CN101787439A - Method for recovering valuable metals from metallurgical waste - Google Patents
Method for recovering valuable metals from metallurgical waste Download PDFInfo
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- CN101787439A CN101787439A CN201010117128A CN201010117128A CN101787439A CN 101787439 A CN101787439 A CN 101787439A CN 201010117128 A CN201010117128 A CN 201010117128A CN 201010117128 A CN201010117128 A CN 201010117128A CN 101787439 A CN101787439 A CN 101787439A
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Abstract
The invention relates to a method for recovering valuable metals from metallurgical waste (which mainly comprises smelting slag, leaching slag and smoke dust with ferro-silicate and ferrate containing the valuable metals such as copper, cobalt, nickel, zinc and the like). The method comprises the following steps: 1) acidolysis: mixing the metallurgical waste with water and sulfuric acid for reaction to be decomposed, and releasing the valuable metals combined in the metallurgical waste; 2) calcinating: mixing material obtained by the acidolysis with sulfur material to obtain mixed material, and then filling air into the mixed material for calcination, wherein the sulfur material contains non-oxide-state sulfur component; and controlling the calcination temperature to be 450-800 DEG C; 3) leaching: after calcination, adding water into calcine sand for leaching; and 4) recovery: carrying out solid-liquid separation on ore pulp after leaching, obtaining solution containing the valuable metals, and recovering the valuable metals. The recovery method has high leaching efficiency, and is economical, simple and convenient.
Description
Technical field
The present invention relates to the metals resources cyclic regeneration, specifically, relate to the regeneration of industrial solid castoff and secondary resource, more particularly, relate to a kind of method that from metallurgical waste, reclaims valuable metal, particularly non-ferrous metal (non-ferrous metal).
Background technology
Metallurgical industry all produces a large amount of solid wastes every year, mainly is smelting slag, blowing slag and the flue dust of slag and flue dust, especially non-ferrous metals smelting works; The leached mud of hydrometallurgy factory; And steel stainless steel slag, ferroalloy slag and electric arc furnace flue dust etc.These solid wastes all contain valuable elements such as copper, cobalt, nickel, zinc.Different according to the smelting process of used raw material and employing, copper, nickel smelted furnace cinder cupric can be from some thousandths of to percent ten even higher, containing cobalt, nickel can all enter more than the cobalt in the concentrate even 90% in smelting slag or the blowing slag from the some thousandths of to percentum.Neutral leached mud of zinc hydrometallurgy factory calcining and electric arc furnace dirt contain zinc 10 to 20.These metallurgical wastes are the economically valuable because containing valuable metal both, also is put into hazardous waste because of the existence of these elements simultaneously.Because enormous amount, also increasing year by year in addition, wherein the resource circulation existing huge economic of regenerating has urgent environment protection significance again.
In the smelted furnace cinder of non-ferrous metal, it is lower that smelting slag (as blast furnace slag and reflection slag) contains the valency metal, generally all directly discarded at present.It is higher that blowing slag (as converter slag) and intensified smelting slag (flash slag) contain the valency metal, can not discard.The processing of blowing slag in the practice, the one, dilution, the 2nd, flotation all only considers to reclaim copper wherein, and valuable metals such as contained cobalt, nickel, zinc all do not reclaim.Dilution is about to slag and returns melting or melting separately in the stove, reclaims wherein part valuable metal, and the finishing slag that the output containing metal is lower is discarded.But except that melting inherent energy consumption height, generation exhaust emission, the cobalt in the slag, nickel still enter the finishing slag loss mostly when Returning smelting, and slag returns in a large number, has reduced the productivity of equipment, has improved cost.Flotation is used to reclaim copper from converter slag more and more, but also can only reclaim the cupric sulfide and the metallic copper of mechanical entrainment in the slag, be sosoloid in the slag and be dispersed in metal oxide, particularly cobalt in ferrosilicate and the wustite, almost complete loss is in flotation tailings.And, require the control slag slowly to cool off, thereby the old slag flotation effect of storing up in the past is bad for making the copper gathering in the slag grow up independently phase.Because still lack the effective ways of recycling valuable element in the slag economically, most of metallurgical slags all are stacked in the slag field.Add over the slag of old accumulation, not only take valuable soil, resource in the waste slag also pollutes the environment.
For from metallurgical slag, extracting valuable metal better, the whole bag of tricks has been proposed so far.Directly leach with sulfuric acid or iron(ic) chloride and to wish to get good leaching effect, need oxygen depress operation (Anand, S., Sarveswara Rao, K., Jena, P.K., 1983.Hydrometallurgy 10,305-312), investment is all high with process cost.Can not reclaim the cobalt that wherein is dispersed in ferrosilicate (fayalite) and the ferrous acid salt face, nickel, zinc etc. with the sulfatizing roasting metallurgical slag, roasting also needs to add fuel or outside heat supply.Also there is a same problem with pyrite and slag are baking mixed, just no longer needs fuel (Bulut G., Perek K.T.and Gul A., 2007.Minerals ﹠amp; MetallurgicalProcessing 24,13-18).With sulfuric acid 150 ℃ of following roasting slags again water leach that (USPatent 3868440,1975), can obtain valuable metal leaching yields such as higher cobalt, nickel, zinc, but the iron in the slag almost all enters solution, bring very big difficulty for follow-up metal recovery operation, and vitriolic consumption very big (mass ratio of sulfuric acid and slag is about 1: 1).Simultaneously, being the valuable metal that sulfide exists in the slag is difficult to leach.Adopt sulfuric acid two-stage roasting (Sukla, L.B., Panda, S.C., Jena, P.K., 1986.Hydrometallurgy 16,153-165.), be slag and sulfuric acid compound after 150 ℃ of following roastings, again this product of roasting is heated to 650 ℃ of following roastings, roasting material is used water logging again, can obtain satisfied copper, cobalt, nickel leaching yield, and the dissolving of iron is also controlled.But second section roasting needs fuel or outer heating, increased energy consumption and Carbon emission.
Summary of the invention
The purpose of this invention is to provide a kind of can be cost-effectively from metallurgical waste, as reclaiming the method for valuable metals such as copper and cobalt, nickel, zinc in metallurgical slag, leached mud and the flue dust (valuable metals such as ferrosilicate and wustite cupric, cobalt, nickel, zinc are wherein arranged).
The invention provides a kind of method that from metallurgical waste, reclaims valuable metal, comprise the steps:
1) acidolysis: metallurgical waste is mixed with water, sulfuric acid, make its reaction decomposes to discharge wherein bonded valuable metal;
2) roasting: with the material after the acidolysis (be called for short the acidolysis material, down with) mix with the sulphur material compound, bubbling air roasting then; Wherein said sulphur material is the material that contains non-oxide attitude sulphur component, and maturing temperature is controlled at 450 ℃～800 ℃;
3) leach: after the roasting, calcining adds water logging and goes out;
4) reclaim: the solid-liquid separation on ore pulp after will leaching obtains containing the solution of valency metal, for the valuable metal that reclaims wherein.
Described metallurgical waste is metallurgical slag, leached mud and flue dust etc., contains the valency metal in ferrosilicate that wherein comprises and the wustite, as copper, cobalt, nickel, zinc etc.
In the described step 1) acidolysis, the mass ratio of described metallurgical waste, sulfuric acid and water is 1: (0.3～3): (0.5～2).
Described step 1) acidolysis comprises the steps:
(1) sizes mixing: ground metallurgical waste and water are mixed and made into ore pulp; The mass ratio of water and slag is (0.5～2): 1;
(2) mix acid: admix sulfuric acid to the ore pulp that mixes up, ore pulp and sulfuric acid are mixed; Wherein the mass ratio of the described metallurgical waste of sulfuric acid and step (1) is (0.3～3): 1;
(3) slaking: the material of mixing after the acid is placed more than the 0.5h.
Step 2) in, described sulphur material comprises sulphide concentrate or sulphur for containing the material of non-oxide attitude (negative divalence or zeroth order) sulphur component.
Described step 2) in, the ratio that the total amount of non-oxide attitude sulphur accounts for the compound total amount in the compound is not less than 6%w/w, preferred 8%～20%w/w.
Described step 2) in, maturing temperature is controlled at 450 ℃～800 ℃, and preferred temperature range is 500 ℃～750 ℃; Roasting time is at 0.5h～5h, preferred 1h～2h.
In the described step 3), extraction temperature maintains more than 20 ℃, and extraction time is 0.5h～5h.
In the described step 3), free vitriolic concentration is less than 30g/L in the gained leach liquor.
The filtrate of gained in the described step 4) through carrying out aftertreatment, is reclaimed wherein valuable metal with the prior art that comprises precipitation, solvent extraction or ion-exchange etc.
The main points of aforesaid method are earlier with ferrosilicate (as fayalite) and wustite in the sulfuric acid decompose slag, discharge wherein valuable metal such as bonded cobalt, nickel, zinc and change vitriol into; Mixture with acidolysis mixes with the sulphur material then, relies on sulphur material oxidation heat liberation to carry out autogenous roasting, finishes the sulphating of valuable metal, and makes the sulphate decomposition of iron become water-fast ferric oxide; After the calcining water leaches, again with the valuable metal in the existing technology recovery leach liquor.
Its technological process specifically describes as follows:
1) acidolysis: will grind good metallurgical slag and mix, and make the slag reaction decomposes, and discharge wherein bonded valuable metal with water, sulfuric acid.Hybrid mode can be determined on a case-by-case basis, and takes following steps generally speaking: size mixing (1): will grind good metallurgical slag and a certain amount of water mixed pulp, the mass ratio of water and slag is about (0.5～2): 1.Adopt wet-milling as pomace, then can control the liquid-solid ratio of ore grinding slurry in aforementioned range by mode such as dense.(2) mix acid: admix sulfuric acid to the ore pulp that mixes up, ore pulp and sulfuric acid are mixed.The mass ratio of sulfuric acid and slag is (0.3～3): 1.At this moment vigorous reaction can take place and emit a large amount of heat, temperature of reaction system is increased to rapidly and boils, and with moisture evaporation.(3) slaking: the material that will mix after the acid is placed more than the 0.5h, during placement insulation better, it is more abundant that mineral in the slag are decomposed, and the silica gel of generation is further dewatered, help improving the leaching yield of metal and leach after solid-liquid separation.
2) roasting: the acidolysis material is mixed the roasting of gained compound bubbling air with the sulphur material.Described sulphur material is the best to contain identical valuable metal person with processed metallurgical waste, as the sulphide concentrate and the sulphur of valuable metals such as copper, cobalt, nickel, iron.The ratio that the total amount of non-oxide attitude sulphur accounts for the compound total amount in the described compound is not less than 6%w/w, and the non-oxide attitude sulphur total amount of optimization accounts for compound total amount 8%～20%w/w.The blowing air roasting then of described compound, maturing temperature should be controlled at 450 ℃～800 ℃ scopes, preferably at 550 ℃～750 ℃.Baking flue gas collection system sulfuric acid or sulphite, sulfurous gas product.
3) leach: behind the mixture roasting, add water logging and go out calcining to extract valuable metal wherein, extraction temperature should maintain more than 20 ℃.As leaching while hot, calcining itself with heat generally can satisfy the extraction temperature requirement, do not need other heating; Calcining can be levigate before leaching if seriously lump, and avoids macrobead can not fully suspend at leaching vat; Leach water and be generally the recirculated water that each operation is returned in the flow process, replenish the part fresh water when not enough; The acid content of looking the content of raw materials used neutral and alkali mineral during leaching and leaching the recirculated water of usefulness is added or is not added (in most cases not adding) sulfuric acid, and leach liquor free sulfuric acid concentration should be less than 30g/L.
4) reclaim: after leaching the solid-liquid separation on ore pulp that obtains, contain the solution existing conventional isolation technique of valency metal, comprise that precipitation, solvent extraction or ion exchange method reclaim valuable constituent wherein, produce metal or its esters product.
The present invention reclaims valuable metal from metallurgical waste side has following beneficial effect:
1) in the inventive method, starches with sulfuric acid and metallurgical waste and to mix acidolysis, discharge valuable metal wherein and make it sulphating, created the prerequisite that these valuable metals are extracted in water logging.The heat of dilution that sulfuric acid is emitted when mixing with water has promoted acidolysis reaction.Without this acidolysis process direct roasting, then be combined in valuable metals such as cobalt in ferrosilicate or the wustite, nickel, zinc and be difficult to change into corresponding vitriol and leached.
2) among the present invention, the wherein sulphating of valuable metal has further been strengthened in the roasting of acidolysis material, the high metal leaching rate of water logging operation after having guaranteed; The vitriol of the iron that acidolysis simultaneously produces resolves into the oxide compound of iron in roasting, the overwhelming majority is stayed in the leached mud when water logging, helps the purification of leach liquor thereafter and the recovery of metal.The sulfurous gas that the vitriol thermolysis of iron is emitted has not only been strengthened the sulphating of valuable metal, and majority enters the flue gas by-product and make sulfuric acid, greatly reduces actual sulfuric acid consumption; Add thermal bake-out in addition and also can promote the silica gel that produces in the acidolysis to dewater, help later solid-liquid separation.
3) among the present invention, big calorimetric is emitted in the sulphur material oxidation that roasting metallurgical waste acid hydrolysate is allocated into, has realized autogenous roasting, need not other refuel or outside heat supply, provides cost savings, and has also reduced Carbon emission.Simultaneously the contained valuable metal of sulphur material self in roasting also the oxidation because of sulphur be transformed into corresponding vitriol, when water logging and the valuable metal in the metallurgical waste leached equally, realized the METAL EXTRACTION of sulphur material self, and increased total metal quantum of output, thereby reduced the cost of unit metal.
4) among the present invention, after metallurgical waste and described sulphur material were baking mixed, the product of roasting that obtains (calcining) leached while hot, and leaching vat do not need outer heating, also made and leached operation economy, easy.
Embodiment
Following examples are used to illustrate the present invention, but are not used for limiting the scope of the invention.
Embodiment 1
To contain (butt, mass percent, be milled to 60% by the 0.15mm sieve aperture after down together) copper smelting slag of 2.33%Cu, 0.44%Co, 4.03%Zn, 51.04%Fe, 0.54%S is weighed, be divided into three equal parts again, add the water stirring respectively and make even ore pulp, add the 95%w/w industrial sulphuric acid then rapidly.First part slag: water: acid (mass ratio) is 1: 1.2: 1.2; Second part slag: water: acid (mass ratio) is 1: 0.5: 0.3; The 3rd part slag: water: acid (mass ratio) is 1: 2: 3.Slag, water, acid mixture generation vigorous reaction and follow a large amount of heat releases when mixing acid, temperature moment is increased to boils, the slag loose drying that becomes.After the slag of three parts of acidolysis was all placed 2h, first part of acidolysis slag mixed with the cobalt concentrate mass ratio by former slag with the cobalt concentrate that contains 1.62%Cu, 0.58%Co and 28.55%S at 1: 1, and bubbling air is 620 ℃ of following roastings then, and roasting time is 2h.Second part of acidolysis slag mixes the back bubbling air at 500 ℃ of following roastings, roasting time 3h with described cobalt concentrate by former slag and cobalt concentrate mass ratio at 1: 1.75.The 3rd part of acidolysis slag mixes the back bubbling air at 750 ℃ of following roastings, roasting time 50min with described cobalt concentrate by former slag and cobalt concentrate mass ratio at 1: 0.75.Calcining adds water to liquid-solid ratio leaching while hot in 3: 1 respectively after the roasting, and extraction time is 3h.Slurries filtration after the leaching, filtrate and filter residue are analyzed its chemical constitution respectively, and analytical results calculating leaching yield is as follows in view of the above: first part of slag Co 92.42%, Cu96.33%; Second part of slag Co 90.85%, Cu 94.18%; The 3rd part of slag Co 91.27%, Cu95.16%.
Embodiment 2
The nickel converter slag that portion is contained 0.89%Co, 2.05%Ni, 0.40%Cu, 0.28%S adds the water wet-milling, after being milled to solid all the slurry of the slag by the 0.3mm sieve aperture removing portion water, admix commercially available 95%w/w industrial sulphuric acid and mix, make slag: water: acid (mass ratio) is 1: 1: 1.Place behind the 36h and contain the pyrrhotite concentrate of 1.26%Ni, 0.50%Cu, 0.033%Co, 22.76%S by former slag and 1: 1.2 baking mixed 1h of copper ore concentrates mass ratio, 650 ℃ of maturing temperatures.Add water to liquid-solid ratio then and while hot leach extraction time 3h at 3: 1.Slurries filtration after the leaching, filtrate and filter residue are analyzed its chemical constitution respectively, and analytical results calculating leaching yield is as follows in view of the above: Ni 90.25%, Co 91.13%, Cu 87.87%.
Embodiment 3
The zinc hydrometallurgy factory leached mud that portion is contained 15.67%Zn, 38.78%Fe and 0.28%S adds water and stirs and make even ore pulp, and admix 95% industrial sulphuric acid rapidly and mix slag: water: acid (mass ratio) is 1: 0.8: 1.2.Mix than 1: 0.8 with concentrate quality by former slag with the zinc ore concentrate that contains 45.34%Zn and 36.37%S after placing 24h, behind 720 ℃ of following roasting 1h, add water to liquid-solid ratio and while hot leach 3h at 3: 1.Leach the back solid-liquid separation, solid slag and leach liquor are analyzed its chemical constitution respectively, and analytical results calculates zinc leaching rate 96.45% in view of the above.
Embodiment 4
The zinc hydrometallurgy factory leached mud of embodiment 3 is added water stir and make even ore pulp, admix 93% industrial sulphuric acid rapidly and mix slag: water: acid (mass ratio) is 1: 0.5: 0.6.Mix than 1: 1.2 with concentrate quality by former slag with the zinc ore concentrate that contains 45.34%Zn and 36.37%S after placing 0.5h, behind 800 ℃ of following roasting 0.5h, add water to liquid-solid ratio and while hot leach 5h at 5: 1.Leach the back solid-liquid separation, solid slag and leach liquor are analyzed its chemical constitution respectively, and analytical results calculates zinc leaching rate 89.63% in view of the above.
Embodiment 5
Water wet-milling to 60% solid of weight such as the electric-arc furnace steelmaking flue dust that portion is contained 22.78%Zn, 31.28%Fe adds is admixed commercially available 95%w/w industrial sulphuric acid rapidly to mix by behind the 0.074mm sieve aperture, makes slag: water: acid (mass ratio) is 1: 1: 1.Mix than 1: 1 with concentrate quality by former slag with the copper and zinc bulk concentrate that contains 13.02%Cu, 11.23%Zn and 38.49%S after placing 1h, at 450 ℃ of following roasting 2h, calcining directly enters and leaches 3h in the water while hot, leaches solid-to-liquid ratio 1: 3.The leaching yield of calculating by the chemical analysis results of leach liquor and slag is: zinc 93.85%, copper 95.13%.
Embodiment 6
The electric-arc furnace steelmaking flue dust of a embodiment 5 is added water wet-milling to 60% solid of 2 times of weight by behind the 0.074mm sieve aperture, dense to slag: water (mass ratio) is 1: 0.6, and admix commercially available 95%w/w industrial sulphuric acid rapidly and mix, make slag: water: acid (mass ratio) is 1: 0.6: 0.8.Mix than 1: 1 with concentrate quality by former slag with the copper and zinc bulk concentrate that contains 13.02%Cu, 11.23%Zn and 38.49%S after placing 0.5h, at 400 ℃ of following roasting 3h, calcining directly enters and leaches 3h in the water while hot, leaches solid-to-liquid ratio 1: 8.The leaching yield of calculating by the chemical analysis results of leach liquor and slag is: zinc 90.56%, copper 94.08%.
Claims (9)
1. a method that reclaims valuable metal from metallurgical waste comprises the steps:
1) acidolysis: metallurgical waste is mixed with water, sulfuric acid, make its reaction decomposes to discharge wherein bonded valuable metal;
2) roasting: with the material after the acidolysis mix with the sulphur material compound, bubbling air roasting then; Wherein said sulphur material is the material that contains non-oxide attitude sulphur component, and maturing temperature is controlled at 450 ℃～800 ℃;
3) leach: after the roasting, calcining adds water logging and goes out;
4) reclaim: the solid-liquid separation on ore pulp after will leaching obtains containing the solution of valency metal, for the valuable metal that reclaims wherein.
2. recovery method according to claim 1 is characterized in that, described metallurgical waste is metallurgical slag, leached mud and the flue dust that contains the valency metal.
3. recovery method according to claim 1 is characterized in that, in the described step 1) acidolysis, the mass ratio of described metallurgical waste, sulfuric acid and water is 1: (0.3～3): (0.5～2).
4. according to claim 1 or 3 described recovery methods, it is characterized in that described step 1) acidolysis comprises the steps:
1) sizes mixing: ground metallurgical waste and water are mixed and made into ore pulp;
2) mix acid: admix sulfuric acid to the slip that mixes up, slip and sulfuric acid are mixed;
3) slaking: the material that will mix after the acid is placed more than the 0.5h.
5. recovery method according to claim 1 is characterized in that step 2) described sulphur material package draws together sulphide concentrate or sulphur.
6. recovery method according to claim 1 is characterized in that, described step 2) in, the ratio that the total amount of non-oxide attitude sulphur accounts for the compound total amount in the compound is not less than 6%w/w.
7. recovery method according to claim 1 is characterized in that, described step 2) in, maturing temperature is controlled at 500 ℃～750 ℃.
8. recovery method according to claim 1 is characterized in that, described step 2) in, roasting time is controlled at 0.5h～5h.
9. recovery method according to claim 1 is characterized in that, in the described step 3), free vitriolic concentration is less than 30g/L in the gained leach liquor.
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| CN102851511A (en) * | 2012-08-29 | 2013-01-02 | 江西自立资源再生有限公司 | Method for reducing contents of zinc and copper in smelting ash acid leaching residues |
| CN103060566A (en) * | 2011-10-21 | 2013-04-24 | 湖南创元铝业有限公司 | Method for recovering aluminum oxide, chlorine salt and villiaumite from aluminum ash |
| CN103182346A (en) * | 2011-12-30 | 2013-07-03 | 北京有色金属研究总院 | Novel process for improving grade of sulfate cinder iron |
| CN105112676A (en) * | 2015-09-09 | 2015-12-02 | 中南大学 | Method for recovering iron in fayalite metallurgy slag through roasting of iron pyrite |
| CN107217146A (en) * | 2017-04-20 | 2017-09-29 | 云南永昌铅锌股份有限公司 | The method containing Zn scrap returns of processing |
| CN108384946A (en) * | 2018-03-16 | 2018-08-10 | 湖南腾驰环保科技有限公司 | A kind of leaded secondary material balling technique |
| CN108929954A (en) * | 2018-05-31 | 2018-12-04 | 西北矿冶研究院 | Method for efficiently recovering zinc from zinc leaching residues |
| CN115874062A (en) * | 2022-11-30 | 2023-03-31 | 新疆有色金属研究所 | Efficient acid-mixing curing leaching process for non-ferrous metal copper smelting slag |
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| CN103060566A (en) * | 2011-10-21 | 2013-04-24 | 湖南创元铝业有限公司 | Method for recovering aluminum oxide, chlorine salt and villiaumite from aluminum ash |
| CN103182346A (en) * | 2011-12-30 | 2013-07-03 | 北京有色金属研究总院 | Novel process for improving grade of sulfate cinder iron |
| CN102851511A (en) * | 2012-08-29 | 2013-01-02 | 江西自立资源再生有限公司 | Method for reducing contents of zinc and copper in smelting ash acid leaching residues |
| CN105112676A (en) * | 2015-09-09 | 2015-12-02 | 中南大学 | Method for recovering iron in fayalite metallurgy slag through roasting of iron pyrite |
| CN105112676B (en) * | 2015-09-09 | 2018-03-30 | 中南大学 | A kind of method of pyrite roasting fayalite class metallurgical slag recovery iron |
| CN107217146A (en) * | 2017-04-20 | 2017-09-29 | 云南永昌铅锌股份有限公司 | The method containing Zn scrap returns of processing |
| CN108384946A (en) * | 2018-03-16 | 2018-08-10 | 湖南腾驰环保科技有限公司 | A kind of leaded secondary material balling technique |
| CN108929954A (en) * | 2018-05-31 | 2018-12-04 | 西北矿冶研究院 | Method for efficiently recovering zinc from zinc leaching residues |
| CN115874062A (en) * | 2022-11-30 | 2023-03-31 | 新疆有色金属研究所 | Efficient acid-mixing curing leaching process for non-ferrous metal copper smelting slag |
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