JPH0324238A - Lead smelting method - Google Patents
Lead smelting methodInfo
- Publication number
- JPH0324238A JPH0324238A JP15704989A JP15704989A JPH0324238A JP H0324238 A JPH0324238 A JP H0324238A JP 15704989 A JP15704989 A JP 15704989A JP 15704989 A JP15704989 A JP 15704989A JP H0324238 A JPH0324238 A JP H0324238A
- Authority
- JP
- Japan
- Prior art keywords
- lead
- furnace
- concentrate
- oxygen
- desulfurization
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Pending
Links
- 238000000034 method Methods 0.000 title claims description 52
- 238000003723 Smelting Methods 0.000 title claims description 36
- 239000012141 concentrate Substances 0.000 claims abstract description 40
- 238000007664 blowing Methods 0.000 claims abstract description 24
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 claims abstract description 23
- 229910052760 oxygen Inorganic materials 0.000 claims abstract description 23
- 239000001301 oxygen Substances 0.000 claims abstract description 23
- 239000002994 raw material Substances 0.000 claims abstract description 19
- 239000002893 slag Substances 0.000 claims abstract description 17
- 229940056932 lead sulfide Drugs 0.000 claims abstract description 14
- 229910052981 lead sulfide Inorganic materials 0.000 claims abstract description 14
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 10
- 239000003638 chemical reducing agent Substances 0.000 claims abstract description 10
- 229910000464 lead oxide Inorganic materials 0.000 claims abstract description 7
- 239000007789 gas Substances 0.000 claims description 21
- 239000003575 carbonaceous material Substances 0.000 claims description 7
- 238000006477 desulfuration reaction Methods 0.000 abstract description 48
- 230000023556 desulfurization Effects 0.000 abstract description 48
- 238000006243 chemical reaction Methods 0.000 abstract description 21
- HTUMBQDCCIXGCV-UHFFFAOYSA-N lead oxide Chemical compound [O-2].[Pb+2] HTUMBQDCCIXGCV-UHFFFAOYSA-N 0.000 abstract description 15
- 239000000779 smoke Substances 0.000 abstract description 15
- KEQXNNJHMWSZHK-UHFFFAOYSA-L 1,3,2,4$l^{2}-dioxathiaplumbetane 2,2-dioxide Chemical compound [Pb+2].[O-]S([O-])(=O)=O KEQXNNJHMWSZHK-UHFFFAOYSA-L 0.000 abstract description 4
- 229910052924 anglesite Inorganic materials 0.000 abstract description 4
- 239000000155 melt Substances 0.000 abstract description 2
- 235000002918 Fraxinus excelsior Nutrition 0.000 abstract 2
- 239000002956 ash Substances 0.000 abstract 2
- 238000006722 reduction reaction Methods 0.000 description 31
- 239000000243 solution Substances 0.000 description 18
- 239000003245 coal Substances 0.000 description 14
- 238000005266 casting Methods 0.000 description 12
- 238000002347 injection Methods 0.000 description 9
- 239000007924 injection Substances 0.000 description 9
- 238000012360 testing method Methods 0.000 description 8
- ATJFFYVFTNAWJD-UHFFFAOYSA-N Tin Chemical compound [Sn] ATJFFYVFTNAWJD-UHFFFAOYSA-N 0.000 description 5
- 238000005516 engineering process Methods 0.000 description 5
- 239000000446 fuel Substances 0.000 description 5
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 5
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 5
- ATUOYWHBWRKTHZ-UHFFFAOYSA-N Propane Chemical compound CCC ATUOYWHBWRKTHZ-UHFFFAOYSA-N 0.000 description 4
- 239000003795 chemical substances by application Substances 0.000 description 4
- 229910052802 copper Inorganic materials 0.000 description 4
- 239000010949 copper Substances 0.000 description 4
- 230000007423 decrease Effects 0.000 description 4
- 230000000694 effects Effects 0.000 description 4
- 239000000203 mixture Substances 0.000 description 4
- 230000003647 oxidation Effects 0.000 description 4
- 238000007254 oxidation reaction Methods 0.000 description 4
- 239000002245 particle Substances 0.000 description 4
- 230000003068 static effect Effects 0.000 description 4
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 description 3
- 238000001816 cooling Methods 0.000 description 3
- 238000000605 extraction Methods 0.000 description 3
- PIJPYDMVFNTHIP-UHFFFAOYSA-L lead sulfate Chemical compound [PbH4+2].[O-]S([O-])(=O)=O PIJPYDMVFNTHIP-UHFFFAOYSA-L 0.000 description 3
- 239000007788 liquid Substances 0.000 description 3
- 239000007787 solid Substances 0.000 description 3
- 239000011449 brick Substances 0.000 description 2
- 238000010586 diagram Methods 0.000 description 2
- 238000010438 heat treatment Methods 0.000 description 2
- 229910052742 iron Inorganic materials 0.000 description 2
- 238000004519 manufacturing process Methods 0.000 description 2
- 239000000463 material Substances 0.000 description 2
- 239000002184 metal Substances 0.000 description 2
- 229910052751 metal Inorganic materials 0.000 description 2
- 239000000843 powder Substances 0.000 description 2
- 239000001294 propane Substances 0.000 description 2
- 238000005245 sintering Methods 0.000 description 2
- 238000007711 solidification Methods 0.000 description 2
- 230000008023 solidification Effects 0.000 description 2
- 229910001220 stainless steel Inorganic materials 0.000 description 2
- 239000010935 stainless steel Substances 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- 229910052717 sulfur Inorganic materials 0.000 description 2
- UGFAIRIUMAVXCW-UHFFFAOYSA-N Carbon monoxide Chemical compound [O+]#[C-] UGFAIRIUMAVXCW-UHFFFAOYSA-N 0.000 description 1
- 229910001018 Cast iron Inorganic materials 0.000 description 1
- 241001062472 Stokellia anisodon Species 0.000 description 1
- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 description 1
- 239000002253 acid Substances 0.000 description 1
- 230000002411 adverse Effects 0.000 description 1
- 239000003570 air Substances 0.000 description 1
- 229910052787 antimony Inorganic materials 0.000 description 1
- 229910052785 arsenic Inorganic materials 0.000 description 1
- 239000013626 chemical specie Substances 0.000 description 1
- 238000002485 combustion reaction Methods 0.000 description 1
- 239000002131 composite material Substances 0.000 description 1
- 230000006378 damage Effects 0.000 description 1
- 238000000354 decomposition reaction Methods 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 230000003628 erosive effect Effects 0.000 description 1
- 238000001704 evaporation Methods 0.000 description 1
- 230000008020 evaporation Effects 0.000 description 1
- 239000003546 flue gas Substances 0.000 description 1
- 238000010574 gas phase reaction Methods 0.000 description 1
- 230000017525 heat dissipation Effects 0.000 description 1
- 238000007654 immersion Methods 0.000 description 1
- 229910052745 lead Inorganic materials 0.000 description 1
- XCAUINMIESBTBL-UHFFFAOYSA-N lead(ii) sulfide Chemical compound [Pb]=S XCAUINMIESBTBL-UHFFFAOYSA-N 0.000 description 1
- 150000002739 metals Chemical class 0.000 description 1
- 238000013508 migration Methods 0.000 description 1
- 230000005012 migration Effects 0.000 description 1
- 239000011812 mixed powder Substances 0.000 description 1
- 230000000877 morphologic effect Effects 0.000 description 1
- 230000001590 oxidative effect Effects 0.000 description 1
- 230000035515 penetration Effects 0.000 description 1
- 238000005502 peroxidation Methods 0.000 description 1
- 238000012545 processing Methods 0.000 description 1
- 239000002904 solvent Substances 0.000 description 1
- 238000003756 stirring Methods 0.000 description 1
- 239000011593 sulfur Substances 0.000 description 1
- 230000003685 thermal hair damage Effects 0.000 description 1
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
Description
【発明の詳細な説明】
（イ）技術分野
本発明は、硫化鉛精鉱や主成分がＰｂＳＯａやＰｂＯで
ある鉛滓等の鉛原料（以下、鉛精鉱という）から粗鉛を
得る連続鉛製錬法に関し、更に詳しくは脱硫炉中に粗鉛
を存在させる必要がなく、装入方法がランスを溶体に浸
漬させない上吹ランス方式であり、煙灰の発生暖が極め
て少ない鉛製錬技術に関するものである．
（口）従来技術
鉛精鉱の標準的な製錬法は、焼結一溶鉱炉法であり、該
方法は鉛精鉱を．焙焼，焼結して硫化鉛（Ｐ ｂ Ｓ）
を酸化鉛（ＰｂＯ）に変える．次に、得られた焼結鉱（
ｐｂｏ）を炭剤とともに溶鉱炉に装入して還元させ粗鉛
な得る２つの工程から構或されている．
しかしながら、該製錬法には焼結工程で発生するＳＯｌ
！ガス濃度が低いこと、焼結の繰返し焼鉱リが多いので
そのハンドリングが煩雑であること，また消費エネルギ
ーが大きいこと等の多くの欠点があり、その対策として
各種鉛製錬法が研究され提案されている．
その１つとして，近年特に銅製錬に広く適用されて来た
自溶炉法が鉛精鉱の溶錬法として試みられている．
しかし，該自溶炉法は高温反応の起こるシャフト部レン
ガの浸食が著るしいことに加え、気相反応であるため、
蒸気圧の高いＰｂＳを伴なう鉛精鉱の溶錬に際しては煙
灰の発生が膨大であり、従って副次的な処理が必要とな
り、低収率．高コスト化を招〈という大きな問題がある
．その例としては、Ｏｕｔｏ Ｋｕｍｐｕ法やＫｉｖ
ｃｓｔ法に代表される通りである．
また，横型の単一炉を用い、その炉内の左右で反応条件
を区分けして、酸化と還元反応を別々に行なう鉛精鉱の
連続製錬法が提案され、該法はＱＳＬ法として知られて
いる．
しかし、この方法は酸化と還元の全く逆向きの反応を単
一炉内で行なわせるといった困難さを伴なうため、極め
て高度な操炉技術が必要であり、そのため：９備が複雑
化してコストが著し〈高〈なるという欠点がある．
史に、最近ではインジェクション製錬法、即ち溶体中へ
原料を吹き込み溶錬する製錬技術が研究開発され、銅製
錬においては既に一部実用化されており、この方法を鉛
製錬分野へ適用する試みも既になされ、本出願人の出願
に係る特開昭６２−２２２０３２号「鉛製錬法」や特開
昭５９−２２６１３０号「鉛の連続直接製錬法」および
特開昭６３−１２８１３８号「電気炉鉛製錬法」等が提
案されている．
まず、特開昭６２−２２２０３２号の鉛製錬法は非自燃
物の硫酸鉛（ＰｂＳＯ４）を主戊分とする鉛原料の製錬
法であって、該プロセスは硫酸鉛原料に熱補償用の微粉
炭等を加え、ランスを通して炉内溶体中に吹き込んで分
解脱硫反応を行なう脱硫炉と、該脱硫炉で生戊した酸化
鉛（Ｐ ｂ Ｏ）を含むスラッグを受けて微粉炭などで
吹込み還元反応を行なう還元炉から構威されている．そ
して、脱硫炉出のスラッグが連続的に還元炉に装入され
るように通常２炉は樋で連結されているのである．
このプロセスの基本反応は硫酸鉛を分解する脱硫炉内で
の次の反応（１） （２）式と還元炉内での反応（３）
式からなるが，脱硫炉での相鉛の生成も可能である．
’ｐｂｓｏ４＋ｃ＋ｐｂｏ＋ｓｏ，＋ｃｏ （１）Ｓ
Ｏ，，＋ＣＯ−＋ Ｓｏ，，＋ＣＯ， （２
）ＰｂＯ＋局Ｃ − Ｐ ｂ ＋ｙ２ＣＯｐ
（３）そして，このプロセスの長所としては
■）低廉な石炭を熱源として使用できるので、エネルギ
ーの低コスト化ができること．
■）脱硫と還元という２反応を別々の炉で行うので，冶
金反応的に合理的であること．
■）設備が簡素で操業が容易で生産性が良好であること
．および
■）排ガス中のＳＯ２を硫酸として回収できること等が
挙げられている．
しかしながら、特開昭６２−２２２０３２号は硫酸鉛を
処理対象とするものであって、鉛精鉱（ＰｂＳ）の処理
ができるか否かについては何の記載もな〈，示唆する何
ものもないのである．次に、特開昭５９−２２６１３０
号の方法は，溶錬炉と電気炉である還元炉とから構成さ
れ，鉛精鉱をランスで溶錬炉の溶融鉛浴中に吹込み、生
成した溶融粗鉛とスラグを還元帯域に移すことにより，
スラグの還元を能率化せしめるプロセスであって、溶錬
炉には常に粗鉛が存在していることが必須の要件となっ
ている鉛製錬法である．即ち、粗鉛のＰｂＳに対する大
きな溶解能で溶錬反応が可能であることを要旨とし、鉛
精鉱を処理対象として開発された製錬法ではあるが、溶
錬炉で常に粗鉛を生戊せしめる必要があるため，当然に
ＰｂＳ活竜を相当に大きくしなければならないのでＰｂ
Ｓ形態での煙灰の発生が著しく，有価金属類の一次収率
の低下、ハンドリング等の操業面での生産性を低下させ
るなど多〈の問題がある．
次に、特開昭６３−１２８１３８号は主にｐｂＳＯ４か
らなる鉛滓と燃料，溶剤及び還元剤とを混合した混合粉
体を空気と共に電気炉内の高温溶体内に特定の吹込み圧
で吹き込むことにより、鉛さいを電気炉を用いて製錬す
る方法である．しかしながら、この方法は還元性雰囲気
でｐｂＳＯ４をｔ体とする鉛滓を直接高温溶体内に吹き
込むため、発生する煙灰量が増大し、その対策に重大な
問題がある．即ち、還元性雰囲気でｐｂとＳとが共存す
る場合には，優勢化学種がＰｂＳであり，非常に揮発し
易い形態となるため、大部分のＰｂ分はＰｂＳとなって
煙灰に移行すると推察され、しかも吹込みにより高温溶
体が撹拌されるので４煙灰の揮発が更に促進されるとい
う重大な問題が残されている．
（ハ）発明の開示
本発明は，上記のような諸問題を解消することを目的と
してなされた硫化鉛精鉱や主威分がＰｂＳＯ＆やＰｂＯ
である鉛滓等の鉛原料（本文中、鉛精鉱という）から粗
鉛を得る連続製錬法であり、即ち酸化鉛を含有するスラ
グ溶体に鉛精鉱を酸素又は酸素富化空気と共にランスを
該溶体に浸漬させることなく吹込んで該鉛精鉱中の鉛分
を酸化物とした後，還元炉に装入し、該溶体中に還元剤
を吹込んで粗鉛を得る煙灰発生量の極めて少ない鉛製辣
法を提供するものである．
以下、本発明を詳細に説明する．
本発明は、鉛精鉱を酸素もしくは酸素富化空気と共に鉛
酸化物を含む溶融訪中に吹き込むことによって，脱硫反
応が瞬間的に終了するとの知見を得，開発された鉛製錬
技術であって、画期的な鉛製錬法である．
本発明の主要反応装置は、脱硫炉と還元炉とを直列した
２炉からなり、該脱硫炉への鉛精鉱の装入方法が溶融鶴
内への上吹き吹込み方法（ランスを溶融鏝に浸漬させる
ことなく吹込む）であって．従って該脱硫炉では強酸化
性雰囲気で反応が進行するため、煙灰の発生率を著し〈
低く抑えることができるのである．
また、脱硫炉で生成する溶体は鉛酸化物（ＰｂＯ主体）
を含有する鎚のみであり，これを連続的に還元炉に還元
剤と共に装入供給して鉛分を粗鉛に還元生成させるので
ある．これが，本発明の特徴である．
吹込み方式による溶錬は，装入する鉛精鉱が溶体内にも
ぐって分散されるため，熱伝導が極めて良く、短時間で
反応に必要な高温状態となるので、反応が著しく加速さ
れる．
従ってＰｂＳは瞬間的に酸化されて溶融鋳に吸収される
．そのために、蒸発し易いＰｂＳ形態で存在する時間が
著し〈短縮されるので、蒸発が抑制され、煙灰発生率を
低く抑えることができるのである．
また、ＰｂＳをＰｂＯまで完全に酸化させるため，粗鉛
を生成させる場合のような反応上の制限がな〈、酸素ポ
テンシャル（ＰＯ２）を任意に太き〈取れるので、酸化
速度を加速できることもＰｂＳの形態比率の低下に効果
的である．さらに、鉛精鉱が溶体内に吹き込まれて溶融
鋳と充分に固液接触するので．ｆｌａｓｈ法の場合と異
なり，メカニカルなキャリーオーバーも減少するのであ
る．
鉛精鉱は脱硫炉における脱硫反応で硫＃原料となる排ガ
スや煙灰となる以外は鉛酸化物を含有する鎚に変わるが
，脱硫反応を継続させるために鋳の持つべき条件として
は、鐸中のＰｂ品位が４０％以上であるならば１０５０
℃以上の温度が保持できれば充分に吹込み溶錬が継続可
能である．Ｆｅ，Ｓｉ０２ ，ＢａＯおよびＣａＯ等の
造鐸成分に関しては、後続の還元炉で生成する鋳が適当
な組成ＷＡＨに収まるような含有比が確保されさえすれ
ばよいのである．
次に、ランスの吹込み速度は１２０鵬／ｓｅｃ以Ｌあれ
ば充分であるが、好ましくは１４０−１６０ｒａ／ｓｅ
ｃが適当である．吹込み速度が早すぎると，エネルギー
面で不利なこと、炉底レンガの損傷をＶめること、溶体
がジャンビングして炉から飛散し、排ガス処理系統の機
器に付着固化して各種トラブルの原因となること等の問
題があり好ましくない．
逆に、遅過ぎる場合には、装入物が溶融鋳中へ侵入でき
なくなり、溶錬反応が起きずに炉外ヘキャリーオーバー
してしまう．
また、吹込用ランスは通常単なるステンレスパイプであ
り、吹込み用の酸素または酸素富化空気で内壁から自己
冷却して炉内の熱アタックから保護していることから，
ランス内での風速が低下すれば当然のことながら冷却効
果が低下し、ランスが熱損傷しやすくなる．
また，吹込み用ランスの位置は，吹込みを停止した状態
即ち静止湯而から９０ｃ園又はこれより少なく，上方に
ランスの先端が位置するのが好まし１，Ｘ．
吹込み！ｆＸ料等の水分と粒度は、鉛精鉱で水分１％以
下．粒径２■■φ以下、粉状炭素質性物質で水分１％以
下、粒径２７５メッシュ以下が適当である．
次に．′：ｊＩＪ２工程の還元炉での還元剤の吹込み条
件は上記の脱硫炉の場合に準ずるが、吹込みの固体／ガ
ス比率が小さいので、ランス吹込み速度は若干速めで、
ランスの先端位置は低目の方が良い．
また、脱硫炉からの排ガスはＳＯｚ濃度が高＜．ＳＯｓ
／５０２比が小さいので，！＆酸工場での硫酸製造に適
している．
本発明法は上記の通りであり、次のような利点がある．
ａ）脱硫炉において粗鉛を生成させる必要がないので、
低品位の鉛精鉱及び低品位鉛滓類等に適用することがで
きる．
ｂ）熱量不足の場合は、粉状の炭素性物質を鉛原刺中へ
添加配合すればよい．
Ｃ）また，脱硫反応における過酸化状態をコントロール
するための脱酸剤にも、同様に粉状の炭素性物質を使用
できる．これは、脱硫反応が過酸化状態になると排ガス
中のＳＯＳ ／３０２比がＬ昇して悪影響を及ぼすから
である．
７ＪＳｌ図は本発明法による設備概念図を示すもので、
ｌは脱硫炉，２は還元炉，３は両炉１．２を連結する脱
硫炉出鋳用樋である．鉛精鉱と脱酸剤及び粉状の炭素性
物質としての微粉炭は，酸素富化空気と共にランスパイ
プ４を経て炉内に吹き込まれる．
５は還元剤や微粉炭及び空気、又は酸素富化空気を吹き
込むためのランスパイプであり，６は含ＳＯ２ガス排出
孔，７は還元炉ガス排出孔である．なお、丘記した第ｌ
工程の粉状炭素性物質に代えて液体燃料や気体燃料とし
てもよく，更に第１工程の脱酸剤や第２工程の還元剤は
粉状炭素性物質でもよいし液体燃料や気体燃料でもよい
．以下，本発明を実施例により説明する．（二）実施例
実施例１
実証炉として、第１図に示すような装置を用いて硫化鉛
精鉱処理試験を行なった．脱硫炉および萌元炉の操業条
件は下記の通りとした．脱硫炉（第１工程）
原料吹込み量：１．８〜３．５ｔ／Ｈ
ｌｇＩ素富化空気量＝８００〜１２００Ｎ嘗３ ／Ｈ酸
素濃度＝４９〜７４％
ランス吹込速度： １ ２ ０ 〜１ ６ ０Ｎｓ／ｓ
ｅｅ微粉炭：原料重量に対して０〜５ｗｔ％ランス位置
：静止湯面上８０ｃｍ（ランス先ｇ）黛元炉（第２工程
）
微粉炭：１２０〜２ ５ ０ Ｋｇ／Ｈ空気量：９００
〜１ １ ５ ０Ｎｍ３／Ｈランス吹込速度： ｌ ８
０〜２ ５ ０ＮＥ／ｓｅｅランス位置：静止湯面上
３０〜６０ｃｍまた，供試硫化鉛精鉱の化学成分品位は
第１表の通りである．
（以下余白）
原料はあらかじめ水分１％以下まで乾燥させ、２■目ス
クリーンで篩分けしておいた．微粉炭は７ ４ ０ ０
Ｋｃａｌ／Ｋｇの発熱量を有する石炭を−２７５メッ
シュに微砕したものを使用し、造鎚用の赤鉱はＦｅ：
４５％，ＳＩＯ，：１８％のものを供試した．
調整された硫化鉛精鉱１００％に対して造鏝剤としての
赤鉱な２０％添加混合し、これに微粉炭をＯ〜５％の比
率で添加混合し．ｌｉ合鉱の水分は１％以下に調整した
．
また，両炉にはあらかじめ銅製錬工程から産出された水
枠鋳を各３ｔづつを装入し，Ａｆｆｉ油バーナーを用い
て両炉天井部のマンホールから加熱し、錫を溶体化した
．
脱硫炉に硫化鉛精鉱を吹込む前に計算量の粗鉛を炉内に
装入し、酸素富化空気のみで炉内吹き込みを行なってＰ
ｂ：５２％，Ｆｅ：１２％，ＳｔＯ，：７％のスタート
用鉛酸化物含有鋳を３５ｃ量厚みに形威させた．
ランスパイプは該パイプ内のガス速度が１２０〜３ ０
０ ｍｌｓｅｃを保持できるように適当な内径のステ
ンレスパイプを用い，ランス先端が静止湯面から８０ｃ
■に維持できるようにした．次に，上記調合硫化鉛原料
を酸素富化空気と共に上記諸条件で脱硫炉に装入し、３
週間連続吹込み溶錬を行なった．その運転効率は９０％
以上であった．
一方、磁元炉は上記脱硫炉から連続的に溶融鐸を受け入
れ，上記条件で還元操業を行なって，バッチ抽出を行な
った．その抽出頻度は鋳で６〜７回／日，粗鉛で３〜４
回７日であった．試験操業中の脱硫炉での溶体温度は１
０３０〜１１５０℃で推移したが、鋳のｐｂ品位が５０
％以七であれば，錫温度は１０５０℃で充分であること
が分った．
一方，還元炉では鶴の残留ｐｂ品位を２％以下に下げる
と、１１００℃以上の温度に保持しなければ錫抽出が容
易でなかった．粗鉛温度は９００〜１１００℃であった
．
脱硫炉からの鋳は還元炉へ連続的に流れ、鋳流量が２
ｔ／Ｉ｛以上の場合は、途中での冷却固化の現象は認め
られなかったが、それ以下では冷却固化現象が認められ
たので，プロパンガスバーナーを脱硫炉出口と還元炉受
口に備え付けて樋及び流出鋳の保温を行なうと共に，樋
上面をフード力バーで覆い、熱放散を防止した．
３週間にわたる実証炉試験の結果（脱硫炉系のみ）を第
２表に示す．
（以下余白）
第２表の結果から分かるように、煙灰発生率は平均で１
１％であり、他のインジェクション方式による鉛精鉱処
理プロセス（例えば前記特開昭５９−２２６１３０号で
は１５〜２０％）よりもはるかに低い．
また，排ガス中のＳＯｚも高濃度であることに加えて、
ＳＯｓ／５０２比で約０．０１とＳＯ３まで酸化される
のは極＜ｍａｔであり、微粉炭が脱酸剤としても有効に
作用していることは明らかである．
Ｓ全体での排ガス中への移行率は９４％と非常に高く，
かつ煙灰のＳ品位も約４％と低いことから，硫化鉛精鉱
中のＰｂＳがほとんど瞬間的に反応してＰｂＯに変わる
ため、ＰｂＳ形態で蒸発する機会が極めて少なかったこ
とが分かる．更に、本実証炉試験の原料比［吹込固体重
量（Ｋｇ）／吹込ガス重１ （Ｋｇ） ］は平均して２
．９であった．
次に，本試験において、吹込み速度は変えずに原料比の
みを変化させて脱硫炉における酸素効率等に及ぼす影響
について試験した．その結果を第３表に示す．
第
３
表
（以下余白）
第３表の結果から、原料比Ａ （１．８）． Ｂ（２．
１）Ｃ （２．３ ）の場合、酸素効率はＡ：９２％，
Ｂ：９５％、Ｃ：９６％であることが分かる．上記の実
施例１に示した実証炉試験で、ランス吹込速度について
みるに、脱硫炉では１２０〜ｌ６ ０ Ｎｙｇ／ ｓｅ
ａ ，還元炉で１ ８ ０ 〜２ ５ ０Ｎ＊／ｓｅｃ
と大きく異なっているが、これは吹込み時の固体重量／
気体重量比（重量比率）が小さいほど溶体への侵入深さ
が浅くなるために，還元炉では脱硫炉の場合に比べて大
きくしなければならないからである．
また、脱硫炉の湯丈（溶体深さ）は静止時で約３０ｃ鵬
厚であるが、吹込み時にはフォーミングによって約６０
ｃｍ厚であった．
一方，還元炉の場合はバッチ抽出を行なうために変動す
るが、静ＩＥ時で３０〜６０ｃｍ厚であった．
実施例２
低品位硫化鉛精鉱につき実施例１と同様に試験を行なっ
た．実証炉としては第１図に示すような装置を用いた．
脱硫炉ならびに還元炉の操業条件は下記の通りとした．
脱硫炉
原料吹込量： ２．５ｔ／｝Ｉ （鉛精鉱２．２ｔ／Ｈ
十繰返し煙灰Ｑ．３ｔ／｝Ｉ）
吹込み空気ｉｔ： ７ ７ ０ＮＭａ ／Ｈ吹込み酸素
量：３３０Ｎ）ｌ”／Ｈ（純度９１％）ランス吹込み速
度： １ ４ ７ Ｎｍｌｓｅｃランス位置：静出湯面
上８０ｃ厘（ランス先端）酸素効率：９０％以上
還元炉
微粉次吹込み量： ２ ４ ０Ｋｇ／Ｈ吹込み空気量：
１０００ＮＭ”／Ｈ
ランス吹込み速度： ２ ４ ０ ｍｌｓｅｃランス位
置：静東場面上４０ｃｍ（ランス先端）供試した低品位
硫化鉛精鉱の化学成分品位は第４表の通りである．
（以下余白）
鉛精鉱はあらかじめ水分０．５％以下まで乾燥させ、２
■目スクリーンで篩分けし、微粉炭は実施例１と同様の
ものを使用した．
また、その他の諸条件は実施例１と同様であるが，脱硫
炉では熱補償用の微粉炭を使用しなくとも２ ． ５
ｔ／｝Ｉ装入で熱バランスが保持され，自己燃焼熱のみ
で反応は充分進行した．
この試験操業において、生成した鋳の温度は約１０７０
℃であり、常時脱硫炉からオーバーフローして鋳流量が
２．０ｔ／｝Ｉと少ないものの，脱硫炉の出口部と逝元
炉の入口部を実施例ｌの場合と同様にプロパンガスバー
ナーで加温することにより，連続的に還元炉に流し込む
ことができた．還元炉の操業も上記条件で実施例ｌと同
様に連続的な還元反応操業を行なった．
その結果，脱硫炉での煙灰発生率はｌ２％であり、また
生成した溶体は鉛酸化物（ＰｂＯ）を含む鋳のみであっ
た．
一方、還元炉では粗鉛と鐸の２種類の溶体が生成するが
、原料中のＣｕ，Ａｓ，Ｓｂ等はほとんど粗船中に吸収
されていた．
還元炉での鋳温度は１１５０±１０℃であり，排ガス組
威はＣＯ：３〜７．５％、Ｇｏ２：１４．５〜２２．０
％であった．
粗鉛の品位は約８Ｂ．０％であり、取鍋でドロッシング
した結果、高品位の精製鉛とドロスとに分離することが
できた．
ドロッシングは４２０〜４６０℃で撹拌して行ったが，
得られたドロスの粒度は−３７メッシュが約８５％であ
った．
（ホ）発明の効果
以上のように、本発明法は通常の高品位硫化鉛粕鉱は勿
論のこと，低品位硫化鉛精鉱や主成分がＰｂＳＯ４やＰ
ｂＯである鉛滓等の鉛原料から低コストで効率よ〈粗鉛
を採収することができ、Ｓ分は第１工程の脱硫炉から高
濃度のＳＯｔガスとして回収して硫酸原料とすることが
できる．特に、脱硫炉において硫化鉛Ｍ鉱を溶融鋳中に
吹き込んで脱硫溶錬し，粗鉛を生成させる必要がないか
ら，煙灰の発生率を著しく低下させることができるので
ある．Detailed Description of the Invention (a) Technical field The present invention relates to continuous lead production in which crude lead is obtained from lead raw materials (hereinafter referred to as lead concentrate) such as lead sulfide concentrate or lead slag whose main components are PbSOa or PbO. Regarding the smelting method, in more detail, it is related to a lead smelting technology that does not require the presence of crude lead in the desulfurization furnace, uses a top-blown lance charging method that does not immerse the lance in the solution, and generates very little heat from smoke ash. It is something. (Example) Conventional technology The standard smelting method for lead concentrate is the sintering-blast furnace method. Roasted and sintered to lead sulfide (PbS)
to lead oxide (PbO). Next, the obtained sintered ore (
The process consists of two steps: charging PBO with carbonaceous powder into a blast furnace and reducing it to obtain crude lead. However, in this smelting method, SOl generated in the sintering process is
! There are many drawbacks, such as low gas concentration, repeated sintering, which makes handling complicated, and high energy consumption. Various lead smelting methods have been researched and proposed as countermeasures. It has been done. As one of these methods, the flash furnace method, which has been widely applied in recent years especially to copper smelting, is being tried as a method for smelting lead concentrate. However, in addition to the fact that the flash furnace method causes significant erosion of the bricks in the shaft part where the high-temperature reaction occurs, and because it is a gas phase reaction,
When lead concentrate is smelted with PbS, which has a high vapor pressure, a large amount of smoke is generated, which necessitates secondary treatment, resulting in low yields. A major problem is that it leads to higher costs. Examples include the Auto Kumpu method and the Kiv
This is typified by the cst method. In addition, a continuous smelting method for lead concentrate has been proposed that uses a single horizontal furnace, separates the reaction conditions on the left and right sides of the furnace, and performs oxidation and reduction reactions separately.This method is known as the QSL method. It is being done. However, this method involves the difficulty of performing the completely opposite reactions of oxidation and reduction in a single reactor, and requires extremely advanced reactor operation technology. The disadvantage is that the cost is significantly high. Recently, the injection smelting method, a smelting technology in which raw materials are injected into a solution and smelted, has been researched and developed, and has already been put into practical use in some copper smelting fields, and this method is being applied to the lead smelting field. Attempts to do so have already been made, including JP-A No. 62-222032 "Lead Smelting Method", JP-A No. 59-226130 "Continuous Direct Smelting Method of Lead", and JP-A No. 63-128138. No. ``Electric Furnace Lead Smelting Method,'' etc., have been proposed. First, the lead smelting method disclosed in JP-A No. 62-222032 is a method for smelting lead raw material mainly containing non-self-combustible lead sulfate (PbSO4). There is a desulfurization furnace in which pulverized coal, etc. is added and blown into the solution in the furnace through a lance to perform a decomposition desulfurization reaction, and slag containing lead oxide (PbO) produced in the desulfurization furnace is received and blown with pulverized coal, etc. It consists of a reduction furnace that performs an integrated reduction reaction. The two furnaces are usually connected by a gutter so that the slag from the desulfurization furnace is continuously charged into the reduction furnace. The basic reactions of this process are the following reactions (1) and (2) in the desulfurization furnace that decomposes lead sulfate, and the reaction (3) in the reduction furnace.
However, it is also possible to generate phase lead in a desulfurization furnace. 'pbso4+c+pbo+so,+co (1)S
O,,+CO−+ So,,+CO, (2
)PbO+station C − P b +y2COp
(3) The advantage of this process is ■) It is possible to use inexpensive coal as a heat source, resulting in lower energy costs. ■) The two reactions, desulfurization and reduction, are carried out in separate furnaces, so it is rational in terms of metallurgical reactions. ■) The equipment is simple, easy to operate, and has good productivity. and ■) the ability to recover SO2 in exhaust gas as sulfuric acid. However, JP-A No. 62-222032 deals with the treatment of lead sulfate, and there is no mention of whether lead concentrate (PbS) can be treated. It is. Next, JP-A-59-226130
The method in question consists of a smelting furnace and a reduction furnace, which is an electric furnace. Lead concentrate is injected into the molten lead bath of the smelting furnace with a lance, and the molten crude lead and slag produced are transferred to the reduction zone. By this,
This lead smelting method is a process that streamlines the reduction of slag, and it is an essential requirement that crude lead always be present in the smelting furnace. In other words, although this smelting method was developed to treat lead concentrate, with the aim of enabling a smelting reaction with the large dissolving power of crude lead for PbS, it is necessary to continuously produce crude lead in a smelting furnace. Since it is necessary to make the PbS active dragon considerably large, Pb
The generation of smoke ash in the S form is significant, and there are many problems such as a decrease in the primary yield of valuable metals and a decrease in productivity in terms of handling and other operations. Next, JP-A No. 63-128138 discloses that a mixed powder consisting of lead slag mainly composed of pbSO4, a fuel, a solvent, and a reducing agent is blown into a high-temperature solution in an electric furnace together with air at a specific blowing pressure. This is a method of smelting lead shavings using an electric furnace. However, in this method, lead slag containing pbSO4 in the t-form is blown directly into the high-temperature solution in a reducing atmosphere, which increases the amount of ash generated, which poses a serious problem. In other words, when Pb and S coexist in a reducing atmosphere, the dominant chemical species is PbS, which is in a highly volatile form, so it is assumed that most of the Pb becomes PbS and transfers to smoke ash. Moreover, the serious problem remains that the high-temperature solution is stirred by blowing, which further accelerates the volatilization of the smoke ash. (C) Disclosure of the Invention The present invention has been made with the aim of solving the above-mentioned problems.
This is a continuous smelting method for obtaining crude lead from lead raw materials such as lead slag (referred to as lead concentrate in the text), in which lead concentrate is lanced with oxygen or oxygen-enriched air in a slag solution containing lead oxide. After converting the lead content in the lead concentrate into oxide by blowing into the solution without immersing it, the lead concentrate is charged into a reduction furnace, and a reducing agent is blown into the solution to obtain crude lead. This provides a less lead-based method. The present invention will be explained in detail below. The present invention is a lead smelting technology that was developed based on the knowledge that the desulfurization reaction ends instantaneously by injecting lead concentrate together with oxygen or oxygen-enriched air into a molten metal containing lead oxide. , an innovative lead smelting method. The main reactor of the present invention consists of two furnaces in which a desulfurization furnace and a reduction furnace are connected in series, and the method for charging lead concentrate into the desulfurization furnace is an upward blowing method into a molten crane (a lance is placed in a molten trowel). (injected without immersion in the water). Therefore, since the reaction proceeds in a strongly oxidizing atmosphere in the desulfurization furnace, the generation rate of smoke ash is significant.
It is possible to keep it low. In addition, the solution produced in the desulfurization furnace is lead oxide (mainly PbO).
This hammer is continuously charged and supplied to a reduction furnace together with a reducing agent to reduce lead to crude lead. This is a feature of the present invention. In smelting using the blowing method, the charged lead concentrate penetrates into the solution and is dispersed, so heat conduction is extremely good and the high temperature required for the reaction can be reached in a short time, so the reaction is significantly accelerated. ．． Therefore, PbS is instantaneously oxidized and absorbed into the molten casting. Therefore, the time that PbS exists in the easily evaporable form is significantly shortened, so evaporation is suppressed and the rate of smoke ash generation can be kept low. In addition, since PbS is completely oxidized to PbO, there are no restrictions on the reaction as in the case of producing crude lead, and the oxygen potential (PO2) can be made arbitrarily large, so the oxidation rate can be accelerated. It is effective in reducing the morphological ratio of Furthermore, the lead concentrate is blown into the solution so that it is in full solid-liquid contact with the molten casting. Unlike the flash method, mechanical carryover is also reduced. In the desulfurization reaction in the desulfurization furnace, lead concentrate turns into a hammer containing lead oxide, except for becoming flue gas and smoke ash, which are the raw materials for sulfur. If the Pb grade of is 40% or more, 1050
Blow smelting can be continued sufficiently if the temperature can be maintained above ℃. As for the brewing components such as Fe, Si02, BaO, and CaO, it is only necessary to ensure a content ratio that allows the cast produced in the subsequent reduction furnace to fall within an appropriate composition WAH. Next, it is sufficient that the lance blowing speed is 120 ra/sec or more, but preferably 140-160 ra/sec.
c is appropriate. If the blowing speed is too high, it may be disadvantageous in terms of energy, damage to the bottom bricks, and the solution may jump and scatter from the furnace, solidifying on equipment in the exhaust gas treatment system and causing various troubles. This is not desirable because it causes problems such as causing problems. On the other hand, if it is too slow, the charge will not be able to enter the molten casting and will carry over to the outside of the furnace without any smelting reaction occurring. In addition, the blowing lance is usually just a stainless steel pipe, and the blowing oxygen or oxygen-enriched air is self-cooled from the inner wall to protect it from heat attack inside the furnace.
Naturally, if the wind speed inside the lance decreases, the cooling effect will decrease, making the lance more susceptible to thermal damage. The position of the lance for blowing is preferably 1. Blow in! The moisture content and particle size of fX materials are lead concentrate with a moisture content of 1% or less. Appropriate particle size is 2■■φ or less, moisture content is 1% or less for powdered carbonaceous material, and particle size is 275 mesh or less. next. ':jThe conditions for injecting the reducing agent in the reducing furnace in the IJ2 process are the same as in the case of the desulfurization furnace described above, but since the solid/gas ratio in the injection is small, the lance blowing speed is slightly faster.
It is better to place the tip of the lance low. In addition, the exhaust gas from the desulfurization furnace has a high SOz concentration. SOs
/502 ratio is small, so! & Suitable for sulfuric acid production in acid factories. The method of the present invention is as described above, and has the following advantages. a) There is no need to generate crude lead in the desulfurization furnace, so
It can be applied to low-grade lead concentrate, low-grade lead slag, etc. b) If the amount of heat is insufficient, powdered carbonaceous material can be added to the lead material. C) Powdered carbonaceous substances can also be used as deoxidizing agents to control the peroxidation state in desulfurization reactions. This is because when the desulfurization reaction becomes overoxidized, the SOS/302 ratio in the exhaust gas increases by L, which has an adverse effect. Figure 7JSl shows a conceptual diagram of the equipment according to the method of the present invention.
1 is a desulfurization furnace, 2 is a reduction furnace, and 3 is a desulfurization furnace casting trough that connects both furnaces 1.2. Lead concentrate, deoxidizer, and pulverized coal as powdered carbonaceous material are blown into the furnace through lance pipe 4 together with oxygen-enriched air. 5 is a lance pipe for blowing reducing agent, pulverized coal and air, or oxygen-enriched air, 6 is an SO2-containing gas exhaust hole, and 7 is a reducing furnace gas exhaust hole. In addition, the hill I mentioned
Liquid fuel or gaseous fuel may be used instead of the powdered carbonaceous material in the process, and the deoxidizing agent in the first step and the reducing agent in the second step may be powdery carbonaceous material, liquid fuel, or gaseous fuel. ．． The present invention will be explained below using examples. (2) Examples Example 1 As a demonstration furnace, a lead sulfide concentrate treatment test was conducted using the apparatus shown in Figure 1. The operating conditions of the desulfurization furnace and Moegen furnace were as follows. Desulfurization furnace (first step) Raw material injection amount: 1.8 to 3.5 t/H IgI enriched air amount = 800 to 1200 N/H Oxygen concentration = 49 to 74% Lance blowing speed: 1 2 0 ~ 1 6 0Ns/s
ee Pulverized coal: 0 to 5 wt% based on the weight of the raw material Lance position: 80 cm above the static hot water level (lance tip g) Mayumoto Furnace (2nd process) Pulverized coal: 120 to 250 Kg/H Air amount: 900
~1 1 5 0Nm3/H lance blowing speed: l 8
0 to 2 5 0 NE/see Lance position: 30 to 60 cm above the static hot water surface. Also, the chemical composition grade of the lead sulfide concentrate tested is as shown in Table 1. (Margins below) The raw materials were dried in advance to a moisture content of 1% or less and sieved using a second screen. Pulverized coal is 7 4 0 0
Coal with a calorific value of Kcal/Kg is pulverized to -275 mesh, and the red ore for hammer making is Fe:
45%, SIO: 18% were tested. To 100% of the prepared lead sulfide concentrate, 20% of red ore as an iron forming agent was added and mixed, and pulverized coal was added and mixed at a ratio of 0 to 5%. The moisture content of the li composite was adjusted to 1% or less. In addition, 3 tons each of water flask produced from the copper smelting process was charged into both furnaces in advance, and tin was dissolved by heating through the manholes in the ceilings of both furnaces using Affi oil burners. Before injecting lead sulfide concentrate into the desulfurization furnace, the calculated amount of crude lead is charged into the furnace, and the furnace is blown with only oxygen-enriched air.
A starting cast containing lead oxide containing b: 52%, Fe: 12%, and StO: 7% was shaped to a thickness of 35 cm. The lance pipe has a gas velocity of 120 to 30
Using a stainless steel pipe with an appropriate inner diameter to maintain 0 mlsec, the tip of the lance should be 80cm above the static water level.
■It is now possible to maintain the following. Next, the above prepared lead sulfide raw material was charged into a desulfurization furnace together with oxygen-enriched air under the above conditions.
Continuous blow smelting was carried out for a week. Its operating efficiency is 90%
That was it. On the other hand, the main furnace continuously received molten slag from the desulfurization furnace, performed reduction operation under the above conditions, and performed batch extraction. The extraction frequency is 6-7 times/day for cast iron and 3-4 times/day for crude lead.
It was the 7th day. The solution temperature in the desulfurization furnace during test operation was 1
The temperature ranged from 030 to 1150℃, but the PB quality of the casting was 50℃.
It was found that a tin temperature of 1050°C is sufficient if the tin temperature is 7% or more. On the other hand, in the reduction furnace, when the residual Pb content of Tsuru was lowered to below 2%, it was not easy to extract tin unless the temperature was maintained at 1100°C or higher. The crude lead temperature was 900-1100°C. The casting from the desulfurization furnace flows continuously to the reduction furnace, and the casting flow rate is 2.
At temperatures above t/I, no mid-cooling solidification phenomenon was observed, but at lower temperatures, cooling solidification phenomena were observed, so propane gas burners were installed at the desulfurization furnace outlet and reduction furnace socket. In addition to keeping the gutter and outflow casting warm, the top surface of the gutter was covered with a hood force bar to prevent heat dissipation. Table 2 shows the results of the three-week demonstration furnace test (desulfurization furnace system only). (Left below) As can be seen from the results in Table 2, the average smoke ash generation rate is 1
1%, which is much lower than other injection-based lead concentrate processing processes (for example, 15 to 20% in the above-mentioned Japanese Patent Application Laid-Open No. 59-226130). In addition to the high concentration of SOz in the exhaust gas,
The SOs/502 ratio is approximately 0.01, which means that the oxidation to SO3 is extremely < mat, and it is clear that pulverized coal also acts effectively as a deoxidizing agent. The overall migration rate of S into the exhaust gas is as high as 94%.
Furthermore, the S content of the smoke ash is low at approximately 4%, indicating that the PbS in the lead sulfide concentrate reacts almost instantaneously and changes to PbO, so there is very little opportunity for it to evaporate in the PbS form. Furthermore, the raw material ratio [injected solid weight (Kg)/injected gas weight 1 (Kg)] in this demonstration reactor test was on average 2.
．． It was 9. Next, in this test, only the raw material ratio was changed without changing the blowing speed, and the effect on oxygen efficiency in the desulfurization furnace was tested. The results are shown in Table 3. Table 3 (blank below) From the results in Table 3, raw material ratio A (1.8). B (2.
1) In the case of C (2.3), the oxygen efficiency is A: 92%,
It can be seen that B: 95% and C: 96%. In the demonstration furnace test shown in Example 1 above, the lance injection rate was 120 to 160 Nyg/se in the desulfurization furnace.
a, 180 to 250N*/sec in reduction furnace
This is very different from the solid weight at the time of injection/
This is because the smaller the gas weight ratio (weight ratio), the shallower the penetration depth into the solution, so the reduction furnace must be larger than the desulfurization furnace. In addition, the desulfurization furnace has a melt depth (solution depth) of about 30 cm when it is stationary, but when blowing, it is about 60 cm thick due to forming.
It was cm thick. On the other hand, in the case of a reduction furnace, the thickness varied from 30 to 60 cm during static IE, although it varied due to batch extraction. Example 2 A test was conducted in the same manner as in Example 1 on low-grade lead sulfide concentrate. As a demonstration reactor, we used the equipment shown in Figure 1.
The operating conditions of the desulfurization furnace and reduction furnace were as follows. Desulfurization furnace raw material injection amount: 2.5t/}I (lead concentrate 2.2t/H
Ten repetitions of smoke ash Q. 3t/}I) Blowing air it: 7 7 0NMa /H Blowing oxygen amount: 330N) l"/H (purity 91%) Lance blowing speed: 1 4 7 Nmlsec Lance position: 80cm above the still surface lance tip) Oxygen efficiency: 90% or more Reduction furnace pulverized powder secondary injection amount: 2 4 0 Kg/H air injection amount:
1000 NM”/H Lance blowing speed: 2 4 0 mlsec Lance position: 40 cm above the Seidou scene (lance tip) The chemical composition grades of the low-grade lead sulfide concentrate tested are as shown in Table 4. (Margins below) ) The lead concentrate is dried in advance to a moisture content of 0.5% or less, and
The coal was sieved using a mesh screen, and the same pulverized coal as in Example 1 was used. In addition, other conditions are the same as in Example 1, but the desulfurization furnace does not require the use of pulverized coal for heat compensation. 5
The heat balance was maintained by charging t/}I, and the reaction proceeded satisfactorily with only self-combustion heat. In this test run, the temperature of the cast produced was approximately 1070
℃, and although the desulfurization furnace constantly overflows and the casting flow rate is as small as 2.0 t/I, the outlet of the desulfurization furnace and the inlet of the die-out furnace are heated with a propane gas burner as in Example 1. By heating it, it was possible to continuously pour it into the reduction furnace. The reduction furnace was operated under the above conditions and a continuous reduction reaction operation was performed in the same manner as in Example 1. As a result, the smoke ash generation rate in the desulfurization furnace was 12%, and the only solution produced was casting containing lead oxide (PbO). On the other hand, in the reduction furnace, two types of solutions, crude lead and lead, were produced, but most of the Cu, As, Sb, etc. in the raw materials were absorbed into the crude lead. The casting temperature in the reduction furnace is 1150±10℃, and the exhaust gas composition is CO: 3 to 7.5%, Go2: 14.5 to 22.0.
%Met. The grade of crude lead is approximately 8B. As a result of drossing in a ladle, it was possible to separate high-grade refined lead and dross. The drossing was performed by stirring at 420-460°C.
The particle size of the obtained dross was approximately 85% -37 mesh. (e) Effects of the invention As described above, the method of the present invention can be applied not only to ordinary high-grade lead sulfide lees, but also to low-grade lead sulfide concentrates whose main components are PbSO4 and PbSO4.
Crude lead can be collected efficiently at low cost from lead raw materials such as lead slag, which is bO, and the S content can be recovered as highly concentrated SOt gas from the desulfurization furnace in the first step and used as a raw material for sulfuric acid. Can be done. In particular, since there is no need to inject lead sulfide M ore into molten casting in a desulfurization furnace to desulfurize and smelt it to generate crude lead, the generation rate of smoke ash can be significantly reduced.
第１図は本発明法を実施する設備の一例を示す概念図で
ある．
符号説明
１一脱硫炉 ２一還元炉 ３一脱硫炉出錫用樋４，５−
ランスパイプ ６一含ＳＯ２ガス排出孔７一還元炉ガス
排出孔
特 許 出 願 人 同和鉱業株式会社６
第１図
手続補正書（自釦
Ｆ威元年 ７月／ソ日Figure 1 is a conceptual diagram showing an example of equipment for carrying out the method of the present invention. Symbol explanation 1- Desulfurization furnace 2- Reduction furnace 3- Desulfurization furnace tin tap trough 4, 5-
Lance pipe 61 Contains SO2 gas discharge hole 71 Reduction furnace gas discharge hole Patent Applicant Dowa Mining Co., Ltd. 6 Figure 1 Procedural Amendment (July 1999/Soviet Union)
Claims (2)
共に吹込んで該硫化鉛精鉱中の鉛分を酸化物とする第１
工程と、該第１工程で得られた溶体を還元炉に装入し、
該溶体中に還元剤を吹込んで粗鉛を得る第２工程とから
なることを特徴とする鉛製錬法。(1) A first step in which lead concentrate is blown into a slag solution together with oxygen or oxygen-enriched air to convert the lead content in the lead sulfide concentrate into oxides.
step, charging the solution obtained in the first step into a reduction furnace,
A lead smelting method comprising a second step of injecting a reducing agent into the solution to obtain crude lead.
剤、粉状炭素性物質ならびに酸素又は酸素富化空気と共
に上吹ランスにより連続的に吹込み、該鉛精鉱中の鉛分
を酸化物とすると共に生成ガスを硫酸原料ガスとする第
１工程と、該第１工程で得られた溶体を連続的に還元炉
に装入し、該溶体中に還元剤を上吹ランスにより吹込ん
で粗鉛を得る第２工程とからなることを特徴とする鉛製
錬法。(2) Continuously blowing lead concentrate into a slag solution containing lead oxides together with a deoxidizer, powdered carbonaceous material, and oxygen or oxygen-enriched air using a top-blowing lance; A first step in which the lead content is made into an oxide and the generated gas is made into a sulfuric acid raw material gas, and the solution obtained in the first step is continuously charged into a reduction furnace, and a reducing agent is top-blown into the solution. A lead smelting method characterized by comprising a second step of obtaining crude lead by injecting it with a lance.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP15704989A JPH0324238A (en) | 1989-06-20 | 1989-06-20 | Lead smelting method |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP15704989A JPH0324238A (en) | 1989-06-20 | 1989-06-20 | Lead smelting method |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| JPH0324238A true JPH0324238A (en) | 1991-02-01 |
Family
ID=15641078
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| JP15704989A Pending JPH0324238A (en) | 1989-06-20 | 1989-06-20 | Lead smelting method |
Country Status (1)
| Country | Link |
|---|---|
| JP (1) | JPH0324238A (en) |
Cited By (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN102031393A (en) * | 2010-11-28 | 2011-04-27 | 郴州市金贵银业股份有限公司 | Continuous lead smelting clean production process |
| CN102230090A (en) * | 2010-06-01 | 2011-11-02 | 中国瑞林工程技术有限公司 | Lead-zinc integrated smelting furnace, and method for recovering lead and zinc |
| CN104988331A (en) * | 2015-07-30 | 2015-10-21 | 长沙有色冶金设计研究院有限公司 | Method for recovering and utilizing crude lead from low-grade lead material |
| CN105483393A (en) * | 2015-11-25 | 2016-04-13 | 中国恩菲工程技术有限公司 | Method for treating secondary lead through improved side-blowing smelting reduction furnace |
| CN105907988A (en) * | 2016-06-17 | 2016-08-31 | 中国恩菲工程技术有限公司 | Device for lead and zinc ore smelting |
| CN106734051A (en) * | 2016-11-21 | 2017-05-31 | 郴州市金贵银业股份有限公司 | The processing method of CRT flint glass |
| CN106756088A (en) * | 2016-11-22 | 2017-05-31 | 云南驰宏锌锗股份有限公司 | A kind of method that Ausmelt stoves process scrap lead cream |
| CN106756087A (en) * | 2016-11-22 | 2017-05-31 | 云南驰宏锌锗股份有限公司 | A kind of method that top side melting processes scrap lead cream |
| CN106893871A (en) * | 2016-12-28 | 2017-06-27 | 呼伦贝尔驰宏矿业有限公司 | A kind of lead concentrate handling process |
| KR102551098B1 (en) * | 2022-11-03 | 2023-07-05 | 고려아연 주식회사 | The method for removing chlorine in process of hydrometallurgy of zinc |
Citations (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JPS58130232A (en) * | 1981-11-26 | 1983-08-03 | マウント・アイザ・マインズ・リミテツド | Method of refining high strength lead |
-
1989
- 1989-06-20 JP JP15704989A patent/JPH0324238A/en active Pending
Patent Citations (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JPS58130232A (en) * | 1981-11-26 | 1983-08-03 | マウント・アイザ・マインズ・リミテツド | Method of refining high strength lead |
Cited By (12)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN102230090A (en) * | 2010-06-01 | 2011-11-02 | 中国瑞林工程技术有限公司 | Lead-zinc integrated smelting furnace, and method for recovering lead and zinc |
| CN102031393A (en) * | 2010-11-28 | 2011-04-27 | 郴州市金贵银业股份有限公司 | Continuous lead smelting clean production process |
| CN104988331A (en) * | 2015-07-30 | 2015-10-21 | 长沙有色冶金设计研究院有限公司 | Method for recovering and utilizing crude lead from low-grade lead material |
| CN105483393A (en) * | 2015-11-25 | 2016-04-13 | 中国恩菲工程技术有限公司 | Method for treating secondary lead through improved side-blowing smelting reduction furnace |
| CN105907988A (en) * | 2016-06-17 | 2016-08-31 | 中国恩菲工程技术有限公司 | Device for lead and zinc ore smelting |
| CN106734051A (en) * | 2016-11-21 | 2017-05-31 | 郴州市金贵银业股份有限公司 | The processing method of CRT flint glass |
| CN106734051B (en) * | 2016-11-21 | 2019-10-08 | 郴州市金贵银业股份有限公司 | The processing method of CRT flint glass |
| CN106756088A (en) * | 2016-11-22 | 2017-05-31 | 云南驰宏锌锗股份有限公司 | A kind of method that Ausmelt stoves process scrap lead cream |
| CN106756087A (en) * | 2016-11-22 | 2017-05-31 | 云南驰宏锌锗股份有限公司 | A kind of method that top side melting processes scrap lead cream |
| CN106893871A (en) * | 2016-12-28 | 2017-06-27 | 呼伦贝尔驰宏矿业有限公司 | A kind of lead concentrate handling process |
| KR102551098B1 (en) * | 2022-11-03 | 2023-07-05 | 고려아연 주식회사 | The method for removing chlorine in process of hydrometallurgy of zinc |
| WO2023234562A1 (en) * | 2022-11-03 | 2023-12-07 | 고려아연 주식회사 | Method for removing chlorine in zinc hydrometallurgy |
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