JPS6117771B2 - - Google Patents

Description

DETAILED DESCRIPTION OF THE INVENTION The present invention relates to the extraction of vanadium from materials containing vanadium, such as gold slag, and in particular from steel refining processes, such as slag produced in basic oxygen converters. This relates to the treatment of lime-rich slag produced in

The process of the present invention comprises combining pulverized slag with sulfur or sulfur-containing substances in a fluidized bed reactor or other suitable reactor in air or oxygen-enriched air.
This is carried out by firing at a temperature of 550°C to 850°C, and leaching the resulting fired product with a dilute sulfuric acid solution to dissolve vanadium valuables. Vanadium can be recovered from concentrated solution by precipitation, solvent extraction, or any other suitable or conventional method.

One of the important sources of vanadium is gold slag and, in particular, steel smelting, where most of the vanadium forms complex spinel-type structures and is in the trivalent state. This is the slag produced. Depending on the method used to smelt steel, the amount of free lime contained in the slag may vary. The slag produced in basic oxygen converters typically contains less than 45% CaO, and the current methods used to extract vanadium values from the slag consume significant amounts of reactants.

A number of methods have been proposed and used to extract vanadium valuables from gold slag.
Typical patent applications relating to extraction of vanadium from gold slag include Peters (U.S. Pat. No. 3,929,460, December 30, 1975) or Burwell (U.S. Pat. No. 3,206,277) , September 14, 1965], in which the slag is heated with sodium carbonate or an alkali salt in a reactor in the presence of oxygen to a temperature of 600°C to 800°C, and then the calcined product is subjected to alkaline leaching. A method is disclosed comprising performing an operation and subsequently harvesting vanadium.

Another method suggested by many people is Hildreth [U.S. Pat. No. 3,227,545, 1966]
[January 4, 2016], and this method involves heating the fired product in chlorine gas or hydrochloric acid gas at 1000°C.
chlorination, followed by selective condensation to separate vanadium chloride from iron chloride, and then purification of vanadium chloride.

Nevertheless, all of these methods are energy intensive since none of the reactants used generate the energy necessary to heat the slag to the required temperature. Moreover, if the slag is lime-rich, such as the slag produced in basic oxygen converters, the consumption of reactants may be high and the vanadium recovery rate may be low. Some methods, such as the method disclosed by Hildreth, are energy intensive and highly corrosive due to the use of hot chlorine or hydrochloric acid.

The method described herein for extracting vanadium values from gold slag or the like has the following advantages over currently practiced methods:

(a) Firing method in actual use All the energy required in the sintering process, which consumes a large amount of energy, is supplied by the oxidation heat of sulfur or sulfur-containing substances that are sintered together with the slag, so the amount of energy required is There are fewer and fewer.

(b) Owing to the low cost and low energy consumption of the reactants primarily used in the process of the invention, the overall cost of production by the process of the invention is lower than other processes currently in use.

(c) All steps of the method of the invention use conventional techniques, so that it is easy to apply or retrofit existing facilities.

(d) The effluent can be treated by well-known techniques at reasonable cost and is less corrosive than most current methods.

The method of the present invention basically consists of the following several steps.

(a) Grinding slag or vanadium-containing materials and sulfur or sulfur-containing materials to the appropriate particle size;

(b) Continuously charging reasonable amounts of these substances into a suitable reactor.

(c) Calcining the slag or vanadium-containing material and the sulfur or sulfur-containing material in the reactor at the desired temperature and for the desired reaction time.

(d) Continuously remove the fired product from the reactor.

(e) Leaching the fired product with a sulfuric acid solution.

(f) Solid-liquid separation of pulp from the leaching process.

(g) extracting vanadium from the liquid by any suitable method;

Next, the steps of the method of the present invention will be explained in more detail.

(a) grinding of gold slag or vanadium-containing substances, and sulfur or sulfur-containing substances;

The slag or vanadium-containing material to be charged to the reactor and the sulfur or sulfur-containing material are ground to a reasonable particle size, usually from -10 mesh 100% to -325 mesh 100 according to American Society for Testing and Materials standards.
%, but I found that -100 mesh 100% is appropriate.

More conveniently, these materials can be milled together or separately. This is because there is no difference in the yield results no matter which method is used.

Sulfating the lime in the slag to form calcium sulfate tends to form a thick layer on the particles, reducing the overall sulfation and thus the spinel to liberate vanadium during the leaching process. decreases the decomposition of This effect becomes more pronounced when the particles are large, so in order to increase the extraction rate of vanadium, it is necessary to pulverize the slag or vanadium-containing material to some extent.

(b) Charge these substances to the reactor.

The substances listed in the preceding paragraph may be mixed or separately,
It can then be fed to the reactor either continuously or in batches, either dry, as a cake after filtration, or as a slurry directly coming out of the attritor. It has been found that there is no substantial difference in the yield results regardless of the preparation method and format used.

In order to satisfy the heat balance and mass balance in the reactor at the same time, sufficient calcium, magnesium, and manganese present in the slag or vanadium-containing material are sulfated and to maintain the reaction at the desired temperature. The ratio of slag or vanadium-containing material to sulfur or sulfur-containing material is adjusted in such a way that sulfur is present. Primarily based on the concentration of sulfur in the material used and the concentration of free lime in the slag or vanadium-containing material, this ratio for the ratio of slag to sulfur or vanadium-containing material to sulfur-containing material is usually 5: It varies from 1 to 1:3.

(c) Firing of the substance.

The substances listed in paragraph (a) may be used in a suitable reactor, such as a fluidized bed reactor, in an amount of 0% less than the stoichiometric amount required to oxidize slag or vanadium-containing substances, and sulfur or sulfur-containing substances. Calcinate with an excess of oxygen up to 200% at a temperature of 550°C to 850°C for a reaction time of 30 minutes to 12 hours. The firing gas used may be air or oxygen-enriched air.

Whether dry or slurry raw materials are used, the next leaching process will yield 80%
To achieve a vanadium extraction rate of over 70%
A calcination temperature of 750° C. and an average reaction time of 3 hours were found to be suitable using a % excess of air.

The sulfation reaction uses sulfur dioxide or sulfur trioxide, or a mixture of both, produced from any external source, and mixes this gas with air or other gases, instead of adding sulfur or sulfur-containing substances. The slag or vanadium-containing material can also be sulphated by blowing it into the reactor. In this case, another source of heat must be supplied to the reactor in order to maintain the material to be sulfated at the desired temperature.

Whether sulfur or sulfur-containing materials are used mixed with the slag, or sulfur dioxide or sulfur trioxide produced by combustion of elemental sulfur outside the reactor is blown into the reactor, the result is gold or gold. are similar.

However, the use of sulfur-containing materials such as pyrite concentrate or other metal sulfides, or elemental sulfur, allows the reaction to maintain the desired temperature without adding any heat to the reactor. It has been found that there is an additional advantage of being able to supply all the heat.

(d) Removal of the fired product from the reactor.

The calcined product produced in the reactor can be removed either continuously or in batches directly into a leaching vessel to dissolve the vanadium values, or it can be cooled in a suitable calcined product cooler before leaching. You can also do it. If the calcined product is removed directly into the leaching tank, due to its heat content, the leaching can be carried out at a higher temperature to give a slightly higher vanadium extraction rate.

(e) Leaching of fired products.

The calcined material, either cold or still hot directly from the reactor, is leached with 10 g/- to 100 g/- sulfuric acid for 30 minutes to 10 hours in a stirring vessel. Nevertheless, in order to achieve an extraction rate of 80% or more of the vanadium valuables contained in the fired product, and to leave substantially all of the iron and calcium contained in the charged product in the form of hematite and snails in the solid residue. It has been found that 50g/2 hours of sulfuric acid is sufficient for this purpose.

(f) Solid-liquid separation.

The pulp obtained by leaching the calcined product can be separated either by concentrating it and passing it through the concentrator effluent, or by passing it directly. Which method is suitable depends only on the overspeed of the pulp obtained by leaching.

(g) Vanadium extraction from leach solution.

Vanadium in solution can be dissolved in any suitable manner by
For example, precipitated sodium heptavanadate,
That is, it can be collected in the form of "red cake", Na ₂ H ₂ V ₆ O ₁₇ .

After solid-liquid separation, the precipitate of the let cake is "air dried" either by drying or by melting.
It is possible to obtain an industrial grade red cake of grade or "fused" grade.

Another method of extracting vanadium from leach solutions is by solvent extraction. In this method, a product with good purity can be obtained. The extractant may be a quaternary amine or a suitable solvent from which vanadium salts such as ammonium vanadate can be precipitated. If this product is calcined at a temperature of 600℃ to 800℃,
A highly pure vanadium pentoxide product is obtained.

Nevertheless, these methods are only examples of the various methods that can be applied to extract vanadium from liquids.

Below are examples of materials tested under the conditions described herein.

Example 1 The following composition produced in a basic oxygen converter VV ₂ O ₅ 6.5% by weight Fe 15.0% Mn.MnO 3.3% by weight PP ₂ O ₅ 3.1% by weight Si.SiO ₂ 11.1% by weight Mg A steel smelting slag containing 3.8% by weight as MgO and 42.1% by weight as Ca.CaO is ground to 100% -150 mesh according to American Society for Testing and Materials standards, and has the following composition: 74.0% by weight as Fe.FeS ₂ -200 mesh 100% pyrite flotation concentrate with American Society for Testing and Materials specifications having 19.3% by weight as SiO ₂ . The slag and pyrite concentrate are mixed in a 2/1 ratio of slag/pyrite, dried and placed in a fluidized bed reactor operating at 750°C with a 70% excess of air over the stoichiometric amount. Continuously prepared. The average reaction time inside the reactor was 4 hours and the charging rate was 7.5 metric tons ^per square meter of hearth per day. The calcined product discharged from the reactor was cooled to room temperature and leached with 45 g/solution of sulfuric acid at 30% solids for 3 hours. The pulp was further strained and the cake was washed with water. Mix the solution and washing to add 24.2% vanadium.
A solution was obtained in which 82% of vanadium was dissolved and extracted. The solution also contained 1.2 g/m iron, 0.6 g/m manganese, and small amounts of other impurities.

The solution is further oxidized with a stoichiometric amount of sodium hypochlorite to convert trivalent vanadium into a pentavalent state.
Add 30g/Na ₂ SO ₄ and 60g/NaC, adjust pH
In step 2, sodium heptavanadate was precipitated by heating at 90°C for 8 hours. After precipitation, the solution was filtered and the resulting red cake was air-dried. The dried cake contains 95.5% V ₂ O ₅ and 0.15% Fe. The precipitation efficiency is 91%, and the overall recovery rate of vanadium from slag to reed cake is 74.6%.
It was hot.

Example 2 In another example, the following composition generated in the steel refining process, 4.6% by weight as VV ₂ O ₅ Fe 13.8% by weight as Mn/MnO 2.9% by weight as PP ₂ O ₅ 3.8% by weight as Mg/MgO A vanadium-containing slag having a content of 13.2% by weight as Ca and 46.1% by weight as CaO is ground to -150 mesh 100% according to the American Society for Testing and Materials standards, and -100 mesh 100% according to the American Society for Testing and Materials standards.
of sulfur concentrate containing 62.0% by weight of elemental sulfur, 23.3% by weight of SiO ₂ and small amounts of other inert materials. Mixing slag in a ratio of slag/sulfur concentrate = 3/1 and continuously charging it into a fluidized bed reactor at a rate of 10 metric tons ^per square meter of hearth area of the reactor per day, 750 for an average reaction time of 3.5 hours
The reaction was carried out at ℃.

The calcined product removed from the reactor was cooled to room temperature and leached with 45 g/solution of H ₂ SO ₄ at 30% solids. The resulting pulp was further filtered and the cake was washed with water. The extraction rate of vanadium was 79%, and 27.7 g of vanadium and iron were extracted by mixing the filter and washing liquid.
A solution was obtained containing 0.9 g/manganese, 0.8 g/manganese, and low concentrations of other impurities.

Vanadium was then precipitated from the solution in the same manner as described in Example 1 to obtain a red cake containing 97.2% by weight of V ₂ O ₅ and 0.10% by weight of Fe. The precipitation efficiency is 94%, and the overall recovery rate of vanadium from slag to red cake is
It was 74.3%.

Claims (1)

[Claims] 1. Grinding a vanadium-containing material or slag to a fine particle size and reacting it with sulfur or a sulfur-containing material or with sulfur dioxide gas and/or sulfur trioxide gas in a suitable reactor. The resulting calcined product is leached with a sulfuric acid solution to produce a solution with a high concentration of vanadium and low concentrations of other dissolved metals, and a solid residue containing iron, calcium, and other elements, and any In extracting vanadium from the solution by any suitable or conventional method, (a) grind the slag or vanadium-containing material to approximately -10 mesh according to American Society for Testing and Materials standards;
100% to 100% -325 mesh and blended with the sulfur or sulfur-containing material, such as elemental sulfur or pyrite concentrate, with a slag to sulfur-containing material ratio of 5:1 to 1:3. (b) charging the mixture into a suitable reactor; (c) converting the mixture in the reactor to between 0% and 200% using air or air with a low oxygen content;
(d) Calcinate the resulting calcined product continuously from said calcining step at a temperature of from 550°C to 850°C for a reaction time ranging from 1/2 hour to 10 hours in excess oxygen. (e) The calcined product is treated with a sulfuric acid solution containing 5 to 100 g of sulfuric acid at a temperature of 10 to 100 °C,
(f) subjecting the pulp obtained from said leaching to solid-liquid separation; (g) extracting soluble vanadium from said solution by any suitable or conventional method; A method for extracting vanadium from a vanadium-containing material such as steel smelting slag,
A method of treating lime-rich slag, such as that produced in basic oxygen converters, especially for steel smelting. 2 The particle size of the slag or vanadium-containing substance and the sulfur or sulfur-containing substance is according to American Society for Testing and Materials standards,
The method of paragraph 1, wherein the method is in the range of about 100% -10 meshes to 100% -325 meshes. 3. Dry or moisten the mixture.
The method according to item 1 above, wherein the method is charged into a reactor. 4 Slag: sulfur, or slag: sulfur-containing substance,
Alternatively, the mixing ratio of vanadium-containing substance: sulfur-containing substance is within the range of 5:1 to 1:3,
The method according to item 1 above. 5. The method according to item 1 above, wherein the firing temperature is within the range of 550°C to 850°C. 6. The method according to item 1 above, wherein the firing time is within the range of 1/2 hour to 10 hours. 7 The firing gas is air, oxygen-rich air, air and sulfur dioxide, air and sulfur trioxide, oxygen-rich air and sulfur dioxide, oxygen-rich air and sulfur trioxide, air and sulfur dioxide and sulfur trioxide, or oxygen. 2. The method according to item 1, wherein the method is selected from air rich in air and sulfur dioxide and sulfur trioxide. 8. The method according to item 1 above, wherein the sulfur dioxide and/or sulfur trioxide may be produced inside the reactor or may be produced externally and blown into the reactor. 9. Add 10g to 100g of sulfuric acid to the fired product from the firing process while it is still hot or after it has cooled down.
The method according to item 1 above, wherein the method is leached with a solution of up to 10. The method according to item 1 above, wherein the leaching temperature is within the range of 10°C to 100°C. 11. The method according to item 1 above, wherein the pulp produced in the leaching step is subjected to liquid-solid separation treatment. 12 The vanadium contained in the liquid obtained from the solid-liquid separation process is collected and converted into sodium heptavanadate, i.e., “red cake”.
2. The method according to item 1 above, wherein the method is in the form of ``Cake)'', Na ₂ H ₂ V ₆ O ₁₇ . 13. The method according to item 1, wherein the vanadium contained in the liquid obtained from the solid-liquid separation step is collected by solvent extraction and converted into ammonium vanadate, NH ₄ VO ₃ .