WO2009110819A1

WO2009110819A1 - Method for producing chemically active metals and a device for carrying out said method

Info

Publication number: WO2009110819A1
Application number: PCT/RU2009/000040
Authority: WO
Inventors: Анатолий Евгеньевич ВОЛКОВ
Original assignee: Volkov Anatoly Evgenievich
Priority date: 2008-02-26
Filing date: 2009-01-28
Publication date: 2009-09-11
Also published as: RU2401874C2; RU2008107403A

Abstract

The invention relates to a method for producing chemically active metals and a device for carrying out said method. The inventive method involves feeding a charge, consisting of a reducible metal compound and a reducing metal, into the inner cavity of a graphite electrode-anode. During heating, at first, metallothermic reduction is carried out inside an electrode, and when the charge exits on the end face of the electrode-anode in the arc burning area, the plasmochemical and electrochemical metal reduction is prolonged in such a way that a liquid metal and charge bath is formed. The metal is cooled on a crystalliser in the form of a cathode. The device comprises a body, a mechanical charge feeding device, a hopper and a branch pipe. The device is also provided with a crystalliser which is connected to the negative pole and is used as a cathode. The device is also provided with an graphite electrode which is connected to the positive pole and used as a anode with a central hole which is used for displacing the charge consisting of a reducible metal compound and a reducing metal. The mechanical device is placed inside the hopper.

Description

METHOD FOR PRODUCING CHEMICALLY ACTIVE METALS AND A DEVICE FOR ITS IMPLEMENTATION

Technical field

The present invention relates to the production of chemically active metals and can be used to produce titanium, zirconium, vanadium, tungsten, niobium, etc. metals.

State of the art

A close technical solution, as the first analogue, is the Cambridge method of titanium production [1]. (Scientific and technical journal "Titan" N ° 2 (19) 2006, pg. 9 ÷ 10). To date, this is one of the new methods in the production of reactive metals.

In recent years, considerable interest has been shown in the possibility of electrochemical production of titanium from oxide raw materials. It has been established that pre-stoichiometric titanium dioxide has high electrical conductivity and can serve as a cathode both for the reduction of titanium oxides to metal and for the removal of dissolved oxygen from it.

In 2000, based on the results obtained earlier by Frey, Farting, and Chen at the Department of Materials Science at Cambridge University, a direct electrochemical reduction of TiO ₂ to titanium in molten CaCl _{2 was} proposed. To introduce a new process, which in the scientific literature was called the Cambridge process or the Frey, Farthing and Chen process (FFC), a new company British Titanium (BTi) was created for commercial production.

The process uses electrochemical deoxidation of solid titanium dioxide, which was originally used by Okabe et al. For refining titanium metal. The developers demonstrated on a laboratory scale that the reduction reaction took place at 950 ° C on a cathode made of solid TiO ₂ , while the oxidation of oxygen anions takes place on a graphite anode with the release of CO ₂ . Removing a small amount of oxygen from the rutile insulator turns it into a highly conductive Magnelli phase (Tiθ _2- χ). Pure calcium chloride (CaCl ₂ ) was chosen as the molten salt electrolyte because of the high solubility of calcium oxide and the excellent migration properties of oxygen anions. The production process of pure titanium consists of the following sequence of operations. Pure titanium dioxide powder is mixed with a suitable binder to a paste or slip and cast into rectangular cathodes using one of the methods traditional for ceramic production, for example, by rolling or slip casting. The unburnt cathode is then calcined in an oven, where the sintering process begins to obtain a solid ceramic material. After sintering, shaped products are used as solid cathodes. Titanium is reduced in a closed electrolyzer filled with an inert gas. The electrolyzer is designed for continuous operation with cathodes installed and removed at different stages along the process through sealed sluices filled with inert gas. By controlling the cathode potential, it is possible to remove oxygen from titanium dioxide, while obtaining a high-purity metal similar in morphology from the sponge obtained by the magnesium thermal method. The voltage on the bath is approximately 3B, which is slightly lower than the decomposition voltage of CaCl ₂ (3.25V at 950 ⁰ C), but much higher than the decomposition voltage of TiO ₂ (1.85V at 950 ⁰ C), so chlorine gas is not formed on the anode. Sufficient overvoltage is necessary to reduce the oxygen content in titanium. The developers claim that mixtures of oxides of other metals with TiO ₂ in the initial cathode are also reduced to metals, making it possible to produce titanium alloys, despite a different microstructure. The process was demonstrated in a pilot reactor (capacity 1 kg per day).

The second analogue of the invention is a method of aluminothermic reduction of titanium dioxide [2]. (A.N. Zelikman - Metallurgy of refractory rare metals, - Moscow: Izd. "Metallurgy", 1986, pp. 417 ÷ 419).

In accordance with the affinity of oxygen to aluminum and titanium in its various oxides, titanium dioxide at temperatures up to 1400 ⁰ C can be reduced to TiO. However, aluminum forms solid solutions and intermetallic compounds with titanium with an additional release of Gibbs energy, which helps to restore to the formation of a titanium-aluminum alloy.

The possibility of aluminothermic reduction in an out-of-furnace process using the generated reaction heat simplifies and cheapens its hardware design. When TiO _{2 is} reduced by aluminum, the total thermal effect, taking into account the additional heat generated during the formation of the titanium-aluminum alloy, is about 1590 kJ / kg of a charge containing 25% RU2009 / 000040

excess aluminum versus stoichiometric calculation. This amount of heat is not enough for spontaneous development of the reaction with the melting of the alloy and slag. Therefore, in order to increase heat dissipation, Berthollet salt is added to the mixture and calcium chloride is used to reduce the melt viscosity. The mixture, stuffed into a graphite chamotte crucible, is ignited with the ignition mixture, and after the reaction spread rapidly throughout the entire volume at the bottom of the crucible, an alloy ingot is obtained that can be easily separated after cooling from the slag.

As titanium dioxide reducing agents, the use of magnesium, calcium, aluminum, and carbon is thermodynamically possible. However, aluminum and carbon are readily soluble in titanium, and therefore aluminothermic and carbothermic processes, if necessary, can only be used to produce titanium alloyed with aluminum or carbon.

A correct assessment of the thermodynamic probability of titanium oxide reduction processes should take into account the energy of formation of oxygen solid solutions in titanium. In this regard, the metallothermal reduction reaction of titanium oxides in general form can be represented by the equation

TiO _x + xMe = Ti + xMeO (1)

Titanium-aluminum alloys can be used as ligatures in the production of alloys, as well as for obtaining pure titanium from them by electrolytic refining. The combination of aluminothermic reduction of titanium dioxide with subsequent electrolytic refining can be one of the production methods for producing titanium.

The third analogue [2] (A.N. Zelikman - Metallurgy of refractory rare metals, - Moscow: Because of Metallurgy, 1986, pp. 414–416), zirconium reduction from tetrafluoride can be adopted. Unlike tetrachloride, zirconium tetrafluoride is stable in air and has a relatively high sublimation temperature (vapor pressure above solid ZrF ₄ reaches 0.1 MPa at 908 ⁰ C). This allows metallothermal reduction of tetrafluoride to be carried out, allowing the development of a high temperature at which smelting of a zirconium ingot is possible (t _PL Zr 1850 ⁰ C).

Get ZrF ₄ . due to the interaction of fluorine or hydrogen fluoride with zirconium dioxide at 500-550 ⁰ C:

ZrO ₂ + 2F ₄ == ZrF ₄ + O ₂ ; (2)

ZrO ₂ + 4H F = ZrF ₄ + 2H O ₂ . (3) The equipment material is nickel or Ni - Cu alloys (monel).

Zirconium tetrafluoride produced by this method contains an unacceptably high oxygen content and requires purification. Cleaning is carried out by sublimation in vacuum at 650-850 ° C.

Zirconium tetrafluoride is reduced by calcium by exothermic reaction

ZrF ₄ + 2Ca == Zr + 2CaF ₂ ; ΔH ° i _100K = - 613 kJ, (4) or 2480 kJ / cr charge.

The resulting CaF ₂ melts at 1419 ⁰ C, boils at 2630 ° C, which is favorable for the smelting of a zirconium ingot.

Recovery is carried out in a sealed steel reactor lined with calcium oxide, calcined at high temperature. Use high-purity calcium (purified by double distillation), which is introduced into the mixture in the form of chips. To ensure the smelting of the ingot according to one of the options, iodine (the heat of reaction of iodine with calcium 533 kJ / mol I ₂ ) is used as a heating additive that increases the thermal process. The reactor is evacuated, filled with argon and heated in an oven to 850 ° C, after which the process proceeds spontaneously at high speed. With a reactor diameter of 460 mm, a zirconium ingot weighing up to 450 kg is obtained. Consumable electrodes for arc remelting, which removes impurities, are made from ingots.

The disadvantages of the process include the significant consumption of expensive iodine and the possibility of contamination of the ingot due to contact with the lining. To reduce the temperature of the smelted ingot, it is proposed to introduce zinc into the charge, which leads to the formation of a relatively low-melting alloy of zirconium with zinc (an alloy with a zinc content of 25% melts at 1400 ° C). After distillation of the zinc in a vacuum at 1500 ° C, zirconium remains in the form of a sponge, which enters the smelter. The approximate content of impurities in ingots after arc melting is as follows,%: C 0.03; About 0.1; N, 0.005; Si 0.02; Al 0.01; Ca 0.003; Fe 0.02; Ti 0.001. Impurities of other metals in the range of 0.001 - 0.003%.

If necessary, alloying metals are added to the mixture of calcium-thermal reduction, obtaining ingots of doped zirconium.

Another variety of the third analogue is the following method. If zirconium powder is not intended for the production of ductile metal, but for direct use in pyrotechnics or, in some cases, in vacuum technology (for example, as a getter) and is not presented to it high purity requirements, it can be obtained by reducing K ₂ ZrF ₆ sodium similar to the production of tantalum and niobium powders from complex fluoride salts.

Potassium fluorozirconate is non-hygroscopic, stable in air. The sodium and potassium fluorides resulting from the reduction are washed off from zirconium with water.

The specific thermal effect of the reaction

K ₂ ZrF ₆ + 4Na = Zr + 4NaF + 2KF. (5) is about 1105 kJ / kg of charge. For the development of the reaction requires an additional supply of heat by heating the mixture in an electric furnace.

Recovery is carried out in steel glasses, where K ₂ ZrF ₆ and sodium pieces are loaded in layers (115 - 120% of the stoichiometric amount). The glass is sealed by attaching the lid to the flange, or, more simply, by welding it. The glass is kept in an oven at 800 - 900 ⁰ C. After cooling, the melted mass is knocked out with a pneumatic hammer, treated with a solution of NH ₄ Cl to dissolve the remaining sodium, ground in water in a ball mill, and KF and NaF are leached with water in stirred reactors. In order to avoid ignition of residual sodium and hydrogen formed during washing, the material is fed into the water gradually, in small portions.

To remove iron, the powder is treated with dilute hydrochloric acid, washed with water, filtered and dried at 60 ° C. When drying, care should be taken, since the powder is prone to self-ignition, therefore, drying is preferable to be carried out in vacuum. Powders contain approximately 98 - 98.5% Zr. [2].

The closest technical solution, as a prototype, is a method for reducing titanium and zirconium with calcium.

The reduction of titanium and zirconia with calcium is used to obtain fine-grained powders of titanium and zirconium, which can be used to obtain products by powder metallurgy, as well as in the form of powders (mainly zirconium) in pyrotechnics and as getters in electronic devices.

The reduction of TiO ₂ and ZrO ₂ with calcium proceeds according to the reaction:

MeO ₂ + 2Ca = Me + 2CaO; (1) where Me is Ti or Zr.

The reduction of TiO ₂ and ZrO ₂ by reaction (1) is accompanied by significant heat

= -392 kJ for TiO ₂ and -240 kJ for ZrO ₂ ), but it is not enough for the spontaneous course of the process. Constant heating required reactor at 950 - 1100 ° C. Even with an excess of calcium in the mixture of 50-100% and the process in an airtight apparatus filled with pure argon, it is fundamentally impossible to obtain titanium and zirconium powders with an oxygen content below OD%. This is because, at an oxygen content of 0.005 - 0.07%, the affinity of calcium for oxygen at 1000 - 1100 ⁰ C is equal to the affinity of zirconium and titanium for oxygen dissolved in them. The actual oxygen content in the powders is 0.2 - 0.3%, since part of the oxygen is there in the form of oxide films formed by washing the powders from calcium.

In accordance with thermodynamic calculations for 1100 K in the case of magnetothermal reduction of titanium oxides, the Gibbs energy of the redox process assumes zero value when the oxygen content in titanium is 0.42% (by mass). In practical conditions, due to kinetic difficulties due to the low speed of diffusion processes in solids, an equilibrium state is not achieved. Therefore, the residual oxygen content is higher: during magnetothermic reduction up to 2.5%.

Calcium reduction is carried out in sealed apparatus made of heat-resistant steel, where a briquetted mixture of TiO ₂ or ZrO ₂ with calcium (in the form of chips) is loaded. Distilled calcium must be used. The apparatus is pumped out, filled with argon, heated to 1000 - 1100 ⁰ C and maintained at this temperature for about 1 hour. The reduction product is crushed, treated with a large volume of water [to remove part of CaO in the form of Ca (OH) ₂ ], then diluted with HCl, washed with water and dried in vacuum at 40 - 50 ° C.

As a result of the recovery, finely dispersed powders of titanium and zirconium (fineness of 2–3 μm) are obtained, since interlayers of refractory calcium oxide interfere with their growth. Particle enlargement is facilitated by the addition of CaCl ₂ , which forms the liquid phase. The mechanism of action of calcium chloride is that the resulting calcium oxide is dissolved in molten CaCl ₂ . At 1000 ° C, the solubility of CaO is ~ 25%. Even if the amount of CaCl ₂ added to the mixture is not enough to completely dissolve CaO in it, the remaining undissolved CaO particles recrystallize through the melt, acquiring some mobility, which allows titanium crystals to coalesce into larger particles. By adding various amounts of CaCl ₂ to the initial charge (TiO ₂ + Ca + CaCl ₂ ), it is possible to control the particle size of the resulting titanium powder. If you add calcium chloride in an amount sufficient to dissolve all the resulting CaO (approximately in molar ratios of CaO: CaCl ₂ = 2: 1), metal particles up to 10 microns in size are obtained.

Disclosure of invention

In order to solve the problem of obtaining a chemically active metal from its compounds more cheaply and quickly than with electrolysis and more efficiently and purely on inclusions than with metallothermy, this method was proposed.

This goal is achieved in that the known method for the production of chemically active metals involves heating the reaction mixture, in the presence of a reducing agent, reducing the oxygen at the metal cathode and anode, characterized in that the mixture does not need to be formed into the electrode, since it enters the cancer zone through the internal the cavity of the graphite electrode, which simultaneously serves as the anode, and the metal recovery process consists of the following chain: gradual heating of the restored charge inside the electrode, meta lotermicheskoe metal recovery within the electrode metallothermic, plazmennohimicheskoe and electrochemical reduction of the metal in the arcing zone, the formation of the metal bath and cooling it at the cathode, which is the mold. A device for the production of chemically active metals, containing a crystallizer, an electrode, a housing, a mechanical feed device, a hopper, a nozzle, differs in that the mold is connected to the negative pole and is the cathode, and the electrode to the positive is the anode, and the electrode is made with a central hole through which the charge moves to the mold, due to the mechanical device located inside the hopper. The mold is made in the form of a cone or hemisphere, and there are two nozzles for separate creation of vacuum or gas supply to the hopper and the casing of the device.

Brief Description of the Drawings

Figure 1 shows the installation that implements the proposed method; figure 2 is a General diagram of the implementation of the method; Fig. 3 shows a graph of the dependence of the technological zones of reduction and crystallization of the metal on temperature fields; figure 4 is a graph of the melting temperature of the oxide from the temperature of the onset of the reaction. The best embodiment of the invention

The installation includes a copper cooled crystallizer 1 connected to the negative pole 3, which makes it a cathode. The electrode 6 is connected to the positive pole 14 and is the anode through which the reaction charge 7 is fed through the mechanical device 12. The mold 1 is hermetically closed by the housing 9, which is connected through the tubular corrugation 11 to the hopper 13, in which the reaction charge 7 is located. The mechanical actuator 10 controls the position of the electrode 6 above the bath mirror of the metal 2, above which there is an electrolyte 4. The internal cavity of the installation is evacuated through a pipe 8a, which is directly connected to the housing 9 and the pipe 86, to tory connected to the hopper 13. Due to the fact that the plant can operate in a neutral medium through nozzles 8, it is possible in the process of reducing the metal in the internal cavity feeding helium or argon. In addition, the mode of gas supply and the creation of vacuum can be mixed.

So, for example, in order to facilitate the passage of the charge 7 through the cavity of the anode b into the hopper 13, argon is supplied through a nozzle 86 under a certain pressure, which not only pushes the charge into the melting cavity, but also passes through the melt and electrolyte, cleans the metal from gas and insoluble impurities. In this case, a vacuum can be created through the pipe 8a. When argon is supplied, through the pipe 86, which subsequently exits at the end of the electrode, it is possible to restore metal at this installation when the protective chamber 9 is not used, i.e. when the installation works outdoors. In this case, the reduced metal will be produced of a sufficiently high quality, since argon is supplied to its mirror and protects it from pollution by the gases present in the air. This installation will be quite inexpensive and can be used for a wide range of recoverable metals. The scheme may be the reverse, in order to reduce volatile vapors from the electrolyte mirror, argon is supplied through the nozzle 8a under a certain pressure, which, acting on the electrolyte mirror, drives a part of the melt into the internal cavity of the anode, since a vacuum is created through the nozzle 86. Due to this, metallothermal processes begin to take place inside the anode along its greater length, unlike the reverse circuit, under the action of high temperature, since the electrode b can be heated to a temperature of 2000 ÷ 3000 ⁰ C during the melting process. In addition, the CO and CO gas generated ₂ without getting into the electrolyte and without oxidizing it, it will be removed through the hopper and pipe 86. According to this scheme, CaCl ₂ can be used as an electrolyte and a metal can be reduced by electrolysis at a low voltage. As you know, CaCl _{2 is} very it dissolves metal oxides well, has very good oxygen anion transfer properties. That is, in this case, when the metal oxide and the metal of the reducing agent move in the inner cavity of the anode, metallothermal reduction processes will occur. This process is characterized by high productivity, but the metal being restored is not completely freed from the oxygen present in it. The oxygen content in the reduced metal can vary from 0.2 to 12%. After the metal reaches the end of the electrode, it falls under an arc discharge, in which the heating temperature can reach 2000–3000 ° C; at this heating temperature, the processes of thermal ionization sharply intensify. That is, the molecules of a substance begin to decay into their constituent atoms, which then turn into ions. Ionization of a substance can be enhanced by bombardment of a substance by charged particles, since the reduced metal, going to the end of the anode electrode is exposed to negative ions, under the influence of which the metal dissolves, which will be accompanied by oxygen evolution.

As you know, electrolysis is the basis of the electrochemical method of obtaining pure substances. That is, electrochemical oxidation occurs at the anode, where negatively charged ions become neutral and are released from the solution. A reduction reaction occurs at the cathode, where positive ions receive the missing electrons.

In the exit zone of the recoverable charge, complex electrochemical processes also take place, which in particular include the plasma-chemical process, due to which some chemical reactions can be accelerated. High plasma temperatures lead to a high rate of chemical reactions, both direct synthesis reactions and reverse decomposition reactions. If, in this case, the metal is reduced in the maximum temperature field of the anode core, and then it is moved to the melt pool mirror, where the temperature is lower, then the reverse oxidation reactions of the reduced metal can be complicated and the yield of the desired product can be significantly increased.

The general scheme of the implementation of the method is shown in FIG. 2, where the main technological zones of metal reduction and crystallization are indicated, as well as one of the possible temperature fields that can pass through the substance in these zones, shown in the graph (Fig. C). That is, the charge 7, moving along the inner part of the electrode b, in zone I begins to warm up to high temperatures, the range of which can range from 1000 to 2000 ⁰ C. Continuous heating of the substance in zone II leads to the fact that when the temperature of the charge mixture exceeds the melting temperature of the reduced oxide by a certain amount, according to the graph (Figure 4), a spontaneous reduction reaction begins. [3]. Due to the reaction in this zone, the metal begins to recover, while additional heat is released. If the reduced and unreduced metal gets into zone III, that is, when it reaches the end of the electrode, the metal will additionally heat up due to the electric arc in the anode spot and its temperature will increase even more, while its increase can reach 2500 ÷ 3500 ° C. That is, plasmochemical processes begin to occur in this zone, which significantly accelerate the course of metal reduction reactions. In the same zone, an electrochemical oxidation process begins, due to which oxygen is removed from the reduced and unreduced metal in the form of gas, released at the anode and removed from the reaction due to the pumping system. In the fourth zone, both electrochemical and metallothermal metal reduction processes continue to occur, due to which metal droplets are formed, forming a melt pool 2a. In this zone, the temperature begins to drop and all processes begin to slow down. In zone V, the formation of a metal bath 2a and its crystallization 26 occurs due to cooling by the crystallizer, while the reduction product in the form of metal exits the reaction, shifting it to the right. In zone VI, the resulting metal is finally cooling. In general, the metal reduction process consists of the following chain: gradual heating of the recoverable charge inside the electrode, metallothermal reduction of the metal inside the electrode, metallothermal, plasma-chemical and electrochemical reduction of the metal in the arc burning zone, the formation of the metal bath and its cooling on the crystallizer.

The operation scheme of the installation may include a mode when inert gas pressure is supplied through both nozzles 8a and 86, and the pressure can be different or variable depending on the problem being solved for the smelting of metal from its oxide, as well as from its chlorides, carbides and fluorides. That is, this scheme and the design of the installation allows a very wide range of settings to restore metal recovery. Moreover, it is possible to widely control the current and voltage at the anode and cathode, as well as the atmospheric pressure in the melting chamber and the hopper, from deep vacuum to inert gas pressure of several tens of atmospheres. These opportunities are due to the simplicity of the installation design, where there is nothing superfluous. The charge supply scheme through the inner part of the electrode is similar to the charge supply scheme for shaft furnaces, which in terms of efficiency and productivity rank first among other furnaces. The charge feeding scheme has the following advantages:

1. Briquetting is not required, what is used in ore-thermal furnaces. Briquetting the mixture increases the cost of the process by 27%.

2. It is not required to create a special electrode from oxides of reduced metals, which is used in the Cambridge process, which makes production more expensive.

3. The feed circuit of the charge inside the electrode is direct-flow and does not inhibit the process of slag formation. In the ore-thermal process, the charged charge enters through the slag, cooling it and increasing its viscosity, which leads to a loss in metal yield, since the slag bath is cooled. In the case of feeding the charge through the electrode, the opposite process occurs, the charge is heated, taking heat from the electrode, thereby increasing its service life. The gradual heating of the charge allows the recovery processes to gradually proceed over the course of the passing time, which, all accelerated, result in the fact that when the end of the electrode is reached, a metal is formed from the charge, which continues to be purified from oxygen by electrolysis at the maximum possible temperature. At the same time, metal droplets do not pass through the entire layer of the slag bath in which they could remain, but pass along a short path from the end of the electrode to the molten bath. On this path, in the zone of arc burning, the highest temperature, and therefore the lowest viscosity of the electrolyte, therefore, the resulting metal can easily and without loss to overcome this path.

The technological process for the production of chemically active metals can be disclosed by the example of the reduction of titanium dioxide, according to the Cambridge process and the second analogue, having finalized and accelerated the process by the fact that, together with electrolysis, titanium dioxide will be reduced by aluminum and carbon. As raw materials, it is possible to use both natural and artificial rutile containing at least 95% TiO ₂ . Aluminum is used in powder, the fraction corresponding to the highest reduction of titanium.

[3]. Dry powder flux is added to the powder mixture of TiO ₂ and Al, so that the metal bath formed near the electrode is constantly updated and does not lose its fluidity. Aluminum to titanium dioxide in quantity can enter 10 ÷ 50% less stoichiometric composition. Fluxing additives can be entered from 1 to 10%, which may be CaF ₂ .

In the titanium reduction reaction, the carbon source is an electrode, which is also the anode.

At the first stage of the process, the electrolyte is melted in the mold, due to the graphite electrode or the electrolyte is melted in a separate container, and then poured into the mold. CaF ₂ , CaCl ₂ can serve as an electrolyte (as in the first analogue), but also a combination of them or any other one used for electrolysis of titanium alloys. Through pipe 8a, a vacuum is created in the melting space in the range from 10 ⁴ to 10 ^{~ 4} Pa. An arc is ignited between the crystallizer and the electrode and the reaction chamber 7 is fed through the mechanism 12 due to the mechanism 12. As the feeding mechanism 12, there can be a screw, a piston, a free backfill mechanism, as well as any other mechanism providing an adjustable and uninterrupted supply of the charge .

In order to create a deep vacuum in the melting chamber, which shifts the reduction reaction towards titanium production, the crystallizer is made in one piece, not collapsible. In order to prevent a significant volume of slag forming over crystallizable titanium from being poured into another container and at the same time not to extend the electrode, the mold is made in the form of a cone or hemisphere. Moreover, if one of the lower half of the mold is filled with metal, the upper half of the mold will be able to accommodate slag in volume 2 ÷ 3 times the volume of the recovered metal. In addition, the installation uses a corrugated pipe 11, which provides a deep vacuum through which movement is transmitted to the electrode, due to mechanisms 10. The mechanisms may be pneumatic, hydraulic or screw drives. The hopper 13, after mixing and filling the reaction mixture into it, is hermetically closed. The layout of the proposed installation, due to its simplicity and minimal use of mechanisms and parts, is able to withstand both deep vacuum and high pressure and be quite simple and reliable in operation.

The feed rate of the charge depends on the current strength and voltage on the electrode. For a more complete electrolysis reaction to occur, the voltage at the electrode must be kept within the explosive zone, as in the Cambridge process. Optionally, the feed rate of the charge is selected so as to maximize the release of titanium from oxygen, while the minimum amount of reducing agent that Al and C will serve in this reaction at the maximum metal production rate was used.

In this case, the reduction of titanium will go through the reactions described above in the second analogue. Additional reactions will occur during the formation of alumina ABO3 in the electrolyte due to the titanium reduction reaction. But due to the occurrence of electrolysis reactions, the process of aluminum reduction will proceed in this zone:

At the cathode Al ³⁺ (w) + Зё -> Al (tv)

At the anode 2O ^{2 '} → O ₂ (g) + 4ё

2C (tv) + O ₂ → 2CO (g)

Oxygen and carbon will be removed from the reaction, and the main metal, the reducing agent, aluminum, will be returned to the reaction again. These reactions help to increase the yield of recoverable titanium and reduce the amount of aluminum loaded. An increase in the reducibility of titanium oxides is possible by removing alumina formed from the reaction sphere, for example, by introducing cryolite into the electrolyte, which dissolves the formed alumina.

Additionally, an increase in the concentration of reduced titanium dioxide will contribute to the shift of the reaction equilibrium to the right. This is economically feasible due to the lower cost of titanium dioxide compared to the cost of aluminum powder.

When comparing the Cambridge method with the proposed, the following can be noted:

1. The titanium reduction rate in the proposed method will be significantly higher due to the combination of metallothermal, plasmochemical and electrochemical reduction, since only one electrochemical titanium reduction process is involved in the Cambridge process.

2. The cost of production will also be less, since the process does not require additional production for the manufacture of electrodes.

When comparing the proposed method with the second analogue, where titanium is produced by aluminothermy using additional exothermic heat from the introduced thermite additives, the following can be noted:

1. The production rate of titanium in the proposed method, its purity, the amount will be significantly higher, since it does not use termite additives, compared with the second analogue. 2. The process is controlled during the smelting of titanium, in contrast to the second analogue.

3. Its cost is lower since termite additives are not used. Similar to this scheme of the disclosure of the method on the example of recovery

TiO ₂ aluminum, other processes will go, where the metals described in other analogues and prototype will be restored, namely:

1. When recovering zirconium from zirconium tetrafluoride with calcium, in the proposed method, when they are filled into the hopper, fluxing additives are not required, since CaF ₂ is formed during the course of the reaction, which is an electrolyte. Unlike the third analogue, in the proposed method, the lining and the proposed heating are not required. In addition, the addition of iodine, which increases the thermality, but pollutes the melt, is not required. In the future, if metal additives are not introduced, additional arc remelting, etc. is not required. Therefore, the proposed method is much simpler and more productive, while providing higher quality metal at a lower cost.

2. When restoring the proposed method of zirconium from potassium fluorozirconate sodium, the recovery pattern and benefits will be the same as described above.

3. When reducing Ti and Zr in a new way, in comparison with the prototype of TiO ₂ and ZrO ₂ calcium, the advantages will be the same. Therefore, the proposed method and device for its implementation in the production of metals, especially chemically active, can be considered useful for widespread implementation ..

As can be noted, in recent years, the possibility of reducing titanium from oxide raw materials has sharply increased interest, since the chlorine method is quite time-consuming and environmentally hazardous enough. On the other hand, over the past half century, titanium production technologies have changed far, as evidenced by the newly emerging methods for the electrochemical production of titanium from oxide raw materials, such as the Cambridge process, OS process, OKABE process and others.

Unlike metallothermal reduction of titanium due to magnesium, calcium, aluminum, these methods directly try to solve the problem of obtaining titanium from its oxides. In this regard, these methods are not productive, energy-consuming and time-consuming.

In the present invention, the method of titanium recovery uses a complex effect due to the use of metallothermal, plasma chemical and T / RU2009 / 000040

15 electrochemical processes. Their effectiveness is enhanced by the introduction of a vacuum or inert gas pressure into the reaction. Equipment that is used to implement the method in complexity is similar to equipment for vacuum arc remelting (VDP). Remelting the finished ingot at the VDP furnace increases its cost price by only 2%. In this case, this rise in price will be exactly the same. When chloramagnetothermic reduction of titanium by the Krol method, this appreciation is 26%. That is, the obtained metal for its subsequent purification can be remelted 13 times in order to achieve the same cost.

The economic efficiency of titanium production in the first stage:

For the production of primary titanium it is required: Al - l part; TiO ₂ - Parts; C 0.2 parts; CaF ₂ - 0.5 parts.

Price: Al - $ 4.8 / cr; TiO ₂ (95%) - 0.4 $ / cr; C - 0.2 $ / kg; CaF ₂ - 0.2 $ / cr.

The cost of one kg of the mixture in the mixture will cost l, 3 $ / kg. In production, we take a yield of 70%, which consists of 67% Ti and 10% Al, therefore, for lkg of the resulting product, material will be spent in the amount of $ 1.69. This product can be used without any purification as a substitute for ferrotitanium and ferroaluminium in the production of high-quality steels. With the cost of ferrotitanium 70%, which is equal to $ 7.5 / kg today, the cost of 67% Ti and 10% Al can be the same amount, since in this combination both titanium and aluminum for steel are active deoxidizing agents. Total costs, energy, work, salary, etc. the cost of lkg reaches $ 0.2 / kg. Therefore, the total cost of production will be l, 89 $ / kg, with a final price of 7.5 $ / kg, i.e. total profit will be $ 5.61 / kg. This profit is quite good, considering that only one equipment is involved in the process. To increase profits, the resulting slag is recycled by electrolysis of aluminum, as well as the extraction of the remaining rutile and flux.

LITERATURE

[one]. Scientific and technical journal "Titan" Na2 (19) 2006, pg. 9 ÷ 10.

[2]. A.N. Zelikman - Metallurgy of refractory rare metals, - Moscow: Izd-vo "Metallurgy", 1986, pp. 414 ÷ 419.

[3]. Yu.L. Pliner, S. I. Suchilnikov, E. A. Rubinshtein - Aluminothermic production of ferroalloys and alloys, - Moscow: "State Scientific and Technical Literature on Ferrous and Nonferrous Metallurgy", 1963, p. 28, 33.

Claims

CLAIM

1. A method for the production of chemically active metals, including heating the reaction mixture in the presence of a reducing agent, reducing the oxygen at the metal cathode and anode, characterized in that the mixture does not need to be formed into the electrode, since it enters the reaction zone through the internal cavity of the graphite electrode, which serves as an anode at the same time, and the metal reduction process consists of the following chain: gradual heating of the recoverable charge inside the electrode, metallothermal metal reduction three electrodes metallothermic, plazmennohimicheskoe and electrochemical reduction of the metal in the arcing zone, the formation of the metal bath and cooling it at the cathode, which is the mold.

2. A device for the production of chemically active metals, containing a mold, an electrode, a housing, a mechanical feed device, a hopper, a nozzle, characterized in that the mold is connected to the negative pole and is the cathode, and the electrode to the positive is the anode, and the electrode is made with the central hole through which the charge moves into the mold, due to the mechanical device located inside the hopper.

3. The device according to claim 2, characterized in that the mold is made in the form of a cone or a hemisphere, and two nozzles for separate creation of vacuum or gas supply to the hopper and the housing of the device.