JP2006130387A - Method of recovering platinum and rhenium from waste catalyst - Google Patents

Method of recovering platinum and rhenium from waste catalyst Download PDF

Info

Publication number
JP2006130387A
JP2006130387A JP2004320180A JP2004320180A JP2006130387A JP 2006130387 A JP2006130387 A JP 2006130387A JP 2004320180 A JP2004320180 A JP 2004320180A JP 2004320180 A JP2004320180 A JP 2004320180A JP 2006130387 A JP2006130387 A JP 2006130387A
Authority
JP
Japan
Prior art keywords
rhenium
platinum
leaching
solution
lead
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.)
Granted
Application number
JP2004320180A
Other languages
Japanese (ja)
Other versions
JP4347783B2 (en
Inventor
Yasukatsu Sasaki
康勝 佐々木
Koji Soe
浩二 副
Atsushi Saito
淳 斉藤
Current Assignee (The listed assignees may be inaccurate. Google has not performed a legal analysis and makes no representation or warranty as to the accuracy of the list.)
Nippon Mining Holdings Inc
Eneos Corp
Original Assignee
Nippon Mining and Metals Co Ltd
Nippon Mining Co Ltd
Priority date (The priority date 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 date listed.)
Filing date
Publication date
Application filed by Nippon Mining and Metals Co Ltd, Nippon Mining Co Ltd filed Critical Nippon Mining and Metals Co Ltd
Priority to JP2004320180A priority Critical patent/JP4347783B2/en
Priority to CNB2005100785703A priority patent/CN100348749C/en
Publication of JP2006130387A publication Critical patent/JP2006130387A/en
Application granted granted Critical
Publication of JP4347783B2 publication Critical patent/JP4347783B2/en
Expired - Fee Related legal-status Critical Current
Anticipated expiration legal-status Critical

Links

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
    • 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
    • Y02P20/00—Technologies relating to chemical industry
    • Y02P20/50—Improvements relating to the production of bulk chemicals
    • Y02P20/584—Recycling of catalysts

Landscapes

  • Manufacture And Refinement Of Metals (AREA)
  • Processing Of Solid Wastes (AREA)
  • Inorganic Compounds Of Heavy Metals (AREA)
  • Catalysts (AREA)

Abstract

<P>PROBLEM TO BE SOLVED: To provide a method of recovering platinum and rhenium from a waste catalyst carried on an alumina-containing porous support with a high recovery. <P>SOLUTION: The method of recovering platinum and rhenium from a waste catalyst carried on an alumina-containing porous support comprises causing platinum and rhenium to leach out with an alkali solution, reducing and filtering the platinum in the leached solution, adsorbing the rhenium in the leached solution with an anion exchange resin, eluting rhenium from the resin with hydrochloric acid and sulfurizing the rhenium in the eluted solution to recover as rhenium sulfide. <P>COPYRIGHT: (C)2006,JPO&NCIPI

Description

本発明は、アルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒からの白金及びレニウムを回収する方法に関する。   The present invention relates to a method of recovering platinum and rhenium from a waste catalyst containing platinum and rhenium supported on a porous support containing alumina.

従来、アルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒からのレニウム及び白金の回収は、硫酸溶液にてレニウムを浸出後、陰イオン交換樹脂、又は溶媒抽出にて浸出液からレニウムを抽出し、白金は浸出残渣として回収する方法（特許文献１：USP3672874）が知られている。
この方法ではアルミナを硫酸溶液にて浸出した液中のアルミン酸イオン処理時に非常に濾過性の悪い残渣が生成し、濾過の作業性が非常に煩雑になる。 Conventionally, platinum supported on a porous carrier containing alumina, rhenium and platinum recovery from a waste catalyst containing rhenium are leached with sulfuric acid solution, and then extracted from the leachate by anion exchange resin or solvent extraction. A method of extracting rhenium and recovering platinum as a leaching residue (Patent Document 1: USP3672874) is known.
In this method, a residue with very poor filterability is produced during the treatment with aluminate ions in a solution obtained by leaching alumina with a sulfuric acid solution, and the workability of filtration becomes very complicated.

また、アルミナ担体上に担持している白金、レニウムの回収する方法として、白金、レニウムを担持させているアルミナを粉砕し、王水などの酸にて浸出後、未浸出のアルミナ残渣を濾過し、濾液中の白金は卑金属（鉄、亜鉛、アルミニウムなど）を用いてセメンテーション反応により沈殿させ、セメンテーション後液中のレニウムは硫化処理を行って硫化物として回収する方法などが知られている。 In addition, as a method for recovering platinum and rhenium supported on an alumina carrier, the alumina supporting platinum and rhenium is pulverized and leached with an acid such as aqua regia, and then the unleached alumina residue is filtered. In addition, platinum in the filtrate is precipitated by a cementation reaction using base metals (iron, zinc, aluminum, etc.), and rhenium in the solution after cementation is sulfidized and recovered as a sulfide. .

しかし王水などの酸による白金やレニウムの浸出法は浸出率が低く、かつレニウムの場合、その後の硫化処理時のレニウム回収率が低く、それを解消するためには、繰返し浸出などの処理を行うことが必要となり、浸出する酸が多量になり、白金濃度が希薄なものとなるため、還元して高い回収率とするには労力と時間を多く費やし、経済性に欠ける欠点がある。 However, the leaching method of platinum and rhenium with acids such as aqua regia has a low leaching rate, and in the case of rhenium, the rhenium recovery rate during the subsequent sulfidation treatment is low. To eliminate this, treatment such as repeated leaching is required. This requires a large amount of acid to be leached and a low platinum concentration. Therefore, it takes a lot of labor and time to reduce to a high recovery rate, and there is a disadvantage that it is not economical.

さらに、直接乾式処理法を用いた場合、レニウムは高温にて非常に酸化揮発し易いため、その回収率は低く、かつ酸化揮発したレニウムは排ガス煙道配管などに付着し、吸湿性のある酸化レニウムは排ガス煙道配管を腐食するなどの課題があった。
USP3672874 Furthermore, when using the direct dry treatment method, rhenium is very oxidative and volatile at high temperatures, so its recovery rate is low, and the oxidized and volatilized rhenium adheres to exhaust gas flue pipes, etc. Rhenium has problems such as corroding exhaust gas flue piping.
USP3672874

本発明の目的は、アルミナを含む多孔質担体上に担持された白金、レニウム
を含む廃触媒から高回収率にて白金及びレニウムを回収する方法を提供する。 An object of the present invention is to provide a method for recovering platinum and rhenium at a high recovery rate from a waste catalyst containing platinum and rhenium supported on a porous support containing alumina.

本発明は、上記の課題を解決するものであって、
（１）アルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒からの白金、及びレニウムを回収する方法において、
前記廃触媒中の白金，レニウムをアルカリ溶液により浸出し、浸出液中の白金を還元し、濾過後、浸出液中のレニウムを陰イオン交換樹脂により吸着し、前記樹脂よりレニウムを塩酸溶液により溶離し、溶離後液中のレニウムは硫化処理を行って硫化レニウムとして回収する廃触媒からの白金及びレニウムの回収方法。 The present invention solves the above problems,
(1) In a method of recovering platinum supported on a porous carrier containing alumina, platinum from a waste catalyst containing rhenium, and rhenium,
Platinum and rhenium in the waste catalyst are leached with an alkaline solution, platinum in the leachate is reduced, filtered, rhenium in the leachate is adsorbed with an anion exchange resin, and rhenium is eluted from the resin with a hydrochloric acid solution, A method for recovering platinum and rhenium from a spent catalyst in which rhenium in the solution after elution is sulfurized and recovered as rhenium sulfide.

（２）上記（１）記載のアルカリ浸出後の浸出触媒及び浸出残渣は乾燥後、電気炉において酸化鉛と溶剤を添加し、1350℃~1500℃に保持し、鉛に白金を吸収させた金属鉛（以下、貴鉛と記す）を得、該貴鉛を酸化炉にて1000℃~1200℃において酸化し、白金を10~40mass%含む貴金属鉛を得る廃触媒からの白金及びレニウムの回収方法。
（３）上記（２）記載の貴金属鉛を粉砕後、硝酸溶液にて鉛を浸出、濾過し、その硝酸浸出の濾過残渣を塩酸と過酸化水素により塩化浸出、濾過を行い、その濾液を酸化剤により酸化処理した後、アルカリ剤にて中和処理を行い、塩化浸出液中の不純物を除去し、塩化白金酸溶液を得、更に精製を行ってスポンジ白金を得る廃触媒からの白金及びレニウムの回収方法。 (2) The leaching catalyst and leaching residue after alkali leaching described in (1) above are dried, then added with lead oxide and solvent in an electric furnace, kept at 1350 ° C to 1500 ° C, and lead is absorbed by platinum Method for recovering platinum and rhenium from a waste catalyst that obtains lead (hereinafter referred to as noble lead), oxidizes the noble lead at 1000 ° C to 1200 ° C in an oxidation furnace, and obtains noble metal lead containing 10 to 40 mass% platinum .
(3) After pulverizing the precious metal lead described in (2) above, lead is leached with a nitric acid solution and filtered, and the filtration residue of the nitric acid leaching is leached with hydrochloric acid and hydrogen peroxide and filtered, and the filtrate is oxidized. After oxidizing with an agent, neutralize with an alkaline agent to remove impurities in the chloride leachate, obtain a chloroplatinic acid solution, and further purify to obtain sponge platinum. Collection method.

（４）上記（１）〜（３）において使用するアルカリ剤は、炭酸ナトリウム及びまたは水酸化ナトリウムであり、浸出温度は90℃以上、反応時間は30分以上である廃触媒からの白金及びレニウムの回収方法。
（５）上記（１）〜（４）におけるレニウムを浸出するに必要な炭酸ナトリウム及びまたは水酸化ナトリウムの量は廃触媒中のアルミナ（Al2O3）、Re、Ptを浸出するのに必要な量の0.4倍当量以上である廃触媒からの白金及びレニウムの回収方法。 (4) The alkali agent used in the above (1) to (3) is sodium carbonate and / or sodium hydroxide, the leaching temperature is 90 ° C. or more, and the reaction time is 30 minutes or more. Recovery method.
(5) The amount of sodium carbonate and / or sodium hydroxide required for leaching rhenium in the above (1) to (4) is necessary for leaching alumina (Al 2 O 3 ), Re, and Pt in the waste catalyst. For recovering platinum and rhenium from spent catalyst that is 0.4 equivalents or more of a certain amount.

（６）上記（１）〜（５）において、レニウム浸出液中の白金を還元するには、硫酸第一鉄を理論当量の1.5倍以上添加する廃触媒からの白金及びレニウムの回収方法。
（７）（１）〜（６）における硫酸第一鉄にて白金を還元した後液からレニウムを陰イオン交換樹脂により吸着し、該吸着した陰イオン交換樹脂よりレニウムを7mol/L以上の塩酸溶液を用いて溶離する廃触媒からの白金及びレニウムの回収方法。 (6) In the above (1) to (5), in order to reduce platinum in the rhenium leaching solution, a method for recovering platinum and rhenium from a spent catalyst in which ferrous sulfate is added 1.5 times or more of the theoretical equivalent.
(7) After reducing platinum with ferrous sulfate in (1) to (6), rhenium is adsorbed from the solution by an anion exchange resin, and rhenium is adsorbed at least 7 mol / L from the adsorbed anion exchange resin. A method for recovering platinum and rhenium from a spent catalyst eluting with a solution.

（８）上記（１）〜（７）における溶離液中のレニウムは、塩酸濃度を5〜6.5mol/Lの範囲で、硫化剤を理論量の1.5倍以上添加して硫化レニウムとして回収する廃触媒からの白金及びレニウムの回収方法。
（９）上記（２）〜（８）における、電気炉工程において酸化鉛の添加量は、処理Pt重量の400倍以上であることを特徴とする廃触媒からの白金及びレニウムの回収方法。 (8) The rhenium in the eluent in the above (1) to (7) is recovered as rhenium sulfide by adding 1.5 times or more of the theoretical amount of a sulfiding agent in a hydrochloric acid concentration range of 5 to 6.5 mol / L. A method for recovering platinum and rhenium from a catalyst.
(9) The method for recovering platinum and rhenium from a spent catalyst, wherein the amount of lead oxide added in the electric furnace step in (2) to (8) is 400 times or more the treated Pt weight.

（１０）上記（２）〜（９）における電気炉工程において、原料投入後の保持時間を90分以上である廃触媒からの白金及びレニウムの回収方法。 (10) A method for recovering platinum and rhenium from a spent catalyst in the electric furnace process of (2) to (9) above, wherein the retention time after starting the raw material is 90 minutes or more.

本発明によれば以下の効果を有する。
（１） （１）触媒から白金、レニウムを効率良く高品位で、回収できる。
（２）本発明方法では、従来のようにアルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒からのレニウム及び白金の回収において、多量の酸を使用せず、かつ高回収率で白金、レニウムを分離回収することを極めて効果的に行うことができる画期的な方法である。 The present invention has the following effects.
(1) (1) Platinum and rhenium can be efficiently and highly recovered from the catalyst.
(2) In the method of the present invention, platinum is supported on a porous support containing alumina as in the prior art, rhenium and platinum are recovered from a waste catalyst containing rhenium, and a large amount of acid is not used and high recovery is achieved. It is an epoch-making method that can separate and recover platinum and rhenium at a very high rate.

以下本発明に関して、図１を用いて、より具体的に説明する。
本発明の処理対象物は、アルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒である。 該廃触媒には、白金が0.1〜0.5mass％、レニウムが0.3〜0.6mass％含有している。
アルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒中のアルミナ（Al2O3）、Re、Ptを浸出するのに必要な炭酸ナトリウム及びまたは水酸化ナトリウム量の0.4当量以上含有するアルカリ溶液にて浸出することが好ましい。
又該浸出の温度は、90℃以上であると好適にレニウムを浸出できる。
Re,Pt,Al2O3を浸出するのに必要なNaOH量の0.4当量以上であるNaOH溶液などのアルカリ溶液にて浸出することが好ましい。
又該浸出の温度は、90℃以上であると好適にレニウムを浸出できる。
アルカリ溶液、例えばNaOH溶液による触媒中のレニウム、白金、アルミナを浸出する反応は、次の通りであると推定される。 Hereinafter, the present invention will be described more specifically with reference to FIG.
The object to be treated of the present invention is a waste catalyst containing platinum and rhenium supported on a porous carrier containing alumina. The waste catalyst contains 0.1 to 0.5 mass% platinum and 0.3 to 0.6 mass% rhenium.
0.4 equivalent or more of the amount of sodium carbonate and / or sodium hydroxide required to leach alumina (Al 2 O 3 ), Re, and Pt in the spent catalyst containing platinum and rhenium supported on a porous support containing alumina It is preferable to leach with the contained alkaline solution.
Moreover, rhenium can be suitably leached when the temperature of the leaching is 90 ° C. or higher.
The leaching is preferably performed with an alkaline solution such as a NaOH solution that is 0.4 equivalent or more of the amount of NaOH necessary for leaching Re, Pt, Al 2 O 3 .
Moreover, rhenium can be suitably leached when the temperature of the leaching is 90 ° C. or higher.
The reaction of leaching rhenium, platinum and alumina in the catalyst with an alkaline solution, for example, NaOH solution, is estimated as follows.

HReO4 +
NaOH + 3H2O → NaReO4 + H2O 反応式１
H2PtCl6
+ 6NaOH → Na2(Pt(OH)6 + 2HCl + 4NaCl 反応式２
Al2O3
+ 2NaOH + 6H2O → 2Na〔Al(OH)4〕 + 3H2O 反応式３
その際に白金の一部も浸出するため、所定時間、浸出反応が終了次第、浸出反応槽に硫酸第一鉄を添加し、還元し、白金を浸出残渣として回収する。
その後、浸出した廃触媒を篩目1mmの篩にて粗濾過を行い、浸出後触媒とスラリーに分離して、そのスラリーについては再度濾過を行って固液分離を行う。
その濾液については、1μmのカートリッジフィルターなどで精密濾過を行った後、陰イオン交換樹脂に通液し、前記濾液中のレニウムを吸着し、該樹脂に吸着したレニウムは塩酸溶液により溶離して回収する。 HReO 4 +
NaOH + 3H 2 O → NaReO 4 + H 2 O Reaction formula 1
H 2 PtCl 6
+ 6NaOH → Na 2 (Pt (OH) 6 + 2HCl + 4NaCl Reaction formula 2
Al 2 O 3
+ 2NaOH + 6H 2 O → 2Na [Al (OH) 4 ] + 3H 2 O Reaction formula 3
At this time, a part of platinum is also leached, and as soon as the leaching reaction is completed for a predetermined time, ferrous sulfate is added to the leaching reaction tank and reduced, and platinum is recovered as a leaching residue.
Thereafter, the leached waste catalyst is roughly filtered with a sieve having a mesh size of 1 mm. After leaching, the catalyst and slurry are separated, and the slurry is filtered again for solid-liquid separation.
The filtrate is subjected to microfiltration with a 1 μm cartridge filter or the like, and then passed through an anion exchange resin to adsorb rhenium in the filtrate. The rhenium adsorbed on the resin is recovered by eluting with a hydrochloric acid solution. To do.

この溶離液の塩酸濃度を調整した後、硫化水素ガスを所定量吹込んでレニウムを硫化レニウムとし、一方、液中には硫化水素ガスが溶存しているため、空気にて液中の硫化水素ガスを脱却してから、硫化レニウムをフィルタープレスにて固液分離し回収する。
前記樹脂よりレニウムを吸着回収した後液のアルカリ溶液の一部はレニウム浸出に繰返し、残りは排水処理における中和剤として再利用し、排水処理工程におけるアルカリ剤の代替とする。 After adjusting the hydrochloric acid concentration of this eluent, a predetermined amount of hydrogen sulfide gas is blown to change rhenium to rhenium sulfide. On the other hand, since hydrogen sulfide gas is dissolved in the liquid, the hydrogen sulfide gas in the liquid is in air. Then, the rhenium sulfide is recovered by solid-liquid separation with a filter press.
After the rhenium is adsorbed and recovered from the resin, a part of the alkaline solution in the liquid is repeatedly used for rhenium leaching, and the rest is reused as a neutralizing agent in the wastewater treatment to replace the alkaline agent in the wastewater treatment step.

次に、レニウム浸出工程から産出した浸出触媒、及び浸出残渣は各々ロータリーキルンにて乾燥した後、酸化鉛と溶剤（炭酸カルシウム、珪酸、酸化鉄など）とを混合添加し、電気炉にて1350℃〜1500℃に加熱して溶融する。
上記温度で溶融を続けることで浸出触媒や浸出残渣の主となる成分は溶融したガラス状の酸化物（以下「スラグ」という）の層となり、コークス等の還元剤により還元された酸化鉛は金属鉛となり、比重差により沈降した金属鉛の層を形成する。 Next, the leaching catalyst and leaching residue produced from the rhenium leaching process are each dried in a rotary kiln, and then mixed with lead oxide and a solvent (calcium carbonate, silicic acid, iron oxide, etc.) and 1350 ° C. in an electric furnace. Heat to ~ 1500 ° C to melt.
By continuing to melt at the above temperature, the main component of the leaching catalyst and leaching residue is a layer of molten glassy oxide (hereinafter referred to as “slag”), and lead oxide reduced by a reducing agent such as coke is a metal It becomes lead and forms a layer of metallic lead precipitated due to the difference in specific gravity.

また浸出触媒、あるいは浸出残渣に含まれる白金はスラグ中に分散し、前記金属鉛に吸収されて沈降し、金属鉛の層へ吸収されて廃触媒や浸出残渣から分離される。スラグ中に分散した白金が金属鉛に吸収されて沈降し、金属鉛の層に吸収されるために十分な時間溶融した後、上層の溶融したスラグ層を電気炉より流出させ、次いで下層の溶融した金属鉛の層を抜き出して貴鉛を得る。 Further, platinum contained in the leaching catalyst or leaching residue is dispersed in the slag, absorbed and precipitated by the metallic lead, and absorbed by the metallic lead layer and separated from the waste catalyst and the leaching residue. After the platinum dispersed in the slag is absorbed by the metal lead and settles and melts for a sufficient time to be absorbed by the metal lead layer, the upper molten slag layer is discharged from the electric furnace, and then the lower layer melts The precious lead is obtained by extracting the layer of the metallic lead.

該貴鉛を酸化炉にて投入し、溶湯温度が1000℃〜1200℃の温度範囲にて、空気又は酸素ガスを吹き込んで貴鉛を酸化する。これにより酸化された酸化鉛の層は上層となり、下層には濃縮された白金を吸収している未酸化の金属鉛の層となる。次いで、酸化炉を傾転し、上層の酸化鉛の層を流出して分離した後、下層の白金を含有濃縮している金属鉛の層は流出させて凝固させて貴金属鉛を得る。このプロセスで重要なことは白金を吸収している金属鉛の部分を酸化して酸化鉛の層とし該酸化鉛の層から白金を吸収している金属鉛を分離することで、白金を濃縮する点にある。 The noble lead is charged in an oxidation furnace, and the noble lead is oxidized by blowing air or oxygen gas in a temperature range of 1000 ° C. to 1200 ° C. As a result, the oxidized lead oxide layer becomes an upper layer, and the lower layer becomes an unoxidized metallic lead layer absorbing the concentrated platinum. Next, the oxidation furnace is tilted, and the upper lead oxide layer is flowed out and separated, and then the lower layer platinum-containing metal lead layer is flowed out and solidified to obtain precious metal lead. What is important in this process is to concentrate platinum by oxidizing the portion of metallic lead that absorbs platinum into a lead oxide layer and separating the metallic lead absorbing platinum from the lead oxide layer. In the point.

また、この酸化の過程で金属鉛の表面に酸化鉛の層を形成したまま酸化を継続すると、酸化炉の耐火物が酸化鉛により浸食されるため、酸化鉛の層が炉を傾転して流出分離できる程度の層になった際に分離することが好ましい。
なお、さらに酸化し分離する操作を繰り返して貴金属鉛中の白金品位を高めるようにしてもよい。この貴金属鉛中の白金品位は数mass％から50mass％の範囲で任意にコントロールできるが、この後工程における白金精製工程における白金の回収効率を高めるには白金の品位は10〜40mass％の範囲が好ましい。 In addition, if oxidation is continued with the lead oxide layer formed on the surface of the metal lead during this oxidation process, the refractory in the oxidation furnace is eroded by lead oxide, so the lead oxide layer tilts the furnace. It is preferable to separate the layers when the layers are separated enough to flow out.
Further, the operation of further oxidizing and separating may be repeated to improve the platinum quality in the noble metal lead. The platinum quality in this precious metal lead can be controlled arbitrarily in the range of several mass% to 50 mass%, but the platinum quality is in the range of 10-40 mass% in order to increase the platinum recovery efficiency in the platinum purification process in this subsequent process. preferable.

該貴金属鉛を粉砕機にて1mm以下に粉砕した後、硝酸溶液にて鉛を浸出し、残渣中の白金品位が70mass％以上になるまで、浸出した後、濾過を行う。
この硝酸浸出残渣を塩酸溶液に入れ、過酸化水素を白金の浸出に必要量の1.5倍以上を定量的に添加しながら白金を浸出する。この反応の際に発熱反応で溶液の温度が上昇するため、浸出反応が終了後、放冷してから、溶液と残渣を固液分離して、塩化白金酸溶液を得る。
該塩化白金酸溶液には鉛などの重金属を含有しているため、精製を行ってスポンジ白金を得る。 After pulverizing the noble metal lead to 1 mm or less with a pulverizer, the lead is leached with a nitric acid solution, and leached until the platinum quality in the residue becomes 70 mass% or more, followed by filtration.
This nitric acid leaching residue is put into a hydrochloric acid solution, and platinum is leached while quantitatively adding hydrogen peroxide at least 1.5 times the amount necessary for leaching platinum. Since the temperature of the solution rises due to an exothermic reaction during this reaction, after the leaching reaction is completed, the solution and the residue are separated by solid-liquid separation to obtain a chloroplatinic acid solution.
Since the chloroplatinic acid solution contains heavy metals such as lead, purification is performed to obtain sponge platinum.

次に本発明のアルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒から白金、レニウムの回収方法に関わる実施例を記載する。   Next, examples relating to a method for recovering platinum and rhenium from a spent catalyst containing platinum and rhenium supported on a porous carrier containing alumina of the present invention will be described.

（実施例１）
アルミナを含む多孔質担体上に担持された白金を0.22mass％、レニウムを0.43mass％含む廃触媒15gを、各種浸出液150mLを各種濃度にて室温から90℃の温度条件にて、反応時間30分から120分の間にて浸出した後濾過を行い、濾液量を測定後、濾液を分析してレニウム浸出率を求めたところ表１の結果を得た。 Example 1
15 minutes of waste catalyst containing 0.22 mass% platinum and 0.43 mass% rhenium supported on a porous carrier containing alumina, 150 ml of various leachates at various concentrations from room temperature to 90 ° C, reaction time from 30 minutes After leaching for 120 minutes, filtration was performed, and after measuring the amount of filtrate, the filtrate was analyzed to obtain the rhenium leaching rate. The results shown in Table 1 were obtained.

レニウム浸出率の高いのは、アルカリ溶液で、中でも炭酸ナトリウム（Na2CO3）または水酸化ナトリウム（NaOH）溶液であった。
浸出温度条件は、いずれの条件とも温度が高い方が良く、90℃以上にすれば、レニウム浸出率は90％以上得られる。 The high rhenium leaching rate was an alkaline solution, particularly a sodium carbonate (Na 2 CO 3 ) or sodium hydroxide (NaOH) solution.
As for the leaching temperature condition, it is better that the temperature is high in any condition. If the leaching temperature is 90 ° C. or higher, the rhenium leaching rate is 90% or higher.

4mass％NaOH溶液において、反応時間は30分以上で、レニウム浸出率は90％以上得られた。









In the 4 mass% NaOH solution, the reaction time was 30 minutes or more and the rhenium leaching rate was 90% or more.









（実施例２）
レニウム浸出工程において、レニウム浸出率とNaOH溶液濃度との関係を把握するために、廃触媒15gをNaOH/(Re+Pt+Al2O3)＝0.006〜6.9当量の範囲の各種濃度のNaOH溶液150mLに入れて、１時間浸出した後濾過を行い、濾液量を測定後、濾液を分析して各成分の浸出率を求めたところ、図２の結果を得た。NaOHを理論量の0.4倍以上添加すれば、レニウムは90％以上の浸出率が得られた。 (Example 2)
In the rhenium leaching process, in order to grasp the relationship between the rhenium leaching rate and the NaOH solution concentration, 15 g of spent catalyst was added to NaOH / (Re + Pt + Al 2 O 3 ) = 0.006 to 6.9 equivalents in various NaOH solutions. After putting into 150 mL and leaching for 1 hour, filtration was performed, and after measuring the amount of filtrate, the filtrate was analyzed to determine the leaching rate of each component. The result shown in FIG. 2 was obtained. When NaOH was added more than 0.4 times the theoretical amount, leaching rate of rhenium was over 90%.

（実施例３）
アルミナを含む多孔質担体上に担持された白金を0.22mass％、レニウムを0.43mass％含む廃触媒220gを、理論量の0.4倍当量のNaOH濃度溶液1.1Lに入れ、温度90℃にて１時間レニウムの浸出を行った後、固液分離を行い、濾液中の白金濃度を分析した。 (Example 3)
220g of spent catalyst containing 0.22mass% platinum and 0.43mass% rhenium supported on a porous carrier containing alumina is placed in 1.1L of NaOH concentration solution 0.4 times the theoretical amount and heated at 90 ° C for 1 hour. After leaching rhenium, solid-liquid separation was performed, and the platinum concentration in the filtrate was analyzed.

濾液中の白金濃度は64mg/L、レニウム濃度は800mg/Lであった。該濾液を200mLずつ5つに分取し、90℃に加熱した溶液に、濃度43g/Lの硫酸第一鉄（FeSO4・7H2O）溶液を各条件に必要な量を添加し、10分間反応させた後、濾過を行い濾液中の白金、レニウム濃度を分析したところ、表２の結果を得た。 The platinum concentration in the filtrate was 64 mg / L and the rhenium concentration was 800 mg / L. The filtrate was fractionated into five 200 mL portions, and a ferrous sulfate (FeSO 4 .7H 2 O) solution having a concentration of 43 g / L was added to the solution heated to 90 ° C. in an amount necessary for each condition. After reacting for 5 minutes, filtration was performed and the platinum and rhenium concentrations in the filtrate were analyzed, and the results shown in Table 2 were obtained.

白金の還元に必要な硫酸第一鉄溶液を理論量の1.5倍以上添加すれば、浸出液中の白金は還元され、還元後液中の白金濃度は1mg/L以下になる。
If the ferrous sulfate solution necessary for the reduction of platinum is added more than 1.5 times the theoretical amount, the platinum in the leachate will be reduced, and the platinum concentration in the solution after reduction will be 1 mg / L or less.

（実施例４）
アルミナを含む多孔質担体上に担持された白金を0.22mass％、レニウムを0.43mass％含む廃触媒中のレニウムをNaOH溶液にて浸出した後、硫酸第一鉄にて、白金を還元除去した後液を1μmのフィルターにて濾過した濾液1000mLに、各種陰イオン交換樹脂（三菱化学製のPA408、PA316、オルガノ製IRA400、住友化学製SA20A）20mLをいれ、１６時間吸着攪拌した後、液中のレニウム濃度を分析したところ、いずれの樹脂とも吸着後液中のレニウム濃度は1mg/L以下になった。 Example 4
After leaching out rhenium in a spent catalyst containing 0.22 mass% platinum and 0.43 mass% rhenium on a porous carrier containing alumina with NaOH solution, after reducing and removing platinum with ferrous sulfate To 1000 mL of the filtrate obtained by filtering the liquid with a 1 μm filter, 20 mL of various anion exchange resins (Mitsubishi Chemical PA408, PA316, Organo IRA400, Sumitomo Chemical SA20A) were added and stirred for 16 hours. As a result of analyzing the rhenium concentration, the rhenium concentration in the solution after adsorption was 1 mg / L or less for both resins.

吸着液と樹脂とを濾別し、樹脂を純水200mLに入れて30分間攪拌洗浄した後、樹脂を濾別し、この操作を2回繰り返した。次に、洗浄後の樹脂を8mol/Lの塩酸溶液200mLに入れて１時間攪拌し、吸着したレニウムを溶離し、溶離液と樹脂とを分離した後、溶離後液中のレニウム濃度を分析した。
この溶離操作を2回繰り返して、レニウムの溶離率を求めたところ、表３の結果を得た。 The adsorbed liquid and the resin were separated by filtration, the resin was put into 200 mL of pure water and stirred and washed for 30 minutes, and then the resin was separated by filtration, and this operation was repeated twice. Next, the washed resin was placed in 200 mL of 8 mol / L hydrochloric acid solution and stirred for 1 hour to elute the adsorbed rhenium, separate the eluent and the resin, and then analyzed the rhenium concentration in the liquid after elution. .
This elution operation was repeated twice to obtain the elution rate of rhenium. The results shown in Table 3 were obtained.

レニウム溶離率が最も高かったのは三菱化成製のPA408であった。






The highest rhenium elution rate was PA408 manufactured by Mitsubishi Kasei.






（実施例５）
上記の三菱化成製PA408 20mLをガラス製のカラムに充填し、前述の廃触媒をNaOH溶液にて浸出した後、硫酸第一鉄溶液にて白金を還元した後液を1μmのフィルターにて濾過した濾液（Re濃度570mg/L）3Lを樹脂カラム内に通液速度1.67mL/min(SV=5hr-)にて2L通液してレニウムを吸着した後、純水を通液速度1.67mL/minにて100mL通液して水洗した後、各種濃度の塩酸溶液を通液速度0.67mL/min(SV=2hr-)にて200mL通液し、溶離後液中のレニウム濃度を分析し、レニウムの溶離量を求めたところ、表４の結果を得た。 (Example 5)
20 ml of the above-mentioned Mitsubishi Kasei PA408 was packed in a glass column, and after leaching the above-mentioned waste catalyst with NaOH solution, platinum was reduced with ferrous sulfate solution, and the solution was filtered with a 1 μm filter. the filtrate (Re concentration 570 mg / L) passed through 3L in the resin column velocity 1.67mL / min (SV = 5hr - ) at was adsorbed rhenium with 2L liquid passing pure water liquid passage rate of 1.67 mL / min at washed with water and 100mL liquid passing, hydrochloric acid solutions of various concentrations liquid permeation speed of 0.67mL / min (SV = 2hr - ) and 200mL passed through by analyzes rhenium concentration in the eluting solution after, rhenium When the elution amount was determined, the results shown in Table 4 were obtained.

７moL/L以上の塩酸濃度で溶離すれば、吸着したレニウムは98％以上溶離できる。
Elution with a hydrochloric acid concentration of 7 moL / L or more can adsorb 98% or more of the adsorbed rhenium.

（実施例６）
NaOH濃度15g/Lの溶液2m3を90℃に加温し、その中にアルミナを含む多孔質担体上に担持された白金を0.22mass％、レニウムを0.43mass％含む廃触媒375kgを装入し、30分浸出した後、43g/Lの硫酸第一鉄を理論量の1.5倍量添加して10分間攪拌した後、溶液を60℃以下まで冷却し、先に浸出触媒を1mmのフィルターで濾過を行った。その後、濾過スラリーを固液分離した後、濾液を1μmのフィルターにて精密濾過した。その濾液を上記の条件にて、陰イオン交換樹脂PA408にてレニウムを吸着・溶離し、得られた溶離液を純水にて、各種塩酸濃度に調整した溶液を650Lずつ準備した。 (Example 6)
The solution 2m 3 of NaOH concentration 15 g / L was heated to 90 ℃, 0.22mass% platinum supported on a porous support comprising alumina therein, it was charged with spent catalyst 375kg including 0.43Mass% rhenium After leaching for 30 minutes, add 43g / L ferrous sulfate 1.5 times the theoretical amount and stir for 10 minutes, then cool the solution to below 60 ° C and filter the leaching catalyst first with a 1mm filter Went. Thereafter, the filtration slurry was subjected to solid-liquid separation, and then the filtrate was microfiltered with a 1 μm filter. Under the above conditions, rhenium was adsorbed and eluted with the anion exchange resin PA408, and the obtained eluent was adjusted with pure water to 650 L each with various hydrochloric acid concentrations.

この溶液を硫化反応槽に入れ、硫化水素ガスを15L/minの流量で、レニウムを硫化するために必要量に対して1.5倍当量吹込んでレニウムを硫化して、硫化後液中のレニウム濃度を分析したところ、表５の結果を得た。 This solution is put into a sulfurization reactor, and hydrogen sulfide gas is blown at a flow rate of 15 L / min at a flow rate of 1.5 times equivalent to the amount required to sulfite rhenium to sulfite rhenium, and the rhenium concentration in the solution after sulfidation is reduced. When analyzed, the results in Table 5 were obtained.

硫化前液中の塩酸濃度を5〜6.5mol/Lの間で硫化処理を行えば、レニウムは98％以上の回収率が得られた。
When sulfurization treatment was performed at a hydrochloric acid concentration of 5 to 6.5 mol / L in the pre-sulfurization solution, a recovery rate of 98% or more of rhenium was obtained.

（実施例７）
前述の条件にて、レニウム浸出工程において固液分離し回収した浸出触媒（Pt品位；0.19％）をドライヤにて乾燥し、その乾燥した浸出触媒250kgと、溶剤としての石灰石（炭酸カルシウム）525kg、珪砂（シリカ）150kg、酸化鉄125kg、還元剤としてコークス粒を酸化鉛を還元するのに必要な理論カーボン量の2.5倍量を添加し、白金の吸収媒体である酸化鉛を処理Pt重量に対して各種割合（＝酸化鉛/Pt重量比；10〜900）を混合し、電気炉に投入し1400℃に加熱した。 (Example 7)
Under the above-mentioned conditions, the leaching catalyst (Pt quality; 0.19%) separated and recovered in the rhenium leaching step was dried with a dryer, and the dried leaching catalyst 250 kg and limestone (calcium carbonate) 525 kg as a solvent, Silica sand (silica) 150 kg, iron oxide 125 kg, coke grains as a reducing agent is added 2.5 times the theoretical carbon amount required to reduce lead oxide, and lead oxide, the platinum absorption medium, is treated to the Pt weight Various ratios (= lead oxide / Pt weight ratio; 10 to 900) were mixed, put into an electric furnace and heated to 1400 ° C.

溶融状態で90分間保持した後、電気炉の上層に生成したスラグ層を電気炉の側面より流出させ、次いで下層の金属鉛の層を電気炉の下部よりレードルに注ぎ冷却固化し、貴鉛を得た。
この際、各種酸化鉛の添加量とスラグ中のPtの含有率を分析したところ、表６の結果が得られた。酸化鉛を処理Pt重量に対して400倍以上添加すれば、スラグ中の白金品位は5mass
ppm以下となり、電気炉におけるPt回収率は99％以上になった。 After maintaining in the molten state for 90 minutes, the slag layer formed in the upper layer of the electric furnace is allowed to flow out from the side of the electric furnace, and then the lower layer of metallic lead is poured into the ladle from the lower part of the electric furnace to solidify by cooling. Obtained.
At this time, when the amount of various lead oxides added and the content of Pt in the slag were analyzed, the results shown in Table 6 were obtained. If lead oxide is added more than 400 times the treated Pt weight, the platinum quality in the slag is 5 mass.
The Pt recovery rate in the electric furnace exceeded 99%.





（実施例８）
前述の条件にて、レニウム浸出工程において固液分離し回収した浸出触媒と、浸出残渣を各々ドライヤにて乾燥し、その乾燥した廃触媒（Pt品位；0.18mass％）250kgと、浸出残渣30kg(Pt品位；0.7mass％)、溶剤としての石灰石（炭酸カルシウム）525kg、珪砂(シリカ)100kg、酸化鉄100kg、還元剤としてコークス粒を20kg、酸化鉛270kgを混合し、電気炉に投入して1400℃に加熱し、原料投入後各時間おきに、電気炉上部より溶融状態のスラグを採取して、冷却固化後のスラグ中Pt品位を分析したところ、表７の結果を得た。 (Example 8)
Under the conditions described above, the leaching catalyst separated and recovered in the rhenium leaching process and the leaching residue were each dried in a dryer, and the dried waste catalyst (Pt grade; 0.18 mass%) 250 kg, and the leaching residue 30 kg ( Pt quality: 0.7 mass%), 525 kg of limestone (calcium carbonate) as a solvent, 100 kg of silica sand (silica), 100 kg of iron oxide, 20 kg of coke grains as a reducing agent, 270 kg of lead oxide, mixed in an electric furnace, 1400 When heated to ° C. and each time after the raw material was charged, slag in a molten state was collected from the upper part of the electric furnace and analyzed for Pt quality in the slag after cooling and solidification. The results shown in Table 7 were obtained.

電気炉へ原料投入後、保持持間を60分以上にすれば、スラグ中のPt品位は5mass ppm以下になり、90分以上保持すれば、さらにスラグ中のPt品位は2mass ppmまで低下した。













After the raw material was charged into the electric furnace, the Pt quality in the slag became 5 mass ppm or less when the holding time was 60 minutes or longer, and the Pt quality in the slag further decreased to 2 mass ppm when held for 90 minutes or longer.













本発明方法の各工程を示すフローチャートである。It is a flowchart which shows each process of this invention method. 実施例３におけるレニウム浸出工程におけるレニウム浸出率とNaOH/(Re+Pt+Al2O3)比との関係を表した図である。Is a diagram showing a relationship between rhenium leaching rate and NaOH / (Re + Pt + Al 2 O 3) ratio in the rhenium leaching process in the third embodiment.

Claims (10)

アルミナを含む多孔質担体上に担持された白金、レニウムを含む廃触媒からの白金、及びレニウムを回収する方法において、
前記廃触媒中の白金，レニウムをアルカリ溶液により浸出し、浸出液中の白金を還元し、濾過後、浸出液中のレニウムを陰イオン交換樹脂により吸着し、前記樹脂よりレニウムを塩酸溶液により溶離し、溶離後液中のレニウムは硫化処理を行って硫化レニウムとして回収することを特徴とする廃触媒からの白金及びレニウムの回収方法。 In a method of recovering platinum supported on a porous support containing alumina, platinum from a waste catalyst containing rhenium, and rhenium,
Platinum and rhenium in the waste catalyst are leached with an alkaline solution, platinum in the leachate is reduced, filtered, rhenium in the leachate is adsorbed with an anion exchange resin, and rhenium is eluted from the resin with a hydrochloric acid solution, A method for recovering platinum and rhenium from a waste catalyst, wherein rhenium in the solution after elution is subjected to sulfidation treatment and recovered as rhenium sulfide. 請求項１記載のアルカリ浸出後の浸出触媒及び浸出残渣は乾燥後、電気炉において酸化鉛と溶剤を添加し、1350℃~1500℃に保持し、鉛に白金を吸収させた金属鉛（以下、貴鉛と記す）を得、該貴鉛を酸化炉にて1000℃~1200℃において酸化し、白金を10~40mass%含む貴金属鉛を得ることを特徴とする廃触媒からの白金及びレニウムの回収方法。 The leaching catalyst and the leaching residue after alkaline leaching according to claim 1 are dried, then lead oxide and a solvent are added in an electric furnace, maintained at 1350 ° C to 1500 ° C, and lead is absorbed in platinum (hereinafter referred to as “lead”). Recovery of platinum and rhenium from a waste catalyst characterized in that the noble lead is oxidized at 1000 ° C. to 1200 ° C. in an oxidation furnace to obtain noble metal lead containing 10 to 40 mass% platinum. Method. 請求項２記載の貴金属鉛を粉砕後、硝酸溶液にて鉛を浸出、濾過し、その硝酸浸出の濾過残渣を塩酸と過酸化水素により塩化浸出、濾過を行い、その濾液を酸化剤により酸化処理した後、アルカリ剤にて中和処理を行い、塩化浸出液中の不純物を除去し、塩化白金酸溶液を得、更に精製を行ってスポンジ白金を得ることを特徴とする廃触媒からの白金及びレニウムの回収方法。 After pulverizing the precious metal lead according to claim 2, the lead is leached with a nitric acid solution and filtered, and the filtration residue of the nitric acid leaching is leached with hydrochloric acid and hydrogen peroxide and filtered, and the filtrate is oxidized with an oxidizing agent. The platinum and rhenium from the spent catalyst is characterized by neutralizing with an alkaline agent to remove impurities in the leaching solution, obtaining a chloroplatinic acid solution, and further purifying to obtain sponge platinum. Recovery method. 請求項１〜３において使用するアルカリ剤は、炭酸ナトリウム及びまたは水酸化ナトリウムであり、浸出温度は90℃以上、反応時間は30分以上であることを特徴とする廃触媒からの白金及びレニウムの回収方法。 The alkali agent used in claims 1 to 3 is sodium carbonate and / or sodium hydroxide, the leaching temperature is 90 ° C. or more, and the reaction time is 30 minutes or more. Collection method. 請求項１〜４におけるレニウムを浸出するに必要な炭酸ナトリウム及びまたは水酸化ナトリウムの量は廃触媒中のアルミナ（Al2O3）、Re、Ptを浸出するのに必要な量の0.4倍当量以上であることを特徴とする廃触媒からの白金及びレニウムの回収方法。 The amount of sodium carbonate and / or sodium hydroxide required to leach rhenium in claims 1 to 4 is 0.4 times equivalent to the amount required to leach alumina (Al 2 O 3 ), Re and Pt in the spent catalyst. A method for recovering platinum and rhenium from a waste catalyst characterized by the above. 請求項１〜５において、レニウム浸出液中の白金を還元するには、硫酸第一鉄を理論当量の1.5倍以上添加することを特徴とする廃触媒からの白金及びレニウムの回収方法。 6. The method for recovering platinum and rhenium from a spent catalyst according to claim 1, wherein in order to reduce platinum in the rhenium leaching solution, ferrous sulfate is added at least 1.5 times the theoretical equivalent. 請求項１〜６における硫酸第一鉄にて白金を還元した後液からレニウムを陰イオン交換樹脂により吸着し、該吸着した陰イオン交換樹脂よりレニウムを7mol/L以上の塩酸溶液を用いて溶離することを特徴とする廃触媒からの白金及びレニウムの回収方法。 The rhenium is adsorbed from the solution after reducing platinum with ferrous sulfate in claims 1 to 6 by an anion exchange resin, and the rhenium is eluted from the adsorbed anion exchange resin with a hydrochloric acid solution of 7 mol / L or more. A process for recovering platinum and rhenium from a spent catalyst. 請求項１〜７における溶離液中のレニウムは、塩酸濃度を5〜6.5mol/Lの範囲で、硫化剤を理論量の1.5倍以上添加して硫化レニウムとして回収することを特徴とする廃触媒からの白金及びレニウムの回収方法。 8. The spent catalyst according to claim 1, wherein rhenium in the eluent is recovered as rhenium sulfide by adding 1.5 times or more of a theoretical amount of a sulfurizing agent in a hydrochloric acid concentration range of 5 to 6.5 mol / L. For recovery of platinum and rhenium from lime. 請求項２〜８における、電気炉工程において酸化鉛の添加量は、処理Pt重量の400倍以上であることを特徴とする廃触媒からの白金及びレニウムの回収方法。 9. The method for recovering platinum and rhenium from a spent catalyst, wherein the amount of lead oxide added in the electric furnace step is 400 times or more the weight of the treated Pt. 請求項２〜９における電気炉工程において、原料投入後の保持時間を90分以上であることを特徴とする廃触媒からの白金及びレニウムの回収方法。








































10. A method for recovering platinum and rhenium from a waste catalyst, characterized in that, in the electric furnace process according to claims 2 to 9, the holding time after the raw material is charged is 90 minutes or more.








































JP2004320180A 2004-11-04 2004-11-04 Method for recovering platinum and rhenium from spent catalyst Expired - Fee Related JP4347783B2 (en)

Priority Applications (2)

Application Number Priority Date Filing Date Title
JP2004320180A JP4347783B2 (en) 2004-11-04 2004-11-04 Method for recovering platinum and rhenium from spent catalyst
CNB2005100785703A CN100348749C (en) 2004-11-04 2005-06-17 Method of recovering platinum and rhenium from waste catalyst

Applications Claiming Priority (1)

Application Number Priority Date Filing Date Title
JP2004320180A JP4347783B2 (en) 2004-11-04 2004-11-04 Method for recovering platinum and rhenium from spent catalyst

Publications (2)

Publication Number Publication Date
JP2006130387A true JP2006130387A (en) 2006-05-25
JP4347783B2 JP4347783B2 (en) 2009-10-21

Family

ID=36724395

Family Applications (1)

Application Number Title Priority Date Filing Date
JP2004320180A Expired - Fee Related JP4347783B2 (en) 2004-11-04 2004-11-04 Method for recovering platinum and rhenium from spent catalyst

Country Status (2)

Country Link
JP (1) JP4347783B2 (en)
CN (1) CN100348749C (en)

Cited By (15)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN100489131C (en) * 2007-12-12 2009-05-20 安泰科技股份有限公司 Method for manufacturing difficult-to-form pure rhenium sheet
JP2010077499A (en) * 2008-09-26 2010-04-08 Dowa Metals & Mining Co Ltd Method for treating substance to be treated containing platinum group element, rhenium and arsenic
JP2010077470A (en) * 2008-09-25 2010-04-08 Dowa Metals & Mining Co Ltd Method for treating material to be treated, containing platinum group element, rhenium and/or arsenic
JP2013249503A (en) * 2012-05-31 2013-12-12 Sumitomo Metal Mining Co Ltd Method for recovering rhenium
CN103789552A (en) * 2014-02-27 2014-05-14 西北有色金属研究院 Method for recovering rhenium from high-temperature alloy pickle liquor
JP2014173107A (en) * 2013-03-07 2014-09-22 Jx Nippon Mining & Metals Corp Method for recovering platinum group elements
KR101448168B1 (en) 2012-08-23 2014-10-13 엘에스니꼬동제련 주식회사 The recovery method of rhenium from the weak acid at copper smelter
KR101507792B1 (en) * 2013-08-29 2015-04-07 서울시립대학교 산학협력단 Method of recovering precious metal
EP2547803A4 (en) * 2010-03-16 2016-10-26 World Resources Co A method for recovering rhenium and other metals from rhenium-bearing materials
US9708690B2 (en) * 2015-10-26 2017-07-18 Techemet, Llc Method for platinum recovery from materials containing rhenium and platinum metals
WO2017167553A1 (en) * 2016-03-29 2017-10-05 Heraeus Deutschland GmbH & Co. KG Method for recovering rhenium and at least one platinum group element from catalyst particles
CN113877531A (en) * 2021-11-05 2022-01-04 福州大学 Preparation method and application of acid-resistant nano-alumina-loaded biochar
CN114058849A (en) * 2021-11-03 2022-02-18 核工业北京化工冶金研究院 Method for recovering platinum and aluminum from aluminum-based platinum catalyst
CN115874060A (en) * 2022-12-19 2023-03-31 本源精化环保科技有限公司 Method for extracting precious metal from waste supported composite catalyst
CN116607010A (en) * 2023-07-19 2023-08-18 长沙华时捷环保科技发展股份有限公司 Method for removing and recovering lead from lead-containing solution

Families Citing this family (19)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN101549891B (en) * 2008-04-03 2011-07-06 天津市风船化学试剂科技有限公司 Method for preparing granular platinic chloride
CN102094129B (en) * 2010-12-24 2012-08-01 辽宁大学 Method for enriching rhenium
CN102936658B (en) * 2011-08-15 2014-05-07 深圳市格林美高新技术股份有限公司 Recovery method for platinum, aluminium and rhenium from waste reforming catalyst
CN102703737B (en) * 2012-06-28 2013-08-07 青岛阿库卡稀有金属有限公司 Method for extracting and separating rhenium from rhenium-containing slag
WO2014038048A1 (en) * 2012-09-06 2014-03-13 株式会社キャタラー Method for recovering precious metal from metal carrier catalyst
CN102925713B (en) * 2012-11-05 2013-11-06 中国海洋石油总公司 Method for recovering rhodium from rhodium-containing waste catalyst in hydroformylation through sulfur
CN103388077B (en) * 2013-07-23 2015-04-01 贵研资源（易门）有限公司 Comprehensive recycling method of platinum-rhenium containing reformed spent catalyst
CN104178634A (en) * 2014-08-19 2014-12-03 昆明贵金属研究所 Method for efficiently and cleanly recovering platinum group metals from spent automobile catalyst
CN105483373B (en) * 2014-09-19 2017-11-28 有研稀土新材料股份有限公司 A kind of leaching agent and leaching method for being used to leach ion adsorption type rare earth ore middle rare earth
CN105983707B (en) * 2015-01-27 2019-02-15 昆明冶金高等专科学校 A method of high-purity rhenium powder is prepared from rhenium-containing high arsenic-and copper-bearing sulfide
CN106566929B (en) * 2016-11-10 2018-04-06 永兴贵研资源有限公司 A kind of recovery method of high-carbon rheniforming dead catalyst
CN106756085B (en) * 2016-12-22 2019-02-05 邝贤语 A kind of smelting process of platinum concentrate
CN108277350B (en) * 2018-02-06 2020-02-21 贵研资源（易门）有限公司 Method for recovering platinum and rhenium from spent alumina-based platinum-rhenium catalyst
CN109097584B (en) * 2018-08-27 2020-03-17 浙江特力再生资源有限公司 Process for recovering platinum-rhenium-aluminum from platinum-rhenium waste catalyst of aluminum carrier
AU2020298961A1 (en) * 2019-07-02 2022-02-03 Nihon Medi-Physics Co., Ltd. METHOD FOR PURIFYING 226Ra-CONTAINING SOLUTION, METHOD FOR PRODUCING 226Ra TARGET, AND METHOD FOR PRODUCING 225Ac
CN110724829A (en) * 2019-10-23 2020-01-24 金川集团股份有限公司 Method for efficiently enriching noble metals from difficultly-treated palladium-containing waste catalyst
CN111411240A (en) * 2020-04-23 2020-07-14 贺利氏贵金属技术(中国)有限公司 Method for recovering noble metal from waste catalyst containing noble metal
CN111549226A (en) * 2020-04-24 2020-08-18 金川集团股份有限公司 Method for recovering nickel, platinum and ruthenium from failed Raney nickel catalyst
CN112961986A (en) * 2021-02-06 2021-06-15 中海油(山西)贵金属有限公司 Method for recovering noble metal in platinum online recoverer

Family Cites Families (2)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US3672874A (en) * 1970-11-12 1972-06-27 Universal Oil Prod Co Recovery of rhenium values from a spent catalyst
RU2100072C1 (en) * 1996-09-25 1997-12-27 Омский государственный университет Method of recovering platinum and rhenium from treated platinum/rhenium catalyst

Cited By (17)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN100489131C (en) * 2007-12-12 2009-05-20 安泰科技股份有限公司 Method for manufacturing difficult-to-form pure rhenium sheet
JP2010077470A (en) * 2008-09-25 2010-04-08 Dowa Metals & Mining Co Ltd Method for treating material to be treated, containing platinum group element, rhenium and/or arsenic
JP2010077499A (en) * 2008-09-26 2010-04-08 Dowa Metals & Mining Co Ltd Method for treating substance to be treated containing platinum group element, rhenium and arsenic
EP2547803A4 (en) * 2010-03-16 2016-10-26 World Resources Co A method for recovering rhenium and other metals from rhenium-bearing materials
JP2013249503A (en) * 2012-05-31 2013-12-12 Sumitomo Metal Mining Co Ltd Method for recovering rhenium
KR101448168B1 (en) 2012-08-23 2014-10-13 엘에스니꼬동제련 주식회사 The recovery method of rhenium from the weak acid at copper smelter
JP2014173107A (en) * 2013-03-07 2014-09-22 Jx Nippon Mining & Metals Corp Method for recovering platinum group elements
KR101507792B1 (en) * 2013-08-29 2015-04-07 서울시립대학교 산학협력단 Method of recovering precious metal
CN103789552A (en) * 2014-02-27 2014-05-14 西北有色金属研究院 Method for recovering rhenium from high-temperature alloy pickle liquor
US9708690B2 (en) * 2015-10-26 2017-07-18 Techemet, Llc Method for platinum recovery from materials containing rhenium and platinum metals
WO2017167553A1 (en) * 2016-03-29 2017-10-05 Heraeus Deutschland GmbH & Co. KG Method for recovering rhenium and at least one platinum group element from catalyst particles
CN114058849A (en) * 2021-11-03 2022-02-18 核工业北京化工冶金研究院 Method for recovering platinum and aluminum from aluminum-based platinum catalyst
CN113877531A (en) * 2021-11-05 2022-01-04 福州大学 Preparation method and application of acid-resistant nano-alumina-loaded biochar
CN113877531B (en) * 2021-11-05 2023-02-14 福州大学 Preparation method and application of acid-resistant nano-alumina-loaded biochar
CN115874060A (en) * 2022-12-19 2023-03-31 本源精化环保科技有限公司 Method for extracting precious metal from waste supported composite catalyst
CN116607010A (en) * 2023-07-19 2023-08-18 长沙华时捷环保科技发展股份有限公司 Method for removing and recovering lead from lead-containing solution
CN116607010B (en) * 2023-07-19 2023-10-17 长沙华时捷环保科技发展股份有限公司 Method for removing and recovering lead from lead-containing solution

Also Published As

Publication number Publication date
JP4347783B2 (en) 2009-10-21
CN100348749C (en) 2007-11-14
CN1769504A (en) 2006-05-10

Similar Documents

Publication Publication Date Title
JP4347783B2 (en) Method for recovering platinum and rhenium from spent catalyst
Hoffmann Recovery of platinum-group metals from gabbroic rocks metals from auto catalysts
WO2014181721A1 (en) Method for recovering scandium
JP2008531842A (en) Formation of nickel iron
CN102534244A (en) Method for concentrating precious metal from low-grade precious metal material
US20120148461A1 (en) Process for multi metal separation from raw materials and system for use
WO2013108478A1 (en) Gold recovery method, and gold production method using same
WO2021014946A1 (en) Pgm recovery method
JP5016929B2 (en) Arsenic solution purification method
JP5196095B2 (en) Precious metal recovery method and recovered precious metal
JPS63197521A (en) Method of removing gaseous mercury from gas
RU2306347C1 (en) Method of processing catalysts containing platinum metals and rhenium on aluminum oxide carriers
CN111455192A (en) Method for recovering palladium from low-grade palladium-containing indissolvable waste catalyst
JP5077517B2 (en) Precious metal recovery method and recovered precious metal
JP5263483B2 (en) Precious metal recovery method
JP4457864B2 (en) Method for recovering nickel and / or cobalt sulfide
CN116716493A (en) Method for secondarily recycling germanium from low-grade germanium-containing material
JP3906333B2 (en) Precious metal recovery method
JP6352702B2 (en) Method for producing perrhenic acid aqueous solution and method for producing potassium perrhenate, ammonium perrhenate and rhenium metal using the same
JP4406745B2 (en) Method for processing Sn, Pb, Cu-containing material
JP2023019519A (en) Method for recovering platinum group metals
US20230083521A1 (en) Platinum-group metal recovery method, composition containing platinum-group metals, and ceramic material
CN108425014A (en) A method of extracting high pure metal palladium from useless palladium-aluminium oxide catalyst
JP6314044B2 (en) Method for producing aqueous perrhenic acid solution from crude rhenium sulfide
CN113088704B (en) Method for recovering precious metal from glass fiber refractory brick

Legal Events

Date Code Title Description
A711 Notification of change in applicant

Free format text: JAPANESE INTERMEDIATE CODE: A712

Effective date: 20060510

A621 Written request for application examination

Free format text: JAPANESE INTERMEDIATE CODE: A621

Effective date: 20060829

A977 Report on retrieval

Free format text: JAPANESE INTERMEDIATE CODE: A971007

Effective date: 20090423

A131 Notification of reasons for refusal

Free format text: JAPANESE INTERMEDIATE CODE: A131

Effective date: 20090512

A521 Written amendment

Free format text: JAPANESE INTERMEDIATE CODE: A523

Effective date: 20090616

TRDD Decision of grant or rejection written
A01 Written decision to grant a patent or to grant a registration (utility model)

Free format text: JAPANESE INTERMEDIATE CODE: A01

Effective date: 20090714

A01 Written decision to grant a patent or to grant a registration (utility model)

Free format text: JAPANESE INTERMEDIATE CODE: A01

A61 First payment of annual fees (during grant procedure)

Free format text: JAPANESE INTERMEDIATE CODE: A61

Effective date: 20090716

R150 Certificate of patent or registration of utility model

Free format text: JAPANESE INTERMEDIATE CODE: R150

FPAY Renewal fee payment (event date is renewal date of database)

Free format text: PAYMENT UNTIL: 20120724

Year of fee payment: 3

FPAY Renewal fee payment (event date is renewal date of database)

Free format text: PAYMENT UNTIL: 20120724

Year of fee payment: 3

S111 Request for change of ownership or part of ownership

Free format text: JAPANESE INTERMEDIATE CODE: R313111

FPAY Renewal fee payment (event date is renewal date of database)

Free format text: PAYMENT UNTIL: 20120724

Year of fee payment: 3

R350 Written notification of registration of transfer

Free format text: JAPANESE INTERMEDIATE CODE: R350

LAPS Cancellation because of no payment of annual fees