CN101903543B - Method for refining copper concentrate - Google Patents
Method for refining copper concentrate Download PDFInfo
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- CN101903543B CN101903543B CN200880121165.7A CN200880121165A CN101903543B CN 101903543 B CN101903543 B CN 101903543B CN 200880121165 A CN200880121165 A CN 200880121165A CN 101903543 B CN101903543 B CN 101903543B
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- slag
- suspension smelting
- copper
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- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims abstract description 96
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 60
- 239000010949 copper Substances 0.000 title claims abstract description 60
- 238000000034 method Methods 0.000 title claims abstract description 42
- 239000012141 concentrate Substances 0.000 title claims abstract description 23
- 238000007670 refining Methods 0.000 title claims abstract description 8
- 238000003723 Smelting Methods 0.000 claims abstract description 76
- 239000002893 slag Substances 0.000 claims abstract description 73
- 239000000725 suspension Substances 0.000 claims abstract description 67
- 239000002184 metal Substances 0.000 claims abstract description 38
- 229910052751 metal Inorganic materials 0.000 claims abstract description 38
- 238000006243 chemical reaction Methods 0.000 claims abstract description 19
- 230000004907 flux Effects 0.000 claims abstract description 8
- 239000003638 chemical reducing agent Substances 0.000 claims abstract description 7
- 239000012495 reaction gas Substances 0.000 claims abstract description 7
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims description 10
- 229910052760 oxygen Inorganic materials 0.000 claims description 10
- 239000001301 oxygen Substances 0.000 claims description 10
- 239000000571 coke Substances 0.000 claims description 4
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 3
- 238000002360 preparation method Methods 0.000 claims 1
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 30
- 229910052742 iron Inorganic materials 0.000 description 15
- 239000007789 gas Substances 0.000 description 11
- RAHZWNYVWXNFOC-UHFFFAOYSA-N Sulphur dioxide Chemical compound O=S=O RAHZWNYVWXNFOC-UHFFFAOYSA-N 0.000 description 8
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 7
- 239000003500 flue dust Substances 0.000 description 6
- 239000000428 dust Substances 0.000 description 5
- 239000002245 particle Substances 0.000 description 5
- 239000002918 waste heat Substances 0.000 description 5
- 229910045601 alloy Inorganic materials 0.000 description 4
- 239000000956 alloy Substances 0.000 description 4
- 239000000377 silicon dioxide Substances 0.000 description 4
- 235000010269 sulphur dioxide Nutrition 0.000 description 4
- 238000004140 cleaning Methods 0.000 description 3
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 2
- 239000002253 acid Substances 0.000 description 2
- 238000010923 batch production Methods 0.000 description 2
- 239000008187 granular material Substances 0.000 description 2
- 235000012239 silicon dioxide Nutrition 0.000 description 2
- 239000004291 sulphur dioxide Substances 0.000 description 2
- 239000005864 Sulphur Substances 0.000 description 1
- -1 and in such a case Substances 0.000 description 1
- 238000005266 casting Methods 0.000 description 1
- 239000003795 chemical substances by application Substances 0.000 description 1
- 229910052681 coesite Inorganic materials 0.000 description 1
- 238000001816 cooling Methods 0.000 description 1
- 229910052906 cristobalite Inorganic materials 0.000 description 1
- 230000001419 dependent effect Effects 0.000 description 1
- 238000005868 electrolysis reaction Methods 0.000 description 1
- 238000005516 engineering process Methods 0.000 description 1
- 238000005469 granulation Methods 0.000 description 1
- 230000003179 granulation Effects 0.000 description 1
- 238000002844 melting Methods 0.000 description 1
- 230000008018 melting Effects 0.000 description 1
- 229910001092 metal group alloy Inorganic materials 0.000 description 1
- 230000003647 oxidation Effects 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- 238000011084 recovery Methods 0.000 description 1
- 238000011946 reduction process Methods 0.000 description 1
- 230000000630 rising effect Effects 0.000 description 1
- 229910052682 stishovite Inorganic materials 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- 229910052717 sulfur Inorganic materials 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- 229910052905 tridymite Inorganic materials 0.000 description 1
- 229910000859 α-Fe Inorganic materials 0.000 description 1
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0026—Pyrometallurgy
- C22B15/0028—Smelting or converting
- C22B15/005—Smelting or converting in a succession of furnaces
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0026—Pyrometallurgy
- C22B15/0028—Smelting or converting
- C22B15/0047—Smelting or converting flash smelting or converting
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/04—Working-up slag
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0026—Pyrometallurgy
- C22B15/0028—Smelting or converting
- C22B15/0052—Reduction smelting or converting
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0026—Pyrometallurgy
- C22B15/0054—Slag, slime, speiss, or dross treating
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- Engineering & Computer Science (AREA)
- Chemical & Material Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The present invention relates to a method for refining copper concentrate. In the method, copper concentrate (1), flux (2) and reaction gas (3) are fed together into a reaction shaft (5) of a suspension smelting furnace (4), for example to the reaction shaft (5) of a flash smelting furnace, and separate phases, i.e. blister copper (13) and slag (14), are produced in the suspension smelting furnace (4). In the method, slag (14) from a suspension smelting furnace is conducted into an electric furnace (16), and the slag (14) from the suspension smelting furnace is treated with a reducing agent in the electric furnace (16) so that separate phases, i.e. bottom metal (17) and slag (18), are produced in the electric furnace (16); removing the electric furnace bottom metal (17) from the electric furnace (16), graining the electric furnace bottom metal (17), and obtaining a grained electric furnace bottom metal (22); and feeding the granulated electric furnace bottom metal (22) to the reaction shaft (5) of the suspension smelting furnace (4).
Description
Technical Field
The present invention relates to a method for refining copper concentrate according to the preamble of claim 1.
Background
When refining copper concentrate in a suspension smelting furnace, such as a flash smelting furnace, two phases, i.e. blister copper (blister copper) and suspension smelting slag, are obtained as products from the suspension smelting furnace.
The blister copper obtained from the suspension smelting furnace is further refined in an anode furnace after the suspension smelting furnace, after which the copper is cast into copper anodes and further electrolytically refined in an electrolysis apparatus by using said copper anodes.
However, not all the copper contained in the copper concentrate is transferred from the copper concentrate to blister copper in the suspension smelting furnace, whereas the slag from the suspension smelting furnace contains a large amount of copper, usually even up to 20%, which can be recovered by various slag cleaning methods.
Two different methods are applied for slag cleaning. The first method is based on partial reduction of slag from a suspension smelting furnace in an electric furnace. In this method, the copper metal obtained from the electric furnace is very pure and can even be fed to the anode furnace together with the blister copper obtained from the suspension smelting furnace. In the partial reduction of slag from a suspension smelting furnace in an electric furnace, so-called partially reduced slag obtained from the electric furnace as secondary products other than copper metal also contains copper. However, in order to recover the copper contained in the partially reduced slag obtained from the electric furnace, the partially reduced slag from the electric furnace must be treated in a concentration (concentration) plant, which is expensive both in terms of operating costs and capital costs.
In a second industrially applicable process, the slag from the suspension smelting furnace is reduced in an electric furnace as a batch process, so that after the reduction process the copper content of the suspension smelting slag is very low, so that further treatment of the slag obtained from the electric furnace, except for the bottom metal, is not economically viable. However, after the reduction step has been carried out sufficiently well, the bottom metal (or alloy) formed in the electric furnace process contains a large amount of iron, so that it is disadvantageous to supply the electric furnace bottom metal to the anode furnace together with the blister copper from the suspension smelting furnace, and the iron must first be removed in a separate conversion process in a so-called iron converter before the copper contained in the electric furnace bottom metal is supplied to the anode furnace.
Thus, the above examples of slag cleaning processes each include two steps.
Disclosure of Invention
The object of the present invention is to propose an improved method of refining copper concentrate.
The object of the invention is achieved by a method according to independent claim 1.
Preferred embodiments of the method according to the invention are set forth in the dependent claims.
In this innovation an arrangement is introduced which has two steps per se, but which is more economical than the above-described arrangement in terms of investment costs and in particular in terms of operating costs. The slag produced in the suspension smelting furnace is further processed in an electric furnace, in a separate unit functioning in a continuous operating process or as a batch process. The partial reduction, or subsequent reduction, of the suspension smelting furnace slag in the electric furnace is such that the slag produced in the electric furnace is a so-called rejectable slag, i.e. its copper content is so low that it is economically not feasible to recover the remaining copper in a separate process. The metal alloy obtained from the electric furnace, i.e. the bottom metal, is granulated, for example by means of water. The formed alloy particles are fed to the reaction shaft of the suspension smelting furnace together with the copper concentrate, the flux and the reaction gas, so that the alloy particles melt and, when advancing through the slag in the settler of the suspension smelting furnace, reach a similar thermodynamic equilibrium with the slag as the blister copper formed from the concentrate. At this time, the iron contained in the particles is oxidized and slagged, so that the blister copper obtained as a product from the suspension smelting furnace is advantageously treated directly in the anode furnace. Due to the low amount of slag-forming components (mainly iron) contained in said granulated copper, the amount of slag does not substantially increase and thus does not cause any excess copper to be recycled back into the electric furnace, but the majority of the copper contained in the granules is directly transferred to blister copper coming from the suspension smelting process as product.
Among the advantages of this process, in addition to the reduction of operating and investment costs, the following characteristics can be pointed out:
compared with the existing two-step method, the copper cycle is reduced;
only blister copper of one quality is fed into the anode furnace, in which case the operation of the anode furnace becomes easier;
in direct blister copper smelting, a large amount of heat is often generated, so that oxygen enrichment has to be limited. Since the heat is used here in the method itself for melting the alloy particles, the smelting furnace can be operated at a higher oxygen enrichment level, as a result of which a larger smelting furnace capacity is obtained (or the smelting furnace, in particular the reaction shaft, can be smaller) and the capacity of the gas lines can be smaller.
In a preferred embodiment, two successive electric fires are used. In the first electric furnace, the reduction of the suspension smelting furnace slag only reaches a level of about 4% Cu, i.e. the remaining partially reduced slag contains a level of about 4% copper, whereas in such a case the iron contained in the slag from the suspension smelting furnace has not been reduced and transferred to the bottom metal phase in the first electric furnace, but remains in the first electric furnace as so-called partially reduced slag. As a product from the first electric furnace, the blister copper obtained can be used directly in the anode furnace for further processing and supply to the anode furnace, since the blister copper from the first electric furnace does not contain iron. In the second electric furnace, the partially reduced slag from the first electric furnace is continuously reduced in order to recover the remaining copper contained in the slag, and in such a case, iron is also reduced together with blister copper; the iron-containing bottom metal is granulated and fed back to the reaction shaft of the suspension smelting furnace, where the iron is oxidized in the manner described above.
Drawings
Several preferred embodiments of the present invention are described in detail below with reference to the attached drawing figures, wherein
FIG. 1 shows a first embodiment of the method, and
fig. 2 shows a second embodiment of the method.
Detailed Description
Figure 1 shows a method for refining copper concentrate 1.
In the method, copper concentrate 1, flux 2 and reaction gas 3, such as oxygen-enriched air, are fed together into a reaction shaft 5 of a suspension smelting furnace 4, e.g. to the reaction shaft of a flash smelting furnace.
It is also possible to feed flue dust 9 into the reaction shaft 5 of the suspension smelting furnace 4, the flue dust being obtained from a waste heat boiler 8, from cooling of the exhaust gases 7 discharged through the uptake shaft 6 of the suspension smelting furnace 4, and/or the flue dust 9 being obtained from an electric dust collector arranged after the waste heat boiler 8.
The substances fed into the reaction shaft 5 of the suspension smelting furnace 4 react together and form separate phases on the bottom 12 of the settler 11 of the suspension smelting furnace 4: blister copper 13, and slag 14 on top of the blister copper 13.
The exhaust gases 7 produced in the suspension smelting furnace are discharged through the rising shaft 6 to a waste heat boiler 8, where the thermal energy of the exhaust gases 7 is recovered. The cooled exhaust gases 7 are led from the waste heat boiler 8 to an electric dust collector 10, where flue dust 9 is separated from the exhaust gases 7 and the flue dust 9 is circulated back to the reaction shaft 5 of the suspension smelting furnace 4. The exhaust gas 7 is led from the electric dust separator 10 for further treatment, e.g. in an acid plant (not shown), for recovery of sulphur dioxide.
The blister copper 13 obtained from the suspension smelting furnace is led to an anode furnace 15 for pyrometallurgicrination. In the anode furnace 15, a small amount of sulfur contained in the blister copper 13 is first removed by oxidation, and then, oxygen contained in the blister copper 13 is removed by reduction. After the anode furnace 15, copper is cast into a copper anode in an anode casting device (not shown), and by using the anode, the copper contained in the copper anode (i.e., the copper anode) is further electrolytically refined into electrolytic copper in an electrolytic device (not shown).
The slag 14 from the suspension smelting furnace is preferably, but not necessarily, conducted into the electric furnace 16 in a molten state, which saves energy since the slag from the suspension smelting furnace 14 is already in a molten state when it reaches the electric furnace 16.
The slag from the suspension smelting furnace 14 is treated with a reducing agent, such as coke, in a reduction furnace, such as an electric furnace 16, so that separate phases, i.e. bottom metal 17 and slag 18, are formed in the electric furnace 16. Preferably, but not necessarily, the slag 14 from the suspension smelting furnace is reduced in an electric furnace 16 by means of coke fed into the electric furnace 16.
Preferably, but not necessarily, anode furnace slag 19 from the anode furnace 15 is also fed to the electric furnace 16.
Preferably, but not necessarily, the slag 14 from the suspension smelting furnace is reduced in an electric furnace 16 so that the copper content in the electric furnace slag 18 is kept below 2%, most advantageously below 1%.
The bottom metal 17 of the electric furnace is removed from the electric furnace 16 and the bottom metal 17 of the electric furnace is granulated in a granulation device 21, for example by means of water 20. The furnace bottom metal 17 contains, in particular, iron in addition to copper.
The granulated electric furnace bottom metal 22 is fed to the reaction shaft 5 of the suspension smelting furnace 4 together with the copper concentrate 1, the flux 2 and the reaction gas 3.
Fig. 2 shows a further embodiment of the method, in which instead of only one electric furnace as depicted in fig. 1, two electric furnaces, namely a first electric furnace 23 and a second electric furnace 24, are used here.
In fig. 2, the slag 14 from the suspension smelting furnace is first conducted into an electric furnace 23. The suspension smelting furnace slag 14 is conducted from the suspension smelting furnace 4, preferably but not necessarily in a molten state, to a first electric furnace 23.
In the first electric furnace 23, the suspension smelting furnace slag 14 is subjected to partial reduction by means of a reducing agent, so that separate phases, blister copper 13 and partially reduced slag 25 containing about 4% copper are formed in the first electric furnace 23.
The blister copper 13 from the first electric furnace is supplied from the first electric furnace 23 to the anode furnace 15. The blister copper 13 obtained from the first electric furnace 23 is preferably, but not necessarily, supplied in a molten state from the first electric furnace 23 to the anode furnace 15. As a product from the first electric furnace 23, blister copper 13 is obtained which can be used for further processing in the anode furnace 15, and blister copper can be supplied to the anode furnace 15 because blister copper obtained from the first electric furnace contains no iron, and only partial reduction of the suspension smelting furnace slag 14 is carried out in the first electric furnace 23.
The partially reduced slag 25 is preferably, but not necessarily, supplied in a molten state from the first electric furnace 23 to the second electric furnace 24.
In the second electric furnace 24, the partially reduced slag 25 from the first electric furnace is subjected to reduction by means of a reducing agent, so that separate phases are formed in the second electric furnace 24: bottom metal 17 and slag 18, the residual copper content in the slag 18 being less than 2%, most advantageously less than 1%.
In addition to copper, the bottom metal 17 from the second electric furnace likewise contains in particular iron. The bottom metal 17 is granulated and fed to the reaction shaft 5 of the suspension smelting furnace 4 together with the copper concentrate 1, the flux 2 and the reaction gas 3.
Examples of the invention
The feed into the suspension smelting furnace is:
copper concentrate (concentrate) 111.0t/h
Flue dust (DBF dust) 19.6t/h
Slagging agent, i.e. flux (silica flux) 9.9t/h
Granular bottom metal (electric furnace metal) 16.6t/h
Total 157.2t/h
And (3) copper concentrate analysis:
34.8 percent of copper
26.0 percent of Fe
S29.1%
Silicon dioxide SiO25.0％
In addition, 60680Nm3 of oxygen enriched air was fed into the suspension smelting furnace, and the oxygen enrichment degree was 46.2%.
Oxygen-enriched air is used for suspension smelting, since the heat generated by the reaction between sulphur contained in the concentrate and ferrite is sufficient to melt the concentrate (producing blister copper and slag) and blister copper particles with a fine grain size. Due to the relatively high oxygen enrichment, a sulfur dioxide with a high sulfur dioxide content (about 36% SO) is produced2) The total amount of said gas is lower than if a lower degree of oxygen enrichment was used. The gases were discharged from the furnace at a rate of about 66,900Nm3/h and at a temperature of 1,320 ℃. The main part of the thermal energy of the gas is recovered in a waste heat boiler before the gas is led to a thermoelectric dust collector and further to an acid plant for recovering sulphur dioxide.
The products obtained from the suspension smelting furnace are blister copper (speed of 39 tons per hour, temperature of about 1,280 ℃) and slag (speed of about 77 tons per hour).
The slag obtained from the suspension smelting furnace has a copper content of 20% Cu and, in order to recover the copper, is fed into the electric furnace in a molten state, so that the amount of slag processed there is 1,830 tons per day. In addition, a small amount of anode furnaces are usedThe slag (20 tons per day) and about 91 tons per day of coke required for reduction were fed into the electric furnace. As a result of the reduction, a slag is produced, the copper content of which is so low that its further treatment is not economically viable [ 1,365 tons per day, iron (Fe) approximately 51%, silicon dioxide (SiO)2) About 26%]. As a product, the bottom metal is produced at a rate of about 400 tons per day, and the iron content in the bottom metal is about 8%, the remainder being mainly copper. The bottom metal is granulated at a temperature of 1,240 c and the granules are dried and fed back to the flash smelting furnace together with the concentrate.
Thus, blister copper is formed in the process as described above and can advantageously be further processed in an anode furnace to anode copper.
It is obvious to a person skilled in the art that as technology advances, the main idea of the invention can be implemented in many different ways. The invention and its various embodiments are thus not limited to the examples described above, but they may vary within the scope of the attached claims.
Claims (11)
1. A method for refining copper concentrate, in which method
-feeding the copper concentrate (1), the flux (2) and the reaction gas (3) together into the reaction shaft (5) of the suspension smelting furnace (4), and
separate phases, namely blister copper (13) and slag (14), are formed in the suspension smelting furnace (4),
it is characterized in that the preparation method is characterized in that,
-guiding slag (14) from the suspension smelting furnace into an electric furnace (16),
-treating slag (14) from a suspension smelting furnace with a reducing agent in an electric furnace (16) such that separate phases, i.e. furnace bottom metal (17) and slag (18), are generated in the electric furnace (16),
-removing the furnace bottom metal (17) from the furnace (16),
-graining the electric furnace bottom metal (17) and obtaining a grained electric furnace bottom metal (22), and
-feeding granulated electric furnace bottom metal (22) to a reaction shaft (5) of a suspension smelting furnace (4),
wherein slag (14) from the suspension smelting furnace is conducted in a molten state into an electric furnace (16).
2. A method according to claim 1, characterized in that the bottom metal (17) of the electric furnace from the electric furnace is granulated by means of water (20).
3. A method according to claim 1, characterized in that the slag (14) from the suspension smelting furnace is reduced in the electric furnace (16) by means of coke fed into the electric furnace (16).
4. A method according to claim 1, characterized in that anode furnace slag (19) from the anode furnace (15) is also fed into the electric furnace (16).
5. A method according to claim 1, characterized by reducing the slag (14) from the suspension smelting furnace in an electric furnace (16) so that the copper content in the electric furnace slag (18) is below 2%.
6. The method of claim 1,
-using two electric furnaces in the method, a first electric furnace (23) and a second electric furnace (24),
-first guiding slag (14) from the suspension smelting furnace to a first electric furnace (23),
-in the first electric furnace (23), the slag (14) of the suspension smelting furnace is subjected to partial reduction by means of a reducing agent, so that separate phases, i.e. blister copper (13) and partially reduced slag (25) containing 4% copper, are generated in the first electric furnace (23),
-feeding partially reduced slag (25) from the first electric furnace (23) into a second electric furnace (24),
-in the second electric furnace (24), the partially reduced slag (25) obtained from the first electric furnace is subjected to reduction by means of a reducing agent, so that separate phases, i.e. bottom metal (17) and slag (18), are formed in the second electric furnace (24), in which slag (18) the copper content remains below 2%,
-removing the bottom metal (17) of the second electric furnace from the second electric furnace (24),
-graining the bottom metal (17) from the second electric furnace and obtaining a grained electric furnace bottom metal (22), and
-feeding granulated electric furnace bottom metal (22) to a reaction shaft (5) of a suspension smelting furnace (4),
wherein slag (14) from the suspension smelting furnace is conducted in a molten state from the suspension smelting furnace (4) to a first electric furnace (23).
7. A method according to claim 6, characterized in that the blister copper (13) obtained from the first electric furnace is fed into an anode furnace (15).
8. A method according to claim 1, characterized in that the reaction gas (3) fed into the reaction shaft (5) of the suspension smelting furnace (4) comprises oxygen-enriched air.
9. A method according to claim 6, characterized in that the copper content remains less than 1% in the slag (18).
10. A method according to claim 1, characterized in that the suspension smelting furnace (4) is a flash smelting furnace.
11. A method according to claim 5, characterized in that the copper content in the electric furnace slag (18) is less than 1%.
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| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN201610207726.1A CN105936980A (en) | 2007-12-17 | 2008-12-15 | Method for refining copper concentrate |
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|---|---|---|---|
| FI20075920A FI120157B (en) | 2007-12-17 | 2007-12-17 | A process for refining copper concentrate |
| FI20075920 | 2007-12-17 | ||
| PCT/FI2008/050735 WO2009077651A1 (en) | 2007-12-17 | 2008-12-15 | Method for refining copper concentrate |
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| CN101903543B true CN101903543B (en) | 2020-07-28 |
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| CN (2) | CN101903543B (en) |
| AU (1) | AU2008337430B2 (en) |
| BR (1) | BRPI0821242B1 (en) |
| CL (1) | CL2008003744A1 (en) |
| EA (1) | EA018279B1 (en) |
| FI (1) | FI120157B (en) |
| PE (1) | PE20091539A1 (en) |
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| CN102605191B (en) | 2012-04-16 | 2013-12-25 | 阳谷祥光铜业有限公司 | Method for directly producing row copper by copper concentrate |
| FI124912B (en) * | 2012-04-16 | 2015-03-31 | Outotec Oyj | A method for treating metallurgical slags of non-ferrous metals |
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| FI115536B (en) * | 2001-09-21 | 2005-05-31 | Outokumpu Oy | A process for producing crude copper |
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| CA1245460A (en) * | 1985-03-20 | 1988-11-29 | Carlos M. Diaz | Oxidizing process for sulfidic copper material |
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| FI115638B (en) * | 2002-12-05 | 2005-06-15 | Outokumpu Oy | Procedure for treating slag |
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2008
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- 2008-12-15 AU AU2008337430A patent/AU2008337430B2/en active Active
- 2008-12-15 PL PL392792A patent/PL213990B1/en unknown
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| FI63441C (en) * | 1976-02-23 | 1983-06-10 | Outokumpu Oy | FOERFARANDE FOER FRAMSTAELLNING AV RAOKOPPAR FRAON KOPPARMALM ELLER -KONCENTRAT INNEHAOLLANDE SKADLIGA ELLER EKONOMISKT SINIFIKANTA MAENGDER ANDRA ICKE-JAERNMETALLER |
| FI115536B (en) * | 2001-09-21 | 2005-05-31 | Outokumpu Oy | A process for producing crude copper |
Also Published As
| Publication number | Publication date |
|---|---|
| FI120157B (en) | 2009-07-15 |
| CN105936980A (en) | 2016-09-14 |
| BRPI0821242A2 (en) | 2015-06-16 |
| JP2011506777A (en) | 2011-03-03 |
| FI20075920A (en) | 2009-06-18 |
| AU2008337430A1 (en) | 2009-06-25 |
| EA201000893A1 (en) | 2010-12-30 |
| WO2009077651A1 (en) | 2009-06-25 |
| BRPI0821242B1 (en) | 2019-09-10 |
| CN101903543A (en) | 2010-12-01 |
| PL392792A1 (en) | 2011-02-28 |
| AU2008337430B2 (en) | 2013-03-28 |
| PE20091539A1 (en) | 2009-10-29 |
| EA018279B1 (en) | 2013-06-28 |
| FI20075920A0 (en) | 2007-12-17 |
| PL213990B1 (en) | 2013-06-28 |
| CL2008003744A1 (en) | 2009-11-27 |
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