JPS60187635A - Collection of metal valuables from substances containing tin and zinc - Google Patents

Abstract

(57) [Summary] This bulletin contains application data before electronic filing, so abstract data is not recorded.

Description

DETAILED DESCRIPTION OF THE INVENTION The present invention is a method of recovering metallic valuables from a substance containing tin and/or zinc by smelting it under oxidizing conditions and reducing the resulting melt. Regarding how to.

The present invention relates to working with all types of starting materials from which metals can be recovered by said method. The materials include primarily lead starting materials containing sulfides, oxides, sulfates and carbonates, and mixtures thereof. Lead starting materials can include mineral concentrates, intermediates and/or waste. Said materials can also be processed by direct lead smelting processes, including soot and!
11 contains at least one lead and other materials derived from various engineering and all operations such as slime, slag and dust.

Many lead smelting processes proposed in recent years consist essentially of an oxidative smelting step and a subsequent reduction of the resulting molten oxide bath.

Therefore, it can be said that all of the methods that belong to the so-called direct lead smelting method, and that produce low-sulfur-containing molten lead water and high-lead Venusian slag, belong to the above-mentioned group of smelting methods. OuLo
kumpu f (e.g. D'E -C-1,179,
004), Com1nco method (IJsA-3,84
7,595), Sj, Josepl+button manufacturing method (,I
Metals, I (12) 26-30. 1969
), Worcra method (US-A-3, 326, 67
1), [1vccL method (US-8-3,555,1
6, No. 1) and Q S-method (tJ S-A-3,
No. 941,587) all belong to this group.

Other lead smelting methods, including smelting reduction, are described in early patent specifications of Boliden, LJS-8-4.017,308 and LJ 5-A-4,008,075, supra. This invention relates to a method for producing metallic lead from oxide and/or sulfate or sulfide materials using a blown rotary converter as a smelting and reduction device. Early publications of Boliden, IE I)-8-0', +007.890 and E? , P-A-0, 00Ei, No. 832, a similar process described in No.
In particular, it relates to a process for producing from those having a high copper and/or arsenic content.

A common feature of these early r3o1iden processes is that the lead is produced in two steps. At the beginning of these stages, the lead starting material and flux are smelted by oxygen-fuel coal delivered to the surface of the material in the furnace, forming a sulfur-poor molten lead phase and a lead oxide-rich slag; The lead oxide content of the slag reaches 20-50%, usually 25-50%.

In the second stage above, coke or some other suitable reducing agent is added to the melting chamber to reduce its contents, the opening is heated and the converter is rotated.

Late Boliden's 1) Patent application SIE-A-
'8.302, 486-43 (this is E I) -Δ-
No. 0,124,497) describes a one-step process in which the reducing agent is charged to the converter together with the button starting material. This process can be considered as a process in which the oxidative smelting of the starting material and the reduction of the resulting melt are carried out simultaneously, and therefore this process also falls within the definition of the lead smelting process encompassed by the present invention.

Hitherto, it has not been possible to recover in one and the same furnace the zinc content of the starting material, in addition to the lead contained therein, and possibly other metals, such as tin.

The usual procedure is to further reduce the zinc present and evaporate it (fu
ming-off), the slag obtained during lead recovery is transferred to another furnace. The tin present is also evaporated (fum(!-of f).Slag fuming (slaH-fumir+B) and such a method is illustrated by, for example, Min, MrH!.

(August 1965) Vol, 113. No. 2
．． p, I14-122.

When the lead starting material is smelted and reduced by the method described in the introduction, the reduction process is very selective, even if the slag hold volume is about 1-2% However, only a small amount of zinc and the soot present evaporate (
fume-off), it is simply distilled. One contributing factor to this may be the presence of a concentration gradient in the melt and the lack of gas flow. This is because for process reasons at least a large portion of the zinc values The effort to recover in one and the same stage has created problems in such a process that the zinc-containing lead starting material can still be finished, although it is difficult.

In order to ensure that no zinc evaporates during the GJ smelting reduction stage, it is necessary to interrupt the reduction process at a content of less than 2-3%, which then takes place in another slag treatment stage. For example, Slugfu.

- This means that problems arise when dealing with slag due to mining. In this second slag treatment stage'(
, the residual lead components of the slag as well as all zinc and tin contained therein are evaporated by the reduction of the slag at high temperatures, and the lead, zinc and tin valuables are oxidized in the gas phase and then subjected to so-called mixed or crude oxidation. Collected in physical form. This mixed oxide requires an additional separation step to remove at least a majority of its lead component before the zinc and tin values can be worked up into useful zinc and soot η compositions.

Surprisingly, however, it has now been found that when the smelting process is carried out according to the method of the invention, which is characterized by the treatment steps indicated in the claims, lead, zinc and It has been found that tin values can be selectively recovered in one and the same furnace.

According to the invention, the starting material is smelted to form a melting ground containing a slag in which lead, zinc, tin, and small amounts of other noble elements are present in the form of oxides.

The required flux is at least 1 to 2 in the slag hold.
It is added in an amount suitable to give the slag a slBgish viscous gourd at the selected base temperature when lowered to a value below %. The melting ground is reduced with a solid carbon η reducing agent, such as coal or coke, which is introduced into the slag so that the suspension is 41%. As shown in the figure, the reducing agent is mixed into the inert (viscous) slag.
d) be done. The reducing agent stirs the slag vigorously (agi
tating or IW stirring, the zinc and tin are kept suspended in the slag for at least the latter part of the reduction period during which reduction of the zinc and tin occurs.

According to the present invention, zinc reduction occurs quickly when the slag hold volume is reduced to 1-2% due to lead reduction and formation of molten lead water. As a result of the action provided by the finely dispersed reducing agent in the slag, the zinc oxide in the slag is reduced to metallic zinc, which is vaporized as zinc vapor under the current reducing conditions and temperatures. The recovery procedure for any tin values present will be described in further detail below.

The soot can be removed before or after zinc reduction, or can be done without zinc reduction at all. If the tin is to be removed prior to zinc reduction, the coke and pyrite, or some other solid carbonaceous reducing agent and sulfur donor and/or chlorine donor, are added to the furnace in finely divided form. The slag is mixed with the slag and kept suspended therein by vigorously agitating or agitating the slag. This agitation of the slag reduces the zinc;
This is an important feature as previously disclosed.

The combination of a solid carbonaceous reducing agent suspended in the slag with a sulfur donor and/or a chlorine donor has the effect of reducing the tin component as well as volatile soot sulfide, SnS,
and/or promoting the production of tin chlorides such as \SllCIg and 5nCI4. The slag is jπ until the desired amount of tin is removed.
It remains in suspension at the soldering temperature.

Alternatively, tin is recovered as volatile tin(II) oxide by supplying only the reducing agent. tin oxide (n)
is evaporated in a separate step after zinc reduction. The reduction periods for zinc and tin may also overlap with each other.

Of course, it is also possible, and sometimes desirable, to retain at least a large portion of the tin content in the molten lead phase once formed, and to recover the tin component from the converter as a lead-tin alloy.

One i as to why zinc can be evaporated by the present invention
A possible explanation is the weak oxidation of slag against solid reducing agents 11
Effect result, reaction: CIO-”-C.

slagmer Carbon monoxide is produced in the slag.

The resulting -carbon oxide then reacts with +]O→-C
O fir Z n →-CO.

The slag reacts with the zinc oxide in the slag. As a result of the inertness of the slag, the carbon dioxide produced by this reaction has a fairly long residence time, which leads to the Bouthar reaction (B o
udouard-reac, tion) :CO2+C
; and 2CO 2 favoring the tendency to be reduced by the solid reducing agent present in suspension in the slag. The carbon monoxide thus also formed in the slag is further active in the reduction of zinc oxide.

The reduction of zinc is thus even more effective and maintains its viscosity even at high temperatures.
If a slag composition that is CoI is chosen, zinc can be eliminated from the slag with a substantial stoichiometric consumption of reducing agent (calculated in terms of the amount of carbon). .

The above discussion of possible zinc reduction mechanisms is also applicable if tin removal is carried out by the formation of volatile tin sulfide or tin chloride. However, tin oxide (I) is extracted from slag.
The removal of tin as I) is simply a direct evaporation step in contact with a reducing agent.

Since slags, especially silicate slags, can exhibit high viscosities with very different compositions, it is difficult to suggest a precise slag composition for implementing this method. However, in general, l, (within the range of 150 to 1.30 (1°C),
In some cases, a slag composition may be selected that maintains a high viscosity even at higher temperatures. When the slag is iron calcium silicate slag, the good effect is the following main analysis: 5iOz 20-30%, Ca(125
~35%, l'oll <25%, Oyobi MgO+AI
zO, 15 to IO%, is obtained with a slag composition of One suitable composition is SiO□about 25
%, CaO about 30%, Fen about 20% and MgO+8
It was observed that h036-8%.

Since lead oxide usually improves the fluidity of the slag, it is neither necessary nor desirable for the slag to be particularly viscous during the lead reduction step. The reduction of zinc and tin is approximately 2
It is therefore the composition of the slag after most of the lead has been reduced that determines the outcome of the zinc and tin removal steps.

The slag is suitably agitated in order to keep the fed reducing agent and sulfur donor and/or chlorine donor material in suspension with the slag. Although agitation can be carried out in many ways, for example by mechanical, pneumatic or electrical means, agitation of the slag by rotation of the furnace is particularly suitable. Therefore, this method
In order to be able to carry out the process in two and the same furnaces, the starting materials are preferably smelted in a rotary converter, for example of the Kaldor type. A further advantage of this type of rotary converter as a smelting and reducing apparatus is that it is particularly suitable for handling viscous sonog. Local blockages 6.2, such as solidified slag agglomerates or tough dough (do
ugby) There is no need to run the risk of breakage during operation of the furnace resulting in slag congealing or getting into the blades 11 and nozzles and blocking them.

When evaporating zinc and tin, the reduction temperature is 1.150
-1.25oC, but depending on the composition of the slag it may be necessary to use lower temperatures. In this latter case the kinetic nature of the reduction process would of course be impaired.

1. Temperatures as high as 300° C. and even higher can be used if the slag viscosity can be kept at a sufficiently high level.

However, the reduction of both lead, zinc, and tin is further enhanced by adding solid carbonate-containing materials to the slag in addition to the reducing agent. Carbonate material is also suspended in the slag to take advantage of its positive influence in the reduction process.

The process gas leaving the furnace contains tin sulfide (11
). The tin component of the gas can be recovered by any number of suitable methods. The soot sulfide initially present in the gas is oxidized to tin dioxide SnO□, which precipitates in the form of solid tf&granule dust. Similarly, the tin chloride present is also oxidized as shown here.

Arsenic and zinc present in the 01 process gas are also
1. As a result of oxidation, a solid fine-grained dust will be precipitated.

The oxide dust is suitably separated from the process gas by contacting the gas with circulating water during a venturi wash in circuit with a thickener. The circulating water is kept at a pl+ below 6, preferably between 2 and 3, so that the arsenic and zinc present in the water dissolve in it, but the tin settles out and produces an oxide slime, which separates from the water in the thickener. ^11 will be done. This separated tin slurry (J,) is filtered and S
An oxide product containing about 50% n is obtained and can be suitably used for the production of metal soot by any known reduction method. For example, slime can be smelted and reduced to produce crude tin, or it can be smelted with lead-containing materials to produce lead-
It can be recovered as a tin alloy. Both of these tin products can be readily purified by conventional methods to produce pure tin metal.

Next, the present invention will be described in more detail with respect to preferred embodiments thereof in the form of flowsheet diagrams and examples.

A lead starting material containing metal valuables such as Cu, Zn, Sn, and impurities such as As and S is fed together with flux to a Kaldor converter, where φj: oxidizing by supplying oxygen to the furnace. Smelting in an atmosphere produces a slag containing at least substantially all zinc and tin components of the smelted material. Depending on the composition of the smelted material, a molten lead phase may also be formed, at least to some extent. The process gas is diluted to a wet gas scrubbing stage, the dust is then separated and recycled as sludge to the smelting stage, and the sulfur dioxide content in the secondary gas is absorbed in a sulfur dioxide facility.

A reducing agent, e.g. coke, is fed to the converter in a star, denoted reduction I, which reduces lead in the slag phase in a first reduction stage. The converter is heated by gas heating during the reduction stage. The reduced lead forms a molten lead phase or is absorbed into the lead phase if already present in the converter. The process gas is directed to the same gas cleaning stage as described above.

During this first reduction step the sulfur donor substance and/or
Tin can also be recovered, at least in part, by feeding the chlorine donor material together with the reducing agent. In such cases, the process gas contains soot sulfide (11) and (
or) oxidized prior to the gas purification step to give oxidation of tin chloride to soot(IV) oxide, which is recovered as said oxide slime.

In the reduction stage 11 a reducing agent is supplied to reduce the zinc present in the slag. Outflowing reducible fluid 1:J
The process gas is combusted and the vaporized zinc metal is converted to zinc oxide, which is separated from the gas in a class 1 electric gas concentrator or wet gas purification facility. Alternatively, the reduction port may be implemented in such a way that the soot present in the slag is reduced, so that the soot is transferred to the molten lead phase.

In the reduction stage (2), the tin content in the slag present as tin oxide (1) is removed by supplying a solid reducing agent to the slag.In the presence of the reducing agent, tin oxide (IT) is removed. Evaporation is promoted. Evaporated tin oxide (I)
I) is passed into the exiting process gas (1'), where it is very rapidly oxidized to soot(IV) oxide, which is solid at practical temperatures and therefore contains tin-containing compounds in wet gas purification equipment. It is separated from the gas as an oxide sludge or slime. Alternatively, the tin component can be recovered as metallic tin absorbed in the molten lead phase by controlling the temperature and the supply of reducing agent.

After the reduction stage, a final slag with low metal content is obtained,
The final slag is discarded.

The resulting molten lead phase may contain impurities or monovalent elements such as arsenic and copper, and in actual cases a large proportion of tin components. Arsenic in the lead phase becomes arsenic after the supply of iron scrap.
It can be removed in the form of iron sorrel gas. Copper present in the lead phase oxidizes the lead phase! I-rugoto copper to I'J
It is also possible to remove ζ. Refined lead, which in practical cases can contain substantially the majority of the tin-containing T72 of the smelted material, is produced as the final product l).

Among the examples, I'b 59.3%, Zn 7.5%, Sn 1
．． 0%, 30.9%, l' e l, 11%, 5i
24 tons of lead concentrate ζ1' containing n2181z (h 3.7% and C3, 8% (present as carbonate), 1%, Zn 8.3%, Snl, 0%
, S3.5%, Fe 1.2%, 5iOz 4 Alz
A mixture consisting of 6 tons of other minerals containing 2% O and 26% C4 (present as carbonate) was heated in a Carteaux converter for 2 hours or bJ in a hatch of 61 tons each. , loaded. Additionally, 1.21 tons of limestone and 0.61 tons of iron oxide were charged into each haunch as flux.

Additionally, 0.8 tons of silica was charged to the first hatch, and 0.3 tons of coke was charged to the remaining hatches.

The charge is heated in an oil-oxygen gas hearth and smelted.
Therefore pond 30757! and 769 ONm of oxygen were consumed during the heating and smelting period. The 1υ1 period of smelting had a duration of 240 minutes and the heating of the charge required 80 minutes. After about 160 minutes of the smelting period, the slag started to become viscous.The temperature was slightly increased to about 1,100°C, and at this level it was further heated to 180°C.
It was held for 30 minutes, resulting in mainly evaporation of the zinc. The zinc content then decreased from about 15% to about 1%. The tin content was also substantially reduced. When the zinc reduction period ends, Pb 1.
5%, 7. nl, 0%, Sn 0.2%, Fe 14.
6%, 5i0227.5%, MgO2, 9%, 8lJa
Eight tons of slag containing 4.1% and 1027.9% C+ were removed from the furnace. Twenty tons of molten water with a purity of 99.6% pb was removed from the furnace. Approximately 2 tons of zinc and 0. 1 ton of soot is discharged from the exhaust furnace gas (exit
iB furnace f! as) was recovered in a gas scrubbing system.

[Brief explanation of the drawing]

Attachment (=1 drawing does not show embodiments of the present invention) IIJ-C1
-Illustration.

Claims (1)

[Claims] +11 A method for recovering metal valuables from a substance containing at least one of zinc and tin by smelting the substance under oxidizing conditions and reducing the resulting molten metal. A solid carbonaceous reducing agent for the slag and Reduction I11° is optionally also carried out with the addition of sulfur donor and/or chlorine donor substances; the reducing agent and the supplied donor substance are then suspended in the lag; the hold volume of the slag is at least during the late stages of the reduction period, when the zinc and soot reduction occurs to about 1-2% or more, and the reduction of zinc and soot occurs in the late stages: retaining the turbidity; converting the zinc to zinc vapor and removing it from the furnace; removing any tin present as volatile tin sulphides, chlorides or oxides;
A process characterized in that lead and other metal values present are removed as a molten lead phase. (2) C3 at current temperature and hold volume less than 2%
iOz20-30%, Ca025-35%, F”ef)
Iron calcium silicate slag ζ containing 25% and Mg (14-812035 to 10%). The method described. (3) The slag is Sio, about 25%, CaO about 30%.
, about 20% of FeO and MgO18I (1,16-8
Claim No. (2) characterized in that it contains %.
The method described in section. (4) Claims (1) to (3) characterized in that the suspension is maintained by vigorously stirring the molten water.
The method described in any one of paragraphs. (5) The reduction step is preferably 1,050 to I, 3
A method according to any one of claims 1 to 4, characterized in that it is carried out at a temperature in the range of 00°C. (0) Slag 4 containing 1 substance containing solid carbonate with reducing agent
．． Claim No. (1) characterized in that two additions are made.
The method according to any one of Items to Items (5). (7) supplying a sulfur donor and/or chlorine donor material to form volatile soot sulfide and/or chloride; converting said sulfide and/or chloride to tin dioxide; 2. A method as claimed in claim 1, characterized in that the oxide dust is formed from the gas present.