Process for recovering valuable metal from red mud

Abstract

The present disclosure provides a series of processes implemented by combining: a dry process involving reduction roasting and magnetic separation for recovering iron in red mud generated as a by-product in a bauxite processing process; and a wet process for additionally recovering each of aluminum and titanium, after separation of iron.

Description

Process for recovering valuable metals from red mud

[0001] The present disclosure relates to a process for recovering valuable metals from red mud. More specifically, the present disclosure relates to a series of processes implemented by combining a dry process involving reduction roasting and magnetic separation to recover iron from red mud, a byproduct of the bauxite processing process, and a wet process to further recover aluminum and titanium, respectively, after separation of the iron.

[0002] Red mud is a red process byproduct generated in the process of manufacturing alumina using bauxite as a raw material (i.e., the Bayer process). It is also called bauxite residue or red mud. In the Bayer process, caustic soda (NaOH) is added and processed under high temperature and high pressure to recover alumina from bauxite. The reaction mechanism by which red mud is generated during this process can be expressed as shown in Scheme 1 below.

[0003]

[0004] [Reaction Formula 1]

[0005] 2NaOH + Al2O3·2H2O → Na2O·Al2O3+ 4H2O + red mud

[0006]

[0007] As demand for alumina has increased, so has the production of red mud. Typically, approximately 1 to 1.5 tons of red mud is produced for every ton of alumina produced. As demand for alumina increases, so too has the production of red mud. Red mud has a pH ranging from approximately 9.2 to 12.8, and its highly alkaline nature means it is treated as designated waste.

[0008] Traditionally, red mud was disposed of through landfills and ocean dumping, rather than through efficient utilization. However, large-scale red mud accumulations have led to soil, surface water, and groundwater contamination due to corrosive leachates, which has also increased treatment costs. Furthermore, conventional treatment methods are limited in their capacity compared to the increasing amount of red mud generated.

[0009] Red mud contains significant amounts of valuable metals, including iron (Fe), aluminum (Al), titanium (Ti), silicon (Si), gallium (Ga), and rare earth elements like scandium (Sc) and yttrium (Y). Therefore, active research is underway to recover and recycle the various valuable metals contained within red mud.

[0010] For example, in order to recover iron in red mud, a method has been proposed in which a carbonaceous substance (e.g., activated carbon, coal, lignite, etc.) as a reducing agent is added to the red mud, heat treatment is performed to convert the iron component contained in the red mud in the form of hematite (Fe2O3) into magnetite (Fe3O4), and then the iron is separated by magnetic separation and then recovered (dry smelting process) (e.g., Steel Trans. 42 (5), 379-386 (2012)). In addition, a technology for separating / recovering valuable metals by leaching them through direct acid treatment of red mud has also been developed (e.g., U.S. Patent No. 9,023,301).

[0011] However, in the case of the dry smelting process, the loss of titanium and aluminum in the slag is caused, while in the wet smelting process, highly corrosive and toxic highly concentrated inorganic acids are mainly used to directly dissolve the metals in the red mud, which may cause secondary environmental pollution.

[0012] In this way, the processing technology according to the conventional technology described above not only has advantages and disadvantages, but it is also practically difficult to implement a single process that can recover the main valuable metals in red mud, such as iron, aluminum, and titanium, with high efficiency.

[0013] In particular, red mud has a variety of compositions depending on the source (e.g., bauxite processing plant), and titanium contained in red mud is particularly difficult to process, making recovery difficult.

[0014] Therefore, a method is required to overcome the limitations of conventional technology and to easily separate and recover valuable metals, especially iron, aluminum, and titanium, contained in red mud with high efficiency and purity.

[0015] One specific example of the present disclosure is to provide a process capable of recovering iron, aluminum and titanium, which are the main components of red mud, with high purity and high efficiency within a single process.

[0016] According to one specific example of the present disclosure,

[0017] a) providing red mud containing iron, aluminum and titanium;

[0018] b) a step of separating the magnetic portion and the non-magnetic portion through magnetic separation after reducing and roasting the red mud; and

[0019] c) A step of recovering iron from the residue of the magnetic portion remaining after performing liquid leaching on each of the magnetic portion and non-magnetic portion, and titanium from the residue of the non-magnetic portion, while recovering aluminum from the leaching liquid of the liquid leaching;

[0020] A process for recovering valuable metals from red mud containing .

[0021] According to an exemplary embodiment, in step c), aluminum can be recovered by adding acid to the leachate and precipitating it.

[0022] According to an exemplary embodiment, the aluminum is precipitated under conditions controlled in the range of pH 3 to 6, wherein the acid can be recovered as a precipitate in the form of AlPO4 using phosphoric acid (H3PO4).

[0023] According to an exemplary specific example, sodium phosphate (Na3PO4) is additionally added together with phosphoric acid, and the amount of sodium phosphate added can be adjusted in a range of 0.5 to 2 in the molar ratio of aluminum to sodium phosphate.

[0024] According to an exemplary embodiment, the precipitation of aluminum can be performed at 50 to 90°C for 0.1 to 5 hours.

[0025] According to an exemplary embodiment, in step c), titanium can be recovered in the form of titania (TiO2).

[0026] According to an exemplary embodiment, the red mud may include, on an elemental basis, 5 to 60 wt% iron (Fe), 3 to 40 wt% aluminum, 2 to 20 wt% titanium, and the balance.

[0027] According to an exemplary embodiment, the content of balance in the red mud may be in the range of 25 to 60 wt% on an elemental basis.

[0028] According to an exemplary embodiment, the loss on ignition (LOI) of the red mud may be in the range of 9 to 50 wt%.

[0029] According to an exemplary embodiment, the reducing roasting of step b) can be performed in the presence of a carbon source as a reducing agent and an alkali metal salt selected from the group consisting of sodium salts and potassium salts as an additive.

[0030] According to an exemplary embodiment, the carbon source is at least one selected from the group consisting of coke, coal, charcoal, biochar, agricultural residues / paper waste and pulp processing waste, and

[0031] The above alkali metal salt may be at least one selected from the group consisting of sodium carbonate, sodium sulfate, borax (Na2B4O7), potassium carbonate, potassium sulfate, and potassium bicarbonate.

[0032] According to an exemplary embodiment, the weight ratio of red mud: alkali metal salt: carbon source can be adjusted in the range of 8 to 20:2 to 8:1.

[0033] According to an exemplary embodiment, the reduction roasting of step b) may be performed at a temperature controlled in the range of 600 to 1200°C.

[0034] According to an exemplary embodiment, the reduction roasting of step b) may be performed for 0.3 to 10 hours.

[0035] According to an exemplary embodiment, the weight ratio of the magnetic portion: non-magnetic portion separated in the above-described Dean system b) may be in the range of 1 to 4:1.

[0036] According to an exemplary embodiment, the liquid-to-solid (L / S) ratio during the liquid leaching of each of the magnetic portion and the non-magnetic portion in step c) can be controlled in the range of 1 to 50.

[0037] According to an exemplary embodiment, the magnetic portion of step c) may contain at least 20 wt.% (by element) of iron.

[0038] According to an exemplary embodiment, the non-magnetic portion in step c) may contain at least 4 wt.% (by element) of titanium.

[0039] According to an exemplary embodiment, the recovery of titanium in step c) may involve a step of mixed-sulfate roasting of the non-magnetic portion using sulfuric acid and sulfate.

[0040] According to an exemplary embodiment, the weight ratio of sulfuric acid / sulfate during mixed-sulfate roasting can be adjusted in the range of 0.5 to 4.

[0041] According to an exemplary embodiment, the sulfate may be at least one selected from the group consisting of sodium sulfate (Na2SO4), ammonium sulfate ((NH4)2SO4), magnesium sulfate (MgSO4), potassium sulfate (K2SO4), and barium sulfate (BaSO4).

[0042] According to an exemplary embodiment, the non-magnetic portion subjected to the above mixed-sulfate roasting is leached with sulfuric acid at a temperature of 70 to 150°C to form a titanium-containing leaching solution, wherein the concentration of sulfuric acid is up to 10 M, and the liquid-to-solid ratio can be controlled in the range of 2 to 100.

[0043] According to an exemplary embodiment, the titanium-containing leachate may be heated to a temperature of 70 to 100°C and titanium hydrolysis may be performed while adjusting the pH to less than 3 to form a titanium-containing precipitate.

[0044] According to an exemplary embodiment, the titanium-containing precipitate can be converted into titania by calcining it under conditions of an oxygen-containing atmosphere and a temperature of 200 to 900° C. for 0.5 to 10 hours.

[0045] The process according to a specific example of the present disclosure can recover each of the valuable metals iron, aluminum, and titanium from red mud, a byproduct of the bauxite processing process, with high efficiency through a single process. In particular, it is configured to be operable with relatively low energy consumption and low reagent usage, thereby being advantageous in terms of environmental and waste reduction. Furthermore, the process according to the specific example offers the advantage of reducing carbon emissions compared to conventional smelting processes while effectively suppressing the generation of secondary waste such as slag. Therefore, widespread commercialization is expected in the future.

[0046] Figure 1 is a flowchart of an exemplary process according to one specific example;

[0047] Figure 2 is an X-ray diffraction (XRD) pattern of a red mud sample used in the example;

[0048] Figure 3 is an X-ray diffraction (XRD) pattern after reduction roasting performed without adding a reducing additive for a red mud sample (mixing ratio (by weight) of coke: red mud is 1:10); and

[0049] Figure 4 is an X-ray diffraction (XRD) pattern after reduction roasting performed under the addition of a reducing additive (sodium carbonate) for a red mud sample (the mixing ratio (by weight) of coke: sodium carbonate: red mud is 1:5:10).

[0050] The present invention can be fully achieved by the following description. It should be understood that the following description describes preferred embodiments of the present invention and is not necessarily limited thereto. Furthermore, the attached drawings are provided to aid understanding and are not intended to limit the present invention. Details regarding individual components can be appropriately understood based on the specific intent of the related descriptions described below.

[0051] The terms used in this specification may be defined as follows.

[0052] "Red mud" can be understood to mean a red process product generated in the process of manufacturing alumina using bauxite as a raw material.

[0053] "Reductive roasting" may refer to a process of reducing a metal oxide to the metal or a compound of lower oxidation number under a reducing atmosphere.

[0054] "Magnetic separation" can broadly mean a process of separating a mixture using a magnet, and specifically a process of applying an electromagnetic field to a mixture containing at least one magnetic component and at least one non-magnetic component to separate the magnetic component from the other components.

[0055] "Leaching" may mean the process of removing a substance by dissolution in a percolating liquid.

[0056] When a numerical range is specified herein with lower and / or upper limits, it is understood that any sub-combination within that numerical range is also disclosed. For example, when "1 to 5" is described, it can include 1, 2, 3, 4, and 5, as well as any sub-combination therebetween.

[0057] Where any component or member in this specification is described as being "connected" to another component or member, unless otherwise stated, it is to be understood that this includes not only the case where it is directly connected to said other component or member, but also the case where it is connected through the intervention of said other component or member.

[0058] Similarly, the term "contact" may be understood to include not only direct contact but also contact under the intervention of other components or absences.

[0059] When we say that something "includes" something, we mean that it may also include other components, unless otherwise stated.

[0060] According to one specific example of the present disclosure, a process is provided for recovering valuable metals, specifically iron (Fe), aluminum (Al) and titanium (Ti), from red mud, which is discharged as a by-product during the processing of bauxite, by a combination of dry and wet treatments in a physical and chemical manner.

[0061] An exemplary overall process sequence for separating and recovering iron, aluminum and titanium from red mud, according to one specific example, is as illustrated in FIG. 1.

[0062] Referring to the above drawing, red mud is provided as a starting material. Red mud is also called bauxite residue and typically contains trace amounts of balance elements, including oxides of silicon (Si), sodium (Na), and calcium (Ca), as well as iron (Fe), aluminum (Al), and titanium (Ti).

[0063] Red mud is typically obtained in the form of fine grains, and prior to processing, a washing process (specifically, water washing) may be performed at least once in order to remove impurities such as various dusts attached to the particle surface in addition to the components inherently contained within the red mud and to lower the high alkalinity. As an example, the amount of water used in the water washing process may be adjusted in the range of, for example, about 5 to 30% (w / v), specifically about 10 to 25% (w / v), and the water washing temperature may be adjusted in the range of, for example, about 50 to 100°C, specifically about 70 to 95°C, and more specifically about 80 to 90°C.

[0064] As described above, after performing the water washing, a drying process may be performed. Optionally, a filtration process may be performed prior to drying to separate solids, and the resulting filtrate may be reused in the water washing process. In addition, the drying process may be performed by any conventional method known in the art, such as heating (hot air) drying or reduced pressure drying.

[0065] Referring back to Figure 1, the particle size of the red mud can be determined by considering factors that affect the processing processes described below (e.g., filtration, transport, etc.), and may be, for example, about 40 ㎛ or less, specifically about 0.8 to 20 ㎛, and more specifically about 3 to 10 ㎛. However, this may be understood as an example.

[0066] According to an exemplary embodiment, the iron (Fe) content in red mud may range, for example, from about 5 to 60 wt%, specifically from about 15 to 50 wt%, and more specifically from about 20 to 40 wt%, on an elemental basis. In this case, a significant portion of the iron (Fe) element may exist in the form of iron oxide (Fe2O3).

[0067] Additionally, the content of aluminum (Al) in the red mud may be, for example, in the range of about 3 to 40 wt%, specifically about 5 to 20 wt%, and more specifically about 7 to 15 wt%, on an elemental basis.

[0068] In addition, the content of titanium (Ti) in the red mud may be, for example, in the range of about 2 to 20 wt%, specifically about 3 to 10 wt%, and more specifically about 5 to 7 wt%, on an elemental basis.

[0069] Meanwhile, red mud may contain balance in addition to the aforementioned components. These balance components may be, for example, silicon (Si), calcium (Ca), sodium (Na), silicon (Si), carbon (C), oxygen (O), etc., and the components constituting the balance components and their contents may vary depending on the source of the red mud, etc.

[0070] According to an exemplary embodiment, the balance content in the red mud may range, for example, from about 25 to 60 wt%, specifically from about 30 to 55 wt%, and more specifically from about 40 to 50 wt%, on an elemental basis.

[0071] Additionally, according to exemplary embodiments, the loss on ignition (LOI) of the red mud may range from about 9 to 50 wt%, specifically from about 20 to 45 wt%, and more specifically from about 25 to 40 wt%.

[0072]

[0073] Reduction roasting

[0074] In the illustrated specific example, red mud is roasted in the presence of a carbon source as a reducing agent (reductive roasting). This roasting process can reduce hematite (Fe2O3) to magnetite (Fe3O4) according to the following reaction scheme 1, and as further reduction progresses, it can also be converted into metallic iron (Fe) as shown in reaction schemes 2 and 3.

[0075]

[0076] [Reaction Formula 1]

[0077] 3Fe2O3+ C → 2Fe3O4+ CO

[0078] [Reaction Formula 2]

[0079] Fe3O4+ C → 3FeO + CO

[0080] [Reaction Formula 3]

[0081] FeO + C → Fe + CO

[0082]

[0083] In this regard, the carbon source is not particularly limited as long as it has reducing capacity. For example, the carbon source may be coke, coal, charcoal, biochar, agricultural residues / waste paper, pulp processing waste, etc., and at least one selected from these may be used. According to a specific embodiment, coke may be used as the carbon source, as coke has a higher carbon calorific value (32 MJ / kg) than coal (24 MJ / kg) and does not exhibit smoke properties, which may be advantageous.

[0084] Meanwhile, according to the illustrated specific example, red mud can be reduced roasted in the presence of an alkali metal salt together with a carbon source as a reducing agent. In this regard, according to an exemplary specific example, the alkali metal salt can be at least one selected from the group consisting of sodium salts and potassium salts, and more specifically, a sodium salt.

[0085] According to the illustrated specific example, red mud is subjected to reduction roasting using a carbon source as a reducing agent in the presence of an alkali metal salt (specifically, a sodium salt). In this case, the roasting can be performed by filling a reduction furnace with red mud, a carbon source, and an alkali metal salt and raising the temperature. According to an exemplary embodiment, a furnace known in the art can be used as a device for performing reduction roasting, for example, at least one selected from a natural gas furnace, an oil furnace, an electric furnace, a waste oil furnace, a wood burning furnace, a dual fuel furnace, a single-stage furnace, a two-stage furnace, a bell furnace, a box furnace, a forging furnace, a pit furnace, a quenching furnace, a rotary furnace, a tempering furnace, etc.

[0086] When reduction roasting is performed without the addition of alkali metal salts, specifically sodium salts, sodium aluminum silicate (AlNa) 12 While the addition of alkali metal salts can promote the conversion of Fe2O3 in red mud into Fe3O4, the process of forming Fe2O3 in red mud can be accelerated.

[0087] In this regard, among alkali metal salts, sodium salt is an additive that can provide surface modification properties, and its addition can promote a reduction reaction during the reduction roasting process. According to exemplary specific examples, sodium salts include sodium carbonate, sodium sulfate, borax (Na2B4O7), etc., and potassium salts include potassium carbonate, potassium sulfate, potassium bicarbonate, etc., and at least one of the types listed above can be used. According to a specific specific example, a sodium salt, specifically sodium carbonate, can be used, and since sodium carbonate has a small ionic radius (about 1.02 Å), it can contribute to effectively converting hematite into magnetite during the reduction roasting process.

[0088] According to an exemplary specific example, when performing reduction roasting, the weight ratio of red mud: alkali metal salt: carbon source may be, for example, in the range of about 8 to 20: about 2 to 8: 1, specifically about 9 to 15: about 3 to 7: 1, and more specifically about 10 to 12: about 4 to 6: 1. In this regard, if the content ratio of the carbon source is too low or high, the reaction may be insufficient or an excessive reaction phenomenon may occur, and if the content ratio of the alkali metal salt is too low or high, the problem of insufficient reaction or an excessive reaction may also occur. Therefore, it may be advantageous to appropriately adjust within the above-mentioned range. However, the range of the content ratio is not necessarily limited thereto, as it may change depending on the types of the carbon source and alkali metal salt used, the composition of the red mud, etc.

[0089] According to an exemplary specific example, reduction roasting can be performed under elevated temperature conditions for reduction of iron (Fe) components in red mud, for example, in a range of about 600 to 1200°C, specifically about 700 to 1100°C, more specifically about 800 to 1000°C, and particularly specifically about 850 to 950°C. In addition, reduction roasting time can be determined in consideration of reaction efficiency, etc., for example, in a range of about 0.3 to 10 hours, specifically about 0.5 to 6 hours, and more specifically about 0.8 to 3 hours.

[0090] Through the above-described reduction roasting, for example, at least about 40 wt%, specifically at least about 50 wt%, more specifically at least about 55 wt%, and particularly specifically at least about 60 wt% of the Fe2O3 in the red mud can be reduced to Fe3O4 and / or elemental iron (Fe). Accordingly, the roasted red mud exists in the form of a mixture of magnetic and non-magnetic portions.

[0091]

[0092] Magnetic sorting

[0093] Referring back to Fig. 1, red mud that has undergone reduction roasting (heat treatment) can be separated into a magnetic portion and a non-magnetic portion through a magnetic separation step based on the principle of magnetic induction. The device used for magnetic separation is not limited to a specific type, and a magnetic separation device known in the art can be used. Specifically, the magnetic separation can be based on a mechanism that separates the magnetic portion and the non-magnetic portion by passing the powder of reduction roasted red mud through a magnetic field. In this regard, the magnitude of the magnetic field applied during magnetic separation can be set in consideration of magnetic induction characteristics, etc., and can be adjusted in the range of, for example, about 1,200 to 10,000 Gauss, specifically about 2,000 to 8,000 Gauss, and more specifically about 4,000 to 6,000 Gauss, but this can be understood as an example.

[0094] According to an exemplary specific example, the ratio (by weight) between the magnetic portion and the non-magnetic portion separated by magnetic separation may vary depending on the composition and properties of the starting material, red mud, reduction roasting treatment conditions, etc., and may be, for example, in the range of about 1 to 4:1, specifically about 1.2 to 3:1, and more specifically about 1.4 to 2:1, but is not limited thereto.

[0095] According to an exemplary embodiment, the iron content within the magnetic portion may be increased, on an elemental basis, by at least about 20 wt%, specifically at least about 25 wt%, and more specifically at least about 30 wt%. According to a specific embodiment, the iron content within the magnetic portion may be, on an elemental basis, by at least about 40 to 80 wt%, specifically at least about 50 to 75 wt%, and more specifically at least about 60 to 70 wt%.

[0096] In contrast, the titanium content in the magnetic portion may be significantly lower than in the initial red mud, since the titanium component tends to be concentrated in the non-magnetic portion. In this regard, the titanium content (on an elemental basis) in the magnetic portion may be, for example, about 15 wt% or less, specifically about 10 wt% or less, more specifically about 5 wt% or less, and particularly specifically about 2 wt% or less. According to a specific embodiment, the titanium content (on an elemental basis) in the magnetic portion may be, for example, about 1 wt% or less, specifically about 0.8 wt% or less, and more specifically about 0.5 wt% or less.

[0097] Meanwhile, the non-magnetic portion may have a reduced iron content compared to the initial red mud, for example, about 60 wt% or less, specifically about 50 wt% or less, more specifically about 30 wt% or less, and particularly specifically about 20 wt% or less, on an elemental basis. In certain specific embodiments, the iron content in the non-magnetic portion may be, for example, about 15 wt% or less, more specifically about 10 wt% or less, and even more specifically about 8 wt% or less, on an elemental basis.

[0098] On the other hand, the titanium content in the non-magnetic portion may be increased compared to the initial red mud, for example, at least about 4 wt% (based on element), specifically about 5 to 40 wt%, and more specifically about 7 to 30 wt%.

[0099] In the case of aluminum, since it can be contained in a balanced amount in each of the magnetic and non-magnetic portions, the reason why aluminum is not concentrated in the magnetic or non-magnetic portion by magnetic separation can be seen as factors such as the size of the external magnetic field. As an example, in each of the magnetic and non-magnetic portions, the aluminum content (based on element) can be determined in a range of, for example, about 3 to 10 wt%, specifically about 4 to 8 wt%, and more specifically about 5 to 7 wt%, but this can be understood as an example. As an example, aluminum in the magnetic and non-magnetic portions can typically exist in the form of sodium aluminate, free alkali metal (specifically, sodium cation), etc., which can be relatively easily dissolved during liquid leaching as described below with a soluble sodium component.

[0100]

[0101] Liquid leaching

[0102] According to the illustrated specific example, liquid leaching can be performed on each of the magnetic and non-magnetic portions of the red mud separated by magnetic separation, thereby removing most of the aluminum components contained therein (i.e., aluminum is dissolved in the leaching liquid). At this time, the leaching medium used in the liquid leaching can be an aqueous medium, and more specifically, water. According to a specific specific example, leaching can be performed using an aqueous medium without adding an acid or base, etc. (i.e., water leaching).

[0103] As an example, the leaching efficiency of aluminum may be, for example, at least about 80% (on an element basis), specifically the leaching efficiency of aluminum in the magnetic portion may be, for example, about 90 to 99%, more specifically about 93 to 96%, while the leaching efficiency of aluminum in the non-magnetic portion may be, for example, about 82 to 92%, specifically about 83 to 87%, but this may be understood to be by way of example only.

[0104] According to an exemplary embodiment, the liquid-to-solid (L / S) ratio during leaching for each of the magnetic portion and the non-magnetic portion may be, for example, in the range of about 50 or less, specifically about 1 to 20, more specifically about 3 to 15, and particularly specifically about 5 to 12.

[0105] In addition, the liquid leaching can be performed under elevated temperature conditions, and the leaching temperature can be controlled in the range of, for example, about 50 to 100°C, specifically about 70 to 98°C, and more specifically about 85 to 95°C. In addition, the water leaching time can be in the range of, for example, about 15 minutes to about 8 hours, specifically about 0.5 to 4 hours, and more specifically about 1 to 2 hours.

[0106] Through the above-described liquid leaching process, a liquid leaching solution (specifically, an aqueous leaching solution) containing aluminum ions (or compounds), and residues of the magnetic portion and non-magnetic portion can be formed, respectively. As in the illustrated specific example, solid-liquid separation techniques known in the art, such as filtration, centrifugation, sedimentation, decantation, and evaporation, can be applied to obtain the leaching solution and residues of the magnetic portion and non-magnetic portion, respectively.

[0107] In this regard, the concentration of aluminum in the liquid leachate for each of the magnetic portion and the non-magnetic portion may be set in a range of, for example, about 3 to 20 g / L, specifically about 4 to 18 g / L, and more specifically about 5 to 15 g / L, but is not limited thereto.

[0108] According to an exemplary embodiment, the iron content in the residue of the magnetic portion after leaching may be significantly increased compared to before leaching, for example, at least about 20 wt% (on an elemental basis), specifically at least about 25 wt%, and more specifically in a range of about 30 to 70 wt%, based on the residue of the magnetic portion. In particular, the relative amount of iron (on an elemental basis) in the residue of the magnetic portion may contain high-grade iron, for example, at least about 80 wt%, specifically at least about 85 wt%, and more specifically at least about 90 wt%, compared to the iron content contained in the initial red mud.

[0109] On the other hand, the contents of iron and titanium in the residue of the non-magnetic portion may not significantly change compared to before leaching, and compared to the non-magnetic portion before leaching, the rate of change in the iron content may be, for example, about 12% or less (specifically, about 10% or less, more specifically, about 6% or less), and the rate of change in the titanium content may be, for example, about 15% or less (specifically, about 10% or less, more specifically, about 8% or less).

[0110] According to an exemplary embodiment, the contents of aluminum and alkali metal (specifically, sodium) contained in the residue of each of the magnetic portion and the non-magnetic portion may be significantly reduced, such that the change in the aluminum content may be, for example, at least about 80% (specifically, at least about 85%, more specifically, at least about 90%), and the change in the alkali metal content may be, for example, at least about 80% (specifically, at least about 85%, more specifically, at least about 90%), compared to before the liquid leaching, but this may be understood as an exemplary intent.

[0111]

[0112] aluminum recovery

[0113] According to the illustrated specific example, a step of recovering aluminum contained in the liquid leachate of each of the magnetic portion and the non-magnetic portion or a combination (mixture) thereof is performed.

[0114] For this purpose, an acid may be added to a liquid leachate (specifically, an aluminum-containing aqueous leachate) to precipitate aluminum. As an example, an acid used for precipitating aluminum may be at least one selected from phosphoric acid (H3PO4), sulfuric acid (H2SO4), hydrochloric acid (HCl), nitric acid (HNO3), and the like, more specifically, phosphoric acid.

[0115] Specifically, the formation of a precipitate by the reaction of aluminum and phosphoric acid in the leachate can be explained as follows: First, aluminum ions react with phosphoric acid to form a soluble intermediate (compound), for example, AlH2PO4. 2+ , AlHPO 4+ This can be explained by the formation of AlPO4, which subsequently releases hydrogen ions and is converted to insoluble AlPO4.

[0116] According to an exemplary specific example, when adding aluminum by adding phosphoric acid, sodium phosphate (Na3PO4) may be additionally added together with phosphoric acid. In this case, the amount of sodium phosphate to be added may be determined in consideration of the molar concentration of aluminum in the solution, for example, the molar ratio of aluminum to sodium phosphate (Al / Na2PO4) may be adjusted in the range of about 0.5 to 2, specifically about 1 to 1.5. According to a specific specific example, sodium phosphate may be added in a stoichiometric amount.

[0117] According to an exemplary embodiment, during the precipitation reaction of aluminum in the leachate, the acid may be added so that the pH is, for example, in the range of about 3 to 6, specifically about 3.5 to 5.5, more specifically about 4 to 5. In addition, when adding the acid, it may be added in the form of an aqueous solution, in which case the acid concentration may be adjusted in the range of, for example, about 5 to 20% (v / v), specifically about 7 to 15% (v / v), more specifically about 9 to 12% (v / v), but the present disclosure is not limited thereto.

[0118] In addition, the temperature during the precipitation of aluminum can be set in a range of, for example, about 50 to 90°C, specifically about 55 to 85°C, more specifically about 60 to 80°C, and the precipitation time can be controlled in a range of, for example, about 0.1 to 5 hours, specifically about 0.5 to 4 hours, more specifically about 1 to 3 hours.

[0119] Through the above-described process, most of the aluminum in the leachate can be precipitated, and the aluminum precipitation efficiency in the leachate can be, for example, at least about 95 wt%, specifically at least about 99 wt%, and more specifically at least about 99.9 wt%. After the aluminum is precipitated in this way, the solid AlPO4 can be recovered using a solid-liquid separation means known in the art. Such a solid-liquid separation means can be, for example, at least one selected from filtration, centrifugation, decantation, evaporation, etc., and specifically, filtration.

[0120] Additionally, the purity of the AlPO4 precipitate may be, for example, at least about 99%, specifically at least about 99.5%, more specifically at least about 99.8%, and particularly specifically at least about 99.9%.

[0121]

[0122] Separation / recovery and utilization of iron

[0123] The magnetic residue obtained after liquid leaching contains a high concentration of iron. In the illustrated specific example, this magnetic residue can be recovered by forming metallic iron through a recovery process known in the art, such as a smelting process, which can then be separated from the slag. Since the details of this process are well known in the art, a detailed description thereof will be omitted.

[0124]

[0125] Titanium recovery

[0126] The non-magnetic residue obtained after liquid leaching contains a significant amount of titanium, and the titanium in the residue can be recovered by converting it into titania.

[0127] To this end, in the illustrated specific example, mixed sulfate roasting is first performed on the residue. In this regard, mixed sulfate roasting may be performed to convert titanium into a form that can be easily leachable in a subsequent process.

[0128] According to an exemplary embodiment, mixed sulfate roasting can be performed using sulfuric acid (H2SO4) and a sulfate, and can be performed in a manner of first mixing the residue of the non-magnetic portion with sulfuric acid and a sulfate, and then roasting under elevated temperature conditions. In this regard, the sulfate can be at least one selected from sodium sulfate (Na2SO4), ammonium sulfate ((NH4)2SO4), magnesium sulfate (MgSO4), potassium sulfate (K2SO4), barium sulfate (BaSO4), and the like. Among them, it can be advantageous to use sodium sulfate, which has a good reaction rate. In this way, titanium in the residue of the non-magnetic portion can be converted into a water-soluble form through mixed sulfate roasting. At this time, sulfuric acid can be used in a concentrated form (e.g., about 98% or more) during mixed sulfate roasting.

[0129] According to an exemplary embodiment, the weight ratio of sulfuric acid / sulfate during mixed sulfate roasting can be adjusted in the range of, for example, about 0.5 to 4, specifically about 1.5 to 2.5, more specifically about 1 to 2.

[0130] According to an exemplary embodiment, the ratio (by weight) of sulfuric acid and sulfate: residue of the non-magnetic portion can be adjusted taking into account the relative ratio with the reactants, for example, in the range of about 0.5 to 3:1, specifically about 0.7 to 2:1, more specifically about 0.9 to 1.5:1.

[0131] In addition, the mixed sulfate roasting temperature can be adjusted to a degree that can destroy the refractory matrix of the titanium compound among the non-magnetic residues, for example, in a range of about 180 to 400°C, specifically about 200 to 300°C, or about 220 to 270°C. In addition, the mixed sulfate roasting time can be adjusted to a range of, for example, about 0.5 to 5 hours, specifically about 1 to 4 hours, or more specifically about 1.5 to 3 hours, but this can be understood as an example.

[0132] According to the illustrated specific example, after mixed sulfate roasting, an aqueous medium or an acid solution (e.g., aqueous sulfuric acid solution) may be added to leach titanium in a water-soluble form in the roasted residue. At this time, the concentration of the sulfuric acid solution may be adjusted, for example, up to 10 M, specifically in a range of about 1 to 5 M, more specifically in a range of about 1.5 to 3 M. In addition, the liquid-to-solid ratio during the leaching process may be adjusted, for example, in a range of about 2 to 100, specifically in a range of about 3 to 20, more specifically in a range of about 5 to 10. In addition, the leaching may be performed at a temperature of, for example, about 70 to 150°C, specifically in a range of about 90 to 140°C, more specifically in a range of about 100 to 130°C, for about 0.2 to 5 hours, specifically in a range of about 0.5 to 4 hours, more specifically in a range of about 0.8 to 3 hours, but this may be understood to be illustrative. At this time, the leaching efficiency of titanium may be, for example, at least about 90%, specifically at least about 95%, and more specifically at least about 99%.

[0133] Through the above-described leaching process, a leaching solution containing a significant amount of titanium (specifically titanium sulfate) can be formed, wherein the titanium concentration in the leaching solution can be, for example, in the range of at least about 3 to 10 g / L, specifically in the range of at least about 4 to 9 g / L, and more specifically in the range of about 6 to 8 g / L.

[0134] Referring back to Figure 1, a hydrolysis step may be performed to precipitate titanium in the titanium-containing leachate. The pH during the hydrolysis step may be adjusted in consideration of the degree of precipitation, for example, less than about 3, specifically about 1 to 2, and more specifically about 1.2 to 1.8. At this time, the hydrolysis reaction may be performed under elevated temperature conditions, for example, in the range of about 70 to 100°C, specifically about 75 to 95°C, and more specifically about 85 to 92°C.

[0135] Through the hydrolysis process described above, the titanium compound (specifically, titanium sulfate) in the leachate can be precipitated in the form of metatitanic acid according to the following reaction scheme 2.

[0136]

[0137] [Reaction Formula 2]

[0138] TiOSO4+ 2H2O → H2TiO3↓+ H2SO4

[0139]

[0140] According to an exemplary embodiment, at least about 90 wt%, specifically at least about 92 wt%, and more specifically at least about 95 wt% of the titanium in the leachate can be converted to a precipitate.

[0141] Meanwhile, after the formation of titanium precipitate involving a hydrolysis reaction, the metatitanic acid precipitate can be recovered using a solid-liquid separation method known in the art. Such a solid-liquid separation method may be at least one selected from, for example, filtration, centrifugation, decantation, evaporation, etc., and specifically, filtration.

[0142] The metatitanic acid precipitate obtained as described above can be optionally subjected to a drying process, and then converted into titania through heat treatment in an oxygen-containing atmosphere, i.e., calcination.

[0143] According to an exemplary embodiment, the calcination temperature may be in the range of, for example, about 200 to 900°C, specifically about 400 to 800°C, and more specifically about 500 to 700°C. In addition, the calcination treatment time may be controlled in the range of, for example, about 0.5 to 10 hours, specifically about 1 to 5 hours, and more specifically about 2 to 4 hours.

[0144] The purity of the titania converted through the above-described calcination treatment may be, for example, at least 95%, specifically at least about 99%, and more specifically at least about 99.5%. In addition, the obtained titania may be in the rutile form.

[0145] As described above, according to the present disclosure, valuable metals including iron, aluminum, and titanium contained in red mud, a by-product of a bauxite treatment process, can be separated and recovered through a single treatment process. This is noteworthy in that it provides a method for recovering at least three types of valuable metals with high efficiency and high purity, compared to conventionally known dry or wet processes.

[0146] The present invention can be more clearly understood by the following examples, which are provided merely for the purpose of illustrating the present invention and are not intended to limit the scope of the invention.

[0147]

[0148] Example

[0149]

[0150] a. Material

[0151] The materials used in this example are as shown in Table 1 below.

[0152]

[0153] Material Name Manufacturer Na2CO3 Sigma-Aldrich Coke Powder H3PO4 Na3PO4 H2SO4 Na2SO4

[0154]

[0155] B. Analysis method

[0156] The analysis performed in this example was performed as follows.

[0157]

[0158] - XRD patterns were analyzed using X'pert PRO PANalytical.

[0159] - The composition was analyzed by inductively coupled plasma-induced emission spectroscopy using iCAP 7400 Duo (Thermo Scientific, USA).

[0160]

[0161] Example 1

[0162] In this example, red mud, a byproduct from a commercial alumina manufacturer, was obtained and used as the starting material. To homogenize and remove excess washable alkalinity, 1 kg of red mud cake was washed with water at 90°C for 1 h in a glass beaker under mechanical stirring at 300 rpm to a bulk density of 20% (w / v). This process was repeated for three different batches, which were dried overnight at 105°C with hot air, and a sufficient amount of sample was collected. The dried red mud was ground by hand, mixed appropriately, and a uniform powder was prepared for use in the next set of experiments. The weight loss was observed to be 42.02% of the initial cake used for washing. After aqua regia digestion, wet chemical analysis (composition analysis) was performed using inductively coupled plasma spectroscopy (ICP-OES, iCAP 7400 Duo, Thermo Scientific) as the analytical instrument, and the results are shown in Table 2.

[0163]

[0164] Main element FeAlNaSiTiC balance (including oxygen) (wt.%) 24.9 9.4 8.5 6.2 3.5 2.3 45.2

[0165]

[0166] According to the above table, red mud contains multiple balance components including the main components iron, aluminum, sodium, silicon, titanium, and carbon, and in the case of balance components, it was analyzed that in addition to oxygen, trace amounts of calcium, magnesium, yttrium, niobium, and strontium are contained.

[0167] Additionally, the physical properties of red mud were evaluated using XRD (X-ray diffraction) analysis, and the results are shown in Fig. 2.

[0168] Referring to the above drawing, Cancrinite [Na7(Al5Si7O 24)CO3.3H2O], Gibbsite [Al(OH)3], Anatase [TiO2], Quartz [SiO2], Hematite [Fe2O3] and Sodalite [Na6(Al6Si6O 24 )2NaF.xH2O] was found to be the main mineral phase in red mud.

[0169]

[0170] Example 2

[0171] 50 g of red mud powder and 25 g of Na2CO3 powder were mixed with coke powder at various weight ratios (coke powder: varied in the range of 1 to 5 g).

[0172] For another experiment, 50 g of red mud and 5 g of coke powder were mixed with various weight ratios of Na2CO3 powder (Na2CO3 powder: varied in the range of 0 to 25 g).

[0173] Reduction roasting was performed in an electric furnace at a heating rate of 10°C / min at 900°C under an inert atmosphere. The roasting time was 1 hour after reaching 900°C.

[0174] The phase transformation of the reduction roasted sample was analyzed using XRD, and the results are shown in Figs. 3 and 4, respectively.

[0175] Referring to Fig. 3, it was confirmed that reduction roasting performed without adding Na2CO3 formed sodium aluminum silicate, maghemite, and perovskite phases.

[0176] On the other hand, according to Fig. 4, for the sample roasted in the presence of 5 g of coke and 25 g of Na2CO3, it was confirmed that hematite was converted to magnetite, and the other major element was confirmed to be a sodium compound.

[0177]

[0178] Example 3

[0179] As in Example 2, red mud was subjected to reduction roasting at 900°C for 1 hour, resulting in a 27.5% weight loss compared to the initial red mud weight. Subsequently, 58 g of the reduction roasted sample was ground in a mortar and pestle, and then subjected to a magnetic separation experiment.

[0180] The magnetic and non-magnetic fractions were 64.7% and 35.3%, respectively, during magnetic separation. The compositions of these two fractions are shown in Table 3 below.

[0181]

[0182] Content (wt%)Element (wt%)FeAlNaSiTiCMagnetic part64.730.36.922.15.30.3not foundNon-magnetic part35.35.65.817.75.47.2not found

[0183]

[0184] Referring to the table above, it was confirmed that >91% of the iron content in the tested sample was separated through magnetic separation. However, the magnetically separated iron grade was found to be somewhat low due to the presence of other elements.

[0185]

[0186] Example 4

[0187] In order to improve the quality of iron, 37.5 g of a sample of the magnetic portion obtained through magnetic separation in Example 3 was leached with an aqueous medium (water) at a temperature of 90°C and a liquid-to-solid ratio of 10% (w / v) for 1 hour. The solid obtained was then filtered using Whatman No. 42 filter paper (pore size: 2.5 μm), and the solid was dried overnight under hot air conditions to obtain a water-leached residue. The analysis results confirmed that the obtained residue had a weight loss of approximately 48.6 wt% compared to the sample before water leaching. It is believed that this weight loss was due to the dissolution of soluble aluminum compounds such as sodium aluminate and other free alkali metals (mainly sodium).

[0188] After water leaching of the magnetic portion, elemental analysis of the residue was performed, and the results are shown in Table 4 below.

[0189]

[0190] Main element FeAlNaSiTi content (wt%) 58.5<0.7<0.2<0.2<0.6

[0191]

[0192] According to the table above, the elemental analysis of the water-leached residue from the magnetic portion showed a significant increase in the iron content from less than 25% to 58.5%. In particular, the recovered high-grade iron accounted for approximately 91 wt% of the total iron content in the initial red mud used for reduction roasting.

[0193]

[0194] Example 5

[0195] A sample of the non-magnetic portion obtained in Example 3 (containing 5.8 wt% aluminum) was subjected to water leaching under the same conditions as in Example 4 (10% (w / v) of the ore density, 90°C, 1 hour). Subsequently, the solid obtained by filtration using Whatman No. 42 filter paper (pore size: 2.5 μm) was dried overnight under hot air conditions to obtain a water-leaching residue. The analysis results confirmed that the obtained residue had a weight loss of approximately 17.1 wt% compared to the sample before water leaching. It is believed that this weight loss was due to the dissolution of soluble aluminum compounds such as sodium aluminate and other free alkali metals (mainly sodium).

[0196] After leaching of the non-magnetic portion, elemental analysis of the residue was performed, and the results are shown in Table 5 below.

[0197]

[0198] Main element FeAlNaSiTi content (wt%) 5.9 1.2 0.8 5.6 7.8

[0199]

[0200] According to the above table, the elemental analysis results for the water leaching residue of the non-magnetic portion showed that the titanium grade increased from 3.5% to 7.8%.

[0201]

[0202] Example 6

[0203] Aluminum was recovered by mixing together 360 mL of the leachate (containing 6.8 g / L aluminum) obtained in Example 4 and 165 mL of the leachate (containing 5.93 g / L aluminum) obtained in Example 5.

[0204] The pH of the mixed solution was adjusted to 4.7 under stirring conditions using a 10% (v / v) H3PO4 solution, and then Na3PO4 was added in a stoichiometric amount and a precipitation reaction was performed for 1 hour in a solution heated to 70°C to form a white precipitate of AlPO4.

[0205] The slurry containing the precipitate was filtered under high temperature conditions, and the separated precipitate was washed twice with hot water and dried to obtain AlPO4. Elemental analysis was performed on the aqueous leachate of the magnetic portion, the aqueous leachate of the non-magnetic portion, the filtrate after precipitation, and the precipitate after drying, respectively, and the results are shown in Table 6 below.

[0206]

[0207] Volume / WeightAl ConcentrationAl Absolute AmountPrecipitation EfficiencyProduct PurityMagnetic Part Leachate360 mL6.80 g / L2.45 g--Nonmagnetic Part Leachate165 mL5.93 g / L0.98 g--Filtrate after Precipitation505 mL0.06 g / L0.03 g99.12%-AlPO4 Precipitate after Drying15.35 g22.1%3.39 g98.83%99.89%

[0208]

[0209] According to the above table, 15.3 g of AlPO4 precipitate was obtained with a precipitation efficiency of 99.12%. In addition, chemical analysis results showed that the final product (i.e., AlPO4) with a purity of 99.1% was obtained with a recovery rate of approximately 83% compared to the initial aluminum content in the red mud.

[0210]

[0211] Example 7

[0212] In Example 5, 17 g of titanium-containing residue obtained by water leaching of the non-magnetic portion (containing 7.8 wt% Ti) was roasted with a concentrated sulfate mixture (H2SO4 + Na2SO4) (mixed-sulfate roasting). At this time, a sulfate source was prepared at a ratio (by weight) of H2SO4:Na2SO4 of 2:1. This sulfate mixture was placed in a porcelain crucible so that a paste was prepared with the titanium residue at a mixture:residue ratio of 1:1.

[0213] Next, the crucible was roasted for 2 hours in an electric furnace preheated to 240°C, and then the crucible was cooled inside the electric furnace.

[0214] The cooled crucible was washed with 170 mL of 2 M H2SO4 solution in a 250 mL leaching reactor, maintaining a 10% (w / v) slurry density (liquid-to-solid ratio: 10). The slurry was stirred for 1 h at a boiling temperature of 120 °C in a closed system equipped with a condenser. Afterwards, the slurry was cooled to room temperature (approximately 25 °C) and the filtrate was recovered.

[0215] After heating 200 mL of the filtrate (containing 6.02 g / L Ti) at 90°C, the pH in the solution was adjusted to 1.5 for titanium hydrolysis.

[0216] After precipitation for 1 hour, the titanium precipitate was recovered through filtration with a precipitation efficiency of 95%, and the obtained precipitate was calcined in an electric furnace at 600°C for 2 hours to recover titania (2.2 g) with a purity of >99.5%.

[0217] Simple modifications or changes of the present invention can be easily utilized by those skilled in the art, and all such modifications or changes can be considered to be included in the scope of the present invention.

Claims

1. a) A step of providing red mud containing iron, aluminum and titanium; b) a step of separating the magnetic portion and the non-magnetic portion through magnetic separation after reducing and roasting the red mud; and c) A step of recovering iron from the residue of the magnetic portion remaining after performing liquid leaching on each of the magnetic portion and non-magnetic portion, and titanium from the residue of the non-magnetic portion, while recovering aluminum from the leaching liquid of the liquid leaching; A process for recovering valuable metals from red mud, including 2. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, in step c), aluminum is recovered by adding acid to the leachate and precipitating it.

3. A process for recovering valuable metals from red mud, characterized in that in the second paragraph, the aluminum is precipitated under conditions controlled in the range of pH 3 to 6, and at this time, the acid is phosphoric acid (H3PO4) and recovered as a precipitate in the form of AlPO4.

4. A process for recovering valuable metals from red mud, characterized in that, in the third paragraph, sodium phosphate (Na3PO4) is additionally added together with phosphoric acid, and the amount of sodium phosphate added is controlled in a range of 0.5 to 2 in the molar ratio of aluminum to sodium phosphate.

5. A process for recovering valuable metals from red mud, characterized in that, in the second paragraph, the precipitation of aluminum is performed at 50 to 90°C for 0.1 to 5 hours.

6. A process for recovering valuable metals from red mud, characterized in that, in the first paragraph, in step c), titanium is recovered in the form of titania (TiO2).

7. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the red mud contains, on an element basis, 5 to 60 wt% of iron (Fe), 3 to 40 wt% of aluminum, 2 to 20 wt% of titanium, and the balance.

8. A process for recovering valuable metals from red mud, characterized in that, in paragraph 7, the content of balance in the red mud is in the range of 25 to 60 wt% on an elemental basis.

9. A process for recovering valuable metals from red mud, characterized in that the LOI (loss on ignition) of the red mud in the first paragraph is in the range of 9 to 50 wt%.

10. A process for recovering valuable metals from red mud, characterized in that, in the first paragraph, the reduction roasting of step b) is performed in the presence of a carbon source as a reducing agent and an alkali metal salt selected from the group consisting of sodium salts and potassium salts as an additive.

11. In paragraph 10, the carbon source is at least one selected from the group consisting of coke, coal, charcoal, biochar, agricultural residue / paper waste, and pulp processing waste, and A process for recovering valuable metals from red mud, characterized in that the alkali metal salt is at least one selected from the group consisting of sodium carbonate, sodium sulfate, borax (Na2B4O7), potassium carbonate, potassium sulfate, and potassium bicarbonate.

12. A process for recovering valuable metals from red mud, characterized in that in paragraph 10, the weight ratio of red mud: alkali metal salt: carbon source is controlled in the range of 8 to 20:2 to 8:

1.

13. A process for recovering valuable metals from red mud, characterized in that the reduction roasting of step b) in paragraph 1 is performed at a temperature controlled in the range of 600 to 1200°C.

14. A process for recovering precious metals from red mud, characterized in that the weight ratio of the magnetic portion and the non-magnetic portion separated in the step b) in the first paragraph is in the range of 1 to 4:

1.

15. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the liquid-to-solid (L / S) ratio during liquid leaching of each of the magnetic portion and the non-magnetic portion in step c) is controlled in the range of 1 to 50.

16. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the magnetic part in step c) contains iron of at least 20 wt% (based on element) of iron.

17. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the non-magnetic portion in step c) contains at least 4 wt% (based on element) of titanium.

18. A process for recovering valuable metals from red mud, characterized in that the recovery of titanium in step c) of paragraph 1 involves a step of mixed-sulfate roasting of the non-magnetic portion using sulfuric acid and sulfate.

19. A process for recovering valuable metals from red mud, characterized in that in paragraph 18, the sulfate is at least one selected from the group consisting of sodium sulfate (Na2SO4), ammonium sulfate ((NH4)2SO4), magnesium sulfate (MgSO4), potassium sulfate (K2SO4), and barium sulfate (BaSO4).

20. A process for recovering valuable metals from red mud, characterized in that in paragraph 18, the weight ratio of sulfuric acid / sulfate during mixed-sulfate roasting is controlled in the range of 0.5 to 4.

21. A process for recovering valuable metals from red mud, characterized in that, in paragraph 18, the non-magnetic portion subjected to the mixed-sulfate roasting is leached with sulfuric acid at a temperature of 70 to 150°C to form a titanium-containing leachate, wherein the concentration of sulfuric acid is up to 10 M, and the liquid-solid ratio is controlled in the range of 2 to 100.

22. A process for recovering valuable metals from red mud, characterized in that in claim 21, titanium-containing leachate is heated to a temperature of 70 to 100°C and titanium hydrolysis is performed while adjusting the pH to less than 3 to form a titanium-containing precipitate.

23. A process for recovering valuable metals from red mud, characterized in that the titanium-containing precipitate is converted into titania by calcining it in an oxygen-containing atmosphere and at a temperature of 200 to 900°C for 0.5 to 10 hours in accordance with paragraph 22.

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