CN103184346A - Method for selectively separating vanadium, lead and zinc from vanadinite - Google Patents

Abstract

The invention discloses a method for selectively separating vanadium, lead and zinc from vanadium-lead ore. The method is to finely grind the vanadium-lead ore into ore powder and mix it with an alkali metal salt flux and a carbonaceous reducing agent to obtain a charge; The obtained charge is pelletized and then placed in a smelting furnace for smelting at 1100-1250°C; the slag obtained after smelting is leached with water at 85-90°C for vanadium, and filtered to obtain a vanadium leaching solution; the method is simple, easy to operate, and The metal comprehensive utilization rate is high, the production efficiency is high, the environment is friendly, the production cost is low, and the economic benefit is good.

Description

A method for selectively separating vanadium, lead and zinc from vanadinite

technical field

The invention relates to a method for selectively separating vanadium, lead and zinc from vanadium-lead ore, belonging to the technical field of mineral extraction and metallurgy.

technical background

Vanadinite is mainly produced as secondary minerals in the oxidation zone of lead-zinc deposits, and is symbiotic with lead-zinc oxide minerals. The vanadinite concentrate obtained by beneficiation often contains zinc, which is a kind of polymetallic oxide mineral. Vanadinite concentrate usually contains Pb32～42%, V ₂ O ₅ 8～13%, Zn10～13%, some also contain 10-60g/t Ag, which has important utilization value.

At present, there are two methods for the utilization of vanadium-lead ore concentrate in China:

One is as raw material for lead smelting, using traditional lead smelting methods to directly smelt crude lead. For polymetallic vanadic-lead ores, this method only utilizes the lead therein, while the vanadium and zinc are not utilized, and the resource utilization rate is extremely low. Since the zinc in vanadinite concentrate is as high as 10-13%, vanadinite concentrate can only be used for ore blending as raw material for lead smelting.

The second is to extract vanadium pentoxide by acid leaching. Vanadinite concentrate is a mineral raw material with very complex components. During acid leaching, in addition to the dissolution of vanadium compounds, many impurity compounds will also dissolve in acid and enter the acid leaching solution. These impurity ions will combine with polyvanadate ions to form insoluble polyvanadate or heteropolyacid salt, so that the dissolved vanadium will re-enter the acid leaching residue and cannot be recovered. Moreover, the acid leaching reaction of vanadium compounds is reversible. When the acidity of the solution decreases, the reaction will proceed in the opposite direction. In order to improve the leaching rate of vanadium, vanadium-lead ore acid leaching vanadium must be carried out under high acidity, but when the acidity is too high, the impurities entering the solution will also increase accordingly, which not only affects the yield of vanadium, but also affects the pentoxide The quality of divanadium also brings great troubles to the subsequent precipitation operation and refining of vanadium. The vanadium-lead ore concentrate is used to extract vanadium pentoxide, and the acid leaching residue retains almost all the lead in the vanadin-lead ore, which can be used as a lead smelting raw material, which improves the utilization rate of valuable metals in the vanadin-lead ore. However, there are problems that the acid leaching efficiency of vanadium-lead ore is low, the yield of vanadium is only 60%, the process flow is long, the consumption of acid and alkali is large, the acid leaching method requires high equipment, the operating environment is poor, and the process will A large amount of waste acid water is produced, which is difficult to deal with and high in cost. Moreover, the residue after vanadium extraction can only be used as raw material for lead smelting, and its economic added value is not high.

Contents of the invention

The invention aims at the low resource utilization rate in the traditional vanadium-lead ore smelting process and the low acid leaching efficiency in the process of extracting vanadium pentoxide by acid leaching, difficult to remove impurities, long process flow, large acid and alkali consumption, and high equipment requirements. The purpose is to provide a method for selectively separating vanadium, lead and zinc from vanadium-lead ore that can make full use of metal resources and has high recovery rate, low cost, environmental friendliness, short flow process and simple process. method.

The invention provides a method for selectively separating vanadium, lead and zinc from vanadium-lead ore. The method is to finely grind vanadand-lead ore into ore powder and mix it with an alkali metal salt flux and a carbonaceous reducing agent to obtain a charge; The obtained charge is pelletized and then placed in a smelting furnace and smelted at 1100-1250°C; the slag obtained after smelting is leached with water at 85-90°C and then filtered to obtain a vanadium leaching solution; wherein, during the smelting process, zinc is The gaseous volatilizes, the lead sinks in the bottom of the furnace in a liquid state, and the vanadium remains in the slag in the form of alkali metal vanadate.

The alkali metal salt flux is one or a mixture of Na ₂ CO ₃ or K ₂ CO ₃ .

The amount of the alkali metal salt flux added is 1.2 to 2.0 times the theoretical molar amount required for the respective reactions of vanadium, silicon and phosphorus in the vanadinite to generate the corresponding M ₃ VO ₄ , M ₂ SiO ₃ , and M ₃ PO ₄ , where M is Na or K.

The carbonaceous reducing agent is coke.

The added amount of _the reducing agent is 1.2-1.5 times of the theoretical molar amount required for the respective reduction reaction of PbO, ZnO and _Fe2O3 in the vanadinite to generate corresponding Pb, Zn and FeO.

The smelting time is 20-60 minutes.

The liquid-solid volume ratio of the leach water and the slag is 3-5:1.

The leaching time is 75-100 minutes.

After the vanadinite is finely ground, the particle size is not greater than 0.18mm.

The particle size of the carbonaceous reducing agent is not greater than 0.18mm.

During the smelting process, the zinc in the vanadinite is recovered in the form of smoke after volatilization in the gaseous state, the liquid sinks to the bottom of the furnace and is recovered in the form of crude lead, and the vanadium is retained in the slag in the form of vanadate and is leached with water.

The pelletizing described in the method of the present invention is formed by using a high-pressure double-roller briquetting machine, the size of the pellets is 20-40mm, the molding pressure is 17-20Mpa, and the moisture content of the dried pellets is not more than 2.0%.

The smelting process of the method of the present invention can be smelted in an electric furnace, or can be smelted in a reverberatory furnace; when the reverberatory furnace is used for smelting, the charge does not need to be pelletized.

In the method of the present invention, if the vanadinite contains Ag, it will be enriched in crude lead and recovered in the lead refining process.

The concrete method of the present invention selectively separates vanadium, lead and zinc from vanadinite, comprises the steps:

The first step: Grinding, mixing, pelletizing

Finely grind vanadinite ore to an ore powder with an average particle size of no more than 0.18mm, add a flux and a carbonaceous reducing agent with an average particle size of no more than 0.18mm to the obtained ore powder, mix them uniformly, and pelletize to obtain pellets;

The second step: reduction smelting

The pellets obtained in the first step are placed in a smelting furnace, heated to a smelting temperature of 1100-1250°C, and kept for 20-60 minutes;

The lead and zinc in vanadinite are selectively reduced, and the lead is reduced to lead metal, which is released from the bottom of the furnace to obtain crude lead ore; zinc is volatilized with zinc vapor, oxidized in the air, and finally recovered in the form of soot; and vanadium Then it remains in the basic slag in the form of soluble vanadate; through the reduction smelting process, the effective separation of vanadium, lead and zinc is realized at one time;

Step Three: Extract Vanadium

The vanadium-containing basic slag obtained in the second step of reduction smelting is extracted with water to extract vanadium. The liquid-solid ratio of water and slag is 3-5:1, the leaching temperature is 85-90°C, and the leaching time is 75-100 minutes; after the leaching is completed, filter, Vanadium-containing leaching solution and leaching slag are obtained; the obtained vanadium-containing leaching solution can be purified by conventional chemical impurity removal methods in vanadium metallurgy, and then subjected to hydrolysis or ammonium salt vanadium precipitation to obtain vanadium products.

The technical principle of the present invention: the inventor overcomes the waste of resources in the traditional vanadium-lead ore smelting process and the defects of low vanadium pentoxide leaching rate and difficult to remove impurities in the process of extracting vanadium pentoxide by acid leaching in the smelting process of vanadium-lead ore, and proposes a new method for the first time. A method for the simultaneous selective separation of vanadium, lead and zinc from vanadinite ore by reduction smelting. In the smelting process of the method of the present invention, a strong base and a weak acid salt are added as a flux, combined with a coke reducing agent, and smelted in an environment with a temperature of 1100-1250 ° C, a series of complex reactions occur, and vanadium, silicon and phosphorus react respectively to form corresponding chemical compounds. Stable K ₃ VO ₄ , K ₂ SiO ₃ and K ₃ PO ₄ , or their sodium salts, while PbO, ZnO and Fe ₂ O ₃ are reduced to corresponding Pb, Zn and FeO respectively. At this melting temperature , Zn volatilizes in the form of steam and is oxidized to zinc oxide when it encounters air during the volatilization process, and finally it is collected as zinc oxide fume through a cloth bag, lead is stored at the bottom of the furnace to generate crude lead, and vanadium is left in the slag in the form of water-soluble salt In the process, only need to further use water as the leaching agent can be leached; the generated flue gas is dust-collected by a bag to obtain a crude zinc oxide product containing no less than 48% zinc. When the lead content in the crude zinc oxide product is high, sulfuric acid can be used to treat the It is leached to recover zinc, and after liquid-solid separation, zinc sulfate leaching solution and acid leaching residue are obtained. The zinc sulfate leaching solution can be processed according to the conventional process of purification and impurity removal and electrolytic deposition in zinc metallurgy to prepare new products, and the acid leaching residue returns to the present invention. The mixing process further recovers the lead in the residue.

Beneficial effects of the present invention:

1. The process of the present invention can effectively separate lead, vanadium and zinc at one time after reduction smelting in an alkaline flux, has high production efficiency, low cost, good economic benefits, simple and reasonable method, and convenient operation;

2. The recovery rate of metal in the reduction smelting process of vanadinite ore in the present invention is high, the reduction rate of lead is not less than 95%, the lead content in crude lead is not less than 96%, and the reduction volatilization rate of zinc is not less than 80%. The conversion rate of vanadium is not less than 95%, so that the resources of polymetallic vanadium-lead ore can be fully recycled;

3. In the vanadium-containing basic slag obtained by the reduction smelting process of vanadium-lead ore in the present invention, vanadium mainly exists in the form of sodium salt or potassium salt that is easily soluble in water. Only by simple water leaching, the vanadium-containing Vanadium is extracted by leaching from the alkali slag, avoiding the use of a large amount of acid for leaching, which is environmentally friendly and cheap.

Detailed ways

The following examples are intended to illustrate the present invention, but not to further limit the present invention.

Example 1:

Extract vanadium, lead and zinc from vanadium-lead ore containing Pb38.84%, _V2O510.34 %, _Zn11.23 %, Fe13.62%, _SiO22.94 %, Ag47g/t. The implementation process is as follows: grind the vanadinite ore to a particle size of ≤0.18mm, and then add flux sodium carbonate and reducing agent coke powder. The ratio is: vanadinite powder: sodium carbonate: coke powder is 100:28.0:8.5 (the amount of sodium carbonate added is 1.2 times the theoretical amount, the amount of reducing agent added is 1.2 times the theoretical amount), coke The powder contains 83% fixed carbon. Mix the accurately weighed materials uniformly in the mixer for 30 minutes. After mixing evenly, press the mixture into pellets. The diameter of the pellets is 30mm, and the molding pressure is 20Mpa. The moisture content of the pellets is less than 2% after natural drying. The pellets are sent into a melting furnace for reduction smelting. The melting furnace used in this embodiment is a reduction electric furnace with a transformer rated capacity of 400KVA and a controlled melting temperature of 1250°C. After all the pellets in the furnace are melted and the liquid level of the melt drops, continue to add pellets until the furnace is finally filled with the melt, and keep warm for 50 minutes thereafter. The crude lead is released from the metal discharge port at the bottom of the furnace, the vanadium-containing alkali slag is released from the slag discharge port, and the zinc-containing fume is collected by the bag dust collector during the smelting process. The resulting products are as follows:

1. Crude lead: lead in the obtained crude lead is 98.58%, lead reduction rate is 98.89%, silver in crude lead is 116g/t, and silver yield is 96.03%.

2. Vanadium-containing alkali slag: the obtained vanadium-containing alkali slag contains 26.52% vanadium pentoxide, and the slag is almost enriched with all the vanadium in the raw material. After the molten slag is condensed, it is crushed to a particle size of 0-8mm, and wet-milled in a ball mill. Transfer the ball-milled pulp into the leaching tank, add water to control the liquid-solid ratio to 3:1, the leaching time is 90 minutes, and the leaching temperature is 90°C. After leaching, filter and wash the filter cake with clean water. The resulting leaching residue contains 0.54% V ₂ O ₅ , and the leaching rate of vanadium reaches 99%. The obtained vanadium-containing leaching solution can be treated by conventional vanadium metallurgy process.

3. Smoke and dust: the reduction evaporation rate of zinc in this embodiment is 82.44%, and the resulting smoke and dust contain 54.32% zinc and 11.46% lead. Zinc is leached from smoke and dust with 180g/L sulfuric acid for 90 minutes at a liquid-to-solid ratio of 6:1 and a leaching temperature of 90°C. After the leaching is finished, filter to obtain lead-containing leaching residue and zinc sulfate solution. The leaching slag contains 2.5% zinc, and the leaching rate of zinc is 98.7%. The leaching slag contains 40.60% lead, and the leaching slag is returned to the mixing operation, and the lead in it is recovered by resmelting. The zinc sulfate solution can be treated according to the conventional process of zinc smelting.

Example 2:

Extract vanadium, lead and zinc from vanadium-lead ore containing Pb38.84%, _V2O510.34 %, _Zn11.23 %, Fe13.62%, _SiO22.94 %, Ag47g/t. The implementation process is as follows: grind the vanadinite ore to a particle size of ≤0.18mm, and then add flux sodium carbonate and reducing agent coke powder. The ratio is: vanadinite powder: sodium carbonate: coke powder is 100:46.0:10.5 (the amount of sodium carbonate added is 1.97 times the theoretical amount, the amount of reducing agent added is 1.45 times the theoretical amount), coke The powder contains 83% fixed carbon. Mix the accurately weighed materials uniformly in the mixer for 30 minutes. After mixing evenly, press the mixture into pellets. The diameter of the pellets is 30mm, and the molding pressure is 20Mpa. The moisture content of the pellets is less than 2% after natural drying. The pellets are sent into a melting furnace for reduction smelting. The melting furnace used in this embodiment is a reduction electric furnace with a transformer rated capacity of 400KVA and a controlled melting temperature of 1150°C. After all the pellets in the furnace are melted and the liquid level of the melt drops, continue to add pellets until the furnace is finally filled with the melt, and then keep warm for 25 minutes. The crude lead is released from the metal discharge port at the bottom of the furnace, the vanadium-containing alkali slag is released from the slag discharge port, and the zinc-containing fume is collected by the bag dust collector during the smelting process. The resulting products are as follows:

1. Crude lead: lead in crude lead is 97.50%, lead reduction rate is 98.85%, silver in crude lead is 114g/t, and silver yield is 95.82%.

2. Vanadium-containing alkali slag: vanadium-containing alkali slag contains 24.68% vanadium pentoxide, and the slag is almost enriched with all the vanadium in the raw material. After the molten slag is condensed, it is crushed to a particle size of 0-8mm, and wet-milled in a ball mill. Transfer the ball-milled pulp into the leaching tank, add water to control the liquid-solid ratio to 5:1, the leaching time is 90min, and the leaching temperature is 90°C. After leaching, filter and wash the filter cake with clean water. The resulting leaching residue contains 0.32% V ₂ O ₅ , and the leaching rate of vanadium is 99.4%. The obtained vanadium-containing leaching solution can be treated by conventional vanadium metallurgy process.

3, soot, the reduction evaporation rate of zinc in the present embodiment is 80.05%, and the gained soot contains 52.58% of zinc and 5.25% of lead.

Example 3:

Extract vanadium, lead and zinc from vanadium-lead ore containing Pb38.84%, _V2O510.34 %, _Zn11.23 %, Fe13.62%, _SiO22.94 %, Ag47g/t. The implementation process is as follows: grind the vanadinite ore to a particle size of ≤0.18mm, and then add flux potassium carbonate and reducing agent coke powder. The ratio is: vanadinite powder: potassium carbonate: coke powder = 100:45.0:9.4 (the amount of potassium carbonate added is 1.48 times the theoretical amount, and the amount of reducing agent added is 1.3 times the theoretical amount), coke The powder contains 83% fixed carbon. Mix the accurately weighed materials uniformly in the mixer for 30 minutes. After mixing evenly, press the mixture into pellets. The diameter of the pellets is 30mm, and the molding pressure is 20Mpa. The moisture content of the pellets is less than 2% after natural drying. The pellets are sent into a melting furnace for reduction smelting. The melting furnace used in this embodiment is a reduction electric furnace with a transformer rated capacity of 400KVA and a controlled melting temperature of 1200°C. After all the pellets in the furnace are melted and the liquid level of the melt drops, continue to add pellets until the furnace is finally filled with the melt, and then keep warm for 40 minutes. The crude lead is released from the metal discharge port at the bottom of the furnace, the vanadium-containing alkali slag is released from the slag discharge port, and the zinc-containing fume is collected by the bag dust collector during the smelting process. The resulting products are as follows:

1. Crude lead: lead in crude lead is 98.50%, lead reduction rate is 98.52%, silver in crude lead is 115g/t, and silver yield is 96%.

2. Vanadium-containing alkali slag: vanadium-containing alkali slag contains 25.45% vanadium pentoxide, and the slag is almost enriched with all the vanadium in the raw material. After the molten slag is condensed, it is crushed to a particle size of 0-8mm, and wet-milled in a ball mill. Transfer the ball-milled pulp into the leaching tank, add water to control the liquid-solid ratio to 4:1, the leaching time is 90min, and the leaching temperature is 90°C. After leaching, filter and wash the filter cake with clean water. The resulting leaching residue contains 0.39% V ₂ O ₅ , and the leaching rate of vanadium is 99.3%. The obtained vanadium-containing leaching solution can be treated by conventional vanadium metallurgy process.

3, soot, the reducing evaporation rate of present embodiment zinc is 81%, and gained soot contains zinc 53.6%, lead 9.1%.

Example 4:

Extract vanadium, lead and zinc from vanadium-lead ore containing Pb38.84%, _V2O510.34 %, _Zn11.23 %, Fe13.62%, _SiO22.94 %, Ag47g/t. The implementation process is as follows: grind the vanadinite ore to a particle size of ≤0.18mm, and then add flux sodium carbonate and reducing agent coke powder. The ratio is: vanadinite powder: sodium carbonate: coke powder = 100:28.0:8.5 (the amount of sodium carbonate added is 1.2 times the theoretical amount, the amount of reducing agent added is 1.2 times the theoretical amount), coke The powder contains 83% fixed carbon. The accurately weighed material is uniformly mixed in the mixer for 30 minutes, and the mixed charge is directly sent to the reverberatory furnace for reduction smelting. The hearth area of the reverberatory furnace is 5 square meters, and the ratio of the hearth to the furnace area is 5:1. The reverberatory furnace smelting is intermittent smelting, the feed temperature is 700-900°C, and the temperature is gradually raised to 1250°C during the melting stage, and the high-temperature melting is maintained until the charge is completely Melting, heat preservation and precipitation stage temperature is maintained at 1100 ~ 1150 ℃, heat preservation and precipitation time is 60 minutes, the crude lead is released from the metal discharge port at the bottom of the furnace, the vanadium-containing alkali slag is released from the slag discharge port, and the zinc-containing fume is released during the smelting process collected by a bag filter. The resulting products are as follows:

1. Crude lead: lead in crude lead is 98.27%, and the reduction yield of lead is 98.38%. The silver in crude lead is 115g/t, and the silver yield is 96%.

2. Vanadium-containing alkali slag: the obtained vanadium-containing slag contains 23.36% of vanadium pentoxide, and the slag is almost enriched with all the vanadium in the raw material. After the molten slag is condensed, it is crushed to a particle size of 0-8mm, and wet-milled in a ball mill. Transfer the ball-milled pulp into the leaching tank, add water to control the liquid-solid ratio to 3:1, the leaching time is 90 minutes, and the leaching temperature is 90°C. After leaching, filter and wash the filter cake with clean water. The resulting leaching residue contains 0.58% V ₂ O ₅ , and the leaching rate of vanadium reaches 98.7%. The obtained vanadium-containing leaching solution can be treated by conventional vanadium metallurgy process.

3. Smoke and dust: The reduction evaporation rate of zinc in this example is 82.03%, and the obtained smoke contains 50.32% zinc and 8.83% lead.

Example 5:

Extract vanadium, lead and zinc from vanadium-lead ore containing Pb45.2%, V ₂ O ₅ 6.88%, Zn8.6%, Fe12.5%, SiO ₂ 5.8%, Ag65g/t. The implementation process is as follows: grind the vanadinite ore to a particle size of ≤0.18mm, and then add flux sodium carbonate and reducing agent coke powder. The ratio is: vanadinite powder: sodium carbonate: coke powder = 100:44.6:8.3 (the amount of sodium carbonate added is 2.0 times the theoretical amount, the amount of reducing agent added is 1.5 times the theoretical amount), coke The powder contains 83% fixed carbon. Mix the accurately weighed materials uniformly in the mixer for 30 minutes. After mixing evenly, press the mixture into pellets. The diameter of the pellets is 30mm, and the molding pressure is 20Mpa. The moisture content of the pellets is less than 2% after natural drying. The pellets are sent into a melting furnace for reduction smelting. The melting furnace used in this embodiment is a reduction electric furnace with a transformer rated capacity of 400KVA and a controlled melting temperature of 1150°C. After all the pellets in the furnace are melted and the liquid level of the melt drops, continue to add pellets until the furnace is finally filled with the melt, and then keep warm for 25 minutes. The crude lead is released from the metal discharge port at the bottom of the furnace, the vanadium-containing alkali slag is released from the slag discharge port, and the zinc-containing fume is collected by the bag dust collector during the smelting process. The resulting products are as follows:

1. Crude lead: lead in crude lead is 96.80%, lead reduction rate is 98.9%, silver in crude lead is 138g/t, and silver yield is 95%.

2. Vanadium-containing alkali slag: Vanadium-containing alkali slag contains 17.5% vanadium pentoxide, and the slag is almost enriched with all the vanadium in the raw material. After the molten slag is condensed, it is crushed to a particle size of 0-8mm, and wet-milled in a ball mill. Transfer the ball-milled pulp into the leaching tank, add water to control the liquid-solid ratio to 5:1, the leaching time is 90min, and the leaching temperature is 90°C. After leaching, filter and wash the filter cake with clean water. The resulting leaching residue contains 0.26% V ₂ O ₅ , and the leaching rate of vanadium is 99.2%. The obtained vanadium-containing leaching solution can be treated by conventional vanadium metallurgy process.

3, soot, the reducing evaporation rate of present embodiment zinc is 81.5%, and gained soot contains zinc 51.5%, lead 5.8%.

Claims (10)

1. the method for selective separation vanadium, lead and zinc from a vanadinite is characterized in that, becomes the vanadinite fine grinding behind the breeze and an alkali metal salt flux, carbonaceous reductant mix and obtain furnace charge; The furnace charge that obtains is made ball be placed in the smelting furnace, 1100～1250 ℃ of following meltings; With the slag water that obtains after the melting 85～90 ℃ leach vanadium down after, filter, obtain vanadium leachate; Wherein, zinc is the gaseous state volatilization in the fusion process, and lead is in a liquid state and sinks to furnace bottom, and vanadium remaines in the slag with the form of basic metal vanadate.

2. the method for claim 1 is characterized in that, described an alkali metal salt flux is Na ₂CO ₃Or K ₂CO ₃In one or both mixing.

3. method as claimed in claim 2 is characterized in that, the add-on of described an alkali metal salt flux is that vanadium, silicon and the phosphorus in the vanadinite reacts the corresponding M of generation respectively ₃VO ₄, M ₂SiO ₃, M ₃PO ₄1.2～2.0 times of required theoretical molar amount, wherein, M is Na or K.

4. the method for claim 1 is characterized in that, described carbonaceous reductant is coke.

5. method as claimed in claim 4 is characterized in that, the add-on of described reductive agent is PbO, ZnO and the Fe in the vanadinite ₂O ₃Reduction reaction generates corresponding Pb, Zn and the required theoretical molar amount of FeO 1.2～1.5 times respectively.

6. the method for claim 1 is characterized in that, the liquid-solid volume ratio of described leaching water and slag is 3～5:1.

7. method as claimed in claim 6 is characterized in that, described extraction time is 75～100min.

8. the method for claim 1 is characterized in that, after the described vanadinite fine grinding, granularity is not more than 0.18mm.

9. the method for claim 1 is characterized in that, the granularity of described carbonaceous reductant is not more than 0.18mm.

10. as each described method of claim 1～9, it is characterized in that the zinc in the described fusion process in the vanadinite is the back form with flue dust of gaseous state volatilization and is recovered, liquid state sinks to furnace bottom lead and reclaims with the lead bullion form, vanadium remaines in the slag with the form of vanadate, and water leaches.