scandium-containing high-titanium slag;

[0009] There is no restriction on the order of step b) and step c).

[0010] Preferably, in step a), the particle size of the vanadium-titanium magnetite concentrate powder and the red mud powder is ≤150 μm.

[0011] Preferably, in step a), after the vanadium-titanium magnetite concentrate powder is mixed with the red mud powder, the molar ratio of Na2O to V2O5 in the mixture is ≥1.2:1.

[0012] Preferably, in step a), the calcination temperature is 600-900° C. and the calcination time is 1-5 hours.

[0013] Preferably, in step a), the liquid-to-solid ratio of the water immersion is (3-5) mL:1 g, the temperature is 80-90° C., the stirring speed is 100-500 r / min, and the time is 30-180 min.

[0014] Preferably, in step b), the precipitant is one or more of calcium chloride, lead sulfate, sodium phosphate, disodium hydrogen phosphate, sodium sulfide, ferrous sulfate, ferric chloride, magnesium sulfate, magnesium chloride, ammonium chloride, ammonium sulfate, hydrogen sulfide, oxalic acid, oxalate, 8-hydroxyquinoline and tannic acid; the molar ratio of the precipitant to V in the leachate is (1-1.5):1; the temperature of the vanadium precipitation is 50-90°C, the stirring speed is 100-500 r / min, and the time is 30-150 min.

[0015] Preferably, in step b), the dissolution temperature is 30-90°C and the time is 50-150 min; the pH value of the mixed reaction with ammonium salt is 1.5-3, and the NH4 + The molar ratio of V is 1:(1.2~3).

[0016] Preferably, in step c), the pellet preparation process includes:

[0017] After the leaching residue is washed and dried, it is mixed with a binder to form pellets, and then roasted to obtain natural alkalinity pellets;

[0018] Alternatively, after the leaching residue is washed and dried, it is mixed with a binder and a calcium flux to form pellets, and then roasted to obtain alkaline pellets.

[0019] Preferably, in step c), the carbonaceous reducing agent is one or more of coke, coal and activated carbon; the amount of the carbonaceous reducing agent is 2-5% of the mass of the pre-reduced pellets; the temperature of the smelting treatment is 1500-1650°C, and the time is 1-3 hours.

[0020] Preferably, in step c), the process of recovering the Ti and Sc comprises:

[0021] The scandium-containing high-titanium slag is prepared into titanium dioxide powder by adopting a sulfuric acid method for preparing titanium dioxide;

[0022] Resin is used to adsorb and desorb scandium ions in the waste acid of titanium dioxide generated during the preparation of titanium dioxide. After the pH of the desorption liquid is adjusted to alkaline, oxalic acid precipitation and roasting are performed to obtain scandium oxide.

[0023] Compared with the prior art, the present invention provides a method for comprehensively recovering iron, vanadium, titanium, scandium and gallium from vanadium-titanium magnetite in conjunction with red mud, comprising the following steps: a) mixing vanadium-titanium magnetite concentrate powder with red mud powder, roasting, leaching in water, and performing solid-liquid separation to obtain leaching residue and leachate; b) mixing the leachate with a precipitant to precipitate vanadium, and performing solid-liquid separation to obtain vanadium-rich slag and a vanadium-precipitated clear liquid; dissolving the vanadium-rich slag in a sulfuric acid solution, reacting with an ammonium salt, and performing solid-liquid separation to roast the obtained solid product, ammonium polyvanadate, to obtain V2O5; using a resin to adsorb and desorb gallium ions in the vanadium-precipitated clear liquid, adjusting the pH of the desorption liquid to alkaline, and then electrolyzing to obtain crude gallium; c) washing and drying the leaching residue to prepare pellets; pre-reducing the pellets, mixing a carbonaceous reducing agent, and performing smelting treatment to obtain scandium-containing high-titanium slag and molten iron; recovering Ti and Sc in the scandium-containing high-titanium slag; there is no order restriction for steps b) and c). This method maximizes the inherent properties of red mud and vanadium-titanium magnetite, using the alkalinity of red mud to synergistically leach vanadium from the vanadium-titanium magnetite, forming soluble vanadium salts and removing alkaline substances contained in the red mud. This produces a vanadium- and gallium-rich leachate. The remaining leached residue is used as an iron-rich concentrate for pelletizing. Using a pre-reduction-melting process, molten iron and scandium-containing high-titanium slag are obtained. The scandium-containing high-titanium slag can be used to produce sulfuric acid titanium dioxide and obtain a scandium-rich solution. This method can achieve the selective separation and extraction of iron, vanadium, titanium, scandium, and gallium from vanadium-titanium magnetite and red mud, offering advantages such as simplicity, cleanliness, and efficiency. DETAILED DESCRIPTION

[0024] The following is a clear and complete description of the technical solutions in the embodiments of the present invention. Obviously, the described embodiments are only part of the embodiments of the present invention, not all of the embodiments. Based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without making any creative efforts are within the scope of protection of the present invention.

[0025] The present invention provides a method for comprehensively recovering iron, vanadium, titanium, scandium and gallium from vanadium-titanium magnetite in collaboration with red mud, comprising the following steps:

[0026] a) mixing vanadium-titanium magnetite concentrate powder with red mud powder, roasting, leaching, and solid-liquid separation to obtain leaching residue and leachate;

[0027] b) mixing the leachate with a precipitant to precipitate vanadium, and performing solid-liquid separation to obtain vanadium-rich slag and a vanadium-precipitated clear solution; dissolving the vanadium-rich slag in a sulfuric acid solution, reacting the mixture with an ammonium salt, and performing solid-liquid separation to obtain a solid product, ammonium polyvanadate (APV), and roasting the resulting solid product to obtain V2O5; adsorbing and desorbing gallium ions in the vanadium-precipitated clear solution using a resin, and adjusting the pH of the desorption solution to alkaline before electrolysis to obtain crude gallium;

[0028] c) washing and drying the leached residue to prepare pellets; pre-reducing the pellets, mixing them with a carbonaceous reducing agent, and performing a smelting treatment to obtain scandium-containing high-titanium slag and molten iron; and recovering Ti and Sc in the scandium-containing high-titanium slag;

[0029] There is no restriction on the order of step b) and step c).

[0030] In the method provided by the present invention, in step a), the total iron (TFe) content of the vanadium-titanium magnetite concentrate powder is preferably 50-65wt%, more preferably 56-59wt%; the FeO content of the vanadium-titanium magnetite concentrate powder is preferably 30-36wt%, more preferably 33-34wt%; the Fe2O3 content of the vanadium-titanium magnetite concentrate powder is preferably 40-50wt%, more preferably 45-46wt%; the V2O5 content of the vanadium-titanium magnetite concentrate powder is preferably 0.5-1wt%, more preferably 0.7-0.8wt%. ; The TiO2 content of the vanadium-titanium magnetite concentrate powder is preferably 7-15wt%, more preferably 10-11wt%; the Sc2O3 content of the vanadium-titanium magnetite concentrate powder is preferably 10-30g / t, more preferably 15-20g / t; the Ga2O3 content of the vanadium-titanium magnetite concentrate powder is preferably 30-60g / t, more preferably 40-50g / t; the particle size of the vanadium-titanium magnetite concentrate powder is preferably ≤150μm, more preferably ≤120μm, further preferably ≤100μm, and most preferably ≤74μm (200 mesh).

[0031] In the method provided by the present invention, in step a), the total iron (TFe) content of the red mud powder is preferably 35-45wt%, more preferably 38-42wt%; the Fe2O3 content of the red mud powder is preferably 50-60wt%, more preferably 55-56wt%; the V2O5 content of the red mud powder is preferably 0.01-0.1wt%, more preferably 0.02-0.08wt%; the TiO2 content of the red mud powder is preferably 3-15wt%, more preferably 6-10wt%; the Sc2O3 content of the red mud powder is preferably 50-150g / t, more preferably 80-100g / t; the Ga2O3 content of the red mud powder is preferably 70-300g / t, more preferably 110-200g / t; the particle size of the red mud powder is preferably ≤150μm, more preferably ≤120μm, more preferably ≤100μm, and most preferably ≤74μm (200 mesh).

[0032] In the method provided by the present invention, in step a), after the vanadium-titanium magnetite concentrate powder is mixed with the red mud powder, the molar ratio of Na2O to V2O5 in the mixture is preferably ≥1.2:1, more preferably (2-5):1, specifically 2:1, 2.3:1, 2.5:1, 2.7:1, 3:1, 3.2:1, 3.5:1, 3.7:1, 4:1, 4.2:1, 4.5:1, 4.7:1 or 5:1.

[0033] In the method provided by the present invention, in step a), the calcination temperature is preferably 600-900°C, specifically 600°C, 620°C, 650°C, 670°C, 700°C, 720°C, 750°C, 770°C, 800°C, 820°C, 850°C, 870°C or 900°C; the calcination time is preferably 1-5h, specifically 1h, 1.5h, 2h, 2.5h, 3h, 3.5h, 4h, 4.5h or 5h.

[0034] In the method provided by the present invention, in step a), the liquid-to-solid ratio of the water immersion is preferably (3-5) mL:1g; specifically, it can be 3 mL:1g, 3.2 mL:1g, 3.5 mL:1g, 3.7 mL:1g, 4 mL:1g, 4.2 mL:1g, 4.5 mL:1g, 4.7 mL:1g or 5 mL:1g; the temperature of the water immersion is preferably 80-90°C, specifically, it can be 80°C, 81°C, 82°C, 83°C, 84°C, 85°C, 86°C, 87°C, 88°C, 89°C or 90°C; the stirring speed of the water immersion is preferably 100-500 r / min, specifically, it can be The immersion time is preferably 30 to 180 min, specifically 30 min, 40 min, 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min, 140 min, 150 min, 160 min, 170 min or 180 min.

[0035] In the method provided by the present invention, in step b), the precipitant is preferably one or more of calcium chloride, lead sulfate, sodium phosphate, disodium hydrogen phosphate, sodium sulfide, ferrous sulfate, ferric chloride, magnesium sulfate, magnesium chloride, ammonium chloride, ammonium sulfate, hydrogen sulfide, oxalic acid, oxalate, 8-hydroxyquinoline and tannic acid; the molar ratio of the precipitant to V in the leachate is preferably (1-1.5):1, specifically 1:1, 1.05:1, 1.1:1, 1.15:1, 1.2:1, 1.25:1, 1.3:1, 1.35:1, 1.4:1, 1.45:1 or 1.5:1.

[0036] In the method provided by the present invention, in step b), the temperature of the vanadium precipitation is preferably 50-90°C, specifically 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C or 90°C; the stirring speed of the vanadium precipitation is preferably 100-500 r / min, specifically 100 r / min, 150 r / min, 200 r / min, 250 r / min, 300 r / min, 350 r / min, 400 r / min, 450 r / min or 500 r / min; the time of the vanadium precipitation is preferably 30-150 min, specifically 30 min, 40 min, 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min, 140 min or 150 min.

[0037] In the method provided by the present invention, in step b), the dissolution temperature is preferably 30-90°C, specifically 30°C, 35°C, 40°C, 45°C, 50°C, 55°C, 60°C, 65°C, 70°C, 75°C, 80°C, 85°C or 90°C; the dissolution time is preferably 50-150 min, specifically 50 min, 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, 130 min, 140 min or 150 min.

[0038] In the method provided by the present invention, in step b), the ammonium salt is preferably ammonium sulfate; when the ammonium salt is mixed with the ammonium salt, the NH4 + The molar ratio of V to V is preferably 1:(1.2-3), specifically 1:1.2, 1:1.3, 1:1.4, 1:1.5, 1:1.6, 1:1.7, 1:1.8, 1:1.9, 1:2, 1:2.1, 1:2.2, 1:2.3, 1:2.4, 1:2.5, 1:2.6, 1:2.7, 1:2.8, 1:2.9 or 1:3; when reacting with the ammonium salt, the pH value of the reaction system is preferably 1.5-3, specifically 1.5, 1.6, 1.7, 1.8, 1.9, 2, 2.1, 2.2, 2.3, 2.4, 2.5, 2.6, 2.7, 2.8, 2.9 or 3, and the pH value is preferably regulated by adding sulfuric acid to the reaction system.

[0039] In the method provided by the present invention, in step b), after the mixing reaction with the ammonium salt, the liquid phase obtained by solid-liquid separation is preferably returned to the dissolution process of the vanadium-rich slag, thereby realizing the recycling of the acid solution.

[0040] In the method provided by the present invention, in step b), the resin is a porous material having a selective affinity for gallium, such as a polymer porous material containing nitrogen or sulfur ligands; the particle size of the resin is preferably 0.3 to 1.2 mm, specifically 0.3 mm, 0.4 mm, 0.5 mm, 0.6 mm, 0.7 mm, 0.8 mm, 0.9 mm, 1 mm, 1.1 mm or 1.2 mm; the adsorption is preferably carried out in an adsorption column, and the flow rate of the vanadium precipitation supernatant through the column is preferably 5 to 15 times the column volume / hour (BV / h), specifically 5 BV / h, 6 BV / h, 7 BV / h, 8 BV / h, 9 BV / h, 10 BV / h, 11 BV / h, 12 BV / h, 13 BV / h, 14 BV / h or 15 BV / h; the desorbent used for the desorption is preferably an acidic solution, more preferably a hydrochloric acid solution and / or a sulfuric acid solution; the concentration of the desorbent is preferably 1-4 mol / L, specifically 1 mol / L, 1.5 mol / L, 2 mol / L, 2.5 mol / L, 3 mol / L, 3.5 mol / L or 4 mol / L; the desorbent temperature during the desorption process is preferably 40-60°C, specifically 40°C, 45°C, 50°C, 55°C or 60°C; the desorption is preferably carried out directly in the adsorption column, and the flow rate of the desorbent through the column is preferably 2-5 BV / h, specifically 2 BV / h, 2.5 BV / h, 3 BV / h, 3.5 BV / h, 4 BV / h, 4.5 BV / h or 5 BV / h.

[0041] In the method provided by the present invention, in step b), the main purpose of adjusting the pH of the desorption solution to alkaline is to precipitate a trace amount of aluminum in the desorption solution; the reagent for adjusting the pH of the desorption solution is preferably sodium hydroxide and / or ammonia water; the pH value of the desorption solution is preferably adjusted to 9 to 10.5, specifically 9, 9.5, 10 or 10.5.

[0042] In the method provided by the present invention, in step b), a conductive salt is preferably added during the electrolysis process, and the conductive salt includes but is not limited to sodium sulfate; the cathode current density of the electrolysis is preferably 50 to 200 A / m 2 , specifically 50A / m 2 , 60A / m 2 , 70A / m 2 , 80A / m 2 , 90A / m 2 , 100A / m 2 , 110A / m 2 , 120A / m 2 , 130A / m 2 、140A / m 2 、150A / m 2 、160A / m 2、170A / m 2 、180A / m 2 、190A / m 2 or 200A / m 2 ; The cell voltage of the electrolysis is preferably 3 to 5V, specifically 3V, 3.2V, 3.5V, 3.7V, 4V, 4.2V, 4.5V, 4.7V or 5V.

[0043] In the method provided by the present invention, in step c), the preparation process of the pellets includes:

[0044] After the leaching residue is washed and dried, it is mixed with a binder to form pellets, and then roasted to obtain natural alkalinity pellets;

[0045] Alternatively, after the leaching residue is washed and dried, it is mixed with a binder and a calcium flux to form pellets, and then roasted to obtain alkaline pellets.

[0046] In the pellet preparation process provided by the present invention, the binder is preferably bentonite and / or an organic binder; the calcium flux is preferably one or more of quicklime, limestone and slaked lime; the calcium flux preferably adjusts the alkalinity of the mixture within the range of natural alkalinity to 0.6; the roasting temperature is preferably 1200-1300°C, specifically 1200°C, 1210°C, 1220°C, 1230°C, 1240°C, 1250°C, 1260°C, 1270°C, 1280°C, 1290°C or 1300°C; the roasting time is preferably 15-25 min, specifically 15 min, 16 min, 17 min, 18 min in, 19min, 20min, 21min, 22min, 23min, 24min or 25min; the equipment used for the roasting is preferably preheated before roasting, and the preheating temperature is preferably 850-950°C, specifically 850°C, 860°C, 870°C, 880°C, 890°C, 900°C, 910°C, 920°C, 930°C, 940°C or 950°C, and the preheating time is preferably 10-20min, specifically 10min, 11min, 12min, 13min, 14min, 15min, 16min, 17min, 18min, 19min or 20min.

[0047] In the method provided by the present invention, in step c), the average compressive strength of the pellets is preferably 2000-2600N, specifically 2000N, 2050N, 2100N, 2150N, 2200N, 2250N, 2300N, 2350N, 2400N, 2450N, 2500N, 2550N or 2600N.

[0048] In the method provided by the present invention, in step c), the equipment for pre-reduction of the pellets is preferably a gas-based shaft furnace; the metallization rate of the pellets after pre-reduction is preferably controlled at 85-95%, specifically 85%, 86%, 87%, 88%, 89%, 90%, 91%, 92%, 93%, 94% or 95%.

[0049] In the method provided by the present invention, in step c), the carbonaceous reducing agent is preferably one or more of coke, coal and activated carbon; the amount of the carbonaceous reducing agent is preferably 2-5% of the mass of the pre-reduced pellets, specifically 2%, 2.3%, 2.5%, 2.7%, 3%, 3.2%, 3.5%, 3.7%, 4%, 4.2%, 4.5%, 4.7% or 5%.

[0050] In the method provided by the present invention, in step c), the equipment for the melting treatment is preferably a melting furnace; the temperature of the melting treatment is preferably 1500-1650°C, specifically 1500°C, 1510°C, 1520°C, 1530°C, 1540°C, 1550°C, 1560°C, 1570°C, 1580°C, 1590°C, 1600°C, 1610°C, 1620°C, 1630°C, 1640°C or 1650°C; the time of the melting treatment is preferably 1-3h, specifically 1h, 1.2h, 1.5h, 1.7h, 2h, 2.3h, 2.5h, 2.7h or 3h.

[0051] In the method provided by the present invention, in step c), the molten iron is cooled to form cast iron, which can be used for steelmaking.

[0052] In the method provided by the present invention, in step c), the process of recovering the Ti and Sc preferably comprises:

[0053] The scandium-containing high-titanium slag is prepared into titanium dioxide powder by adopting a sulfuric acid method for preparing titanium dioxide;

[0054] Resin is used to adsorb and desorb scandium ions in the waste acid of titanium dioxide generated during the preparation of titanium dioxide. After the pH of the desorption liquid is adjusted to alkaline, oxalic acid precipitation and roasting are performed to obtain scandium oxide.

[0055] In the above-mentioned process of recovering Ti and Sc provided by the present invention, the specific process of preparing titanium dioxide by the sulfuric acid method preferably includes: mixing the high-titanium slag containing scandium with concentrated sulfuric acid for acid hydrolysis to obtain titanium liquid; mixing the titanium liquid with iron element to reduce the Fe in the titanium liquid. 3+ Reduction to Fe 2+, then cooling and crystallizing to precipitate ferrous sulfate; diluting the titanium liquid after filtering out the precipitate, then heating and hydrolyzing to generate metatitanic acid precipitate, solid-liquid separation, respectively obtaining metatitanic acid solid and titanium dioxide waste acid; finally, calcining the metatitanic acid solid to obtain titanium dioxide powder. Wherein, the concentration of the concentrated sulfuric acid is preferably 90-98wt%, specifically 90wt%, 91wt%, 92wt%, 93wt%, 94wt%, 95wt%, 96wt%, 97wt% or 98wt%; the mass ratio of the scandium-containing high-titanium slag to the concentrated sulfuric acid is preferably 1:(1.5-1.8), specifically 1:1.5, 1:1.6, 1:1.7 or 1:1.8; the temperature of the acidolysis is preferably 160-180°C, specifically 160°C, 165°C, 170°C, 175°C or 180°C; the time of the acidolysis is preferably 2-4h, specifically 2h, 2.5h, 3h, 3.5h or 4h; the temperature of the cooling crystallization is preferably 10-15°C, Specifically, it can be 10℃, 11℃, 12℃, 13℃, 14℃ or 15℃; the titanium content of the titanium liquid after dilution in terms of TiO2 is preferably 200-220g / L, specifically 200g / L, 205g / L, 210g / L, 215g / L or 220g / L; the temperature of the heating hydrolysis is preferably the boiling point; the calcination temperature is preferably 850-950℃, specifically 850℃, 860℃, 870℃, 880℃, 890℃, 900℃, 910℃, 920℃, 930℃, 940℃ or 950℃; the calcination time is preferably 1.5-2h, specifically 1.5h, 1.6h, 1.7h, 1.8h, 1.9h or 2h.

[0056] In the above-mentioned process for recovering Ti and Sc provided by the present invention, before the adsorption and desorption of scandium ions, the titanium dioxide waste acid is preferably subjected to adsorption and removal of organic matter in the titanium dioxide waste acid using activated carbon. The particle size of the activated carbon is preferably 2 to 4 mm, specifically 2 mm, 2.5 mm, 3 mm, 3.5 mm or 4 mm; the adsorption and removal of organic matter is preferably carried out in an activated carbon bed, and the air velocity is preferably 5 to 8 h -1 , specifically 5h -1 , 6h -1 , 7h -1 or 8 hours -1 .

[0057] In the above-mentioned process of recovering Ti and Sc provided by the present invention, the adsorption and desorption of scandium ions in the titanium white waste acid are preferably carried out in a D001 type cationic resin tower; the resin particle size in the D001 type cationic resin tower is preferably 0.4-0.6 mm, specifically 0.4 mm, 0.45 mm, 0.5 mm, 0.55 mm or 0.6 mm; the bed height-to-diameter ratio of the D001 type cationic resin tower is preferably (2-5):1, specifically 2:1, 3:1, 4:1 or 5:1; the flow rate of the titanium white waste acid through the tower is preferably 3-5 BV / h, specifically 3 BV / h, 3.5 BV / h, 4 BV / h, 4.5 BV / h or 5 BV / h.

[0058] In the above-mentioned process of recovering Ti and Sc provided by the present invention, the desorption process preferably includes: first eluting the residual iron in the resin tower with a low-concentration hydrochloric acid solution, and then eluting the scandium adsorbed in the resin tower with a high-concentration hydrochloric acid solution to obtain a scandium-rich desorption liquid. Wherein, the concentration of the low-concentration hydrochloric acid solution is preferably 0.5-2 mol / L, specifically 0.5 mol / L, 1 mol / L, 1.5 mol / L or 2 mol / L; the flow rate of the low-concentration hydrochloric acid solution through the tower is preferably 1-3 BV / h, specifically 1 BV / h, 1.5 BV / h, 2 BV / h, 2.5 BV / h or 3 BV / h; the amount of the low-concentration hydrochloric acid solution is preferably 3-5 BV, specifically 3 BV, 3.5 BV, 4 BV, 4.5 BV or 5 BV; the high ... The concentration of the solution is preferably 3 to 6 mol / L, specifically 3 mol / L, 3.5 mol / L, 4 mol / L, 4.5 mol / L, 5 mol / L, 5.5 mol / L or 6 mol / L; the flow rate of the high concentration hydrochloric acid solution through the tower is preferably 0.5 to 2 BV / h, specifically 0.5 BV / h, 1 BV / h, 1.5 BV / h or 2 BV / h; the amount of the high concentration hydrochloric acid solution is preferably 4 to 6 BV, specifically 4 BV, 4.5 BV, 5 BV, 5.5 BV or 6 BV.

[0059] In the above-mentioned process for recovering Ti and Sc provided by the present invention, the reagent for adjusting the pH of the desorption solution is preferably sodium hydroxide; the pH value of the desorption solution is preferably adjusted to 4-4.5, specifically 4, 4.1, 4.2, 4.3, 4.4 or 4.5.

[0060] In the above-mentioned process for recovering Ti and Sc provided by the present invention, when performing oxalic acid precipitation, the oxalic acid is preferably added to the desorption liquid in the form of an oxalic acid solution; the molar ratio of the oxalic acid to the Sc in the desorption liquid is preferably (2-4):1, more preferably 3:1; the precipitation temperature is preferably 40-80°C, specifically 40°C, 50°C, 60°C, 70°C, or 80°C; the precipitation time is preferably 0.5-2 hours, specifically 0.5 hours, 1 hour, 1.5 hours, or 2 hours. After the precipitation is completed, a scandium oxalate precipitate is obtained.

[0061] In the above-mentioned process of recovering Ti and Sc provided by the present invention, the temperature for calcining the scandium oxalate precipitate is preferably 600-1000°C, specifically 600°C, 650°C, 700°C, 750°C, 800°C, 850°C, 900°C, 950°C or 1000°C; the calcination time is preferably 1-5h, specifically 1h, 1.5h, 2h, 2.5h, 3h, 3.5h, 4h, 4.5h or 5h.

[0062] The method provided by the present invention maximizes the inherent properties of red mud and vanadium-titanium magnetite. It uses the alkalinity of red mud to synergistically leach vanadium from the vanadium-titanium magnetite, forming soluble vanadium salts and removing alkaline substances contained in the red mud to obtain a vanadium- and gallium-rich leachate. The remaining leached residue is used as an iron-rich concentrate for pelletizing. A pre-reduction-melting process is used to obtain molten iron and scandium-containing high-titanium slag. The scandium-containing high-titanium slag can be used to produce sulfuric acid titanium dioxide and obtain a scandium-rich solution. This method can achieve the selective separation and extraction of iron, vanadium, titanium, scandium, and gallium from vanadium-titanium magnetite and red mud, and has the advantages of simple process, cleanliness, and high efficiency.

[0063] For the purpose of greater clarity, the present invention is described in detail with reference to the following examples.

[0064] Example 1

[0065] The composition of vanadium-titanium magnetite concentrate in a mining area in Panxi is shown in Table 1:

[0066] Table 1 Panxi vanadium-titanium magnetite concentrate composition (wt%)

[0067] TFe FeO <![CDATA[Fe2O3]]> <![CDATA[SiO2]]> CaO MgO <![CDATA[Al2O3]]> <![CDATA[V2O5]]> <![CDATA[TiO2]]> 58.25 34.09 45.37 2.26 0.19 2.66 3.19 0.75 10.02

[0068] Among them, the Ga2O3 content in the vanadium-titanium magnetite concentrate is 40g / t, and the Sc2O3 content is 15g / t.

[0069] The composition of Bayer red mud from a certain factory is shown in Table 2:

[0070] Table 2 Red mud composition (wt%)

[0071] TFe <![CDATA[Fe2O3]]> <![CDATA[SiO2]]> CaO <![CDATA[Al2O3]]> <![CDATA[V2O5]]> <![CDATA[TiO2]]> <![CDATA[Na2O]]> 38.92 55.76 5.43 0.80 16.01 0.08 6.87 2.16

[0072] Among them, the Ga2O3 content in red mud is 110g / t and the Sc2O3 content is 100g / t.

[0073] The vanadium-titanium magnetite concentrate and red mud were ground and passed through a 200-mesh sieve. The mixture was then mixed at a mass ratio of 7:3 (mass ratio of vanadium-titanium magnetite concentrate to red mud). The molar ratio of Na₂O to V₂O₅ was approximately 3.6. The mixed material was then roasted at 800°C for 2 hours. The roasted clinker was then water-soaked at a liquid-to-solid ratio (mL / g) of 3:1, a temperature of 90°C, a duration of 40 minutes, and a stirring speed of 200 rpm. After the leaching, the mixture was filtered and separated to obtain the leached residue and filtrate A.

[0074] The precipitating agent CaCl2 was added to the filtrate A to precipitate vanadium. The molar ratio of the precipitating agent to vanadium was controlled to be 1.2:1, the precipitation temperature was 60°C, the precipitation time was 60 minutes, and the stirring speed was 200 r / min. After the precipitation reaction was completed, the vanadium-rich slag and filtrate B were obtained by filtration. The vanadium-rich slag was dissolved in sulfuric acid solution at a temperature of 80°C and a dissolution time of 60 minutes. Then, an appropriate amount of sulfuric acid and ammonium sulfate were added to control the pH value of the reaction system to 2.0, and NH4 + The molar ratio of V to V is controlled to be 1:2. After the reaction is completed, APV and filtrate C are obtained by filtration and separation. Filtrate C can be returned to the vanadium-rich slag for dissolution. APV can be roasted to obtain V2O5. Filtrate B is adsorbed in a resin column. The saturated resin is eluted with an acidic solution to obtain a gallium-containing desorption liquid. The pH of the gallium-containing desorption liquid is then adjusted to 9.5 and electrolyzed to obtain crude gallium. The resin filled in the resin column is a polymer porous material containing sulfur ligands, the resin particle size is 0.5 mm, the flow rate of filtrate B through the column is 10 BV / h, the acidic solution used for desorption is a 2 mol / L hydrochloric acid solution, the temperature of the acidic solution is controlled at 50 ° C during the desorption process, the flow rate of the acidic solution through the column is 3 BV / h, the reagent for adjusting the pH is NaOH, and Na2SO4 is added during the electrolysis process to increase the conductivity. The cathode current density of the electrolysis is 100 A / m 2 , the cell voltage is 4 V. In this process, the vanadium recovery rate is about 75%, and the gallium recovery rate is about 60%.

[0075] The leached slag is washed with distilled water and dried at 120°C, then mixed with an appropriate amount of bentonite to form pellets. A rotary kiln is preheated at 900°C for 15 minutes. The pellets are then added to the preheated rotary kiln for calcination at 1250°C for 15 minutes. After calcination, natural basicity pellets with an average compressive strength of 2300N are obtained. The pellets are pre-reduced in a gas-based shaft furnace, with the metallization ratio of the pre-reduced pellets controlled at 90%. The pre-reduced pellets are then added to a melting furnace, mixed with 5% of the pellet weight (fixed carbon content ≥80%). The melting furnace is then operated to separate the slag and iron from the pre-reduced pellets at a melting temperature of 1600°C for 2 hours. After smelting, scandium-containing high-titanium slag and molten iron are obtained. The molten iron cools to form cast iron suitable for steelmaking. The high-titanium slag containing scandium (TiO2 content ≥ 80wt%, Sc2O3 0.02 ~ 0.1wt%) and 98wt% concentrated sulfuric acid were mixed at a mass ratio of 1:1.6, and acid-hydrolyzed at 160℃ for 4 hours to generate titanium liquid (containing TiOSO4, Sc 3+ 、Fe 2+ / Fe 3+ ). Add iron chips to the titanium liquid to 3+ Reduction to Fe 2+ , cooled to 10℃ to crystallize out ferrous sulfate (FeSO4·7H2O). After filtering out the ferrous sulfate precipitate from the titanium liquid, dilute it to a titanium concentration (in terms of TiO2) of 200g / L, and hydrolyze it at the boiling point to generate metatitanic acid (H2TiO3) precipitate, which is filtered to obtain metatitanic acid solid and titanium dioxide waste acid respectively. The metatitanic acid solid is calcined at 900℃ for 2 hours to obtain titanium dioxide (TiO2≥99wt%). Titanium dioxide waste acid (containing H2SO415~20wt%, Sc 3+ 50~100mg / L, Fe 2+ 5~10g / L) through the activated carbon bed (particle size 3mm, air velocity 6h -1 ) adsorbs organic matter. The waste acid after removing organic matter flows through a D001 cationic resin tower (particle size 0.5mm, bed height-diameter ratio 3:1) at a flow rate of 4BV / h. The scandium adsorption rate is greater than 95%, and the iron adsorption rate is less than 1%. After adsorption saturation, the resin tower is first eluted with 4BV 1mol / L HCl at a flow rate of 2BV / h to elute residual iron, and then eluted with 5BV 4mol / LHCl at a flow rate of 1BV / h to obtain a Sc-containing 3+ The desorption solution was adjusted to pH 4.5 using NaOH. A 10 wt% oxalic acid solution was added to the desorption solution at a molar ratio of oxalic acid to Sc of 3:1. The solution was aged at 60°C for 1 hour to obtain a scandium oxalate precipitate. The scandium oxalate precipitate was calcined at 800°C for 2 hours to obtain scandium oxide with a purity of ≥99.5%. During this process, the iron recovery rate was approximately 90%, the titanium recovery rate was approximately 80%, and the scandium recovery rate was approximately 70%.

[0076] Example 2

[0077] The composition of vanadium-titanium magnetite concentrate in a mining area in Panxi is shown in Table 3:

[0078] Table 3 Panxi vanadium titanium magnetite concentrate composition (wt%)

[0079] TFe FeO <![CDATA[SiO2]]> CaO MgO <![CDATA[Al2O3]]> <![CDATA[V2O5]]> <![CDATA[TiO2]]> MnO 56.9 33.58 2.44 0.24 2.92 3.17 0.737 10.3 0.434

[0080] Among them, the Ga2O3 content in the vanadium-titanium magnetite concentrate is 50g / t, and the Sc2O3 content is 20g / t.

[0081] The composition of Bayer red mud from a certain factory is shown in Table 4:

[0082] Table 4 Red mud composition (wt%)

[0083] TFe <![CDATA[Fe2O3]]> <![CDATA[SiO2]]> CaO <![CDATA[Al2O3]]> <![CDATA[V2O5]]> <![CDATA[TiO2]]> <![CDATA[Na2O]]> 41.2 55.76 2.43 2.80 12.15 0.02 9.5 1.21

[0084] Among them, the Ga2O3 content in red mud is 200g / t and the Sc2O3 content is 80g / t.

[0085] The vanadium-titanium magnetite concentrate and red mud were ground and passed through a 200-mesh sieve. The mixture was then mixed at a mass ratio of 5:5 (mass ratio of vanadium-titanium magnetite concentrate to red mud). The molar ratio of Na₂O to V₂O₅ was approximately 4.8. The mixed material was then roasted at 700°C for 4 hours. The roasted clinker was then water-leached at a liquid-to-solid ratio (mL / g) of 4.5:1, at 80°C for 120 minutes, and at a stirring speed of 400 rpm. After the leaching, the mixture was filtered to obtain the leached residue and filtrate A.

[0086] The precipitating agent PbSO4 was added to the filtrate A to precipitate vanadium. The molar ratio of the precipitating agent to vanadium was controlled at 1.5:1. The precipitation temperature was 80°C, the precipitation time was 120 minutes, and the stirring speed was 400 r / min. After the precipitation reaction was completed, the vanadium-rich slag and filtrate B were obtained by filtration. The vanadium-rich slag was dissolved in sulfuric acid solution at a temperature of 70°C and a dissolution time of 120 minutes. Then, an appropriate amount of sulfuric acid and ammonium sulfate were added to control the pH value of the reaction system to 1.5, and NH4 +The molar ratio of V to V is controlled to be 1:2.5. After the reaction is completed, APV and filtrate C are obtained by filtration and separation. Filtrate C can be returned to the vanadium-rich slag for dissolution. APV can be calcined to obtain V2O5. Filtrate B is adsorbed in a resin column. The saturated resin is eluted with an acidic solution to obtain a gallium-containing desorption liquid. The pH of the gallium-containing desorption liquid is then adjusted to 9.5 and electrolyzed to obtain crude gallium. The resin filled in the resin column is a polymer porous material containing sulfur ligands, the resin particle size is 0.5 mm, the flow rate of filtrate B through the column is 10 BV / h, the acidic solution used for desorption is a 2 mol / L hydrochloric acid solution, the temperature of the acidic solution is controlled at 50 ° C during the desorption process, the flow rate of the acidic solution through the column is 3 BV / h, the reagent for adjusting pH is NaOH, and Na2SO4 is added during the electrolysis process to increase conductivity. The cathode current density of the electrolysis is 100 A / m 2 , the cell voltage is 4 V. During this process, the vanadium recovery rate is about 60%, and the gallium recovery rate is about 65%.

[0087] The leached slag is washed with distilled water and dried at 110°C. It is then mixed with an appropriate amount of organic binder and limestone to form pellets with an alkalinity of 0.6. A belt roaster is preheated at 950°C for 20 minutes. The pellets are then added to the preheated belt roaster and roasted at 1280°C for 20 minutes. After roasting, alkaline pellets with an average compressive strength of 2500N are obtained. The pellets are pre-reduced in a gas-based shaft furnace, with a metallization ratio of 85%. The pre-reduced pellets are then added to a melting furnace, where activated carbon (3% by weight) is added. The furnace is then operated to separate the slag and iron from the pre-reduced pellets at a melting temperature of 1650°C for 1.5 hours. After melting, scandium-containing high-titanium slag and molten iron are obtained. The molten iron cools to form cast iron suitable for steelmaking. The high-titanium slag containing scandium (TiO2 content ≥ 80wt%, Sc2O3 0.02 ~ 0.1wt%) and 98wt% concentrated sulfuric acid were mixed at a mass ratio of 1:1.6, and acid-hydrolyzed at 160℃ for 4 hours to generate titanium liquid (containing TiOSO4, Sc 3+ 、Fe 2+ / Fe 3+ ). Add iron chips to the titanium liquid to 3+ Reduction to Fe 2+ , cooled to 10℃ to crystallize out ferrous sulfate (FeSO4·7H2O). After filtering out the ferrous sulfate precipitate from the titanium liquid, dilute it to a titanium concentration (in terms of TiO2) of 200g / L, and hydrolyze it at the boiling point to generate metatitanic acid (H2TiO3) precipitate, which is filtered to obtain metatitanic acid solid and titanium dioxide waste acid respectively. The metatitanic acid solid is calcined at 900℃ for 2 hours to obtain titanium dioxide (TiO2≥99wt%). Titanium dioxide waste acid (containing H2SO415-20%, Sc 3+ 50~100mg / L, Fe2+ 5~10g / L) through the activated carbon bed (particle size 3mm, air velocity 6h -1 ) adsorbs organic matter. The waste acid after removing organic matter flows through a D001 cationic resin tower (particle size 0.5mm, bed height-diameter ratio 3:1) at a flow rate of 4BV / h. The scandium adsorption rate is greater than 95%, and the iron adsorption rate is less than 1%. After adsorption saturation, the resin tower is first eluted with 4BV 1mol / L HCl at a flow rate of 2BV / h to elute residual iron, and then eluted with 5BV 4mol / L HCl at a flow rate of 1BV / h to obtain a Sc-containing 3+ The desorption solution was adjusted to pH 4.5 using NaOH. A 10 wt% oxalic acid solution was added to the desorption solution at a molar ratio of oxalic acid to Sc of 3:1. The solution was aged at 60°C for 1 hour to obtain a scandium oxalate precipitate. The scandium oxalate precipitate was calcined at 800°C for 2 hours to obtain scandium oxide with a purity of ≥99.5%. During this process, the iron recovery rate was approximately 92%, the titanium recovery rate was approximately 85%, and the scandium recovery rate was approximately 75%.

[0088] The above is only a preferred embodiment of the present invention. It should be pointed out that for ordinary technicians in this technical field, several improvements and modifications can be made without departing from the principles of the present invention. These improvements and modifications should also be regarded as within the scope of protection of the present invention.

Claims

1. A method for comprehensive recovery of iron, vanadium, titanium, scandium and gallium from vanadium-titanium magnetite and red mud, characterized in that: The following steps are involved: a) mixing vanadium-titanium magnetite concentrate powder with red mud powder, roasting, leaching, and solid-liquid separation to obtain leaching residue and leachate; b) mixing the leachate with a precipitant to precipitate vanadium, and performing solid-liquid separation to obtain vanadium-rich slag and a vanadium-precipitated clear solution; dissolving the vanadium-rich slag in a sulfuric acid solution, reacting the mixture with an ammonium salt, and performing solid-liquid separation to obtain a solid product, ammonium polyvanadate, and roasting to obtain V2O5; adsorbing and desorbing gallium ions in the vanadium-precipitated clear solution using a resin, and adjusting the pH of the desorption solution to alkaline before electrolysis to obtain crude gallium; c) washing and drying the leached residue to prepare pellets; pre-reducing the pellets, mixing them with a carbonaceous reducing agent, and performing a smelting treatment to obtain scandium-containing high-titanium slag and molten iron; and recovering Ti and Sc in the scandium-containing high-titanium slag; There is no restriction on the order of step b) and step c).

2. The method according to claim 1, characterized in that In step a), the particle size of the vanadium-titanium magnetite concentrate powder and the red mud powder is ≤150 μm.

3. The method according to claim 1, characterized in that In step a), after the vanadium-titanium magnetite concentrate powder is mixed with the red mud powder, the molar ratio of Na2O to V2O5 in the mixture is ≥1.2:

1.

4. The method according to claim 1, wherein In step a), the calcination temperature is 600-900° C. and the calcination time is 1-5 hours.

5. The method according to claim 1, wherein In step a), the liquid-to-solid ratio of the water immersion is (3-5) mL:1 g, the temperature is 80-90° C., the stirring speed is 100-500 r / min, and the time is 30-180 min.

6. The method according to claim 1, characterized in that In step b), the precipitant is one or more of calcium chloride, lead sulfate, sodium phosphate, disodium hydrogen phosphate, sodium sulfide, ferrous sulfate, ferric chloride, magnesium sulfate, magnesium chloride, ammonium chloride, ammonium sulfate, hydrogen sulfide, oxalic acid, oxalate, 8-hydroxyquinoline and tannic acid; the molar ratio of the precipitant to V in the leachate is (1-1.5):1; the temperature of the vanadium precipitation is 50-90° C., the stirring speed is 100-500 r / min, and the time is 30-150 min.

7. The method according to claim 1, characterized in that In step b), the dissolution temperature is 30-90°C and the time is 50-150 min; the pH value of the mixed reaction with ammonium salt is 1.5-3, and the NH4 + The molar ratio of V is 1:(1.2~3).

8. The method according to claim 1, characterized in that In step c), the preparation process of the pellets includes: After the leaching residue is washed and dried, it is mixed with a binder to form pellets, and then roasted to obtain natural alkalinity pellets; Alternatively, after the leaching residue is washed and dried, it is mixed with a binder and a calcium flux to form pellets, and then roasted to obtain alkaline pellets.

9. The method according to claim 1, characterized in that In step c), the carbonaceous reducing agent is one or more of coke, coal and activated carbon; the amount of the carbonaceous reducing agent is 2-5% of the mass of the pre-reduced pellets; the temperature of the smelting treatment is 1500-1650° C., and the time is 1-3 hours.

10. The method according to claim 1, characterized in that In step c), the process of recovering the Ti and Sc comprises: The scandium-containing high-titanium slag is prepared into titanium dioxide powder by adopting a sulfuric acid method for preparing titanium dioxide; Resin is used to adsorb and desorb scandium ions in the waste acid of titanium dioxide generated during the preparation of titanium dioxide. After the pH of the desorption liquid is adjusted to alkaline, oxalic acid precipitation and roasting are performed to obtain scandium oxide.