Short-process high-quality utilization method for vanadium titano-magnetite resources

Abstract

The invention belongs to the field of recycling of metal mineral resources, and particularly relates to a short-process high-quality utilization method of vanadium titano-magnetite resources, which comprises the following steps: co-enriching raw vanadium titano-magnetite to obtain iron-vanadium-titanium co-separation concentrate; uniformly mixing the iron-vanadium-titanium co-separation concentrate, a solid carbonaceous reducing agent, a binder and a magnesium-based regulator, pelletizing or briquetting to prepare carbon-containing pellets; the carbon-containing pellets are subjected to pre-reduction treatment in a hydrogen-containing atmosphere in advance, and pre-reduced pellets are prepared; and the pre-reduced pellets are smelted and separated, and molten iron and vanadium-rich titanium slag are obtained. According to the method, the ilmenite and the vanadium titano-magnetite in the raw ore are co-enriched and co-smelted, the problems of excessive fine grinding, toxic flotation reagent use, resource waste and the like in a raw iron / titanium separation process are avoided, a calcium-free smelting new slag system of adjusting concentrate MgO through the mineral separation process and controlling FeO through the pre-reduction process is optimized, controllable separation of iron and vanadium and titanium is achieved, and the method is suitable for industrial production. And further selective extraction and recovery of vanadium and titanium are facilitated, and short-process high-quality utilization of vanadium titano-magnetite resources is achieved.

Description

Technical Field

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

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

[0003] The main idea of ​​the current vanadium-titanium magnetite resource development and utilization process is to separate the vanadium-titanium magnetite into iron concentrate and titanium concentrate, and then smelt the iron concentrate and titanium concentrate separately. The typical steps are: pre-separate the vanadium-titanium magnetite ore with weak magnetic separation to separate the iron concentrate (containing TFe ~55%, V 2 O 5 ~0.5%, TiO 2 ~10%), and the iron tailings are flotated to produce titanium concentrate (containing TFe ~35%, V 2 O 5 ~0.1%, TiO 2 ~48%). Iron concentrate is smelted in a blast furnace, and the iron element is recovered in the form of molten iron for steelmaking. After the vanadium element enters the molten iron during the smelting process, it is recovered through processes such as vanadium slag blowing in a converter and sodium vanadium extraction. 2 O 5 , titanium (accounting for about 52% of the total titanium resources) is transformed into stable perovskite and enters the blast furnace slag. This part of titanium cannot be used at all. At the same time, some of the co-existing rare elements such as nickel and cobalt also enter the blast furnace slag and cannot be used. The resulting titanium concentrate is used to produce titanium dioxide by the sulfuric acid method, and the iron and vanadium elements in the titanium concentrate cannot be recycled. The current vanadium-titanium magnetite resources adopt the route of separate separation and utilization of iron and titanium, which leads to the titanium, nickel and cobalt elements in the iron concentrate entering the blast furnace slag and cannot be recovered, the vanadium extraction process is cumbersome and the recovery rate is low, and the iron and vanadium in the titanium concentrate cannot be recovered. In the end, the utilization rates of iron, vanadium and titanium are only 79%, 44% and 29% respectively. In addition, the Chinese patent document with publication number CN115522074A discloses a slag making method for vanadium-titanium magnetite electric furnace smelting, which includes the following steps: S1, pre-reducing the vanadium-titanium magnetite concentrate oxidation pellets to obtain vanadium-titanium magnetite metallized pellets; S2, adding the vanadium-titanium magnetite metallized pellets, calcium slag-making agent and reducing agent to the electric furnace in proportion for deep reduction; S3, smelting after deep reduction, and separating the slag and iron to obtain vanadium-containing molten iron and titanium-containing molten slag; S4, vanadium-containing molten iron is subjected to converter vanadium extraction treatment, and titanium-containing molten slag is subjected to chemical method titanium extraction treatment. This method is complex, and vanadium enters the molten iron, and the treatment process has a large metal process loss.

[0004] In summary, the existing process for recovering vanadium-titanium magnetite metal resources is relatively complicated, and the process loss of valuable metal elements is 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 valuable metals from 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 flow, high selectivity and high recovery rate that does not require separation of iron, titanium and vanadium, and directly conducts reduction-smelting separation to selectively separate iron and vanadium-titanium.

[0006] There are mainly two types of minerals in the vanadium-titanium magnetite raw ore, namely 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 conduct smelting 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 situation, 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 realize 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 regulation 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, the present invention has conducted in-depth research and provides the following improvement scheme:

[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 titanomagnetite) 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) Melting 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 ideas 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. It first performs countercurrent gravity separation tailing rejection on the raw ore and reverse flotation to obtain a co-selected concentrate (co-enriched minerals) of ilmenite and vanadium-titanium magnetite in the raw ore, and then pre-charges it with carbon and hydrogen for auxiliary pre-reduction, and strictly controls the metallization degree of iron, the contents of FeO and MgO during the pre-reduction process. Based on the pre-reduced minerals with special physical and chemical characteristics during pre-reduction, subsequent smelting separation treatment is carried out, thereby regulating 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, thus 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 desilication and de-aluminum of vanadium-titanium magnetite, and performs special pre-reduction - smelting separation co-smelting treatment on the special co-enriched iron-vanadium-titanium co-selected concentrate. In this way, not only is the treatment process greatly shortened, avoiding the use of blast furnaces and harmful arsenic agents in the prior art, 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 intergrown 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, fayalite, and pyrite, and the occurrence phases of Ti include titanomagnetite and ilmenite. In addition, in the raw ore, at least one valuable element such as Ni, Co, scandium, etc. 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 V 2 O 5 、5 - 8 wt.% of TiO 2 . 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 the vanadium-titanium magnetite raw ore is crushed and ground, the particle size should be controlled to have a -0.15 mm content of 70 - 95%.

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

[0017] In the present invention, the countercurrent gravity concentration step in step (a) is as follows: countercurrent inclined plane gravity concentration column is used for gravity concentration, wherein 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.

[0018] Preferably, the number of times of countercurrent gravity concentration is 1-3 times.

[0019] In the present invention, desilication and dealumination 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 as needed.

[0024] In the present invention, the activator is a conventional component capable of achieving reverse flotation of aluminum and silicon, such as calcium chloride.

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

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

[0027] In the present invention, the tailings of the countercurrent gravity concentration can be subjected to magnetic separation as needed, 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-concentration process of countercurrent gravity concentration - reverse flotation, a co-concentrate of ilmenite - vanadium-titanium magnetite, that is, the co-concentrate of vanadium-titanium iron, can be obtained. The weight ratio of ilmenite to vanadium-titanium magnetite in the co-concentrate of vanadium-titanium iron is 1:5 - 3:4.

[0029] Preferably, in the co-selected concentrate of iron, vanadium and titanium, the content of TFe is 40 wt.% to 50 wt.%, and the content of TiO 2 is 13 wt.% to 18 wt.%, and the content of V 2 O 5 is 0.4 wt.% to 1.0 wt.%.

[0030] Preferably, the co-selected concentrate of iron, vanadium and titanium may further contain 0.4 wt.% to 8 wt.% of MgO.

[0031] In the present invention, for the co-enriched co-selected concentrate of iron, vanadium and titanium, 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. In this way, 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 biomass charcoal.

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

[0036] Preferably, in the blended material, a magnesium-based regulator is added as needed, and the additive includes at least one of magnesium oxide, hydroxide, carbonate, chloride, nitrate, and organic acid salt. For example, in the present invention, the weight ratio of the co-selected concentrate of iron, vanadium and titanium to the magnesium-based additive can be 100:1 to 3.

[0037] In the present invention, in the pre-reduction stage, the volume fraction of H 2 in the hydrogen-containing atmosphere is greater than or equal to 15%; further, it can be hydrogen.

[0038] Aiming at the problems faced in the combined smelting of vanadium-titanium-iron co-selected concentrate, such as the selectivity of iron-vanadium-titanium and the difficulty in controlling the slag properties of vanadium-titanium, the present invention innovatively conducts combined pre-reduction of the pre-reduced pellets in a carbon and hydrogen atmosphere, further combined with the combined control of temperature, metallization rate of the pre-reduced pellets, FeO, and MgO. In this way, synergy can be unexpectedly achieved, which is beneficial for the subsequent selective smelting separation of iron and vanadium-titanium. In addition, it is also beneficial for regulating the activity of vanadium-titanium slag and facilitating the recovery of vanadium-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, better short-process and high-efficiency recovery effects can be obtained.

[0040] The present invention innovatively conducts subsequent reduction smelting treatment on the pre-reduced pellets with the metallization rate, FeO, and MgO contents. 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 smelting separation can be optimized. In addition, it is also beneficial for regulating and obtaining a vanadium-titanium active slag phase that is conducive to the extraction and separation of vanadium-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 subjected to acid leaching treatment after being crushed and finely ground to obtain 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 being crushed and finely ground 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 acid to 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 vanadium pentoxide products.

[0047] In the present invention, the TiO 2 grade of the titanium slag ≥50%;

[0048] Preferably, titanium slag is used to prepare titanium white.

[0049] Beneficial effects:

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

[0051] Co-enrichment:

[0052] The present invention innovates the beneficiation process of iron / titanium separation and iron first then titanium. It directly synchronously enriches and recovers iron, titanium, and vanadium in the raw ore in their intrinsic substances through rough grinding - gravity separation - reverse flotation to remove silicon and aluminum, obtaining an iron-vanadium-titanium co-selected concentrate with a TFe content of 40 wt.% - 50 wt.%, a TiO 2 content of 13 wt.% - 18 wt.%, and a V 2 O 5 content of 0.4 wt.% - 1.0 wt.%. 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 caused by the current beneficiation process of separate iron / titanium separation. And for iron / titanium separation, 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 arsonic acid toxic agents in the flotation process increases the environmental protection pressure. Therefore, the present invention shortens the beneficiation process of vanadium-titanium magnetite, 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 and vanadium-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, which can greatly improve 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 (TiO 2 grade 45%). During the recovery process, TiO 2The loss rate is relatively high. In the present invention, the useful minerals ilmenite and magnetite in vanadium-titanium magnetite are dissociated from minerals such as feldspar, olivine, and pyroxene with relatively coarse dissemination sizes through crushing and grinding, and then fed into a countercurrent inclined heavy separation column for heavy separation operations. The separation is mainly carried out by utilizing the specific gravity differences of mineral particles, achieving the co-enrichment of iron-vanadium-titanium elements. In the countercurrent water flow that overcomes gravity and flows upward, the associated mineral particles of ilmenite, and the particles of feldspar, olivine, and pyroxene have relatively small specific gravities and form an overflow with the water flow, while the ilmenite and magnetite particles have relatively large specific gravities and settle downward against the countercurrent. Therefore, the TFe, TiO 2 , V 2 O 5 recovery rates in the underflow obtained by heavy separation all exceed 90%. Compared with ordinary heavy separation technologies, the recovery rates are increased significantly.

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

[0059] Traditional blast furnace smelting of vanadium-titanium iron concentrate requires that the TiO 2 content is not more than 13%, and the titanium slag type is not ideal, making it difficult to recycle; in addition, vanadium enters the hot metal 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. By synergistically regulating FeO-MgO in pre-reduced pellets, a new slag system for smelting is constructed, which can facilitate 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 conducive to the slag system regulation and slag-iron separation in the subsequent smelting process, and can reduce / avoid the addition of extra magnesia flux, achieving 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 hot metal. Subsequently, wet separation and extraction can be used to 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 smelting process all exceed 70%, and the comprehensive recovery rate is greatly improved compared with the current process.

[0061] (4) The difficulty of pellet pre-reduction is reduced

[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 original ore and the complex dissemination relationship are maintained, and the reduction difficulty is greater than that of the conventional sorted iron ore pellets. However, in view of 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. Moreover, with the consumption of the solid carbonaceous reducing agent, a large number of pores can be left inside the pellets, providing space for the consumption of the reduction expansion stress while improving the reduction kinetic conditions, and 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 kinetic conditions for reduction 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 titanomagnetite 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 the iron-vanadium-titanium-nickel-cobalt rough concentrate also contains scandium, through the said technology, scandium, nickel and 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 limiting the present invention.

[0068] An optional short-process high-quality utilization method for vanadium-titanium magnetite resources of 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 original vanadium-titanium magnetite ore is crushed and ground, and then subjected to countercurrent inclined-plane current gravity separation to obtain pre-concentrated rough concentrate. The rough concentrate is directly subjected to reverse flotation to obtain iron-vanadium-titanium co-selection concentrate, or after desulfurization flotation to recover valuable elements such as nickel and cobalt therein, and then subjected to reverse flotation treatment to obtain iron-vanadium-titanium co-selection concentrate; in addition, the tailings after gravity separation pre-concentration are further subjected to magnetic separation to recover the iron minerals therein and merge them into the co-selection concentrate;

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

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

[0072] (4) The pre-reduced pellets in step (3) are fed into an electric furnace for smelting separation to obtain hot metal and vanadium-titanium-rich slag, and the hot metal can be used for converter steelmaking;

[0073] (5) The vanadium-titanium-rich slag obtained in step (4) is crushed and finely ground, and then subjected to acid leaching - extraction separation to extract vanadium to prepare vanadium pentoxide products;

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

[0075] In the above method, the particle size of the original vanadium-titanium magnetite ore after crushing and grinding in step (1) should be controlled so that the content of -0.15 mm is ≥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 iron-vanadium-titanium co-selection rough concentrate with a TFe content of 40 wt.% - 50 wt.%, a TiO 2 content of 13 wt.% - 18 wt.%, a V 2 O 5 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 - 0.4 T.

[0078] In the desulfurization flotation - de-aluminum and silicon 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 desulfurization regulators and desulfurization collectors.

[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 sulfur-nitrogen 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 scavenging is 200 - 400 g / t, and the addition amount in the second scavenging is 0 - 200 g / t;

[0083] The addition amount of the roughing desulfurization collector is 80 - 200 g / t, the addition amount in the first scavenging is 40 - 80 g / t, and the addition amount in the second 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 scavenging is 50 - 600 g / t (further can be 450 - 500 g / t), and the dosage of the inhibitor in the second scavenging is 50 - 400 g / t (further can be 250 - 300 g / t).

[0089] The dosage of the roughing collector in 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 scavenging is 50 - 600 g / t (further can be 450 - 500 g / t), and the dosage of the collector in the second scavenging is 50 - 400 g / t (further can be 250 - 300 g / t).

[0090] In the reagents for reverse flotation desilication and de-alumination, 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 de-alumination and desilication, an activator is also allowed to be added, which can be, for example, calcium chloride.

[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 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 magnesia fluxes.

[0095] The mixed concentrate co-enriched with 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 a mass ratio of 100:(2.5 - 3.5):(1 - 1.5):(1.25 - 1.75) to obtain a mixed material.

[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 a mass ratio of 100:3:1:1.5 to obtain a mixed material.

[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 magnesia 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 reduction gas contains H 2 with a volume fraction greater than or equal to 15%. As a further preference, the reduction gas is a mixture of one or two of H 2 and CO, where the volume fraction of H 2 is greater than or equal to 15%. Further preferably, H 2 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 excessively reduced to metallic iron, and the FeO content in the pre-reduced pellets is too low to play a tempering 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, the FeO content in the slag during smelting separation will be too high, and the iron recovery rate will decrease.

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

[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 ≥98 wt.%, and the TiO 2 content in the vanadium-rich titanium material is ≥40 wt.%, and further above 50 wt.%.

[0104] The particle size of the vanadium-rich titanium slag should be controlled such that the content of -0.074 mm 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 TiO 2 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, a gravity separation device disclosed in a 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 hydrogen content 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, which is different from the conventional separation process. The weight ratio of ilmenite to vanadium-titanium magnetite therein can be 1:2 to 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 is ≥70 wt.%.

[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 120 L / min, and the water flow rate at 16 L / 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 120 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 tailings of the first-stage TARC re-election and the tailings of 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 operations. The steps are as follows: perform flotation roughing on the re-election concentrate to obtain the roughing tailings; perform the first scavenging 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 roughing, first sweeping and second sweeping include frother, collector, activator and inhibitor, wherein the frother is 2# oil; the inhibitor is starch; the activator is calcium chloride; the collector is oleic acid;

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

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

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

[0122] (4) The iron-vanadium-titanium co-selected concentrate of step (3) and biomass (wood chips), a composite binder (moisture content Mad=8.49%, ash content Ad=28.73%, volatile matter Vdaf=60.86%), and magnesite are mixed in a mass ratio of 100:5:1:1.5 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.

[0123] (5) The carbon-containing pellets in step (4) are heated in H 2 The pre-reduced pellets with a metallization rate of 94.43wt.%, a FeO content of 5.63wt.%, and a MgO content of 4.42wt.% were obtained by reduction at a reduction temperature of 1050°C for 30 minutes under an atmosphere.

[0124] (6) The pre-reduced pellets in step (5) were mixed with 0.3% of the weight of sawdust and reduced and melted at 1450°C for 1 h to separate molten iron and vanadium-rich titanium material. The obtained iron sample TFe was 99.8wt.%, and the iron recovery rate was 88.73%. The vanadium-rich titanium material (XRD Figure 1 )TiO 2 The grade is above 50wt.%, and the titanium recovery rate is 95wt.%, V 2 O 5 The grade is 1.3 wt.%, and the vanadium recovery rate is 85wt.%.

[0125] (7) After crushing and fine grinding, the particle size of vanadium-rich titanium slag should be controlled at -0.074mm and the content should be ≥90%, and then acid leaching treatment should be carried out. The initial hydrochloric acid concentration of the acid leaching is 20%, the acid-slag mass ratio is 3:1, the leaching temperature is 60 ºC, the time is 2 h, and then the solid-liquid separation is carried out to obtain vanadium leaching solution and vanadium leaching slag. The vanadium leaching rate is 86.8%.

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

[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) 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 are proportioned and mixed according to the 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.

[0130] (5) The carbon-containing pellets in step (4) are reduced at a reduction temperature of 1050 °C for 15 min in an H 2 atmosphere 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) The pre-reduced pellets in step (5) and the reducing agent (wood chips) are proportioned according to the mass ratio of 100:0.5, and reduced and melted at 1450 °C for 1 h to separate molten iron and vanadium-rich titanium material. The obtained iron sample has a TFe of 99.8 wt.%, an iron recovery rate of 80.91%, and the TiO 2 grade in the vanadium-rich titanium material is above 50 wt.%.

[0132] The titanium recovery rate is 97.3 wt.%, and the V 2 O 5 grade is 1.2 wt.%, and the vanadium recovery rate is 92.5 wt.%.

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

[0134] The acidolysis rate of vanadium-bearing slag (titanium slag) with 95% concentrated sulfuric acid 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 mixed material, and prepare the mixed material into carbon-containing pellets with a diameter of 10 mm to 14 mm through a pelletizing process.

[0138] (5) Reduce the carbon-containing pellets in step (4) at a reduction temperature of 1000 °C for 60 min in an H 2 atmosphere 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) Charge the pre-reduced pellets in step (5) and the reducing agent in a mass ratio of 100:0.5, 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 93.09%. The TiO 2 grade in the vanadium-titanium-rich material is above 50 wt.%.

[0140] The titanium recovery rate is 94.6 wt.%, and the V 2 O 5 grade is 1.3 wt.%, and the vanadium recovery rate is 84.5 wt.%.

[0141] (7) Same as 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 plane gravity separation 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 rough concentrate of the one-stage TARC gravity separation, and the overflow is used as tailings and directly discarded.

[0147] Feed a section of TARC reselected iron rough concentrate into the second-stage TARC re-election 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 tailings from the first-stage TARC re-election and the tailings from the second-stage TARC re-election as tailings. The indexes of the mixed concentrate and tailings obtained according to the above steps are shown in Table 3.

[0148]

[0149] Treat the mixed concentrate through steps (3) to (7), and the effects are as follows: molten iron and vanadium-titanium-rich material. The obtained iron sample has a TFe of 99.6 wt.%, an iron recovery rate of 90%, and the TiO 2 grade in the vanadium-titanium-rich material is above 50 wt.%, and the V 2 O 5 grade is 1.45%, and the recovery rates of titanium and vanadium are 98.4% and 90.2% respectively.

[0150] Example 5

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

[0152] (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 100:6:1:3 for batching and mixing 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 H 2 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 amount of wood chips added is 0.5% of the pellet weight) for 1 h to obtain molten iron and vanadium-titanium-rich material. The obtained iron sample has a TFe of 99.8 wt.%, an iron recovery rate of 92.58%, and the TiO 2 grade in the vanadium-titanium-rich material is above 50 wt.%.

[0155] The titanium recovery rate is 95.3 wt.%, and the V 2 O 5 grade is 1.35 wt.%, and the vanadium recovery rate is 86.5 wt.%.

[0156] (7) Similar to Example 1, where the vanadium leaching rate is 83.7%.

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

[0158] Example 6

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

[0160] (6) The pre-reduced pellets in step (5) (with the amount of sawdust added being 0.5% of the pellet weight) are reduced and smelted at 1600 °C for 0.5 h to separate 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 TiO 2 grade is above 50 wt.%.

[0161] The titanium recovery rate is 94.1 wt.%, and the V 2 O 5 grade is 1.16 wt.%, and the vanadium recovery rate is 82.2 wt.%.

[0162] (7) Similar to Example 1, where the vanadium leaching rate is 84.6%.

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

[0164] Comparative Example 1

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

[0166] After smelting 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] It can be seen from Example 1 and Comparative Example 1 that adopting the countercurrent gravity separation enrichment method of the present invention helps to achieve the co-enrichment and co-smelting of ilmenite and vanadium-titanium magnetite in the raw ore, and is conducive to 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 reduced smelting separation yields hot metal and vanadium- and titanium-rich materials. The obtained iron sample has a TFe of 98.5 wt.%, and the iron recovery rate is 70%. The TiO2 grade in the vanadium- and titanium-rich materials is below 40 wt.%, making it unsuitable as a raw material for titanium extraction.

[0171] In this case, parameters such as the metallization rate of the pre-reduced pellets and FeO were not controlled within the required ranges, resulting in unsatisfactory selectivity in the subsequent smelting separation. Thus, it can also be known that by using the pre-reduction process described in the present invention and jointly controlling the described pre-reduction conditions, it is unexpectedly beneficial for successfully realizing the separation behavior and sorting selectivity of iron-vanadium-titanium in the short process of raw ore - co-enrichment - 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 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 unsatisfactory. When leached using the leaching process of Example 1, the vanadium leaching rate is 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 the subsequent smelting, and in addition, it is not conducive to the regulation of the vanadium-titanium active slag phase.

[0176] Comparative Example 4

[0177] Compared with Example 1, the difference is only that in step (4), the amount of biomass used 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. In this case, pellets cannot be formed, and subsequent H2 reduction cannot be carried out.

[0179] Group B: The described biomass is not added.

[0180] The results are as follows: In step 6, the obtained iron sample has a TFe of 99.5 wt.%, the iron recovery rate is 80.22%, and the TiO 2 grade in the vanadium- and titanium-rich materials 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 atmosphere for pre-reduction 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 separation 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-bearing - gas-solid reduction process combination, the smelting separation behavior of iron, vanadium, and titanium in co-enrichment - co-smelting can be achieved, the separation selectivity between iron and vanadium-titanium slag can be improved, and the phase and physical and chemical characteristics of vanadium-titanium slag can be optimized, which is beneficial to the recovery rate in 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 metallized pellets of vanadium-titanium magnetite.

[0188]

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

[0190]

[0191] In this comparative example, the titanium-containing smelting slag contains 49.4 wt% TiO 2 , and during the smelting separation process, a large amount of calcium-containing flux needs to be added to lower 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 molten iron, and the subsequent vanadium extraction process is the same as the cumbersome vanadium extraction process in blast furnace smelting.

[0192] Example 7

[0193] According to Figure 2 the process flow for treatment, the steps are as follows:

[0194] (1) According to the process of Example 4, crushing, grinding, and countercurrent gravity separation are carried out to obtain gravity separation concentrate (the composition is shown in Table 3) and gravity separation tailings.

[0195] (2) Conduct sulfide flotation of Ni / Co on the reselected concentrate obtained in step (1), controlling the pulp concentration of the flotation at 40% and pH = 6. The addition amount of sulfuric acid for roughing is 600 g / t, for the first scavenging is 200 g / t, and for the second scavenging is 100 g / t; the addition amount of butyl xanthate for roughing is 200 g / t, and for the first scavenging is 80 g / t; the foaming agent is 2 # The addition amount of # oil for roughing is 40 g / t, and for the first scavenging is 20 g / t; obtain the sulfide concentrate containing Ni / Co, which can be used for extracting and smelting Ni / Co products.

[0196] (3) Conduct reverse flotation of silicate minerals on the flotation tailings in step (2) (the one-roughing and two-scavenging process is the same as in Example 1). The dosage of water glass as the depressant is 800 g / t, for the first scavenging is 500 g / t, and for the second scavenging is 300 g / t; the dosage of oleic acid as the collector is 1000 g / t, for the first scavenging is 500 g / t, and for the second scavenging is 300 g / t; the foaming agent is 2 # The addition amount of # oil for roughing is 80 g / t, for the first scavenging is 40 g / t, and for the second scavenging is 20 g / t, obtain the reverse flotation iron-vanadium-titanium co-selection concentrate with a TFe grade exceeding 40 wt.%, a TiO 2 grade exceeding 18 wt.%, and a V 2 O 5 content exceeding 0.6 wt.%.

[0197] Conduct magnetic separation on the reselected tailings to obtain magnetic separation concentrate; mix the secondary separation concentrate and the reverse flotation iron-vanadium-titanium co-selection concentrate to form the iron-vanadium-titanium co-selection concentrate;

[0198] (4) Charge and mix the iron-vanadium-titanium co-selection concentrate in step (3), biomass (redwood sawdust), composite binder (Mad = 8.49%, Ad = 28.73%, Vdaf = 60.86%), and magnesite in a mass ratio of 100:3:1:1.5 to obtain a mixed material, and prepare the mixed material into carbon-containing pellets with a diameter of 10 mm - 14 mm through a pelletizing process.

[0199] (5) Reduce the carbon-containing pellets in step (4) in an H 2 atmosphere at a reduction temperature of 1050 °C for 60 min 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.%.

[0200] (6) The pre-reduced pellets and the reducing agent in step (5) were mixed in a mass ratio of 100:0.5, and reduced and melted at 1450°C for 1 h to obtain molten iron and vanadium-rich titanium material. The obtained iron sample TFe was 98.4 wt.%, and the iron recovery rate was 92.65%. The TiO in the vanadium-rich titanium material was 1.34 wt.%. 2 The grade is above 40wt.%.

[0201] Titanium recovery rate is 96.8wt.%, V 2 O 5 The grade is 1.2wt.%, and the vanadium recovery rate is 90.5wt.%.

[0202] (7) Same as Example 1, wherein the vanadium leaching rate is 88.6%.

[0203] The acidolysis rate of vanadium leaching slag (titanium slag) with 95% concentrated sulfuric acid is 96.5%.

[0204] This case realizes the simultaneous recovery of elements such as Ni / Co.

Claims

1. A short-process high-quality utilization method of vanadium-titanium magnetite resources, characterized in that the steps include: (a) pre-enriching the vanadium-titanium magnetite ore by crushing, grinding and countercurrent gravity separation; The rough concentrate obtained by pre-enrichment is then subjected to reverse flotation to remove aluminum and silicon minerals, thereby obtaining an iron-vanadium-titanium co-selected concentrate enriched with ilmenite and vanadium-titanium magnetite in the original ore; (b) mixing the iron-vanadium-titanium co-selected concentrate, a solid carbonaceous reducing agent, a binder, and a magnesium-based regulator to obtain a mixed material, and pelletizing or agglomerating the mixed material to prepare carbon-containing pellets; The carbon-containing pellets are pre-reduced in a hydrogen-containing atmosphere to obtain pre-reduced pellets with a metallization degree of 85-98%, a FeO content of 1-15%, and a MgO content of 0.5-10%; (c) Smelting and separating the pre-reduced pellets to obtain molten iron and vanadium-titanium-rich slag.

2. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 1 is characterized in that: The vanadium-titanium magnetite ore includes ilmenite, magnetite and titanomagnetite, and is interspersed with at least one of feldspar, olivine, pyroxene, serpentine, chlorite and amphibole; wherein the Fe occurrence phase includes at least one of magnetite, ilmenite, ferromagnetite and pyrite, and the Ti occurrence phase includes titanomagnetite and ilmenite.

3. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 1 is characterized in that: The countercurrent gravity separation is TARC countercurrent slope 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-selected concentrate, the TFe content is 40wt.%~50wt.%, the TiO2 content is 13wt.%~18wt.%, and the V2O5 content is 0.4wt.%~1.0wt.%.

4. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 1 is characterized by: The reverse flotation reagents include reverse flotation inhibitors and reverse flotation collectors; The reverse flotation depressant 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, hexadecyl quaternary ammonium salt, tetradecyl quaternary ammonium salt, and sodium dimethylaminodecyl carboxylate.

5. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to any one of claims 1 to 4, characterized in that: In step (b), the solid carbonaceous reducing agent comprises one or a mixture of anthracite, bituminous coal, lignite, biomass and biochar; The binder is a composite binder, modified starch, cellulose or a mixture of the two or more thereof; The weight ratio of the iron-vanadium-titanium co-selected concentrate, the solid carbonaceous reducing agent and the binder is 100:(1-10):(0.5-3).

6. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 5 is characterized by: The magnesium-based regulator includes at least one of magnesium oxide, hydroxide, carbonate, chloride, nitrate and organic acid salt.

7. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 1 is characterized by: 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. to 1200° C., and the pre-reduction time is 10 to 90 minutes.

8. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 1 is characterized by: The smelting and separation temperature is 1300℃~1600℃, and the smelting time is 0.5h~4h.

9. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 1 is characterized by: The vanadium-rich titanium slag is crushed, finely ground and then acid-leached to obtain vanadic acid leaching solution and titanium slag; vanadium is recovered from the vanadic acid leaching solution, and titanium is recovered from the titanium slag.

10. The short-process high-quality utilization method of vanadium-titanium magnetite resources according to claim 9 is characterized in that: The particle size of the vanadium-titanium-rich slag after crushing and fine grinding should be controlled at -0.074mm and the content should be ≥90%; The acid solution for acid leaching is hydrochloric acid, the concentration of which is 10% to 25%, the acid-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; Extracting and separating vanadium from vanadic acid leaching solution and preparing vanadium pentoxide product; The TiO2 grade of titanium slag is ≥50%; Titanium slag is used to prepare titanium dioxide.

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