A short-process high-quality utilization method for vanadium-titanium magnetite resources

Abstract

The present invention belongs to the field of recovery and utilization of metal mineral resources, and particularly relates to a method for short-process high-quality utilization of vanadium-titanium magnetite resources. The original vanadium-titanium magnetite ore is co-enriched to obtain an iron-vanadium-titanium co-selected concentrate; the iron-vanadium-titanium co-selected concentrate, a solid carbonaceous reducing agent, a binder, and a magnesium-based regulator are mixed and pelletized or compacted to prepare a carbon-containing pellet; the carbon-containing pellet is pre-reduced in a hydrogen-containing atmosphere to obtain a pre-reduced pellet; the pre-reduced pellet is smelted and separated to obtain hot metal and vanadium-titanium-rich slag. The present invention co-enriches and co-smelts ilmenite and vanadium-titanium magnetite in the original ore, avoiding problems such as over-fine grinding in the original iron / titanium separation process, use of toxic flotation reagents, and resource waste. Through the optimization of the new slag system without calcium by adjusting MgO in the concentrate in the beneficiation process and controlling FeO in the pre-reduction process, the controllable separation of iron and vanadium-titanium is achieved, which is conducive to the further selective extraction and recovery of vanadium and titanium therein, realizing the short-process high-quality utilization of vanadium-titanium magnetite resources.

Description

Technical Field

[0001] The present invention belongs to the field of mineral utilization, and particularly relates to the technical field of the development and utilization of vanadium-titanium magnetite resources. Background Art

[0002] Vanadium-titanium magnetite is a composite ore in which multiple metal elements such as iron, vanadium, titanium, chromium, nickel, and cobalt coexist, and has extremely high comprehensive utilization value and is an important mineral resource recognized internationally.

[0003] The main idea of the current development and utilization process of vanadium-titanium magnetite resources is to separate vanadium-titanium magnetite into iron concentrate and titanium concentrate, and then conduct subsequent smelting on the iron concentrate and titanium concentrate respectively. Its typical steps are as follows: First, the original vanadium-titanium magnetite ore is weakly magnetically separated to obtain iron concentrate (containing TFe ~55%, V2O5 ~0.5%, TiO2 ~10%), and the tailings after iron separation are flotated to obtain titanium concentrate (containing TFe ~35%, V2O5 ~0.1%, TiO2 ~48%). The iron concentrate is smelted in a blast furnace, and the iron element is recovered in the form of hot metal for steelmaking. The vanadium element enters the hot metal during smelting and then V2O5 is recovered through processes such as blowing vanadium slag in a converter and sodiumizing vanadium extraction. The titanium element (accounting for about 52% of the total titanium resources) is transformed into stable perovskite and enters the blast furnace slag, and this part of the titanium cannot be utilized at all. At the same time, some coexisting nickel and cobalt rare elements also enter the blast furnace slag and cannot be utilized. The obtained titanium concentrate is used for the production of titanium white by the sulfuric acid method, and the iron and vanadium elements in the titanium concentrate cannot be recovered. The current vanadium-titanium magnetite resources adopt a route of separating iron and titanium and using them separately, resulting in the titanium, nickel, and cobalt elements in the iron concentrate entering the blast furnace slag and being unable to be recovered, the vanadium extraction process being cumbersome and having a low recovery rate, the iron and vanadium in the titanium concentrate not being recovered, and finally the utilization rates of iron, vanadium, and titanium being only 79%, 44%, and 29% respectively. In addition, Chinese patent document with publication number CN115522074A discloses a slag-making method for electric furnace smelting of vanadium-titanium magnetite, which includes the following steps: S1. Pre-reducing the oxidized pellets of vanadium-titanium magnetite concentrate to obtain metallized pellets of vanadium-titanium magnetite; S2. Adding the metallized pellets of vanadium-titanium magnetite, calcareous slag-making agent, and reducing agent into the electric furnace in proportion for deep reduction; S3. Conducting smelting after deep reduction and separating the slag and iron to obtain hot metal containing vanadium and molten fraction slag containing titanium; S4. Conducting converter vanadium extraction treatment on the hot metal containing vanadium, and conducting chemical method titanium extraction treatment on the molten fraction slag containing titanium. This method has a complex process, and vanadium enters the hot metal, and there are relatively large metal process losses in the treatment process.

[0004] In summary, the existing processes for the recovery of vanadium-titanium magnetite metal resources are relatively complex, and the process losses of valuable metal elements are relatively large. Summary of the Invention

[0005] Aiming at the problems of long process flow and low utilization rate of valuable metals in the existing process for utilizing vanadium-titanium magnetite, the object of the present invention is to provide a method for short-process and high-quality utilization of vanadium-titanium magnetite resources, aiming to provide a brand-new smelting method with a short process, high selectivity and high recovery rate, which does not require separation of iron, titanium and vanadium, and directly performs reduction-smelting separation to selectively separate iron and vanadium-titanium.

[0006] In the vanadium-titanium magnetite raw ore, there are mainly two types of minerals: titanomagnetite and ilmenite, and the minerals are closely intergrown and have fine crystal grains. However, the reaction behaviors of titanomagnetite and ilmenite are quite different, making it difficult to co-smelt. The reaction behaviors of iron, vanadium and titanium elements during the co-smelting process are difficult to control. For example, vanadium among them easily enters the molten iron, resulting in difficult control of the selectivity of element separation and being unfavorable for the selective smelting separation of metals. Therefore, the existing idea for resource utilization of vanadium-titanium magnetite raw ore is usually to separate the raw ore into titanomagnetite and ilmenite, and then smelt them separately. This conventional idea will significantly increase the process flow and also increase the process losses of metals. For example, the final utilization rates of iron, vanadium and titanium in the existing technology are only 79%, 44% and 29% respectively. In view of this technical status, the present invention breaks away from the existing conventional idea of pre-separating the raw ore into iron concentrate-ilmenite concentrate, and for the first time provides a short-process treatment idea of directly co-enriching and co-smelting titanomagnetite and ilmenite in the vanadium-titanium magnetite raw ore. However, to implement this short-process idea, it is necessary to face problems such as difficult control of the reaction behaviors of iron, vanadium and titanium elements, difficult control of the vanadium-titanium slag phase, and unsatisfactory separation selectivity of iron, vanadium and titanium elements. In view of the problems faced by the brand-new co-enrichment-co-smelting idea of the present invention, through in-depth research, the present invention provides the following improvement solutions:

[0007] A method for short-process and high-quality utilization of vanadium-titanium magnetite resources, the steps include:

[0008] (a) Crushing, grinding and countercurrent gravity separation pre-enrichment of the vanadium-titanium magnetite raw ore; the gravity separation rough concentrate obtained by pre-enrichment is further subjected to reverse flotation to remove aluminum and silicon minerals, and an iron-vanadium-titanium co-selection concentrate enriched with iron, titanium and vanadium minerals (ilmenite and vanadium-titanium magnetite) in the raw ore is obtained;

[0009] (b) Mixing the iron-vanadium-titanium co-selection concentrate, solid carbonaceous reducing agent, binder and magnesium-based regulator to obtain a mixed material, and pelletizing or briquetting the mixed material to prepare a carbon-containing pellet;

[0010] The carbon-containing pellet is pre-reduced in a hydrogen-containing atmosphere in advance to obtain a pre-reduced pellet with a metallization degree of 85-98%, an FeO content of 1-15% and an MgO content of 0.5-10%;

[0011] (c) Smelting and separating the pre-reduced pellet to obtain molten iron and vanadium-titanium-rich slag.

[0012] The present invention breaks away from the conventional iron-titanium separation idea in existing raw ores and innovatively provides a short-process recovery idea of co-enrichment (countercurrent gravity separation - reverse flotation) - co-smelting (magnesium oxide-assisted internal carbon gas-solid pre-reduction - smelting separation) of raw ores. Firstly, it throws away the tailings of the raw ore by countercurrent gravity separation and obtains the co-selected concentrate (co-enriched minerals) of ilmenite and vanadium-titanium magnetite in the raw ore by reverse flotation. Then, it performs pre-carbon addition and hydrogen-assisted pre-reduction, strictly controls the degree of iron metallization, and the contents of FeO and MgO during the pre-reduction process, and conducts subsequent smelting separation treatment on the pre-reduced minerals with special physical and chemical characteristics. Furthermore, it regulates the reaction behaviors of iron, vanadium, and titanium components and the element trends, enabling iron to be highly selectively enriched in the hot metal and vanadium and titanium to be highly selectively enriched in the vanadium-titanium slag, thereby realizing the selective smelting separation of iron and vanadium-titanium and helping to regulate the slag phase, and further improving the selective recovery effect of vanadium and titanium in the vanadium-titanium slag. The method of the present invention provides a brand-new co-enrichment idea of special countercurrent gravity separation - reverse flotation for desiliconization and de-aluminum of vanadium-titanium magnetite, and conducts special pre-reduction - smelting separation co-smelting treatment on the co-selected concentrate of iron, vanadium, and titanium with such special co-enrichment. In this way, not only is the treatment process greatly shortened, and the use of blast furnaces and harmful arsenic agents in the prior art is avoided, but also the total recovery rate of vanadium-titanium magnetite can be significantly improved. For example, the final utilization rates of vanadium and titanium can be increased from the existing 44% and 29% to over 80% respectively.

[0013] The vanadium-titanium magnetite raw ore described above includes ilmenite, magnetite, and titanomagnetite, and is disseminated with at least one of feldspar, olivine, pyroxene, serpentine, chlorite, and amphibole; among them, the occurrence phases of Fe include at least one of magnetite, ilmenite, ferromagnesium olivine, and pyrite, and the occurrence phases of Ti include titanomagnetite and ilmenite. In addition, in the above-mentioned raw ore, at least one valuable element such as Ni, Co, and scandium is also allowed to exist, and it can coexist with pyrite and pyrrhotite.

[0014] In the present invention, the vanadium-titanium magnetite raw ore contains 20 - 30 wt.% of Fe, 0.1 - 0.5 wt.% of V2O5, and 5 - 8 wt.% of TiO2. In the present invention, 0.01 - 0.1 wt.% of Ni and 0.01 - 0.1 wt.% of Co are also allowed to exist in the raw ore.

[0015] Preferably, after being crushed and ground, the particle size of the vanadium-titanium magnetite raw ore should be controlled such that the content of -0.15 mm is 70 - 95%.

[0016] In the present invention, the raw ore is subjected to simple rough grinding treatment and subsequent countercurrent gravity separation enrichment to obtain the co-selected concentrate of vanadium, titanium, and iron. The combination of the countercurrent gravity separation process in this step helps to make full use of the physical and chemical characteristics of the raw ore, thus facilitating the combination with subsequent processes, helping the selective smelting separation of iron - vanadium-titanium, and facilitating the optimization of the physical and chemical activity of the subsequent vanadium-titanium slag, which is conducive to the recovery of vanadium and titanium.

[0017] In the present invention, the countercurrent re-election step in step (a) is as follows: countercurrent inclined-plane re-election column is used for re-election, wherein the ore feeding concentration is 18-22%, the ore feeding speed is controlled at 100-150 L / min, and the water flow speed is controlled at 14-18 L / min.

[0018] Preferably, the number of times of countercurrent re-election is 1-3 times.

[0019] In the present invention, desilication and dealumination treatment can be carried out by a conventional reverse flotation process.

[0020] The reagents for the reverse flotation include a reverse flotation inhibitor and a reverse flotation collector.

[0021] The reverse flotation inhibitor includes one or more of starch, water glass, carboxymethyl cellulose, sodium hexametaphosphate, sodium silicate, and sodium fluorosilicate.

[0022] The reverse flotation collector is one or more of oleic acid, linoleic acid, cetyl quaternary ammonium salt, tetradecyl quaternary ammonium salt, and sodium dimethylaminodecyl carboxylate.

[0023] In the reverse flotation for desilication and dealumination, a certain amount of frother and activator can also be added according to needs.

[0024] In the present invention, the activator is a conventional component capable of achieving reverse flotation of aluminum and silicon, for example, it can be calcium chloride.

[0025] The process flow of the reverse flotation for desilication and dealumination can include roughing and scavenging treatment in 1-3 stages.

[0026] In the present invention, when the raw ore further contains sulfide ores of at least one metal among nickel, cobalt, and scandium, according to the desulfurization requirement, the countercurrent re-election concentrate can be pre-treated for desulfurization, and then subsequent reverse flotation for desilication and dealumination treatment can be carried out.

[0027] In the present invention, according to needs, the tailings of the countercurrent re-election can be subjected to magnetic separation treatment, and the minerals obtained by magnetic separation can be recycled to the reverse flotation process and / or the pre-reduction process.

[0028] In the present invention, through the co-enrichment process of the countercurrent re-election - reverse flotation, an iron-vanadium-titanium co-election concentrate co-enriched with ilmenite - vanadium-titanium magnetite can be obtained. The weight ratio of ilmenite to vanadium-titanium magnetite in the iron-vanadium-titanium co-election concentrate is 1:5 - 3:4.

[0029] Preferably, in the iron-vanadium-titanium co-election concentrate, the content of TFe is 40 wt.% - 50 wt.%, the content of TiO2 is 13 wt.% - 18 wt.%, and the content of V2O5 is 0.4 wt.% - 1.0 wt.%.

[0030] Preferably, in the iron-vanadium-titanium co-concentrate, 0.4 wt.% to 8 wt.% of MgO is also allowed to be included.

[0031] In the present invention, for the co-enriched iron-vanadium-titanium co-concentrate, an internal carbon-injected gas-solid pre-reduction treatment assisted by magnesium oxide is carried out, in combination with the joint control of parameters such as the metallization rate, FeO, and MgO content of the pre-reduced product. In this way, it is unexpectedly beneficial to the reaction behavior of iron, vanadium, and titanium and the regulation of the enrichment direction in the subsequent reduction process. Thus, it is beneficial to achieve the highly selective separation of hot metal and vanadium-titanium slag. In addition, it is also beneficial to the joint control of the slag phase and physical and chemical characteristics of vanadium-titanium slag, and further beneficial to the selective separation of vanadium and titanium in the subsequent vanadium-titanium slag.

[0032] In the present invention, in step (b), the solid carbonaceous reducing agent includes one or a mixture of several of anthracite, bituminous coal, lignite, biomass, and biochar.

[0033] Preferably, the solid carbonaceous reducing agent is biomass, and its fixed carbon content FC d is 15 wt.% to 20 wt.%, and the volatile matter content V daf ≥80 wt.%.

[0034] Preferably, the binder is a mixture of one or several of a composite binder, modified starch, and cellulose. As a preference, the binder is a coal-based colloidal composite binder.

[0035] Preferably, the weight ratio of the iron-vanadium-titanium co-concentrate, the solid carbonaceous reducing agent, and the binder is 100:(1 - 10):(0.5 - 3); further, it can be 100:3 - 6:1 - 1.5.

[0036] Preferably, in the mixed material, a magnesium-based regulator is added as needed. The additive includes at least one of magnesium oxides, hydroxides, carbonates, chlorides, nitrates, and organic acid salts. For example, in the present invention, the weight ratio of the iron-vanadium-titanium co-concentrate to the magnesium-based additive can be 100:1 - 3.

[0037] In the present invention, in the pre-reduction stage, the volume fraction of H2 in the hydrogen-containing atmosphere is greater than or equal to 15%; further, it can be hydrogen.

[0038] Aiming at the problems such as the difficulty in controlling the iron-vanadium-titanium selectivity and the vanadium-titanium slag properties faced by the combined smelting of vanadium-titanium iron co-concentrate, the present invention innovatively conducts a combined pre-reduction of the pre-reduced pellet carbon and hydrogen atmosphere, further in combination with the joint control of temperature, the metallization rate of the pre-reduced pellet, FeO, and MgO. In this way, it is unexpectedly possible to achieve synergy, which is beneficial to the subsequent selective smelting of iron and vanadium-titanium. In addition, it is also beneficial to regulating the activity of vanadium-titanium slag and is beneficial to the recovery of vanadium and titanium.

[0039] Preferably, the pre-reduction temperature is 900 °C to 1200 °C, and the pre-reduction time is 10 to 90 minutes. Further preferably, the pre-reduction temperature is 950 to 1050 °C, and the pre-reduction time is 15 to 60 minutes; further 30 to 60 minutes. Under the preferred pre-reduction process, a better short-process and high-efficiency recovery effect can be obtained.

[0040] In the present invention, the pre-reduced pellets with the metallization rate, FeO, and MgO contents are innovatively subjected to subsequent reduction smelting treatment. In this way, with the synergistic assistance of the FeO / MgO contents, the reduction and distribution behaviors of Fe, vanadium, and titanium in the subsequent reduction melting and separation can be optimized. In addition, it is also beneficial to regulate and obtain a vanadium-titanium active slag phase that is conducive to the extraction and separation of vanadium and titanium.

[0041] Preferably, the metallization rate of the pre-reduced pellets is 90 to 97%, the FeO content therein is 1 to 15%, and the MgO content is 0.5 to 10%. The total content of FeO and MgO does not exceed 20 wt.%. Further preferably, the FeO content of the pre-reduced pellets is 3 wt.% to 10 wt.%, and the MgO content is 2 wt.% to 8.5 wt.%.

[0042] In the present invention, the smelting separation temperature is 1300 °C to 1600 °C, and the smelting time is 0.5 h to 4 h.

[0043] In the present invention, the vanadium-rich titanium slag is crushed and finely ground and then subjected to acid leaching treatment to obtain a vanadium acid leaching solution and titanium slag; vanadium is recovered from the vanadium acid leaching solution, and titanium therein is recovered from the titanium slag.

[0044] In the present invention, the particle size of the vanadium-rich titanium slag after crushing and fine grinding should be controlled such that the content of -0.074 mm is ≥90%.

[0045] The acid solution for the acid leaching is hydrochloric acid, and its concentration is preferably 10% to 25%, the mass ratio of the acid to the slag is 2:1 to 6:1, the leaching temperature is 30 to 160 °C, and the time is 0.5 to 5 h.

[0046] Preferably, vanadium is extracted and separated from the vanadium acid leaching solution to prepare a vanadium pentoxide product.

[0047] In the present invention, the TiO2 grade of the titanium slag is ≥50%;

[0048] Preferably, titanium white is prepared from the titanium slag.

[0049] Beneficial effects:

[0050] (1) Process innovation (short process for co-enrichment and co-smelting of iron, vanadium, and titanium)

[0051] Co-enrichment:

[0052] The present invention innovates the beneficiation process of iron / titanium separation, with iron separated first and then titanium. Through coarse grinding - gravity separation - reverse flotation for desilication and dealumination, iron, titanium, and vanadium in the raw ore are synchronously enriched and recovered in their intrinsic substances, obtaining an iron-vanadium-titanium co-beneficiation concentrate with a TFe content of 40wt.% - 50wt.%, a TiO2 content of 13wt.% - 18wt.%, and a V2O5 content of 0.4wt.% - 1.0wt.%. The recovery rates of Fe, V, and Ti are greatly improved. It avoids the loss of titanium in the iron concentrate and iron in the titanium concentrate in the current beneficiation process of separate iron / titanium separation. Moreover, for iron / titanium separation, an extremely fine grinding particle size is required. On the one hand, the grinding energy consumption is high. On the other hand, the subsequent fine-grained titanium concentrate is difficult to recover and has a large loss, and the use of arsenic acid-based toxic agents in the flotation process increases the environmental protection pressure. Therefore, while shortening the beneficiation process of vanadium-titanium magnetite, the present invention reduces the grinding energy consumption, eliminates the use of toxic agents, and improves the recovery rates of iron, titanium, and vanadium elements.

[0053] Co-smelting:

[0054] The present invention also innovates the extraction and smelting process of vanadium-titanium iron concentrate. It innovatively realizes the separation of iron from vanadium and titanium through a short process of pre-reduction - smelting separation. Vanadium is enriched in the smelting separation slag (instead of hot metal), simplifying the vanadium extraction process, reducing energy consumption while increasing the recovery rate of vanadium. The titanium slag after wet vanadium extraction can be used as a raw material for titanium white production, significantly improving the utilization rate of titanium resources.

[0055] Based on the above-mentioned co-enrichment - co-smelting short process, the present invention can effectively reduce the process flow and significantly enhance the final recovery rates of iron, vanadium, and titanium. For example, the final utilization rates of vanadium and titanium in the process of the present invention can be respectively increased from the existing 44% and 29% to over 80%.

[0056] (2)High recovery efficiency of iron-vanadium-titanium co-enrichment beneficiation

[0057] The existing beneficiation process of vanadium-titanium magnetite is relatively complex and requires multi-stage grinding and multi-stage separation to obtain iron concentrate (iron grade 11% - 12%) and titanium concentrate (TiO2 grade 45%). During the recovery process, the TiO2 loss rate is relatively high. The present invention dissociates the useful minerals ilmenite and magnetite in vanadium-titanium magnetite from minerals with relatively coarse dissemination sizes such as feldspar, olivine, and pyroxene through crushing and grinding, and feeds them into a countercurrent inclined plane gravity separation column for gravity separation operations. The separation is mainly carried out by utilizing the specific gravity differences of mineral particles, realizing the co-enrichment of iron-vanadium-titanium elements. In the countercurrent water flow against gravity upward, the associated mineral particles of ilmenite, feldspar, olivine, and pyroxene particles have a relatively small specific gravity and form an overflow with the water flow, while the ilmenite and magnetite particles have a relatively large specific gravity and settle downward against the countercurrent. Therefore, the recovery rates of TFe, TiO2, and V2O5 in the underflow obtained by gravity separation all exceed 90%. Compared with ordinary gravity separation technology, the recovery rate is greatly improved.

[0058] (3) High resource utilization rate of pre-reduction - smelting separation of titanium and vanadium

[0059] Traditional blast furnace smelting of vanadium-titanium iron concentrate requires that the TiO2 content is not more than 13%, and the titanium slag type is not ideal, making it difficult to recycle; in addition, vanadium enters the molten iron during blast furnace smelting, resulting in a complex and lengthy subsequent separation and vanadium extraction process with low recovery rate.

[0060] The present invention abandons traditional blast furnace smelting and proposes a short process of pre-reduction - electric furnace melting and separation for the characteristics of the mixed concentrate composition. By synergistically regulating FeO-MgO in the pre-reduced pellets to construct a new slag system for smelting, it is beneficial for the selective melting and separation of iron and (vanadium-titanium) during the melting and separation stage. In addition, it can optimize the phase and active form of vanadium-titanium slag, create conditions for the subsequent utilization of titanium in the slag, and improve the utilization rate of titanium resources; at the same time, by controlling the separation of magnesium-containing minerals in the raw ore through the co-enrichment beneficiation system, the co-selected concentrate has an appropriate MgO content range, which is beneficial for the slag system regulation and slag-iron separation in the subsequent smelting process, can reduce / avoid the addition of extra magnesia flux, and achieve the beneficial effect of beneficiation - metallurgy coordination. During the smelting process, by controlling the reduction potential and the FeO content in the slag phase, vanadium is enriched in the slag rather than in the molten iron, and subsequent wet separation and extraction can achieve the efficient preparation of vanadium products and improve the utilization rate of vanadium resources. The comprehensive recovery rates of iron, vanadium, and titanium in the entire beneficiation and metallurgy process all exceed 70%, showing a significant increase in the comprehensive recovery rate compared with the current process.

[0061] (4) Reduction of the difficulty of pellet pre-reduction

[0062] In the ilmenite, magnetite and titanomagnetite in the iron-vanadium-titanium co-selected concentrate described in the present invention, the mineral phases of the raw ore are maintained and the complex dissemination relationship exists, and the reduction difficulty is greater than that of the conventional sorted iron ore pellets. However, aiming at the problem of difficult pre-reduction faced by the short-process technology of the present invention, the present invention innovatively adopts the method of gasification reduction with internal solid carbonaceous reducing agent and external gas-based reduction to prepare pre-reduced pellets for electric furnace smelting separation. On the one hand, through the cooperation with external gas-based reduction, the dosage of the solid carbonaceous reducing agent can be reduced, and defects such as low strength and easy pulverization of the carbon-containing pellets can be effectively avoided. And with the consumption of the solid carbonaceous reducing agent, a large number of pores can be left inside the pellets, which can improve the reduction kinetic conditions and provide space for the consumption of the reduction expansion stress, effectively alleviating the phenomenon of pellet pulverization due to the reduction expansion stress. On the other hand, reduction occurs simultaneously inside and outside the pellets, and the reduction kinetic conditions are superior, which can avoid the stress concentration phenomenon caused by uneven reduction rates inside and outside the pellets. The electric furnace smelting separation of the present invention requires that the pre-reduced pellets contain a certain amount of FeO, so as to reduce the viscosity of the slag and improve the slag-iron separation conditions. On this basis, the requirement for the metallization degree of the pre-reduced pellets can be appropriately relaxed, and the reduction difficulty of the pellets is greatly reduced. In addition, magnesia additives beneficial to smelting separation can effectively improve the reduction expansion and reduction pulverization properties of vanadium-titanium magnetite pellets during the reduction stage.

[0063] (5)Recycling of Ni / Co rare elements

[0064] In the existing vanadium-titanium magnetite beneficiation and smelting process, Ni / Co elements finally enter the blast furnace slag and cannot be recycled. In the preferred process of the present invention, by adding a flotation operation process at the source, Ni / Co sulfide minerals are enriched and recovered, and Ni / Co products are produced through extraction and smelting. If scandium is also contained in the iron-vanadium-titanium-nickel-cobalt rough concentrate, through the said technology, scandium and nickel-cobalt can also be enriched and recovered together. The process of the present invention can realize the recycling of elements such as nickel and cobalt, and has no adverse impact on the subsequent beneficiation and smelting of iron-vanadium-titanium minerals. Description of the drawings

[0065] Figure 1 XRD pattern of the vanadium-titanium-rich slag in Example 1;

[0066] Figure 2 Process flow diagram of vanadium-titanium magnetite in Example 7. Detailed implementation manners

[0067] The following examples are intended to further illustrate the present invention rather than limit the present invention.

[0068] An optional method for high-quality utilization of vanadium-titanium magnetite resources in a short process according to the present invention aims to solve the problem of low recovery rates of components such as titanium, vanadium, nickel, and cobalt in the existing beneficiation and smelting processes. It mainly includes the following steps:

[0069] (1) The vanadium-titanium magnetite ore is crushed and ground, and then subjected to countercurrent inclined flow gravity separation to obtain a pre-enriched coarse concentrate. The coarse concentrate is directly subjected to reverse flotation to obtain an iron-vanadium-titanium co-selected concentrate, or it is subjected to desulfurization flotation to recover valuable elements such as nickel and cobalt, and then subjected to reverse flotation to obtain an iron-vanadium-titanium co-selected concentrate; in addition, the tailings after gravity pre-enrichment are then subjected to magnetic separation to recover the iron minerals and merged into the co-selected concentrate;

[0070] (2) the iron-vanadium-titanium co-selected concentrate obtained in step (1) is mixed with a solid carbonaceous reducing agent, a binder, and a magnesium-based regulator in a mass ratio of 100:(1-10):(0.5-3):(1-3) to obtain a mixed material, and the mixed material is pelletized or pressed into carbon-containing pellets with a diameter of 10-20 mm;

[0071] (3) Using hydrogen-rich reducing gas to reduce carbon-containing pellets to obtain pre-reduced pellets with a certain degree of metallization;

[0072] (4) feeding the pre-reduced pellets from step (3) into an electric furnace for smelting and separation to obtain molten iron and vanadium-titanium-rich slag, wherein the molten iron can be used for converter steelmaking;

[0073] (5) crushing and finely grinding the vanadium-rich titanium slag obtained in step (4) and then subjecting it to acid leaching-extraction separation to extract vanadium to prepare a vanadium pentoxide product;

[0074] (6) The titanium-rich material remaining after acid leaching in step (5) can be used for titanium dioxide production.

[0075] In the above method, the particle size of the vanadium-titanium magnetite ore after crushing and grinding in step (1) should be controlled to be -0.15mm and the content should be ≥95%.

[0076] In the above method, the gravity separation feed concentration in step (1) is 18-22%, the feed rate is controlled at 100-150 L / min, and the water flow rate is controlled at 14-18 L / min, and an iron-vanadium-titanium co-selected coarse concentrate with a TFe content of 40 wt.%-50 wt.%, a TiO2 content of 13 wt.%-18 wt.%, a V2O5 content of 0.4 wt.%-1.0 wt.%, and a MgO content of 0.4-8 wt.% can be obtained.

[0077] In step 1, the magnetic separation intensity is 0.2~0.4T.

[0078] In desulfurization flotation-dealuminization and siliconization reverse flotation, the pulp concentration is controlled at 20%~40% and pH=6~11.

[0079] In the present invention, the flotation reagents used in the desulfurization flotation process include a desulfurization regulator and a desulfurization collector.

[0080] The desulfurization regulator includes at least one of sulfuric acid, hydrochloric acid, sodium hydroxide, and lime.

[0081] The desulfurization collector is at least one of xanthate collectors, dithiophosphate collectors, and sulfonamide collectors; preferably, it includes at least one of ethyl xanthate, butyl xanthate, isopropyl xanthate, isobutyl xanthate, amyl xanthate, hexyl xanthate, ammonium butyl dithiophosphate, and ethyl thionocarbamate.

[0082] The desulfurization flotation process includes roughing and two-stage scavenging. Among them, the addition amount of the desulfurization regulator in the roughing stage is 400 - 800 g / t, the addition amount in the first-stage scavenging is 200 - 400 g / t, and the addition amount in the second-stage scavenging is 0 - 200 g / t.

[0083] The addition amount of the desulfurization collector in roughing is 80 - 200 g / t, the addition amount in the first-stage scavenging is 40 - 80 g / t, and the addition amount in the second-stage scavenging is 0 - 40 g / t.

[0084] In the present invention, the reverse flotation for removing silicon and aluminum can remove minerals such as Al and Si in the gravity separation concentrate in the form of flotation foam. In addition, it can also optimize the composition of components such as Mg in the minerals of the reverse flotation tailings.

[0085] Preferably, the reagents for the reverse flotation for removing aluminum and silicon (also known as reverse flotation for removing aluminum and silicon) include a reverse flotation inhibitor and a reverse flotation collector.

[0086] The reverse flotation inhibitor includes one or more of starch, water glass, carboxymethyl cellulose, sodium hexametaphosphate, sodium silicate, and sodium fluorosilicate.

[0087] Preferably, the reverse flotation collector is one or more of oleic acid, linoleic acid, cetyl quaternary ammonium salt, tetradecyl quaternary ammonium salt, and sodium dimethylaminodecyl carboxylate.

[0088] The reverse flotation for removing aluminum and silicon includes roughing and two-stage scavenging. Among them, the dosage of the inhibitor in roughing is 250 - 2000 g / t (further can be 750 - 850 g / t), the dosage of the inhibitor in the first-stage scavenging is 50 - 600 g / t (further can be 450 - 500 g / t), and the dosage of the inhibitor in the second-stage scavenging is 50 - 400 g / t (further can be 250 - 300 g / t).

[0089] The dosage of the collector in roughing of the reverse flotation for removing aluminum and silicon is 250 - 2000 g / t (further can be 900 - 1100 g / t), the dosage of the collector in the first-stage scavenging is 50 - 600 g / t (further can be 450 - 500 g / t), and the dosage of the collector in the second-stage scavenging is 50 - 400 g / t (further can be 250 - 300 g / t).

[0090] In the reagents for reverse flotation desilication and dealumination, a foaming agent is also allowed to be added. Among them, the dosage of the foaming agent in the roughing stage can be 40 - 120 g / t (further can be 70 - 90 g / t). The dosage of the foaming agent in the first scavenging is 20 - 60 g / t (further can be 30 - 50 g / t), and the dosage of the foaming agent in the second scavenging is 10 - 30 g / t (further can be 15 - 25 g / t).

[0091] In the reagents for reverse flotation dealumination and desilication, an activator is also allowed to be added, which can be calcium chloride for example.

[0092] Among them, the dosage of the activator in the roughing stage can be 400 - 600 g / t (further can be 450 - 550 g / t). The dosage of the activator in the first scavenging is 100 - 300 g / t (further can be 150 - 250 g / t), and the dosage of the activator in the second scavenging is 50 - 150 g / t (further can be 80 - 120 g / t).

[0093] In the above method, in step (2), the solid carbonaceous reducing agent is one or a mixture of several of anthracite, bituminous coal, lignite, biomass, and biomass charcoal. Preferably, the solid carbonaceous reducing agent is biomass, and the fixed carbon content FC of this biomass d is 15 wt.% - 20 wt.%, and the volatile matter content V daf ≥80 wt.%.

[0094] The additive is a mixture of one or more of dolomite and its processed products such as calcined dolomite, magnesite and other common magnesian fluxes.

[0095] The mixed concentrate with co - enrichment of iron, vanadium and titanium obtained in step (2), the solid carbonaceous reducing agent, the binder, and the magnesium - based additive are proportioned and mixed according to the mass ratio of 100:(2.5 - 3.5):(1 - 1.5):(1.25 - 1.75) to obtain a mixture.

[0096] The co - selected concentrate of iron, vanadium and titanium obtained, the solid carbonaceous reducing agent, the binder, and the magnesium - based additive are proportioned and mixed according to the mass ratio of 100:3:1:1.5 to obtain a mixture.

[0097] The binder is a mixture of one or several of a composite binder, modified starch, and cellulose. Preferably, the binder is a coal - based colloidal composite binder, and its preparation and composition are recorded in the patent document with the publication number CN117305581A.

[0098] The purpose of adding magnesian flux in the present invention is to replace calcareous flux to reduce the slag viscosity: when the reduction degree of the pre-reduced pellets is relatively high, the FeO content of the slag separated by subsequent smelting is relatively low. Increasing the magnesium oxide in the slag can play a role in reducing the slag viscosity, increasing the iron recovery rate, and changing the occurrence state of titanium in the slag.

[0099] The volume fraction of H2 in the reducing gas is greater than or equal to 15%. As a further preference, the reducing gas is a mixture of one or two of H2 and CO, and the volume fraction of H2 is greater than or equal to 15%. Further preferably, H2 is used to pre-reduce the carbon-containing pellets.

[0100] The pre-reduction temperature is 900°C to 1200°C, and the pre-reduction time is 10 to 90 minutes. Further preferably, the pre-reduction temperature is 950 to 1050°C, and the pre-reduction time is 15 to 60 minutes. If the pre-reduction temperature is too high, high-valent iron oxides (including trivalent iron and divalent iron) will be overly reduced to metallic iron. If the FeO content in the pre-reduced pellets is too low, it cannot play a conditioning role in the subsequent process. If the pre-reduction temperature is too low, the metallization rate of the pre-reduced pellets will be too low, and the FeO content in the slag during smelting separation will be too high, resulting in a decrease in the iron recovery rate.

[0101] The FeO content of the pre-reduced pellets is 1wt.% to 15wt.%, and the MgO content is 0.5wt.% to 10wt.%. Further preferably, the FeO content of the pre-reduced pellets is 3wt.% to 10wt.%, and the MgO content is 2wt.% to 8.5wt.%.

[0102] The smelting separation temperature is 1300°C to 1600°C, preferably 1350°C to 1450°C, and the reduction time is 0.5 h to 4 h.

[0103] Through the above settings, the temperature at which the slag viscosity reaches 0.5 Pa·s can be reduced to 1350°C to 1450°C. The TFe content of the obtained iron sample is ≥98wt.%, and the TiO2 content in the vanadium-rich titanium material is ≥40wt.%, further above 50wt.%.

[0104] The particle size of the vanadium-rich titanium slag should be controlled such that the content of -0.074mm is ≥90% after crushing and fine grinding. The initial hydrochloric acid concentration for acid leaching is 10% to 25%, the acid-to-slag mass ratio is 2:1 to 6:1, the leaching temperature is 30 to 160 °C, and the time is 0.5 to 5 h.

[0105] The TiO2 grade of the titanium-rich material is ≥50%, and it can be directly used for titanium white production.

[0106] In the above method, the re-election equipment used is a countercurrent inclined plane flow re-election column. For example, as an alternative option, in this case, the gravity separation device disclosed in the Chinese patent document with the publication number CN117414938A is selected.

[0107] In the following case, the hydrogen gas selected is a commercial hydrogen-rich atmosphere. For example, the content of hydrogen gas is above 95%.

[0108] In the present invention, ilmenite and vanadium-titanium magnetite in the original ore are co-enriched in the iron-vanadium-titanium co-selected concentrate. It is different from the conventional separation process, and the weight ratio of ilmenite to vanadium-titanium magnetite can be 1:2 - 3.

[0109] Example 1

[0110] In this example, the vanadium-titanium magnetite original ore from a certain concentrator in Baima, Panzhihua is selected, and its chemical composition analysis is shown in Table 1.

[0111]

[0112] This implementation case specifically includes the following steps:

[0113] (1) Feed the vanadium-titanium magnetite original ore into the first-stage crushing operation for crushing and grinding, and control the fineness of the grinding product so that the content of -0.15mm ≥ 70wt.%.

[0114] (2) Feed the grinding product in step (1) into the first-stage TARC countercurrent inclined plane re-election for separation. Control the feed concentration at 25%, the feed rate at 120L / min, and the water flow rate at 16L / min. The underflow obtained by separation is used as the first-stage TARC re-election rough concentrate, and the overflow is used as tailings and directly discarded.

[0115] Feed the first-stage TARC re-election iron rough concentrate into the second-stage TARC re-election operation. Control the feed concentration at 25%, the feed rate at 120L / min, and the water flow rate at 16L / min. The underflow obtained by separation is used as the final iron-vanadium-titanium co-enriched mixed concentrate, and the overflow obtained by separation is used as tailings and directly discarded. Combine the tailings from the first-stage TARC re-election and the second-stage TARC re-election as tailings. The indexes of the mixed concentrate (re-election concentrate) and tailings obtained according to the above steps are shown in Table 2.

[0116]

[0117] (3) Perform reverse flotation desilication and dealumination on the re-election concentrate in step (2), which includes one roughing and two scavenging steps. The steps are as follows: perform rough flotation treatment on the re-election concentrate to obtain roughing tailings; perform the first scavenging treatment on the roughing tailings, and then perform the second scavenging on the tailings of the first scavenging. The collected tailings are the iron-vanadium-titanium co-selected concentrate;

[0118] The flotation reagents for rough selection, first scavenging, and second scavenging include a foaming agent, a collector, an activator, and an inhibitor. Among them, the foaming agent is No. 2 oil; the inhibitor is starch; the activator is calcium chloride; the collector is oleic acid;

[0119] During the rough selection process, the dosage of the foaming agent is 80 g / t; the dosage of the inhibitor is 800 g / t; the dosage of the activator is 500 g / t; the dosage of the collector is 1000 g / t;

[0120] During the first scavenging process, the dosage of the foaming agent is 40 g / t; the dosage of the inhibitor is 500 g / t; the dosage of the activator is 200 g / t; the dosage of the collector is 500 g / t;

[0121] During the second scavenging process, the dosage of the foaming agent is 20 g / t; the dosage of the inhibitor is 300 g / t; the dosage of the activator is 100 g / t; the dosage of the collector is 300 g / t.

[0122] (4) Charge the iron-vanadium-titanium co-selected concentrate from step (3), biomass (wood chips), a composite binder (Mad = 8.49%, Ad = 28.73%, Vdaf = 60.86%), and magnesite in a mass ratio of 100:5:1:1.5, mix them evenly to obtain a mixture, and prepare the mixture into carbon-containing pellets with a diameter of 10 mm to 14 mm through a pelletizing process.

[0123] (5) Reduce the carbon-containing pellets from step (4) at a reduction temperature of 1050 °C for 30 min under a H2 atmosphere to obtain pre-reduced pellets with a metallization rate of 94.43 wt.%, an FeO content of 5.63 wt.%, and an MgO content of 4.42 wt.%.

[0124] (6) Mix the pre-reduced pellets from step (5) with 0.3% of their weight of wood chips and reduce and melt-separate them at 1450 °C for 1 h to obtain hot metal and vanadium-titanium-rich material. The obtained iron sample has a TFe of 99.8 wt.% and an iron recovery rate of 88.73%. The vanadium-titanium-rich material (XRD see Figure 1 ) has a TiO2 grade of over 50 wt.%, a titanium recovery rate of 95 wt.%, a V2O5 grade of 1.3 wt.%, and a vanadium recovery rate of 85 wt.%.

[0125] (7) After the vanadium-titanium-rich slag is crushed and finely ground, the particle size should be controlled such that the content of -0.074 mm is ≥ 90%, and then it is subjected to acid leaching treatment. The initial hydrochloric acid concentration for the acid leaching is 20%, the acid-to-slag mass ratio is 3:1, the leaching temperature is 60 °C, and the time is 2 h. Subsequently, solid-liquid separation is carried out to obtain a vanadium leaching solution and vanadium leaching slag, and the vanadium leaching rate is 86.8%.

[0126] The 95% concentrated sulfuric acid acidolysis rate of the vanadium leaching slag (titanium slag) is 94.0%, which can be used as a raw material for the production of titanium white by the sulfuric acid method.

[0127] Example 2

[0128] Compared with Example 1, the difference is only that the steps of step (4) to step (6) are changed, specifically:

[0129] (4) Charge the iron-vanadium-titanium co-selected concentrate, biomass (wood chips), composite binder (moisture Mad = 8.49%, ash Ad = 28.73%, volatile matter Vdaf = 60.86%), and magnesite in a mass ratio of 100:3:1:1.5, mix them evenly to obtain a mixture, and prepare the mixture into carbon-containing pellets with a diameter of 10 mm to 14 mm through a pelletizing process.

[0130] (5) Reduce the carbon-containing pellets in step (4) under an H2 atmosphere at a reduction temperature of 1050°C for 15 min to obtain pre-reduced pellets with a metallization rate of 87.87 wt.%, an FeO content of 11.22 wt.%, and an MgO content of 3.09 wt.%.

[0131] (6) Charge the pre-reduced pellets in step (5) and the reducing agent (wood chips) in a mass ratio of 100:0.5, and carry out reduction melting separation at 1450°C for 1 h to obtain hot metal and vanadium-titanium-rich material. The obtained iron sample has a TFe of 99.8 wt.%, an iron recovery rate of 80.91%, and the TiO2 grade in the vanadium-titanium-rich material is above 50 wt.%.

[0132] The titanium recovery rate is 97.3 wt.%, the V2O5 grade is 1.2 wt.%, and the vanadium recovery rate is 92.5 wt.%.

[0133] (7) The same as Example 1, where the vanadium leaching rate is 86.2%.

[0134] The acidolysis rate of 95% concentrated sulfuric acid for the vanadium leaching residue (titanium slag) is 97%.

[0135] Example 3

[0136] Compared with Example 1, the difference is only that the steps of step (4) to step (6) are changed, specifically:

[0137] (4) Charge the iron-vanadium-titanium co-selected concentrate, biomass (wood chips), composite binder (moisture Mad = 8.49%, ash Ad = 28.73%, volatile matter Vdaf = 60.86%), and magnesite in a mass ratio of 100:3:1:1.5, mix them evenly to obtain a mixture, and prepare the mixture into carbon-containing pellets with a diameter of 10 mm to 14 mm through a pelletizing process.

[0138] (5) The carbon-containing pellets in step (4) are reduced at a reduction temperature of 1000 °C in an H2 atmosphere for 60 min to obtain pre-reduced pellets with a metallization rate of 96.57 wt.%, an FeO content of 3.54 wt.%, and an MgO content of 5.87 wt.%.

[0139] (6) The pre-reduced pellets in step (5) and the reducing agent are proportioned at a mass ratio of 100:0.5, and are reduced and smelted at 1450 °C for 1 h to separate hot metal and vanadium- and titanium-rich materials. The obtained iron sample has a TFe of 99.8 wt.% and an iron recovery rate of 93.09%. The TiO2 grade in the vanadium- and titanium-rich materials is above 50 wt.%.

[0140] The titanium recovery rate is 94.6 wt.%, the V2O5 grade is 1.3 wt.%, and the vanadium recovery rate is 84.5 wt.%.

[0141] (7) Similar to Example 1, wherein the vanadium leaching rate is 85.2%.

[0142] The sulfuric acid digestion rate of 95% concentrated sulfuric acid for the vanadium leaching residue (titanium slag) is 94.2%.

[0143] Example 4

[0144] Compared with Example 1, the difference is only that the process steps of steps (1) to (2) are changed, and other operations and parameters are the same as those in Example 1. The different steps (1) to (2) are as follows:

[0145] (1) Feed the original vanadium-titanium magnetite ore into the first-stage crushing operation for crushing and grinding, and control the fineness of the grinding product so that the content of -0.15 mm is ≥70 wt.%.

[0146] (2) Feed the grinding product in step (1) into a one-stage TARC countercurrent inclined table concentration for separation. Control the feed concentration at 20%, the feed rate at 100 L / min, and the water flow rate at 16 L / min. The underflow obtained by separation is used as the one-stage TARC concentration rough concentrate, and the overflow is used as tailings and directly discarded.

[0147] Feed the one-stage TARC concentration iron rough concentrate into the two-stage TARC concentration operation. Control the feed concentration at 20%, the feed rate at 100 L / min, and the water flow rate at 16 L / min. The underflow obtained by separation is used as the final iron-vanadium-titanium co-enriched mixed concentrate, and the overflow obtained by separation is used as tailings and directly discarded. Combine the one-stage TARC concentration tailings and the two-stage TARC concentration tailings as tailings. The indexes of the mixed concentrate and tailings obtained according to the above steps are shown in Table 3.

[0148]

[0149] After the mixed concentrate is processed through steps (3) to (7), the effects are as follows: for the molten iron and vanadium-rich titanium material, the obtained iron sample has a TFe of 99.6 wt.%, an iron recovery rate of 90%, the TiO2 grade in the vanadium-rich titanium material is above 50 wt.%, the V2O5 grade is 1.45%, and the titanium and vanadium recovery rates are 98.4% and 90.2% respectively.

[0150] Example 5

[0151] Compared with Example 1, the difference is only that the steps of steps (4) to (6) are changed. Specifically:

[0152] (4) Charge and mix the iron-vanadium-titanium co-selected concentrate, biomass (wood chips), composite binder (moisture Mad = 8.49%, ash Ad = 28.73%, volatile matter Vdaf = 60.86%), and magnesite in a mass ratio of 100:6:1:3 to obtain a mixed material. Prepare the mixed material into carbon-containing pellets with a diameter of 10 mm to 14 mm through a pelletizing process.

[0153] (5) Reduce the carbon-containing pellets in step (4) in an H2 atmosphere at a reduction temperature of 1000 °C for 60 min to obtain pre-reduced pellets with a metallization rate of 96.56 wt.%, an FeO content of 3.26 wt.%, and an MgO content of 8.42 wt.%.

[0154] (6) Reduce and melt-separate the pre-reduced pellets in step (5) at 1450 °C (the wood chip addition amount is 0.5% of the pellet weight) for 1 h to obtain molten iron and vanadium-rich titanium material. The obtained iron sample has a TFe of 99.8 wt.%, an iron recovery rate of 92.58%, and the TiO2 grade in the vanadium-rich titanium material is above 50 wt.%.

[0155] The titanium recovery rate is 95.3 wt.%, the V2O5 grade is 1.35 wt.%, and the vanadium recovery rate is 86.5 wt.%.

[0156] (7) The same as in Example 1, where the vanadium leaching rate is 83.7%.

[0157] The 95% concentrated sulfuric acid acidolysis rate of the vanadium-leaching slag (titanium slag) is 95.7%.

[0158] Example 6

[0159] Compared with Example 1, the difference is only that the process of step (6) is changed. Specifically:

[0160] (6) Reduce and melt-separate the pre-reduced pellets in step (5) (the wood chip addition amount is 0.5% of the pellet weight) at 1600 °C for 0.5 h to obtain molten iron and vanadium-rich titanium material. The obtained iron sample has a TFe of 99.4 wt.%, an iron recovery rate of 90.27%, and the TiO2 grade in the vanadium-rich titanium material is above 50 wt.%.

[0161] The titanium recovery rate is 94.1 wt.%, the V2O5 grade is 1.16 wt.%, and the vanadium recovery rate is 82.2 wt.%.

[0162] (7) The same as Example 1, wherein the vanadium leaching rate is 84.6%.

[0163] The acid digestion rate of 95% concentrated sulfuric acid for vanadium leaching slag (titanium slag) is 95.0%.

[0164] Comparative Example 1

[0165] Compared with Example 1, the difference is only that in step (2), the shaking table gravity separation process is used for the gravity separation step, and other operations and parameters are the same as those in Example 1.

[0166] After the smelting separation in step (6), the obtained iron sample has a TFe of 94.2 wt.%, an iron recovery rate of 80.73%, and the TiO2 grade in the vanadium-rich titanium material is below 40 wt.%, which is not suitable as a raw material for titanium extraction.

[0167] From Example 1 and Comparative Example 1, it can be seen that by using the countercurrent gravity separation enrichment method of the present invention, it is helpful to realize the co-enrichment and co-smelting of ilmenite and vanadium-titanium magnetite in the raw ore, and is conducive to realizing the selective separation of iron and vanadium-titanium.

[0168] Comparative Example 2

[0169] Compared with Example 1, the difference is only that in step (5), the parameters of the pre-reduced pellets are adjusted to: pre-reduced pellets with a metallization rate of 72.56 wt.%, an FeO content of 24.4 wt.%, and an MgO content of 4.75 wt.%.

[0170] The molten iron and vanadium-rich titanium material are obtained by reduction smelting separation. The obtained iron sample has a TFe of 98.5 wt.%, an iron recovery rate of 70%, and the TiO2 grade in the vanadium-rich titanium material is below 40 wt.%, which is not suitable as a raw material for titanium extraction.

[0171] In this case, the metallization rate of the pre-reduced pellets and parameters such as FeO are not controlled within the required range, resulting in unsatisfactory selectivity in the subsequent molten separation. It can also be known from this that by using the pre-reduction process described in the present invention and the combined control of the pre-reduction conditions, it is unexpectedly beneficial to successfully realize the separation behavior and sorting selectivity of iron-vanadium-titanium in the short process of raw ore co-enrichment and co-smelting.

[0172] Comparative Example 3

[0173] Compared with Example 1, the difference is only that in step (4), an equal weight of calcium oxide is used to replace the magnesite. Other operations and parameters are the same as those in Example 1.

[0174] The reduced smelting separation yields hot metal and vanadium- and titanium-rich materials. The obtained iron sample has a TFe of 98.9 wt.%. The slag type of the vanadium- and titanium-rich slag is not ideal, and it is leached using the leaching process of Example 1, with a vanadium leaching rate of only 20%.

[0175] From Example 1 and Comparative Example 3, it can be seen that without controlling the MgO content in the pre-reduced pellets, it is not conducive to the separation of iron and vanadium-titanium in subsequent smelting. In addition, it is also not conducive to the regulation of the active slag phase of vanadium-titanium.

[0176] Comparative Example 4

[0177] Compared with Example 1, the only difference is that in step (4), the amount of biomass is changed. The experimental groups are as follows:

[0178] Group A: The weight ratio of the iron-vanadium-titanium co-selected concentrate to biomass is 100:15. Other operations and parameters are the same as in Example 1. This case cannot form pellets and cannot be reduced by H2 subsequently.

[0179] Group B: The above-mentioned biomass is not added.

[0180] The results are as follows: In step 6, the obtained iron sample has a TFe of 99.5 wt.%, an iron recovery rate of 80.22%, and the TiO2 grade in the vanadium- and titanium-rich material is 43 wt.%.

[0181] The titanium recovery rate is 95 wt.%, and the vanadium recovery rate is 85 wt.%. In step 7, the vanadium leaching rate is 80.5%.

[0182] Comparative Example 5

[0183] Compared with Example 1, the only difference is that in step (4), the pre-reduction atmosphere is nitrogen. Other operations and parameters are the same as in Example 1. The metallization rate of the pre-reduced product is too low (<50%), and the smelting selectivity is not ideal.

[0184] From Example 1, Comparative Example 4, and Comparative Example 5, it can be seen that by using the magnesia-assisted internal carbon-gas solid reduction process in combination, the smelting behavior of iron, vanadium, and titanium in co-enrichment-co-smelting can be achieved, the separation selectivity of iron and vanadium-titanium slag can be improved, and the phase and physical and chemical characteristics of the vanadium-titanium slag can be optimized, which is conducive to the recovery rate in the vanadium-titanium slag.

[0185] Comparative Example 6

[0186] Compared with Example 1, the only difference is that the iron-vanadium-titanium co-selected concentrate obtained in step (3) is prepared into oxidized pellets and then pre-reduced. The different steps are as follows:

[0187] (3) The oxidized pellets prepared from the mixed concentrate obtained by gravity separation in step (2) are pre-reduced. The components of the oxidized pellets are shown in Table 4 to obtain vanadium-titanium magnetite metallized pellets.

[0188]

[0189] (4) Add hot metallized pellets, quicklime, and Shanxi Yilin coal into an electric furnace according to a mass ratio of 100:8.93:3.8 for deep reduction; after deep reduction, smelt at 1500 °C for 60 min, separate the slag from the iron, and obtain hot metal containing vanadium (the recovery rate of iron is 92%) and titanium-containing molten slag with compositions shown in Tables 5 and 6.

[0190]

[0191]

[0192] In this comparative example, the titanium-containing molten slag contains 49.4 wt% TiO2. Moreover, during the smelting and separation process, a large amount of calcareous flux needs to be added to reduce the melting point of the slag phase. Titanium in the slag mainly exists in the form of perovskite, which is difficult to utilize; and vanadium mainly enters the hot metal, and the subsequent vanadium extraction process is the same as the cumbersome vanadium extraction process in blast furnace smelting.

[0193] Example 7

[0194] Process according to Figure 2 the following process, and the steps are as follows:

[0195] (1) According to the process of Example 4, perform crushing, grinding, and countercurrent gravity separation to obtain gravity separation concentrate (composition shown in Table 3) and gravity separation tailings.

[0196] (2) Perform sulfide flotation of Ni / Co on the gravity separation concentrate obtained in step (1), control the pulp concentration of the flotation at 40% and pH = 6. The addition amount of sulfuric acid for roughing is 600 g / t, the addition amount for the first scavenging is 200 g / t, and the addition amount for the second scavenging is 100 g / t; the addition amount of butyl xanthate for roughing is 200 g / t, and the addition amount for the first scavenging is 80 g / t; the frother is 2 # The addition amount of oil for roughing is 40 g / t, and the addition amount for the first scavenging is 20 g / t; obtain sulfide concentrate containing Ni / Co, which can be used for extracting and smelting Ni / Co products.

[0197] (3) Perform reverse flotation of silicate minerals on the flotation tailings in step (2) (the one-roughing and two-scavenging process is the same as that in Example 1), the dosage of water glass as the inhibitor is 800 g / t, the dosage of the inhibitor for the first scavenging is 500 g / t, and the dosage of the inhibitor for the second scavenging is 300 g / t; the dosage of oleic acid as the collector is 1000 g / t, the dosage of the collector for the first scavenging is 500 g / t, and the dosage of the collector for the second scavenging is 300 g / t; the frother is 2 #The rough selection addition amount of oil is 80 g / t, the first scavenging addition amount is 40 g / t, and the second scavenging addition amount is 20 g / t, obtaining a reverse flotation iron-vanadium-titanium co-selection concentrate with a TFe grade exceeding 40 wt.%, a TiO2 grade exceeding 18 wt.%, and a V2O5 content exceeding 0.6 wt.%.

[0198] The gravity separation tailings are subjected to magnetic separation to obtain magnetic separation concentrate; the secondary selection concentrate and the reverse flotation iron-vanadium-titanium co-selection concentrate are mixed to form an iron-vanadium-titanium co-selection concentrate;

[0199] (4) The iron-vanadium-titanium co-selection concentrate in step (3), biomass (redwood sawdust), composite binder (moisture Mad = 8.49%, ash Ad = 28.73%, volatile matter Vdaf = 60.86%), and magnesite are proportioned and mixed at a mass ratio of 100:3:1:1.5 to obtain a mixture, and the mixture is prepared into carbon-containing pellets with a diameter of 10 mm to 14 mm through a pelletizing process.

[0200] (5) The carbon-containing pellets in step (4) are reduced at a reduction temperature of 1050 °C for 60 min in an H2 atmosphere to obtain pre-reduced pellets with a metallization rate of 90.53 wt.%, an FeO content of 12.58 wt.%, and an MgO content of 9.82 wt.%.

[0201] (6) The pre-reduced pellets in step (5) and the reducing agent are proportioned at a mass ratio of 100:0.5, and reduced and melted for 1 h at 1450 °C to separate into hot metal and vanadium-titanium-rich material. The obtained iron sample has a TFe of 98.4 wt.% and an iron recovery rate of 92.65%. The TiO2 grade in the vanadium-titanium-rich material is above 40 wt.%.

[0202] The titanium recovery rate is 96.8 wt.%, the V2O5 grade is 1.2 wt.%, and the vanadium recovery rate is 90.5 wt.%.

[0203] (7) The same as Example 1, wherein the vanadium leaching rate is 88.6%.

[0204] The 95% concentrated sulfuric acid acidolysis rate of the vanadium leaching residue (titanium slag) is 96.5%.

[0205] This case realizes the synchronous recovery of elements such as Ni / Co.

Claims

1. A short-process high-quality utilization method for vanadium-titanium magnetite resources, characterized in that the steps Comprising: (a) Crushing, grinding, and pre-concentrating the vanadium-titanium magnetite raw ore by countercurrent gravity separation; The gravity separation rough concentrate obtained by pre-concentration is further subjected to reverse flotation to remove aluminum and silicon minerals, obtaining an iron-vanadium-titanium co-selection concentrate enriched with ilmenite and vanadium-titanium magnetite in the raw ore; the vanadium-titanium magnetite raw ore includes ilmenite, magnetite, and titanomagnetite, and is disseminated with at least one of feldspar, olivine, pyroxene, serpentine, chlorite, and amphibole; among them, the occurrence phases of Fe include at least one of magnetite, ilmenite, ferro-magnesium olivine, and pyrite, and the occurrence phases of Ti include titanomagnetite and ilmenite; The weight ratio of the iron-vanadium-titanium co-selection concentrate, solid carbonaceous reducing agent, and binder is 100:(1-10):(0.5-3); (b) Mixing the iron-vanadium-titanium co-selection concentrate, solid carbonaceous reducing agent, binder, and magnesium-based regulator to obtain a mixed material, and pelletizing or briquetting the mixed material to prepare a carbon-containing pellet; Pre-reducing the carbon-containing pellet in a hydrogen-containing atmosphere in advance to obtain a pre-reduced pellet with a metallization degree of 85-98%, an FeO content of 1-15%, and an MgO content of 0.5-10%; (c) Melting and separating the pre-reduced pellet to obtain hot metal and vanadium-titanium-rich slag.

2. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to claim 1, wherein: The countercurrent gravity separation is TARC countercurrent inclined plane gravity separation; During the countercurrent gravity separation process, the feed concentration is 18-22%, the feed rate is controlled at 100-150 L / min, and the water flow rate is controlled at 14-18 L / min; In the iron-vanadium-titanium co-selection concentrate, the TFe content is 40 wt.% - 50 wt.%, the TiO2 content is 13 wt.% - 18 wt.%, and the V2O5 content is 0.4 wt.% - 1.0 wt.%.

3. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to claim 1, wherein: The reagents for the reverse flotation include a reverse flotation inhibitor and a reverse flotation collector; The reverse flotation inhibitor includes one or more of starch, water glass, carboxymethyl cellulose, sodium hexametaphosphate, sodium silicate, and sodium fluorosilicate; The reverse flotation collector is one or more of oleic acid, linoleic acid, cetyl quaternary ammonium salt, tetradecyl quaternary ammonium salt, and sodium dimethylaminodecyl carboxylate.

4. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to any one of claims 1 to 3, characterized in that: In step (b), the solid carbonaceous reducing agent includes a mixture of one or several of anthracite, bituminous coal, lignite, biomass, and biomass charcoal; The binder is a mixture of one or several of a composite binder, modified starch, and cellulose.

5. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to claim 4, characterized in that: The magnesium-based regulator includes at least one of magnesium oxides, hydroxides, carbonates, chlorides, nitrates, and organic acid salts.

6. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to claim 1, characterized in that: In the pre-reduction stage, the volume fraction of H2 in the hydrogen-containing atmosphere is greater than or equal to 15%; The pre-reduction temperature is 900 °C - 1200 °C, and the pre-reduction time is 10 - 90 min.

7. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to claim 1, characterized in that: The melting and separation temperature is 1300 °C - 1600 °C, and the melting time is 0.5 h - 4 h.

8. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to claim 1, characterized in that: Crush and finely grind the vanadium-titanium-rich slag, then perform acid leaching treatment to obtain a vanadium acid leaching solution and titanium slag; recover vanadium from the vanadium acid leaching solution and recover titanium from the titanium slag.

9. The method for short-process high-quality utilization of vanadium-titanium magnetite resources according to claim 8, characterized in that: After the vanadium-titanium-rich slag is crushed and finely ground, the particle size should be controlled such that the content of -0.074 mm is ≥ 90%; Among them, the acid solution for acid leaching is hydrochloric acid, with a concentration of 10% - 25%, the mass ratio of acid residue is 2:1 - 6:1, the leaching temperature is 30 - 160 °C, and the time is 0.5 - 5 h; Extract and separate vanadium from the vanadium acid leaching solution and prepare vanadium pentoxide products; The TiO2 grade of the titanium slag is ≥50%; Prepare titanium white from the titanium slag.

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