JP2861483B2 - Dry smelting of zinc and lead - Google Patents

Description

DETAILED DESCRIPTION OF THE INVENTION

[0001]

BACKGROUND OF THE INVENTION 1. Field of the Invention The present invention relates to the simultaneous smelting of zinc and lead.

[0002]

2. Description of the Related Art The zinc smelting method is roughly classified into a wet method and a dry method. However, in either method, zinc sulfide concentrate, which is a main raw material, is roasted to be zinc oxide. afterwards,
In the wet method, zinc is recovered by acid leaching and electrolytic recovery. In the dry method, a treatment such as charging the zinc oxide together with coke into a furnace is performed, and the zinc is reduced and volatilized and recovered.

The only practical method of the wet method is the electrolytic purification method. However, this method has to take as gentle a roasting method as possible and usually uses a fluidized roasting furnace. For this reason, zinc concentrate having a high lead content cannot be used for briquetting, and when the obtained zinc oxide is leached, impurities such as copper, cobalt, nickel and cadmium are leached. Therefore, these impurities must be removed prior to the electrolytic recovery of zinc.

[0004] Dry methods include a horizontal distillation method, a vertical distillation method, an electrothermal distillation method and an ISP method. The horizontal distillation method is a batch type, and therefore has almost disappeared in the late 1970's due to its extremely labor-intensive, poor working environment, and small scale merits.

[0005] The vertical distillation method utilizes the principle of the horizontal distillation method, and improves the poor productivity which is a drawback of the horizontal distillation method. However, this method uses an externally heated vertical furnace, so that 200-300 t-Zn /
This method is also used for zinc smelting because of the maximum processing capacity of the moon and furnace, the complexity of the process, and the need to process high-lead copper-bearing slag generated from the furnace. lost.

[0006] The electrothermal distillation method has a processing capacity of 1000 to 3000 t-Zn compared to the above two methods.
Although it is as large as / month / furnace, it has disadvantages such as long pretreatment for obtaining a sintered mass to be charged into the furnace, and limitation of reduction in power consumption per unit due to the electric heating furnace. For this reason,
This approach is becoming less common in areas with high electricity rates.

On the other hand, the ISP method has a feature that zinc and lead are smelted simultaneously, and is the mainstream of the dry method. In this method, zinc sulfide concentrate is roasted together with lead concentrate or zinc concentrate containing lead and sintered to obtain sintered ore having sufficient strength, and reoxidation rich in carbon dioxide gas. Despite the easy atmosphere, the development and adoption of a technology to condense zinc with a molten lead splash condenser at a high temperature of 1000 ° C or higher while containing gas containing zinc vapor has resulted in a production volume per furnace of 6,000 to 6,000.
It has been increased to 10,000 t-Zn / month / furnace.

Although the ISP method has advantages such as productivity, thermal efficiency, and variety of processing materials, the ISP method requires about four times the ore in the roasting and sintering process in order to obtain a sintered mass to be supplied to the blast furnace. In addition, the operation of the roasting and sintering process requires skill, and a blast furnace requires the use of expensive massive coke.

[0009]

SUMMARY OF THE INVENTION The present invention relates to a dry process for recovering metallic zinc and / or metallic lead from sulfide concentrate at a low cost without using a roasting and sintering process for zinc concentrate in the ISP method. The purpose is to provide a refining method.

[0010]

In order to solve the above-mentioned problems, the present invention provides a dry smelting method for obtaining zinc and lead from a sulfide concentrate containing zinc sulfide, lead sulfide, and iron sulfide. In the oxidation step, iron / silicate slag or lime containing iron / silicate slag is formed or supplied in an oxidation furnace, and the sulfide concentrate and non-combustible material are contained in the slag. And the flux, and any one of industrial oxygen, oxygen-enriched air, and air are blown and allowed to react, whereby most of zinc in sulfide concentrate and non-combustibles and a part of lead are converted into slag. And the temperature of the slag is 115
0 to 1300 ° C, Fe / SiO ₂ of 0.70 to 1.4
6, forming a slag of 15% by weight or less of CaO, 15 to 25% by weight of Zn, and 0.5 to 3% by weight of S and a mat and / or metal composed of a part of lead in the raw material; In the reduction step, heavy oil is added to the slag obtained from the oxidation step,
Pulverized coal, and blowing a reducing agent selected from coke fines, volatilize zinc and lead in the slag, and then condense to obtain molten zinc and molten lead. Provide a smelting method.

In the present invention, preferably, valuable substances containing zinc and lead in the gas generated in the oxidation step are recovered as non-natural substances, and the non-self-combustible substances are returned to the oxidation step, and the non-combustible substances are returned to the reduction step. A part of the residual slag is cooled as it is or solidified, and then pulverized and used as the slag for the oxidation furnace.

Further, the raw materials are prepared so that the total amount of zinc in the raw materials supplied into the oxidation furnace is larger than that of lead,
Then, it is preferable to blow one of oxygen, oxygen-enriched air, and air into the mat and / or the metal to lower the S quality.

[0013]

In the present invention, after sulfide concentrate and non-combustibles (smoke ash) are dissolved in slag in the oxidation step, zinc and lead are volatilized in the reduction step and are recovered as molten zinc and molten lead. Mats and metals generated in the oxidation process are separated and recovered from the slag, and non-combustible materials are returned to the oxidation process.

The oxidation step and the reduction step may be performed in one furnace, or may be performed using two furnaces corresponding to each step. The gas used for the reaction in the oxidation step may be any of industrial oxygen, oxygen-enriched air, and air.

The sulfide concentrate used as a raw material contains Fe and SiO.
₂ is included, and the flux is adjusted in consideration of this in order to obtain a slag of a target composition. However, the amount of flux obtained in this way cannot usually absorb all of the zinc in the concentrate. Therefore, a part of the slag corresponding to the amount of zinc in the concentrate must be newly supplied to the furnace.
The most suitable slag as the supply slag is slag after Zn and Pb are reduced and volatilized in the reduction step of the present invention. After solidifying, it may be pulverized and blown into the slag together with the raw material. Of course, the above-mentioned slag amount can be ensured by increasing the amount of the flux containing the slag component.

In the present invention, iron, silicate slag or lime containing iron, silicate slag or lime is used as the slag because the raw material contains a relatively large amount of iron sulfide and SiO ₂ as described above. This is because it is advantageous for the formation of slag and the melting point of slag can be reduced by CaO and the volatilization rate of Zn in the reduction step can be increased.

When the temperature of the slag is low, the reactivity between the concentrate blown into the slag and the slag is sharply reduced, and a large amount of unmelted material is generated in the furnace.
Not only that, most of the Zn also becomes fumes, which are non-combustibles, and scatters outside the furnace, increasing the amount of fumes returned to the furnace and significantly reducing the smelting efficiency. Therefore, in the present invention, the temperature of the slag is set to 1150 to 1300 ° C.

Fe / SiO ₂ in the slag is related to the quality of the magnetite in the slag and the melting point of the slag.
When O ₂ is less than 0.7, the magnetite grade is lowered but the melting point is 1300 ° C. or higher. When it exceeds 1.46, the melting point of slag is lowered, but the magnetite grade is raised and precipitated from the slag layer, It may accumulate on the furnace bottom and push up the furnace bottom.

When the CaO grade exceeds 15% by weight (hereinafter referred to as “%”), Fe / SiO ₂ becomes 0.7%.
In any range from 0 to 1.46, the melting point of the slag is increased. Therefore, the CaO grade must be 15% or less. Note that CaO is present in the concentrate or in the smoke ash in a small amount, so that the CaO grade in the slag cannot be reduced to 0%.

By the way, the concentration of Zn in the concentrate is usually 50%.
It is about. Therefore, lowering the zinc quality in the slag increases the amount of treated slag in the reduction furnace accordingly. The lower limit of the zinc quality in slag is a matter of production efficiency. Usually, the range that can be tolerated is about 3 to 4 times the raw material, and in consideration of this, the Zn grade in the slag must be 15% or more. In the slag system of the present invention, the melting limit of Zn is about 25%, and does not substantially exceed 25%.

The reason why the S grade in the slag is set to 0.5 to 3% is that when the S grade is lower than 0.5%, the amount of magnetite in the slag rises remarkably and precipitates from the slag layer, and If it exceeds 3%, although the precipitation of magnetite can be prevented, sulfur volatilizes in the reduction step,
This is because when mixed with the gas and condensed by the condenser, it reacts with Zn to form ZnS, and the ZnS precipitates and solidifies at the condenser entrance, which hinders the operation. In order to avoid such adverse effects more reliably, the S grade must be 1-2.
% Is desired.

When a raw material containing Pb is used and a reaction gas is blown to generate the slag, a part of Pb present in the raw material becomes mat and / or metal. These mats and metals are higher in S quality than those obtained by the ISP method, and as such, it is not possible to obtain metal lead by direct electrolytic treatment and purification. Therefore, it is necessary to react the mat and / or the metal with the oxidizing gas until the metal lead has a low sulfur grade enough to be directly electrorefined. This oxidation treatment may be performed in the oxidation furnace in parallel with the oxidation of the concentrate, or the mat or metal may be extracted from the oxidation furnace, and the mat or metal may be oxidized in another furnace. When the oxidation treatment is performed in accordance with the former, the oxidizing gas must be directly blown into the mat or metal layer so as not to contact the slag layer.

In the exhaust gas generated by the reaction, Zn, Pb and the like are present in the form of oxides and sulfates. Therefore, these must be collected as smoke (non-combustible). The device for recovery is not particularly limited, and may be a general electric dust collector, a back filter, or the like.
Since the recovered smoke ash generally has a high S grade, it is inappropriate to return it to the reduction furnace, and returns it to the oxidation furnace. In the return of the smoke ash, the smoke ash may be mixed with the concentrate, or may be supplied to the furnace in a separate system from the concentrate. The oxidizing gas used may be any of industrial oxygen, oxygen-enriched air, and air.

[0024] In the slag produced from the oxidation step, most of Zn and part of Pb in the concentrate are dissolved mainly in the form of oxides. In order to recover zinc and lead from this slag, the slag is reduced with a reducing agent, and Zn and Pb are recovered.
Need to be reduced and volatilized and condensed. The reduction of the slag is basically the same as the slag fuming method, and heavy oil, pulverized coal, coke, reducing gas and the like can be used as a reducing agent. And, as described above, the oxidation treatment is first performed using one furnace, the mat or the metal is extracted, and then the remaining slag cannot be used for reduction treatment, and the two furnaces can be used in one furnace. Oxidation treatment,
The slag reduction treatment may be performed in another furnace.

The exhaust gas generated by the reduction contains Zn and P
b is present as a metal vapor. For this reason, using the Pb splash condenser used in the ISP method,
It is preferable to recover the vapor of n or Pb. The Zn and Pb thus recovered are treated according to the usual ISP method. On the other hand, a part of the slag which has been reduced and volatilized is returned to the oxidation step as it is, or cooled and solidified, then pulverized, mixed with the raw material, or blown into the oxidation furnace alone.

In general, Pb is more likely to become smoke than Zn. Therefore, if the proportion of Pb in the raw material is too high, smoke ash increases, and the amount of adhesion to the exhaust heat boiler increases, making it difficult to operate exhaust gas treatment equipment. In order to prevent this, it is preferable that the total amount of Zn charged in the oxidation furnace is larger than the total amount of Pb, and more desirably, the total amount of Zn is at least twice the total amount of Pb. Is preferred.

[0027]

The present invention will be further described with reference to the following examples.

First, a test smelting furnace for performing the dry smelting method of the present invention will be described.

The test smelting furnace shown in FIG. 1 has a height of 2.8.
m, a reaction tower 10 having an inner diameter of 1.5 m, and a settler 12 having a length of 5.25 m and an inner diameter of 1.5 m. One end of the settler 12 is connected to the reaction tower 10, and the other end is connected to the flue duct 14. Have been.

A first lance 16 having a diameter of 2.5 cm is inserted from above the reaction tower 10. First lance 1
An oxygen / raw material mixer 18 for mixing oxygen and a raw material is connected to 6, and a raw material transfer facility 20 is connected to the oxygen / raw material mixer 18.

On the opposite side wall of the settler 12, oxygen
A heavy oil burner 22 and a heavy oil burner 24 for heat insulation are provided.

A slag hole 26 is provided below the heat retaining heavy oil burner 24 so that the slag 28 can flow out.

Mats and / or deposits under the slag 28
Alternatively, a tap hole 32 for extracting the metal 30 is provided in a part of the side wall of the settler 12.

In the test smelting furnace shown in FIG. 2, a second lance 40 for blowing coke breeze is provided at the upper center of the settler 12 of the smelting furnace shown in FIG.

A coke breeze feeder 42 for storing coke breeze for reducing slag and maintaining the furnace temperature at a target value is connected to the first lance 16 and the second lance 40 via a distributor 44. doing. A slag hole 48 for letting out the slag 46 is provided on a part of the side wall of the settler 12. The test smelting furnace shown in FIG.
There is no heavy oil burner.

Using the raw materials having the compositions shown in Table 1 and the conditions shown in Table 2, the test smelting furnace shown in FIG.
Was tested. In the test, the charged ore, auxiliary fuel, and oxygen-enriched air prepared under the specified conditions were first blown into the reactor from the top of the reactor in the same manner as in a normal flash smelting furnace, and the molten slag was discharged. Manufactured.

Next, as shown in FIG. 1, a blowing lance having a diameter of 2.5 cm from the upper part of the reaction tower was placed 30 mm from the upper surface of the slag.
The mine was placed so that the injection port was located at a position of 1 cm, and the charged ore was blown into the slag together with 70% oxygen-enriched air from this lance. The heat required for the ore melting and the heat dissipated from the settler were compensated for by using a fuel oil burner installed on the side wall opposite to the settler. In addition, 70% oxygen-enriched air was used as combustion air for the heavy oil burner on the reaction tower side, and air was used for the heavy oil burner on the slag hole side.

The ore charged was prepared by drying each concentrate, smoke ash, and flux shown in Table 1 and mixing them according to Table 2. In determining the mixing ratio, the concentrate processing amount was set to 300 Kg / Hr, and the amounts of smoke ash, flux, heavy oil, and oxygen were adjusted so that the intended operation was possible.

The produced slag was withdrawn from the slag hole 26 shown in FIG. 1 into a ladle approximately every four hours, and the temperature was measured and a sample for X-ray fluorescence analysis was collected in the first and second half of the withdrawal. The mat and / or metal was extracted when it became possible to extract it from the tap hole 32, and about 0.5 t was discharged each time, and a sample for analysis was collected at the same time. The mat or metal was confirmed by inserting a measuring rod into the melt from a measuring hole provided on the settler ceiling, pulling it up, and checking the state of the melt adhered to the measuring rod.

The obtained results are shown in Tables 3 and 4. Although the output is all intermittent, the slag interval is 3 ~
Relatively short, 4 hours, and 1.6 emissions per time
Since it was as large as ~ 2.0 t, evaluation could be made with fairly good accuracy.

The smoke ash was continuously collected by a dust chamber and an electric dust collector, discharged out of the system every day, and weighed. Therefore, there is no problem in the accuracy of the smoke ash amount.

However, since the mat amount cannot be extracted unless it reaches a certain amount and cannot be completely discharged, the measurement accuracy is poor.

Further, since the metal cannot be extracted in the same amount as the mat, the adherence to the measuring rod and the solidification of the mat were examined, and the bottom of the ladle was inspected to determine whether or not the metal had formed.

The following is a description of each embodiment shown in Tables 2 and 3.

[0045]

[Table 1]

[0046]

[Table 2]

[0047]

[0048]

[Table 3]

[0049]

Example 1 In Example 1, the slag temperature was set to 1
250 ° C., 1.5% S grade, 0.1% Fe / SiO ₂ .
9. The operation was carried out after mixing and preparing CaO at 5% and Zn at 20%, and slag almost as intended was obtained. In this example, although a small amount of mat and smoke ash were obtained, the generation of metal could not be confirmed.

[Embodiment 2] This embodiment is intended to lower the Cao quality of the slag in the embodiment 1.
The charging of the flux E was stopped. Since the charged amount of flux A was smaller than the target and the charged amount of concentrate A was slightly increased, the slag temperature increased by 10 ° C. and the S grade became 2.6%. And flux A originally contains CaO
Was contained at 2.4%, so that the CaO grade in the slag was reduced only to 1.5%. From this result, it was found that the concentrate could be treated without substantially adding CaO. Note that, when viewed as a whole, it can be determined that the same operation results as those of the first embodiment are obtained in the second embodiment.

[Embodiment 3] In this embodiment, the CaO grade is 1
This is an operation when the temperature is increased to 5%. Expecting a decrease in the melting point of slag due to CaO, the target slag temperature is set to 1250.
To 1180 ° C. In operation,
Since the addition amount of the flux E was increased, the fuel oil combustion amount in the fuel oil burner on the reaction tower side was increased to 28 l / h.

Although there was no problem in discharging the slag, the amount of Zn and Pb dissolved in the slag decreased, and the quality of magnetite also increased. As a result, a spatter, which is a magnetite-rich semi-solid, was formed between the slag and the mat. Further, the Zn grade in the slag was 15.0%. In this example, generation of lead metal was confirmed.

Subsequently, when the CaO grade was increased to 20%, the magnetite grade further increased by about 3%, the melting point of the slag also increased, some of the slag solidified, and the pool in the settler began to decrease. . In addition, when the slag was withdrawn, the slag swelled on the gutter, making it difficult to discharge. From this result, the CaO grade must be 15% or less.

[Embodiment 4] In this embodiment, the slag CaO
It was intended to eliminate spoilers at about 15% quality. That is, the flux A is decreased and the flux D is increased to reduce the Fe / SiO ₂ from 0.9 to 0.7.
It has been reduced to. Since a significant increase in the melting point of slag was expected due to a decrease in Fe / SiO ₂ , the target slag temperature was set to 1300 ° C.

As a result, the spatter disappeared and the magnetite grade in the slag decreased by 2.5%.
The grade remained at 15% and most of the Pb in the feed was distributed to smoke ash. As described above, when Fe / SiO ₂ is 0.7
Below, the slag temperature must be increased, and as a result, Zn and Pb tend to volatilize, and the tendency is high Ca
It can be seen that the O grade is more remarkable. Therefore, Fe
/ SiO ₂ must be 0.7 or more.

Example 5 Next, in order to operate at a low CaO grade, the addition of flux E was stopped, the CaO concentration was kept low, and the operation was performed with Fe / SiO _{2 set} to 0.7.
In Example 5, despite the high slag temperature of 1273 ° C., the absorption of Zn and Pb into the slag was good, and slag of 19.6% and 4.9% quality was obtained, respectively. . As a result, smoke ash was significantly reduced. The magnetite grade was low due to the low lime grade. Nevertheless, there was a slight rise in the hearth.
Therefore, it is necessary to set the slag temperature to 1300 ° C. or higher in order to continue stable operation under this condition, which is not practical. Therefore, from this point, Fe /
It can be said that SiO ₂ needs to be 0.7% or more.

[Embodiment 6] In this embodiment, concentrate B is used instead of concentrate A, and the feature is that the amount of Pb is small. In this embodiment, the target slag temperature is 1170 ° C.
And Although spoilage occurred between the slag and the mat, build-up (swelling) of the furnace bottom and removal of the slag could be performed without any trouble. However, since the slag temperature was as low as 1167 ° C., the flammability of the concentrate was slightly poor, and a small amount of unmelted lumps was confirmed on the slag, but there was no obstacle to the operation. After the extracted mat was solidified in the ladle, it was taken out of the ladle and confirmed, and metal was confirmed.

Under these conditions, when the slag temperature was lowered to 1145 ° C., a large amount of unmelted material was found under the blowing lance. Therefore, the slag temperature needs to be 1150 ° C. or higher.

[Embodiment 7] This embodiment is a continuation of Embodiment 6. As described above, after lowering the slag temperature to 1145 ° C and confirming the unmelted material, charging of the flux E was stopped, and the slag temperature was increased to about 1260 ° C. Also disappeared. In this example, metal was produced together with the mat, but smoke ash was reduced. In this example, a Zn grade in the slag of 25.1% was obtained, which was the highest Zn grade obtained in a series of test runs. Therefore, the upper limit of the grade of Zn in the slag is set to 25%.

Example 8 In order to further reduce the load of Pb, charging of smoke ash A was stopped, and flux B was used instead of flux A. As a result, the load of Pb was reduced, and the quality of Zn in the slag was significantly reduced due to the increase in the charged amount of the flux. However, there was no change in the furnace maneuverability such as raising the furnace bottom and extracting slag.

In this embodiment, the P contained in the raw material
Since the amount of b was small, no mat or metal was formed. Therefore, it can be seen that under special raw material conditions, only slag may exist in the furnace. But,
Generally, it is rare that the melt in the furnace is only slag because the ash containing Pb generated in the oxidation step is repeated in the oxidation step.

At the end of this example, when the amount of oxygen-enriched air for concentrate was increased and the S grade in the slag was gradually lowered, the magnetite grade in the slag was 0.4% at 0.4%.
8.3%, a large amount of spoilage occurred, and a sharp rise in the furnace bottom was observed. This indicates that the S quality in the slag needs to be 0.5% or more.

Example 9 In this example, the same kind of raw material as in Example 8 was used, the S grade in the slag was maintained at about 1%, and the Fe / SiO ₂ was adjusted to 1.5. did. As a result, when the S grade was 1.1% and the Fe / SiO ₂ was 1.46, the magnetite grade was 16.4%, and the same phenomenon as when the S grade was 0.4% was observed. . This indicates that Fe / SiO ₂ must be 1.46 or less.

Example 10 In this example, the smoke ash B produced in Example 8 was charged, and concentrate B and fluxes B, D, E
And a test operation was performed using Example 10
In Example ₂ , the S grade in the slag was 2.7, and Fe / SiO ₂ was 0.89, and the same operation as in Example 8 was possible. Thus, it can be seen that the treatment of smoke ash containing oxides and sulfates is also possible.

[Example 11] In this example, concentrate A, smoke ash produced in Example 1 (smoke ash C), slag (flux C) obtained after the reduction test described later, fluxes D and E were performed.
And processed together. From Tables 3 and 4, it was found that there was no operational problem even if smoke ash C and flux C were used. Therefore, it can be seen that most of the slag after reduction and volatilization can be repeated in the oxidation step. In this example, the slag after reduction and volatilization was solidified and pulverized for use.
If this is repeated in a molten state, it is estimated that energy costs can be significantly reduced.

Example 12 A second lance 40 for blowing coke breeze into the center of the settler part of the furnace used in Example 1 was used.
2, the slag obtained in Example 1 is solidified and pulverized, and the slag is charged into the raw material transfer facility 20, a predetermined amount of slag powder is cut out, and the slag is flown by air. Reaction tower 10
I blew it down. The coke breeze for reducing the slag and maintaining the temperature in the furnace at the target value is cut out from the coke breeze feeding equipment (injection tank) 42, fed with air, and sent to the first lance 16 by the distributor 44.
Most of the coke breeze was blown into the lower portion of the reaction tower together with the slag powder.

The remainder of the coke breeze was blown into the settler 12 from the second lance 40. Then, industrial oxygen was supplied into the furnace together with slag powder and coke breeze from the first lance 16 provided in the reaction tower 10.

The slag temperature in the furnace was maintained at 1300 ° C., and the CO ₂ / CO ratio in the exhaust gas was adjusted to 0.5.
A test run of time was performed. The reduced and volatilized Zn and Pb are reacted by blowing air in an exhaust gas treatment facility for convenience.
O and PbO were collected. CO in gas is CO
Detoxified as ₂ . Table 4 shows the operating conditions and the results obtained.
It was shown to.

Table 4 shows that the slag obtained from the oxidation furnace
It turns out that n and Pb can be reduced and volatilized. Therefore, it is clear that Zn and Pb can be recovered as metal by using a capacitor used in the existing ISP method.

[0071]

[Table 4]

[0072]

According to the method of the present invention, the roasting and sintering step of zinc fine required in the conventional ISP method can be omitted, Zn and Pb can be simultaneously recovered as metal, and reduction can be achieved. Inexpensive coke breeze can be used as an agent.

[Brief description of the drawings]

FIG. 1 is a schematic sectional view of a smelting furnace used in Examples 1 to 11 of the present invention.

FIG. 2 is a schematic sectional view of a smelting furnace used in Example 12 of the present invention.

[Explanation of symbols]

 DESCRIPTION OF SYMBOLS 10 Reaction tower 12 Settler 16 Injection 1st lance 18 Oxygen / raw material mixer 20 Raw material feeding equipment 22 Oxygen / heavy oil burner 24 Heating heavy oil burner 26, 48 Slag hole 42 Coke breeze feeding equipment

Continuation of front page (58) Field surveyed (Int. Cl. ⁶ , DB name) C22B 19/20 C22B 13/02

Claims (5)

(57) [Claims] 1. A vulcanization zinc, lead sulfide, zinc from sulphide concentrate containing iron sulfide, in pyrometallurgical process to obtain a lead has an oxidation step and the reduction step, the oxidation step, the oxidation furnace iron silicate - Tosuragu or iron-silicate containing lime - forming a Tosuragu, or supplies, said sulfide concentrate in said slag, and a non-self-combustibles,
The flux is blown with any one of industrial oxygen, oxygen-enriched air, and air to react, thereby melting most of zinc in sulfide concentrate and non-combustibles and part of lead in the slag. And the temperature of the slag is 1150-1300 ° C, F
e / SiO ₂ is 0.70 to 1.46, CaO is 15% by weight or less, Zn is 15 to 25% by weight, and S is 0.5 to 3% by weight, and is composed of a part of lead in the raw material. Mat and / or
Or forming a metal, and in the reducing step, a reducing agent selected from heavy oil, pulverized coal , and coke fines is blown into the slag obtained from the oxidizing step to volatilize zinc and lead in the slag. then pyrometallurgical method of zinc及beauty lead, characterized in that to obtain the molten lead and the molten zinc by condensing. 2. The zinc according to claim 1, further comprising a step of recovering valuable substances containing zinc and lead in the gas generated in the oxidation step as non-combustible substances, and returning the non-combustible substances to the oxidation step. >及beauty pyrometallurgical method of lead. Wherein the reduction step as a part of in the residual slag, or after cooling and solidifying, and pulverizing, the further the step of using a slag for oxidizing furnace including claim 1, wherein the zinc及beauty Dry smelting method for lead . 4. A dry zinc及beauty lead of claims 1 to 3, wherein the total amount of zinc in the raw material to be supplied to the oxidation furnace is characterized by compounding the raw materials to be larger than that of lead Smelting method. 5. A mat and / or oxygen into the metal, oxygen-enriched air, blown into any one of air, claims 1, characterized in that reducing the S grade 3 wherein zinc及beauty Dry smelting method for lead .