WO2025005481A1 - Process for recovering valuable metal from red mud - Google Patents

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

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 for recovering iron in red mud generated as a by-product in a bauxite processing process, and a wet process for further recovering aluminum and titanium, respectively, after separation of iron.

Red mud is a red process byproduct generated in a process of manufacturing alumina using bauxite as a raw material (i.e., the Bayer process), and is also called bauxite residue, red mud, etc. In the case of the Bayer process, in order to recover alumina from bauxite, caustic soda (NaOH) is added and processed under high temperature and high pressure. The reaction mechanism by which red mud is generated in the process can be expressed as shown in the following reaction scheme 1.

[Reaction Formula 1]

2NaOH + Al ₂ O ₃ ·2H ₂ O → Na ₂ O ·Al ₂ O ₃ + 4H ₂ O + red mud

As the demand for alumina has increased recently, the amount of red mud generated has also increased. Generally, red mud is generated at a rate of about 1 to 1.5 tons per ton of alumina produced. As the demand for alumina increases, the amount of red mud generated has also increased proportionally. The pH of red mud is approximately 9.2 to 12.8, and due to its strong alkaline properties, it is treated as designated waste.

In the past, rather than efficiently utilizing the red mud generated, it was disposed of by methods such as underground landfill and ocean dumping. However, problems have arisen due to the contamination of soil, surface water, and groundwater by corrosive leachate from the red mud accumulated in large quantities, and the cost of its disposal is also increasing. Furthermore, the amount of disposal is limited compared to the increasing amount of red mud generated by the existing disposal methods.

Red mud contains a significant amount of valuable metals such as iron (Fe), aluminum (Al), titanium (Ti), silicon (Si), gallium (Ga), and rare earth elements such as scandium (Sc) and yttrium (Y). Therefore, research is actively being conducted to recover and recycle multiple valuable metals contained in red mud.

For example, in order to recover iron in red mud, a method has been proposed in which a carbonaceous material (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 (Fe ₂ O ₃ ) into magnetite (Fe ₃ O ₄ ), and then 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., US Patent No. 9,023,301).

However, in the case of the dry smelting process, the loss of titanium and aluminum in the slag is induced, 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.

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 for highly efficiently recovering the main valuable metals in red mud, such as iron, aluminum, and titanium.

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.

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.

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

According to one specific example of the present disclosure,

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 the non-magnetic portion, and titanium from the residue of the non-magnetic portion, while recovering aluminum from the leachate of the liquid leaching;

A process for recovering valuable metals from red mud including .

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

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 AlPO ₄ using phosphoric acid (H ₃ PO ₄ ).

According to an exemplary specific example, sodium phosphate (Na ₃ PO ₄ ) is additionally added together with phosphoric acid, and at this time, the amount of sodium phosphate added can be adjusted in a range where the molar ratio of aluminum to sodium phosphate is 0.5 to 2.

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

According to an exemplary embodiment, in step c), titanium can be recovered in the form of titania (TiO ₂ ).

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.

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.

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

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

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 and pulp processing waste, and

The above alkali metal salt may be at least one selected from the group consisting of sodium carbonate, sodium sulfate, borax (Na ₂ B ₄ O ₇ ), potassium carbonate, potassium sulfate, and potassium bicarbonate.

According to an exemplary specific example, the weight ratio of red mud: alkali metal salt: carbon source can be controlled in the range of 8 to 20:2 to 8:1.

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

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

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

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.

According to an exemplary embodiment, the magnetic portion in step c) may contain at least 20 wt.-% (based on element) of iron.

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

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.

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.

According to an exemplary embodiment, the sulfate may be at least one selected from the group consisting of sodium sulfate (Na ₂ SO ₄ ), ammonium sulfate ((NH ₄ ) ₂ SO ₄ ), magnesium sulfate (MgSO ₄ ), potassium sulfate (K ₂ SO ₄ ), and barium sulfate (BaSO ₄ ).

According to an exemplary specific example, 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 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.

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.

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

The process according to the specific example of the present disclosure can recover each of iron, aluminum, and titanium, which are valuable metals, from red mud, which is a by-product of a bauxite processing process, with high efficiency through a single process, and is particularly advantageous in terms of the environment and waste reduction since it is configured to be operable with relatively low energy consumption and low reagent usage. In addition, the process according to the specific example of the present disclosure provides the advantage of effectively suppressing the generation of secondary waste such as slag while reducing carbon emissions compared to a conventional smelting process. Therefore, widespread commercialization is expected in the future.

FIG. 1 is a flow chart of an exemplary process according to one specific example;

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

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

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).

The present invention can be achieved by the following description. It should be understood that the following description describes preferred embodiments of the present invention, and the present invention is not necessarily limited thereto. In addition, the attached drawings are provided to aid understanding, and the present invention is not limited thereto, and details regarding individual components can be appropriately understood by the specific intent of the related description described below.

The terms used in this specification can be defined as follows.

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

"Reductive roasting" may refer to a process in which a metal oxide is reduced to the metal or a compound of lower oxidation number under a reducing atmosphere.

"Magnetic separation" may broadly mean a process of separating a mixture using magnets, 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 other components.

"Leaching" may mean a process of removing a substance by dissolution in a percolating liquid.

When a numerical range is specified in this specification as a lower limit and/or an upper limit, it is to be understood that any sub-combination within that numerical range is also disclosed. For example, when described as "1 to 5", it can include 1, 2, 3, 4, and 5, as well as any sub-combination therebetween.

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

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

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

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, through a combination of dry treatment and wet treatment in a physical and chemical manner.

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.

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

Red mud is typically obtained in the form of fine grains, and prior to processing, a washing process (specifically, water washing) can be performed at least once in order to remove impurities such as various dusts attached to the surface of the particles in addition to the components inherently contained in the red mud and to lower the high alkalinity. As an example, the amount of water used in the water washing process can 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 can 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.

After performing the water washing as described above, a drying process may be performed, and optionally, a filtration process may be performed prior to drying to separate solids, and the filtrate produced at this time may be reused in the water washing process. In addition, the drying process may be performed in a conventional manner known in the art, for example, heating (hot air) drying, reduced pressure drying, etc. may be applied.

Referring back to Fig. 1, the particle size of the red mud may be determined in consideration of factors affecting the treatment process 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.

According to an exemplary specific example, the content of iron (Fe) in the red mud may be, for example, in a range of about 5 to 60 wt%, specifically about 15 to 50 wt%, and more specifically 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 appropriate iron (Fe ₂ O ₃ ).

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

In addition, the content of titanium (Ti) in the red mud may be, for example, in a 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.

Meanwhile, red mud may contain balance in addition to the aforementioned components, and such 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.

According to an exemplary embodiment, the balance content in the red mud can be, for example, in a range of about 25 to 60 wt%, specifically about 30 to 55 wt%, more specifically about 40 to 50 wt%, on an elemental basis.

Additionally, according to exemplary embodiments, the loss on ignition (LOI) of the red mud may be in the range of, for example, about 9 to 50 wt%, specifically about 20 to 45 wt%, and more specifically about 25 to 40 wt%.

Reduction roasting

In the illustrated specific example, red mud is roasted in the presence of a carbon source as a reducing agent (reductive roasting). This roasting treatment can reduce hematite (Fe ₂ O ₃ ) to magnetite (Fe ₃ O ₄ ) 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.

[Reaction Formula 1]

3Fe ₂ O ₃ + C → 2Fe ₃ O ₄ + CO

[Reaction Formula 2]

Fe ₃ O ₄ + C → 3FeO + CO

[Reaction Formula 3]

FeO + C → Fe + CO

In this regard, the carbon source is not particularly limited as long as it has a reducing ability. For example, the carbon source may be coke, coal, charcoal, biochar, agricultural residues/waste paper, pulp processing waste, etc., and may be at least one selected from these. 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.

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 the exemplary specific example, the alkali metal salt can be at least one selected from the group consisting of a sodium salt and a potassium salt, and more specifically, can be a sodium salt.

According to the illustrated specific example, red mud is subjected to reduction roasting by 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 increasing 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.

When reduction roasting is performed without adding an alkali metal salt, specifically a sodium salt, sodium aluminum silicate (AlNa ₁₂ SiO ₅ ), maghemite (Fe ₂ O ₃ , γ-Fe ₂ O ₃ )), perovskite, etc. are formed, whereas when an alkali metal salt is added, the process of converting Fe ₂ O ₃ in red mud to Fe ₃ O ₄ can be promoted.

In this regard, among the 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 embodiments, examples of the sodium salt include sodium carbonate, sodium sulfate, borax (Na ₂ B ₄ O ₇ ), and the like, and examples of the potassium salt include potassium carbonate, potassium sulfate, potassium bicarbonate, and the like, and at least one of the types listed above can be used. According to a specific embodiment, 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.

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 a 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, when the content ratio of the carbon source is excessively low or high, the reaction may be insufficient or an excessive reaction phenomenon may occur, and also when the content ratio of the alkali metal salt is excessively low or high, the problem of insufficient reaction or excessive reaction may 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 since it may change depending on the types of the carbon source and alkali metal salt used, the composition of the red mud, etc.

According to an exemplary specific example, the 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, the reduction roasting time can be determined in consideration of the reaction efficiency, etc., and can be, 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.

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 Fe 2 O 3 in the red mud can be reduced to Fe ₃ O ₄ and/or elemental iron (Fe). Accordingly, the roasted red mud exists in the form of a mixture of magnetic portions and non-magnetic portions.

Magnetic sorting

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 for separating 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 size 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.

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.

According to an exemplary embodiment, the iron content in the magnetic portion can be increased, on an elemental basis, compared to the initial red mud, for example, 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 in the magnetic portion can be, on an elemental basis, for example, about 40 to 80 wt %, specifically about 50 to 75 wt %, and more specifically about 60 to 70 wt %.

In contrast, the titanium content in the magnetic portion may be significantly lower than in the initial red mud, because 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.

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. According to a specific specific example, 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 more specifically about 8 wt% or less, on an elemental basis.

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%.

In the case of aluminum, since it can be contained in a balanced amount in each of the magnetic portion and the non-magnetic portion, the reason why aluminum is not concentrated in the magnetic portion or the non-magnetic portion by magnetic separation can be considered to be due to factors such as the size of the external magnetic field. As an example, in each of the magnetic portion and the non-magnetic portion, 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 portion and the non-magnetic portion can typically exist in the form of sodium aluminate, a free alkali metal (specifically, a sodium cation), etc., which can be relatively easily dissolved during the liquid leaching described below as a soluble sodium component.

Liquid leaching

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

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 is to be understood as an example.

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 a range of about 50 or less, specifically about 1 to 20, more specifically about 3 to 15, and particularly specifically about 5 to 12.

In addition, the liquid leaching can be performed under elevated temperature conditions, and the leaching temperature can be controlled in a 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 a 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.

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 residues of the non-magnetic portion can be formed, respectively. As in the illustrated specific example, the leaching solution and residues of each of the magnetic portion and the non-magnetic portion can be obtained by applying a solid-liquid separation technique known in the art, such as filtration, centrifugation, sedimentation, decantation, evaporation, etc.

In this regard, the concentration of aluminum in the liquid leachate for each of the magnetic portion and the non-magnetic portion may be determined 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.

According to an exemplary embodiment, the iron content in the residue of the magnetic portion after leaching can be significantly increased compared to before leaching, for example, at least about 20 wt% (based on element), 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 (based on element) in the residue of the magnetic portion can 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%, relative to the iron content contained in the initial red mud.

On the other hand, the contents of each 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).

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 can be significantly reduced, such that the change in the aluminum content can 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 can 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 can be understood as an exemplary intent.

Recovery of aluminum

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

For this purpose, an acid may be added to a liquid leachate (specifically, an aluminum-containing aqueous leachate) to precipitate aluminum. As an example, examples of the acid used for precipitating aluminum include at least one selected from phosphoric acid (H ₃ PO ₄ ), sulfuric acid (H ₂ SO ₄ ), hydrochloric acid (HCl), nitric acid (HNO ₃ ), and the like, more specifically, phosphoric acid.

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 soluble intermediates (compounds), such as AlH ₂ PO ₄ ²⁺ , AlHPO ^{4+ ,} etc., and subsequently convert to insoluble AlPO ₄ while releasing hydrogen ions.

According to an exemplary specific example, when adding aluminum by adding phosphoric acid, sodium phosphate (Na ₃ PO ₄ ) 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/Na ₂ PO ₄ ) may be adjusted in a 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.

According to an exemplary specific example, during the precipitation reaction of aluminum in the leachate, the acid may be added such that the pH is in a range of, for example, about 3 to 6, specifically about 3.5 to 5.5, more specifically about 4 to 5. In addition, the acid may be added in the form of an aqueous solution, in which case the acid concentration may be adjusted in a 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.

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.

Through the above-described process, most of the aluminum in the leachate can be precipitated, and the efficiency of precipitating aluminum 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 AlPO ₄ can be recovered using a solid-liquid separation means known in the art. The solid-liquid separation means can be, for example, at least one selected from filtration, centrifugation, decantation, evaporation, and the like, and specifically, filtration.

Additionally, the purity of the AlPO ₄ precipitate can be, for example, at least about 99%, specifically at least about 99.5%, more specifically at least about 99.8%, and especially specifically at least about 99.9%.

Separation/recovery and utilization of iron

The residue of the magnetic portion obtained after the liquid leaching contains a high concentration of iron, and in the illustrated specific example, the residue of the magnetic portion can be recovered by forming metallic iron through a recovery process known in the art, for example, a smelting process, and separating it from the slag. Since the details of this process are known in the art, a description of the details is omitted.

Titanium recovery

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 form.

For this purpose, in the illustrated specific example, mix sulfation roasting is first performed on the residue. In this regard, mixed sulfation roasting can be performed to convert titanium into a form that can be easily leached in a subsequent process.

According to an exemplary embodiment, mixed sulfate roasting can be performed using sulfuric acid (H ₂ SO ₄ ) and a sulfate, wherein the residue of the non-magnetic portion is first mixed with sulfuric acid and sulfate, and then roasted under elevated temperature conditions. In this regard, the sulfate can be at least one selected from sodium sulfate (Na ₂ SO ₄ ), ammonium sulfate ((NH ₄ ) ₂ SO ₄ ), magnesium sulfate (MgSO ₄ ), potassium sulfate (K ₂ SO ₄ ), barium sulfate (BaSO ₄ ), and the like. Among them, it can be advantageous to use sodium sulfate having 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 (for example, about 98% or more) during the mixed sulfate roasting.

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

According to an exemplary embodiment, the ratio (by weight) of the sulfuric acid and sulfate:residue of the non-magnetic portion can be controlled taking into account the relative ratio with respect to 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.

In addition, the mixed sulfate roasting temperature can be adjusted to an extent capable of destroying the refractory matrix of the titanium compound among the residues of the non-magnetic portion, 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, for example, in a range of 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.

According to the illustrated specific example, after the mixed sulfate roasting, an aqueous medium or an acid solution (e.g., an 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 controlled, 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 controlled, 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 is to be understood as an exemplary embodiment. 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%.

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

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

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

[Reaction Formula 2]

TiOSO ₄ + 2H ₂ O → H ₂ TiO ₃ ↓+ H ₂ SO ₄

In exemplary embodiments, 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.

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

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.

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

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

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 by 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.

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.

Example

a. Material

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

Material name manufacturing company Na ₂ CO ₃ Sigma-Aldrich coke powder H ₃ PO ₄ Na ₃ PO ₄ H ₂ SO ₄ Na ₂ SO ₄

B. Analysis method

The analysis performed in this example was performed as follows.

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

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

Example 1

In this example, red mud, which is a by-product 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 to a bulk density of 20% (w/v) at 90 °C for 1 hour in a glass beaker under mechanical stirring at 300 rpm. This 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 properly, and prepared into a uniform powder to be used in the next set of experiments. The weight loss was observed to be 42.02% from the initial cake used for washing. After undergoing aqua regia decomposition, wet chemical analysis (composition analysis) was performed using inductive coupled plasma spectroscopy (ICP-OES, iCAP 7400 Duo, Thermo Scientific) as an analyzer, and the results are given in Table 2 below.

Main Element Fe Al Na Si Ti C Balance (including oxygen) (wt.%) 24.9 9.4 8.5 6.2 3.5 2.3 45.2

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.

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

Referring to the above drawing, Cancrinite [Na ₇ (Al ₅ Si ₇ O ₂₄ )CO ₃ .3H ₂ O], Gibbsite [Al(OH) ₃ ], Anatase [TiO ₂ ], Quartz [SiO ₂ ], Hematite [Fe ₂ O ₃ ], and Sodalite [Na ₆ (Al ₆ Si ₆ O ₂₄ )2NaF. x H ₂ O] were found to be the main mineral phases in the red mud.

Example 2

Fifty grams of red mud powder and 25 g of Na ₂ CO ₃ powder were mixed with various weight ratios of coke powder (coke powder: varied in the range of 1 to 5 g).

In another experiment, 50 g of red mud and 5 g of coke powder were mixed with various weight ratios of Na ₂ CO ₃ powder (Na ₂ CO ₃ powder: varied in the range of 0 to 25 g).

Reduction roasting was performed in an electric furnace at a heating rate of 10 ℃/min at 900 ℃ under an inert atmosphere. At this time, the roasting time was 1 hour after reaching 900 ℃.

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

Referring to Fig. 3, it was confirmed that reduction roasting performed without adding Na ₂ CO ₃ formed sodium aluminum silicate, maghemite, and perovskite phases.

In contrast, according to Fig. 4, for the sample roasted in the presence of 5 g of coke and 25 g of Na ₂ CO ₃ , it was confirmed that hematite was converted to magnetite, and the other major elements were confirmed to be sodium compounds.

Example 3

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

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

division Content (wt%) Element (wt%) Fe Al Na Si Ti C magnetic part 64.7 30.3 6.9 22.1 5.3 0.3 not found Non-magnetic part 35.3 5.6 5.8 17.7 5.4 7.2 not found

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 grade of the magnetically separated iron was confirmed to be somewhat low due to the presence of other elements.

Example 4

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

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

Main Element Fe Al Na Si Ti Content (wt%) 58.5 <0.7 <0.2 <0.2 <0.6

According to the above table, the elemental analysis results for the water-leached residue of the magnetic portion showed that the iron grade significantly increased from less than 25% to 58.5%. In particular, the recovered high-grade iron accounted for approximately 91 wt% of the total iron element in the initial red mud introduced into the reduction roasting.

Example 5

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

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

Main Element Fe Al Na Si Ti Content (wt%) 5.9 1.2 0.8 5.6 7.8

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

Example 6

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.

The pH of the mixed solution was adjusted to 4.7 under stirring conditions using 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 ℃ to form a white precipitate of _AlPO4 .

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 AlPO ₄ . 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.

division Volume/Weight Al concentration Al absolute quantity Sedimentation efficiency product
water Extraction of magnetic part 360 mL 6.80 g/L 2.45 g - - Leachate of non-magnetic part 165 mL 5.93 g/L 0.98 g - - Filtrate after sedimentation 505 mL 0.06 g/L 0.03 g 99.12% - AlPO ₄ precipitate after drying 15.35 g 22.1% 3.39 g 98.83% 99.89%

According to the above table, 15.3 g of AlPO ₄ precipitate was obtained with a precipitation efficiency of 99.12%. In addition, the chemical analysis results showed that the final product (i.e., AlPO ₄ ) with a purity of 99.1% was obtained with a recovery rate of about 83% with respect to the initial aluminum in the red mud.

Example 7

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 (H ₂ SO ₄ + Na ₂ SO ₄ ) (mixed-sulfate roasting). At this time, a sulfate source was prepared at a ratio of H ₂ SO ₄ : Na ₂ SO ₄ of 2 : 1 (by weight). This sulfate mixture was placed in a porcelain crucible and a paste was prepared with the titanium residue such that the ratio of the mixture : residue was 1 : 1.

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.

The cooled crucible was washed with 170 mL of 2 MH ₂ SO ₄ solution in a 250 mL leaching reactor to maintain 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 under a closed system equipped with a condenser. After that, the slurry was cooled to room temperature (about 25 °C) and the filtrate was recovered.

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

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%.

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 (23)

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 the non-magnetic portion, and titanium from the residue of the non-magnetic portion, while recovering aluminum from the leachate of the liquid leaching; A process for recovering valuable metals from red mud, including A process for recovering valuable metals from red mud, characterized in that in the first paragraph, in step c), aluminum is recovered by precipitating it by adding acid to the leachate. 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 (H ₃ PO ₄ ) and recovered as a precipitate in the form of AlPO ₄ . A process for recovering valuable metals from red mud, characterized in that in the third paragraph, sodium phosphate (Na ₃ PO ₄ ) is additionally added together with phosphoric acid, and the amount of sodium phosphate added is controlled in a range of a molar ratio of aluminum to sodium phosphate of 0.5 to 2. 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. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, titanium in the step c) is recovered in the form of titania (TiO ₂ ). 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. A process for recovering valuable metals from red mud, characterized in that in claim 7, the content of balance in the red mud is in the range of 25 to 60 wt% on an element basis. A process for recovering valuable metals from red mud, characterized in that in claim 1, the LOI (loss on ignition) of the red mud is in the range of 9 to 50 wt%. 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 a sodium salt and a potassium salt as an additive. In claim 10, the carbon source is at least one selected from the group consisting of coke, coal, charcoal, biochar, agricultural residues/paper 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 (Na ₂ B ₄ O ₇ ), potassium carbonate, potassium sulfate, and potassium bicarbonate. A process for recovering valuable metals from red mud, characterized in that in claim 10, the weight ratio of red mud: alkali metal salt: carbon source is controlled in a range of 8 to 20:2 to 8:1. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the reduction roasting of step b) is performed at a temperature controlled in the range of 600 to 1200°C. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the weight ratio of the magnetic portion: non-magnetic portion separated in the step b) is in the range of 1 to 4:1. 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. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the magnetic portion in step c) contains iron of at least 20 wt% (based on element) of iron. 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. A process for recovering valuable metals from red mud, characterized in that in the first paragraph, the recovery of titanium in step c) involves a step of mixed-sulfate roasting of the non-magnetic portion using sulfuric acid and sulfate. A process for recovering valuable metals from red mud, characterized in that in claim 18, the sulfate is at least one selected from the group consisting of sodium sulfate (Na ₂ SO ₄ ), ammonium sulfate ((NH ₄ ) ₂ SO ₄ ), magnesium sulfate (MgSO ₄ ), potassium sulfate (K ₂ SO ₄ ), and barium sulfate (BaSO ₄ ). A process for recovering valuable metals from red mud, characterized in that in claim 18, the weight ratio of sulfuric acid/sulfate during mixed-sulfate roasting is controlled in the range of 0.5 to 4. A process for recovering valuable metals from red mud, characterized in that in claim 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-to-solid ratio is controlled in the range of 2 to 100. A process for recovering valuable metals from red mud, characterized in that in claim 21, the 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. A process for recovering valuable metals from red mud, characterized in that in claim 22, the titanium-containing precipitate is converted into titania by calcining it under the conditions of an oxygen-containing atmosphere and a temperature of 200 to 900° C. for 0.5 to 10 hours.

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