Method for co-processing red mud by using titanium white waste acid

Abstract

The invention discloses a method for cooperatively treating red mud by using titanium white waste acid, and belongs to the technical field of industrial solid waste recycling and pollution cooperative treatment. The invention provides a method for synergistically treating red mud by using titanium dioxide waste acid in order to efficiently and synergistically recycle various metals in the red mud and the titanium dioxide waste acid, which comprises the following steps: pretreating the red mud; activating and leaching titanium white waste acid; multi-element separation and enrichment: co-precipitating iron and aluminum, extracting scandium and gallium by solvent extraction, and separating aluminum and gallium; and finally, titanium, iron, scandium, aluminum and gallium are recycled. According to the method disclosed by the invention, through a magnetic separation-acid leaching-coprecipitation-extraction-electrodeposition multi-stage combination technology, by utilizing the characteristics of high alkalinity of the red mud and high acidity of the titanium dioxide waste acid and optimizing the process, effective separation of various metals is realized, efficient synergistic recovery of Fe, Ti, Sc, Ga and Al in the red mud is realized, and Fe, Ti and Al in the titanium dioxide waste acid are synchronously recovered.

Description

TECHNICAL FIELD

[0001] The present application belongs to the technical field of industrial solid waste resource utilization and pollution co-treatment, and particularly relates to a method for treating red mud and synchronously recovering valuable metals by using waste acid generated in the production process of titanium dioxide by the sulfuric acid method, and is particularly suitable for synchronous resource utilization of red mud containing iron, scandium, gallium and other elements and titanium dioxide waste acid containing iron, titanium, aluminum and other elements. BACKGROUND

[0002] Red mud is an industrial waste discharged in the production process of alumina. The red mud discharge amount is comprehensively affected by the bauxite grade, production process and technical level. On average, 1-2 tons of red mud can be produced for every ton of alumina produced. At present, the global stock of red mud is about 5 billion tons, of which more than 60% is in China, and the stock is still increasing, with an annual discharge of 120 million tons. The high alkalinity and heavy metal content of red mud can cause environmental problems such as soil pollution and groundwater alkalization if it is stored for a long time, and also occupies a large amount of land resources. However, red mud is rich in valuable metals such as iron (Fe), aluminum (Al), titanium (Ti), scandium (Sc) and gallium (Ga), and has very high recycling value.

[0003] Low-concentration sulfuric acid with a concentration of about 20% is generated in the production process of titanium dioxide, and in addition to sulfuric acid, it also contains FeSO4, Al2(SO4)3 and TiOSO4 and other sulfate salts. According to statistics, 6-8 tons of waste acid are generated for every ton of titanium dioxide produced, and a large amount of gypsum hazardous waste is generated by the traditional lime neutralization method. The existing technologies have not realized the co-treatment of waste acid and red mud and the full-element recovery.

[0004] CN103131854A discloses a method for comprehensively recovering scandium and titanium by leaching red mud with titanium dioxide waste acid, which comprises the following steps: ① taking Bayer process red mud and waste acid by-produced in the production of titanium dioxide by the sulfuric acid method as raw materials, and leaching the red mud with titanium dioxide waste acid at normal pressure, with a liquid-solid ratio of 4.0-6.0 by volume weight ratio; ② using sulfuric acid for secondary acid leaching of the leaching residue; ③ after the secondary acid leaching residue is dealkalized, an iron-containing ore and a tailing are obtained by one-step strong magnetic separation, and the iron-containing ore is further subjected to two-step strong magnetic separation to obtain an iron concentrate and a middling; ④ the leaching liquid is subjected to extraction, back extraction, acid-dissolution hydrolysis, secondary extraction, impurity removal, secondary back extraction, hydrochloric acid dissolution, oxalic acid precipitation and calcination processes to obtain a scandium oxide product; and ⑤ the back extraction residue is dissolved with sulfuric acid, hydrolyzed for 2-3 hours under heating and boiling conditions, the metatitanic acid obtained is dried, placed in a muffle furnace at 700-850℃ for calcination for 2-3 hours, and then crushed to obtain titanium yellow powder.

[0005] CN110358937A discloses a method for selectively enriching scandium from titanium white waste acid and red mud, comprising the following steps: (1) pulp preparation: mix titanium white waste acid with red mud to form ore pulp; (2) low-temperature sulfuric acid curing: add 98% concentrated sulfuric acid to the ore pulp to cure the red mud; (3) high-temperature roasting: constant temperature roasting of the low-temperature cured material obtained in step (2) to obtain high-temperature roasted material and recover SO2 or SO3 gas; (4) mechanical grinding activation: mechanical activation of the high-temperature roasted material after cooling to room temperature; (5) clinker leaching: adding titanium white waste acid to the activated clinker for leaching, then performing solid-liquid separation to obtain leaching residue and leaching liquor; (6) water washing: adding water to the leaching residue, thoroughly mixing and stirring, and then performing solid-liquid separation to obtain water-washed leaching residue and leaching washing liquor; (7) low-temperature crystallization iron removal: adding a certain amount of ferrous sulfate seed crystal to the leaching liquor, adding an appropriate amount of ethanol, and crystallizing at low temperature, and then filtering to obtain purified leaching liquor; (8) selective adsorption: using porous silicon-based carrier adsorbent to selectively adsorb scandium in the iron-removed leaching liquor.

[0006] CN117004836A discloses a method for extracting scandium element from titanium white waste acid and red mud solution, comprising the following steps: (a) loading silicon-based tributyl phosphate resin into an adsorption resin column, and flowing the titanium white waste acid and red mud solution through the scandium extraction resin column; (b) washing the saturated scandium extraction resin in step (a) with ultrapure water; (c) first washing the resin after adsorbing scandium in step (b) with sodium hydroxide solution, and then washing it with high-purity water to wash off the sodium hydroxide on its surface; (d) acid solution washing: connecting the resin column loaded with silicon-based tributyl phosphate extraction resin and the resin column loaded with silicon-based tributyl phosphate enrichment resin with the acid solution storage tank; controlling the acid solution to flow into the resin column by a pressure pump, and slowly flowing the acid solution into the resin column loaded with silicon-based tributyl phosphate extraction resin first, and then into the resin column loaded with silicon-based tributyl phosphate enrichment resin to wash the extraction resin washed in step (c), and using a certain concentration of hydrochloric acid solution to desorb the extraction resin; (e) desorption: using resin desorbent to desorb scandium from the resin washed in step (d); (f) directly electrolyzing the desorption liquid to obtain metallic scandium.

[0007] It can be seen that the current red mud recycling technology mainly aims at single metal recovery, and lacks a systematic full-component utilization scheme, resulting in resource waste. The high alkalinity and complex mineral composition of red mud make the traditional metallurgical method less adaptable and the recovery cost high. In view of the above problems, it is urgent to develop a high-efficiency, low-consumption and full-component recycling method for red mud resource utilization. SUMMARY

[0008] The present application aims at the defects of the prior art, and develops a multi-stage combined technology of "magnetic separation-acid leaching-co-precipitation-extraction-electrodeposition", so as to realize efficient and synergistic recovery of various metal elements in red mud and titanium white waste acid.

[0009] The present application provides a method for synergistically treating red mud by using titanium white waste acid, which comprises the following steps: A, red mud pretreatment: after the red mud is dried and ball milled, magnetic separation is carried out at 3000-5000 Gauss to recover Fe3O4, then 1-20 wt.% NaOH aqueous solution and the red mud after magnetic separation are stirred at a liquid-solid ratio of 1-10 L:1 kg at 30-90 DEG C, after solid-liquid separation, drying is carried out again; B, titanium white waste acid activation leaching: the titanium white waste acid and the red mud after magnetic separation and drying in step A are leached at a liquid-solid ratio of 2-5 L:1 kg, fluoride is added as a cosolvent, the leaching temperature is controlled at 40-90 DEG C, the leaching process and the leaching end point pH are controlled at 1.0-2.0, solid-liquid separation is carried out, and the leaching liquid and the titanium-containing leaching residue are obtained; C, multi-element separation and enrichment: C1, iron-aluminum co-precipitation: an alkaline reagent is added to the leaching liquid obtained in step B to adjust the pH to 1.0-4.0, and reaction is carried out at 40-90 DEG C, solid-liquid separation is carried out, and aluminum-iron co-precipitation (yellow sodium iron alum [Na2Fe6(SO4)4(OH) 12 ] and Al(OH)3 co-precipitation) and scandium-gallium solution 1 are obtained; C2, solvent extraction for scandium and gallium: scandium and gallium solution 1 is subjected to extraction operation using an extraction solvent containing an extractant, O / A is controlled at 1:1-5, and scandium and gallium loaded organic phase and aluminum-gallium solution 2 are obtained; C3, aluminum-gallium separation: aluminum-gallium solution 2 is subjected to resin adsorption, gallium is adsorbed on the resin, and aluminum-containing aqueous phase is obtained; D, recovery of valuable elements: D1, titanium recovery: the titanium-containing leaching residue obtained in step B is subjected to flotation, and titanium concentrate is obtained; D2, iron recovery: the co-precipitation product obtained in step C1 is subjected to calcination and magnetic separation, and Fe2O3 is obtained (the Fe2O3 obtained by magnetic separation can be returned to the blast furnace ironmaking system as iron concentrate together with the Fe3O4 obtained by magnetic separation in step A); D3, scandium recovery: the scandium and gallium loaded organic phase obtained in step C2 is subjected to back extraction using 1-4 mol / L NaOH aqueous solution, O / A is controlled at 1:1-5, scandium hydroxide-containing slurry is obtained, solid-liquid separation is carried out, Sc(OH)3 precipitate and gallium-containing organic phase are obtained, and Sc2O3 is obtained by calcination; D4, aluminum recovery: the aluminum-containing aqueous phase obtained in step C3 is subjected to evaporation crystallization, and Al2(SO4)3·18H2O is obtained; D5, gallium recovery: the gallium-containing organic phase obtained in step D3 is back-extracted with sulfuric acid to obtain a gallium-containing solution 3; the resin adsorbed with gallium in step C3 is desorbed with sulfuric acid to obtain a gallium-containing solution 4; the gallium-containing solution 3 and the gallium-containing solution 4 are combined and electrodeposited to obtain metallic gallium.

[0010] In the above method, in step A, the red mud contains SiO2 9~23 wt.%, Al2O3 6.5~25 wt.%, Fe2O3 / Fe3O4 4~38 wt.%, CaO 15~46 wt.%, Na2O 2~9 wt.%, TiO2 2~8 wt.%, K2O 0.1~1.7 wt.%, Sc 0.001~0.01 wt.%, Ga 0.002~0.02 wt.%.

[0011] In the above method, in step B, the titanium dioxide waste acid is a waste acid generated in the production process of titanium dioxide by the sulfuric acid method, and contains H2SO4 15~25 wt.%, Fe 2+ / Fe 3+ 5~20 g / L, Ti 4+ 1~10 g / L, Al 3+ 0.5~5 g / L, Mg 2+ 0.1~3 g / L, Mn 2+ 0.05~2 g / L, SiO2 0.1~1 g / L.

[0012] In the above method, in step A, in the drying ball milling, drying is carried out at 105±5℃ until the water content is ≤5%, and ball milling is carried out until the particle size is ≤75μm.

[0013] In the above method, in step A, the mass of the recovered Fe3O4 accounts for 15~23 wt.% of the total amount of iron in the red mud.

[0014] In the above method, in step A, the stirring time is 0.5~4 hours.

[0015] In the above method, in step A, in the re-drying, drying is carried out at 105±5℃ until the water content is ≤5%.

[0016] In step A of the present application, the aqueous NaOH solution and the red mud after magnetic separation are mixed and stirred to remove the surface-adsorbed Na + , and reduce the subsequent acid consumption.

[0017] In the above method, in step B, the amount of fluoride is 0.1~1.0 wt.% of the mass of the red mud after magnetic separation and drying in step A.

[0018] In the above method, in step B, the fluoride is selected from at least one of NaF, NH4F, MgF2, CaF2, K2HF5, HF, sodium fluorosilicate, a fluoride mineral, fluorite (CaF2), or fluorapatite (Ca5[PO4]3F).

[0019] In the above method, in step B, the leaching time is 1-4 h.

[0020] In step B of the present application, valuable metals such as Al, Sc, Ga, etc. in red mud are dissolved by using titanium white waste acid (containing H2SO4 15-25%), fluoride is added as a solubilizing agent, and the pH during the reaction process and at the end of the reaction is controlled to ensure that Fe 3+ and Ti 4+ hydrolyze, realizing effective separation of titanium and other metals.

[0021] In the above method, in step C1, the amount of the alkaline reagent is 2-10 wt.% of the mass of the red mud after magnetic separation and drying.

[0022] In the above method, in step C1, the alkaline reagent is selected from at least one of CaCO3, Na2CO3, NaOH, CaO, KOH, K2CO3, NaHCO3, (NH4)2CO3, NH4HCO3, or Ca(OH)2.

[0023] In the above method, in step C1, the reaction time is 10-60 minutes.

[0024] In the above method, in step C2, the extraction solvent containing the extraction agent is composed of 10-2 vol.% extraction agent and 90-98 vol.% organic solvent; the extraction agent is selected from at least one of P204, TBP, P507, C272, TOPO, TRPO, N235, C923, or C927; and the organic solvent is at least one of dimethylbenzene, sulfonated kerosene, dichloroethane, aviation kerosene, cyclohexane, isopentane, isopentanol, or n-butane.

[0025] In the above method, in step C3, the resin is an amine oxime resin. The amine oxime resin is a resin containing an oxime group and an amino group, such as LSC600, LSC700, JK, etc.

[0026] In the above method, in step D1, the TiO2 content in the titanium-containing leaching residue is 8-12 wt.%.

[0027] In the above method, in step D1, the TiO2 content in the titanium concentrate is ≥45 wt.%.

[0028] In the above method, in step D1, during the flotation, the pH of the system is controlled to be 5-10.

[0029] In the above method, during step D1, the collector used in flotation is selected from either a non-polar collector or an ionic collector. The non-polar collector is a hydrocarbon oil collector, such as those for kerosene, diesel, or fuel oil. Ionic collectors include sodium dodecyl sulfonate, dodecylamine, hexadecyltrimethylammonium bromide, and carboxylic acids (oleic acid, paraffin, hydroxamic acid), among others.

[0030] In the above method, in step D1, during flotation, the amount of collector added is 100~500 g / t of titanium-containing leaching residue.

[0031] In the above method, during step D5, the sulfuric acid concentration is 0.5~3 mol / L and the O / A ratio is controlled to be 1:1~5.

[0032] In the above method, during step D5, the sulfuric acid concentration is 1~4 mol / L and the amount of sulfuric acid used is 2~10 BV.

[0033] The beneficial effects of this invention are: This invention utilizes a multi-stage combined technology of "magnetic separation-acid leaching-co-precipitation-extraction-electrodeposition" to effectively dissolve valuable metals such as Al, Sc, and Ga in red mud by taking advantage of the high alkalinity of red mud and the high acidity of titanium dioxide waste acid. Through process optimization, it achieves effective separation of multiple metals, overcomes the defects of existing processes, and realizes the efficient synergistic recovery of Fe, Ti, Sc, Ga, and Al in red mud. At the same time, it also recovers Fe, Ti, and Al from titanium dioxide waste acid. The recovery rates are no less than 75% for iron, no less than 85% for titanium, no less than 80% for scandium, no less than 80% for gallium, and no less than 70% for aluminum, which has significant economic and environmental benefits. Detailed Implementation

[0034] The present invention will be further described in detail below through embodiments, but the scope of protection of the present invention is not limited to the embodiments described herein.

[0035] Example 1 Red mud (21.30 wt.% SiO2, 18.32 wt.% Al2O3, 23.48 wt.% Fe2O3 / Fe3O4, 17.80 wt.% CaO, 5.20 wt.% Na2O, 2.89 wt.% TiO2, 0.21 wt.% K2O, 0.005 wt.% Sc, 0.01 wt.% Ga) was dried at 105℃ to a moisture content of 4.2%, ball-milled to 70 μm, and Fe3O4, accounting for 18 wt.% of the total iron in the red mud, was recovered by magnetic separation at 3000 Gauss. A 5 wt.% NaOH aqueous solution and the magnetically separated red mud were mixed at a liquid-to-solid ratio of 3 L: 1 kg and stirred at 60℃ for 1 hour to remove surface-adsorbed Na. +, reduce the subsequent acid consumption, after solid-liquid separation, continue to dry at 105℃ until the moisture content is less than or equal to 5%.

[0036] Titanium white waste acid (20wt.% H2SO4, Fe 2+ / Fe 3+ 12g / L, Ti 4+ 5g / L, Al 3+ 2g / L, Mg 2+ 1g / L, Mn 2+ 0.5g / L, SiO21g / L) and the liquid-solid ratio of the magnetic separation and dried red mud is 2L:1kg, 0.2wt.% NaF is added, and leaching is carried out at 45℃ for 1h, the solution leaching and end point pH control is at 1.5, and solid-liquid separation obtains leaching solution and titanium-containing leaching residue. The titanium-containing leaching residue (containing TiO210%) is subjected to flotation (diesel oil as collector, pH=6.5) to obtain titanium concentrate (TiO2≥45%).

[0037] 8wt.% CaCO3 of the mass of the magnetic separation and dried red mud is added to the leaching solution, the solution pH is adjusted to 1.0, the temperature is controlled at 40℃, and the reaction is carried out for 60min, after the reaction, solid-liquid separation obtains iron-aluminum co-precipitation product and gallium-scandium-containing solution 1. The co-precipitation product is calcined, and magnetic separation can obtain Fe2O3, which can be returned to the blast furnace ironmaking system with the previous magnetic separation product Fe3O4 as iron concentrate.

[0038] Gallium-scandium-containing solution 1 is subjected to extraction at room temperature using 2vol.% P204+98vol.% xylene with O / A being 1:1 to obtain scandium and gallium-loaded organic phase and solution 2 containing a small amount of gallium and aluminum, the scandium-loaded organic phase is first subjected to back extraction of scandium using 1.5mol / L NaOH solution to obtain a scandium hydroxide-containing slurry, solid-liquid separation is carried out to obtain Sc(OH)3precipitate and gallium-containing organic phase, and the solid is calcined to obtain scandium oxide.

[0039] The gallium-containing organic phase after back extraction of scandium is subjected to gallium back extraction using sulfuric acid to obtain gallium-containing solution 3. Solution 2 flows through an adsorption column filled with resin to separate gallium and aluminum to obtain aluminum-containing aqueous phase, and the adsorption column is subjected to sulfuric acid desorption to obtain gallium-containing solution 4. Solution 3 and 4 are mixed and subjected to electrodeposition experiment to obtain metallic gallium.

[0040] The aluminum-containing aqueous phase after passing through the adsorption column is subjected to evaporation crystallization to obtain Al2(SO4)3·18H2O.

[0041] The recovery rate of iron in the whole process is 75%, the recovery rate of titanium is 85%, the recovery rate of scandium is 80%, the recovery rate of gallium is 80%, and the recovery rate of aluminum is 70%.

[0042] Example 2 Red mud (17.85wt.%SiO2, 24.62wt.%Al2O3, 13.76wt.%Fe2O3 / Fe3O4, 20.12wt.%CaO, 5.95wt.%Na2O, 3.73wt.%TiO2, 1.62wt.%K2O, 0.008wt.%Sc, 0.015wt.%Ga) was dried at 105℃ to moisture content of 5.0%, ball milled to 70μm, and 3500 gauss magnetic separation was used to recover 19.2wt.% of Fe3O4 in the total amount of iron in the red mud. 5wt.% NaOH aqueous solution and the magnetic separation red mud were stirred at 60℃ for 1 hour at a liquid-solid ratio of 3L:1kg, and the NaOH adsorbed on the surface of the red mud was removed. The red mud was dried at 105℃ to a moisture content of ≤5% after solid-liquid separation. + , reducing the subsequent acid consumption, and the red mud was dried at 105℃ to a moisture content of ≤5% after solid-liquid separation.

[0043] Titanium white waste acid (20wt.%H2SO4, Fe 2+ / Fe 3+ 12g / L, Ti 4+ 5g / L, Al 3+ 2g / L, Mg 2+ 1g / L, Mn 2+ 0.5g / L, SiO21g / L) was mixed with the magnetic separation and dried red mud at a liquid-solid ratio of 3L:1kg, 0.8wt.% NH4F was added, and the mixture was leached at 55℃ for 1.5h. The solution leaching and end point pH were controlled at 1.2, and the leaching solution and titanium-containing leaching residue were obtained after solid-liquid separation. The titanium-containing leaching residue (containing TiO211%) was subjected to flotation (diesel oil as collector, pH=6.5) to obtain a titanium concentrate (TiO2≥45%).

[0044] 6wt.% Na2CO3 of the magnetic separation and dried red mud was added to the leaching solution, the solution pH was adjusted to 1.5, the temperature was controlled at 50℃, and the reaction was carried out for 50min. After the reaction, the iron and aluminum co-precipitation product and gallium and scandium-containing solution 1 were obtained after solid-liquid separation. The co-precipitation product was calcined, and Fe2O3 was obtained by magnetic separation, which can be returned to the blast furnace ironmaking system together with the previous magnetic separation product Fe3O4 as an iron concentrate.

[0045] Gallium and scandium-containing solution 1 was extracted at room temperature using 2vol.% P204+2vol.% TBP+96vol.% sulfonated kerosene at an O / A ratio of 1:2 to obtain an organic phase loaded with scandium and gallium and solution 2 containing a small amount of gallium and aluminum. The scandium and gallium-loaded organic phase was first stripped of scandium using 2.0mol / L NaOH solution to obtain a slurry containing scandium hydroxide, which was subjected to solid-liquid separation to obtain Sc(OH)3precipitate and gallium-containing organic phase. The solid was calcined to obtain scandium oxide.

[0046] The gallium-containing organic phase after stripping scandium is subjected to sulfuric acid stripping of gallium to obtain gallium-containing solution 3. Solution 2 is allowed to flow through an adsorption column filled with resin to separate gallium and aluminum, to obtain an aluminum-containing aqueous phase. The adsorption column is subjected to desorption with sulfuric acid to obtain gallium-containing solution 4. Solutions 3 and 4 are mixed and subjected to electrodeposition to obtain metallic gallium.

[0047] The aluminum-containing aqueous phase after passing through the adsorption column is subjected to evaporation crystallization to obtain Al2(SO4)3·18H2O.

[0048] The recovery rates of iron, titanium, scandium, gallium and aluminum in the whole process are 78%, 88%, 90%, 85% and 75%, respectively.

[0049] Example 3 Red mud (9.58wt.%SiO2, 19.20wt.%Al2O3, 25.93wt.%Fe2O3 / Fe3O4, 28.65wt.%CaO, 3.15wt.%Na2O, 4.40wt.%TiO2, 0.50wt.%K2O, 0.008wt.%Sc, 0.015wt.%Ga) is dried at 105°C to a water content of 4.5%, ball milled to 70μm, and 20.3wt.% of Fe3O4 in the total amount of iron in the red mud is recovered by 4000 gauss magnetic separation. 5wt.% NaOH aqueous solution and the red mud after magnetic separation are stirred at 60°C for 1 hour at a liquid-solid ratio of 3L:1kg, and the NaOH adsorbed on the surface of the red mud is removed. + , the acid consumption is reduced, and after solid-liquid separation, the red mud is dried at 105°C to a water content of ≤5%.

[0050] Titanium white waste acid (25wt.%H2SO4, Fe 2+ / Fe 3+ 20g / L, Ti 4+ 3g / L, Al 3+ 1g / L, Mg 2+ 0.5g / L, Mn 2+ 0.1g / L, SiO20.8g / L) is mixed with the red mud after magnetic drying at a liquid-solid ratio of 3L:1kg, 0.6wt.% CaF2 is added, and leaching is carried out at 65°C for 2.5h, with the solution leaching and end point pH controlled at 1.8, to obtain a leaching solution and a titanium-containing leaching residue. The titanium-containing leaching residue (containing TiO212%) is subjected to flotation (diesel oil as collector, pH=6.5) to obtain a titanium concentrate (TiO2≥45%).

[0051] The leaching solution is added with 2wt.% of NaOH of the mass of the dried red mud after magnetic separation, the pH of the solution is adjusted to 2.0, the reaction is controlled at 65°C for 40 minutes, and after the reaction, the iron and aluminum co-precipitation product and the gallium and scandium containing solution 1 are obtained by solid-liquid separation. The co-precipitation product is calcined, and the Fe2O3 is obtained by magnetic separation, which can be returned to the blast furnace ironmaking system together with the previous magnetic separation product Fe3O4 as iron concentrate.

[0052] The gallium and scandium containing solution 1 is extracted at room temperature by using 4vol.% P507+4vol.% TBP+92vol.% dichloroethane with O / A of 1:3, and the organic phase loaded with scandium and gallium and the solution 2 containing a small amount of gallium and aluminum are obtained. The organic phase loaded with scandium and gallium is first subjected to back extraction of scandium by using 3.0mol / L NaOH solution, and the scandium hydroxide slurry is obtained. After solid-liquid separation, the Sc(OH)3 precipitate and the gallium containing organic phase are obtained. The solid is calcined to obtain scandium oxide.

[0053] The gallium containing organic phase after back extraction of scandium is subjected to gallium back extraction by sulfuric acid to obtain the gallium containing solution 3. The solution 2 flows through the adsorption column filled with resin to separate gallium and aluminum, and the aluminum containing aqueous phase is obtained. After desorption of the adsorption column by sulfuric acid, the gallium containing solution 4 is obtained. The solution 3 and 4 are mixed and subjected to electrodeposition experiment to obtain metallic gallium.

[0054] The aluminum containing aqueous phase after passing through the adsorption column is subjected to evaporation crystallization to obtain Al2(SO4)3·18H2O.

[0055] The recovery rate of iron in the whole process is 82%, the recovery rate of titanium is 90%, the recovery rate of scandium is 95%, the recovery rate of gallium is 88%, and the recovery rate of aluminum is 80%.

[0056] Example 4 The red mud (21.30wt.% SiO2, 18.32wt.% Al2O3, 23.48wt.% Fe2O3 / Fe3O4, 17.80wt.% CaO, 5.20wt.% Na2O, 2.89wt.% TiO2, 0.21wt.% K2O, 0.005wt.% Sc, 0.01wt.% Ga) is dried at 105°C to a water content of 3.8%, and ball milled to 70μm. 21.6wt.% of Fe3O4 in the total amount of iron in the red mud is recovered by 4500 gauss magnetic separation. 5wt.% NaOH aqueous solution and the red mud after magnetic separation are mixed at a liquid-solid ratio of 3L:1kg, stirred at 60°C for 1 hour, and the NaOH solution is removed. The water content of the red mud is reduced to ≤5% after drying at 105°C. + , and the acid consumption in the subsequent process is reduced. After solid-liquid separation, the red mud is dried at 105°C to a water content of ≤5%.

[0057] Titanium white waste acid (18wt.% H2SO4, Fe 2+ / Fe 3+ 10g / L, Ti 4+10g / L, Al 3+ 5g / L, Mg 2+ 3g / L, Mn 2+ The liquid-to-solid ratio of red mud (containing 2 g / L of TiO2 and 0.1 g / L of SiO2) to magnetically separated and dried red mud was 4 L: 1 kg. 0.5 wt.% MgF2 was added, and leaching was carried out at 75℃ for 3 h. The pH of the leaching solution and the endpoint were controlled at 1.6. Solid-liquid separation yielded leachate and titanium-containing leaching residue. The titanium-containing leaching residue (containing 9.5% TiO2) was subjected to flotation (using diesel oil as the collector, pH=6.5) to obtain titanium concentrate (TiO2≥45%).

[0058] Add 5 wt.% CaO (by weight of magnetically separated and dried red mud) to the leachate, adjust the pH of the solution to 2.5, and react at 75℃ for 30 min. After the reaction, solid-liquid separation yields an iron-aluminum coprecipitate and a gallium-scandium-containing solution 1. The coprecipitate is roasted and then magnetically separated to obtain Fe2O3, which, along with the previously magnetically separated Fe3O4, can be returned to the blast furnace ironmaking system as iron concentrate.

[0059] Gallium-scandium solution 1 was extracted with 5 vol.% P2O4 + 5 vol.% P5O7 + 90 vol.% sulfonated kerosene at room temperature with an O / A ratio of 1:4 to obtain an organic phase loaded with scandium and gallium and solution 2 containing a small amount of gallium and aluminum. The scandium-gallium organic phase was first back-extracted with 2.0 mol / L NaOH solution to obtain a scandium hydroxide slurry. Solid-liquid separation was performed to obtain Sc(OH)3 precipitate and gallium-containing organic phase. The solid was calcined to obtain scandium oxide.

[0060] After scandium back-extraction, the gallium-containing organic phase is further back-extracted with sulfuric acid to obtain gallium-containing solution 3. Solution 2 flows through a resin-filled adsorption column to separate gallium and aluminum, obtaining an aluminum-containing aqueous phase. After desorption with sulfuric acid, gallium-containing solution 4 is obtained. Mixing solutions 3 and 4 and performing electrodeposition experiments yields metallic gallium.

[0061] Al2(SO4)3·18H2O can be obtained by evaporation and crystallization of the aluminum-containing aqueous phase after adsorption column.

[0062] Throughout the entire process, the recovery rates are as follows: iron 85%, titanium 92%, scandium 96%, gallium 90%, and aluminum 85%.

[0063] Example 5 Red mud (20.50wt.%SiO2, 7.15wt.%Al2O3, 8.10wt.%Fe2O3 / Fe3O4, 45.86wt.%CaO, 2.36wt.%Na2O, 7.30wt.%TiO2, 0.50wt.%K2O, 0.009wt.%Sc, 0.018wt.%Ga) was dried at 105℃ to moisture content of 3.5%, ball milled to 70μm, 5000 Gauss magnetic separation was used to recover 22wt.% of Fe3O4 in total iron in red mud, 5wt.% NaOH aqueous solution and magnetic separation red mud were added with liquid-solid ratio of 3L:1kg, stirred at 60℃ for 1h, NaOH solution was removed, and the surface adsorbed Na + , reduced the subsequent acid consumption, after solid-liquid separation, it was dried at 105℃ to moisture content of ≤5%.

[0064] Titanium white waste acid (23wt.%H2SO4, Fe 2+ / Fe 3+ 18g / L, Ti 4+ 8g / L, Al 3+ 4g / L, Mg 2+ 0.1g / L, Mn 2+ 0.05g / L, SiO20.9g / L) and the liquid-solid ratio of the magnetic dried red mud was 5L:1kg, 1.0wt.% NaF was added, leaching at 85℃ for 4h, the solution leaching and end point pH was controlled at 2.0, solid-liquid separation obtained leaching solution and titanium-containing leaching residue. The titanium-containing leaching residue (containing TiO212%) was subjected to flotation (diesel oil as collector, pH=6.5) to obtain titanium concentrate (TiO2≥45%).

[0065] KOH was added to the leaching solution at a mass of 3wt.% of the magnetic dried red mud, the solution pH was adjusted to 3.5, the temperature was controlled at 90℃ for 10min, after the reaction, solid-liquid separation obtained iron-aluminum co-precipitation product and gallium-scandium-containing solution 1. The co-precipitation product was calcined, and the Fe2O3 obtained by magnetic separation could be returned to the blast furnace ironmaking system as iron concentrate together with the Fe3O4 obtained by previous magnetic separation.

[0066] Gallium-scandium-containing solution 1 was subjected to extraction at room temperature using 3vol.% P204+2vol.% P507+5vol.% C272+90vol.% aviation kerosene with O / A ratio of 1:4 to obtain scandium and gallium-loaded organic phase and solution 2 containing a small amount of gallium and aluminum, the scandium and gallium-loaded organic phase was first subjected to back extraction of scandium using 2.0mol / L NaOH solution to obtain a slurry containing scandium hydroxide, solid-liquid separation, Sc(OH)3precipitate and gallium-containing organic phase were obtained, and the solid was calcined to obtain scandium oxide.

[0067] The organic phase containing gallium after stripping scandium is subjected to sulfuric acid stripping to obtain a gallium-containing solution 3. Solution 2 is allowed to flow through an adsorption column filled with resin to separate gallium and aluminum, thereby obtaining an aluminum-containing aqueous phase. The adsorption column is subjected to sulfuric acid desorption to obtain a gallium-containing solution 4. The solutions 3 and 4 are mixed and subjected to electrodeposition to obtain metallic gallium.

[0068] The aluminum-containing aqueous phase after passing through the adsorption column is subjected to evaporation crystallization to obtain Al2(SO4)3·18H2O.

[0069] In the whole process, the recovery rates of iron, titanium, scandium, gallium and aluminum are 88%, 95%, 93%, 92% and 90%, respectively.

Claims

1. A method for co-treating red mud using titanium dioxide waste acid, characterized in that: Includes the following steps: A. Red mud pretreatment: After drying and ball milling, the red mud is magnetically separated at 3000~5000 Gauss to recover Fe3O4. Then, 1~20wt.% NaOH aqueous solution and magnetically separated red mud are stirred at 30~90℃ at a liquid-solid ratio of 1~10L:1kg. After solid-liquid separation, it is dried again. B. Titanium dioxide waste acid activation leaching: Titanium dioxide waste acid and red mud dried by magnetic separation in step A are leached at a liquid-to-solid ratio of 2~5L:1kg, with fluoride added as a co-solvent. The leaching temperature is controlled at 40~90℃, and the pH during the leaching process and at the leaching endpoint is 1.0~2.

0. Solid-liquid separation is performed to obtain leachate and titanium-containing leaching residue. C. Multi-element separation and enrichment: C1. Iron-aluminum coprecipitation: Add an alkaline reagent to the leachate obtained in step B to adjust the pH to 1.0~4.0, react at 40~90℃, separate the solid and liquid, and obtain aluminum-iron coprecipitate and scandium-gallium solution 1; C2. Solvent extraction to extract scandium gallium: Scandium gallium solution 1 is extracted using an extraction solvent containing an extractant, with O / A ratio controlled at 1:1~5, to obtain scandium gallium-loaded organic phase and aluminum gallium solution 2; C3. Aluminum-gallium separation: Aluminum-gallium solution 2 is subjected to resin adsorption, and gallium is adsorbed onto the resin to obtain an aluminum-containing aqueous phase; D. Valuable elements are recycled separately: D1. Titanium recovery: The titanium-containing leaching residue obtained in step B is subjected to flotation to obtain titanium concentrate; D2. Iron recovery: The coprecipitate obtained in step C1 is roasted and magnetically separated to obtain Fe2O3. D3. Scandium recovery: The scandium-gallium-loaded organic phase obtained in step C2 is back-extracted with 1~4 mol / L NaOH aqueous solution, controlling O / A=1:1~5, to obtain scandium hydroxide slurry. Solid-liquid separation is performed to obtain Sc(OH)3 precipitate and gallium-containing organic phase, which are then calcined to obtain Sc2O3. D4. Aluminum recovery: The aluminum-containing aqueous phase obtained in step C3 is evaporated and crystallized to obtain Al2(SO4)3·18H2O; D5. Gallium recovery: The gallium-containing organic phase obtained in step D3 is back-extracted with sulfuric acid to obtain gallium-containing solution 3; the resin adsorbing gallium in step C3 is desorbed with sulfuric acid to obtain gallium-containing solution 4; gallium-containing solution 3 and gallium-containing solution 4 are combined and electrodeposited to obtain metallic gallium.

2. The method according to claim 1, characterized in that: In step A, the red mud contains 9-23 wt.% SiO2, 6.5-25 wt.% Al2O3, 4-38 wt.% Fe2O3 / Fe3O4, 15-46 wt.% CaO, 2-9 wt.% Na2O, 2-8 wt.% TiO2, 0.1-1.7 wt.% K2O, 0.001-0.01 wt.% Sc, and 0.002-0.02 wt.% Ga.

3. The method according to claim 1, characterized in that: In step B, the titanium dioxide waste acid contains 15-25 wt.% H2SO4 and Fe. 2+ / Fe 3+ 5~20 g / L, Ti 4+ 1~10 g / L, Al 3+ 0.5~5 g / L, Mg 2+ 0.1~3 g / L, Mn 2+ 0.05~2 g / L, SiO2 0.1~1 g / L.

4. The method according to claim 1, characterized in that: In step A, at least one of the following must be satisfied: In the drying ball milling process, the material is dried at 105±5℃ until the moisture content is ≤5%, and then ball milled until the particle size is ≤75μm. The recovered Fe3O4 accounts for 15-23 wt.% of the total iron content in the red mud. The stirring time is 0.5 to 4 hours; During the subsequent drying process, the product is dried at 105±5℃ until the moisture content is ≤5%.

5. The method according to claim 1, characterized in that: In step B, at least one of the following must be satisfied: The amount of fluoride used is 0.1~1.0 wt.% of the mass of the red mud after magnetic separation and drying in step A. The fluoride is selected from at least one of NaF, NH4F, MgF2, CaF2, K2HF5, HF, sodium fluorosilicate, fluoride minerals, fluorite or fluorapatite; The leaching time is 1 to 4 hours.

6. The method according to claim 1, characterized in that: In step C1, at least one of the following must be satisfied: The amount of alkaline reagent used is 2-10 wt.% of the mass of the red mud after magnetic separation and drying in step A. The alkaline reagent is selected from at least one of CaCO3, Na2CO3, NaOH, CaO, KOH, K2CO3, NaHCO3, (NH4)2CO3, NH4HCO3 or Ca(OH)2; The reaction time is 10 to 60 minutes.

7. The method according to claim 1, characterized in that: In step C2, the extraction solvent containing the extractant consists of 10-2 vol.% extractant and 90-98 vol.% organic solvent; the extractant is selected from at least one of P204, TBP, P507, C272, TOPO, TRPO, N235, C923 or C927; the organic solvent is at least one of xylene, sulfonated kerosene, dichloroethane, aviation kerosene, cyclohexane, isopentane, isopentyl alcohol or n-butane.

8. The method according to claim 1, characterized in that: In step C3, the resin is a methylamine oxime resin.

9. The method according to claim 1, characterized in that: In step D1, at least one of the following conditions must be met: The titanium-containing leaching residue contains 8-12 wt% TiO2. Titanium concentrate contains ≥45 wt% TiO2; During flotation, the pH of the system should be controlled at 5-10; During flotation, the collector is selected from non-polar collectors or ionic collectors; During flotation, the amount of collector added is 100~500 g / t of titanium-containing leaching residue.

10. The method according to claim 1, characterized in that: In step D5, at least one of the following conditions must be met: During back-extraction, the sulfuric acid concentration is 0.5~3 mol / L, and the O / A ratio is controlled at 1:1~5; During desorption, the sulfuric acid concentration is 1~4 mol / L, and the amount of sulfuric acid used is 2~10 BV.