JPS6158530B2 - - Google Patents
Info
- Publication number
- JPS6158530B2 JPS6158530B2 JP12559483A JP12559483A JPS6158530B2 JP S6158530 B2 JPS6158530 B2 JP S6158530B2 JP 12559483 A JP12559483 A JP 12559483A JP 12559483 A JP12559483 A JP 12559483A JP S6158530 B2 JPS6158530 B2 JP S6158530B2
- Authority
- JP
- Japan
- Prior art keywords
- slurry
- leaching
- copper
- precipitate
- silver
- 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
- 229910052802 copper Inorganic materials 0.000 claims description 35
- 239000010949 copper Substances 0.000 claims description 35
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 claims description 33
- 238000002386 leaching Methods 0.000 claims description 32
- 238000000034 method Methods 0.000 claims description 31
- 239000002244 precipitate Substances 0.000 claims description 29
- KZBUYRJDOAKODT-UHFFFAOYSA-N Chlorine Chemical compound ClCl KZBUYRJDOAKODT-UHFFFAOYSA-N 0.000 claims description 16
- 229910052751 metal Inorganic materials 0.000 claims description 7
- 239000002184 metal Substances 0.000 claims description 7
- 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 4
- 239000002002 slurry Substances 0.000 description 33
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 description 19
- 239000010931 gold Substances 0.000 description 19
- 229910052737 gold Inorganic materials 0.000 description 16
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 description 15
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 description 15
- 229910052709 silver Inorganic materials 0.000 description 15
- 239000004332 silver Substances 0.000 description 14
- 238000011084 recovery Methods 0.000 description 11
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 10
- 239000011780 sodium chloride Substances 0.000 description 9
- BASFCYQUMIYNBI-UHFFFAOYSA-N platinum Chemical group [Pt] BASFCYQUMIYNBI-UHFFFAOYSA-N 0.000 description 8
- 239000000243 solution Substances 0.000 description 8
- KDLHZDBZIXYQEI-UHFFFAOYSA-N Palladium Chemical compound [Pd] KDLHZDBZIXYQEI-UHFFFAOYSA-N 0.000 description 7
- 239000000460 chlorine Substances 0.000 description 7
- 239000010970 precious metal Substances 0.000 description 7
- 229910052711 selenium Inorganic materials 0.000 description 7
- 239000011669 selenium Substances 0.000 description 7
- 230000000052 comparative effect Effects 0.000 description 6
- 239000010944 silver (metal) Substances 0.000 description 6
- BUGBHKTXTAQXES-UHFFFAOYSA-N Selenium Chemical compound [Se] BUGBHKTXTAQXES-UHFFFAOYSA-N 0.000 description 5
- 229910052801 chlorine Inorganic materials 0.000 description 5
- HKZLPVFGJNLROG-UHFFFAOYSA-M silver monochloride Chemical compound [Cl-].[Ag+] HKZLPVFGJNLROG-UHFFFAOYSA-M 0.000 description 5
- 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
- 239000006227 byproduct Substances 0.000 description 4
- 238000010828 elution Methods 0.000 description 4
- 229910001510 metal chloride Inorganic materials 0.000 description 4
- 229910052763 palladium Inorganic materials 0.000 description 4
- 229910052697 platinum Inorganic materials 0.000 description 3
- 238000007670 refining Methods 0.000 description 3
- 229910052714 tellurium Inorganic materials 0.000 description 3
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 description 2
- WCUXLLCKKVVCTQ-UHFFFAOYSA-M Potassium chloride Chemical compound [Cl-].[K+] WCUXLLCKKVVCTQ-UHFFFAOYSA-M 0.000 description 2
- 229910021607 Silver chloride Inorganic materials 0.000 description 2
- 239000007864 aqueous solution Substances 0.000 description 2
- 229910052785 arsenic Inorganic materials 0.000 description 2
- 239000007789 gas Substances 0.000 description 2
- 239000012535 impurity Substances 0.000 description 2
- 238000002347 injection Methods 0.000 description 2
- 239000007924 injection Substances 0.000 description 2
- 150000002500 ions Chemical class 0.000 description 2
- 150000002739 metals Chemical class 0.000 description 2
- FGDZQCVHDSGLHJ-UHFFFAOYSA-M rubidium chloride Chemical compound [Cl-].[Rb+] FGDZQCVHDSGLHJ-UHFFFAOYSA-M 0.000 description 2
- 238000003756 stirring Methods 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- UMGDCJDMYOKAJW-UHFFFAOYSA-N thiourea Chemical compound NC(N)=S UMGDCJDMYOKAJW-UHFFFAOYSA-N 0.000 description 2
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Chemical compound O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 2
- UXVMQQNJUSDDNG-UHFFFAOYSA-L Calcium chloride Chemical compound [Cl-].[Cl-].[Ca+2] UXVMQQNJUSDDNG-UHFFFAOYSA-L 0.000 description 1
- 229910004298 SiO 2 Inorganic materials 0.000 description 1
- XSQUKJJJFZCRTK-UHFFFAOYSA-N Urea Natural products NC(N)=O XSQUKJJJFZCRTK-UHFFFAOYSA-N 0.000 description 1
- RQNWIZPPADIBDY-UHFFFAOYSA-N arsenic atom Chemical compound [As] RQNWIZPPADIBDY-UHFFFAOYSA-N 0.000 description 1
- WDIHJSXYQDMJHN-UHFFFAOYSA-L barium chloride Chemical compound [Cl-].[Cl-].[Ba+2] WDIHJSXYQDMJHN-UHFFFAOYSA-L 0.000 description 1
- 229910001626 barium chloride Inorganic materials 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 229910001627 beryllium chloride Inorganic materials 0.000 description 1
- LWBPNIJBHRISSS-UHFFFAOYSA-L beryllium dichloride Chemical compound Cl[Be]Cl LWBPNIJBHRISSS-UHFFFAOYSA-L 0.000 description 1
- 229910052797 bismuth Inorganic materials 0.000 description 1
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 description 1
- 239000001110 calcium chloride Substances 0.000 description 1
- 229910001628 calcium chloride Inorganic materials 0.000 description 1
- 235000011148 calcium chloride Nutrition 0.000 description 1
- 239000003638 chemical reducing agent Substances 0.000 description 1
- 239000000306 component Substances 0.000 description 1
- 150000001875 compounds Chemical class 0.000 description 1
- 150000001879 copper Chemical class 0.000 description 1
- 230000007423 decrease Effects 0.000 description 1
- 239000006185 dispersion Substances 0.000 description 1
- 239000012153 distilled water Substances 0.000 description 1
- 239000003792 electrolyte Substances 0.000 description 1
- 238000000605 extraction Methods 0.000 description 1
- 229910052745 lead Inorganic materials 0.000 description 1
- 239000007788 liquid Substances 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 230000001590 oxidative effect Effects 0.000 description 1
- -1 platinum group metals Chemical class 0.000 description 1
- 238000012805 post-processing Methods 0.000 description 1
- 238000003672 processing method Methods 0.000 description 1
- 239000000047 product Substances 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- 230000000717 retained effect Effects 0.000 description 1
- 238000000926 separation method Methods 0.000 description 1
- 229910052708 sodium Inorganic materials 0.000 description 1
- 239000011734 sodium Substances 0.000 description 1
- ZNCPFRVNHGOPAG-UHFFFAOYSA-L sodium oxalate Chemical compound [Na+].[Na+].[O-]C(=O)C([O-])=O ZNCPFRVNHGOPAG-UHFFFAOYSA-L 0.000 description 1
- 229940039790 sodium oxalate Drugs 0.000 description 1
- 239000007787 solid Substances 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
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、銅の電解精製工程において銅陽極か
ら副生するアノードスライム即ち銅電解殿物から
金、銀、白金族金属等に代表される有価元素類を
効率的に回収する為の浸出方法に関するものであ
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
しにくい物質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。
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
施する方法が行なわれているが、複雑な化合物の
集合体である殿物の溶錬であるため、直接採取率
にばらつきがあり、繰返物の溶錬を不可避的に必
要とするので、収率及びコンテスト面からはもと
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。
加圧浸出等を用いた湿式法が試みられたことも
あつたが、所期の成果をあげえず、結局乾式法に
とつて代わることはなかつた。
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、そこに塩素を吹込むことに
より貴金属を浸出する方法が提唱されている。例
えば、銅電解スライムを水性スラリーとして塩素
ガスを吹込む方法や、銅電解殿物を塩素水溶液中
でスラリー状とし、塩素ガスを吹込んで浸出を行
う方法がある。これら塩素浸出法は金、銀等の早
期回収という点から見て非常に簡単かつ効率の良
いプロセスであると考えられ、従来からの乾式法
に代替しうるものとして有望視される。
しかしながら、本件出願人が上記方法を銅製錬
所の銅電解工程において副生する実際の銅電解殿
物に適用してみたところ、いずれも満足な貴金属
回収率を示さなかつた。例えば、塩酸／塩素浸出
法では、殿物のスラリー化に際して用いる濃厚な
塩酸溶液に対して塩化銀（AgCl）がかなりの量
再溶解することからAgCl残渣としての銀の回収
率が最大限でも98.2％どまりとなり、また金浸出
率も95.7％どまりとなつて、それ以上の改善を為
しえなかつた。更に、白金やパラジウムの回収率
も低かつた。
従つて、実操業においてもつと高い貴金属回収
率を保証しうるよう、上記塩素浸出法について
種々の検討を加えた結果、銅電解殿物のスラリー
化において周期表第族或いは族の金属塩化物
の水溶液を用いることにより、金、銀ともに99％
以上で浸出回収されることが判明し、また白金や
パラジウムの回収率も向上することも判明した。
本発明は、この知見に基く銅電解殿物浸出方法
を提供するものであり、周期表第族又は族金
属の塩化物水溶液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ことを本旨とする。
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いわゆる脱銅殿物を用
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、硫酸化焙焼、
Fe3+イオン添加等の方法いずれをも使用しう
る。脱銅後、酸化焙焼による脱セレン処理を実施
するのが通常であるが、本発明の場合には脱セレ
ン処理の実施は任意である。セレンは本発明以降
の工程で採取しうる。
脱銅殿物は、Au、Ag、Cu、As、Se、Te、
Pb、Fe、Sb、Ｓ、SiO2等を出所源及び処理方法
に応じて様々な範囲で含んでいる。本発明に従え
ば、銅電解殿物、好ましくは脱銅殿物は周期表第
族或いは第族の金属の塩化物水溶液を使用し
てスラリー化される。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、
RbCl、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、200〜1500c.c.／分／スラリーの割合で５〜
７時間の吹込みで99％以上の金溶出が可能であ
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、銀や鉛
の溶解度が増し、従つて浸出液中に銀や鉛が溶出
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、
液固分離後、それぞれに含まれる有価元素回収の
為爾後処理に供される。溶出液からの金、その他
の貴金属の回収法及び浸出残渣からの銀の回収法
については、既に多くの方法が確立されている。
例えば、浸出液から過酸化水素水、蓚酸ソーダ等
の還元剤で処理して金を回収し、脱金液からPt、
Pd、Te、Se等が回収される。溶出残渣からAgは
チオ尿素浸出或いは乾式法にて回収しうる。
実施例 １及び比較例
銅製錬所において副生される銅電解殿物Fe3+
イオンで脱銅処理して表１の化学組成の脱銅殿物
を得た。
The present invention relates to a leaching method for efficiently recovering valuable elements such as gold, silver, and platinum group metals from anode slime, that is, copper electrolytic precipitate, which is produced as a by-product from a copper anode in a copper electrolytic refining process. It is something. In the copper electrolytic precipitate deposited at the bottom of the electrolytic tank during the copper electrolytic refining process, all the metals nobler than copper that were present in the copper smelting raw materials are concentrated and present, and the copper electrolyte that is present in the copper anode is also present. As a result of concentration of substances that are difficult to dissolve in dilute sulfuric acid, which is the main component, gold, silver, platinum group elements, selenium, tellurium, bismuth, lead, copper, and gangue are mixed. 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 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 is currently in use, but since it involves smelting 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 competition, but also in terms of interest rates because it took a long time to collect. A wet method using pressure leaching was attempted, but it did not produce the desired results and ultimately did not replace the dry method. In recent years, a new and noteworthy method has been proposed in which copper electrolytic precipitates are made into a slurry and chlorine is blown into the slurry to leach out precious metals. For example, there is a method in which copper electrolytic slime is made into an aqueous slurry and chlorine gas is blown into it, and a method in which copper electrolytic precipitate is made into a slurry in a chlorine aqueous solution and chlorine gas is blown into the slurry to perform leaching. These chlorine leaching methods are considered to be very simple and efficient processes in terms of early recovery of gold, silver, etc., and are considered promising as a substitute for the conventional dry method. However, when the present applicant applied the above method to actual copper electrolytic precipitates produced as a by-product in the copper electrolytic process at a copper smelter, none of them showed a satisfactory precious metal recovery rate. For example, in the hydrochloric acid/chlorine leaching method, a considerable amount of silver chloride (AgCl) is redissolved in the concentrated hydrochloric acid solution used to slurry the precipitate, so the recovery rate of silver as AgCl residue is at most 98.2%. %, and the gold leaching rate also remained at 95.7%, and no further improvement could be made. Furthermore, the recovery rate of platinum and palladium was also low. Therefore, in order to ensure a high recovery rate of precious metals in actual operations, we have conducted various studies on the above chlorine leaching method, and have found that metal chlorides from group 1 or group of the periodic table are used in slurrying copper electrolytic precipitates. By using an aqueous solution, both gold and silver are 99%
It has been found that the above method can be leached and recovered, and that the recovery rate of platinum and palladium is also improved. The present invention provides a method for leaching copper electrolytic precipitates based on this knowledge, and includes blowing chlorine gas into copper electrolytic precipitates that have been made into a slurry using an aqueous chloride solution of a group metal of the periodic table. The main purpose. 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 so-called decoppered precipitate. Various methods have been established for decopper treatment, including sulfuric acid leaching, sulfated roasting,
Any method such as adding Fe 3+ ions can be used. After copper removal, selenium removal treatment is usually performed by oxidative roasting, but in the case of the present invention, implementation of selenium removal treatment is optional. Selenium can be collected in steps subsequent to the present invention. De-coppered precipitates include Au, Ag, Cu, As, Se, Te,
It contains Pb, Fe, Sb, S, SiO 2 etc. in varying ranges depending on the source and processing method. According to the present invention, a copper electrolytic precipitate, preferably a decoppered precipitate, is slurried using an aqueous chloride solution of a Group or Group metal of the periodic table. NaCl as a metal chloride
and MgCl 2 are typically used, but in addition, KCl,
RbCl, CaCl2 , BaCl2 , BeCl2 may 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%. 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 blown is considered to be an appropriate amount to bring about the specified gold elution, but it is recommended to
More than 99% of gold can be eluted after 7 hours of blowing. The preferred method of blowing is to blow the first half more than the second half.
It has been found to be beneficial to inject 1.5 to 3 times more. For example, the first 2 to 4 hours 400 to 600 c.c./
minutes/slurry, and half the amount for the remaining 1 to 4 hours. Slurry concentration is 200~400g/
It is said that If the slurry concentration is too low, the solubility of silver and lead will increase, making it easier for silver and lead to dissolve into the leachate. The precipitate slurry that has been injected with chlorine gas for a predetermined period of time is leached with over 99% of the gold and 99% of the silver.
% or more of the residue retained as AgCl and converted into
After liquid-solid separation, each liquid is subjected to post-processing to recover the valuable elements contained therein. Many methods have already been established for the recovery of gold and other precious metals from the eluate and for the recovery of silver from the leaching residue.
For example, gold can be recovered from the leachate by treating it with a reducing agent such as hydrogen peroxide or sodium oxalate, and Pt can be recovered from the degold solution.
Pd, Te, Se, etc. are recovered. Ag can be recovered from the elution residue by thiourea leaching or a dry method. Example 1 and Comparative Example Copper electrolytic precipitate Fe 3+ produced as a by-product in a copper smelter
A decoppered precipitate having a chemical composition shown in Table 1 was obtained by decoppering with ions.
【表】
この脱銅殿物をスラリー元液として(i)蒸留水、
(ii)１〜5N HCl及び(iii)１〜5N NaClの３種類を用
いて375ｇ/のスラリー濃度にスラリー化した。
ここに塩素ガスを吹込むことにより、３者の浸出
挙動を比較した。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500cc／分／スラリーとし、残りの
時間をその半分量とした。実施例、比較例それぞ
れの浸出率を表２に示す。表２においては、HCl
及びNaClの濃度として3Nのものも代表的に示し
た。[Table] This copper-removed precipitate was used as a slurry source solution (i) distilled water,
Three types of slurry were used: (ii) 1-5N HCl and (iii) 1-5N NaCl to a slurry concentration of 375 g/.
By blowing chlorine gas into this, the leaching behavior of the three materials was compared. The leaching temperature was 60°C and the leaching time was fixed at 6 hours. The amount of chlorine gas blown was 500 cc/min/slurry for the first 3 hours, and half that amount for the remaining time. Table 2 shows the leaching rates of Examples and Comparative Examples. In Table 2, HCl
A representative NaCl concentration of 3N is also shown.
【表】【table】
【表】
実施例と比較例との比較からわかるように、金
の浸出率は水スラリーの場合は94.19％と低く、
塩酸溶液スラリーを使用しても95.72％までしか
向上しないのに対し、本発明によれば金浸出率は
99％以上となる。銀についても、比較例では銀が
溶出残渣から1.76〜1.77％溶出液中に溶解するの
に対し、本発明では溶解量はその1/3程度に抑え
られる。パラジウムについても比較例の89％台か
ら99％以上へのまた白金についても比較例の73％
から96.54％への著しい向上が見られる。
表３は1N、3N及び5Nの濃度のNaClの場合の浸
出率を示す。3Nの場合が1N及び5Nより良好な結
果を示す。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選択されるべきであ
る。[Table] As can be seen from the comparison between Examples and Comparative Examples, the leaching rate of gold was as low as 94.19% in the case of water slurry;
Although the use of hydrochloric acid solution slurry only improves the gold leaching rate to 95.72%, according to the present invention, the gold leaching rate is
More than 99%. Regarding silver, in the comparative example, 1.76 to 1.77% of silver was dissolved in the eluate from the elution residue, whereas in the present invention, the amount dissolved was suppressed to about 1/3 of that amount. Palladium also increased from 89% in the comparative example to over 99%, and platinum also increased to 73% in the comparative example.
A remarkable improvement can be seen from 96.54% to 96.54%. Table 3 shows the leaching rates for NaCl concentrations of 1N, 3N and 5N. The case of 3N shows better results than 1N and 5N. The NaCl concentration should be selected to be optimal depending on the slurry concentration, leaching conditions, etc.
【表】【table】
【表】
このように、NaCl型スラリーの浸出実績は
HCl型スラリーのそれよりかなり向上することが
わかる。更に、有害不純物元素としてのSbの浸
出率はHClスラリーの場合の74％からNaClスラ
リーでは27％へと著しく低下する。これは、不溶
性のSbCl3・NaCl塩が生成して溶出残渣中に多く
残留する為であるが、有害不純物が早い時期にし
かもかなりの量系外に除去しうることは、後の有
価金属回収工程がその分だけ楽になるというメリ
ツトを与える。
実施例 ２
スラリー濃度を250ｇ/と低くそしてCl2ガス
吹込量を前半３時間を１／分／スラリーそし
て後半３時間を0.5／分／スラリーとした以
外は、実施例１と同条件で試験を行つた。反応時
間の経過に伴うAgのAgCl残渣としての回収率及
びAuの浸出率を測定した。結果を表４に示す。[Table] As shown above, the leaching results of NaCl type slurry are
It can be seen that this is considerably improved over that of the HCl type slurry. Moreover, the leaching rate of Sb as a harmful impurity element decreases significantly from 74% in the case of HCl slurry to 27% in NaCl slurry. This is because insoluble SbCl 3 and NaCl salts are generated and remain in large amounts in the elution residue, but the fact that harmful impurities can be removed from the system in a considerable amount at an early stage means that valuable metals can be recovered later. This provides the advantage of making the process that much easier. Example 2 The test was carried out under the same conditions as in Example 1, except that the slurry concentration was as low as 250g/, and the Cl 2 gas injection rate was 1/min/slurry for the first 3 hours and 0.5/min/slurry for the latter 3 hours. I went. The recovery rate of Ag as AgCl residue and the leaching rate of Au were measured as the reaction time progressed. The results are shown in Table 4.
【表】
これより、スラリー濃度が250ｇ/でもAgと
Auの浸出率は本質的に変らないことがわかる。
実施例 ３
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、3NMgCl2溶液
を用い、前記脱銅殿物を250ｇ/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５に示す。[Table] From this, even if the slurry concentration is 250g/Ag
It can be seen that the leaching rate of Au remains essentially unchanged. Example 3 Mg from group 3 of the periodic table as a chloride other than NaCl
A precipitate leaching test using MgCl 2 aqueous slurry was conducted on selected representative samples. Here, the decoppered precipitate was slurried using a 3NMgCl 2 solution to a concentration of 250 g/ml. Increase the leaching temperature to 80℃ and add Cl2 gas to 60℃.
It was injected continuously for hours. The blowing rate is 1/min/slurry for the first 0 to 3 hours and the second half for 3 to 6 hours.
0.5/min/slurry. The obtained leaching rates are shown in Table 5.
【表】
Auの浸出率を見ると3N NaCl使用の場合とほ
とんど同じ挙動を示す。一方、Agの浸出残渣へ
の移行率（スラリーAgの回収率）は６時間後で
98.27％と多少おちるようである。スラリー濃度
が250ｇ/と低いためにAgClの溶液への溶解度
が多少高まつたためと考えられる。
いずれにせよ金属塩化物／Cl2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。
以上説明した通り、本発明は先に提唱されてい
る銅殿物塩素ガス吹込浸出法に較べて金、銀その
他の貴金属の回収率を高めることに成功したもの
であり、本方法により高回収率で得られた金、銀
が製品となるのに約１週間と現行法に較べ非常に
短縮される。[Table] Looking at the Au leaching rate, it shows almost the same behavior as when using 3N NaCl. On the other hand, the transfer rate of Ag to the leaching residue (slurry Ag recovery rate) was
It seems to have fallen somewhat to 98.27%. This is thought to be because the solubility of AgCl in the solution increased somewhat due to the low slurry concentration of 250 g/ml. In any case, good results can be achieved by selecting an appropriate element from the groups (Na, K, Rb, etc.) and groups (Be, Mg, etc.) of the periodic table in precipitate leaching in a metal chloride/Cl 2 system. This has been proven. As explained above, the present invention has succeeded in increasing the recovery rate of gold, silver, and other precious metals compared to the previously proposed copper precipitate chlorine gas injection leaching method. It takes about one week for the gold and silver obtained in this process to become finished products, which is much shorter than with the current method.
Claims (1)
を用いてスラリ−化した銅電解殿物あるいは脱銅
殿物に塩素ガスを吹込むことを特徴とする銅電解
殿物の浸出方法。1. A method for leaching a copper electrolytic precipitate, which comprises blowing chlorine gas into a copper electrolytic precipitate or a decoppered precipitate that has been slurried using an aqueous chloride solution of a metal of group 1 or group of the periodic table.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP58125594A JPS6021340A (en) | 1983-07-12 | 1983-07-12 | Method for leaching precipitate after copper electrolysis |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP58125594A JPS6021340A (en) | 1983-07-12 | 1983-07-12 | Method for leaching precipitate after copper electrolysis |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| JPS6021340A JPS6021340A (en) | 1985-02-02 |
| JPS6158530B2 true JPS6158530B2 (en) | 1986-12-12 |
Family
ID=14914012
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| JP58125594A Granted JPS6021340A (en) | 1983-07-12 | 1983-07-12 | Method for leaching precipitate after copper electrolysis |
Country Status (1)
| Country | Link |
|---|---|
| JP (1) | JPS6021340A (en) |
Families Citing this family (5)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US4979986A (en) * | 1988-02-22 | 1990-12-25 | Newmont Gold Company And Outomec U.S.A., Inc. | Rapid oxidation process of carbonaceous and pyritic gold-bearing ores by chlorination |
| JP4216307B2 (en) | 2006-09-27 | 2009-01-28 | 日鉱金属株式会社 | Process for electrolytically precipitated copper |
| JP2009082815A (en) * | 2007-09-28 | 2009-04-23 | Central Res Inst Of Electric Power Ind | Washing method and apparatus for coal gasification slag |
| CN113215412B (en) * | 2021-05-10 | 2022-07-29 | 上海第二工业大学 | Method for selectively leaching and recovering silver on surface of waste silver-plated part or silver-containing solid waste |
| CN113308606B (en) * | 2021-06-04 | 2022-10-18 | 昆明理工大学 | Method for leaching and separating valuable metals from silver-gold-rich selenium steaming slag |
-
1983
- 1983-07-12 JP JP58125594A patent/JPS6021340A/en active Granted
Also Published As
| Publication number | Publication date |
|---|---|
| JPS6021340A (en) | 1985-02-02 |
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