How to Produce Pure Lead from Galena Concentrates

How to Produce Pure Lead from Galena Concentrates

A major cost factor in the sintering and smelting process for producing Pb is the control needed to meet existing environmental standards for Pb emissions. Another issue is the current concern over acid rain, which will in all probability result in even more stringent controls on emission of S gases.

The method consists of (1) oxidative leaching with H2O2, PbO2, and waste fluosilicic acid at 95° C to produce a solution of PbSiF6 and a residue containing elemental S, (2) electrowinning the PbSiF6 solution at 35° C to produce 99.99 pct Pb metal and H2SiF6, and (3) solvent extraction to recover S, leaving a residue containing Cu, Ag, and other metal values.

Researchers at the Bureau of Mines have developed an effective low-temperature leaching-electrowinning method to produce Pb metal and elemental S from galena concentrates. The method reduces Pb emissions and totally eliminates the formation of S gases. The elemental S produced is more economical to store and ship than the sulfuric acid (H2SO4) generated by the high-temperature smelting process.

This hydrometallurgical method consists of leaching galena concentrates in waste fluosilicic acid (H2SiF6) with hydrogen peroxide (H2O2) and lead dioxide (PbO2) as oxidants:

2PbS + 3H2SiF6 + H2O2 + PbO2 → 3PbSiF6 + 4H2O + 2S°………………………………(A)

After solid-liquid separation to obtain the Pb fluosilicate solution and the residue containing elemental S, electrowinning of the aqueous PbSiF6 solution at 35° C produces Pb metal and H2SiF6:

PbSiF6 + 2H+ + 2e- → Pb + H2SiF6……………………………………………………..(B)

The H2SiF6 is then recycled to leach more PbS. Sulfur is recovered from the leach residue by solvent extraction.
Several galena leaching processes have been investigated, including processes using ferric chloride, ferric sulfate, nitric acid, and ammonium acetate solutions. The leached PbCl2 and PbSO4 salts have a very limited solubility in aqueous solution, making aqueous electrolysis difficult. Lead metal was recoverable from PbCl2 by molten-salt electrolysis operated at 450° C. It is known that electrowinning of Pb in HNO3 and H2SiF6 solutions yields Pb metal at the cathodes and at the same time PbO2 at the anodes.

Recent investigations on Pb recovery from battery sludge have shown that small amounts of a phosphate or an arsenic compound in the H2SiF6 electrolyte will prevent the formation of PbO2 on anodes, making Pb electrowinning a viable process.

Figure 1 is the flow diagram of the oxidative leaching-electrowinning process. The parameters for each unit process were investigated in bench-scale experiments and tested on a larger scale. The unit process included oxidative leaching, lead electrowinning, activated carbon treatment of spent electrolyte, and S removal by solvent extraction.

hydrometallurgical-process flow diagram

Materials Equipment and Procedures

Oxidative Leaching

Galena concentrates, obtained several times from a Missouri Pb smelter, were analyzed and found to contain 74.4 to 78.3 pct Pb depending on the acquisition date. The results of chemical and screen analysis of typical samples are given in tables 1 and 2, respectively. Sample 1 was used in bench-scale tests and sample 2 in large-scale tests.

An ore mineralization study showed that the galena was present largely as free particles, and only small amounts were found in association with chalcopyrite, quartz, pyrite, dolomite, and sphalerite.

Initially, technical-grade H2SiF6 was used to leach the PbS concentrates.

hydrometallurgical-process analyses of galena concentrate

Later, waste H2SiF6 and recycled spent electrolyte (H2SiF6 + PbSiF6) were employed. The waste H2SiF6 was a byproduct of phosphate fertilizer manufacturing. The oxidants used were hydrogen peroxide (35 pct H2O2 ) solution and PbO2 powder (86.1 pct Pb), both technical grade.

A redox meter with a Pt electrode and an Ag-AgCl reference electrode were used to measure the redox potential of the oxidative leaching reaction.

The amounts of galena concentrates and H2SiF6 acid used in each leach experiment were calculated so that the resulting leachate would contain about 90 g/L Pb and 90 g/L free H2SiF6.

The leach procedure, for bench-scale tests, was to mix 100 mL H2SiF6 (200 g/L) solution with 2 mL H2O2 and 0.5 g PbO2 in a 250-mL tall form beaker, followed by the addition of 20 g PbS concentrate. As the mixture was heated to the desired temperature, H2O2 was slowly added to initiate the leach reaction. Subsequently, PbO2 powder was added to maintain the redox potential. When the redox potential of the leach mixture dropped below 480 mV, more PbO2 was added to raise it to at least 480 mV, but not higher than 550 mV. The leach reaction was complete when the redox potential remained above 480 mV. The mixture was then filtered to separate the residue from the leachate.

Leach parameters, such as temperature, time, concentration of acid, and amounts of oxidants used, were investigated in

hydrometallurgical-process screen analysis of galena concentrate

bench-scale experiments using 20- to 200-g samples of PbS.

Large-scale leaching, using 10 kg PbS and 75 L H2SiF6 solution, was conducted in a 100-L reactor described previously.

Lead Electrowinning

Electrolyte was prepared from the PbS leachates, which contained 80 to 100 g/L Pb as PbSiF6 and 90 to 100 g/L free H2SiF6. Prior to electrolysis, glue and calcium lignin sulfonate (lignosite) were added as leveling agent and grain refiner, respectively, and H3PO4 was added to prevent excess PbO2 formation on the anodes.

Lead electrowinning experiments were conducted in a 1-L cell for 8 h using two PbO2-coated titanium anodes and one pure Pb sheet cathode. The electrolyte temperature was 35° C, and the current density was varied from 180 to 250 A/m².

Large-scale electrowinning tests were performed in a 20-L cell for 5-day runs; 200 L of electrolyte was stored in a tank and circulated through the cell as described in RI 8857. The cell had four PbO2-Ti anodes (10 by 26 by 0.4 cm) and three Pb cathodes (12 by 28 by 0.16 cm).

Recycling of Electrolyte

Lignosite and glue were removed from spent electrolyte by mixing the electrolyte with activated C (minus 325 mesh) in a beaker at ambient temperature for 30 min and then filtering to obtain the clear H2SiF6 solution and C residue. The amount of C used was 12.5 g/L of spent electrolyte for bench-scale tests, or 1.25 kg/100 L for large-scale treatment in the 100-L reactor.

Sulfur Extraction

The typical galena leach residue contained about 40 to 50 pct elemental S. Solvent extraction was investigated for recovering S from the residue.

The apparatus used for recovering S by solvent extraction consisted of a 2-L glass filter reactor heated with an electric mantle (fig. 2). The reactor head was equipped with three necks to accomodate a thermocouple, a glass stirrer, and a Graham condenser to recover solvent vapor. The bottom of the reactor was sealed with a fritted glass disc to retain particles larger than 10 µm, and a stopcock opening was used for solvent drainage. The solvent was removed from the residue by vacuum filtering. A flask was used to collect the hot S-carrying solvent. Two cold traps were used in the vacuum line to collect solvent fumes during the hot filtration.

The S extraction procedure was to slurry 100 g of dried residue with 800 mL of mixed solvents of tetrachloroethylene and kerosene in the glass reactor, heat to 110° to 118° C, and hold for 5 min before filtering. This arrangement facilitated rapid removal of the hot S-carrying solvent and prevented solidification of the S during filtering. After cooling to ambient temperature to allow the S to crystallize out of solution, S was filtered from the solvent, which was recycled for S extraction. The recovered S was dried at 90° C to remove any residual solvent.

Results and Discussion

The chemical equations for PbS leaching in acid solution with and without oxidants [O] follow:

PbS + 2H+ + [O] → Pb²+ + S° + H2O………………………………………………………….(C)

PbS + 8H+ + 8[O] → PbSO4 + 4H2O………………………………………………………….(D)

PbS + 2H+ → Pb²+ + H2S………………………………………………………………………….(E)

Reaction C shows that oxidative leaching of PbS will yield Pb salt and elemental S. Reaction D suggests PbSO4 may form if the redox potential of the solution is too high. Reaction E indicates H2S will form when leaching in acid solution if the redox potential is too low.

hydrometallurgical-process sulfur extraction apparatus

The leach procedure was developed based on the above information. To avoid the generation of H2S, one-fourth of the required oxidant was added to the H2SiF6 solution prior to the addition of PbS. During leaching, the redox potential in the leach mixture was controlled between 480 and 500 mV to prevent oxidizing the S into PbSO4 as in reaction D. The reaction is exothermic, and it is necessary to add H2O2 slowly through a burette to avoid overheating the leach solution. After adding the H2O2, PbO2 was added slowly to control the redox potential above 480 mV until the reaction was complete.

The reactions occurring during the oxidative leaching of PbS concentrates with H2SiF6 solution follow:

PbS + H2O2 + H2SiF6 → PbSiF6 + 2H2O+ S°……………………………………………………………..(F)

PbS + PbO2 + H2SiF6 → PbSiF6 + PbO + HgO + S°……………………………………………………..(G)

The PbO reacts with additional H2SiF6 :

PbO + H2SiF6 → PbSiF6 + H2O………………………………………………………………………………..(H)

At the end of leaching, the mixture was filtered to separate the leachate from the residue. The residue consisted of elemental S and metal values. The leachate was sent to electrowinning to recover pure Pb metal.

Leaching Parameters—Bench-Scale Tests

Oxidants

Lead extraction was 98 pct on the minus 400-mesh PbS samples and 96 pct on the as-received concentrates if H2O2 and PbO2 were used as oxidants. Other oxidants, such as air, oxygen, ozone, HNO3, and MnO2, used alone or in combinations with H2O2, resulted in less than 80-pct Pb extraction during screening tests.

The effect of using different combinations of H2O2 and PbO2 on PbS leaching was insignificant, as shown in table 3. Leaching experiments showed that H2O2 was a more efficient oxidizer to initiate the leach reaction; also, it was less expensive than PbO2 ($0.51/kg versus $1.54/kg, as listed in the Chemical Marketing Reporter, April 28, 1986). Thus, it is beneficial to use H2O2 to leach PbS and only use PbO2 at the end of the leach to avoid oxidizing PbS into PbSO4.

hydrometallurgical-process-lead-extraction

Temperature

Leaching temperatures had a great influence on reaction rate and Pb extraction, as shown in table 4. When leaching below 80° C, the reaction rate was thought to be too slow for any practical application. Lead extraction was 96 pct when leaching at 95° C for 35 min using H2O2 and PbO2 as oxidants. The leaching rate increased greatly, and the required leaching time was reduced from 90 to 35 min as the temperature increased from 90° to 95° C.

Time

Table 5 shows that Pb extraction was increased from 92 to 96 pct as leaching time increased from 30 to 60 min at 95° C.

Initial leaching was rapid as shown in figure 3, but as the elemental S formed

hydrometallurgical-process-effect-of-time

hydrometallurgical-process-effect-of-leach-time

hydrometallurgical-process dissolution of pb

hydrometallurgical-process dissolution of impurities

and coated the PbS particles, further reaction was probably diffusion controlled and the leach rate was reduced. However, the effect of the S coating was not critical, because of the fine particle size of the PbS.

The dissolution of impurities (at 90° C) increased with leaching time as shown in figure 4.

Fluosilicic Acid

The amounts of PbS, PbO2 , and H2SiF6 used in a leach test determined the concentration of PbSiF6 and free H2SiF6 in the pregnant leachate. Increasing the concentration of free H2SiF6 above 56 g/L had no significant effect on Pb extraction, as shown in table 6; extraction of impurities decreased with decreasing concentration of free H2SiF6.

Lead extractions of 96, 91, and 96 pct were achieved using H2SiF6 solutions made from technical-grade, waste, and recycled acid, respectively, under similar conditions. Waste H2SiF6 contained HCl (1.7 g/L) and H2SO4 (0.64 g/L) as impurities, which formed some insoluble Pb salts during leaching, resulting in lower Pb extraction. Recycled electrolyte, in which impurities were removed during prior leaching, was as reactive as technical- grade H2SiF6.

Large-Scale Leaching Tests

Results of large-scale leaching experiments are given in table 7. Lead recoveries of 93.6, 91.1, and 95.6 pct were achieved when leaching with technical, waste, and recycled H2SiF6, respectively. The results are in good agreement with the bench-scale tests, except longer times were required. The consumption of oxidants in the large-scale leach was about 30 pct less than in the bench-scale tests. Hydrogen peroxide seemed to react more efficiently in a large reactor, probably because it was introduced into the bottom of the reactor through a capillary tube.

hydrometallurgical-process concentration

hydrometallurgical-process large-scale leaching tests

Electrowinning

Bench Scale (1-L Cell)

Lead cathode deposits of 99.9- to 99.99-pct purity were obtained from electrowinning the electrolyte prepared from the leach solution at 35° C and 200 A/m² for 8 h. The purity of electrowon Pb depended on the amount of Cu²+ ions in the electrolyte, as shown in table 8.

Table 9 gives the results of 10 consecive recyclings of electrolyte in bench-scale leach-electrowinning experiments.

hydrometallurgical-process lead deposits

hydrometallurgical-process results

The success of these repeated leaching-electrowinning tests demonstrated the feasibility of the process. Average current efficiency and energy consumption were 86 pct and 0.72 kW·h/kg, respectively.

The effects of the electrolyte concentration and current density on Pb electrowinning were studied in an artificial electrolyte made from technical H2SiF6 and PbO to avoid any complication. Concentrations of 60 to 160 g/L Pb and 40 to 140 g/L free H2SiF6 had little effect on current efficiency and energy consumption, which ranged from 93 to 97 pct and 0.62 to 0.64 kWh/kg, respectively. Table 10 gives the effect of current density; the current efficiency, energy consumption, and Pb production rate all increased with current density increasing from 150 to 300 A/m².

Large Scale (20-L Cell)

The 200-L electrolyte used for the 20-L cell test was prepared from solutions from large-scale leaching experiments. The solution contained higher amounts of Cu²+ (0.04 to 0.09 g/L) than the bench-scale leachates because a longer leaching time was required. The resulting electrowon Pb contained 99.96 Pb, 0.04 Cu, and <0.001 Zn and Fe. Table 11 gives the analysis of electrolyte (before and after use), the parameters used, and the results of electrowinning.

Impurities

The major impurities in the PbS leach liquors are ions of Ca, Mg, Fe, Zn, Al, and Cu with trace amounts of Co and Ni. Results of tests showed Al, Ca, Mg, and

hydrometallurgical-process-effect-of-current-density

hydrometallurgical-process electrowinning tests

Zn had no adverse effect on Pb electrowinning, except for Zn when its concentration was greater than 10 g/L. Copper ions tend to codeposit with the Pb on the cathode, especially at current densities greater than 200 A/m². The presence of ferric ions in the electrolyte reduces the current efficiency as plotted in figure 5. For example, as the concentration of ferric ions increases from 0 to 1 and then to 2 g/L, the current efficiency decreases from 95 to about 87 and then to less than 80 pct. The effect of trace amounts of Ni and Co was not studied.

Table 9 gives the results of 10 consecutive bench-scale leach, electrowinning experiments using recycled electrolyte with C treatment indicated that the buildup of major impurities could be controlled by controlling the leach parameters, e.g., by reducing the leaching time and lowering the free H2SiF6 concentration. As shown in figure 6, the concentration of major impurities reached an equilibrium value after the fourth recycling of, in g/L, 2.67 Ca, 1.35 Mg, 0.88 Fe, 0.72 Zn, 0.38 Al, 0.008 Cu. Current efficiency and energy consumption for the 4th to 10th recycling were all close to 86 pet and 0.72 kW·h/kg, respectively, as compared to 95 pct and 0.64 kW·h/kg for single-cycle electrolyte.

Sulfur

Elemental S was separated from the leach residues by treating with a mixture of tetrachloroethylene (TCE) and kerosene at 115° C. There was 99-pct removal of the elemental S and over 90-pct recovery. Approximately 9 pct of the S did not crystallize out of solution upon cooling of the solvent. The average solvent recovery was about 92 pct after each

hydrometallurgical-process effect of ferric ions

hydrometallurgical-process impurity buildup

extraction, with most solvent loss occurring during filtration. The solvent was subsequently lost in the drying of elemental S. Fresh mixed solvent was added to the used to make up the required volume before recycling. The major impurity in the solvent was Pb at 0.005 g/L after the 10th recycling.

Less than 50 pct of the S was recovered from fresh solvent, but S recoveries averaged over 90 pct by the second recycling. One percent S° was left in the residue after treating the residue with solvent that was saturated with S° at ambient temperature. Table 12 details the effects of different ratios of TCE and kerosene on S extraction and indicates that the 1:1 ratio gives the highest recovery of S.

Wet residue, as received from the H2SiF6 leach, was heated in a 1:1 ratio TCE-kerosene. A heavy froth developed during heating, and an azeotropic mixture formed at 99.2° C. This temperature was too low for effective S extraction, and attempts to raise it only resulted in loss of solvent. Pure wet S under the same conditions did not produce the froth, and 113° C was easily obtained. The reason for the different reactions was not determined, but it was decided to use dried residue in all tests.

Residue

As shown in table 13, removing the elemental S upgraded the H2SiF6 leach

hydrometallurgical-process-sulfur-recovery

hydrometallurgical-process comparison of galena concentrate

residue to 11 tr oz/mt Ag and 50 pct Pb, PbS, and PbSO4 compounds.

The S-free residue was subjected to a three-step releach consisting of (1) oxidative leach with H2SiF6 to form PbSiF6 and S°, (2) carbonation leach to convert PbSO4 into PbCO3 (2), and (3) H2SiF6 leach to dissolve PbCO3. As a result of releaching, the residue was upgraded to contain 26 tr oz/mt Ag and, in pct, 12.2 Zn, 7.6 Cu, 1.26 Ni, and 0.9 Co, while Pb was reduced from 50 to 5.4 pct.

This final residue became a marketable raw material. The analyses of each residue were given in table 13 to compare with the 1,000-g-PbS-concentrates analysis.

Conclusions

A process has been developed to produce Pb and elemental S of 99.99-pct purity from galena concentrates by hydrometallurgical and electrometallurgical methods. This process eliminates S gases and Pb emissions. The elemental S produced is easier to transport and store than is the H2SO4 generated by the pyrometallurgical method.

The unit processes investigated and optimized by bench-scale testing included oxidative leaching of PbS with H2SiF6, electrowinning the leach solution to produce high-purity Pb metal, carbon treatment of spent electrolyte for recycling, and S removal from the leach residue by solvent extraction. The first three processes were also tested in larger scale experiments to demonstrate that they were technically feasible.

Pressure leaching of PbS concentrates in H2SiF6 acid under oxygen pressure is being investigated to reduce the need for H2O2 and PbO2 as oxidants. Methods to remove ferric ions from electrolyte are being studied to reduce the energy consumption for electrowinning.