hydro of lead

8/13/2019 Hydro of Lead

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NEW TECHNOLOGY FOR LEAD

The Hydrometallurgy of Lead

Fathi HabashiDepartment of Mining, Metallurgical, and Materials Engineering

Laval University, Quebec City, Canada G1V 0A6e-mail: [email protected]

ABSTRACT

Lead is an ancient metal, has been produced to-date from its ores exclusively bypyrometallurgical route. The process suffers from high operating cost and excessive pollutionproblems. Extensive research has been going on since the beginning of the twentieth century tofind a non-polluting process for its production and a solution for its complex refining scheme. Ithas been assumed correctly that the hydrometallurgical route should be the most promising. Pilotplants have been constructed and operated for a reasonable periods, but no process has proved tobe fully satisfactory. As a result, a number of smelters and refineries have been shut down. Thepresent review analyzes the situation and suggests that the best route to treat lead sulfideconcentrates is by leaching with fluoroboric acid, HBF4, containing ferric fluoroborate, Fe(BF4)3,and electrolyzing the solution in a diaphragm cell - - a process recently invented by Italianmetallurgists. In this process, any silver present in the concentrate remains in the residue togetherwith elemental sulfur and can be recovered by known methods.

INTRODUCTION

Lead has been produced for thousands of years exclusively by thermal methods. The complexrefining steps and the pollution in the neighborhood of smelters are causing much trouble to thenearby population. For example, the high content of lead in crab collected from the ocean in thevicinity of a lead smelter worries consumers. The appreciable amounts of lead in wine producedfrom a vineyard near a lead smelter causes concern to the industry. The maximum permissiblelimit for lead in the vicinity of a smelter is 0.05 mg Pb/m3of air or 6 ppb, a limit that is difficultto achieve in many plants. In addition, lead smelters produce SO2and this must be converted tosulfuric acid. A nearby market for the acid must exist otherwise SO2will have to be emitted inthe environment, which is unacceptable. Both problems can only be solved by using ahydrometallurgical route to process the sulfide concentrate. No lead fumes will be emitted in theenvironment and elemental sulfur could be produced, which is easy to store or ship to sulfuricacid manufacturers.


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PRESENT TECHNOLOGY

The present process for lead production suffers from numerous steps, high operating cost, andexcessive pollution problems (Figure 1).

Figure 1 - General scheme for the pyrometallurgy of lead

RefiningMolten lead is sent for refining in kettles at controlled temperature to collect the dross, otherimpurities are then removed, then silver is recovered by zinc, and finally silver is recovered bycupellation while zinc is recovered by vacuum distillation (Figures 2-4).

Figure 2 - General scheme of lead refining


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Figure 3 - Refining of lead

Figure 4 - Removal of silver

HYDROMETALLURGY OF LEAD

The first review on the hydrometallurgy of lead was published in 1924 by Ralston of the US

Bureau of Mines in Berkeley, California 1. He drew attention to the solubility of PbCl2 andPbSO4in brine solutions and summarized their solubility data. The most important process thatwas operating on a small industrial scale using this technology was that developed by Taintonand his co-workers at Bunker Hill & Sullivan in Kellogg, Idaho (Figure 5).

PbS concentrate

Pb

Ca(OH)2

H2O

Cl2

O2

CaCl2

Roasting

Leaching

Filtration

PbSO4, impurities

SO2

Absorption

Impurities

Gangue, CaSO4

Ag

PbCl42-

Leaching

Filtration

Aq. Electrolysis

Refining

Figure 5- Tainton process (1924)


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Tainton himself never published a description of his process, but an account of it was given in

1945 by Bownan in Liddells Handbook of Nonferrous Metallurgy 2 and in 1953 by Van

Arsdale in Hydrometallurgy of Base Metals3. The process, however, was not successful andthe plant was dismantled. The failure of the process seems to have been in the electrowinningstep since chlorine was formed and was not disposed of properly beside its corrosion problems.

In addition, lead powder obtained was not satisfactorily handled and was contaminated by silver.Intensive research on the hydrometallurgy of lead was undertaken in 1970s and 1980s whenenvironmental problems due to lead smelters became of much concern. However, no viableprocess resulted. In the present review it will be shown that the production of lead by ahydrometallurgical route is certainly possible.

Electrometallurgy of lead

Lead may be deposited from aqueous media at high cathode efficiencies. Lead electrodepositedfrom aqueous solutions of the acetate and nitrate baths have no commercial application becausethe deposits tend to be acicular and porous hence difficult to wash. The fluoborate, fluosilicate,

and sulfamate baths have achieved technical importance for commercial application because ofthe smoothness of the deposits.

Electrodeposition of lead from fluoroborate bath was patented in 1886 by G. Leuchs in Germany[4]. The solution is prepared from basic lead carbonate and fluoroboric acid. The bath containing100g/L Pb, 42 g/L fluoroboric acid, 10g/L boric acid, and 0.2 g/L glue operated at 15-40C.Electrowinning of lead in nitrate and fluorosilicate media yields lead metal at the cathodes andPbO2 at the anodes. Addition of small amounts of phosphate or arsenic compound influorosilicate system will prevent the formation of PbO2 on anodes. Impurities should beremoved from solution before electrowinning. It was also proposed to precipitate lead as PbSO4leaving impurities in solution converting it to PbCO3by ammonium carbonate, then dissolving

the carbonate in H2SiF6for electrolysis. Ammonium sulfate generated during this process can bemarketed as a fertilizer. This proposal, however, was not attractive to industry.

The electrolytic refining of lead was invented in 1901 by Anson G. Betts [5]. Crude lead wascast in form of anodes and immersed in an electrolyte of lead fluorosilicate solution. Thecathodes were made of pure lead. When the current was passed, pure lead deposited on thecathode while the impurities remained as slimes on the bottom of the cell. The process is usedtoday for refining of about 20% of the lead bullion. Bullion usually contains 12-15% impurities.It is necessary to control the level of arsenic, antimony, and bismuth in the anode to maintain itsintegrity and slime adherence during electrorefining.

Electrolysis of fused PbCl2 NaCl at 500C has been successfully operated and chlorinegenerated can be used in the oxidation of ferrous chloride to ferric, which is used in the leachingstep, or for the direct attack of lead sulfide concentrate. Lead forms as a molten pool at thebottom of the cell [6].

BASIC CHEMISTRY

The fact that PbSO4is insoluble in water while ZnSO4and CuSO4are soluble suggested to early


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metallurgists a method for separating these metals, which are usually occurring together in ores,by roasting at about 500C:

Sg 42SMSOOMS

followed by leaching with water. A small amount of SO2 is formed during roasting due to thepresence of other sulfides, which form oxides instead of sulfates [1]. Later, it was found that

slurrying the concentrate with water then heating in an autoclave at 220C:

aqS,aq 42SMSOOMS

followed by filtration resulted in a solution containing all the copper and zinc sulfates while allthe lead and silver sulfates remained in the residue and can be sent to the blast furnace. This wasthe process adopted on industrial scale in 1960s by Bunker Hill in Kellogg, Idaho (Figure 6 andTable 1) [7]. This technology was thought to be more advantageous than the roasting processbecause no SO2would be evolved but in fact it was not, since SO2was generated in the blastfurnace gases due to the decomposition of lead sulfate. It had the disadvantage further thatsulfate ion in solution had to be disposed of by precipitation with lime.

oncentrate

Pb

Ag

PbSO4, Ag2SO4

Crude Pb

Refining

H2OO2

CuSO4, ZnSO4

CO2, SO2

Aq. Oxidation

Filtration

Blast Furnace

Figure 6- Processing of lead sulfide concentrates at Bunker Hill, Kellogg, Idaho (1960s).High temperature aqueous oxidation (220oC)

Further research demonstrated that there are a large number of options available to choose fromfor the hydrometallurgical treatment of lead sulfide concentrates without the production of SO2.The following points should be taken in consideration:

Silver, zinc, and copper are usually important by-products. One should recall the following solubility data to help selecting the proper leaching

system: PbCl2, AgCl, PbSO4, and Ag2SO4 are insoluble in water but soluble inconcentrated brine solution. Lead fluorosilicate, PbSiF6, is soluble in water while Ag2SiF6is not. Sulfates and chlorides of copper (II) and zinc are soluble in water. All nitrates aresoluble while carbonates are not.


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Galena is easily attacked by dilute acids generating H2S [8-10]:PbS + 2H+ Pb

2+ + H2S

which can be collected and converted by standard technology to elemental sulfur at400C using alumina as catalyst:

H2S + O2 S + H2O

However, the toxicity of H2S and its explosive nature renders this route undesirable.

Galena is attacked by concentrated H2SO4at 100C to form SO2and elemental sulfur:PbS + 2H2SO4 PbSO4 + SO2 + S + 2H2O

The reaction is simple but offers no special advantage since SO2is generated.

Table 1 Aqueous oxidation of lead sulfide concentrate at 220C(Bunker Hill Process)

Feed%

Solutiong/L

Residue%

Pb 50.6 Trace 51.2Zn 8.7 49.5 1.20Cu 6.46 47.8 0.83Fe 7.8 4.0 6.7S 16.4 67.6 9.4As 2.2 1.36 2.0Sb 2.0 0.15 1.8

Bi 0.006 0.0 0.006

Sn 0.07 Trace 0.06Cd 0.15 0.85 0.02

Co 0.025 0.13 0.01

Ni 0.40 1.78 0.11Mn 0.14 - 0.10In 0.08 0.06 0.04Ca 0.2 0.0 0.2Mg 0.15 - 0.10Insol. 2.3 - 3.0

Ag

(oz/ton) 131.7 Trace 140.1

SULFATE AND CHLORIDE SYSTEMS

Aqueous oxidation of PbS in acid results in the formation of elemental sulfur [11,12]:

PbS + O2 + 2H+ Pb

2+ + S + H2O


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When H2SO4 is used, PbSO4will be formed and when HCl is used then PbCl2will be formedtogether with an appreciable amount of PbSO4 since a portion of sulfide sulfur is oxidized toSO4

2- and other components of the concentrate will form soluble sulfates. There is no advantagein using the sulfate system because of the difficulties encountered in the recovery step. WhenFe

3+ion is used instead of oxygen the following reaction takes place [13-16]:

PbS + 2Fe3+ Pb2+ + 2Fe2+ + S

Solution purification can be achieved by cementation of the impurities with lead powder. Ferrousion can then be oxidized back to ferric for recycle. Oxidation may take place by oxygen:

Fe2+

+ O2 + 2H+ 2Fe

3++ H2O

Or by chlorine when FeCl3is used:

2Fe2+ + Cl2(aq) 2Fe3+ + 2Cl-

In this case chlorine could be obtained from the electrolysis of PbCl2either in aqueous solution(complexed with NaCl) or in the molten state (Figure 7). This was the basis of the processesdeveloped by researchers at US Bureau of Mines at Reno, Nevada [17-22] and others [23]. Thisshows again that the chloride system is more preferable than the sulfate system, since in the lattercase the sulfate ion must be disposed of. Oxidation of Fe

2+ may also be achieved in the

electrolytic step whereby the evolution of chlorine is suppressed as proposed by Frenchresearchers [24].

PbS concentrate

Aq. Oxidation

Filtration

Flotation S

Leaching

Filtration

Fe2+

oxidation Crystallization

PbCl2

Electolysis

Pb

HCl Fe

3+

Brine

Soluble chlorides

Gangue

Cl2

Figure 7- Processing lead sulfide concentrates with formation

of elemental sulfur. Chloride system


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Other researchers added a variety of reagents during the aqueous oxidation step to solubilizePbSO4formed. For example, ammonium acetate [25,26], ammonia [27], sodium hydroxide [28],and amines [29]. Instead of using acid and ferric ion, aqueous chlorine solutions are used [30].In all these systems there is always formation of variable amounts of PbSO4.

CARBONATE SYSTEM

To avoid the formation of PbSO4or PbCl2, aqueous oxidation of PbS was conducted by Chinese

researchers 31-33 in presence of ammonium carbonate at about 50C. Lead carbonate andelemental sulfur are formed:

PbS + (NH4)2CO3 + O2 + H2O PbCO3 + S + 2NH4OH

Residence time about 6 hours and yield of sulfur is 60%. After flotation of elemental sulfur,PbCO3 was solubilized in fluorosilicic acid, the solution purified, then electrolyzed for

electrowinning of lead (Figure 8).concentrate

Pb

Impurities

Solids

PbCO3, gangue

Residue containing AgFiltration

Purification

Electrolysis

Flotation

Dissolution

O2

H2SiF6

Recovery

(NH4)2CO3

S, Ag

Aq. Oxidation

Filtration

Figure 8 - Processing lead sulfide concentrates in carbonate system

NITRATE SYSTEM

The nitrate system has the advantage that both lead and silver will go into solution and henceseparation can be readily achieved. Using HNO3, however, has the disadvantage of generatingnitric gases, which must be re-converted to HNO3 [34]. The use of ferric nitrate was already

proposed in 1939 by Fande [35and was studied further in 1987 by Fuerstenau et al. 36who


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showed that the following reaction takes place:

PbS + 2Fe(NO3)3 Pb(NO3)2 + 2Fe(NO3)2 + S

Complete dissolution of galena took place at 70C in 0.25 MFe(NO3)3solution at pH 1.2-1.4 in

100 minutes. Recently Pashkov and co-workers [37] demonstrated that lead could be selectivelyextracted from lead sulfide concentrates into solution at room temperature in less than 30minutes using ferric nitrate. It was proposed to remove iron via ferrous nitrate oxidation andferric nitrate self-decomposition in autoclave, yielding hematite and NO + NO2 gases; thisprocess has been previously studied by Van Weert and co-workers [38,39]:

6Fe(NO3)2+ 5H2O3Fe2O3+ 2NO + 10HNO3

Nitrous gases are employed to produce nitric acid and to dissolve hematite forming ferric nitratesolution for recycle:

Fe2O3+ 6HNO32Fe(NO3)3+ 3H2O

After purification of the solution from copper and bismuth by cementation on lead and theseparation, if necessary, of zinc by organic solvents, lead can be recovered by electrowinning atthe cathode as Pb and at the anode as PbO2(Figure 9):

Pb(NO3)2+ 2e-Pb + 2NO3

-

Pb(NO3)2+ 2H2O + 2NO3-PbO2+ 4HNO3+ 2 e

-

overall reaction

2Pb(NO3)2+ 2H2O Pb + PbO2+ 4HNO3

concentrate

Gangue S

HNO3 Fe2O3

Regeneration

Dissolution

Solution

Bi, Cu, Ag, Zn

NO + NO2

Fe(NO3)3

Leaching

Filtration

Heating

Flotation

Solution

Pb + PbO2

Filtration

Purification

Electrolysis

Figure 9 - Processing lead sulfide concentrates in nitrate system


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This system, however, will not solve the problem of lead metallurgy because of the need torecover nitric gases to form nitric acid, the use of diaphragm cell is undesirable, and only half ofthe lead is recovered in the metallic state. Cementation of lead by iron from nitrate solution isnot possible because of the oxidizing power of the nitrate ion and the formation of iron oxide[40].

FLUOROSILICATE SYSTEM

Since lead fluorosilicate is soluble in water, it was suggested by workers at US Bureau of Mines

in Rolla, Missouri 41,42to leach lead sulfide concentrate in fluorosilicic acid:

PbS + H2SiF6+ O2PbSiF6+ S + H2O

After filtration of the residue and solution purification, lead can be recovered by electrolysis. Theresidue contained elemental sulfur, silver, zinc, and copper (Figure 10). A recent study in thisdirection was published by Taylor et al. [43]. Attempts to electrowin lead from the leach solution

were not successful because of low current efficiency and undesirable cathode morphology dueto impurities present. To overcome this problem, the Bureau of Mines researchers suggestedadding H2SO4 to precipitate PbSO4, transform the sulfate into carbonate, dissolving PbCO3 inH2SiF6, then electrolyzing the pure lead fluorosilicate solution. High purity lead was obtainedbut the procedure suffers from the numerous steps involved and the generation of ammoniumsulfate as a by-product.

PbS concentrate

Silver recover

Pb

Impurities

H2SiF6 O2

Flotation S

Solids

Solution

Leaching

Filtration

Purification

Electrolysis

Figure 10 - Processing lead sulfide concentrates in fluorosilicate system

CHLORINATION

The use of gaseous chlorine has the advantage over the aqueous chloride system is the absence ofPbSO4 formation since all the sulfide sulfur is transformed to the elemental form. There havebeen early attempts in this direction at the beginning of the twentieth century but without success

owing to the difficulties encountered in handling chlorine 47. About sixty years later,


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researchers at the US Bureau of Mines in Rolla, Missouri re-examined this technology and

recommended the recovery of lead by the electrolysis of fused PbCl248. A pilot plant was later

operated at Hazen Research Center in Golden, Colorado based on such technology 49(Figure11). Chlorine gas is fed to rotating reactor counter-current to the flow of fresh concentrate so thatany sulfur monochloride formed becomes the chlorinating agent:

PbS + Cl2PbCl2+ S

2S + Cl2S2Cl2

PbS + S2Cl2PbCl2+ 3S

The temperature in the reactor is 155-175C.

PbS

Cl2 concentrate

Chlorination

Leaching

Filtration S + gangue

Crystallization Fe powder

Centrifuge Cementation

PbCl2

Electrolysis Filtration

Pb Silver recovery

Solution for Cu, Zn recovery

Brine

Figure 11 Processing lead sulfide concentrate by chlorination

Lead chloride and other chlorinated compounds are then solubilized in hot brine solution. Leadchloride is then crystallized and, the anhydrous PbCl2is fed in a fused salt cell containing 90%

PbCl2and 10% NaCl and operating at 500C. High purity lead was obtained. Mother liquor fromcrystallization step is treated with sponge iron to remove silver. A bleed stream is treated withNaOH or Na2CO3to remove other impurities. Work at Universal Oil Products laboratory in Des

Plaines, Illinois also confirmed this technology 50.

FLUOROBORATE SYSTEM

Researchers at Engitec Tehnologies in Milan, Italy found out that galena concentrates wassolubilized in fluoroboric acid containing ferric fluoroborate at 80

oC liberating elemental sulfur

while silver, copper, bismuth, and antimony remain in the residue [44-46] (Figure 12):

PbS + 2Fe(BF4)3Pb(BF4)2+ 2Fe(BF4)2+ S


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The solution is then electrolyzed in a diaphragm cell where pure compact lead is deposited at thecathode and the ferrous fluoroborate is oxidized at the anode to ferric fluoroborate for recycle:

Pb2+

+ 2e-Pb

Fe2+Fe3++ e-

PbS concentrate

Silver recover

Bleed for zinc recovery

HBF4+ Fe(BF4)3

Flotation S

Solids

Solution

Leaching

Filtration

Electrolysis

Lead

Figure 12 - Processing lead sulfide concentrates in fluoroborate system

No oxygen is formed at the anode and this is a great advantage because the voltage will be lower.Air is sparged at the anodes to prevent the formation of PbO2. A typical electrolyte will have thefollowing composition in g/L: Pb2+50-70, Fe2+25-30, Fe3+35-30, fluoroboric acid 35-45, totalHBF4325-335. Temperature of electrolysis 35

oC, anodic and cathodic current densities 200-250A/m

2, and graphite anodes can be used. Cell voltage at 300 A/m

2 is 2.10 volts, the current

efficiency is 96%, and energy consumption is 570 kWh/tonne Pb cathode. Lead produced is99.99% and the overall recovery is 96.1%. Zinc is partially solubilized in the leaching step but

not electrodeposited; it can be recovered from the bleed solution.

Fluoroboric acid, HBF4, is prepared industrially by reacting hydrofluoric acid with boric acid:

H3BO3+ 4HFHBF4+ 3H2O

The reaction is exothermic and the acid is available only as a 48% solution. Ferric fluoroborate isprepared by reacting the acid with Fe2O3. Fluoroboric acid is more expensive than fluorosilicicacid but it has the advantage of being more stable on heating and its solutions have higherelectrical conductivity. Fluoroboric acid was used for more than fifty year by NorddeutscheAffinerie in Germany for refining lead bullion with high bismuth content. Presently it is used in

electrodeposition of lead and its alloys, especially in printed circuit board manufacture.

Doe Run Company [former St. Joe Lead] had a demonstration plant in southeast Missouri,running for more than 3 years. Now it is in stand-by. (Figures 13 and 14). It is operating the cellat 8000 A/m2with an efficiency of 80 85 % and a current consumption between 900 and 1000kWh/t of lead, and lead recovery is about 99 %. In the same location in Missouri there is a smallunit (50 kg/d) that is available for testing. It is expected to invest more than $150 million.


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Figure 13 - Doe Run pilot plant in Missouri for electrolysis of lead in fluoroborate system

Figure 14 - Doe Run pilot plant in Missouri for electrolysis of lead in fluoroborate system

CONCLUSIONS

There have been many hydrometallurgical methods proposed for treating lead sulfide

concentrates. It seems that the best method is by leaching in fluoroboric acid containing ferricfluoroborate at 80

oC liberating elemental sulfur while silver remains in the residue. The solution

is then electrolyzed in a diaphragm cell where pure compact lead is deposited at the cathode andthe ferrous fluoroborate is oxidized at the anode to ferric fluoroborate for recycle. Any silverpresent in the concentrate can then be recovered from the residue after flotation of elementalsulfur. The recent work by Asarco engineers [51] based on granulating the lead bullion obtainedby smelting, leaching the granules in ferric fluoroborate, purification of solution, thenelectrowinning using a diaphragm cell is not justified because pollution problems during


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smelting have not been solved (Figure 15). The problem of lead metallurgy should be attackedfrom the very beginning, starting from the concentrate.

Pb concentrate

Smelting SO2

Granulation

Fe(BF4)3

Leaching

Purification Impurities

Electrowinning

Pb

Bullion

Figure 15 Schematic diagram for the proposed process by Asarco engineers [51] based on granulating

the lead bullion obtained by smelting, leaching the granules in ferric fluoroborate, purification of solution,then electrowinning using a diaphragm cell

Zinc had a similar history like lead up to 1980s. It was produced exclusively by roasting thesulfide concentrate to form zinc oxide, thermal reduction of the oxide, then refining the crudemetal by vacuum distillation. During World War I, leaching of the oxide and electrowinning ofzinc from the purified solution replaced the reduction and vacuum distillation steps. In 1980, thetotal hydrometallurgical route, i.e., aqueous oxidation of the sulfide concentrate to get zincsulfate solution and elemental sulfur was introduced on industrial scale. Copper had also asimilar history until at the beginning of the twenty first century when the aqueous oxidation ofcopper sulfide concentrates was also introduced on industrial scale by Phelps-Dodge in Arizona.Will lead follow a similar situation? The present writer is convinced that the answer is yes.

REFERENCES

[1] O. C. Ralston, Hydrometallurgy of Lead, Trans. AIME70, 447-466 (1924), Discussions pp.466-470

[2] R. G. Bowman, Lead, pp. 144-215 in Handbook of Nonferrous Metals, edited by D. M.Liddell, McGraw-Hill, New York 1945

[3] G. D. Van Arsdale,Hydrometallurgy of Base Metals, McGraw-Hill, New York 1953

[4] G. Leuchs,Ber. Deutsche Chem. Ges. 20, III, 152 (1887), German Patent 38, 193 (1886)

[5] G. Betts, Electrolytic Refining of Lead, 1902


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[6] A.T. Kuhn, editor, The Electrometallurgy of Lead, Academic Press, New York 1979

[7] J. E. McKay, Lead, pp. 207-267 in Kirk-Othmer Encyclopedia of Chemical Technology,Wiley, New York 1966

[8] Y.Awakura, S. Kamei, and H. Majima,A Kinetic Study of Non-oxidative Dissolution ofGalena in Aqueous Acid Solutions,Metall. Trans. 11B, 377-381(1980)

[9] G. L. Pashkov, E.V. Mikhlina, A.G.Khlmogorov, and Y.L. Mikhlin,Effect of Potential andFerric Ions on Lead Sulfide Dissolution in Nitric Acid,Hydrometallurgy63, 171-179 (2002)

[10] P.D.S. Scott and M.J. Nicol, Kinetics of Non-oxidative Dissolution of Galena in AcidicChloride Solutions. Trans. Inst. Min. & Met.85C, 40-44 (1976)

[11] Vizsoly, H. Veltman and F. A. Forward, Aqueous Oxidation of Galena in Acid Media

under Pressure, Trans. Am. Inst. Min. Engrs227, 215-220 (1963)

[12] F. Oprea and P. Moldovan, Le mcanisme et la cintique du lessivage des sulfures de zincet de plomb dans lautoclave sous pression doxygne, Revue Roum. Sci. Techn. Mtall.16,(2), 129-40 (1971)

[13] R. A. Agracheva, and A. N. Volskii, Treatment of Lead Sulfide Concentrates by FerricChloride Solution, Sb. Nauch. Trud. Mosk. Inst. Tsvet. Metall. Zolota 33, 26-33 (1960)(Russian text)

[14] R. J. Murray, The Dissolution of Galena Concentrates in Aqueous Ferric ChlorideSolutions, Ph. D. Thesis, University of Idaho, Moscow, Idaho 1972

[15] R. A. Agracheva, A. N. Volskii and A. M. Egorov, Treatment of Lead SulfideConcentrates by the Application of Ferric Chloride Solutions, Metallurgiya Topl. 3, 37-46(1959) (Russian text)

[16] J.E. Dutrizac, The Dissolution of Galena in Ferric Chloride Media, Met. Trans. 17B, 5-17(1986)

[17] F. P. Haver, K. Uchida and M. M. Wong, Recovery of Lead and Sulfur From GalenaConcentrate Using a Ferric Sulfate Leach, US Bureau of MinesRep. Invest. 7360 (1970)

[18] J. E. Murphy, F. P. Haver and M. M. Wong, Recovery of Lead From Galena by a LeachElectrolysis Procedure, US Bureau of Mines,Rep. Invest. 7913 (1974)

[19] F. P. Haver et al., Recovery of Lead from Lead Chloride by Fused-Salt Electrolysis, USBureau of MinesRep. Invest. 8166 (1976)


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[20] F. P. Haver, and M. M. Wong, Ferric Chloride-Brine Leaching of Galena Concentrate, USBureau of MinesRep. Invest. 8105 (1976)

[21] F. P. Haver, D. L. Bixby, and M. M. Wong, Aqueous Electrolysis of Lead Chloride, USBureau of Mines,Report No. 703-103/5 (1978)

[22] M. M. Wong, F. P. Haver and R. G. Sandberg, Ferric Chloride Leach ElectrolysisProcess for Production of Lead, pp. 445-54, in Cigan J. M., Mackey T. S. and OKeefe T. J.,eds.Lead-Zinc-Tin 80, The Metallurgical Society of AIME, Warrendale, Pennsylvania 1980

[23] F. Umar and Irawadi ,Double-stage Chloride Leaching of Lead Sulfide Concentrate, pp.91-96 in MMIJ/IMM Joint Symposium, Kyoto 1989, published by the Institution of Mining &Metallurgy, London: To-days Technology for the Mining and Metallurgical Industries

[24] J.M. Demarthe and A. Georgeau, Hydrometallurgical Treatment of Lead Concentrates,pp. 426-444 in Cigan J. M., Mackey T. S. and OKeefe T. J., eds. Lead-Zinc-Tin 80, The

Metallurgical Society of AIME, Warrendale, Pennsylvania 1980

[25] D. P. Seraphim, Oxidation of Galena in Ammonium Acetate Solutions under OxygenPressure,M. Sc. thesis, University of British Columbia, Vancouver 1952

[26] D. P. Seraphim and C. S. Samis, Kinetics of Oxidation of Galena in Ammonium AcetateSolutions under Oxygen Pressure, Trans. Am. Inst. Min. Engrs, 206, 1956, 1096-9.

[27] A. Vizsolyi, H. Veltman and F. A. Forward, Aqueous Oxidation of Galena under Pressurein Ammonia Solutions, pp. 326-344 in Unit Processes in Hydrometallurgy, edited by M. E.Wadsworth and F. T. Davis, Gordon & Breach, New York 1964

[28] J. E. Anderson, J. Halpern and C. S. Samis, Kinetics of the Oxidation of Galena in SodiumHydroxide Solutions under Oxygen Pressure, Trans. Am. Inst. Min. Engrs. 197, 554-558(1953)

[29] F. A. Forward, H. Veltman and A. Vizsoly, Production of High Purity Lead by AmineLeaching, pp. 823-837 inInternational Mineral Processing Congress, Institution of Mining& Metallurgy, London 1960

[30] M.I. Sherman and J. D. H. Strickland, Dissolution of Lead Sulfide Ores in Acid ChlorineSolutions, Trans. Am. Inst. Min. Engrs, 211,795-800 (1957)

[31] K.-Y. Lu and C.-Y. Chen, Conversion of Galena to Lead Carbonate in AmmoniumCarbonate Solution,Hydrometallurgy17, 73-83 (1986)

[32] De-Yu Liu, Carbonation Conversion. Hydrometallurgical Extraction of Lead from LeadSulfide Concentrates Kuangye Gongcheng6 (3), 49-53 (1986)

[33] Ya-Jun Gong, Qian Gong and Chia-Yung Chen, Study on Effect of Sulfide Impurities on


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Conversion of Lead Sulfide into Lead Carbonate in Ammonium Carbonate Solution,Nonferrous Metals(Beijing) 39 (4), 64-39 (1987)

[34] F. Habashi, Nitric Acid in Hydrometallurgy, pp. 357-364 in Extraction & ProcessingDivision 1999edited by B. Mishra, TMS-AIME, Warrendale, Pennsylvania 1999

[35] J. Fande, Leaching Sulfidic Ores (by Ferric Nitrate), German Patent680 518 (August 10,1939)

[36] M. C. Fuerstenau et al., The Kinetics of Leaching Galena with Ferric Nitrate, Metall.Trans.B 18, 25-30 (1989)

[37] G. L. Pashkov, A. G. Kholmogorov, E. V. Mikhlina, Yu. L. Mikhlin, Fundamentals andTechnology of Leaching of Gorevski Lead Concentrate in Nitric Acid Solutions, pp.276-277;S. V. Ostanova, G. L. Pashkov, V. V. Patrushev, S. V. Drozdov, A. V. Chubarov,Electrowinning of Lead from Nitric Acid Media, pp.282-283, inMetallurgy of Nonferrous

and Rare Metals, 2

nd

International Conference, volume 2, Krasnoyarsk, Russia 2003. Also,G.L.Pashkov et al, Method for Hydrometallurgical Treatment of Lead-ContainingConcentrates, Russian Federation Patent 2 002 117 528 (February 10, 2004)

[38] G. Van Weert and Y. Shang, Iron Control in Nitrate Hydrometallurgy by (Auto)Decomposition of Iron (II) Nitrate,Hydrometallurgy 33, 255-271 (1993)

[39] G. Van Weert and Y. Shang, Iron Control in Nitrate Hydrometallurgy by AutoclavingHydrolysis of Iron (III) Nitrate,Hydrometallurgy33, 273-290 (1993)

40 J. C. Fanning, The Chemical Reduction of Nitrate in Aqueous Solution, Coordination

Chem. Rev.199, 159-179 (2000)

[41] A.Y. Lee et al., Hydrometallurgical Process for Producing Lead and Elemental Sulfur fromGalena Concentrates, US Bureau of Mines,Rept. Invest.9055 (1986)

[42] A.Y. Lee et al., Pressure Leaching of Galena Concentrates to Recover Lead Metal andElemental Sulfur, US Bureau of Mines,Rept. Invest.9314 (1990)

[43] P.R.Taylor et al.,Fluosilicic Acid Leaching of Galena, pp. 461-473 in Hydrometallurgy2003, edited by C.A. Young et al., The Minerals, Metals, and Materials Society, Warrendale,Pennsylvania 2003

[44] M. Olper and M. Maccagni, The Production of Electrolytic Lead and Elementary Sulfurfrom Lead Sulfide Concentrates, pp. 1147-1167 International Symposium onHydrometallurgy, edited by J.B. Hiskey and G.W. Warren. TMS-AIME 1993

[45] M. Olper, Fluoborate Technology. A New Challenging Way for Primary and SecondaryLead Processing, pp. 185-198 in Zinc and Lead Processingedited by J.E. Dutrizac, J.A.Gonzalez, G.L. Bolton, and P. Hancock , Canadian Institute of Mining, Metallurgy, and


18/18

Petroleum, Montreal 1998

[46] M. Olper and M. Maccagni, Atmospheric Leaching and Electrowinning of Lead fromPrimary Sulfide Concentrate. The Flubor Technology, pp. 259-269 inEuropean MetallurgyConference, Friedrichshafen, Germany , published by GDMB 2001

47F. Habashi, The Recovery of Elemental Sulfur from Sulfide Ores, Bulletin 51, MontanaBureau of Mines and Geology, Butte, Montana 1966

[48] A. G. Starliper and H. Kenworthy, Recovery of Lead and Sulfur by Combined Chlorinationand Electrolysis of Galena,US Bureau of Mines,Rep. Invest. 6554 (1964)

49 J. E. Reynolds, D. N. Goens, and C. W. Kenney, Pilot Plant Development of ChlorideProcess for Lead-Zinc Concentrates, pp. 301-325 inLead-Zinc Update, edited by D. O.Rausch, F. M. Stephens, Jr., and B. C. Mariacher, Society of Mining Engineers AIME,New York 1977

[50] R.T. Um, W.K. Tolley, and J.C. Stauter, The Recovery of Lead from Sulfide ConcentratesUsing a Chlorination/Brine Leach/Electrolysis Process, pp.109-113 in ProcessFundamentals of Selected Hydrometallurgical Systems, edited by M.C. Kuhn, Society ofMining Engineers-AIME, New York 1981

[51] F. Ojebuoboh, S. Wang and M. Maccagni, Refining Primary Lead by Granulation-Leaching-Electrowinning,JOM 55(4), 19-23 (2003)

hydro of lead

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