Method for completely recovering valuable components in low-grade vanadium titano-magnetite based on hydrogen metallurgy

Abstract

The invention discloses a total recovery method for valuable components in low-grade vanadium titano-magnetite based on hydrogen metallurgy, and belongs to the technical field of comprehensive utilization of metallurgical resources. According to the method, vanadium titano-magnetite oxidized pellets serve as raw materials, low-temperature selective reduction is carried out under the pure hydrogen atmosphere, and metallized pellets with the metallization ratio ranging from 88% to 94% are obtained; then the metallized pellets are fed into a multi-electrode electric melting furnace for melting separation, and efficient separation of molten iron and vanadium-rich titanium slag is achieved through optimized arrangement of an electrode array and precise injection of a flux; the vanadium-rich titanium slag is roasted through a complexing agent, so that vanadium is converted into water-soluble vanadate, then vanadium is extracted through normal-pressure acid leaching, and a vanadium product and the titanium-rich slag are obtained. Through system integration of three technologies of hydrogen metallurgy reduction, electric melting furnace melting separation and wet vanadium extraction, efficient separation and all-component recovery of iron, vanadium and titanium are achieved, and the method has the outstanding advantages of being short in process, low in energy consumption, small in pollution, high in resource utilization rate and the like, and is particularly suitable for green and efficient utilization of low-grade vanadium titano-magnetite.

Description

TECHNICAL FIELD

[0001] The present application belongs to the technical field of metallurgical engineering and comprehensive utilization of resources, and particularly relates to an efficient and green smelting method of vanadium-titanium magnetite. In particular, the present application relates to a method for realizing efficient separation and full recovery of valuable components of iron, vanadium and titanium in low-grade vanadium-titanium magnetite, based on hydrogen metallurgy technology and combined with new electric smelting furnace melting separation and roasting-acid leaching process. BACKGROUND

[0002] Vanadium-titanium magnetite is an extremely important strategic resource, which is rich in iron, vanadium, titanium and other valuable elements. How to realize the efficient comprehensive utilization of these valuable components has been a major technical problem in the field of metallurgy. The traditional smelting process of vanadium-titanium magnetite mainly adopts the "blast furnace-converter" process: that is, vanadium-containing molten iron is obtained by blast furnace ironmaking, and then vanadium slag is obtained by converter blowing, and vanadium oxide is extracted by sodium roasting and water leaching, while titanium enters the blast furnace slag during the blast furnace smelting process, which is difficult to be economically recovered due to low grade and complex phase, resulting in a huge waste of titanium resources. The traditional process has problems such as high requirement for raw material grade, huge energy consumption, high carbon emission, extremely low titanium resource recovery rate (usually less than 50%), and serious environmental pollution.

[0003] In order to overcome the drawbacks of the traditional process, the industry has developed a new "pre-reduction-electric furnace melting separation" process. In this process, vanadium-titanium magnetite is first made into pellets or sintered ore, which is reduced by coal-based or gas-based reduction to obtain metallized pellets, and then high-temperature melting separation is carried out in an electric furnace to obtain vanadium-containing molten iron and titanium-containing melting slag. However, the "pre-reduction-electric furnace melting separation" process in the prior art still has many bottlenecks. (1) In the reduction process, the traditional carbon thermal reduction has a high temperature (usually > 1100℃), which will partially reduce vanadium, titanium and other oxides, and easily form low-melting eutectic phases with SiO2, Al2O3 and other impurities, resulting in small and dispersed metal iron particles, poor crystallization and polymerization, and subsequent melting separation difficulty, and incomplete slag-iron separation. In addition, a large amount of CO2 is generated during the carbon thermal reduction process, which does not meet the green and low-carbon development direction of metallurgy. (2) In the melting process, the conventional electric furnace (such as three-phase three-electrode electric arc furnace) has problems such as uneven thermal field, local overheating, large slag viscosity, poor electrical conductivity, easy foaming and spattering when dealing with high-titanium melting slag, resulting in high melting energy consumption, low efficiency and unstable operation, which is actually difficult to adapt to the melting of low-grade ore. And the composition of the obtained titanium slag fluctuates greatly, which is not conducive to the subsequent extraction of titanium. (3) The atmosphere of the electric furnace is difficult to control, resulting in vanadium in the slag, thereby causing low recovery rate of vanadium.

[0004] To solve the above problems, although there are relevant researches trying to improve. For example, patent CN120041686A discloses a method for removing impurities and extracting vanadium from calcified roasting vanadium slag leaching solution, which focuses on the optimization of impurity removal in the wet process, and does not solve the fundamental problem of front-end reduction and melting separation. Patent CN117737335A discloses a method for recovering iron and titanium from vanadium extraction slag by hydrogen-based reduction-electric furnace melting separation, which deals with "sodiumized vanadium extraction tailings", and its innovation lies in the dealkalization pretreatment and hydrogen-based reduction of vanadium extraction tailings to solve the problem of reduction pulverization caused by high alkali metal content in the tailings. However, its process is complex, additional steps such as acid leaching and dealkalization are required, and its final goal is to recover the residual iron and titanium in the tailings, not the whole process recovery of the raw ore. In addition, patent ZL202211424324.9 proposes a method for gradient reduction and separation to recover iron and vanadium-titanium resources in vanadium-titanium magnetite, which uses CH4-H2-N2mixed gas for medium-low temperature selective reduction to convert iron oxides into Fe3C, and then separates the iron phase from the vanadium-titanium-rich material by magnetic separation, and extracts vanadium titanium chloride by carbonitriding-low temperature chlorination. This technical route has its characteristics in realizing the separation of iron and vanadium-titanium, especially suitable for scenarios where Fe3C is the target product. However, this process still uses carbon-containing gas as the reducing medium, and cannot achieve zero carbon emission in the whole process; at the same time, it relies on carbonitriding-chlorination for subsequent vanadium-titanium extraction, the process is relatively long, and there is a risk of dispersion of vanadium elements in the iron phase and the slag phase, which is not conducive to the directional enrichment and efficient recovery of vanadium in the slag phase.

[0005] Therefore, in the prior art, there is still a lack of a mature solution that can start from low-grade vanadium-titanium magnetite raw ore, go through the whole process of "reduction-melting separation-vanadium extraction", and realize the recovery of all valuable components (Fe, V, Ti) in a low-carbon, efficient and clean manner. Developing a completely new and integrated technical route to completely break through the above technical bottlenecks is of great significance for realizing the high-value and green utilization of vanadium-titanium magnetite resources. SUMMARY

[0006] To solve the above technical problems existing in the prior art, the purpose of the present application is to provide a method for full recovery of valuable components in low-grade vanadium-titanium magnetite based on hydrogen metallurgy. The method integrates three core technologies of full hydrogen reduction, electric furnace melting separation, and roasting-acid leaching to form a complete process chain that is efficient, green and low-cost, and realizes efficient separation and full recovery of iron, vanadium and titanium.

[0007] The purpose of the present application is to disclose a method for full recovery of valuable components in low-grade vanadium-titanium magnetite based on hydrogen metallurgy. The valuable components include iron (Fe), vanadium (V) and titanium (Ti). The method comprises the following steps:

[0008] Pelletizing: Vanadium-titanium magnetite concentrate is used as raw material, bentonite is added at 1-2wt.% of the mass of the raw material, and calcium ferrite is added at 1%-5% of the mass of the raw material as a composite additive, the mixture is uniformly mixed and then balling is performed to obtain green balls; the green balls are dried at 100-110℃ for 4-6h, preheated at 900-1000℃ for 11-17min, and roasted at 1250-1350℃ for 9-15min to obtain vanadium-titanium magnetite oxidized pellets with a compressive strength of ≥2500N / pellet.

[0009] Full-hydrogen reduction: the pretreated vanadium-titanium magnetite oxidized pellets are subjected to a reduction reaction in a pure hydrogen (H2) atmosphere, and the reduction temperature is 750-900℃, to obtain metallized pellets with a metallization rate of 88%-94%.

[0010] Multi-electrode electric furnace smelting: the metallized pellets are subjected to a smelting process in a multi-electrode electric furnace, the slag basicity is controlled by spraying CaO flux during smelting, and the oxidation of the molten pool is adjusted by spraying carbon powder, to separate molten iron and vanadium-titanium-rich slag.

[0011] Roasting: the vanadium-titanium-rich slag is mixed with a calcium-magnesium-containing composite agent and subjected to roasting, the roasting temperature is 800-900℃, and the roasting time is 60-180min, to selectively convert vanadium into water-soluble vanadate.

[0012] Vanadium precipitation by acid leaching: the calcined fine powder is subjected to atmospheric acid leaching to separate vanadium from titanium, to obtain a vanadium-containing leaching solution and a titanium-rich slag. Vanadium pentoxide is precipitated from the vanadium-containing leaching solution to obtain a vanadium product.

[0013] The three core technologies of the present application promote and cooperate with each other to form a synergistic effect.

[0014] (1) Full-hydrogen reduction lays the foundation for electric furnace smelting: full-hydrogen reduction is carried out in a pure hydrogen atmosphere and at a relatively low temperature (750-900℃). From the reaction principle, hydrogen only reduces iron oxides (such as Fe2O3 and Fe3O4) to metallic iron, avoiding the reduction of vanadium-titanium oxides (such as V2O5 and TiO2), ensuring that vanadium and titanium remain in the form of high-valence oxides in the slag, and providing pellets with a moderate metallization rate (88%-94%) for the subsequent smelting process. If the metallization rate is too high, vanadium will be reduced into the molten iron, causing vanadium loss; if the metallization rate is too low, it will increase the energy consumption of smelting and cause iron loss. This step eliminates CO2 emissions from the source, meets the requirements of green metallurgy, and saves energy costs through low-temperature operation.

[0015] (2) The electric smelting furnace is used for smelting and separating in a top-down manner, and the slag system is optimized to create conditions for roasting: the electric smelting furnace with multiple electrodes adopts a six-electrode arrangement (double triangle or symmetric array) to provide a uniform and stable thermal field of the molten pool, and to overcome the problems of high viscosity of high-titanium slag and easy local overheating. The slag system basicity R (R = CaO / SiO2 mass ratio 0.8-1.2) is controlled by spraying CaO flux, and the oxidizing property is adjusted by spraying carbon powder to control the FeO content in the range of 9%-14%. From the reaction principle, the basicity adjustment improves the flowability of the slag and the separation efficiency of slag and iron, and the control of the FeO content ensures the moderate oxidizing property of the smelted slag, prevents the vanadium oxides from being excessively reduced, and makes the vanadium enriched in the slag. This step provides the vanadium-rich titanium slag with optimized composition for roasting: the low FeO content reduces the interference of iron on the vanadium conversion, and the appropriate basicity avoids the silicate phase in the slag from wrapping vanadium, thereby improving the subsequent conversion rate of vanadium. The efficient heating of the multiple-electrode system reduces the smelting time and the power consumption, thereby saving energy from the cost.

[0016] (3) Roasting-acid leaching realizes efficient separation of vanadium and titanium and closed-loop recovery: the roasting uses a composite agent (the molar ratio of CaO / MgO in the composite agent is 1:4-6) to selectively convert vanadium into water-soluble vanadate (such as Ca(VO3)2, Mg(VO3)2) at 800-900°C, while titanium remains in the form of stable oxides (such as TiO2). The introduction of the magnesium agent improves the mass transfer conditions of roasting, inhibits the dissolution of titanium, and improves the selectivity of vanadium conversion. Subsequently, the vanadium is efficiently leached (leaching rate > 95%) by atmospheric acid leaching (pH = 3-4, temperature 60-80°C), and high-grade titanium-rich slag (TiO2≥40%) is obtained. From the reaction principle, the roasting-acid leaching is closely related to the properties of the smelted slag at the front end: the low FeO and optimized basicity of the smelted slag reduce the consumption of additives in roasting, thereby reducing the cost; and the titanium-rich slag after acid leaching is stable in phase and can be directly used as a raw material for titanium dioxide or titanium extraction, thereby realizing the full recovery of titanium resources. This step is energy-saving and environmentally friendly, and reduces the burden of wastewater treatment.

[0017] In the full-hydrogen reduction step, the preferred reduction temperature is 750-900°C, and the reduction time is 60-120 min. This low-temperature hydrogen reduction process eliminates CO2 emissions from the source, and by precisely controlling the metallization rate (88%-94%), it avoids the excessive reduction of vanadium and titanium, and ensures that vanadium and titanium oxides remain in the slag, thereby laying a foundation for the subsequent smelting and vanadium extraction processes.

[0018] In the smelting step of the multiple-electrode electric smelting furnace, the electric smelting furnace is a six-electrode electric smelting furnace, and the electrodes are arranged in a double-triangle or symmetric array distribution. The multiple-electrode system can provide a uniform and stable thermal field of the molten pool, effectively solving the technical bottlenecks of local overheating, high viscosity of the slag, and unstable operation when processing high-titanium slag.

[0019] The smelting separation process in the multi-electrode electric smelting furnace is carried out at 1500-1650 DEG C, the slag basicity R (R=CaO / SiO2 mass ratio 0.8-1.2) is controlled by spraying CaO flux, and the FeO content in the smelting separation slag is controlled between 9%-14% by spraying carbon powder to regulate the oxidation of the molten pool. The control ensures the moderate reducibility of the slag, prevents vanadium oxides from being reduced into molten iron, and makes V2O5 preferentially enriched in the slag, thereby providing high-quality raw materials with high vanadium content and few impurities for subsequent roasting. The precise regulation of the slag property by spraying flux and carbon powder online improves the smelting dynamic conditions, promotes the separation of slag and iron, and reduces energy consumption.

[0020] In the roasting step, the fluxing agent is a composite agent of calcium oxide (CaO) and magnesium oxide (MgO), wherein the molar ratio of CaO to MgO is 1:(4-6), and the addition amount is 2%-5% of the mass of the vanadium-rich titanium slag. The introduction of the magnesium agent improves the mass transfer conditions of the roasting process, promotes the conversion of vanadium by forming a eutectic phase, significantly improves the conversion rate and selectivity of vanadium, and at the same time, maximally inhibits the dissolution of titanium, avoiding the loss of titanium resources.

[0021] In the acid leaching step, the roasted clinker is crushed to a particle size of less than 0.074 mm (-200 mesh) of more than 90%, and the normal pressure acid leaching is carried out under the conditions of pH value of 3-4, leaching temperature of 60-80 DEG C and leaching time of 30-60 min. The process ensures efficient leaching of vanadium (leaching rate greater than 95%) and efficient enrichment of titanium, is simple to operate, low in energy consumption, and easy to treat wastewater, and meets the requirements of green chemical industry.

[0022] The titanium-rich slag obtained in the acid leaching step has a TiO2 grade of not less than 40%. The titanium-rich slag is stable in phase, and titanium mainly exists in the form of rutile or anatase, which is suitable as raw material for titanium dioxide production by sulfuric acid method or chlorination method or as raw material for further purification, thereby realizing effective recovery and high-value utilization of titanium resources.

[0023] The vanadium obtained by the present application comes from pure vanadium-containing leaching liquor, and high-purity V2O5 or other vanadium products can be obtained after precipitation and calcination, the vanadium recovery rate is high, and the product quality is stable.

[0024] Compared with the prior art, the present application has at least the following beneficial effects: the present application forms a brand new technical route through the innovation and integration of three core technologies of full hydrogen reduction, multi-electrode electric furnace smelting, and roasting-acid leaching. The three technologies complement each other and optimize each other: full hydrogen reduction realizes green reduction from the source and provides ideal raw materials for smelting; multi-electrode electric furnace smelting provides optimized slag type for roasting by precisely controlling the slag system; and roasting-water leaching realizes efficient separation of vanadium and titanium and recycles all valuable components in a closed loop. This overall scheme successfully breaks through the multiple bottlenecks of traditional processes and existing "pre-reduction-electric furnace smelting" processes in terms of energy consumption, emissions, titanium recovery rate, and environmental pollution. The whole process has less CO2 emissions, low energy consumption, less reagent consumption, and significant cost savings, providing a feasible implementation scheme for the high-value and green comprehensive utilization of low-grade vanadium-titanium magnetite resources. BRIEF DESCRIPTION OF DRAWINGS

[0025] Figure 1 The process flow chart of the method for full recovery of valuable components in low-grade vanadium-titanium magnetite based on hydrogen metallurgy according to the present application. DETAILED DESCRIPTION

[0026] The technical solutions of the present application are illustrated by specific examples below. It should be understood that the one or more steps mentioned in the present application do not exclude the existence of other methods and steps before and after the described combination steps, or other methods and steps can be inserted between these explicitly mentioned steps. It should also be understood that these examples are only used to illustrate the present application and not to limit the scope of the present application. Unless otherwise specified, the numbering of each method step is only for the purpose of identifying each method step, and is not intended to limit the arrangement order of each method or to limit the scope of the implementation of the present application. Changes or adjustments of the relative relationship, without substantial technical content changes, can also be considered as the scope of implementation of the present application.

[0027] The following content is intended to further illustrate the standard process flow and core process parameters of the method according to the present application, and to verify the feasibility and effect of the method on the full recovery of iron, vanadium, and titanium valuable components in low-grade vanadium-titanium magnetite. The specific steps are as follows:

[0028] S1, raw material preparation and oxidized pellet preparation: take a low-grade vanadium-titanium magnetite concentrate, add 1-2 wt.% bentonite and 1-5 wt.% calcium ferrite as a binder, wherein the amount of bentonite added can be selected from 1 wt.%, 1.2 wt.%, 1.5 wt.%, 1.6 wt.%, 1.8 wt.% or 2 wt.%, and the amount of calcium ferrite added can be selected from 1 wt.%, 2 wt.%, 3 wt.%, 4 wt.% or 5 wt.%, mix uniformly, then ball by disc balling machine, generate green pellets, the moisture of the obtained green pellets is 8%-10%, the green ball drop strength is higher than 4 times per ball (0.5 m), and the green ball compression strength is higher than 12 N per ball. After drying the green pellets at 100-110℃ for 4-6h, the drying temperature and time can be selected according to the drying condition of the pellets, which can be 100℃ for 6h, 105℃ for 5h or 110℃ for 4h, preheat in the grate-kiln system at 900-1000℃ for 11-17min, the preheating temperature and time can be 900℃ for 17min, 950℃ for 14min or 1000℃ for 11min, and roast at 1250-1350℃ for 9-15min to obtain oxidized pellets, the roasting temperature and time can be 1250℃ for 15min, 1300℃ for 10min or 1350℃ for 9min. According to this step, oxidized pellets with a compression strength of ≥2500 N per ball and stable chemical composition can be obtained.

[0029] S2, total hydrogen reduction: load the above oxidized pellets into a shaft furnace hydrogen reduction reactor. Pure hydrogen (H2 purity ≥ 99.99%) is introduced into the reactor, and the reduction temperature is 750-900℃, and the reaction time is 60-120min. The specific reduction temperature can be selected from 750℃, 850℃ or 900℃, and the reduction time can be selected from 60min, 90min or 120min. According to this step, metallized pellets with a metallization rate of 88%-94% can be obtained, and their microstructure is loose and porous, which is beneficial to subsequent smelting.

[0030] S3, multi-electrode electric smelting furnace smelting: continuously add the metallized pellets into a closed electric smelting furnace with a six-electrode arrangement (electrodes adopt double-triangle distribution). The smelting temperature is controlled at 1500-1650℃, and the smelting temperature can be 1500℃, 1580℃ or 1650℃. CaO flux powder is sprayed into the molten pool through a spray gun, and the spraying amount is dynamically adjusted according to the online detection results of the molten pool slag sample to control the final slag basicity R, R = CaO / SiO2 mass ratio is 0.8-1.2, and the specific R can be selected from 0.8, 1.0 or 1.2. This step needs to keep the smelting process of the electric smelting furnace stable and maintain good slag and iron fluidity to achieve efficient separation. According to this, qualified molten iron for steelmaking ([C] 4.2%, [Si] 0.2%) and vanadium-titanium-rich slag are obtained.

[0031] S4, roasting: the vanadium-rich titanic slag is ground to a uniform powder with a particle size of 200 mesh or more than 90%. A composite agent is prepared according to the molar ratio CaO:MgO = 1:(4-6), and the addition amount is 2%-5% of the mass of the vanadium-rich titanic slag. After the slag powder is mixed with the composite agent, it is roasted in a rotary kiln at 800-900°C for 60-180 min. The roasting temperature can be selected as 800°C, 850°C or 900°C, and the roasting time can be selected as 60 min, 120 min or 180 min. In the roasting step, vanadium is mainly converted into water-soluble calcium vanadate (Ca(VO3)2) and magnesium vanadate (Mg(VO3)2), and titanium exists in the form of water-insoluble perovskite (CaTiO3).

[0032] S5, acid leaching of vanadium and recovery of titanium-rich slag: the roasted slag powder is subjected to atmospheric acid leaching at 60-80°C and pH 3-4, and mechanically stirred for 30-60 min. The acid leaching temperature can be selected as 60°C, 70°C or 80°C, and the pH value can be 3 or 4. After leaching, filtration is performed to obtain a vanadium-containing leaching solution and a leaching residue. After vanadium extraction in this step, the vanadium-containing leaching solution can be sent to the precipitation process, and a high-purity V2O5 product can be obtained by the ammonium salt vanadium precipitation-calcination process. The leaching residue is washed and dried to obtain a titanium-rich slag, which has a TiO2 grade that meets the quality requirements for raw materials for the production of titanium dioxide by the sulfuric acid method.

[0033] Example 1

[0034] This example completely implements a method for the full recovery of valuable components in low-grade vanadium-titanium magnetite based on hydrogen metallurgy, and the specific steps are as follows:

[0035] S1, pelletizing: a certain low-grade vanadium-titanium magnetite concentrate is used as the raw material, and 1.5wt.% of bentonite is added as a binder, and 3wt.% of calcium ferrite is added as a composite additive. After mixing, the green pellets are dried at 105°C for 5h, and then preheated at 950°C for 14 min in a grate-kiln system, and then roasted at 1300°C for 12 min to obtain oxidized pellets. The average compressive strength of the obtained oxidized pellets is 2950N / pellet. Through the optimization of the composite additive and the precise control of the heat treatment system, high-strength and composition-uniform oxidized pellets are obtained, which provide a good gas diffusion channel and reaction stability for the subsequent hydrogen reduction process, and physically ensure the smooth progress of the whole process.

[0036] S2, total hydrogen reduction: the above oxidized pellets are placed in a shaft hydrogen reduction reactor, and pure hydrogen gas with a purity of ≥99.99% is introduced. The reduction is carried out at 850°C for 90 min. After the reduction is completed, metallized pellets with a metallization rate of 91.5% are obtained. The pellet microstructure is loose and porous, which is beneficial to mass transfer and separation in the subsequent smelting process. This step realizes efficient reduction of iron oxides under pure hydrogen and low temperature conditions, while effectively inhibiting the reduction of vanadium and titanium oxides, ensuring that vanadium and titanium are enriched in the slag phase in the form of oxides. Precise control of the metallization rate provides an ideal raw material composition for the subsequent smelting process, avoiding the loss of vanadium and reducing the energy consumption of smelting.

[0037] S3, smelting: the metallized pellets are sent to a closed electric smelting furnace with a six-electrode arrangement (double triangular array) and smelted at 1580°C for 30 min. During the smelting process, the final slag basicity R(CaO / SiO2 mass ratio) is controlled to be 1.0 by spraying CaO flux, and the FeO content in the smelting slag is adjusted to 11.5% by spraying carbon powder. Finally, molten iron (composition: [C] 4.2%, [Si] 0.2%) and vanadium-titanium-rich slag (composition: TiO2 42.5%, V2O5 1.8%) are successfully separated. This step realizes efficient separation of slag and iron under high-titanium smelting conditions through uniform heating by a multi-electrode system and precise control of the slag system. The obtained vanadium-titanium-rich slag has stable composition and moderate FeO content, providing an optimized composition and controllable impurities for subsequent roasting-acid leaching of vanadium, laying the foundation for efficient separation of vanadium and titanium.

[0038] S4, roasting: the vanadium-titanium-rich slag is ground to a particle size of -200 mesh, and a composite agent (CaO and MgO molar ratio of 1:5) is added at a dosage of 2% of the slag mass. The mixture is roasted in a rotary kiln at 850°C for 120 min. Vanadium is selectively converted into water-soluble vanadates (such as Ca(VO3)2, Mg(VO3)2), while titanium remains in the form of stable perovskite (CaTiO3) and rutile (TiO2). This step realizes high selectivity of vanadium conversion at moderate temperatures through the synergistic effect of CaO-MgO composite agents, while effectively inhibiting the dissolution of titanium. This process creates favorable conditions for subsequent atmospheric acid leaching to achieve efficient separation of vanadium and titanium, taking into account reaction efficiency and process economy.

[0039] S5, acid leaching separation: the calcined clinker fine powder is subjected to atmospheric pressure acid leaching under the condition of 70°C and pH value of 4, and mechanical stirring for 60 min. After leaching, filtration is performed to obtain a vanadium-containing leaching solution and a leaching residue. The vanadium leaching rate of this step reaches 96.8%, and the titanium dissolution rate is less than 1.5%. The leaching residue is washed and dried to obtain a titanium-rich residue, and the TiO2 grade of the titanium-rich residue is increased to ≥40%. This step realizes efficient leaching of vanadium and effective enrichment of titanium under mild conditions, and has high vanadium leaching rate and low titanium loss rate, which reflects the good matching of the roasting-acid leaching process. The obtained titanium-rich residue meets the requirements of subsequent utilization of titanium resources, and realizes the full recovery of titanium components.

[0040] S6, vanadium precipitation: the vanadium-containing leaching solution is treated by using an ammonium salt precipitation-calcination process, and after precipitation, filtration and calcination, a vanadium pentoxide (V2O5) product with purity of >98% is obtained, realizing efficient recovery of vanadium. This step successfully prepares a high-purity vanadium product from a pure vanadium-containing leaching solution at the front end, verifies the effective control and closed-loop recycling ability of the whole process for vanadium element, and reflects the technical advantages of the present application in the high-value recycling of strategic metals.

[0041] This embodiment successfully constructs a complete, efficient and green valuable component full recovery process chain of low-grade vanadium-titanium magnetite by system integration and parameter optimization of the six steps of "balling-full hydrogen reduction-electric furnace smelting-roasting-acid leaching-vanadium precipitation". The steps are closely connected, and the parameters are optimized in coordination, so that the directional separation and high-value recovery of iron, vanadium and titanium are gradually realized. The whole process eliminates carbon emissions at the source, realizes the double control of energy consumption and reagent consumption in the process, and achieves resource full recovery and pollution minimization at the end, fully reflecting the overall innovativeness, technical feasibility and economic environmental synergy of the present application.

[0042] Example 2

[0043] This embodiment has the same process steps as Example 1, and the parameters of S2-S6 in Example 1 are fixed, only the parameters of the S1 oxidation pellet preparation step are adjusted to observe the influence of S1 step on the compressive strength, and the pellet performance is optimized through experiments to provide a basis for the subsequent reduction and smelting process. The strength of the oxidized pellet directly affects the gas diffusion efficiency of the reduction process and the slag-iron separation effect of the smelting process, so the optimization of the pellet preparation parameters is crucial. The effects of calcium ferrite addition amount, preheating temperature, preheating time, calcination temperature and calcination time on the compressive strength are investigated. All experiments are based on the standard process parameters in Example 1.

[0044] In Example 2, the baseline parameters are calcium ferrite addition amount of 3 wt.%, preheating temperature of 950°C, preheating time of 14 min, calcination temperature of 1300°C, and calcination time of 12 min. Under this baseline, the average compressive strength of the oxidized pellet is 2950 N.

[0045] Table 1 Oxidized pellet preparation process parameters

[0046]

[0047] In a specific embodiment, when the calcium ferrite addition amount is 1 wt.%, the pellet preheating temperature is 900°C, the preheating time is 11 min, the roasting temperature is 1300, and the roasting time is 9 min, the average compressive strength of the obtained oxidized pellets is 2600N. Too little calcium ferrite and low roasting temperature are not conducive to the formation of a liquid phase, and shorter preheating and roasting times result in incomplete oxidation of magnetite, with the pellets mainly connected by metallic bonds inside, leading to relatively low pellet strength.

[0048] In a specific embodiment, when the calcium ferrite addition amount is 3 wt.%, the pellet preheating temperature is 950°C, the preheating time is 14 min, the roasting temperature is 1300, and the roasting time is 12 min, the average compressive strength of the obtained oxidized pellets is 2950N. Due to the appropriate amount of calcium ferrite addition, higher preheating and roasting temperatures, and the extension of the reaction time, the liquid phase calcium ferrite inside the pellets diffuses sufficiently, forming a strong slag phase bond between the particles, greatly enhancing the compressive strength of the pellets. At the same time, as the degree of oxidation of magnetite increases, atomic diffusion energy is strengthened, and metallic bond connections are also more rapid.

[0049] In a specific embodiment 2, when the calcium ferrite addition amount is 5 wt.%, the pellet preheating temperature is 1000°C, the preheating time is 17 min, the roasting temperature is 1300, and the roasting time is 15 min, the average compressive strength of the obtained oxidized pellets is 2790N. Too much calcium ferrite liquid phase forces small pores inside the pellets to merge into large voids, on the one hand leading to uneven stress and negatively affecting strength, and on the other hand also worsening the H2 diffusion kinetics in the subsequent reduction process, reducing the reduction efficiency. In addition, too high a temperature and too long a reaction time result in overfiring of the pellet surface and the formation of a hard shell, hindering further oxidation and the entry of subsequent reduction gas, and also worsening the metallurgical properties of the pellets.

[0050] Example 3

[0051] The process steps of this example are the same as those of Example 1, with the parameters of S1, S3-S6 in Example 1 fixed, and only the key process parameters (reduction temperature, reduction time) in the S2 full-hydrogen reduction step adjusted to observe the effect of S2 full-hydrogen reduction on the metallization rate of the metallized pellets and verify the feasibility and controllability of achieving a metallization rate in the range of 88%-94% under a pure hydrogen atmosphere. The metallization rate is a key indicator that affects the subsequent melting and separation effect and the movement of valuable elements, and its precise control is crucial for the optimization of the entire process. This example is based on the standard process in Example 1, and the effects of reduction temperature and reduction time on the metallization rate are systematically investigated.

[0052] In Example 3, the reference parameters are a reducing atmosphere of pure hydrogen (H2 purity ≥ 99.99%), a reducing temperature set at 850°C, and a reducing time of 90 min. Under this reference condition, the average metallization rate of the obtained metallized pellets is 91.5%, and the pellet microstructure is loose and porous, which is beneficial for subsequent smelting.

[0053] Table 2: Full-hydrogen reduction process parameters

[0054]

[0055] In a specific embodiment, under the condition of a reducing time of 90 min, when the reducing temperature is 750°C, the average metallization rate of the obtained metallized pellets is 88.2%. Due to the lower temperature, the hydrogen reduction kinetics is limited, and part of Fe3O4 fails to be completely reduced to metallic iron, resulting in a lower metallization rate, but still meeting the basic requirement of the subsequent smelting on the metallization rate of the raw material. When the reducing temperature is 850°C, the metallization rate increases to 91.5%. At this temperature, the hydrogen reduction reaction rate is moderate, Fe2O3 and Fe3O4 can be fully reduced to metallic iron, while excessive reduction of vanadium and titanium oxides is avoided, the pellet microstructure is uniform, and the pore development is good, which is the ideal reduction state. When the reducing temperature is further increased to 900°C, the metallization rate reaches 93.8%. High temperature significantly accelerates the reduction reaction, but at the same time, there is a risk of partial reduction of some vanadium oxides, and if the metallization rate exceeds 94%, vanadium elements may enter the iron phase, causing a decrease in vanadium recovery rate.

[0056] In a specific embodiment, under the condition of a reducing temperature of 850°C, when the reducing time is 60 min, the metallization rate is 89.0%. The reaction time is short, and part of the pellet core is not completely reduced, resulting in a slightly lower metallization rate. When the reducing time is extended to 90 min, the metallization rate stabilizes at 91.5%. At this time, the reduction degree of the pellets inside and outside is balanced, and vanadium and titanium oxides are basically not reduced, meeting the requirements of the subsequent smelting and vanadium extraction of the raw material. When the reducing time is further extended to 120 min, the metallization rate increases to 94.1%. Although the metallization rate increases, the long reduction time may cause slight sintering of the pellet surface layer, a decrease in porosity, which is not conducive to the diffusion of hydrogen and subsequent smelting, and there is a risk of vanadium loss.

[0057] In Example 3, by systematically adjusting the reducing temperature and time, it is verified that the precise control of the metallization rate in the range of 88%-94% can be achieved under a pure hydrogen atmosphere. It is clear how to ensure efficient reduction of iron while avoiding excessive reduction of vanadium and titanium, which provides metallized pellet raw materials with ideal composition and suitable structure for subsequent multi-electrode electric smelting furnace smelting and roasting-acid leaching processes.

[0058] In Example 3, in addition to temperature and time, hydrogen flow rate is an important parameter affecting the gas-solid reaction interface renewal and reduction kinetics. Under the benchmark conditions, when the hydrogen flow rate is low, the H2O generated by the reaction is difficult to remove from the surface of the pellets in time, which hinders the diffusion of hydrogen into the pellet and inhibits the forward progress of the reduction reaction, which may result in insufficient reduction of the core of the pellet, low metalization rate and unevenness inside and outside. When the hydrogen flow rate is moderate, an effective reducing atmosphere can be formed and the product water vapor can be removed in time, ensuring uniform reduction of the pellet from the surface to the core and achieving the ideal metalization rate. If the hydrogen flow rate is too high, although the reduction can be accelerated, the marginal effect of the reaction rate is decreasing, and at the same time, the gas consumption and operating cost will be significantly increased, which is not optimal from the economic point of view. Therefore, in actual operation, the hydrogen flow rate needs to be controlled within a reasonable range that can ensure the reduction efficiency and economic efficiency, according to the type of reactor and the amount of charge.

[0059] In Example 3, the particle size of the oxidized pellets determines the length of the hydrogen diffusion path, which is a key physical factor affecting the uniformity of reduction and the final metalization rate. Under the same reduction conditions, smaller pellets have shorter internal diffusion paths and larger specific surface areas, which are beneficial to the rapid penetration and reaction of hydrogen, thereby achieving a higher and more uniform metalization rate. However, too small a particle size will worsen the permeability of the material layer and may cause pulverization during the reduction process, affecting the stability of the operation. On the contrary, larger pellets improve the permeability of the material layer, but significantly increase the resistance of hydrogen diffusion to the core of the pellet, which is prone to form a "metallic iron shell-unreacted core" structure, resulting in a low overall metalization rate and internal gradient. Therefore, controlling the particle size of the oxidized pellets within a moderate and concentrated range through the balling and screening process is an important prerequisite for ensuring the efficient and uniform performance of the subsequent hydrogen reduction process.

[0060] Example 4

[0061] This example has the same process steps as Example 1, with the parameters of S1-S2, S4-S6 in Example 1 fixed, only the key process parameters (melting and separating temperature, slag basicity, FeO content in the melting and separating slag) in the S3 multi-electrode electric furnace melting and separating step are adjusted to observe the effect of S3 step on the composition of molten iron and vanadium-titanium-rich slag. The melting and separating effect is directly related to the quality of molten iron and the recovery rate of valuable elements vanadium and titanium, and its precise control is crucial for the optimization of the whole process. Based on the standard process in Example 1, this example systematically investigates the effects of melting and separating temperature, slag basicity and FeO content on the carbon content in molten iron and the TiO2 and V2O5 grades in vanadium-titanium-rich slag.

[0062] In Example 4, the reference parameters are: a separation temperature of 1580 °C, a slag basicity (CaO / SiO2 mass ratio) of 1.0, and an FeO content of 11.5% in the separation slag. Under these reference conditions, the carbon content in the obtained hot metal is 4.2%, the TiO2 grade in the vanadium-titanium-rich slag is 42.5%, and the V2O5 content is 1.8%. In this example, the metallized pellets prepared under the reference conditions of Example 3 are used as the separation raw material to ensure consistent input conditions. During the separation process, the slag basicity is controlled by spraying CaO flux, and the oxidation of the molten bath is adjusted by spraying carbon powder to maintain the FeO content within the target range.

[0063] Table 3. Separation process parameters of the multi-electrode electric smelting furnace

[0064]

[0065] In a specific example, when the separation temperature is 1500 °C under the condition of a slag basicity of 1.0, the carbon content in the obtained hot metal is 4.0%, the TiO2 grade in the vanadium-titanium-rich slag is 41.0%, and the V2O5 content is 1.6%. Due to the lower temperature, the separation dynamics are slow, and the separation of slag and iron is not complete, resulting in a slightly lower carbon content in the hot metal and a decrease in the enrichment of vanadium and titanium in the slag. When the separation temperature is increased to 1580 °C, the carbon content in the hot metal reaches 4.2%, the TiO2 grade is increased to 42.5%, and the V2O5 content is 1.8%. At this temperature, the separation reaction is complete, and the slag has good fluidity, achieving efficient separation. When the separation temperature is further increased to 1650 °C, the carbon content in the hot metal increases to 4.4%, but the V2O5 content in the vanadium-titanium-rich slag decreases to 1.7%, and the TiO2 grade is 43.0%. Although high temperature promotes the separation of slag and iron and the enrichment of titanium, it may cause excessive reduction of some vanadium oxides, slightly increasing the risk of vanadium loss.

[0066] In a specific example, when the separation temperature is 1580 °C, and the slag basicity is 0.8, the carbon content in the obtained hot metal is 4.1%, the TiO2 grade in the vanadium-titanium-rich slag is 41.5%, and the V2O5 content is 1.7%. When the basicity is lower, the slag viscosity is higher, and the fluidity is poor, affecting the separation efficiency of slag and iron, resulting in a decrease in the enrichment of titanium and vanadium. When the slag basicity is adjusted to 1.0, the carbon content in the hot metal stabilizes at 4.2%, the TiO2 grade increases to 42.5%, and the V2O5 content is 1.8%. An appropriate basicity significantly improves the fluidity and reaction kinetics of the slag, promoting the enrichment of valuable components. When the slag basicity is further increased to 1.2, the carbon content in the hot metal remains at 4.2%, the TiO2 grade reaches 43.0%, and the V2O5 content is 1.9%. The high basicity slag system has better fluidity, but attention should be paid to avoid excessive basicity leading to an increase in the melting point of the slag system and an increase in energy consumption.

[0067] In Example 4, the controllability of the multi-electrode electric smelting furnace smelting process on the composition of molten iron and the grade of vanadium-titanium-rich slag was verified by systemically regulating the tapping temperature and the basicity of the slag. It was clarified how to optimize the vanadium-titanium enrichment effect while ensuring efficient separation of slag and iron, providing a stable composition and low impurity content of vanadium-titanium-rich slag for the subsequent roasting-acid leaching process. In addition, the uniform thermal field and precise injection system of the multi-electrode electric smelting furnace provide a basic guarantee for the parameter regulation of this experiment, ensuring the stability and reproducibility of the smelting process.

[0068] In Example 4, in addition to the above parameters, the metallization rate of the metallized pellets is also a key factor affecting the smelting effect. Under the baseline smelting conditions, when the metallization rate is low (such as 88%), there are more unreduced iron oxides in the pellets, and more energy is consumed for reduction during the smelting process, which may lead to fluctuations in the carbon content of molten iron and an increase in FeO in the slag; when the metallization rate is high (such as 94%), although the smelting energy consumption is reduced, there is a risk of excessive reduction of vanadium into molten iron. Therefore, controlling the metallization rate in the ideal range of 88%-94% is an important prerequisite to ensure the economy of smelting and the recovery rate of valuable elements.

[0069] In Example 4, the rate and uniformity of carbon powder injection are crucial to the regulation of FeO content. If the injection is insufficient, the oxidation of the molten pool will be too high, which will lead to a low carbon content of molten iron and insufficient vanadium enrichment; if the injection is excessive, it may cause the molten pool to boil and the slag and iron to emulsify, worsening the separation effect. By optimizing the design of the injection gun and the control system, precise dynamic regulation of the FeO content can be achieved, further improving the smelting efficiency.

[0070] Example 5

[0071] This example has the same process steps as Example 1, with the parameters of S1-S3, S5-S6 in Example 1 fixed, only the CaO / MgO molar ratio of the composite agent in the roasting step S4 is adjusted to observe the effect of the CaO / MgO molar ratio of the composite agent on the vanadium conversion selectivity and leaching effect. The conversion rate of vanadium and the dissolution rate of titanium are directly related to the efficiency of vanadium recovery and the enrichment of titanium resources, and their optimization control is crucial to the resource utilization rate of the whole process. Based on the standard process in Example 1, this example systematically investigates the effect of the CaO / MgO molar ratio on the vanadium leaching rate and titanium dissolution rate, and verifies the synergistic effect of the composite agent in selective vanadium extraction and inhibition of titanium dissolution.

[0072] In Example 5, the reference parameters are: roasting temperature 850°C, roasting time 120 min, total composite agent addition amount 2% of the mass of vanadium-rich titanium slag, and vanadium-rich titanium slag ground to more than 90% of -200 mesh. Under this reference condition, the system changes the molar ratio of CaO to MgO in the composite agent to conduct experiments. The same batch of vanadium-rich titanium slag obtained from the melting and separation step in Example 1 is used as the raw material in this embodiment. After roasting, the vanadium leaching rate and titanium dissolution rate are evaluated under standard acid leaching conditions.

[0073] In a specific embodiment, when pure CaO is used as the roasting additive, the conversion rate and leaching rate of vanadium are 89.5% and 87.2%, respectively, but the dissolution rate of titanium is as high as 8.5%. This is because the perovskite (CaTiO3) structure formed by the pure calcium agent during roasting is unstable, and part of the titanium is dissolved in ionic form during acid leaching, resulting in loss of titanium resources, and the conversion selectivity of vanadium is low, which is not conducive to the enrichment of titanium and subsequent utilization.

[0074] In a specific embodiment, when the molar ratio of CaO / MgO is 1:4, the conversion rate of vanadium is increased to 92.8%, the leaching rate reaches 90.5%, and the dissolution rate of titanium is reduced to 4.2%. The introduction of magnesium agent improves the mass transfer conditions of the roasting system, promotes the selective conversion of vanadium by forming a eutectic phase (such as CaO-MgO-V2O5), and at the same time, the magnesium ion is solid-solved in the slag phase to enhance the stability of titanium and inhibit the dissolution of titanium.

[0075] In a specific embodiment, when the molar ratio of CaO / MgO is 1:5, the conversion rate of vanadium is 96.5%, the leaching rate is as high as 96.8%, and the dissolution rate of titanium is further reduced to less than 1.5%. At this ratio, CaO and MgO form the best composite reaction interface, significantly improving the generation efficiency of vanadates (such as Ca(VO3)2 and Mg(VO3)2), and at the same time, titanium mainly remains in the form of stable rutile (TiO2) and perovskite (CaTiO3), achieving high selectivity extraction of vanadium and efficient enrichment of titanium.

[0076] In a specific embodiment, when the molar ratio of CaO / MgO is 1:6, the conversion rate of vanadium is 94.2%, the leaching rate is 93.5%, and the dissolution rate of titanium is 2.1%. Although further increasing the proportion of magnesium agent helps to inhibit the dissolution of titanium, too high a content of MgO may lead to an increase in the melting point of the slag system and a decrease in the kinetics of vanadium conversion, which is not conducive to the complete conversion and leaching of vanadium.

[0077] In Example 5, the key influence of composite agent on vanadium-titanium separation behavior during roasting was verified by systemically regulating CaO / MgO molar ratio. It was explicitly pointed out that controlling CaO / MgO molar ratio could optimize vanadium conversion and leaching effect, while reducing titanium dissolution rate, providing a key process optimization window for efficient vanadium-titanium separation and full recovery. In addition, the introduction of magnesium agent not only improved the thermodynamic and kinetic conditions of roasting reaction, but also reduced the reagent consumption and wastewater treatment load in the subsequent acid leaching process, meeting the requirements of green chemical industry.

[0078] In Example 5, the addition amount of composite agent directly affects the alkalinity and vanadate generation kinetics during roasting. When the addition amount is too low (such as less than 1%), the conversion of vanadium oxide in the slag is insufficient, the vanadium conversion rate decreases significantly, and the titanium dissolution rate increases, mainly due to insufficient alkalinity, leading to selective conversion of vanadium. When the addition amount increases to 2%, the vanadium conversion rate increases rapidly, and the titanium dissolution rate is effectively inhibited, indicating that this addition amount can optimize the alkalinity of the slag system and promote the selective conversion of vanadium without causing excessive reaction of titanium. Further increasing the addition amount can slightly improve the vanadium conversion rate, but will increase the viscosity of the slag system and energy consumption, and may introduce impurities, affecting the economic efficiency of subsequent acid leaching. Therefore, the addition amount of composite agent should be controlled within the range of 2%-5% to achieve efficient recovery of vanadium and balance of process economy.

[0079] In Example 5, roasting temperature and time are also important factors affecting vanadium conversion rate. Under the benchmark conditions, too low temperature or too short time will result in incomplete conversion of vanadium; too high temperature or too long time may cause sintering of slag phase and transformation of titanium phase, increasing the risk of titanium dissolution. Therefore, by controlling the roasting temperature within the range of 800-850℃ and the time within the range of 60-120min, and combining with the composite agent ratio of CaO / MgO molar ratio 1:4-1:6, the optimal balance of vanadium-titanium separation effect and full-process economy can be ensured.

[0080] In Example 5, the particle size of vanadium-rich titanium slag is also a key physical factor affecting the uniformity of roasting reaction and the conversion efficiency of vanadium. When the particle size is too coarse (such as less than 85% of-200 mesh), the reaction interface is insufficient, and the vanadium conversion rate decreases; when the particle size is too fine (such as more than 95% of-200 mesh), it is easy to cause material sticking and mass transfer obstruction during roasting. By controlling the slag powder particle size to be more than 90% of-200 mesh, the reaction efficiency and operation stability can be considered, providing a guarantee for the smooth implementation of roasting-acid leaching process.

Claims

1. A method for the complete recovery of valuable components from low-grade vanadium-titanium magnetite based on hydrogen metallurgy, characterized in that, The method includes the following steps: S1. Pelletizing: Using vanadium-titanium magnetite concentrate as raw material, bentonite is added at 1-2 wt.% of the raw material mass, and calcium ferrite is added at 1%-5% of the raw material mass as a composite additive. After mixing evenly, pellets are formed to obtain green pellets. The green pellets are dried at 100-110℃ for 4-6 hours, preheated at 900-1000℃ for 11-17 minutes, and calcined at 1250-1350℃ for 9-15 minutes to obtain oxide pellets with a compressive strength ≥2500N / pellet. S2, Full hydrogen reduction: The oxidized pellets are placed in a pure hydrogen atmosphere with a volume concentration of ≥99% and reduced at 750-900℃ for 60-120 min to obtain metallized pellets with a metallization rate of 88%-94%. S3, Melting and Separation: The metallized pellets are placed in a multi-electrode electric melting furnace and melted at 1500-1650℃ for 20-40 minutes. During the melting and separation process, the final slag basicity R is controlled between 0.8 and 1.2 by injecting CaO into the molten pool, and the FeO content in the melted slag is controlled between 9% and 14% by injecting carbon powder, so as to separate molten iron and vanadium-titanium-rich slag. S4. Calcination: The vanadium-rich titanium slag is mixed with a composite agent containing CaO and MgO. The amount of the composite agent added is 2-5% of the mass of the vanadium-rich titanium slag. The molar ratio of CaO to MgO in the composite agent is 1:(4-6). The mixture is calcined at 800-900℃ for 60-180 min to selectively convert the vanadium component in the slag into water-soluble vanadate. S5. Acid leaching separation: The clinker after roasting in step S4 is crushed to a particle size of less than -200 mesh (more than 90%). It is then subjected to normal pressure acid leaching under the conditions of pH 3-4, temperature 60-80℃, and time 30-60min. After solid-liquid separation, vanadium-containing leachate and titanium-rich slag with TiO2 content of not less than 40% are obtained. S6. Vanadium precipitation: Vanadium pentoxide is precipitated from the vanadium-containing leachate to obtain the vanadium product.

2. The method as described in claim 1, characterized in that, In step S1, the pelleting time is 12 minutes, and the moisture content of the raw pellets is 8%-10%.

3. The method as described in claim 1, characterized in that, The electric furnace mentioned in step S3 is a six-electrode electric furnace, with the electrodes arranged in a double triangle or symmetrical array.

4. The method as described in claim 1, characterized in that, In step S5, the vanadium leaching rate is greater than 95%.

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