JPS6240407B2 - - Google Patents
Info
- Publication number
- JPS6240407B2 JPS6240407B2 JP6508984A JP6508984A JPS6240407B2 JP S6240407 B2 JPS6240407 B2 JP S6240407B2 JP 6508984 A JP6508984 A JP 6508984A JP 6508984 A JP6508984 A JP 6508984A JP S6240407 B2 JPS6240407 B2 JP S6240407B2
- Authority
- JP
- Japan
- Prior art keywords
- silver
- leaching
- residue
- slurry
- chlorine gas
- 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.)
- Expired
Links
- 238000000034 method Methods 0.000 claims description 65
- 238000002386 leaching Methods 0.000 claims description 60
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 claims description 54
- 229910052709 silver Inorganic materials 0.000 claims description 54
- 239000004332 silver Substances 0.000 claims description 53
- HKZLPVFGJNLROG-UHFFFAOYSA-M silver monochloride Chemical group [Cl-].[Ag+] HKZLPVFGJNLROG-UHFFFAOYSA-M 0.000 claims description 34
- 229910052802 copper Inorganic materials 0.000 claims description 33
- 239000010949 copper Substances 0.000 claims description 33
- 239000002244 precipitate Substances 0.000 claims description 32
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims description 31
- 239000002002 slurry Substances 0.000 claims description 30
- KZBUYRJDOAKODT-UHFFFAOYSA-N Chlorine Chemical compound ClCl KZBUYRJDOAKODT-UHFFFAOYSA-N 0.000 claims description 25
- AKHNMLFCWUSKQB-UHFFFAOYSA-L sodium thiosulfate Chemical compound [Na+].[Na+].[O-]S([O-])(=O)=S AKHNMLFCWUSKQB-UHFFFAOYSA-L 0.000 claims description 17
- 235000019345 sodium thiosulphate Nutrition 0.000 claims description 17
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 16
- 229910021607 Silver chloride Inorganic materials 0.000 claims description 14
- 229910052751 metal Inorganic materials 0.000 claims description 13
- 239000002184 metal Substances 0.000 claims description 13
- 238000011084 recovery Methods 0.000 claims description 10
- 238000005406 washing Methods 0.000 claims description 8
- 238000007664 blowing Methods 0.000 claims description 6
- 230000000737 periodic effect Effects 0.000 claims description 6
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 claims description 5
- 239000007864 aqueous solution Substances 0.000 claims description 4
- 238000011033 desalting Methods 0.000 claims 1
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 description 22
- 230000008569 process Effects 0.000 description 12
- 239000011780 sodium chloride Substances 0.000 description 11
- 239000000243 solution Substances 0.000 description 11
- DHCDFWKWKRSZHF-UHFFFAOYSA-N sulfurothioic S-acid Chemical compound OS(O)(=O)=S DHCDFWKWKRSZHF-UHFFFAOYSA-N 0.000 description 11
- 239000010931 gold Substances 0.000 description 10
- 239000000460 chlorine Substances 0.000 description 8
- 229910052737 gold Inorganic materials 0.000 description 8
- 150000002739 metals Chemical class 0.000 description 8
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 description 7
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 5
- 238000004140 cleaning Methods 0.000 description 5
- 230000009467 reduction Effects 0.000 description 5
- WQZGKKKJIJFFOK-GASJEMHNSA-N Glucose Natural products OC[C@H]1OC(O)[C@H](O)[C@@H](O)[C@@H]1O WQZGKKKJIJFFOK-GASJEMHNSA-N 0.000 description 4
- TWRXJAOTZQYOKJ-UHFFFAOYSA-L Magnesium chloride Chemical compound [Mg+2].[Cl-].[Cl-] TWRXJAOTZQYOKJ-UHFFFAOYSA-L 0.000 description 4
- 238000003723 Smelting Methods 0.000 description 4
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 4
- 229910052797 bismuth Inorganic materials 0.000 description 4
- 239000008121 dextrose Substances 0.000 description 4
- 229910001510 metal chloride Inorganic materials 0.000 description 4
- 239000000203 mixture Substances 0.000 description 4
- 239000011669 selenium Substances 0.000 description 4
- SQGYOTSLMSWVJD-UHFFFAOYSA-N silver(1+) nitrate Chemical compound [Ag+].[O-]N(=O)=O SQGYOTSLMSWVJD-UHFFFAOYSA-N 0.000 description 4
- 229910052787 antimony Inorganic materials 0.000 description 3
- 239000003638 chemical reducing agent Substances 0.000 description 3
- 150000001875 compounds Chemical class 0.000 description 3
- 230000000694 effects Effects 0.000 description 3
- 150000002500 ions Chemical class 0.000 description 3
- 229910052745 lead Inorganic materials 0.000 description 3
- HWSZZLVAJGOAAY-UHFFFAOYSA-L lead(II) chloride Chemical compound Cl[Pb]Cl HWSZZLVAJGOAAY-UHFFFAOYSA-L 0.000 description 3
- 239000007788 liquid Substances 0.000 description 3
- 229910052711 selenium Inorganic materials 0.000 description 3
- 239000011734 sodium Substances 0.000 description 3
- 238000003756 stirring Methods 0.000 description 3
- 239000000126 substance Substances 0.000 description 3
- CIWBSHSKHKDKBQ-JLAZNSOCSA-N Ascorbic acid Chemical compound OC[C@H](O)[C@H]1OC(=O)C(O)=C1O CIWBSHSKHKDKBQ-JLAZNSOCSA-N 0.000 description 2
- ZAMOUSCENKQFHK-UHFFFAOYSA-N Chlorine atom Chemical compound [Cl] ZAMOUSCENKQFHK-UHFFFAOYSA-N 0.000 description 2
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 2
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 description 2
- BUGBHKTXTAQXES-UHFFFAOYSA-N Selenium Chemical compound [Se] BUGBHKTXTAQXES-UHFFFAOYSA-N 0.000 description 2
- 229910004298 SiO 2 Inorganic materials 0.000 description 2
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 2
- 229910052785 arsenic Inorganic materials 0.000 description 2
- WQZGKKKJIJFFOK-VFUOTHLCSA-N beta-D-glucose Chemical compound OC[C@H]1O[C@@H](O)[C@H](O)[C@@H](O)[C@@H]1O WQZGKKKJIJFFOK-VFUOTHLCSA-N 0.000 description 2
- 229910052801 chlorine Inorganic materials 0.000 description 2
- 239000003792 electrolyte Substances 0.000 description 2
- 238000005363 electrowinning Methods 0.000 description 2
- 238000010828 elution Methods 0.000 description 2
- 229910001629 magnesium chloride Inorganic materials 0.000 description 2
- 229910017604 nitric acid Inorganic materials 0.000 description 2
- 239000000843 powder Substances 0.000 description 2
- 239000010970 precious metal Substances 0.000 description 2
- 238000007670 refining Methods 0.000 description 2
- 229910001961 silver nitrate Inorganic materials 0.000 description 2
- 229910052714 tellurium Inorganic materials 0.000 description 2
- 229910020756 KI—KOH Inorganic materials 0.000 description 1
- 230000009471 action Effects 0.000 description 1
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 1
- 235000010323 ascorbic acid Nutrition 0.000 description 1
- 229960005070 ascorbic acid Drugs 0.000 description 1
- 239000011668 ascorbic acid Substances 0.000 description 1
- 230000008901 benefit Effects 0.000 description 1
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 description 1
- -1 but in addition Inorganic materials 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- 238000006243 chemical reaction Methods 0.000 description 1
- 238000005660 chlorination reaction Methods 0.000 description 1
- 229910052681 coesite Inorganic materials 0.000 description 1
- 150000001879 copper Chemical class 0.000 description 1
- 229910052906 cristobalite Inorganic materials 0.000 description 1
- 230000001186 cumulative effect Effects 0.000 description 1
- 239000006185 dispersion Substances 0.000 description 1
- 238000004090 dissolution Methods 0.000 description 1
- 238000000605 extraction Methods 0.000 description 1
- 229910052742 iron Inorganic materials 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- BASFCYQUMIYNBI-UHFFFAOYSA-N platinum Chemical group [Pt] BASFCYQUMIYNBI-UHFFFAOYSA-N 0.000 description 1
- 238000012805 post-processing Methods 0.000 description 1
- 238000001556 precipitation Methods 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 238000011946 reduction process Methods 0.000 description 1
- 230000000717 retained effect Effects 0.000 description 1
- 238000000926 separation method Methods 0.000 description 1
- 239000000377 silicon dioxide Substances 0.000 description 1
- 235000012239 silicon dioxide Nutrition 0.000 description 1
- 239000010944 silver (metal) Substances 0.000 description 1
- 238000005245 sintering Methods 0.000 description 1
- 229910052708 sodium Inorganic materials 0.000 description 1
- 229910052682 stishovite Inorganic materials 0.000 description 1
- 230000019635 sulfation Effects 0.000 description 1
- 238000005670 sulfation reaction Methods 0.000 description 1
- 229910052717 sulfur Inorganic materials 0.000 description 1
- PORWMNRCUJJQNO-UHFFFAOYSA-N tellurium atom Chemical compound [Te] PORWMNRCUJJQNO-UHFFFAOYSA-N 0.000 description 1
- 229910052905 tridymite Inorganic materials 0.000 description 1
Classifications
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Description
本発明は、銅電解殿物からの銀の回収方法に関
するものであり、特には塩素ガス浸出法により高
収率で銀をAgClの形で固定した浸出残渣からチ
オ硫酸浸出法により銀を回収する方法の改善に関
する。
銅の電解精製工程において電解槽底に沈積する
銅電解殿物（アノードスライム）には、銅製錬原
料中に存在した銅より貴な金属がすべて濃縮され
て存在し、更に銅陽極中に存在し銅電解液の主成
分である希硫酸に溶解しにくい元素が濃縮する結
果として、金、銀、白金族元素、セレン、テル
ル、ビスマス、鉛、銅及び脈石類が混在してい
る。この銅電解殿物から貴金属等の有価元素類を
短時日で収率良くしかも低コストで回収すること
は、その製錬所の収益の改善に役立つのみでな
く、資源に乏しい我国においてはきわめて望まし
いことである。本発明は上記有価金属のうち特に
銀の効率的回収を計るものである。
我国における従来からの銅電解殿物の処理方法
として、銅電解殿物から銅及びセレンを大部分除
去した殿物を乾式熔錬することによつて貴金属類
を粗銀メタル中に収集し、分銀及び分金工程を実
施する方法が実施されているが、複雑な化合物の
集合体である殿物の熔錬であるため、直接採取率
にばらつきがあり、繰返物の熔錬を不可避的に必
要とするので、収率及びコスト面からはもとよ
り、回収に長時日を要するため金利面から不利で
あつた。
近年、新たな注目すべき方法として、銅電解殿
物をスラリー状とし、そこに塩素ガスを吹込むこ
とにより金その他の有価金属が溶出した浸出液と
銀をAgClの形で固定した浸出残渣とに分離する
塩素ガス浸出法が提唱されている。塩素ガス浸出
法としては、銅電解殿物を水性スラリーとして塩
素ガスを吹込む方法、銅電解殿物を塩素水溶液中
でスラリー状とし塩素ガスを吹込む方法及び周期
表第族及び族金属の塩化物（NaCl、MgCl2
等）を用いてスラリー化した銅電解殿物に塩素ガ
スを吹込む方法があり、中でも最後に挙げた方法
（Cl2／金属塩化物浸出法と表示する）は本件出願
人の提唱に係るものであり、殿物中の銀の99.5％
以上がAgClとして浸出残渣に固定でき、金その
他の有価金属も高収率で浸出液中に回収しうる点
で前２者の方法よりも優れている。
いずれにせよ、これら塩素ガス浸出法は、金、
銀等の早期回収という点から見て非常に簡単且つ
効率の良いプロセスであり、従来からの乾式法に
代替しうるものである。塩素ガス浸出法の浸出残
渣特にCl2／NaCl浸出法による浸出残渣には銀が
高濃度で濃縮されており、重要な銀回収源であ
る。しかしながら、この浸出残渣中には、AgCl
に加えて、通常、Pb、Sb、Bi等の化合物及び
SiO2が決して少くはない量共存している為に、
この浸出残渣から銀を高純度の形で収率よく回収
するのはそう容易ではない。これまで幾つかの方
法が提唱されているが、有力な方法としてチオ硫
酸浸出法がある。
チオ硫酸浸出法は、
(1) 銅電解殿物を塩素ガス浸出して発生する塩化
銀残渣をPH＝10においてリパルブし、このPH値
を維持しつつチオ硫酸ソーダ（Na2S2O3）で銀
を浸出する段階及び
(2) この浸出液にデキストローズや鉄粉等の還元
剤を加えて還元銀を生成させる段階
を要旨とするものである。還元銀は、その後、例
えば硝酸に再溶解し、浄液後、電解採取によつて
高純度銀が回収される。
しかしながら、現在実施されているチオ硫酸浸
出法ではチオ硫酸ソーダを還元段階後廃棄してい
る。これは、チオ硫酸ソーダによる浸出作用は現
状ではあまり良くなく、いわんやそれを繰返して
使用すると還元銀の収率が著しく低下するためで
ある。ところがチオ硫酸ソーダの価格は非常に高
く、このような使い捨ては工程コストの負担増を
招いている。従つて、殿物に対するチオ硫酸ソー
ダの浸出作用を改善して銀回収率を向上すると共
に、チオ硫酸浸出後液を塩化銀残渣の浸出工程に
繰返し使用することによつてコスト低減化を計る
必要がある。
本発明者は、銅電解殿物から生成する塩化銀残
渣に対するチオ硫酸ソーダの浸出作用を妨げる原
因を究明するべく検討を重ねた。その結果、銅電
解殿物中にはPbが含まれ、それが塩化銀残渣中
にも塩化鉛の形で多量に残存している事実が根本
的原因となつていることが判明した。塩化銀残渣
中に存在する鉛がチオ硫酸ソーダの浸出作用を妨
害して結局還元銀の収率を低下せしめ、またその
中に介入する鉛品位も高くなつて後の硝酸銀電解
液の浄液を困難としている。チオ硫酸ソーダを繰
返し使用すると塩化鉛の累積量が高くなりすぎ、
銀収率は著しく低下してしまう。従つて、塩化銀
残渣中に多量に含まれる鉛をチオ硫酸ソーダ浸出
段階前に出来るだけ低い水準に除去しておくこと
によつて、上記問題の抜本的な解決が可能とな
る。
脱鉛法としては、NaCl法、HNO3洗浄法等が考
慮しうるが、脱鉛と同時に銀も一部溶出する方法
ではその回収工程が必要となる。検討の結果、温
水或いは熱水による洗浄法が好ましいことが判明
した。
斯くして、本発明は、銅電解殿物或いはそれか
ら脱銅及び脱砒した脱銅殿物を塩素ガス浸出し、
銀を塩化銀の形で濃縮した塩化銀残渣を生成し、
チオ硫酸ソーダにより銀を浸出しそして生成浸出
液を還元して還元銀を生成する銀回収法におい
て、前記チオ硫酸ソーダ浸出前に塩化銀残渣中に
含まれる鉛を除去する脱鉛操作を行うことを特徴
とする銅電解殿物からの銀回収法を提供する。脱
鉛操作は、加熱水（温水或いは熱水）により塩化
銀残渣を洗浄することにより実施することが好ま
しい。
以下、本発明について詳述する。
本発明の対象は銅の電解精製工程において副生
する銅電解殿物であるが、これはまだかなりの銅
を含んでいるので脱銅処理を施すことにより脱
銅、併せて脱砒をも行つた脱銅殿物を用いること
が好ましい。脱銅処理としては様々の方法が確立
されており、硫酸浸出、硫酸化熔焼、Fe3+イオ
ン添加等の方法いずれをも使用しうる。脱銅殿物
は、その出所源及び処理方法に応じてAu、Ag、
Cu、As、Se、Te、Pb、Bi、Fe、Sb、Ｓ、SiO2
等を様々の範囲で含んでいる。これらのうち有価
金属を回収するシステムの一プロセスとして本発
明は銀を回収することを目的とする。
本発明に従えば、銅電解殿物或いは脱銅電物、
好ましくは脱銅殿物は、塩素ガス浸出工程におい
てスラリー状態で塩素ガス浸出される。殿物をス
ラリー化する媒体としてはこれまで水、塩酸溶液
及び周期表第族乃至族の金属の塩化物水溶液
が提唱されていることは前述したが、水や塩酸溶
液を使用した場合、銀の固定化率が悪いため、本
発明においてはNaClやMgCl2に代表される周期
表第族乃至族の金属の塩化物水溶液を使用し
て殿物のスラリー化を計るのが好都合である。例
えば、Cl2／HCL浸出法では塩化銀のかなりの量
が再溶解してAgCl残渣としての銀の回収率が最
大限でも98.2％どまりとなるのに対し、Cl2／
NaCl浸出法では残渣中に99.5％以上の銀をAgCl
として固定することができる。
上記金属塩化物を使用しての塩素ガス浸出法に
おいて、金属塩化物としてはNaClやMgCl2が代
表的に使用されるが、この他KCl、CaCl2、
BaCl2、BeCl2も好適に使用しうる。金属塩化物
濃度は一般に１〜5N、好ましくは2.5〜3.5Nとさ
れる。開放或いは密閉型の容器において、上記ス
ラリーが60〜80℃の温度の下で塩素ガスを吹込ま
れる。スラリーは容器に設置された撹拌羽根によ
つて例えば200〜1000rpmの撹拌速度で撹拌され
ることが好ましい。塩素ガス吹込量は所定の金溶
出をもたらすに適当量とされるが、200〜1500
c.c.／分／スラリーの割合で５〜７時間の吹込み
で99.5％以上の銀の残渣への固定化と99％以上の
金その他の有価金属の溶出が可能である。好まし
い吹込方法として前半の方を後半より1.5〜３倍
多量に吹込むのが有益であることが判つた。例え
ば、最初の２〜４時間を400〜600c.c.／分／スラ
リーとし、残る１〜４時間をその半分量とするの
がよい。スラリー濃度は200〜400ｇ／とされ
る。スラリー濃度が低すぎると、液PHが下り、銅
や鉛が溶出しやすくなる。
こうして所定期間塩素ガスを吹込まれた殿物ス
ラリーは、金が99％以上溶出した浸出液と銀を99
％以上AgClとして保持した残渣とに変換され、
固液分離後、それぞれに含まれる有価値元素回収
の為爾後処理に供される。塩素ガス浸出法は、工
程の早期において、殿物から銀をAgClの形で高
純度の浸出残渣として入手しうる点で優れた方法
である。浸出残渣中の金含量の低いことも特筆す
べき利点である。
こうして得られた、gCl残渣は、一般に40〜43
％の銀含量及び10％前後のPb含量を有し、他に
Sb、Bi、SiO2等を含んでいる。そして、このPb
こそが、PbCl2の形で存在して、爾後のチオ硫酸
ソーダ浸出段階を妨害しているのである。
そこで、本発明に従えば、チオ硫酸ソーダ浸出
段階の前に、脱鉛操作が行われる。脱鉛操作は、
NaCl法、HNO3洗浄法、酒石酸−KI−KOH洗浄
法、加熱水洗浄法等の適宜の方法で行いうるが、
後に参考例で示すように加熱水洗浄法が、その簡
便さ、安価さ及び銀の同時溶出を伴わないことの
点で一番有利な方法である。HNO3洗浄法も実施
しうるが、同時溶出した銀の回収工程が必要とさ
れよう。
こうして、塩化銀残渣中の鉛を大部分除去した
後の塩化銀残渣は、従来実施態様に従つてチオ硫
酸ソーダ浸出及び還元段階に併せられる。チオ硫
酸ソーダ浸出段階においては、AgCl＋2S2O3 2-→
〔Ag（S2O3）2〕3-＋Cl-の反応に従つて銀が錯イオ
ンの形で抽出される。還元段階は、Fe粉、Zn
粒、Mg粉等の金属還元剤やデキストローズ又は
アスコルビン酸等の有機還元剤を用いて実施され
る。
こうして還元銀が析出される。得られた還元銀
は、これを硝酸に再溶解し、浄液後、硝酸銀溶液
からの電解採取によつて高純度銀に変換される
し、また還元銀を鋳込んで原銀板アノード板とし
た後で電解精製によつて高純度銀に変換されう
る。最終的に、99.999％以上の高純度銀が回収し
うる。
実施例
(A) 塩素ガス浸出工程
銅製錬所において副生される銅電解殿物を
Fe3+イオンで脱銅処理して第１の化学組成の
脱銅殿物を得た。
The present invention relates to a method for recovering silver from copper electrolytic deposits, and in particular, a method for recovering silver from a leaching residue in which silver is fixed in the form of AgCl in a high yield by a chlorine gas leaching method, by a thiosulfuric acid leaching method. Concerning method improvements. In the copper electrolytic precipitate (anode slime) deposited at the bottom of the electrolytic tank during the copper electrolytic refining process, all metals nobler than copper that were present in the copper smelting raw materials are concentrated and present, and furthermore, metals that are more noble than copper that are present in the copper anode are present. As a result of the concentration of elements that are difficult to dissolve in dilute sulfuric acid, which is the main component of the copper electrolyte, gold, silver, platinum group elements, selenium, tellurium, bismuth, lead, copper, and gangue are present. Recovering valuable elements such as precious metals from this copper electrolytic precipitate in a short period of time, with high yield, and at low cost not only helps improve the profitability of the smelter, but is also extremely desirable in Japan, which is poor in resources. It is. The present invention aims to efficiently recover silver among the valuable metals mentioned above. The conventional treatment method for copper electrolytic precipitates in Japan is to collect precious metals into coarse silver metal by dry smelting the precipitates from which most of the copper and selenium have been removed, and then separate them. A method of carrying out a silver and metal dispersion process has been implemented, but since it involves smelting of precipitates, which are a collection of complex compounds, there are variations in the direct extraction rate, and repeated smelting is unavoidable. This was disadvantageous not only in terms of yield and cost, but also in terms of interest rates because it took a long time to collect. In recent years, a new and noteworthy method has been developed in which copper electrolytic precipitates are made into a slurry, and chlorine gas is blown into the slurry to create a leachate in which gold and other valuable metals are eluted, and a leachate residue in which silver is fixed in the form of AgCl. A separate chlorine gas leaching method has been proposed. Chlorine gas leaching methods include a method in which the copper electrolytic precipitate is made into an aqueous slurry and chlorine gas is blown into it, a method in which the copper electrolytic precipitate is made into a slurry in a chlorine aqueous solution and chlorine gas is blown into it, and a method in which chlorine gas is blown into the copper electrolytic precipitate in the form of a slurry in a chlorine aqueous solution, and chlorination of group metals in the periodic table. substances (NaCl, MgCl2
There is a method of blowing chlorine gas into slurry-formed copper electrolytic precipitate using Cl 2 /metal chloride leaching method. and 99.5% of the silver in the treasure
The above method is superior to the former two methods in that it can be fixed in the leaching residue as AgCl, and gold and other valuable metals can also be recovered in the leachate at a high yield. In any case, these chlorine gas leaching methods
This is a very simple and efficient process in terms of early recovery of silver, etc., and can replace the conventional dry method. The leaching residue from the chlorine gas leaching method, especially the leaching residue from the Cl 2 /NaCl leaching method, is highly concentrated in silver and is an important source of silver recovery. However, this leaching residue contains AgCl
In addition, compounds such as Pb, Sb, Bi and
Because SiO 2 coexists in a not small amount,
It is not so easy to recover silver in high purity form from this leaching residue in good yield. Several methods have been proposed so far, but one of the most promising is the thiosulfuric acid leaching method. The thiosulfuric acid leaching method is as follows: (1) The silver chloride residue generated by leaching the copper electrolytic precipitate with chlorine gas is repulsed at pH=10, and while maintaining this pH value, it is leached with sodium thiosulfate (Na 2 S 2 O 3 ). (2) adding a reducing agent such as dextrose or iron powder to this leaching solution to produce reduced silver. The reduced silver is then redissolved in, for example, nitric acid, and after purification, high-purity silver is recovered by electrowinning. However, in the currently practiced thiosulfate leaching process, the sodium thiosulfate is discarded after the reduction step. This is because the leaching effect of sodium thiosulfate is currently not very good, and if it is used repeatedly, the yield of reduced silver will drop significantly. However, the price of sodium thiosulfate is extremely high, and such disposable use increases the burden of process costs. Therefore, it is necessary to improve the leaching effect of sodium thiosulfate on precipitates to increase the silver recovery rate, and to reduce costs by repeatedly using the solution after thiosulfate leaching in the leaching process of silver chloride residue. There is. The inventors of the present invention have conducted repeated studies in order to find out the cause of hindering the leaching effect of sodium thiosulfate on silver chloride residues generated from copper electrolytic precipitates. As a result, it was found that the fundamental cause was the fact that the copper electrolytic precipitate contained Pb, which remained in large amounts in the form of lead chloride in the silver chloride residue. The lead present in the silver chloride residue interferes with the leaching action of sodium thiosulfate, ultimately reducing the yield of reduced silver, and the lead content involved also increases, making it difficult to clean the silver nitrate electrolyte afterward. It is considered difficult. Repeated use of sodium thiosulfate will cause the cumulative amount of lead chloride to become too high;
The silver yield will drop significantly. Therefore, the above problem can be fundamentally solved by removing the large amount of lead contained in the silver chloride residue to the lowest possible level before the sodium thiosulfate leaching step. As a deleading method, NaCl method, HNO 3 washing method, etc. can be considered, but methods that partially elute silver at the same time as deleading require a recovery step. As a result of the study, it was found that a cleaning method using hot water or hot water is preferable. Thus, the present invention provides chlorine gas leaching of copper electrolytic precipitate or decopper-removed precipitate which has been decoppered and arsenized therefrom,
produces a silver chloride residue, which is concentrated silver in the form of silver chloride;
In a silver recovery method in which silver is leached with sodium thiosulfate and the resulting leachate is reduced to produce reduced silver, a deleading operation to remove lead contained in the silver chloride residue is performed before the sodium thiosulfate leaching. A method for recovering silver from copper electrolytic deposits is provided. The deleading operation is preferably carried out by washing silver chloride residue with heated water (hot water or hot water). The present invention will be explained in detail below. The object of the present invention is copper electrolytic precipitate, which is produced as a by-product in the copper electrolytic refining process, but since this still contains a considerable amount of copper, copper removal treatment is performed to remove copper and also remove arsenic. It is preferable to use a decoppered precipitate. Various methods have been established for copper removal treatment, and any of the methods such as sulfuric acid leaching, sulfation sintering, and Fe 3+ ion addition can be used. Decoppered precipitates can contain Au, Ag, or
Cu, As, Se, Te, Pb, Bi, Fe, Sb, S, SiO 2
etc. to various extents. Among these, the present invention aims to recover silver as one process of a system for recovering valuable metals. According to the present invention, a copper electrolytic precipitate or a decoppered electrical material,
Preferably, the decoppered precipitate is leached with chlorine gas in a slurry state in the chlorine gas leaching step. As mentioned above, water, a hydrochloric acid solution, and an aqueous chloride solution of a metal from group 1 or group of the periodic table have been proposed as media for slurrying precipitates. Since the immobilization rate is poor, it is convenient in the present invention to slurry the precipitate using an aqueous chloride solution of a metal of group 1 or group of the periodic table, such as NaCl or MgCl 2 . For example, in the Cl 2 /HCL leaching method, a significant amount of silver chloride is redissolved, resulting in a maximum silver recovery rate of only 98.2% as AgCl residue, whereas Cl 2 /HCL leaching
In the NaCl leaching method, more than 99.5% of silver is added to the residue using AgCl.
It can be fixed as In the above chlorine gas leaching method using metal chlorides, NaCl and MgCl 2 are typically used as metal chlorides, but in addition, KCl, CaCl 2 ,
BaCl 2 and BeCl 2 can also be suitably used. The metal chloride concentration is generally 1-5N, preferably 2.5-3.5N. In an open or closed container, the slurry is blown with chlorine gas at a temperature of 60-80°C. The slurry is preferably stirred by a stirring blade installed in the container at a stirring speed of, for example, 200 to 1000 rpm. The amount of chlorine gas injected is said to be an appropriate amount to bring about the specified gold elution, but it is 200 to 1500.
By blowing at a rate of cc/min/slurry for 5 to 7 hours, it is possible to immobilize more than 99.5% of silver in the residue and to elute more than 99% of gold and other valuable metals. It has been found that a preferred method of blowing is to blow 1.5 to 3 times more into the first half than the second half. For example, it is preferable to use 400 to 600 c.c./min/slurry for the first 2 to 4 hours, and half that amount for the remaining 1 to 4 hours. The slurry concentration is 200 to 400g/. If the slurry concentration is too low, the liquid pH will drop, making it easier for copper and lead to elute. The precipitate slurry, which has been blown with chlorine gas for a predetermined period of time, is combined with a leachate in which more than 99% of gold has been eluted and 99% of silver.
% or more of the residue retained as AgCl and converted into
After solid-liquid separation, they are subjected to post-processing to recover valuable elements contained in each. The chlorine gas leaching method is an excellent method in that silver can be obtained from the precipitate in the form of AgCl as a highly pure leaching residue at an early stage of the process. The low gold content in the leaching residue is also a notable advantage. The gCl residue thus obtained is generally 40 to 43
% silver content and around 10% Pb content, and other
Contains Sb, Bi, SiO2, etc. And this Pb
It is this, present in the form of PbCl2 , that interferes with the subsequent sodium thiosulfate leaching step. According to the invention, therefore, a deleading operation is carried out before the sodium thiosulfate leaching step. The deleading operation is
It can be carried out by any appropriate method such as NaCl method, HNO 3 washing method, tartaric acid-KI-KOH washing method, heated water washing method, etc.
As shown later in Reference Examples, the heated water washing method is the most advantageous method in terms of its simplicity, low cost, and absence of simultaneous elution of silver. A HNO 3 wash method could also be performed, but a co-eluted silver recovery step would be required. Thus, after removing most of the lead in the silver chloride residue, the silver chloride residue is combined with a sodium thiosulfate leaching and reduction step in accordance with conventional embodiments. In the sodium thiosulfate leaching step, AgCl+2S 2 O 3 2- →
According to the [Ag(S 2 O 3 ) 2 ] 3- + Cl - reaction, silver is extracted in the form of complex ions. In the reduction stage, Fe powder, Zn
It is carried out using a metal reducing agent such as grains or Mg powder, or an organic reducing agent such as dextrose or ascorbic acid. In this way, reduced silver is deposited. The obtained reduced silver is redissolved in nitric acid, purified, and converted into high-purity silver by electrowinning from the silver nitrate solution.The reduced silver is also cast to form raw silver plates and anode plates. After that, it can be converted to high purity silver by electrorefining. Ultimately, more than 99.999% high purity silver can be recovered. Example (A) Chlorine gas leaching process
A decoppered precipitate having a first chemical composition was obtained by decoppering treatment with Fe 3+ ions.
【表】
(i) この脱銅殿物をスラリー元液として１〜
5N NaClを用いて375ｇ／のスラリー濃度
にスラリー化し、ここに塩素ガスを吹込むこ
とにより塩素ガス浸出を行つた。浸出温度
は、60℃としそして浸出時間は６時間と固定
した。塩素ガス吹込量は最初の３時間に500
c.c.／分／スラリーとし、残りの時間をその
半分量とした。処理後の浸出液の化学組成を
表２に示す。浸出液中のAg濃度は非常に低
く、それだけAgがAgClとして浸出残渣中に
固定されていることを示す。ちなみに、Au
の浸出率は3N NaClの場合99％以上もの高い
値を示している。NaCl濃度は、スラリー濃
度、浸出条件等に応じて最適となるように選
択されるべきである。[Table] (i) This decopper-removed precipitate is used as slurry source liquid for 1 to
A slurry was prepared using 5N NaCl to a slurry concentration of 375 g/ml, and chlorine gas was leached by blowing chlorine gas into the slurry. The leaching temperature was fixed at 60°C and the leaching time was fixed at 6 hours. The amount of chlorine gas injected is 500 in the first 3 hours.
cc/min/slurry, and half the amount for the remaining time. The chemical composition of the leachate after treatment is shown in Table 2. The Ag concentration in the leachate is very low, indicating that Ag is fixed in the leach residue as AgCl. By the way, Au
The leaching rate of 3N NaCl is as high as 99% or more. The NaCl concentration should be selected to be optimal depending on the slurry concentration, leaching conditions, etc.
【表】
(ii) NaCl以外の塩化物として周期表第族か
らMgを代表的に選び、MgCl2水溶液スラリ
ーによる殿物浸出試験を行つた。ここでは、
3N MgCl2溶液を用い、前記脱銅殿物を250
ｇ／の濃度にスラリー化した。浸出温度を
80℃に上げ、Cl2ガスを６時間連続して吹込
んだ。吹込量は前半０〜３時間は１／分／
スラリーそして後半３〜６時間は0.5／
分／スラリーとした。得られた浸出率を表
３に示す。[Table] (ii) Mg was selected as a representative chloride from Group 3 of the periodic table as a chloride other than NaCl, and a precipitate leaching test using a MgCl 2 aqueous solution slurry was conducted. here,
Using 3N MgCl2 solution, remove the copper-decoppered precipitate at 250%
The slurry was made into a slurry at a concentration of g/g/g. leaching temperature
The temperature was raised to 80° C., and Cl 2 gas was continuously bubbled for 6 hours. The blowing amount is 1/min/for the first 0 to 3 hours.
Slurry and the second half 3 to 6 hours 0.5/
minute/slurry. The obtained leaching rates are shown in Table 3.
【表】【table】
【表】
スラリー濃度が250ｇ／と低いためAgCl
の再溶解度が多少高まつたようである。スラ
リー濃度を適正に選択することによりAgCl
回収率を増大しうる。
いずれにせよ、Cl2／金属塩化物系での殿
物浸出において周期律表の族（Na、Ｋ、
Rb等）、第族（Be、Mg等）の中から適当
な元素を選び好成績を収め得ることが実証さ
れた。
(B) 脱鉛工程
(A)工程と同様にして得られた次の組成の塩化
銀残渣を使用した。
成分 組成（重量％）
Ag 25.7
Pb 5.2
Au 0.16
Se 1.46
Te 0.34
As 0.46
Sb 0.9
Bi 0.58
Ｓ 0.37
この塩化銀残渣を60℃の温水において50ｇ／
スラリー濃度で30分の浸出時間における脱鉛
操作を２回繰返した。78％の鉛が除去された。
(C) チオ硫酸浸出工程
こうして脱鉛された塩化銀残渣を次の条件で
チオ硫酸浸出処理した：
パルプ濃度 100ｇ／
チオ硫酸ソーダ添加量
２モルNa2S2O3／１モルAg
浸出温度 室温
浸出時間 1hr
(D) 還元工程
得られた浸出液をデキストローズを用いて還
元処理し、還元銀を得た。処理条件は次の通り
とした：
デキストローズ添加量 0.84ｇ／ｇAg
還元温度 60℃
還元時間 10〜30分
(E) チオ硫酸浸出液の繰返し使用
還元銀析出後のチオ硫酸浸出液を(C)及び(D)と
同条件で繰返し使用した。
この結果得られた銀収率を脱鉛処理を行わな
かつた場合と併せて示す。[Table] Because the slurry concentration is as low as 250g/AgCl
It appears that the re-solubility of the compound has increased somewhat. By properly selecting the slurry concentration, AgCl
Recovery rates can be increased. In any case, the groups of the periodic table (Na, K,
It has been demonstrated that it is possible to achieve good results by selecting an appropriate element from the groups (Rb, etc.) and group groups (Be, Mg, etc.). (B) Deleading step Silver chloride residue obtained in the same manner as in step (A) and having the following composition was used. Component composition (wt%) Ag 25.7 Pb 5.2 Au 0.16 Se 1.46 Te 0.34 As 0.46 Sb 0.9 Bi 0.58 S 0.37 This silver chloride residue was mixed with 50 g/g in hot water at 60°C.
The deleading operation at slurry concentration and leaching time of 30 minutes was repeated twice. 78% of lead was removed. (C) Thiosulfuric acid leaching process The silver chloride residue thus deleaded was subjected to thiosulfuric acid leaching treatment under the following conditions: Pulp concentration 100g/Additional amount of sodium thiosulfate
2 mol Na 2 S 2 O 3 / 1 mol Ag Leaching temperature Room temperature leaching time 1 hr (D) Reduction process The obtained leaching solution was reduced using dextrose to obtain reduced silver. The treatment conditions were as follows: Added amount of dextrose 0.84g/gAg Reduction temperature 60℃ Reduction time 10-30 minutes (E) Repeated use of thiosulfuric acid leachate The thiosulfuric acid leachate after reduced silver precipitation was used as (C) and ( It was used repeatedly under the same conditions as D). The silver yield obtained as a result is shown together with the case where no deleading treatment was performed.
【表】
この結果から、脱鉛処理を予め行うことによ
つて初回の銀収率が上るだけでなく、繰返し使
用しても充分の銀収率が確保されることがわか
る。80℃熱水で脱鉛を行うと一層良好な結果が
生じる。
参考例
脱鉛操作は適宜の方法で行いうるが、ここでは
加熱水により好適に為しうることの例を示す。次
の３種の洗浄液を用いて塩化銀残渣をスラリー状
態として洗浄（即ち鉛の浸出）を行つた：[Table] From this result, it can be seen that by performing deleading treatment in advance, not only the initial silver yield increases, but also a sufficient silver yield is ensured even after repeated use. Even better results are obtained when deleading is carried out with hot water at 80°C. Reference Example Although the deleading operation can be carried out by any suitable method, an example will be shown here in which it can be suitably carried out using heated water. The silver chloride residue was slurried and cleaned (i.e., lead leached) using the following three types of cleaning solutions:
【表】
〓
〓第2回 〃 91％
このように、上記条件で温水或いは熱水により
２回の洗浄を行うことによつて脱Pb率を充分に
高めることができ、しかも銀の同時溶解を伴わな
いので、また費用も一番安くつくので、加熱水洗
浄が一番効果的である。もちろん、スラリー濃
度、撹拌条件、浸出時間を変えることにより１回
の水洗でもつて所望水準への脱Pbを実施しう
る。その他の方法も、洗浄方式及び条件を変える
ことによりもつと高い水準への脱Pbを画ること
ができ、本発明においてその使用を排除すること
を意図するものでない。
以上説明した通り、本発明は、銅電解殿物から
発生する塩化銀残渣をチオ硫酸浸出するに際し、
残渣の脱鉛を行うことによつて、銀の回収率を著
しく向上し、爾後の浄液の負担を軽減し、同時に
高価なチオ硫酸浸出後液を使い捨てずに繰返し使
用することにより工程コストを著しく低減するこ
とに成功したものであり、その意義はきわめて大
きい。[Table] 〓
〓2nd 〃 91%
In this way, by performing washing twice with warm water or hot water under the above conditions, the Pb removal rate can be sufficiently increased, and since it does not involve simultaneous dissolution of silver, it is also the cheapest method. Therefore, heated water cleaning is the most effective. Of course, by changing the slurry concentration, stirring conditions, and leaching time, it is possible to remove Pb to a desired level with just one water wash. Other methods can achieve a higher level of Pb removal by changing the cleaning method and conditions, and the present invention does not intend to exclude their use. As explained above, in the present invention, when leaching silver chloride residue generated from copper electrolytic precipitate with thiosulfuric acid,
By deleading the residue, the recovery rate of silver is significantly improved and the burden of cleaning solution is reduced. At the same time, the process cost is reduced by repeatedly using the expensive thiosulfuric acid leaching solution instead of throwing it away. We have succeeded in significantly reducing this, and its significance is extremely significant.
Claims (1)
脱銅殿物を周期表第族乃至族の金属の塩化物
の水溶液によりスラリー化し、生成スラリーに塩
素ガスを吹込むことにより塩素ガス浸出して、銀
を塩化銀の形で高度に濃縮した塩化銀残渣を生成
し、該塩化銀残渣中に含まれる鉛を除去する脱鉛
操作を行つた後、チオ硫酸ソーダにより銀を浸出
し、そして後生成浸出液を還元して還元銀を生成
することから成る銀回収法。 ２ 脱塩操作が加熱水により塩化銀残渣を洗浄す
ることにより実施される特許請求の範囲第１項記
載の方法。[Claims] 1. Slurrying a copper electrolytic precipitate or a decoppered precipitate obtained by decoppering and dearsenizing it with an aqueous solution of a chloride of a metal from Group 1 or Group 3 of the periodic table, and blowing chlorine gas into the slurry produced. After leaching chlorine gas to produce a silver chloride residue in which silver is highly concentrated in the form of silver chloride, and performing a deleading operation to remove lead contained in the silver chloride residue, the silver is leached with sodium thiosulfate. A method of silver recovery consisting of leaching and reducing the post-produced leachate to produce reduced silver. 2. The method according to claim 1, wherein the desalting operation is carried out by washing silver chloride residue with heated water.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP59065089A JPS60208434A (en) | 1984-04-03 | 1984-04-03 | Method for recovering silver from precipitate of copper electrolysis |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP59065089A JPS60208434A (en) | 1984-04-03 | 1984-04-03 | Method for recovering silver from precipitate of copper electrolysis |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| JPS60208434A JPS60208434A (en) | 1985-10-21 |
| JPS6240407B2 true JPS6240407B2 (en) | 1987-08-28 |
Family
ID=13276853
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| JP59065089A Granted JPS60208434A (en) | 1984-04-03 | 1984-04-03 | Method for recovering silver from precipitate of copper electrolysis |
Country Status (1)
| Country | Link |
|---|---|
| JP (1) | JPS60208434A (en) |
Families Citing this family (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US6627149B1 (en) | 1996-06-21 | 2003-09-30 | Dowa Mining Co., Ltd. | High-purity silver wires for use in recording, acoustic or image transmission applications |
| JP3725621B2 (en) | 1996-06-21 | 2005-12-14 | 同和鉱業株式会社 | High-purity silver wire for recording or sound or image transmission |
| AUPQ456299A0 (en) * | 1999-12-09 | 2000-01-13 | Geo2 Limited | Recovery of precious metals |
| US6660059B2 (en) | 2000-05-19 | 2003-12-09 | Placer Dome Technical Services Limited | Method for thiosulfate leaching of precious metal-containing materials |
| HUE060079T2 (en) | 2010-12-07 | 2023-01-28 | Barrick Gold Corp | Co-current and counter current resin-in-leach in gold leaching processes |
| AR086933A1 (en) | 2011-06-15 | 2014-01-29 | Barrick Gold Corp | METHOD FOR RECOVERING PRECIOUS METALS AND COPPER OF LIXIVIATE SOLUTIONS |
| US10161016B2 (en) | 2013-05-29 | 2018-12-25 | Barrick Gold Corporation | Method for pre-treatment of gold-bearing oxide ores |
| JP6368672B2 (en) * | 2014-05-12 | 2018-08-01 | Ｊｘ金属株式会社 | Silver smelting method |
| JP6444784B2 (en) * | 2015-03-19 | 2018-12-26 | Ｊｘ金属株式会社 | Method for treating solution containing silver, thiosulfuric acid and impurities, method for recovering thiosulfate, and method for leaching silver |
| PE20211512A1 (en) | 2019-01-21 | 2021-08-11 | Barrick Gold Corp | METHOD FOR CARBON-CATALYZED THOSULFATE LEACHING OF MATERIALS CONTAINING GOLD |
Citations (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JPS5785942A (en) * | 1980-11-18 | 1982-05-28 | Sumitomo Metal Mining Co Ltd | Recovering method for gold from slime after copper electrolysis |
| JPS57177941A (en) * | 1981-04-22 | 1982-11-01 | Sumitomo Metal Mining Co Ltd | Collecting method of silver from electrolytic slime |
-
1984
- 1984-04-03 JP JP59065089A patent/JPS60208434A/en active Granted
Patent Citations (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JPS5785942A (en) * | 1980-11-18 | 1982-05-28 | Sumitomo Metal Mining Co Ltd | Recovering method for gold from slime after copper electrolysis |
| JPS57177941A (en) * | 1981-04-22 | 1982-11-01 | Sumitomo Metal Mining Co Ltd | Collecting method of silver from electrolytic slime |
Also Published As
| Publication number | Publication date |
|---|---|
| JPS60208434A (en) | 1985-10-21 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| US4293332A (en) | Hydrometallurgical process for recovering precious metals from anode slime | |
| JP3616314B2 (en) | Method for treating copper electrolytic deposits | |
| JP3879126B2 (en) | Precious metal smelting method | |
| CN106119554B (en) | High Purity Gold and the method for being enriched with silver, platinum and palladium are prepared from silver anode slime | |
| JP5237120B2 (en) | Gold collection method | |
| US4229270A (en) | Process for the recovery of metal values from anode slimes | |
| US4002544A (en) | Hydrometallurgical process for the recovery of valuable components from the anode slime produced in the electrolytical refining of copper | |
| JP2001316736A (en) | Method for recovering silver | |
| JP6810887B2 (en) | Separation and recovery methods for selenium, tellurium, and platinum group elements | |
| JPS5952218B2 (en) | Method for recovering gold from copper electrolytic slime | |
| JPS6240407B2 (en) | ||
| CN106048233A (en) | Leaching method of silver | |
| PL205994B1 (en) | Method for processing anode sludge | |
| JP7206142B2 (en) | Method for separating and recovering valuable metals | |
| JPS6139383B2 (en) | ||
| US4662938A (en) | Recovery of silver and gold | |
| US4666514A (en) | Hydrometallurgical process for recovering silver from copper-electrolysis anode sludge | |
| JP4016680B2 (en) | Method for dissolving selenium platinum group element-containing material | |
| EP0266337B1 (en) | Hydrometallurgical process for recovering silver from electrolytic copper refinery slimes and from similar raw materials | |
| JP4715598B2 (en) | Chloride leaching method of lead electrolysis slime | |
| JP2791161B2 (en) | Method for recovering silver from silver-containing nitric acid solution | |
| US5939042A (en) | Tellurium extraction from copper electrorefining slimes | |
| JPH0781172B2 (en) | Silver refining ore mud purification method | |
| JPS6158530B2 (en) | ||
| JPS6221851B2 (en) |