EP0042702A1 - Procédé pour la récupération de plomb et d'argent à partir de résidus - Google Patents

Procédé pour la récupération de plomb et d'argent à partir de résidus Download PDF

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Publication number
EP0042702A1
EP0042702A1 EP81302614A EP81302614A EP0042702A1 EP 0042702 A1 EP0042702 A1 EP 0042702A1 EP 81302614 A EP81302614 A EP 81302614A EP 81302614 A EP81302614 A EP 81302614A EP 0042702 A1 EP0042702 A1 EP 0042702A1
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EP
European Patent Office
Prior art keywords
lead
silver
chloride
process according
brine
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Application number
EP81302614A
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German (de)
English (en)
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EP0042702B1 (fr
Inventor
Robert Steven Salter
Roy Slater Boorman
Ross David Gilders
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Provincial Holdings Ltd
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Provincial Holdings Ltd
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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B11/00—Obtaining noble metals
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B13/00—Obtaining lead

Definitions

  • This invention relates to the extraction of metal compounds from metal bearing materials and more particularly to the.extraction and recovery of lead values in. a calcium plumbate and/or oxide product from minerals or lead bearing materials.
  • Silver which may be present in association with the lead may be recovered as native silver, silver chloride, sulfide or sulphate, or a silver complex with other metals, or in some other form from which it can be recovered by conventional techniques.
  • West German Patent 2,500,453, (1976) describes a method of leaching lead sulphate containing material in sodium chloride solution and after residue separation, precipitating the Pb from solution with milk of lime.
  • the lead precipitates contain greater than 10% chloride and 11% sulphate and are not acceptable to conventional lead smelters except in small amount and at depressed prices due to the deleterious effects of chlorides on. acid plant catalysts, baghouses, and refractories.
  • Australian Patent 28,957,(1971) describes a method of precipitating lead chloride from brine solution by cooling followed by reacting said lead chloride with water and calcium sulphate to produce a lead sulphate precipitate and calcium chloride solution.
  • low chloride levels in the lead sulphate were obtainable with rigorous washing, the product is again suitable to lead smelters in limited quantities and must be treated on a sinter machine to remove sulphur before the blast.
  • Capital and operating costs are projected to be high for the process since the brine solution must be heated for high lead solubility and then cooled for lead chloride precipitation.
  • Canadian Patent 13918 describes a method of precipitating lead and silver from brine solution by cementation on metallic zinc. Recently other researchers have rediscovered the cementation techique and advocate either zinc or iron as cementation media. High grade metallic lead and silver cementates are produced in these processes which are acceptable to lead smelters at premium prices at high tonnage. Considerable economic penalties are incurred to produce good quality cementates, however, since the cementation reagents are expensive and the zinc or iron in the lean brine resulting from cementation must be recovered in a form acceptable for sale or reuse. This can be accomplished only at considerable cost.
  • a process comprising the steps of (1) preparing a solution of lead chloride by dissolving lead sulphate contained in an ore or process residue in an acidic concentrated chloride brine; (2) separating the solution so formed from insoluble gangue or other residue; (3) forming a precipitate of lead oxychloride by adding lime.to said solution and separating said lead oxychloride precipitate from the residual lean brine solution; (4) reacting the said oxychloride precipitate with oxygen and lime in a reactor at an elevated temperature to produce a calcine containing most of the lead as calcium plumbates and/or lead oxides; (5) washing said calcine in water and/or dilute chloride brine to dissolve soluble chlorides; (6) separating the resulting residue obtained from the resulting chloride brine; and, (7) washing said residue containing calcium plumbates and/or lead oxides with fresh water to remove residual chlorides.
  • a process for gaining lead and silver values comprising the steps of (1) preparing a solution of lead chloride and silver compounds by dissolving lead sulphate and silver compounds contained in an ore or process residue in an acidic concentrated chloride brine; (2) separating the solution so formed from insoluble gangue or other residue; (3) forming a precipitate of lead oxychloride and silver compounds by adding lime to said solution and separating said precipitate from the residual lean brine solution; (4) reacting the said oxychloride precipitate with oxygen and lime in a reactor at an elevated temperature to produce a calcine containing most of the lead as calcium plumbates and/or lead oxides and most of the silver as silver or silver compounds; (5) washing said calcine in water and/or dilute chloride brine to dissolve soluble chlorides; (6) separating the resulting residue from the resulting chloride brine; and, (7) washing said residue containing calcium plumbates and/or lead oxides, as well as silver and silver compounds, with fresh water to remove
  • a process comprising the step of (1) preparing a solution of lead chloride by dissolving lead sulphate contained in ore or process residues in concentrated chloride brine, thereby also dissolving any silver associated with the lead; (2) separating the solution so formed from the insoluble gangue and other residues; (3) forming a precipitate of lead oxychloride (ana any silver which may be present) by adding lime to the solution and separating the lead oxychloride and silver precipitate from the residual lean brine solution; (4) recycling the lean brine, normally after concentration thereof such as by evaporation or by addition of further chloride and also normally after re-acidification by the addition of further acid, for reuse in the further extraction of lead sulphate as under steps (1) and (2).
  • the improvement comprises (5) reacting the said precipitate containing lead oxychloride with oxygen such as by air and with excess lime present in the precipitate, and if desired adding fresh lime, in a reactor at a temperature above 325°C for longer than one half hour to produce a calcine containing most of the lead as calcium plumbates and lead oxides, and containing any silver present as native silver, silver chloride, sulfide or sulphate, and complexes of silver with other materials; (6) repulping said calcine in water and/or dilute chloride brine to remove soluble chlorides;(7) separating the residue obtained in step (6) from the resulting chloride brine; (8) recycling the brine resulting from step (7), with the optional treatment mentioned above, for further extraction of lead sulphate under the previous steps; (9) washing the said residue from step (7) with fresh water to remove residual chlorides; and (10) recycling the chloride brine obtained in step (9) to step (6) and/or recycling the said chloride brine, again with the optional treatment mentioned
  • Lead and silver extractions into brine can be accelerated by increasing the acidity by addition of an acid such as hydrochloric or sulfuric acid which will ensure at least'mildly acidic conditions.
  • an acid such as hydrochloric or sulfuric acid which will ensure at least'mildly acidic conditions.
  • the optimum p H in the brine leach for high lead and silver extraction, efficient residue washing, and low lime consumption appears to be about 1.5.
  • Extractions of lead from lead sulphate material into brine are very high and may approach 99% with the proper choice of retention time, temperature, brine composition, and residue washing techniques as long as the solubility limit of the lead is not approached.
  • Lead extractions fall from 99% at 75% of lead chloride saturation to 96% at 86% of saturation to 91% at 94% of saturation for brine leaching in 269 gpl NaCl- 33 gpl CaC1 2 - pH 1.5 brine at 35°C and 1.5 hours leaching time.
  • the saturation limit of lead as lead chloride in this brine is 18.3 gpl.
  • Silver extraction by brine is very dependent on the nature and prior history of the lead sulphate containing material. Some materials, usually those which have been very recently produced in a roaster or leach process exhibit silver extractions greater than 80%. Other materials, usually stockpiled, exhibit lower . silver extractions of about 50%. Silver recoveries can be increased from these materials by flotation recovery of a silver concentrate and using the present process on the flotation tailings which contain most of the lead and all the remaining silver. The silver flotation concentrate and the plumbate product can then be combined for sale to conventional lead smelters. Flotation processes such as described by Moriyama,E. and Yamamoto,Y: in AIME World Symposium of Mining and Metallurgy of Lead and Zinc, Vol. II, 1970, page 215 have been shown to yield silver concentrates with high silver assays and recoveries from lead sulphate containing materials.
  • Silver can be removed from the brine leach solution by cementation on a suitable metallic medium such as zinc, iron, or lead. With proper stoichiometric conditions, retention time, and pH nearly all the silver can be recovered in a high grade metallic product containing some little lead and copper as contaminants. Lead along with the solubilized cementation agent would then be recovered in the plumbate product as in steps (3) - (9).
  • the lead compounds formed will depend on-the pH or the mole ratio of base to lead chloride and the total chloride concentration. Table 1 shows the effect of these variables on the nature of the lead precipitate when lead is precipitated from a brine solution containing 15 gpl lead as lead chloride at 45°C and a retention time of 1.5 hours. Shorter retention times can be employed but chloride levels in the precipitate will increase unless the temperature is increased above 45°C. Silver is coprecipitated with lead. Most of the brine soluble impurities which are present in the lead sulphate containing starting material such as zinc, copper, iron, bismuth, and arsenic also coprecipitate with lead.
  • the precipitate is repulped with water or any reasonably unsaturated brine solution produced in the process.
  • Lime is added to the pulp to bring the mole ratio of the total lime addition in the process to between 2.5 and 5.5.
  • the pulp is subjected to thermal treatment.
  • the lead oxychloride precipitate may be blended with lime to increase the total mole ratio to between 2.5 and 5.5, without repulping and the blend subjected to thermal treatment.
  • step (3) If the lime addition in step (3) was greater than 2.5 mole ratio to lead, then repulping and/or further lime addition is not necessary before thermal treatment.
  • the lead oxychloride-lime blend is heated in a reactor in the presence of oxygen or air.
  • the reactor can be a rotary kiln, furnace, roaster, autoclave or any device commonly used for thermal treatment.
  • the retention time in the reactor depends upon the desired degree of conversion of oxychloride to plumbate and oxide, the temperature in the reator, the lime to lead mole ratio, and the oxygen partial pressure. The effects of these variables on the nature of the calcine product are shown in Table 2.. At reaction temperatures above 500 C sintering of the product appears and volatilization of lead chloride begins.
  • Preferred conditions appear to be a total process lime addition to lead mole ratio of about 3, a reaction temperature of about 400 o C, a retention time of about 1 hour, and an excess of oxygen for lead oxidation. Pressure above atmospheric is not required for the reaction to be complete within 2 hours. Sufficient air or oxygen can be supplied for the reaction by convection, free or forced, or by pressurizing the reactor.
  • calcium orthoplumbate (Ca 2 Pb0 4 ) is formed at high yields with such low temperatures and short retention times.
  • Calcium orthoplumbate in a pure form (>90%) is a valuable commodity and is used- in the manufacture of primers for steel and galvanized steel, of pigments and of binders for paints. It also has use in the plastics and resin industries.
  • the common commercial production method is the reaction of PbO with lime and air or oxygen at temperatures above 700°C. The reaction kinetics are reportedly slow below this temperature and the reaction will not go below 500°C. It has been reported by Denev, D.G.
  • the present invention differs from the customary practice in that the reagent for calcination is lead oxychloride and not lead oxide. Also, the oxychloride is contaminated with significant quantities of NaCl and CaCl2. Accordingly, the kinetics and energetics of plumbate formation have been altered significantly from commercial experience.
  • Figure 3 is a schematic plant layout for a particular embodiment of the invention relating to example 1.
  • Samples of hot sulphuric acid leach residues obtained from the sulphation roasting and leaching of bulk zinc-lead-copper-silver sulphide concentrates assaying 30-32% Zn, 3.5-10% Pb, 0.7% Cu, 4.4 - 8.8 oz/ST (troy ounce per short ton) silver, and 14-23% iron were processed accordingly to the invention.
  • a sample (example 10) of a hot acid leach residue from a dead roast zinc plant was also processed. The residues assayed as in Table 3.
  • One pregnant brine solution (example #6) was treated with zinc dust at an addition rate of 0.5 gpl producing a cementate containing 99% of the silver, 95% of the copper and 80% of the bismuth in the pregnant brine.
  • the remaining pregnant brines and the solution resulting from the zinc dust cementation test were then treated with lime at 1.5 mole ratio to lead in the brine solution.
  • the temperature and retention time were 45°C and 1.5 hours respectively.
  • Precipitates were allowed to settle and the thickened precipitates filtered.
  • the precipitates in all tests were then blended with lime, as required, to bring the total mole ratio of lime added to the process to lead in the precipitate to 1.75-5.5.
  • the blends were treated in an oven with a slow purge of fresh air for 1.0 hours at 400 0 C.
  • the calcines were repulped to 50% pulp density for 15 minutes in fresh water and filtered and displacement washed with a volume of water equal to the calcine repulp water.
  • the assays (dry basis) of the resulting plumbate products are given in Table 6.
  • a sample of the hot sulfuric acid leach residue used in example 5 was repulped in water at 10% pulp density, American Cyanamid flotation reagents Aero 404 (trade mark) promoter and Aerofroth 77A (trade mark) (frother) were added at 600 g. and 60 g . per metric ton of residue respectively. After 5 minutes conditioning time, flotation was initiated and a stable froth was maintained for about 7.5 minutes. The concentrate obtained assayed 18% Pb, 24% Fe, 1.8% Zn, and 635 oz/ST silver. Silver recovery from the head material was 70%.
  • the process of this invention may be applied to the recovery of metals from a variety of metallurgical products such as ores and concentrates, smelter dusts, metal drosses, middling concentrates from flotation processing, slags and process residues, and other like sources of lead and silver.

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  • Engineering & Computer Science (AREA)
  • Chemical & Material Sciences (AREA)
  • Manufacturing & Machinery (AREA)
  • Materials Engineering (AREA)
  • Mechanical Engineering (AREA)
  • Metallurgy (AREA)
  • Organic Chemistry (AREA)
  • Manufacture And Refinement Of Metals (AREA)
EP81302614A 1980-06-16 1981-06-12 Procédé pour la récupération de plomb et d'argent à partir de résidus Expired EP0042702B1 (fr)

Applications Claiming Priority (2)

Application Number Priority Date Filing Date Title
CA000354083A CA1156048A (fr) 1980-06-16 1980-06-16 Methode de recuperation de plomb et d'argent a partir de mineraux et de residus de procede
CA354083 1980-06-16

Publications (2)

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EP0042702A1 true EP0042702A1 (fr) 1981-12-30
EP0042702B1 EP0042702B1 (fr) 1984-09-26

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EP81302614A Expired EP0042702B1 (fr) 1980-06-16 1981-06-12 Procédé pour la récupération de plomb et d'argent à partir de résidus

Country Status (10)

Country Link
EP (1) EP0042702B1 (fr)
JP (1) JPS5729541A (fr)
AU (1) AU549357B2 (fr)
CA (1) CA1156048A (fr)
DE (1) DE3166293D1 (fr)
ES (1) ES8301284A1 (fr)
FI (1) FI71342C (fr)
IE (1) IE52179B1 (fr)
PT (1) PT73185B (fr)
ZA (1) ZA813982B (fr)

Cited By (4)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN101994007B (zh) * 2009-08-28 2012-08-15 沈阳有色金属研究院 用氯化镁从废铅酸蓄电池膏泥中脱硫的方法
CN104789790A (zh) * 2015-04-08 2015-07-22 吉林吉恩镍业股份有限公司 尼尔森重选含铅金精矿无铅化冶炼工艺
RU2670117C2 (ru) * 2013-09-27 2018-10-18 Текникас Реунидас, С.А. Способ селективного извлечения свинца и серебра и карбонатный концентрат свинца и серебра, полученный вышеуказанным способом
CN112442602A (zh) * 2020-10-09 2021-03-05 超威电源集团有限公司 一种废旧铅膏回收方法

Families Citing this family (3)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
JP5046263B2 (ja) * 2005-06-24 2012-10-10 株式会社吉野工業所 液注出容器のキャップ
JP4781794B2 (ja) * 2005-11-28 2011-09-28 キユーピー株式会社 注出容器
CN109022817A (zh) * 2018-07-27 2018-12-18 郴州雄风环保科技有限公司 高氯铅银渣脱氯的新工艺

Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US1434085A (en) * 1920-01-02 1922-10-31 Niels C Christensen Process of treating ores with chloride solutions
US1745945A (en) * 1924-01-08 1930-02-04 Us Smelting Refining & Mining Process of treating ores or analogous materials
GB365964A (en) * 1930-06-17 1932-01-28 Paul Gamichon Process for converting into soluble salts lead and other metals contained in lead bearing ores
US3477928A (en) * 1966-03-28 1969-11-11 Cerro Corp Process for the recovery of metals
FR2297253A1 (fr) * 1975-01-08 1976-08-06 Duisburger Kupferhuette Procede pour la production de plomb d'oeuvre a partir de matieres variees contenant le metal
FR2459292A1 (fr) * 1980-03-24 1981-01-09 Asua Ind Quim Procede pour la modification physico-chimique de residus argento-plombiferes de l'industrie hydrometallurgique du zinc en vue de la recuperation simultanee de l'argent et du plomb

Patent Citations (6)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US1434085A (en) * 1920-01-02 1922-10-31 Niels C Christensen Process of treating ores with chloride solutions
US1745945A (en) * 1924-01-08 1930-02-04 Us Smelting Refining & Mining Process of treating ores or analogous materials
GB365964A (en) * 1930-06-17 1932-01-28 Paul Gamichon Process for converting into soluble salts lead and other metals contained in lead bearing ores
US3477928A (en) * 1966-03-28 1969-11-11 Cerro Corp Process for the recovery of metals
FR2297253A1 (fr) * 1975-01-08 1976-08-06 Duisburger Kupferhuette Procede pour la production de plomb d'oeuvre a partir de matieres variees contenant le metal
FR2459292A1 (fr) * 1980-03-24 1981-01-09 Asua Ind Quim Procede pour la modification physico-chimique de residus argento-plombiferes de l'industrie hydrometallurgique du zinc en vue de la recuperation simultanee de l'argent et du plomb

Cited By (4)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN101994007B (zh) * 2009-08-28 2012-08-15 沈阳有色金属研究院 用氯化镁从废铅酸蓄电池膏泥中脱硫的方法
RU2670117C2 (ru) * 2013-09-27 2018-10-18 Текникас Реунидас, С.А. Способ селективного извлечения свинца и серебра и карбонатный концентрат свинца и серебра, полученный вышеуказанным способом
CN104789790A (zh) * 2015-04-08 2015-07-22 吉林吉恩镍业股份有限公司 尼尔森重选含铅金精矿无铅化冶炼工艺
CN112442602A (zh) * 2020-10-09 2021-03-05 超威电源集团有限公司 一种废旧铅膏回收方法

Also Published As

Publication number Publication date
PT73185A (en) 1981-07-01
FI811847L (fi) 1981-12-17
FI71342C (fi) 1986-12-19
ES502948A0 (es) 1982-11-16
JPS5729541A (en) 1982-02-17
AU7184681A (en) 1981-12-24
IE811310L (en) 1981-12-16
JPS6352094B2 (fr) 1988-10-18
DE3166293D1 (en) 1984-10-31
AU549357B2 (en) 1986-01-23
PT73185B (en) 1982-07-16
EP0042702B1 (fr) 1984-09-26
ZA813982B (en) 1982-08-25
FI71342B (fi) 1986-09-09
ES8301284A1 (es) 1982-11-16
IE52179B1 (en) 1987-08-05
CA1156048A (fr) 1983-11-01

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