JP4387618B2 - Method for recovering copper from copper converter slag - Google Patents
Method for recovering copper from copper converter slag Download PDFInfo
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- JP4387618B2 JP4387618B2 JP2001253337A JP2001253337A JP4387618B2 JP 4387618 B2 JP4387618 B2 JP 4387618B2 JP 2001253337 A JP2001253337 A JP 2001253337A JP 2001253337 A JP2001253337 A JP 2001253337A JP 4387618 B2 JP4387618 B2 JP 4387618B2
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- 239000002893 slag Substances 0.000 title claims description 108
- 239000010949 copper Substances 0.000 title claims description 87
- 229910052802 copper Inorganic materials 0.000 title claims description 73
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims description 67
- 238000000034 method Methods 0.000 title claims description 44
- 239000003638 chemical reducing agent Substances 0.000 claims description 78
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims description 59
- 229910052742 iron Inorganic materials 0.000 claims description 15
- 239000007787 solid Substances 0.000 claims description 9
- 229910052799 carbon Inorganic materials 0.000 claims description 8
- 239000002245 particle Substances 0.000 claims description 8
- 238000007664 blowing Methods 0.000 claims description 7
- 229910004298 SiO 2 Inorganic materials 0.000 claims description 5
- 238000011084 recovery Methods 0.000 claims description 4
- 229910052710 silicon Inorganic materials 0.000 claims description 4
- 229910018072 Al 2 O 3 Inorganic materials 0.000 claims description 3
- 229910052748 manganese Inorganic materials 0.000 claims description 3
- 229910052698 phosphorus Inorganic materials 0.000 claims description 3
- 238000009924 canning Methods 0.000 claims description 2
- 239000007789 gas Substances 0.000 description 50
- 238000006722 reduction reaction Methods 0.000 description 40
- 239000000571 coke Substances 0.000 description 15
- 238000005336 cracking Methods 0.000 description 12
- ATUOYWHBWRKTHZ-UHFFFAOYSA-N Propane Chemical compound CCC ATUOYWHBWRKTHZ-UHFFFAOYSA-N 0.000 description 8
- 230000000052 comparative effect Effects 0.000 description 7
- 230000005484 gravity Effects 0.000 description 7
- 229910000805 Pig iron Inorganic materials 0.000 description 6
- 238000006243 chemical reaction Methods 0.000 description 6
- 238000003723 Smelting Methods 0.000 description 5
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 5
- 239000003245 coal Substances 0.000 description 5
- 239000000428 dust Substances 0.000 description 5
- 238000002844 melting Methods 0.000 description 5
- 230000008018 melting Effects 0.000 description 5
- 229910052760 oxygen Inorganic materials 0.000 description 5
- 239000001301 oxygen Substances 0.000 description 5
- 238000002347 injection Methods 0.000 description 4
- 239000007924 injection Substances 0.000 description 4
- 239000001294 propane Substances 0.000 description 4
- 229910002091 carbon monoxide Inorganic materials 0.000 description 3
- 238000002485 combustion reaction Methods 0.000 description 3
- 230000000694 effects Effects 0.000 description 3
- 239000000463 material Substances 0.000 description 3
- 239000002184 metal Substances 0.000 description 3
- 229910052751 metal Inorganic materials 0.000 description 3
- PNEYBMLMFCGWSK-UHFFFAOYSA-N aluminium oxide Inorganic materials [O-2].[O-2].[O-2].[Al+3].[Al+3] PNEYBMLMFCGWSK-UHFFFAOYSA-N 0.000 description 2
- ODINCKMPIJJUCX-UHFFFAOYSA-N calcium oxide Inorganic materials [Ca]=O ODINCKMPIJJUCX-UHFFFAOYSA-N 0.000 description 2
- 238000007796 conventional method Methods 0.000 description 2
- 238000001816 cooling Methods 0.000 description 2
- 238000010438 heat treatment Methods 0.000 description 2
- SZVJSHCCFOBDDC-UHFFFAOYSA-N iron(II,III) oxide Inorganic materials O=[Fe]O[Fe]O[Fe]=O SZVJSHCCFOBDDC-UHFFFAOYSA-N 0.000 description 2
- 239000003208 petroleum Substances 0.000 description 2
- 238000004064 recycling Methods 0.000 description 2
- QPLDLSVMHZLSFG-UHFFFAOYSA-N Copper oxide Chemical compound [Cu]=O QPLDLSVMHZLSFG-UHFFFAOYSA-N 0.000 description 1
- 239000005751 Copper oxide Substances 0.000 description 1
- 239000007795 chemical reaction product Substances 0.000 description 1
- 229910000431 copper oxide Inorganic materials 0.000 description 1
- 238000002474 experimental method Methods 0.000 description 1
- 238000005188 flotation Methods 0.000 description 1
- 238000000227 grinding Methods 0.000 description 1
- 230000020169 heat generation Effects 0.000 description 1
- 239000007788 liquid Substances 0.000 description 1
- 238000011068 loading method Methods 0.000 description 1
- 239000000155 melt Substances 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 230000003647 oxidation Effects 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- 238000003672 processing method Methods 0.000 description 1
- 239000000047 product Substances 0.000 description 1
- 230000003252 repetitive effect Effects 0.000 description 1
- 230000006641 stabilisation Effects 0.000 description 1
- 238000011105 stabilization Methods 0.000 description 1
- 238000003756 stirring Methods 0.000 description 1
- 239000002699 waste material Substances 0.000 description 1
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Description
【０００１】
【発明が属する技術分野】
本発明は、銅製錬転炉工程において生成したスラグ中に含まれる銅の回収方法に関するものであり、より詳しく述べるなら、銅転炉スラグ中のFe3O4を還元し、スラグの粘性を低下させることによってスラグ中の銅分を粗銅に回収する銅転炉スラグからの銅の回収方法に関するものである。
【０００２】
【従来の技術】
銅転炉操業では、溶錬炉から送られるマット中のＦｅを酸化・スラグ化により除去している。この時生成する銅転炉スラグには30〜40％のFe3O4が含まれるためにスラグの粘性が高くなっており、これが主原因となって、転炉スラグの銅含有率は5〜10％（本明細書において百分率は、特記しない限り、質量％である）と高くなっている。このため、この銅転炉スラグは銅分の回収を目的として、前工程である溶錬炉への繰り返し処理により、あるいは、別の工程での処理により銅分回収が行われている。国内では、銅転炉スラグを固化してから粉砕し、その後浮選により銅分を回収するスラグ選鉱法が主として採用されている（資源素材学会誌、「資源と素材」1993. 12, Vol 109「非鉄製錬号」第954, 965頁,「資源と素材」1997,12,Vol.113「リサイクリング大特集号」第996頁左欄、最終パラグラフ）。このスラグ選鉱法は、スラグ破砕・摩鉱・選鉱・脱水の多くの工程を要し、より、簡便な処理方法が望まれている。
【０００３】
また、連続製銅炉を用いる第１工程で生成し、粗銅から分離されたスラグに第２工程の処理炉にて、コークス、石炭等の炭素質固体還元剤または気体還元剤を吹込んで、成分や性質が若干異なる各種溶融スラグ中に含まれる酸化銅及びFe3O4を還元し、銅品位１％以下のカラミと粗銅を得るスラグの処理法が特開昭５３−２２１１５号に提案されている。
【０００４】
さらに、溶融状態の銅転炉スラグに吹込まれる石油液化ガス（ＬＰＧ）からなる還元剤が吹込み管内および吹込み管先端においてクラッキング反応を起こすと局部的なスラグ温度の低下を招く。このクラッキング反応を防止するための酸素あるいは空気（以下この空気を「クラッキングエア」と言う）を吹込む方法が特開平０９−８７７６１号に提案されている。この方法によると、溶融スラグ中のFe3O4を還元し、スラグ中の銅を回収する際に、溶融スラグの温度低下が防止されるために回収効率が高められる。
【０００５】
上記のコークス、石炭等の炭素質固体還元剤あるいは石油液化ガス（ＬＰＧ）などの気体還元剤を溶融スラグ中に吹込む場合、溶融スラグに比べ還元剤の比重が小さいため、溶融スラグ中での滞留時間が短く、還元効率が40〜70％と低くなる欠点がある。続いて各還元剤の操業例を説明する。
先ず、コークス還元の場合、一般的に、還元効率は５０％であり、コークス１kgに対してインジェクションエア1.8Nm3の割合でコークスを吹込む操業が行われており、１kgのFe3O4をFeOに還元する過程で N2,CO2,H2O合計で0.16〜0.20Ｎｍ3の排ガスが発生する。
次にＬＰＧ還元剤の場合、一般的に、還元効率は65％であり、ＬＰＧ1kgにクラッキングエア0.5 Nm3、未燃ＬＰＧを燃焼するために炉出口でフリーエアを添加する操業が行われている。クラッキングエアの一部は発熱に寄与する。この条件では、1kgのFe3O4をFeOに還元するためにN2,CO2,H2O合計で0.20〜0.25Nm3の排ガスが発生する。この排ガスは1000〜1200℃から約200〜300℃への冷却処理、還元剤の未燃分であるすすとスラグから発生したZnO等を主成分とするダストの除塵処理を行なう必要があるので、排ガス量は少ない方が好適である。
【０００６】
【発明が解決しようとする課題】
従来の銅転炉スラグから銅を還元することにより銅を回収する方法では、LPG、コークスなどのC,H分から主として構成される還元剤を使用するために多量の排ガスが発生し、これを処理するための後工程の負担が増加する。したがって、本発明は銅転炉スラグからの銅分回収において、還元剤の還元効率が高く、かつ排ガス発生量が少ない銅転炉スラグからの銅の回収方法を提供することを目的とする。
【０００７】
【課題を解決するための手段】
還元効率を高めるためには、還元剤は、容易に溶融スラグ中のFe3O4を還元し、かつスラグよりも比重が大きく溶融スラグ中に沈降することが必要である。また、排ガス量を減少させるには、還元剤はガス成分となるＣ，Ｈ分が少ないことが必要である。さらに、還元剤に含まれるＣｕ等の有価物を含めて安価な物であれば更に好適である。このような認識に基づいて本発明は次の方法を提案する。
(1)銅転炉操業の造カン期に生成しかつFe3O4と銅分を含有する溶融状態のスラグに、気体状還元剤及び炭素質固体還元剤の少なくとも１種からなる第1の還元剤を吹込む銅転炉スラグからの銅の回収方法において、メタリック鉄を７０質量％以上含有する固体からなる第２の還元剤を銅転炉に併せて装入することにより、前記メタリック鉄により前記Fe3O4の一部をFeOに還元することを特徴とする方法。
(2)第２の還元剤の粒径が３０ｍｍ以下である(1)の方法。
(3)第２の還元剤のＣｕ品位が２０質量％以下である(1)の方法。
(4)第２の還元剤がメタリック鉄と、残部C,Si,Cu,P,Mn,Al2O3,SiO2及びCaOからなる群から選択された2種以上とからなる(1)〜(3)の何れか１項記載の方法。
(5)第1の還元剤を溶融状態のスラグに吹込むと同時に第２の還元剤を該溶融状態のスラグに装入する(1)〜(4)の何れか1項記載の方法。
(6)第1の還元剤を溶融状態のスラグに吹込み、この吹込みと交互に第２の還元剤を該溶融状態のスラグに装入する(1)〜(4)の何れか1項記載の方法。
(7)溶融状態のスラグに予め第２の還元剤を投入した後に第1の還元剤を吹込む(1)〜(4)の何れか1項記載の方法。
(8)第２の還元剤の装入をスラグ還元炉の炉口から行う(1)〜(7)の何れか1項記載の方法。以下、本発明の構成を詳しく説明する。
【０００８】
コークス、石炭等の炭素質固体第1の還元剤は銅転炉スラグより比重が小さいことからスラグ中に留まらず、還元効率は一般的に40〜60％である。また、還元に寄与するコークス中のＣ，Ｈはスラグ中のFe3O4から酸素を奪いCO2,CO,H2Oガスを発生する。一方、還元に寄与しないコークス中のＣ，Ｈは溶融スラグから放出されインジェクションエア・炉出口フリーエアにより燃焼しCO2,CO,H2Oガスとなる。炉出口では、これら還元反応生成物と燃焼生成物の両方のCO2,CO,H2Oガスに、インジェクションエア・フリーエア中のＮ2を加えた排ガスが発生する。この排ガスの温度は1000〜1200℃である。なお、排ガス中には、一般に、スラグから発生したZnO等を主成分とするダストが含まれているので、ガス冷却・除塵処理のために排ガス処理設備が設けられている。コークス、石炭などの還元剤を使用すると、上記したガスの他にコークス中の未燃Ｃなどが排ガスに含まれまたガス量が増大するために、ガス量に応じた排ガス処理コストを要する。ＬＰＧ等の気体の還元剤を使用する場合も、コークス還元時と同様に、排ガス量に応じた排ガス処理設備、排ガス処理コストを要する。
【０００９】
本発明において、第1の還元剤は、ＬＰＧ、プロパンなどの炭化水素質気体還元剤、及び石炭、コークス等の炭素質固体還元剤である。第1の還元剤は、気体還元剤はそのまま、あるいはクラッキングエアとともに、固体還元剤はインジェクションエアの吹込みとともに吹き込んで、造カン期で生成し、還元炉に移された溶融スラグの攪拌を行うと同時に該還元剤を溶融スラグと接触させる。吹込みエアはコークス１ｋｇ当り１．５〜２Nm3であることが好ましい。
本発明においてはメタリック鉄を約７０％以上含む第2の還元剤はスラグ還元炉に吹込みではなく装入する。「装入」とはクレーン、スコップなどでの投入、シュートなどからの落下、支持治具での所定位置へのスラグ浴表面への移動など、液体・気体を補助手段としない方法で溶融スラグと接触させる方法である。この手段によると吹込み用ガスを使用しないから、スラグ還元炉スラグ還元処理時の排ガス発生量を大幅に削除できる。しかも、ノズルなどからメタリック鉄を吹込むとノズルの磨耗が起こり好ましくないが、投入法は極めて簡便である。投入法としては転炉の炉口から投入することが最も簡便である。少量の第２の還元剤をコークスなどと混合吹込むことは可能はであるが、吹込みエア量が第２の還元剤混合分だけ増加し、排ガス発生量が多くなるので、好ましくはない。第2の還元剤におけるメタリック鉄の残部は、全体の比重が７を超えない物質であれば、特に制限はなく、また全く含まれなくともよい。残部の成分は例えばＣ、Ｓｉなどの還元剤であってもよい。これらの成分からなる第２の還元剤は銑鉄、スクラップ鉄などである。また、副成分はＡｌ2Ｏ3，ＳｉＯ2及びＣａＯなどのスラグ成分、Ｃｕなどの有価金属であってもよい。さらに，Ｐ，Ｍｎなど製錬に影響の少い成分であってもよい。但し、Ｃｕ品位は２０％以下の範囲であることが好ましい。これらの残部とメタリック鉄からなる還元剤は上記成分の３種以上を含むリサイクル資源であってもよい。
【００１０】
第2の還元剤の投入方法としては前記(5)の方法が(6),(7)の方法よりも操業の安定化の面から好ましい。前記(7)の方法によると添加された第２の還元剤が溶融スラグ層を貫通して溶融物の底に溜まるから、第1の還元剤の吹込みに時に溶融スラグを十分に攪拌することが必要である。
【００１１】
第1の還元剤による還元反応は次のとおりである。
C +2 [O] → CO2 ・・・・(1)
H2 + [O] →H2 O ・・・ (2)
第２の還元剤による還元反応は次のとおりである。
Fe + [O] →[FeO] ・・・・(3)
C +2 [O] → CO2 ・・・・(4)
Si +2 [O] →[ SiO2 ]・・・・(5)
上記反応式における[O] は スラグ中のFe3O4の酸素を指す。これら各反応式によりスラグから除去される酸素の割合が、第２の還元剤のうちガス発生を伴わない(3),(5)式による割合が１０〜５０％、第1の還元剤によるものが９０〜５０％であることが好ましい。第２の還元剤による上記酸素除去割合が１０％未満であると、排ガス発生量削減効果が少なく、一方５０％を超えると溶融スラグの攪拌が困難になる。
【００１２】
本発明による還元処理前の銅転炉スラグは、一般に組成が30 〜35%FeO, 25 〜35% Fe3O4、20〜25% SiO2、 5〜10％Cu、3〜6％ ZnOであり、また温度は1250〜1330℃である。
【００１３】
【作用】
本発明の第2の還元剤に主成分として含まれるFeは、
Fe304＋Fe＝4FeO ・・・・・(6)
の反応によりFe304を還元し、しかもガスを発生しない。また、スラグの比重４〜５に比べて、メタリック鉄を７０％以上含有する第2の還元剤は比重が６．５〜７と重いため、スラグ内に留まり、還元効率はほぼ１００％と推定される。一方、この比重差では第2の還元剤の一部分は溶湯の底部に溜まり、溶融スラグ中Fe3O4還元に寄与しないことも予測される。これを防止するため、第２の還元剤の粒径を３０ｍｍ以下として溶融スラグ中に留まりやすくし、ＬＰＧ、エア等の気体を利用して溶融スラグを攪拌させて溶融スラグと第2の還元剤が十分に接触し、反応することが必要である。このような方法により、銅転炉スラグ還元処理時の発生排ガス量を大幅に削減することができる。あるいは、排ガス発生量を同量とするならば、還元処理時間の短縮、スラグ処理能力の増大を図ることができる。
以下に実施例により本発明を詳しく説明する。
【００１４】
【実施例】
（実験例）
5.2%Cu, 28.0%マグネタイト(Fe3O4 )を含有する銅転炉スラグ1.4kgをアルミナルツボに装入して、外部抵抗加熱方式の電気炉で銅転炉スラグを溶融し1250℃に保持した。このルツボに上方から挿入し,先端を溶融スラグ表面上２０ｍｍに保持した直径16mmのノズルからプロパン0.20g/minとクラッキングエア0.10リットル/minの混合ガスを６０分間吹込むと同時に、銑鉄粒を0.7g/minで連続的に投入した。この銑鉄粒は、一般ゴミの「直接溶融・資源化プラント」から発生したもので、組成は82%Fe, 3.0%C, 1.4%Si,4.5%Cuであり、粒径は1mm〜10mmφであった。
還元処理３０分後のスラグのFe3O4含有率は11.2％、Ｃｕ含有率は0.95％、６０分後のスラグのFe3O4含有率は７．６％、Ｃｕ含有率は０．７４％であった。この時の還元時間の経過に伴うスラグ中のFe3O4含有率の変化を図１に、スラグ中のＣｕ含有率の変化を図２に実験例として示した。
【００１５】
（比較実験例）
5.2%Cu, 28.0%マグネタイト(Fe3O4 )を含有する銅転炉スラグ1.4kgをアルミナルツボに装入して、外部抵抗加熱方式の電気炉で銅転炉スラグを溶融し1250℃に保持した。このルツボに上方から挿入し、先端を溶融スラグ表面上２０ｍｍに保持した直径１６ｍｍのノズルからプロパン0.31g/minとクラッキングエア0.16リットル/minの混合ガスを６０分間吹込んだ。
還元処理３０分後のスラグのFe3O4含有率は13.7％、Ｃｕ含有率は1.29％であり、６０分後のスラグのFe3O4含有率は8.9％,Ｃｕ含有率は0.80％であった。この時の還元時間の経過に伴うスラグ中のFe3O4含有率の変化を図１に、スラグ中のＣｕ含有率の変化を図２に比較例として示した。
以上のデータから算出した推定排ガス量と還元効率の比較を表１に示す。
【００１６】
【表１】
【００１７】
以上説明したFe3O4を還元することにより銅転炉スラグからＣｕ分を回収する方法において、還元剤としてＬＰＧのみを使用する場合に比べて、還元剤の一部に一般ゴミの「直接溶融・資源化プラント」から発生した銑鉄粒を使用することで、図１に示すようにFe3O4還元効率及び図２に示すようにＣｕ回収効率を悪化させることなく、上表に示すように、排ガス量の大幅な削減を達成し、ＬＰＧと銑鉄粒を合計した全体の還元効率の向上を達成できることが判明した。次にスラグ還元炉による実操業においてこれらの効果が具体的にどのように達成されるかを示す。
【００１８】
比較操業例 Fe3O4を28.0%含有する銅転炉スラグ50tが入れられている銅転炉スラグ還元炉においてスラグ中のFe3O4を1時間で8.0％に還元する場合、LPG還元効率６５％、LPG1kgにつきクラッキングエア0.5Nm3の条件では、吹込みLPG297kg/h＋クラッキングエア148Nm3/h、未燃ＬＰＧの炉出口での燃焼のためのフリーエア1330Nm3/hが必要となる。この時の、スラグ還元炉出口では温度が1000〜1200℃の排ガスが2240Nm3/hが発生する。
【００１９】
発明操業例
比較操業例と同じ条件のスラグを還元する場合、スラグ量の２％に相当する銑鉄粒（第２の還元剤）1000kg/hを投入し、かつLPG（第１の還元剤）161kg/h＋クラッキングエア80Ｎｍ3／ｈを吹込む。未燃LPGの炉出口での燃焼のためのフリーエア720Nm3/hが必要となる。この時の、還元炉出口では、1000〜1200℃の排ガスが1280Nm3/h発生する。
【００２０】
以上の比較操業例と発明操業例の対比から、還元処理転炉スラグ量の２％の銑鉄粒（第２の還元剤）を併用することにより、同量のスラグを同時間で処理する場合、ガス冷却、除塵等の排ガス処理量は５５〜６０％に削除でき、排ガス処理設備の縮小、処理コストの削除が可能となることが理解できよう。また、排ガス量を同じ，2240Nm3/hとすると、銅転炉スラグ５０tを３４分で還元処理でき、同規模設備であれば、同時間で１．７〜１．８倍の銅転炉スラグを還元処理できる。
【００２１】
【発明の効果】
本発明の銅転炉スラグからの銅の回収方法によれば、従来法の一つであるＬＰＧ＋クラッキングエアのみを還元剤として使用する場合より、還元効率が向上し、排ガス量を低減できる。
【図面の簡単な説明】
【図１】 実験例、比較実験例での還元時間に対するスラグ中Fe3O4含有率の変化を示すグラフである。
【図２】 実験例、比較実験例での還元時間に対するスラグ中Ｃｕ含有率の変化を示すグラフである。[0001]
[Technical field to which the invention belongs]
The present invention relates to a method for recovering copper contained in slag produced in a copper smelting converter process. More specifically, the present invention reduces Fe 3 O 4 in a copper converter slag and reduces the viscosity of the slag. It is related with the collection | recovery method of the copper from the copper converter slag which collect | recovers the copper content in slag to rough copper by making it.
[0002]
[Prior art]
In the copper converter operation, Fe in the mat sent from the smelting furnace is removed by oxidation and slag conversion. The copper converter slag produced at this time contains 30 to 40% Fe 3 O 4, so the viscosity of the slag is high, which is mainly due to the copper content of the converter slag being 5 to 5%. 10% (in this specification, percentage is mass% unless otherwise specified). For this reason, the copper converter slag is recovered for the purpose of recovering the copper content by repetitive processing to the smelting furnace, which is the previous process, or by processing in another process. In Japan, the slag beneficiation method is mainly adopted, in which the copper converter slag is solidified and then pulverized, and then the copper content is recovered by flotation (Journal of Resources and Materials, “Resources and Materials”, 1993. 12, Vol 109 “Non-ferrous metal smelting”, pages 954, 965, “Resources and materials”, 1997, 12, Vol. 113, “Special issue on recycling”, page 996, left column, last paragraph). This slag beneficiation method requires many steps of slag crushing, grinding, beneficiation, and dewatering, and a simpler processing method is desired.
[0003]
In addition, a carbonaceous solid reducing agent such as coke and coal or a gas reducing agent is blown into the slag produced in the first step using a continuous copper furnace and separated from the crude copper in the second step furnace. Japanese Laid-Open Patent Publication No. 53-22115 proposes a method for treating slag by reducing copper oxide and Fe 3 O 4 contained in various molten slags having slightly different properties and obtaining copper and grade copper or less. Yes.
[0004]
Furthermore, when a reducing agent made of petroleum liquefied gas (LPG) blown into a molten copper converter slag causes a cracking reaction in the blow pipe and at the tip of the blow pipe, a local decrease in slag temperature is caused. Japanese Laid-Open Patent Application No. 09-87761 proposes a method of blowing oxygen or air (hereinafter referred to as “cracking air”) for preventing this cracking reaction. According to this method, when Fe 3 O 4 in the molten slag is reduced and copper in the slag is recovered, the temperature of the molten slag is prevented from being lowered, so that the recovery efficiency is improved.
[0005]
When a carbonaceous solid reducing agent such as coke or coal or a gaseous reducing agent such as petroleum liquefied gas (LPG) is blown into the molten slag, the specific gravity of the reducing agent is smaller than that of the molten slag. There are disadvantages that the residence time is short and the reduction efficiency is as low as 40 to 70%. Then, the example of operation of each reducing agent is demonstrated.
First, in the case of coke reduction, generally, the reduction efficiency is 50%, and the operation of injecting coke at a rate of 1.8 Nm 3 of injection air to 1 kg of coke is performed, and 1 kg of Fe 3 O 4 is supplied. In the process of reducing to FeO, exhaust gas of 0.16 to 0.20 Nm 3 is generated in total in N 2 , CO 2 , and H 2 O.
Next, in the case of an LPG reducing agent, the reduction efficiency is generally 65%, and cracking air 0.5 Nm 3 is added to 1 kg of LPG, and free air is added at the furnace outlet to burn unburned LPG. . Part of the cracking air contributes to heat generation. Under this condition, in order to reduce 1 kg of Fe 3 O 4 to FeO, exhaust gas of 0.20 to 0.25 Nm 3 is generated in total of N 2 , CO 2 , and H 2 O. This exhaust gas needs to be cooled from 1000 to 1200 ° C to about 200 to 300 ° C, and dust removal treatment of dust mainly composed of ZnO generated from slag when it is unburned by the reducing agent. A smaller amount of exhaust gas is preferred.
[0006]
[Problems to be solved by the invention]
In the conventional method of recovering copper by reducing copper from copper converter slag, a large amount of exhaust gas is generated due to the use of a reducing agent mainly composed of C and H components such as LPG and coke. This increases the burden on the subsequent process. Accordingly, an object of the present invention is to provide a method for recovering copper from a copper converter slag having a high reducing agent reduction efficiency and a small amount of exhaust gas generation in recovering the copper content from the copper converter slag.
[0007]
[Means for Solving the Problems]
In order to increase the reduction efficiency, the reducing agent needs to easily reduce Fe 3 O 4 in the molten slag and to have a higher specific gravity than the slag and settle in the molten slag. In order to reduce the amount of exhaust gas, the reducing agent needs to have a small amount of C and H as gas components. Furthermore, it is more suitable if it is an inexpensive thing including valuables, such as Cu contained in a reducing agent. Based on this recognition, the present invention proposes the following method.
(1) A first slag formed of at least one of a gaseous reducing agent and a carbonaceous solid reducing agent is formed in a molten slag containing Fe 3 O 4 and copper, which is formed during the canning stage of the copper converter operation. In the method for recovering copper from a copper converter slag into which a reducing agent is blown, the metallic iron is charged by adding a second reducing agent made of a solid containing 70 % by mass or more of metallic iron to the copper converter. A method characterized in that a part of the Fe 3 O 4 is reduced to FeO.
(2) The method according to (1), wherein the particle size of the second reducing agent is 30 mm or less.
(3) The method of (1), wherein the second reducing agent has a Cu quality of 20% by mass or less.
(4) The second reducing agent is composed of metallic iron and two or more selected from the group consisting of the balance C, Si, Cu, P, Mn, Al 2 O 3 , SiO 2 and CaO (1) to The method according to any one of (3).
(5) The method according to any one of (1) to (4), wherein the second reducing agent is charged into the molten slag at the same time as the first reducing agent is blown into the molten slag.
(6) The first reducing agent is blown into the molten slag, and the second reducing agent is charged into the molten slag alternately with this blowing, any one of (1) to (4) The method described.
(7) The method according to any one of (1) to (4), wherein the first reducing agent is blown after the second reducing agent has been added to the molten slag in advance.
(8) The method according to any one of (1) to (7), wherein the second reducing agent is charged from the furnace port of the slag reduction furnace. Hereinafter, the configuration of the present invention will be described in detail.
[0008]
The carbonaceous solid first reducing agent such as coke and coal has a specific gravity smaller than that of the copper converter slag, so it does not stay in the slag, and the reduction efficiency is generally 40 to 60%. Further, C and H in the coke contributing to reduction take oxygen from Fe 3 O 4 in the slag and generate CO 2 , CO, and H 2 O gas. On the other hand, C and H in coke that do not contribute to reduction are released from the molten slag and burned by injection air / furnace outlet free air to become CO 2 , CO, H 2 O gas. At the furnace exit, exhaust gas is generated by adding N 2 in the injection air / free air to the CO 2 , CO, H 2 O gas of both the reduction reaction product and the combustion product. The temperature of this exhaust gas is 1000-1200 degreeC. In addition, since the exhaust gas generally contains dust mainly composed of ZnO generated from slag, an exhaust gas treatment facility is provided for gas cooling and dust removal treatment. When a reducing agent such as coke or coal is used, the amount of gas containing unburned C in the coke in addition to the above gas increases in the exhaust gas, so that an exhaust gas treatment cost corresponding to the amount of gas is required. When a gaseous reducing agent such as LPG is used, an exhaust gas treatment facility and an exhaust gas treatment cost corresponding to the amount of exhaust gas are required as in the case of coke reduction.
[0009]
In the present invention, the first reducing agent is a hydrocarbonaceous gas reducing agent such as LPG or propane, and a carbonaceous solid reducing agent such as coal or coke. As for the first reducing agent, the gaseous reducing agent is used as it is or with cracking air, and the solid reducing agent is blown together with the injection air, and the molten slag generated in the can-making stage is stirred into the reducing furnace. At the same time, the reducing agent is brought into contact with the molten slag. The blowing air is preferably 1.5 to 2 Nm 3 per kg of coke.
In the present invention, the second reducing agent containing about 70% or more of metallic iron is charged into the slag reduction furnace instead of being injected. “Loading” refers to melting slag by a method that does not use liquid or gas as auxiliary means, such as charging with a crane or scoop, dropping from a chute, or moving to a slag bath surface to a specified position with a support jig. It is a method of contacting. According to this means, since the blowing gas is not used, the amount of exhaust gas generated during the slag reduction furnace slag reduction treatment can be largely eliminated. Moreover, when metallic iron is blown from a nozzle or the like, the nozzle wears and is not preferable, but the charging method is very simple. As the charging method, it is most convenient to input from the furnace port of the converter. Although it is possible to mix and blow a small amount of the second reducing agent with coke or the like, it is not preferable because the amount of blown air increases by the amount of the second reducing agent mixed and the amount of exhaust gas generated increases. The remainder of the metallic iron in the second reducing agent is not particularly limited as long as it has a total specific gravity of not more than 7, and may not be included at all. The remaining component may be a reducing agent such as C or Si. The second reducing agent composed of these components is pig iron, scrap iron and the like. The subcomponent may be a slag component such as Al 2 O 3 , SiO 2 and CaO, or a valuable metal such as Cu. Further, it may be a component having little influence on smelting such as P and Mn. However, the Cu quality is preferably in the range of 20% or less. The reductant comprising these balance and metallic iron may be a recycled resource containing three or more of the above components.
[0010]
As the second reducing agent charging method, the method (5) is more preferable than the methods (6) and (7) in terms of stabilization of operation. According to the method (7), since the added second reducing agent penetrates the molten slag layer and accumulates at the bottom of the melt, the molten slag is sometimes sufficiently stirred for blowing the first reducing agent. is required.
[0011]
The reduction reaction by the first reducing agent is as follows.
C +2 [O] → CO 2 ... (1)
H 2 + [O] → H 2 O (2)
The reduction reaction by the second reducing agent is as follows.
Fe + [O] → [FeO] (3)
C +2 [O] → CO 2 ... (4)
Si +2 [O] → [SiO 2 ] (5)
[O] in the above reaction formula indicates Fe 3 O 4 oxygen in the slag. The proportion of oxygen removed from the slag by each of these reaction formulas is 10 to 50% according to the formulas (3) and (5) with no gas generation in the second reducing agent, which is due to the first reducing agent. Is preferably 90 to 50%. If the oxygen removal ratio by the second reducing agent is less than 10%, the effect of reducing the amount of exhaust gas generated is small, whereas if it exceeds 50%, it is difficult to stir the molten slag.
[0012]
The copper converter slag before the reduction treatment according to the present invention is generally composed of 30-35% FeO, 25-35% Fe 3 O 4 , 20-25% SiO 2 , 5-10% Cu, 3-6% ZnO. And the temperature is 1250-1330 ° C.
[0013]
[Action]
Fe contained as a main component in the second reducing agent of the present invention,
This reaction reduces
Hereinafter, the present invention will be described in detail by way of examples.
[0014]
【Example】
(Experimental example)
Inserting 1.4 kg of copper converter slag containing 5.2% Cu and 28.0% magnetite (Fe 3 O 4 ) into an alumina crucible and melting the copper converter slag in an external resistance heating type electric furnace and maintaining it at 1250 ° C did. This crucible was inserted from above, and a mixed gas of propane 0.20 g / min and cracking air 0.10 liter / min was blown in through a nozzle with a diameter of 16 mm whose tip was held 20 mm above the surface of the molten slag. It was continuously charged at g / min. These pig iron grains originated from the “direct melting and resource recycling plant” of general waste, have a composition of 82% Fe, 3.0% C, 1.4% Si, 4.5% Cu, and a particle size of 1 mm to 10 mmφ. It was.
The Fe 3 O 4 content of the slag after 30 minutes of the reduction treatment is 11.2%, the Cu content is 0.95%, the Fe 3 O 4 content of the slag after 60 minutes is 7.6%, and the Cu content is 0.74. %Met. The change in the Fe 3 O 4 content in the slag with the lapse of the reduction time at this time is shown in FIG. 1, and the change in the Cu content in the slag is shown in FIG. 2 as an experimental example.
[0015]
(Comparative experiment example)
Inserting 1.4 kg of copper converter slag containing 5.2% Cu and 28.0% magnetite (Fe 3 O 4 ) into an alumina crucible and melting the copper converter slag in an external resistance heating type electric furnace and maintaining it at 1250 ° C did. The crucible was inserted from above, and a mixed gas of propane 0.31 g / min and cracking air 0.16 liter / min was blown in for 60 minutes from a nozzle having a diameter of 16 mm whose tip was held 20 mm above the surface of the molten slag.
The Fe 3 O 4 content of the slag after 30 minutes of reduction treatment is 13.7% and the Cu content is 1.29%. The Fe 3 O 4 content of slag after 60 minutes is 8.9% and the Cu content is 0.80%. there were. The change in the Fe 3 O 4 content in the slag with the lapse of the reduction time at this time is shown in FIG. 1, and the change in the Cu content in the slag is shown in FIG. 2 as a comparative example.
Table 1 shows a comparison between the estimated exhaust gas amount calculated from the above data and the reduction efficiency.
[0016]
[Table 1]
[0017]
In the method of recovering Cu from the copper converter slag by reducing Fe 3 O 4 as described above, compared with the case where only LPG is used as the reducing agent, “direct melting” of general garbage is part of the reducing agent. As shown in the above table, without using the Fe 3 O 4 reduction efficiency as shown in FIG. 1 and the Cu recovery efficiency as shown in FIG. It has been found that a significant reduction in the amount of exhaust gas can be achieved, and the overall reduction efficiency can be improved by adding LPG and pig iron particles together. Next, it will be shown how these effects can be achieved in actual operation with a slag reduction furnace.
[0018]
When reducing the comparative operation example Fe 3 O 4 and Fe 3 O 4 in the slag to 8.0% in 1 hour in a copper converter slag reduction furnace copper converter slag 50t it is being input containing 28.0%, LPG reduction efficiency 65%, under the conditions of cracking the air 0.5 Nm 3 per LPG1kg, blow LPG297kg / h + cracking air 148 nm 3 / h, free air 1330 nm 3 / h for the combustion in the unburned LPG furnace exit is required. At this time, 2240 Nm 3 / h of exhaust gas having a temperature of 1000 to 1200 ° C. is generated at the outlet of the slag reduction furnace.
[0019]
Inventive operation example When reducing slag under the same conditions as the comparative operation example, pig iron particles (second reducing agent) equivalent to 2% of the slag amount, 1000 kg / h, and LPG (first reducing agent) 161 kg Blow / h + cracking air 80Nm 3 / h. 720Nm 3 / h of free air for combustion at the furnace outlet of unburned LPG is required. At this time, exhaust gas of 1000 to 1200 ° C. is generated at the reduction furnace outlet at 1280 Nm 3 / h.
[0020]
From the comparison between the above comparative operation example and the invention operation example, by using together the pig iron particles (second reducing agent) of 2% of the reduction treatment converter slag amount, when processing the same amount of slag in the same time, It can be understood that the amount of exhaust gas treatment such as gas cooling and dust removal can be deleted to 55 to 60%, and the exhaust gas treatment facility can be reduced and the processing cost can be deleted. Moreover, if the amount of exhaust gas is the same, 2240 Nm 3 / h, the copper converter slag 50t can be reduced in 34 minutes. Can be reduced.
[0021]
【The invention's effect】
According to the method for recovering copper from the copper converter slag of the present invention, the reduction efficiency can be improved and the amount of exhaust gas can be reduced as compared with the case where only LPG + cracking air, which is one of the conventional methods, is used as the reducing agent.
[Brief description of the drawings]
FIG. 1 is a graph showing a change in Fe 3 O 4 content in slag with respect to reduction time in experimental examples and comparative experimental examples.
FIG. 2 is a graph showing changes in Cu content in slag with respect to reduction time in experimental examples and comparative experimental examples.
Claims (8)
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| JP2001253337A JP4387618B2 (en) | 2000-08-29 | 2001-08-23 | Method for recovering copper from copper converter slag |
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| JP2000-258542 | 2001-02-13 | ||
| JP2001253337A JP4387618B2 (en) | 2000-08-29 | 2001-08-23 | Method for recovering copper from copper converter slag |
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| KR101448147B1 (en) * | 2012-09-10 | 2014-10-08 | 엘에스니꼬동제련 주식회사 | The recovery method of valuble metals included in slag at copper smelter |
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| JP4090219B2 (en) * | 2001-06-04 | 2008-05-28 | 日鉱金属株式会社 | Apparatus for charging iron content into copper smelting furnace and method of using the same |
| GB0313886D0 (en) * | 2003-06-16 | 2003-07-23 | Jha Animesh | Extraction route for Ti02 and alumina from bauxite and bauxitic residues,and titaniferrous deposits and wastes |
| DE102006052181A1 (en) * | 2006-11-02 | 2008-05-08 | Sms Demag Ag | A process for the continuous or discontinuous recovery of a metal or metals from a slag containing the metal or compound of the metal |
| JP5049311B2 (en) * | 2009-03-31 | 2012-10-17 | パンパシフィック・カッパー株式会社 | Method and system for dry treatment of converter slag in copper smelting |
| JP2012012707A (en) * | 2011-09-22 | 2012-01-19 | Pan Pacific Copper Co Ltd | Dry-type treating method and system for converter slag in copper refining |
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| KR101448147B1 (en) * | 2012-09-10 | 2014-10-08 | 엘에스니꼬동제련 주식회사 | The recovery method of valuble metals included in slag at copper smelter |
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