B_Articulos de Consulta

CURSO DE ACTUALIZACION

INNOVACIONES METALÚRGICAS EN LA RECUPERACION DE ORO Y PLATA.

Avances en los métodos de recuperación de oro y plata de minerales refractarios.

Pressure cyanide leaching for precious metals recovery.

Detoxification of cyanide using titanium dioxide and hydrocyclone sparger with chlorine dioxide.

Estudio cinético de la extracción de oro de un mineral piritoso mediante oxidación y lixiviación simultánea a presión.

Continuous laboratory gold solvent extraction from cyanide solutions using LIX 79 reagent.

Enhance cyanide recovery by using air-sparged hydrocyclone.

Estudio de la resina Aurix 100 de intercambio iónico para recuperar oro en las soluciones cianuradas.

Extracción de oro por solventes en un circuito continúo.

Gold solvent extraction from alkaline cyanide solutions, using LIX 79 extractant.

New technology for recovery of gold and silver by pressure cyanidation leaching and electrocoagulation.

Uso de resina de intercambio iónico para la recuperación del complejo oro tiosulfato desde soluciones acuosas.

Tostación de un concentrado refractario de oro y plata.

Zinc-dust cementation of silver from alkaline cyanide solutions- analysis of Merril-Crowe plant data.

Kinetic aspect of gold and silver recovery in cementation with zinc power and electrocoagulation iron process.

Recovery of silver and gold from cyanide solution by magnetic species formed in the electrocoagulation process.

.

.

2007 October • JOM 43

Research SummaryAqueous Processing

A novel method demonstrates that the

oxidation and dissolution of gold and

silver in alkaline cyanide solution can

be conducted simultaneously in the same

autoclave in less than 90 minutes with

a recovery that exceeds 96%. Because

mild operating conditions of 80°C and

0.6 MPa oxygen pressure are used in this

process, low cost materials of construc-

tion can be utilized for the autoclave.

INTRODUCTION

Durango is the second largest gold-silver producing state in México. The Bacís mine in particular produces a pyrite concentrate rich in silver and gold. During the mine’s short history, efforts have been made to reduce costs and increase production through improved process developments.1,2 At the present time, cyanidation has superseded all previous leaching processes, particularly chlorination,3 because of its ability to effectively and economically treat ores containing as little as 1–3 g/ton gold. The precious metals are dissolved in a dilute solution of sodium cyanide, which is recycled after precipitation of the pre-cious metals with zinc dust. A bleed of this barren solution is often required to control the accumulation of deleterious impurities. Refractory gold-silver ores contain precious metals locked up in a matrix of pyrite and/or arsenopyrite. Such ores are not amenable to cyanidation. To liberate the gold-silver from the sulfide matrix and render it accessible to cyanidation, it is a common practice to subject them to a preoxidation in order to enhance the recovery of the precious metals.4,5

Although effective, these techniques result in the production of large quanti-ties of environmentally hazardous sub-stances such as SO

2 or H

2SO

4, which can

cause difficulties in compliance with

Pressure Cyanide Leaching for Precious Metals Recovery

José R. Parga, Jesús L. Valenzuela, and Francisco Cepeda T.

environmental pollution regulations. After review of the literature,4–9 it is clear that of the refractory gold treatment plants installed during the past 25 years, pressure oxidation has been the most popular (see Table I). This process, which is used in many commercial plants around the world, involves the recovery of the precious metals by oxidation pretreatment followed by traditional cyanidation. It should be remembered that the pressure oxidation technology is not used to leach the gold and silver but only to make it accessible to the cyanide in the host mineral. See the sidebar for experimental pro-cedures.

PROCESS CHEMISTRY AND LEACHING EXPERIMENTS

Pyrite, arsenopyrite, sphalerite, and covellite are the most common host minerals of gold, silver, and electrum. It is important from a process optimization standpoint to understand the behavior of

each of these minerals during alkaline pressure oxidation. This process leads to dissolution/destruction and subse-quent liberation of gold and silver which would then be available for cyanidation. Pressure oxidation of pyrite involves reactions yielding ferrous ion, sulfate ion, and elemental sulfur as products.10,11

The primary reactions are shown in Equations 1 and 2. (All equations are presented in the Equations table on page 46.) Ferrous ions produced by Equations 1 and 2 are subsequently oxidized to ferric ions, as shown in Equation 3. The ferric ions can also contribute to the oxidation of silver iron sulfide, argentite, pyrite, pyrrhotite, sphalerite, and chalcopyrite (see Equations 4–9, respectively). The resulting elemental sulfur may be further oxidized to sulfate by oxygen or by ferric sulfate, as given in Equations 10 and 11. This results in the formation of a porous, but nonprotec-tive, elemental sulfur layer, thus allowing

Figure 1. A schematic of the mechanism of gold and silver leaching.

JOM • October 200744

Table A. Analytical Data for the Composite Samples

g/t Percent

Au Ag Pb Zn Cu Fe As S

Concentrate 87.09 12,320 2.6 3.8 0.5 29.2 0.15 32

Ore 4.12 289 0.5 0.4 0.04 3.67 0.1 3.7

EXPERIMENTAL PROCEDURES

Mineralogical analysis indicates that the sulfide minerals for the Bacís concentrate

include silver iron sulfide, silver sulfide or argentite, pyrite, pyrrhotite, arsenopyrite,

chalcopyrite, covellite, hematite, and magnetite. The non-opaque minerals are quartz,

calcite, apatite, gypsum, fluorite, and barite. Also, several photomicrographs indicate

occlusion and/or dissemination of micrometer-sized gold, silver, and electrum particles

in the sulfide minerals such as pyrite, sphalerite, arsenopyrite, and quartz (Figure A).

The compositions are listed in Table A.

Figure A. A photomicro-graph of argentite (ARG) occluded in pyrite, size 60 m.

Figure 3. A comparison of gold extraction at ambient conditions and high pressure. Conditions: % solids = 20; pH = 11.0; rpm = 300.

Figure 2. A comparison of silver extraction at ambient conditions and high pressure. Conditions: % solids = 20; pH = 11.0; rpm = 300.

cyanide and dissolved oxygen to access to the previously locked gold, silver, and electrum. It was thought that by main-taining high concentrations of CN– ions in the pressure leach reactor it would be possible to complex gold and silver (Equation 12) as they became liberated and thus achieve both objectives (i.e., decomposition of refractory minerals and Au/Ag recovery) in a single stage. Both gold and silver extractions achieved were, in fact, greater than 96%. Since silver in Mexican ores occurs as argen-tite the cyanidation and sulfide oxidation reactions are given in Equations 12 through 15. Sulfide ions are oxidized to thiosulfate and can contribute with the dissolution of gold. Also J.S. Graham et al.12 showed that an alkaline or near-neutral solution of thiosulfate dissolved gold metal slowly in the presence of a mild oxidant. The dissolution of the gold can be written as in Equation 16, where oxygen is the oxidant and thiosulfate is the ligand.

Better extractions of gold with thio-sulfate are achieved when elevated temperatures (e.g., 65°C) are used 13 and also, cupric ion has been found to have a strong catalytic effect on the rate of

oxidation. The proposed physiochemical aspects of the leaching reaction mecha-nism for gold and silver dissolution in relatively mild operating conditions of 80°C and 0.6 MPa oxygen pressure are illustrated in Figure 1. In this figure, the ore particle displays different diffusing ions. All experiments were conducted in a 4 L Parr autoclave assembled with impel-ler, thermowell, pressure gauge, gas inlet, and outlet pipes used for simultaneous oxidation and cyanidation. Samples were ground to 84% minus 40 m or finer, and pulped with fresh tap water. Depend-ing on test requirements, variables such as NaCN, lixiviation time, temperature, particle diameter, pH, and pulp density were set and the leaching experiments were undertaken.

RESULTS AND DISCUSSION

Preliminary batch testing had indi-cated that at ambient conditions (25°C, 1 atm.) direct cyanide leaching of the concentrate with air gave poor silver and gold recoveries. The results obtained are shown in Figures 2 and 3. Figure 4 shows that there is an opti-mum leach time to achieve both maxi-mum gold and silver extractions. The optimum leach time appears to be close to 60 min. at 80°C. Increasing time beyond this value leads to chemical degradation of the complex cyanide due to the hydrogen ions produced by the sulfide oxidation. Also, for prolonged leaching time there is a significant loss in cyanide, since the oxygen present


Table I. Existing Commercial Plants Using Pressure Oxidation4–8

Operation Data

Plant Location Feed Medium (t/d) T( C) P (MPa) t (min.)

McLaughlin, California Ore Acid 2,700 180 2.76 90

São Bento, Brazil Concentrate Acid 240 190 1.58 90

Barrick Mercur, Utah Ore Alkaline 790 225 3.26 70

Getchell Mine, Nevada Ore Acid 3,000 210 2.89 130

Goldstrike, Nevada Ore Acid 1,360 225 3.02 75


Porgera, Papua, Concentrate Acid 1,350 190 1.77 180

New Guinea

Campbell, Vancouver, Concentrate Acid 70 195 2.10 120

Barrick, Toronto Ore Acid 2,000 220 3.10 100


Santa Fe Pacific Gold, Ore Acid 4,000 225 3.17 50

Nevada

Sunshine Mining Co., Ore Acid 1,000 120 0.62 60

Idaho

Porgera, Papua, Concentrate Acid 2,700 190 1.72 150

New Guinea

oxidizes the cyanide to cyanate and ultimately, to ammonia and carbon dioxide. Sodium cyanide addition appears to be the critical variable for good recovery. Concentrations of 0.6–1.1 wt.% sodium cyanide were used to leach the concen-trate in the autoclave. When the concen-tration of sodium cyanide was less than 0.6%, gold and silver extraction could not be completed optimally in a single stage of autoclave leaching. Increasing the cyanide concentration resulted in an increase in the gold and silver extraction. The results are shown in Figure 5. The concentrate was leached as 20 wt.% solids slurry with 1 wt.% of NaCN for 60 min. Temperature was varied from 60°C to 200°C. As the temperature was increased, the gold and silver extractions decreased since at temperatures higher than 80 C, the oxidation of cyanide ions is too rapid. This results in the formation of CO

2 gas and ammonium ions (see

Equation 17). In this case, cyanide ions are not available for complexation, as shown in Figure 6. A pH range of 10.5–13 was examined for the extraction of gold and silver. In Figure 7, it can be seen that at a pH of 11.5 or above, optimum extractions could not be maintained. This is because in alkaline-oxidizing medium the pyrite and pyrrhotite in the concentrate compete with gold and silver for consumption of oxygen. This mechanism may be respon-sible for the decrease in extraction, as shown in Equations 18 and 19. Eventu-ally the ferrous sulfate formed is con-verted to stable, insoluble ferric hydrox-ide, as given in Equations 20 and 21. The decrease in the extraction of gold and silver with increasing pH may also be due to a decrease in the formation of ferric sulfate. Ferric sulfate is the reagent

that contributes to the oxidation of sulfide minerals which occlude gold and silver. Also, G.A. Deitz14 found that at pH values between 12 and 13, the silver surface became coated with a whitish film of CaO

2 and thus the leaching process was

retarded. The effect of varying the pulp density in the range of 15–30% solids is shown in Figure 8. Gold and silver extractions gradually decreased as the pulp density increased. This was apparently due to the increased apparent viscosity of the slurry, which impedes good oxygen dispersion in the system, and also due to the excessive acid generation at high pulp density. As seen, direct pressure leaching cyanidation is advantageous over numerous pressure oxidation plants for the pretreatment of gold/silver ores

and concentrates. It is faster, requires lower temperatures and pressures, and uses a stainless-steel autoclave. The optimum conditions as delineated in the laboratory study were: solid 20%; temperature 80°C; pressure 5.6 kg/cm2; rpm 300; cyanidation time 60 min.; NaCN 1%; pH 10.7; and PbO 100 g/t. These conditions were used to study the extraction of gold and silver during a continuous one-month plant campaign. A total of 600 t of concentrate was pro-cessed and recoveries for both gold and silver averaged approximately 96%. The plant data are shown in Figure 9. The results described by the pilot tests were in good agreement with the results obtained in the laboratory autoclave

Figure 6. The effect of temperature on gold and silver extraction. Conditions: % solids = 20; pH = 11.2; pressure 0.6 MPa; rpm = 300; leach time 60 min.

Figure 5. The effect of cyanide concentra-tion on gold and silver extraction. Condi-tions: % solids = 20; pH = 11.0; temp. 80°C: rpm = 300; pressure = 0.6 MPa; leach time 60 min.

Figure 4. The effect of autoclave retention time on gold and silver extraction. Conditions: % solids = 20; pH = 11.0; pressure = 0.6 MPa rpm = 300; temp. 80°C.

▲

▲▲

Capacity

JOM • October 200746

Equations

2FeS2 + 7O

2 + 2H

2O 2Fe2+ + 4SO

42– + 4H+ (1)

FeS2 + 2H+ Fe2+ + H

2S + S0 (2)

2Fe2+ + 1/2O2 + 2H+ 2Fe3+ + H

2O (3)

AgFe2S

3 + Fe3+ Ag+ + 3Fe2+ + 2S2– + S0 (4)

Ag2S + 2Fe3+ 2Ag+ + 2Fe2+ + S0 (5)

FeS2 + 8Fe3+ + 4H

2O 9Fe2+ + 8H+ + S0 + SO

42– (6)

Fe7S

8 + 32H

2O + 62Fe3+ 69Fe2+ + 64H+ + 8SO

42– (7)

ZnS + 2Fe3+ Zn2+ + 2Fe2+ + S0 (8)

CuFeS2 + 4Fe3+ Cu2+ + 5Fe2+ + 2S0 (9)

2S0 + 3O2 + 2H

2O 4H+ + 2SO

42– (10)

S0 + 6Fe3+ + 4H2O 6Fe2+ + 8H+ + SO

42– (11)

Ag2S + 4CN– 2Ag(CN)–

2 + S2– (12)

S2– + CN– + 1/2O2 + H

2O SCN– + 2OH (13)

2S2– + 2O2 + H

2O S

2O

32– + 2OH– (14)

S2O

32– + 2OH– + 2O

2 2SO

42– + H

2O (15)

4 Au + 8S2O

32– + O

2 + 2H

2O 4[Au(S

2O

3)2]3– + 4OH– (16)

CNO– + H2O + 2H+ CO

2 + NH

4+ (17)

2FeS2 + 7O

2 + 2H

2O 2Fe2+ + 4H+ + 4SO

42– (18)

FeS + 2O2 Fe2+ + SO

42– (19)

FeSO4 + Ca(OH)

2 Fe(OH)

2 + CaSO

4 (20)

2Fe(OH)2 + 1/2O

2 + H

2O 2Fe(OH)

3 (21)

Zn + 4CN– + 1/2O2 + H

2O Zn(CN)

42– + 2OH– (22)

2Au(CN)2– + Zn 2Au + Zn(CN)

42– (23)

Zn + 2H2O + 2Au(CN)

2– 2Au + HZnO

2– + 3H+ + 4CN– (24)

with low temperature and low pressure. Under these conditions, less expensive autoclaves can be used. K.G. Thomas found15 that a 1,400 t/d acid pressure

oxidation circuit might cost $30 mil-lion while the corresponding non-acid circuit might have a price of only $15 million.

RECOVERY OF PRECIOUS METALS

The cementation of gold and silver by zinc dust is an electrochemical process, proceeding by localized anodic and cathodic reactions. The main reactions for zinc dissolution are given in Equa-tions 22–24. The Merrill–Crowe process is pre-ferred for a very rich pregnant solution or solutions containing large amounts of silver.16 Powered zinc is added to the clarified, deaerated, pregnant cyanide solution. The zinc particles with gold and silver are removed from the solution using a plate and frame filter press. The mean chemical analysis of the pregnant solution from plant testing found 0.26 g/t CN; 91 ppm silver, and 0.9 ppm gold. For both gold and silver cyanide ions, the recovery by cementation was 99.5%.

CONCLUSIONS

The described study shows that gold and silver values are associated with silver iron sulfide, argentite, pyrite, pyr-rhotite, sphalerite, and chalcopyrite in the Bacís concentrate. The dissolution of gold and silver is due to the strong complexing capabilities of cyanide anions combined with the oxidizing properties of the dissolved molecular oxygen. The kinetics of the direct pres-sure oxidation/cyanidation was found to be strongly dependent on particle size, concentration of sodium cyanide, tem-perature, and pH. Single-stage direct pressure oxidation/cyanidation has proven to be effective in treating pyrite refractory gold and silver concentrates from Bacís mining. For both gold and silver it was found that the precious metals recovery exceeded 96%.

Figure 7. The effect of pH on gold and silver ex t r a c t i o n . Conditions: % solids = 20; pressure = 0.6 MPa; rpm = 300; leach time 60 min.; temp. = 80°C; NaCN = 1%.

Figure 8. The effect of per-cent solids in the extraction of gold and silver. Condi-tions: pressure 0.6 MPa; pH = 11.2; rpm = 300 ; leach time 60 min.; temp. = 80°C; NaCN = 1%.


The relatively mild operating condi-tions of 80°C and 0.6 MPa oxygen pres-sure offer distinct advantages. For example, low cost materials of construc-tion can be utilized for the autoclave. Finally in this process, there is obviously lower gold and silver inventory.

ACKNOWLEDGEMENTS

The author thanks the management

of Bacís Mining S.A. and CONACYT for

support and permission to publish this

paper. Appreciation is also extended to

Prof. Jan D. Miller of the University of

Utah for his interest in this research.

References

1. J.R. Parga and H. Mercado, “Precious Metals

Extraction by Direct Oxidative Pressure Cyanidation of Bacís Concentrates,” Proceedings Randol Gold Forum (Beaver Creek, OR: Randol International Ltd., 1993), pp. 209–212.2. J.R. Parga and H. Mercado, “Comparative Study of Direct Pressure Oxidation/Cyanidation and a Pyrometallurgical Process for the Recovery of Precious Metals from Sulfide Concentrates,” Proceeding of the 17th IPMI, Precious Metals, ed. Rajesh K. Mishra (Allentown, PA: International Precious Metals Institute, 1993), pp. 55–67.3. N. Geoffroy and F. Cardarelli, “A Method for Leaching or Dissolving Gold from Ores or Precious Metal Scrap,” JOM, 57 (8) (2005), pp. 47–50.4. Anonymous, “Refractory Gold Technology,” Mining Magazine (1996), pp. 231–234.5. G.L. Simons et al., “Pressure Oxidation Problems and Solutions: Treating Carbonaceous Gold Ores Containing Trace Amounts of Chlorine (Halogens),” Mining Engineering (1998), pp. 69–73.6. F. Habashi, “Copper Metallurgy at the Crossroads,” Proceedings of the Copper 95-Cobre 96 International Conference, ed. W.C. Cooper and D.B. Dreisinger (Montreal, Canada: CIM, 1995), pp. 493–510.

7. W. Yernberg, “Santa Fe Pacific Gold Targets Millions Ounces/Year Production,” Mining Engineering, 9 (1996), pp. 39–43.8. C.G. Anderson and S.M. Nordwick, “Pretreatment Using Alkaline Sulfide Leaching and Nitrogen Species Catalyzed Pressure Oxidation on a Refractory Gold Concentrate,” EPD Congress 1996, ed. G.W. Warren (Warrendale, PA: TMS, 1996), pp. 323–341.9. J. Chen and K. Huang, “A New Technique for Extraction of Platinum Group Metals by Pressure Cyanidation,” Hydrometallurgy, 82 (2006), pp. 164–171.10. S. Chander and A. Briceño, “Kinetics of Pyrite Oxidation,” Minerals and Metallurgical Processing, 8 (1987), pp. 170–176.11. R.G.M.S Berezowski and D.R. Weir, “Pressure Oxidation Pretreatment of Refractory Gold,” Mine Metallurgy Process (1994), pp. 1–4.12. J.S. Graham and T.W. James, “Cyanide and Other Lixiviant Leaching Systems for Gold with Some Practical Applications,” Mineral Processing and Extractive Metallurgy Review, 14 (1995), pp. 193–247.13. R.Y. Wan, M. Le Vier, and J.D. Miller, “Research and Development Activities for the Recovery of Gold from Non-Cyanide Solutions,” Hydrometallurgy Fundamentals, Technology & Innovations (Littleton, CO: Society for Mining, Metallurgy & Exploration, 1993), pp. 415–436.14. G.A. Deitz and J. Halpern, “Reaction of Silver with Aqueous Solutions of Cyanide and Oxygen,” Journal of Metals, 5 (1953), pp. 1109–1116.15. K.G. Thomas, D. Burns, and A.R. Hill, “Autoclaving Technology and Barrick Gold” (Paper prepared for the International Precious Metals Institute, Reno, Nevada, October 1992).16. G. Chi, M. Fuerstenau, and J. Maesden, “Study of Merril–Crowe Processing. Part I: Solubility of Zinc in Alkaline Cyanide Solution,” International Journal of Mineral Processing, 49 (1997), pp. 171–183.

José R. Parga, Jesús L. Valenzuela, and Francisco Cepeda T. are with the Instituto Tehnologico de Saltillo in Saltillo, Mexico. Dr. Parga can be reached at [email protected].

Figure 9. The results of continuous plant operation.

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for more information: [email protected]

Chemical Speciation and Bioavailability (2012), 24(3) 176

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aDepartment of Metallurgy and Materials Science, Institute Technology of Saltillo, V. Carranza 2400, CP. 25000, Saltillo Coahuila, MexicobDepartment of Chemical and Extractive Metallurgy, University of Sonora, Hermosillo Sonora, Mexico*E-mail: [email protected]

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The extraction of gold and silver from minerals and concentrates with cyanide is an important hydrometallurgy process that has been studied for more than 120 years. This technology, which consists of the dissolutions of the precious metals in cyanide solutions, followed by the recovery of the values by cementation, activated carbon or ion exchange resin. Most of WKH�ZDVWHV�LQ�WKH�LQGXVWULDO�HIÀXHQWV¶�PLOOLQJ�DUH�NQRZQ�WR�FRQWDLQ�KLJK�FRQWHQWV�RI�IUHH�F\DQLGH�DV�ZHOO�DV�PHWDOOLF�F\DQLGH�complexes, which give them a high degree of toxicity. Appropriate methods for the treatment of cyanide solutions include F\DQLGH�GHVWUXFWLRQ�E\�R[LGDWLRQ�XVLQJ�D�SKRWRHOHFWURFDWDO\WLF�GHWR[L¿FDWLRQ�WHFKQLTXH�ZLWK�WLWDQLXP�GLR[LGH�PLFURHOHFWURGHV��This is one of the most innovative ways for the treatment of wastewater containing cyanide. Another is the use of chlorine dioxide (ClO2) with a gas-sparged hydrocyclone as the reactor. The results show that photodegradation of cyanide was 93% in 30 minutes using a 450 W halogen�lamp, and in the case of ClO2, the destruction of cyanides was 99% in 1 minute. In both cases, excellent performances can be achieved with the high capacity of these technologies.

.H\ZRUGV: cyanide, titanium dioxide, ClO2, photoelectrocatalysis, gas sparged hydrocyclone

www.chemspecbio.co.uk

INTRODUCTION

Mexico is the world’s largest producer of silver. According

to annual records, the production of silver in 2011 was 37 x

106 ounces. Mexico, as does the rest of the world, uses the

traditional method of cyanidation to recover gold and silver in

which effluents containing cyanide are generated. This waste

is dangerous for the environment. In the extractive industry,

this method of recovering precious metals is carried out using

a solution of cyanide of 1000–3000 mg L-1 of NaCN. This

concentration is necessary because the cyanidation of gold

and silver ores is a complex heterogeneous process, where the

leaching behaviour of gold and silver depends upon the nature

of the gold and silver bearer as well of its distribution within

the ore with respect to its ability to be reached by the liquid

phase. The pH is greater than 10 and aeration is necessary

to keep the pulp or solution saturated with oxygen (>7 mg

L-1). The overall reaction for the dissolution of silver may be

expressed by the classic Elsner equation (Jefrey, 2000):

4Ag + 8 CN– + O2 + 2H

2O o 4Ag(CN)

2– + 4OH– (1)

The reaction has the following mechanism:

4Ag + 8CN– o 4Ag(CN)2– + 4e– (2)

O2 + 2H

2O + 2e– o 2OH- + H

2O

2 (3)

H2O

2 + 2e– o 2OH– (4)

In these reaction mechanisms, the cyanide ion is the

complexing agent or ligand and oxygen is the oxidant. Also,

in this redox process, oxygen removes four electrons from

the silver simultaneous with the transfer of protons (H+)

from water. At room temperature and standard atmospheric

pressure, approximately 8.2 mg of oxygen are present in 1 L

of water. This corresponds to 0.26 x 10-3 mol L-1. Accordingly,

the sodium cyanide concentration (molecular weight of

NaCN = 49) should be equal to 4 x 0.26 x 10-3 x 49 = 0.05 g

L-1 or approximately 0.01%. This was confirmed in practice

at room temperature by a very dilute solution of NaCN of

CSB1100052.indd 176 7/30/2012 7:05:01 PM

José R. Parga, Victor Vazquez, Jesús L.Valenzuela , Zully Matamoros and Gregorio Gonzalez 177

0.01–0.5% for ores, and for concentrates rich in gold and silver of 0.5 –5% (Parga, 2007). Aat the end of this leaching process, free cyanide enriched solutions, easily dissociable cyanide compounds, and metallic cyanide complexes are generated (Botz, 1998). Thus, the cyanide present in either form has a high degree of toxicity that needs treatment. In normal conditions, cyanide can be reduced by a natural degradation mechanism, by means of exposure of the cyanide

solution to solar radiation in large settling ponds or cyanide removal by volatilisation of free cyanide. Since the natural degradation process needs proper climatic conditions to be efficient, the use of a procedure involving chemical oxidation to accelerate the destruction of cyanide is often mandatory (Dasai, 1998). A review of these technologies along with their distinct advantages and disadvantage is show in Table 1. The most commonly used technology for cyanide destruction is

Table 1 Advantages and disadvantages of various cyanide destruction options (Desai, 1998; Mosher, 1996; Ogutueren, 1999; White, 2000;

Carrillo, 2001; Pavas, 2005)

Detoxification method

Advantages DisadvantagesSuitability for Kg reagent

Kg CN-

Time(min)Low CN- High CN-

SO2/Air

(INCO) Process

Reagent is very inexpensive and operational simplicity

For treating aqueous solutions and gold mine waste sludges

Used over a wide pH range

At least some of reagent savings are offset by license/royalty payments

Process adds sulfates to treated water

If precipitating ferrocyanides with copper, must dispose of precipitate

9 8 2.5 30

Hydrogen peroxide

Excess reagent decomposes to water and oxygen

Relatively simple to operate

Is not as reactive with thiocyanate

Reagent cost

If precipitating ferrocyanides with copper, must dispose of precipitate

Requires accurate measurement of chemical dose

8 9 3.8 60

Biological oxidation

“Natural approach”, received well publicly and by regulators

Uses heaps as a reactor, reducing total wash volumes, and possibly reach low flow areas of the heap more effectively

Relatively inexpensive

Can treat total cyanides without generating another waste stream

Technology is not well established

Requires combination of metallurgy, biology and process engineering

Tends to be very site specific, with each ore type requiring a specific evaluation and study

Viable cell systems may require long adaptation and do not function properly at cyanide concentrations higher than 200 mg L-1

9 8 1.9 106

Ozonation Some regeneration of cyanide possible.

In situ production

Does not produce salts

Solubility in water greater than oxygen

Produces ammonia

High reagent cost/equipment

Because hydroxide decomposes ozone, it becomes less efficient at pH values greater than 11.

9 9 1.9 20

Photocatalytic degradationTiO

2/Air/UV

Abundant solar energy and the process not add any toxic pollutant

Highly economical compared to artificial UV radiation, which require substantial electric power input

Highly reactive and oxidising hydroxyl radicals

Photocatalytic activity of TiO2

influenced by surface area, surface orientation and crystals structure

Difficulty of recovering all titanium dioxide used after the treatment process

UV radiation does not

significantly penetrate water and not available at night and is poor in the rainy season

9 8 1 103

ClO2 with a

GSHStrong oxidant and can treat total cyanides without generating another waste stream

Compact sealed unit for slurry treatment

High cyanide removal in 1 min

Low-cost reagents

Need compressed feed gas (ClO2)

to the GSH

For maximum process efficiency the porous inner tube need cleaner every week

9 5.2 1

CSB1100052.indd 177 7/30/2012 7:05:01 PM

Detoxification of cyanide using TiOs and GSH with ClO2178

the Inco cyanide remediation process and more than 90 mines worldwide now use an Inco SO

2/air detoxification circuit to

convert cyanide to the much less toxic cyanate before waste is discharged to a tailings pond. Typically, this process blows compressed air through the tailings while adding sodium metabisulfite which releases SO

2, lime to maintain the

pH at around 8.5, and copper sulfate as a catalyst if there is insufficient copper in the ore extract. This procedure can reduce concentrations of "weak acid dissociable" (WAD) cyanide to below the 10 ppm, as mandated by the EU's Mining Waste Directive (Robbins, 1996).

All these methods (see Table 1) are based on cyanide recovery by acidification and/or chemical oxidation destruction, however, none of the above-mentioned chemical processes is fully acceptable for cyanide detoxification and a combination of processes, in sequence, is required to meet the targeted cyanide concentration. Any chemical process may leave an excess of other undesirable chemicals in the wastewater stream. Also residence times required for biological process are typically much greater than those required for mass transfer. As mentioned previously, in many cases, the process involves a high cost of chemicals and royalty payments, in this regard, two promising alternatives to these processes can be found in the cyanide detoxification of effluents by the use of photocatalytic detoxification of cyanide (Blanco,2004) and the use of chlorine dioxide (ClO

2)

with a gas-sparged hydrocyclone (GSH) as the reactor.

TiO2 photocatalyst

In past few years, photocatalysis by polycrystalline semiconductors irradiated by near-UV light has been found

to be effective in oxidising certain organic and inorganic

pollutants to less dangerous species in mild reaction conditions

(Shifu et al., 2011; Xiong et al., 2011). This method was

found to be suitable for the oxidation of free and complex

cyanides dissolved in wastewater. The process works by

exposing wastewater to the combined forces of sunlight

and semiconductor catalyst. A commonly used catalyst is

titanium dioxide (TiO2) (Pavas, 2005). This catalyst may be

mixed into the water, creating slurry, or fixed onto lattice-

type structure where the water flows through. The TiO2 is

the semiconductor most used in photocatalysis because it

is chemically and biologically inert, it is not poisonous,

it is stable to photochemical and chemical corrosion, it is

abundant in nature and also possess a band of energy of 3.2

eV that can be excited by UV light of O<387nm which can be

supplied by solar light (Pavas, 2005).

The mechanism of photocatalysis is carried out when the

nanoparticles of TiO2 are illuminated by UV light, in which,

photons excite the band that contains the valence electrons

and these cross the conductance band, leaving spaces or

holes (h+), a characteristic feature of semiconducting metal

oxides is the strong oxidation power of their holes h+. If

the semiconductor (TiO2) is in an aqueous medium, those

spaces will react with the water molecules (H2O) producing a

hydroxyl radical ( OH) (Blanco, 2004), capable of oxidising

cyanide to cyanate (CNO–). Figure 1 shows the interrelation

of energy and the redox mechanism in the surface of TiO2

with the presence of oxygen.

In an aqueous medium with cyanide, the first product

of the photocatalytic oxidation using nanocrystals of TiO2

is CNO- (Augugliaro, 1999). The chemical reactions that

represent this oxidation process are as follows:

Figure 1 Scheme of photocatalytic process over TiO2 surface (Augugliaro, 1999).

CSB1100052.indd 178 7/30/2012 7:05:01 PM


Formation of spaces with UV light (solar light or artificial

source):

TiO2

+ hv o h+ + e– (5)

Reaction of valence band spaces (h) with cyanide ion

CN– + 2 h+ + 2 OH– o CNO– + H2O (6)

Finally oxygen reduction consumes the electrons generated

by reaction (5)

O2 + 2e– + 2 H

2O o H

2O

2 + 2 OH– (7)

Hydroxyl radical oxidation is considered to be the primary

mechanism for the destruction of the cyanide ion. The UV

required for photocatalytic processes may come from an

artificial source or the sun. The abundant solar energy

available in Mexico, South America, African countries,

Australia, etc. can be used effectively as a concentrated solar

radiation in a photocatalytic and photochemical processes for

the degradation of cyanide ions (Pavas, 2005). This radiation,

however, is obviously not available in the night and is poor in

the rainy season.

Gas-sparged hydrocyclone technology

The GSH technology was originally developed at the

University of Utah for the fast and efficient flotation of fine

particles from suspensions (Miller et al., 1989). Recent

studies indicate that the fluid flow conditions inside the GSH

system can be effectively exploited for air stripping of VOCs

from contaminated water (Miller, 1996).

The GSH unit, shown schematically in Figure 2, consists

of two concentric right-vertical tubes and a conventional

cyclone header at the top. The porous inner tube is constructed

of any suitable material such as plastic, ceramic, or stainless

steel and allows for the sparging of air or any other gas or

steam. The outer nonporous tube simply serves to establish

a gas jacket and provides for the even distribution of the

ClO2(g)

through the porous tube. Thus, the GSH reactor can

be used for chemical oxidation in which case the destruction

of cyanide can be achieved with an oxidising gas such as

ClO2(g)

which is considered here. The cyanide solution is fed

tangentially at the top through the cyclone header to develop

a swirl flow adjacent to the inside surface of the porous

tube, leaving an empty air core centred on the axis of the

GSH unit. The high-velocity swirl flow shears the sparged

air to produce a high concentration of small bubbles and

intimate interaction between these numerous fine bubbles

of ClO2(g)

and the cyanide solution. Gaseous products are

then transported radially to the centre of the cyclone (Miller,

1989). The major portions of the gas phase move towards

the vortex finder of the cyclone header, and are vented into

an appropriate post-treatment device. The specific capacity

of the GSH system is at least 300–400 gallons per minute

per cubic foot of equipment volume, 100–600 times that of

conventional gas-stripping equipment. The GSH equipment

requires an operating space significantly less that of a reactor

tank, packed tower or other air stripping devices, which result

in a significant savings in capital cost (Miller, 1996).

MATERIALS AND METHODS

Photocatalytic oxidation test

The cyanide photocatalytic oxidation tests were carried out

in a 400 mL beaker. Catalyst concentrations of 1.25, 1.0,

0.75 and 0.5 g L-1 of TiO2 (Degussa P-25) were used with

an average particle size of 30 nm and BET surface area of

GSH GAS

Tank UnderŇow OverŇow

pH=9 SO2

NaOH

HCN

GAS Barren

SoluƟon of cyanide

Ca(OH)2

Figure 2 Schematic representation of the gas-sparged hydrocyclone reactor.

CSB1100052.indd 179 7/30/2012 7:05:16 PM


about 55 m2 g-1. The initial concentration of cyanide was 400 ppm, the solution was prepared using a NaCN, reagent grade (95%), (Monterrey chemical products), a 450 W halogen lamp was used as the UV light source, the solution tests were equilibrated by stirring for 30 minutes in the dark before exposing the reactor assembly to the concentrated light from the lamp on the top of the beaker. Then, a sample was removed for analysis. The concentration of cyanide in this filtered sample was treated as the zero time concentration in each experiment before exposure to UV light. Cyanide determinations were carried out using the selective ion electrode technique adjusting the pH to 11 in order to avoid HCN formation. The setup used for the photocatalytic cyanide detoxification process is shown in Figure 3. The reactor was enclosed inside an aluminum reflecting compartment.

Experimental procedure of cyanide oxidation by ClO2

The properties of ClO2, like those of chlorine, must necessarily

be considered both in the gas phase and in the aqueous phase (Perry, 1990). ClO

2 has a disagreeable, irritating odour

similar to chlorine. ClO2(g)

and may be readily decomposed on exposure to UV light, therefore, it is always stored in the dark. The sensitivity of gaseous and liquid ClO

2(g) to temperature

and pressure has mitigated against its bulk shipment. ClO2(g)

can be handled safely in the gas phase by diluting with air or nitrogen to keep its concentration below about 10%. For potable water and wastewater treatment process, it is only used in aqueous solutions and is generated on site.

Finally, the environmental impacts associated with the use of ClO

2(g) as wastewater disinfectant or waste treatment are

not well known. It has been reported that the impacts are less adverse than those associated with chlorination (Freeman, 2006). Since it does not react with water, its oxidation potential and strength are not pH dependent. Consequently, ClO

2(g) is

superior to chlorine particularly for wastewater of relatively

high pH in which the concentration of hypochlorous acid is low (Chang, 2000).

Experiments for cyanide oxidation by ClO2, at the Institute

Technology of Saltillo pilot plant included preparation of the barren solution with SO

2 gas (pH=2–11) and stripping with

a 2-inch diameter GSH unit. Chemical analysis for cyanide in the effluent streams was accomplished with a reflux distillation method. The collected cyanide was quantified by titration with silver nitrate standard solution and/or the ion selective electrode technique.

During the experiments two streams had to be delivered to the GSH; the cyanide solution and the air. Cyanide solution was provided by a sump pump mounted on a 300 L retention tank. The cyanide solution flow rate was adjusted using a regulated return flow to the tank. Using an air compressor, airflow was evenly distributed between the upper and lower sections of GSH and all parts were sealed with gaskets. Operators were provided with personal HCN gas monitor/alarm units (Ogutueren, 1999).

The method used for generating on-site ClO2(g)

consists of blending a 45% solution of sodium chlorate with 66°Be sulfuric acid in the top of a reaction vessel. Air containing 10% SO

2(g) was blown into a diffuser at the bottom of this

vessel and ClO2 plus air was extracted at the top of the vessel

(ClO2 content was between 6 and 10%). This is the so-called

Mathieson process (Perry,1990 ; Kesting, 1953). The basic reactions are:

2NaClO3 + H

2SO

4 o 2HClO

3 + Na

2SO

4 (8)

2HClO3 + SO

2(g) o 2ClO

2(g) + H

2SO

4 (9)

Side reactions also take place, including:

2NaClO3 + 5SO

2(g) + 4H

2O o Cl

2(g) +3H

2SO

4 + 2NaHSO

4 (10)

Operating conditions are optimised to suppress the latter reaction as much as possible, since this produces only chlorine. The exit gases are reacted with the cyanide solution in the GSH reactor where the oxidation or destruction of cyanide is expected to occur according to the following reaction:

CN-+ 2ClO2(g)

+ 2 OH- o CNO- + 2ClO2- + H

2O (11)

The stoichiometry of the reaction shows that 5.2 kg., of ClO2

is needed to neutralise 1 kg. CN-.


Results of the photocatalytic technique for destruction of

cyanide

For cyanide determination, a calibration curve was constructed initially, the results are shown in Figure 4. The cyanide oxidation results using the photocatalysis technology with TiO

2 nanoparticles are shown in Table 2. Blank tests

indicated that the cyanide conversion percentage were less than 1% without either UV light or TiO

2 catalyst.Figure 3 Experimental setup for photocatalytic oxidation of cyanide.

CSB1100052.indd 180 7/30/2012 7:05:17 PM


In Table 2 we can observe that cyanide oxidation using the

photocatalysis technology is a viable option to eliminate this

water pollutant at pH = 11.2. Also, it can be ascertained from

Table 2 that efficiency increased rapidly when the weight

of the catalyst is less than 500 mg. Results indicated that

the cyanide oxidation was held at 93% within a 30 minutes

period. This is due to the fact that an optical saturation

happens with an increased concentration of TiO2 catalyst,

a phenomena that was also observed by other investigators

(Blanco, 2004). Laboratory experiments are performed in

small reactors and the results from these experiments cannot

be directly extrapolated for the design of a larger industrial

ponds.

Cyanide destruction with a GSH reactor

Effect of pH on cyanide destruction

One of the primary process variables for cyanide destruction

with ClO2(g)

is the pH of the solution. Table 3 shows the

results of a set of experiments for the effect of pH on cyanide

destruction. The effluent cyanide concentration (both free and

combined cyanide) is plotted as a function of pH in Figure 5.

It is evident that free cyanide is destroyed at all pH values.

Table 2 Percent cyanide destruction results

[TiO2]

(g L-1)[CN-]

IN

(mg L-1)[CN-]

OUT

(mg L-1)Destruction of CN–

(%)

0.5 400 28 93

0.75 400 40 90

1.0 400 52 87

1.25 400 48 88

Figure 4 Calibration curve for measurement of cyanide.

The combined cyanide CNx is destroyed most effectively at

high pH values.

Effect of NaCl in cyanide destruction

The use of NaCl was considered. These experiments were designed to eliminate the use of sulfur dioxide (SO

2) for the

ClO2 generation and minimise the danger of explosion. The

basic reaction for this process is:

2NaClO3 + 2NaCl + 2H

2SO

4 o 2ClO

2(g) + Cl

2(g)

+ 2Na2SO

4 + 2H

2O (12)

Air is blown into the bottom of the vessel driving off the gas formed which is then diluted by the incoming air. Then the mixture of ClO

2/Cl

2 gas, which is extremely volatile, is

stripped from the aqueous solution by the air and is then circulated through the GSH reactor. The data are presented in Table 4. The results show 99.9% destruction of free cyanide and the response is similar to that when SO

2 was used for

ClO2 generation.

CONCLUSIONS

It has been shown that cyanide destruction with photocatalytic oxidation can be achieved in an economical and environmentally acceptable manner.

Results show that the photocatalysed oxidation of cyanide waste with TiO

2 semiconductor is very efficient for the

destruction of cyanide. Over 90% of cyanide depletion was observed in the cyanide concentration range studied.

The use of ClO2 in the GSH reactor has been tested in

bench and pilot plant scale applications and has been proven

Table 3 Effect of pH on cyanide destruction

Feed solution = 250 mg L-1 total cyanide Solution flow rate = 40 L min-1

Gas (6% ClO2) flow rate = 100 L min-1

Cyanide solution (mg L-1) Final effluent (mg L-1) % Destruction

pH Total cyanide Free cyanide Combined cyanide Free cyanide Combined cyanide (CNX)

11.23 250 0.12 52.54 99.9 78.8

9.20 250 0.19 103.0 99.9 58.8

2.57 250 0.11 187.0 99.9 25.2

CNX = CNO-, HCNO, Fe(CN)

6-4, etc.

Figure 5 Effect of pH in cyanide destruction.

CSB1100052.indd 181 7/30/2012 7:05:17 PM


to be effective for the destruction of cyanide in solution and

slurries, the results show 99.9% destruction of free cyanide

and the response is similar to that when SO2 was used for

ClO2 generation.

Finally, the economics of the ClO2

process are quite

attractive because there are cost savings due to the short

residence time, decreases in operating staff and other chemical

utilised (i.e. copper sulfate and NaOH) in the other processes.

ACKNOWLEDGEMENTS

The authors wish to acknowledge support of this project to the

National Council of Science and Technology (CONACYT)

and to Dirección General de Educación Superior Tecnológica

(DGEST).

REFERENCES

Augugliaro, V., Blanco, J. and Caceres, J. (1999) Catalysis Today, 54,

245-253.

Blanco, J. and Malato, S. (2004) PhD thesis, Plataforma Solar de

Almería, 14.

Botz, M. and Stevenson, J. (1995) Eng. Min. J., 44-47.

Chang, C.Y., Hsieh, Y.H., Hsu, S.S., Hu P.Y. and Wang, K.H. ( 2000) J.

Hazard. Mater., B79, 89-102.

Carrillo, F.R., Pedroza, M.J. and Soria, A. (2001) Eur. J. Min. Process.

Envir. Protect., 1, 55-63.

Desai, J.D., Ramakrishna, C., Patel, P.S. and Awasthl, S.K. (1998) Chem.

Eng. World, XXXIII, 6,115-121.

Freeman, H.M. (1988) Standard handbook of hazardous waste treatment

and disposal. McGraw-Hill, New York.

Jeffrey, M.I. and Ritchie, I.M. (2000) J. Electrochem. Soc. I, 147, 3257-

3262.

Kesting, E. ( 1953) TAPPI, 36, 176-171.

Miller, J.D. and Kinneberg, D.J. (1984) International Conference on

Recent Advances in Mineral Science and Technology, pp. 337-338.

Johannesburg.

Miller, J.D. and Ye Yi (1989) Miner. Proc. Extract. Metall. Rev., 307-309.

Miller, J.D., Lelinski, D. and Parga, J.R. (1996), Final report-CX

823711, Advance Process Technology for the Wastepaper Recycling

Plants and Pulp/Paper Plants. Southwest Centre for Environmental

Research and Policy.

Mosher, J.B. and Figueroa, A. (1996) Miner. Eng., 5,573-581.

Robbins, G.H. (1996) CIM Bull., 9, 63-69.

Ogutueren, U.B., Toru, E. and Koparal, S. (1999) Water Res., 33, 1851-

1856.

Parga, J.R., Valenzuela, J.L. and Cepeda, F. (2007) J. Metals, 10, 43-47.

Pavas, E. Camargo, M. and Jones, C. (2005), MSc thesis, Universidad

EAFIT, 11,7.

Perry, W.L. andEckenfelder, W.W. (1990) Toxicity reduction in industrial

effluents. New York: Van Nostrand Reinhold Co.

Shifu, Ch. Yunguang, Y. and Wei, L. (2011) J. Hazard. Mater., 186,

1560-1567.

White, D.M., Pilon, T.A. andWoolard, C. (2000) Water Res., 34, 2105-

2109.

Xiong, Z. Ma, J., Ng, W.J., Waite, T.D. and Zhao, X.S. (2011) Water

Res., 45, 2095-2103.

Table 4 Effect of NaCl in cyanide destruction

Feed solution = 250 mg L-1 Total cyanide. Solution flow rate = 40 L min-1

Gas (ClO2/Cl

2) flow rate = 100 L min-1 pH = 11.2

Barren bleed (mg L-1) Final effluent (mg L-1) % Destruction

NaCl (g L-1) Total cyanide Free cyanide Combined cyanide Free cyanide Combined cyanide

1.23 250 0.02 52.54 99.9 78.8

0.80 250 0.10 149.0 99.9 40.4

0.40 250 0.09 193.0 99.9 22.8

0 250 0.10 222.0 99.9 11.2

CNX = CNO-, HCNO, Fe(CN)

6-4, etc.

CSB1100052.indd 182 7/30/2012 7:05:17 PM

09 a 12 de setembro de 2012 Búzios, RJ

ESTUDIO CINETICO DE LA EXTRACCIÓN DE ORO DE UN MINERAL PIRITOSO MEDIANTE OXIDACIÓN Y LIXIVIACIÓN SIMULTÁNEA A PRESIÓN

J.L. VALENZUELA G.1

, L.S. QUIROZ C.1

, J.R. PARGA T.2

, P.J. VALENZUELA G.3

,

P. GUERRERO G.1

1

Universidad de Sonora, Departamento de Ingeniería Química y Metalurgia, México.

2

Instituto Tecnológico de Saltillo, Departamento de Metalurgia y Materiales, México

3

Minas de Oro Nacional, Mulatos, Sonora, México.

E-mail para contacto: [email protected]

RESUMEN – Oxidación a presión es un pre tratamiento oxidativo

hidrometalúrgico utilizado para menas refractarias sulfurosas. Se realizó un

estudio de la oxidación a presión en medio alcalino y lixiviación simultánea con

cianuro de sodio a un mineral piritoso proveniente de la mina Mulatos, localizada

en el municipio de Sahuaripa, Sonora, Mexico. Se muestran los resultados de

caracterización del material y de los procesos de lixiviación, en los cuales las

variables a modificar utilizadas fueron temperatura, presión y concentración de

cianuro. La caracterización del material indicó la presencia de pirita como mineral

principal. Se llevaron a cabo pruebas en rangos de temperatura de 60 – 80 °C,

presión de 87.5 a 150 psi, utilizando oxígeno, y concentración de cianuro de 0.8%

a 1.2% en peso, utilizando un reactor a presión Parr de 2 litros.

1. INTRODUCCIÓN

En la extracción de metales preciosos utilizando hidrometalurgia, el proceso más

comúnmente utilizado es la lixiviación con cianuro de sodio, se muestran las reacciones para

oro y plata, Fleming (1992), Marsden y House (2006).

4Au + 8CN-

+ 2H2O + O2 → 4 Au(CN)2

-

+ 4 OH-

(1)

4Ag + 8CN-

+ 2H2O + O2 → 4 Ag(CN)2

-

+ 4 OH-

(2)

Existen minerales de oro y plata llamados “refractarios“ los cuales se encuentran en una matriz piritosa, para los que la cianuración no es efectiva, por lo cual se recurre a pre-

tratamientos oxidativos a fin de convertir los minerales sulfuros a óxidos porosos en las que la

cianuración funciona más eficientemente. La reacción (3) muestra la oxidación de pirita en

medio alcalino y la formación del óxido poroso, Koslides y Ciminelli (1992).

2FeS2 (s) + 15/2 O2 (g) + 8H-

(aq) → Fe2O3 (s) + 4SO4

2-

(aq) +4H2O(l) (3)

Se ha utilizado el tratamiento de oxidación química a presión y lixiviación con cianuro

en simultáneo para un concentrado de flotación producido en la mina de Bacís, Durango con

buenos resultados Parga et al. (2007). Se propone utilizar este tratamiento para un mineral

Nicolas Martinez


refractario piritoso de la región de Mulatos, Municipio de Sahuaripa, Sonora, México,

manipulando variables tales como presión, temperatura y concentración de NaCN, para

encontrar las mejores condiciones de oxidación lixiviación.

2. METODOLOGÍA

Se obtuvo la muestra del mineral de Mulatos, la muestra se denominó SAS-1, un sulfuro

silicificado con contenido de oro y plata. La muestra fue reducida de tamaño hasta un

intervalo entre 45 y 53 µm. Se caracterizó el mineral con las técnicas de Difracción de Rayos

X, Microscopía Electrónica de Barrido, Espectroscopía de Absorción Atómica,

Espectroscopía de Dispersión de Energía de Rayos X para la determinación de especies

minerales y elementales, junto con pruebas metalúrgicas de lixiviación agitada, cianuración en

caliente y ensaye al fuego para determinar porcentajes de extracción del oro, metal de interés.

El sistema experimental consistió en un reactor a presión Parr de 2 litros con

intercambiador de calor, agitación y válvulas de control para entrada y salida de gases, Figura

1. El mineral se mezcló con agua para formar una pulpa, a la cual se añadieron los reactivos y

se introdujo al reactor. Se utilizó oxígeno puro para el reactor a presión, se añadieron cianuro

de sodio como lixiviante y cal, para mantener un pH superior a 11, con el propósito de evitar

la reacción de producción de ácido cianhídrico, el cual es volátil y venenoso; ésta ocurre a pH

menores a 9.5, ilustrada en la ecuación (4).

H+

+ NaCN → HCN + Na+

(4)

Las pruebas se realizaron con el reactor completamente agitado. Posterior al tiempo de

reacción dentro del recipiente a presión, el contenido del reactor se filtró al vacío para separar

sólido y líquido para su posterior análisis. Se calculó el porcentaje de extracción de oro en

base a los resultados obtenidos.

Figura 1. Reactor Parr utilizado en la Experimentación, equipado con chaqueta de

calentamiento y agua de enfriamiento para control de temperatura.


2.1 Caracterización del Mineral

Los resultados de la caracterización indicaron la presencia de pirita (FeS2) como principal

componente mineral, así como cuarzo. El mineral tiene un contenido de oro, plata y otras

especies elementales como se muestra en las Tablas 1 y 2. Los análisis de DRX y EDS

confirmaron la presencia de sulfuro de hierro FeS2 y SiO2 como principales especies, como se

muestra en las Figuras 2 y 3.

Tabla 1 – Resultados de Ensaye al Fuego y Espectroscopia de Absorción Atómica para la

Muestra SAS-1

Au (g/t) Ag (g/t) Cu (%) Fe (%) Pb (%) Zn (%)

SAS-1 4 23 0.029 6.33 0.061 0.0068

Tabla 2 – Resultados de Contenido de Oro y Azufre y Consumos de Reactivo

Muestra

Au (g/t)

Extracción Au (%) Consumo (kg/t)

S (%) Ensaye Calculada

Cabeza Cola Cabeza Ens Calc CaO NaCN

SAS-1 3.97 1.65 3.96 57.4 58.43 4.98 0.86 16.32

Figura 2. Imágenes de Microscopía Elecrónica de Barrido y Espectroscopía de Dispersión de

Energía de Rayos X de la muestra SAS-1.


SAS_1

(Room) - Time Started: 18 s - 2-Theta: 5.000 ° - Theta: 2.500 ° - Chi: 0.00 ° -

Lin

(Cou

nts)

0

100

200

300

400

2-Theta - Scale10 20 30 40 50

d=7.

1558

9

d=3.

5752

3

d=7.

1558

9

d=4.

2526

1

d=3.

5752

3

d=3.

3413

9

d=3.

1272

9

d=2.

7083

2 d=2.

4552

9d=

2.42

249

d=2.

2805

8d=

2.23

693

d=2.

2115

6

d=2.

1265

9

d=1.

9789

9

d=1.

9153

7 d=1.

8179

0

d=1.

6713

5

d=1.

6337

7

d=3.

3641

5

Figura 3. Resultado de Difracción de Rayos X para la muestra SAS-1. El espectro indica la

presencia de cuarzo y pirita, SiO2 y FeS2 respectivamente.

Los resultados de SEM y DRX indican altos contenidos de hierro, azufre y silicio, los

cuales se encuentran en las especies de pirita y cuarzo. Como se aprecia en la microfotografía,

las partículas con aspecto brillante son las que contienen mayor cantidad de Fe y S.

3. RESULTADOS Y DISCUSIÓN

3.1. Pruebas experimentales

Las pruebas de lixiviación a presión atmosférica revelan porcentajes de extracción bajos

relativamente, así como las pruebas de cianuración caliente en el rango de 40 – 80°C. Las

pruebas de lixiviación atmosférica se llevaron a cabo bajo las siguientes condiciones:

porcentaje de sólidos 30% w/w, concentración de NaCN 300 mg/L, pH≥ 10.5 (CaO), tiempo lixiviación 72 h, tamaño de partícula p80 3/16” (100%-3/8”). Las pruebas de cianuración caliente se realizaron bajo condiciones de: concentración de NaCN 1%, adición de CaO para

control de pH de 10 Kg/ton, temperatura del rango de 40°C - 80°C, porcentaje de sólidos de

30%, tiempo de 2.5 horas, tamaño de partícula de - 270 mesh (55µm). Las pruebas de

oxidación y lixiviación a presión en simultáneo se realizaron con variables modificadas de

presión en rango de 87.5 – 200 psi, temperatura en el rango de 60 – 150°C, y concentración

de cianuro de 0.8 – 1.2 %w., tiempo de 1 hora y pH superior a 11.


Figura 4. Resultados de lixiviación a presión atmosférica de la muestra SAS-1

Figura 5. Resultados de la cianuración caliente de la muestra SAS-1

0%

5%

10%

15%

20%

25%

30%

35%

40%

0 20 40 60 80

Ext

racc

ión

de A

u

Tiempo, h

48%

49%

50%

51%

52%

53%

54%

55%

0 0.5 1 1.5 2 2.5 3

Ext

racc

ión

de A

u

Tiempo, h

40°C

60°C

80°C


Figura 6. Curvas de extracción contra temperatura a presión constante para

concentración inicial de cianuro de 0.8 %w.

Figura 7. Curvas de extracción contra temperatura a presión constante para concentración

inicial de cianuro de 1.0 %w.

Al comparar la extracción en los procesos de lixiviación a presión atmosférica y

lixiviación oxidación a presión, puede apreciarse que los obtenidos en el reactor a presión se

obtiene una mayor extraccion (%), y obtenidos en un tiempo mucho menor. Se ha apreciado

un efecto similar en el estudio de las operaciones de oxidación y lixiviación con cianuro por

separado, donde los porcentajes de extracción son directamente afectados por la oxidación de

los sulfuros, y ocurría un aumento del 20% al 70% de extracción de oro, como lo presenta Li

et al. (2006).

0

10

20

30

40

50

60

70

80

90

100

0 50 100 150 200

Ext

racc

ión

de A

u, %

Temperatura, °C

P = 87.5 psi

P = 120 psi

P = 150 psi

P = 200 psi

0

10

20

30

40

50

60

70

80

90

100

0 50 100 150 200

Ext

racc

ión

de A

u, %

Temperatura, °C

P = 87.5 psi

P = 120 psi

P = 150 psi

P = 200 psi


Figura 8. Curvas de extracción contra temperatura a presión constante para concentración

inicial de cianuro de 1.2 %w.

Los resultados experimentales arrojaron que el efecto de la temperatura es mayor que el

efecto de la presión en los porcentajes de extracción. Es posible que el tiempo de reacción de

una hora, considerado como suficiente para un caso ya estudiado por Parga et al. (2007), no

haya bastado para formar la capa porosa de producto oxidado vista para la pirita en solución

alcalina Koslides y Ciminelli (1992), debido a diferencias en la mineralogía. Cabe mencionar

que posiblemente por estas condiciones de oxidación parcial, la presión de oxígeno no mostró

tener un efecto mayor en el porcentaje de extracción.

Figura 9. Resultados de pruebas de oxidación y lixiviación a presión, para extracción de oro

contra tiempo a tres distintos valores de temperatura. Condiciones de 0.8%w de NaCN y

87.5 psi.

0

10

20

30

40

50

60

70

80

90

100

0 50 100 150 200

Ext

racc

ión

de A

u, %

Temperatura, °C

P = 87.5 psi

P = 120 psi

P = 150 psi

P = 200 psi

0

10

20

30

40

50

60

70

80

90

100

0 20 40 60 80 100

Extr

acci

ón d

e or

o, %

Tiempo, min

110°C

90°C

120°C


3.2. Ajuste a un Modelo Cinético de Lixiviación

Se presentan los resultados de las pruebas cinéticas continuas en la Figura 9, donde se

muestra la extracción de oro a tres distintas temperaturas. Posteriormente se toma la ecuación

del modelo de núcleo decreciente para control por difusión a través de la capa de producto

Sohn y Wadsworth (1986), Ballester et al.(2007), utilizando α como la fracción reaccionada

de la mena y t como el tiempo de reacción.

( )2

3

21 1

3

kta a- - - = (5)

Al graficar la parte izquierda de la ecuación contra el tiempo (Figura 10), se obtiene una

gráfica recta con una pendiente que equivale a la constante cinética, k, que se utiliza en el

cálculo de la energía de activación (Tabla 3).

Figura 10. Gráfica de los datos de extracción de oro contra tiempo, para los tres valores de

temperatura estudiados en un período de tiempo de 60 minutos, utilizando el modelo de

núcleo decreciente de control por difusión a través de la capa de producto.

Posteriormente se construye una gráfica de Arrhenius (Figura 11) para determinar el

valor de la energía de activación, el cual permite saber si la cinética de la reacción química

está controlada por el transporte por difusión de los reactivos hacia la superficie de la

partícula o por la reacción química superficial, con valores menores a 20 kJ/mol y mayores a

40 kJ/mol para cada etapa controlante respectivamente. La ecuación de Arrhenius (Ec. 6) en

su forma logarítmica nos indica que los datos de las constantes cinéticas k contra el inverso de

los valores de temperatura absoluta, nos da una gráfica cuya pendiente negativa es igual a la

energía de activación de la reacción sobre la constante de los gases ideales, con la cual se hace

el cálculo, expresando en kJ/mol el valor de .

1ln ln

aEk AR T

= - (6)

y = 0.000x + 0.001R² = 0.952

y = 0.001x + 0.017R² = 0.842

y = 0.002x + 0.019R² = 0.941

0

0.05

0.1

0.15

0.2

0.25

0 10 20 30 40 50 60 70

1-2ɑ

/3-(

1-ɑ)

^(2/

3)

Tiempo, min

90°C

110°C

120°C

Lineal (90°C)

Lineal (110°C)

Lineal (120°C)


El índice de correlación de los datos cinéticos en cuanto a si hay concordancia con el

modelo, para las temperaturas de 90 y 110°C, se considera baja, del orden de 0.62 a 0.70 para

90 minutos de reacción. Cabe mencionar que para los primeros 60 minutos a estas dos

temperaturas, la concordancia con el modelo es del orden de 0.92 a 0.96. El comportamiento

de la reacción a esas temperaturas a tiempos mayores de 60 minutos podría deberse a que la

oxidación de la partícula ha llegado a un punto donde la partícula comienza a exponer las

superficies de las partículas de oro y la lixiviación se vuelve más eficiente. Quiroz (2010).

Los valores de las pendientes k para las ecuaciones del modelo de control por difusión y

las tres distintas temperaturas se expresan en la Tabla 3.

Figura 11. Gráfica de Arrhenius para el cálculo de la energía de activación de la oxidación y

lixiviación a presión usando el modelo de control por difusión a través de la capa de producto.

Tabla 3 – Cálculo de la energía de activación según gráfica y ecuación de Arrhenius para el

modelo de control por difusión a través de la capa de producto.

T, °C pendiente, k T, °K 1000/T ln k (-ܧ /R) ܧ, kJ/mol

90 0.0008 363.15 2.75368305 -7.1308988 -4.574 38.0303949

110 0.0011 383.15 2.60994389 -6.8124451

120 0.0023 393.15 2.54355844 -6.0748461

4. CONCLUSIONES

El tratamiento de oxidación y lixiviación simultánea a presión ha mostrado mejoras

considerables en los porcentajes de extracción de oro, en comparación con la lixiviación a

presión atmosférica, en tiempos mucho menores y a condiciones de presión moderadas, lo

cual es un factor importante a considerar al hacer el escalamiento hacia un nivel de planta

piloto.

y = -4.5744x + 5.3842

R² = 0.8227

-7.4

-7.2

-7

-6.8

-6.6

-6.4

-6.2

-6

2.5 2.55 2.6 2.65 2.7 2.75 2.8

ln k

1000/T, K


A condiciones normales de cianuración, la disolución del oro fue de 38-39%, en la

cianuración utilizando temperatura de 40 – 80°C la disolución del oro se incrementó a 50-

54%, y finalmente utilizando oxidación y lixiviación a presión de 87.5 psi, dio una disolución

final del 81-86%.

La oxidación y lixiviación a presión simultánea produjo un aumento en los porcentajes

de extracción de oro a comparación con las pruebas a condiciones normales.

Las condiciones más adecuadas que se determinaron son: T = 90°C, P =87.5 psi,

[NaCN] = 8 g/L, t = 1 hora, pH > 11 y agitación constante.

El análisis cinético de los datos obtenidos de la experimentación nos demuestran que el

proceso controlante de la lixiviación oxidante a presión se ajusta a un modelo de núcleo

decreciente de control mixto, ya que los valores de la energía de activación fueron de 30 – 38

kJ/mol. Finalmente se concluye que para este mineral sulfuro altamente silicificado, la

temperatura y presión oxidante tienen un efecto altamente significativo, ya que la disolución

del oro se incrementó en un 45%.

5. RECONOCIMIENTOS

Se extiende un agradecimiento y reconocimiento al Departamento de Ingeniería

Química y Metalurgia de la Universidad de Sonora, a la empresa Minas de Oro Nacional por

su colaboración con el suministro de muestras y ensayes metalúrgicos, así como al proyecto

PIFI 2011 de Fortalecimiento Institucional de la SEP de Mexico.

6. REFERENCIAS

BALLESTER A., VERDEJA L.F., SANCHO J. Metalurgia Extractiva, Vol.1 Fundamentos.

Madrid: Editorial Síntesis, 2007.

FLEMING, C.A. (). Hydrometallurgy of precious metals recovery. Hydrometallurgy, v.30, p.

127-162, 1992.

KOSLIDES, T.; CIMINELLI, V.S.T. Pressure oxidation of arsenopyrite and pyrite in alkaline

solutions, Hydrometallurgy, v. 30 p. 87-105, 1992.

LI, J.; DABROWSKI, B.; MILLER; J.D., ACAR; S., DIETRICH, M.; LEVIER, K.M.; WAN,

R.Y. The influence of pyrite pre-oxidation on gold recovery by cyanidation. Minerals Engineering, v.19, p. 883-895, 2006.

MARSDEN, J.O.; HOUSE, C.I. The Chemistry of Gold Extraction. Litleton, Colorado, USA:

Ed. SME, 2006.

PARGA, J.R.; VALENZUELA, J. L.; CEPEDA T. F. (). Pressure cyanide leaching for

precious metals recovery, JOM v. 59, 43-47, 2007.

QUIROZ C., L.S. Tesis de Maestría, Depto. Ingeniería Química y Metalurgia, Universidad de

Sonora, 2010.

SOHN H.Y.; WADSWORTH M.E. Cinética de los Procesos de la Metalurgia Extractiva.

México: Editorial Trillas, 1986.

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Time of distribution air (sec.)

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4

New Technology for Recovery of Gold and Silver by Pressure Cyanidation

Leaching and Electrocoagulation José R. Parga1, Jesús L. Valenzuela2 and José A. Díaz1

1Department of Metallurgy and Materials Science Institute

Technology of Saltillo, Saltillo Coahuila 2Department of Chemical Engineering and Metallurgy

University of Sonora, Hermosillo, Sonora

México

1. Introduction

The present chapter describes a new technology of pressure oxidation/cyanidation leaching for the dissolution of gold and silver and the recovery of the precious metals by using the electrochemical process of Electrocoagulation (EC). The novel method demonstrates that the oxidation and dissolution of gold and silver in alkaline cyanide solution can be conducted simultaneously in the same reactor in less than 90 minutes with a recovery that exceeds 96%. Then, the pregnant cyanide solution with gold and silver is sent for recovery of precious metals by using a very promising electrochemical technique (EC) that does not require high concentrations of silver and gold in cyanide solutions.

Gold is classed as a noble metal because of its inertness to most chemical reactions under ordinary conditions. At the present time, cyanidation has superseded all previous leaching processes, particularly chlorination, because of its ability to effectively and economically treat ores containing as little as 1-3 g/ton gold. Cyanidation processes are especially suitable for treatment of gold/silver-bearing sulphidic materials. Gold cyanidation has been reported to involve the chemical reactions shown in Eqs. (1) and (2). Silver is accomplished in the same fashion (Senanayake, 2008; Parga et al., 2007).

2Au + 4NaCN+O2+2H2O ń 2 Na[Au(CN)2]+2NaOH+H2O2 (1)

2Au + 4NaCN+H2O2 ń 2 Na[Au(CN)2]+2NaOH (2)

The Equation (1) proposed by Elsner is stoichiometrically correct but does not describe the cathodic reactions associated with the dissolution. The stoichiometry of the process shows that 4 moles of cyanide are needed for each mole of oxygen used. At room temperature and std. atmospheric pressure, approximately 8.2 mg of oxygen are present in one liter of water. This corresponds to 0.27 x 10-3 mol/L. Accordingly, the sodium cyanide concentration (molecular weight of NaCN = 49) should be equal to 4 x 0.27 x 10-3 x 49 = 0.05 g/L or approximately 0.01%. This was confirmed in practice at room temperature by a very dilute

Noble Metals 72

solution of NaCN of 0.01% - 0.5% for ores, and for concentrates rich in gold and silver of 0.5 % -5 % (Parga et al., 2007). Details of this electrochemical reaction have received considerable attention and under certain circumstances the reaction is limited by the coupled diffusion of CN- and O2 to the gold surface. Lime or sodium hydroxide (caustic) is added to keep the system at an alkaline pH of 10-11. This protective alkalinity is required to counteract the generation of acid during cyanidation, thereby preventing cyanide degradation and the formation of the deadly HCN gas.

Gold and silver ores are classified as refractory when a significant portion of the precious metals cannot be extracted by the conventional cyanidation process. The refractoriness may be of a physical or chemical nature. The former type is usually due to sub-microscopic particles of gold being locked within mineral particles, for example in sulphides or silicates. Refractory gold-silver ores contain precious metals locked up in a matrix of pyrite and/or arsenopyrite (Parga et al., 2007). Such ores are not amenable to cyanidation. To liberate the gold-silver from the sulfide matrix and render it accessible to cyanidation. It is a common practice to subject them to a preoxidation in order to enhance the recovery of the precious metals. Although effective, these techniques result in the production of large quantities of environmental hazardous substances such as sulphur dioxide gas or sulfuric acid, which can cause difficulties in compliance with environmental pollution regulations. After review of the literature (Parga et al., 2007), it was shown that roasting is currently the most cost effective means of oxidizing refractory pyrite and arsenopyrite concentrates to produce a product amenable to the cyanidation process. However, the major drawback of roasting is that it produces large quantities of SO2 gas which is released into atmosphere and this is not acceptable. The sulphur dioxide must be collected, or an alternative technology used. One such alternative process that has previously been used successfully around the world, is that of pressure aqueous pre-oxidation of the sulphide minerals. Pressure oxidation has been the most popular method for the treatment of refractory gold concentrates (see Table 1); This process utilizes oxygen or air at high pressures and temperatures to oxidize an aqueous slurry of the ore or concentrate to produce hematite, iron sulphates and considerable quantities of free sulphuric acid. Because of this, before the cyanidation process excessive lime or caustic soda must be used in order to the pH of the pulp to 10 or 11.

This process, which is used in many commercial plants around the world, involves the recovery of the precious metals by oxidation pretreatment followed by traditional cyanidation. By consequence the saving in time for the gold and silver dissolution from ore or concentrates is limited by the conventional cyanidation leaching step, a process which requires 48 to 72 hours. It should be remembered that the pressure aqueous pre-oxidation technology is not used to leach the gold but only to make the cyanide ions accessible to the gold in the host mineral.

2. Extraction of gold and silver from the argentopyrite / argentite ore 2.1 Pressure oxidation / cyanidation chemistry

The oxidation of gold and silver is a prerequisite for its dissolution in the alkaline cyanide lixiviant. Pyrite along with arsenopyrite, argentopyrite, sphalerite and covellite are the most common host minerals of gold, silver and electrum. It is important from a process optimization standpoint to understand the behavior of each of these minerals during alkaline pressure oxidation/cyanidation which leads to dissolution/destruction and

New Technology for Recovery of Gold and Silver by Pressure Cyanidation Leaching and Electrocoagulation 73

subsequent liberation of gold and silver which would then be available for the cyanide ions. Pressure oxidation of pyrite and argentopyrite at low temperature (700C) and oxygen pressure (60 lb/in2) involves reactions yielding ferrous ion, sulfate ion, and elemental sulfur as products (Anderson and Nordwick, 1996; Chander and Briceño, 1987).

The primary reactions are:

2FeS2 + 7O2 + 2H2O ń 2Fe2+ + 4SO42- + 4H+ (3)

FeS2 + 2H+ ń Fe2+ + H2S + S0 (4)

Plant Location

Feed Medium CapacityT/d

Operation Data

T (qC) P (psi) t (min) McLaughlin, California USA Ore Acid 2700 180 401 90 São Bento Brazil Concentrate Acid 240 190 230 90 Barrick Mercur Utah, USA Ore Alkaline 790 225 473 70 Getchell USA Ore Acid 3000 210 420 130 Goldstrike Nevada, USA Ore Acid 1360 225 439 75 Goldstrike Nevada, USA Ore Acid 5450 225 400 75 Porgera, Papua New Guinea Concentrate Acid 1350 190 258 180 Campbell Canada Concentrate Acid 70 195 305 120 Barrik USA Ore Acid 2000 220 450 100 Goldstrike Nevada, USA Ore Acid 11580 200 200 140 Santa Fe Pacific GoldNevada, USA

Ore Acid 4000 225 460 50

Sunshine Mining & CoIdaho, USA

Ore Acid 1000 120 90 60

Porgera, Papua New Guinea Concentrate Acid 2700 190 250 150

Table 1. Commercial plants using pressure aqueous pre-oxidation (Adams M.D., 2005; Parga et al. 2007).

Ferrous ions produced by reaction (3, 4) are subsequently oxidized to ferric ions.

2Fe2+ + 1/2O2 + 2H+ ń 2Fe3+ + H2O (5)

And contribute to the leaching of sulfides only as a source of ferric ions. The ferric ions can also contribute to the oxidation of argentopyrite, argentite, pyrite, pyrrhotite, sphalerite and chalcopyrite:

AgFe2S3 + Fe3+ ń Ag+ + 3Fe2+ + 2S2- + S0 (6)

Ag2S + 2Fe3+ ń 2Ag+ + 2Fe2+ + S0 (7)

FeS2 + 8Fe3+ + 4H2O ń 9Fe2+ + 8H+ + S0 + SO42- (8)

Fe7S8 + 32H2O + 62Fe3+ ń 69Fe2+ + 64H+ + 8SO42- (9)

ZnS + 2Fe3+ ń Zn2+ + 2Fe2+ + S0 (10)

CuFeS2 + 4Fe3+ ń Cu2+ + 5Fe2+ + 2S0 (11)

Noble Metals 74

Of course silver has been shown to activate the oxidation and dissociation of chalcopyrite because a porous sulfur layer is formed (Parga et al., 2007). Then, elemental sulfur may also be further oxidized to sulfate by oxygen or by ferric sulfate:

2S0 + 3O2 + 2H2O ń 4H+ + 2SO42- (12)

S0 + 6Fe3+ + 4H2O ń 6Fe2+ + 8H+ + SO42- (13)

This results in the formation of a porous, but nonprotective, elemental sulfur layer, thus allowing cyanide and dissolved oxygen to access to the previously locked gold, silver and electrum. It was thought that by maintaining high concentrations of CN- ions in the pressure leach reactor it would be possible to complex gold and silver (Eq. 14) as they became liberated and thus achieve both objectives, i.e. decomposition of refractory minerals and leaching the Au/Ag with the cyanide ions, simultaneously in a single stage. Also, two of the main advantages of this cyanidation process are the selectivity of free cyanide for gold and silver dissolution and the extremely high stability of the cyanide complex as illustrated in Table 2.

Reaction K Reaction K Au+ + 2CNï=Au(CNሻଶ 1038.3–1039.3 Ag2S = 2Ag++S2ï 10ï53.5*f Au(CNሻଶ = AuCN+CNï 10ï7 a H++S2ï = HSï 1017m Au++2HSï = Au(HSሻଶ 1030.1 b HSï+H+ = H2S 107 Au++HSï = Au(HS) 1024.5 b 2S+HSï+OHï = S2S2ï+H2O 102.2 Au++2OHï = Au(OHሻଶ 1022.0 c 3S+HSï+OHï = S3S2ï+H2O 103.9 Au++OHï = Au(OH) 1020.6 c–1010.2 d 4S+HSï+OHï = S4S2ï+H2O 104.6 Au++OHï+CNï = Au(OH)(CN)ï

1023.3 e 5S+HSï+OHï = S5S2ï+H2O 104.6

Au++2CH3CN = Au(CH3CNሻଶכ

101.6 n 6S+HSï+OHï = S6S2ï+H2O 102.3

Au2S = 2Au++S2ï 10ï72.8*f Cu2++4CNï = Cu(CNሻସଶ 1025 Ag++2CNï = Ag(CNሻଶ 1020.1 Fe3++6CNï = Fe(CNሻଷ 1043.6

Ag++3CNï = Ag(CNሻଷଶ 1021.4–1021.8 3Ag++Fe(CNሻଷ = Ag3Fe(CN)6(s)

1018.2

Ag++2OHï = Ag(OHሻଶ 103.6–104.2 Au++2S2�ଷଶ = Au(S2O3ሻଶଷ 1026 Ag++OHï = AgOH 102.3–103.9 Ag++S2�ଷଶ = AgS2O3ï 108.80 Ag++OHï+CNï = Ag(OH)(CN)ï

1012.8–1013.2 Ag++2S2�ଷଶ = Ag(S2O3ሻଶଷ 1013.7

AgCN = Ag++CNï 10ï15.7* Ag++3S2�ଷଶ = Ag(S2O3ሻଷହ 1014.2

Table 2. Equilibrium constants (Marsden and House, 1960).

For this argentopyrite concentrate, both gold and silver extractions achieved were in fact greater than 96%.

Since silver in Mexican ores occurs as argentite the cyanidation and sulfide oxidation reaction are as follows:


Ag2S + 2nCN- ń 2Ag(CN)(n-1)-n + S2- (14)

Ag2S + 4CN- + 0.5H2O + O2 ń 2Ag(CN)2- + 0.5 S2O32- + OH (15)

S2- + CN- + 1/2O2 + H2O ń SCN- + 2OH (16)

2S2- + 2O2 + H2O ń S2O32- + 2OH- (17)

S2O32- + 2OH- + 2O2 ń 2SO42- + H2O (18)

According to the above equations, sulfide ions are oxidized to thiosulphate and can contribute with the dissolution of gold. Also Graham et al. (Graham and James, 1995)

showed that an alkaline or near neutral solution of thiosulphate dissolved gold metal slowly in the presence of a mild oxidant. The dissolution of the gold can be written as in Equation (19), where oxygen is the oxidant and thiosulphate is the ligand.

4 Au + 8S2O32- + O2 + 2H2O ń 4[Au(S2O3)2]3- + 4OH- (19)

Fig. 1. Schematic mechanism of gold and silver leaching.

Better extractions of gold with thiosulphate are achieved when elevated temperatures (e.g. 65 qC) are used (Wan et al.,1993). The proposed leaching reaction mechanism for gold and

Noble Metals 76

silver dissolution in relatively mildly operating conditions of 70 qC and 60 psi oxygen pressure is described in Figure 1. The role of oxygen gas is to dissolve the easily solubilized sulphides in mildly oxidative pressure and temperature by forming sulphate and metals (e.g. Fe2+ to Fe3+) to liberate gold occluded in the refractory matrix.

2.2 Experimental details

Mineralogical analysis indicates that the sulfide minerals for the Bacís Mining Co. concentrate include pyrite, argentopyrite, pyrrhotite, arsenopyrite, chalcopyrite, covellite, hematite and magnetite. The non-opaque minerals were quartz, calcite, apatite, gypsum, fluorite and barite. Also, several photomicrographs indicate occlusion and/or dissemination of micron-size gold, silver, electrum particles in the sulfide minerals such as argentopyrite, pyrite, sphalerite, arsenopyrite and quartz (Figure 2). The compositions are listed in Table 3.

Fig. 2. Photomicrograph, Native gold, size 5 Ǎm , Argentite (ARG) occluded in pyrite, size 60 Ǎm, Electrum (EL) occluded in pyrite, size 10 Ǎm.

gr./ton %

Au Ag Pb Zn Cu Fe As S

Concentrate 87.09 12320 2.6 3.8 0.5 29.2 0.15 32

Ore 4.12 289 0.5 0.4 0.04 3.67 0.1 3.7

Table 3. Analytical data for the composite samples.

2.2.1 Leaching experiments

Experiments were carried out in a four-liter Parr autoclave assembled with impeller, thermowell, pressure gauge, gas inlet and outlet pipes were used for simultaneous oxidation and cyanidation in the same autoclave. Samples were ground to 84% minus 40 Ǎm or finer, pulped with fresh tap water. Depending on test requirements, variables such as NaCN, reaction time, temperature, particle size, pH and pulp density were set and the leaching experiments were undertaken.


2.3 Results and discussion

2.3.1 Ambient condition

Preliminary batch testing had indicated that at ambient conditions (25qC, 1atm.) direct cyanide leaching of the concentrate with air gave poor silver and gold recoveries. The results obtained are shown in Figures 3 and 4.

2.3.2 Effect of the autoclave retention time

Figure 5 shows that there is an optimum leach time to achieve both maximum gold and silver extractions. The optimum leach time appears to lie close to 60 minutes at 80qC.

Increase in time beyond this value leads to chemical degradation of gold and silver cyanocomplexes, due to the hydrogen ions produced by the sulfide oxidation. Also for prolonged leaching time there is significant loss in cyanide, since the oxygen present oxidizes the cyanide to cyanate and ultimately to ammonia and carbon dioxide.

Fig. 3. Comparison of silver extraction at ambient conditions and high pressure.

2.3.3 Effect of cyanide concentration

Concentrations of 0.6 - 1.1 wt. % sodium cyanide were used to leach the concentrate in the autoclave. When the concentration of sodium cyanide was less than 0.6%, gold and silver extraction could not be completed optimally in a single stage of autoclave leaching. Increasing the cyanide concentration resulted in an increase in the gold and silver extraction. The results obtained are shown in Figure 6.

2.3.4 Effect of temperature

The concentrate was leached as 20-wt.% solids slurry with 1 wt.% of NaCN for 60 minutes. Temperature was varied from 60 to 200qC. As the temperature was increased, the gold and silver extractions decreased, since at temperatures higher than 80qC, the oxidation of cyanide ions is too rapid resulting in the formation of CO2 gas and ammonium ion. Cyanide is then not available for complexetion with the gold and silver. See Figure 7.

CNO- + H2O + 2H+ ń CO2 + NH4+ (20)

0

20

40

60

80

100

0 50 100 150 200 250 300 350

EX

TR

AC

TIO

N, %

TIME (min)

Temp.= 80 oCP= 80 psi

Temp.= 25 oC

P= 1 atm

CONDITIONS%Solids = 20

pH= 11.0rpm = 300

Noble Metals 78

Fig. 4. Comparison of gold extraction at ambient conditions and high pressure.

Fig. 5. Effect of autoclave retention time on gold and silver extraction.

Fig. 6. Effect of the cyanide concentration on gold and silver extraction

0

20

40

60

80

100

0 30 60 90 120 150 180 210 240 270 300

EX

TR

AC

TIO

N, %

TIME (min)

P= 80 psiT= 80 oC

P= 1 atmTemp.= 25 RCCONDITIONS

%Solids= 20pH= 11.0rpm= 300

85

90

95

100

15 30 45 60 75 90 105 120

EX

TR

AC

TIO

N, %

AUTOCLAVE RETENTION TIME (minutes)

GOLD

SILVER

CONDITIONS%Solids = 20pH = 11Pressure = 80psirpm = 300Temp. = 80 oC

50

60

70

80

90

100

0.4 0.6 0.8 1 1.2

CYANIDE CONCENTRATION, %

EX

TR

AC

TIO

N, %

GOLD

SILVER

CONDITIONS% Solids = 20pH = 11

Temp. = 80oC

Pressure = 80 psirpm = 300leach time = 60 min.


Fig. 7. Effect of temperature on gold and silver extraction.

2.3.5 Effect of pH

A pH range of 10.5-13, was examinated for the extraction of gold and silver. In Figure 8, it can be seen that at a pH of 11.5 or above, optimum extractions could not be maintained.

This behavior is because in alkaline-oxidizing medium the pyrite and pyrrhotite in the concentrate competes with gold and silver for consumption of oxygen, and this mechanism may be responsible for the decrease in extraction.

2FeS2 + 7O2 + 2H2O ń 2Fe2+ + 4H+ + 4SO42- (21)

FeS + 2O2 ń Fe2+ + SO42- (22)

Eventually the ferrous sulfate formed is converted to the stable, insoluble ferric hydroxide.

FeSO4 + Ca(OH)2 ń Fe(OH)2 + CaSO4 p (23)

2Fe(OH)2 + 1/2O2 + H2O ń 2Fe(OH)3 p (24)

The decrease in the extraction of gold and silver with increasing pH may also be due to a decrease in the formation of ferric sulfate which is the reagent that contributes to the oxidation of sulfide minerals which occlude gold and silver. Also Deitz (Deitz and Halpern, 1953) found that at high pH values between 12 to 13, the silver surface was coated with a whitish film of CaO2 and thus the leaching process was retarded.

2.3.6 Effect of pulp density

The effect of varying the pulp density in the range of 15-30% solids is shown in Figure 9.

Gold and silver extractions gradually decreased as the pulp density increased apparently due to the increased apparent viscosity of the slurry which impedes good oxygen dispersion in the system, and also for the excessive acid generation.

Noble Metals 80

2.3.7 Plant operation

The optimum conditions as delineated in the laboratory study (Table 4) were used to study the extraction of gold and silver during a continuous one-month plant campaign. A total of 600 tons of concentrate was processed. Recoveries for both gold and silver averaged approximately 96%. These plant data are shown in Figure 10.

On the basis of the results described above the plant tests were in good agreement with the results obtained in the laboratory autoclave and relative to any comparable process, low temperature, low pressure and with these conditions cheaper autoclaves can be used. Thomas shows (Thomas et al., 1992) that a 1400 t/d acid pressure oxidation circuit might cost C$30 million while the corresponding non-acid circuit might cost only C$15 million.

Fig. 8. Effect of pH on gold and silver extraction.

Fig. 9. Effect of percent solids in the extraction of gold and silver.


Solid % = 20 Cyanidation time = 60 min. Temperature = 80qC NaCN = 1% Pressure = 5.6 kg./cm2 (psi) pH = 10.7 rpm = 300 PbO = 100gr./ton

Table 4. Optimum parameters for the autoclave

Fig. 10. Result from continuous plant operation.

3. Electrocoagulation process alternative for silver and gold recovery After the extraction of gold and silver from the argentopyrite/argentite ore, the next step is the recovery of the precious metals from the cyanidation process by which gold and silver are recovered from their ores, and it is recognized that the Carbon in Pulp, the Merrill-Crowe (Metal displacements), the Ion Exchange and Electrowinning processes are used for concentration and purification of gold and silver from cyanide solutions. Each recovery method has advantages and disadvantages. Process selection depends on the specific conditions for a particular operation and the facilities already available. The Merrill-Crowe method had been the preferred process for many years for recovery silver and gold from high reach solutions. Only recently, in the past years, has the carbon adsorption process become popular for recovering gold from large volumes of low grade pregnant leach solutions that contain mainly gold. Other processes, Ion-Exchange Resins and Solvent Extraction, have recently been reviewed as an alternative for gold and silver recovery from alkaline cyanide solutions, (Aguayo et al., 2007). Commercially available resins were unable to compete with activated carbon due to poor selectivity, mechanical breakdown of the beads and the requirement for complex elution, generation of HCN and regeneration process.

Among several available options for recovery of precious metals from cyanide solutions, Electrocoagulation (EC) is a very promising electrochemical technique that does not require high concentrations of silver and gold in cyanide solutions in order to recovery them. Also, literature review showed that the potential of EC as an alternative to traditional treatment recovery of precious metals (silver and gold cyanide) has not yet been exploited. Advantages and disadvantages for the different processes are presented in Table 5, along with those of the EC (Mollah et al., 2004; Parga et al., 2005). EC has been proposed since before the turn of the 20th century. A plant was built in London in 1889

Noble Metals 82

Method Advantages Disadvantages

Merrill-Crowe

x Lower capital and operating costs.

x Handles solutions containing high silver and gold content.

x It is highly efficient (99.5%). x Also can treat high-grade

solutions produced by carbon elution.

x Alternative to Electrowinning. Well known technology.

x The pregnant solution need clarification and deoxygenating.

x Low concentrations of metals, increases amount of zinc

x Depends on the pH and concentration of the free cyanide.

x The precipitate contain cyanides like copper and arsenic.

x The precipitate spend one week in the filter press.

Adsorption with Activated Carbon

x Does not require pre-treatment of pregnant solution.

x Not dependent on the concentration of metals.

x Large specific surface. x The pulp needs no clarification.

x Fouled carbon needs to be regenerated by heating.

x Large carbon inventory. x The pregnant solution has

to go through 5 or 6 columns.

x High operating costs.

Ion Exchange Resins

x Does not need: washing, revitalization or heat treatment.

x High abrasion resistance in tanks of adsorption.

x High selectivity.

x High cost of the process. x Lower loading capacity. x Royalty payments. x The resin must be

regenerated in acid medium.

Electrocoagulation

x Low residence time (minutes). x Does not use chemicals. x Handles solutions containing

lower or high silver and gold contents.

x Energy costs per m3 of pregnant solution are lower than conventional treatment systems.

x Sacrificial anode must be placed periodically.

x Precise initial pH control. x New technology. x The product is high in iron. x The sludge need to be

leaching with sulfuric acid.

Table 5. Advantages and disadvantages of methods for recovery of gold and silver (Emamjomeh et al. 2004; Mollah et al., 2009; Parga et al., 2007).

for the treatment of sewage mixing it with seawater and electrolyzing it. In 1906, EC was first patented (Parga et al., 2005) and used to treat bilge water from ships. In 1909, in the United States J.T. Harries received a patent for wastewater treatment by electrolysis using sacrificial aluminum and iron anodes (Vik et al., 1984). Matteson (Matteson et al., 1995), described a device of the 1940’s, the “electronic coagulator” which electrochemically

dissolved aluminum (from the anode) into solution, reacting this with the hydroxyl ion

(from the cathode) to form aluminum hydroxide. The hydroxide flocculates and coagulates


the suspended solids and thereby purifies water. A similar process was used in Britain in 1956 in which iron electrodes were used to treat river water (Holt et al., 2005).

3.1 Electrocoagulation fundamentals

The electrochemical phenomenon of Electrocoagulation has been employed previously for treating many types of wastewater with varying degrees of success. This electrochemical method of contaminant removal requires smaller quantities of salt addition to increase the conductivity of the solution (typically, an aqueous electrolye), and the maintenance and operation of the EC cells are relatively simple. Electrocoagulation processes offer significant potential for removing ionic species from solution, particularly heavy metals (Mollah et al., 2004; Parga et al., 2007). Operating conditions are highly dependent on the chemical composition and properties of the aqueous medium, specifically, conductivity and pH. Other important process variables, such as particle size, type of electrodes and retention time between electrodes, electrode spacing and chemical-constituent concentrations influence the operating process-parameters (Parga et al., 2007). The fundamental operating-principle is that cations produced electrolytically from iron and/or aluminum anodes (by oxidation) provide for the coagulation of contaminants contained in the aqueous electrolyte. Thus, the (sacrificial) metal-anodes provide a continuous supply of polyvalent metal cations in the vicinity of the anode. These cations participate in the coagulation process by neutralizing the negatively-charged ions (anions) that are transported toward the anode by electrophoresis. In a continuous-flow EC system, the production of polyvalent cations from the oxidation of the sacrificial anodes (Fe and/or Al) and the evolution of electrolysis gases (H2 at the cathode and O2 at the anode) are directly proportional to the current (charge) supplied according to Faraday’s Law of Electrolysis. The evolved gases enhance the

flocculation of the coagulant species.

A schematic of the electrocoagulation process for recovery gold and silver is shown in

Figure 11. The gas bubbles produced by electrolysis convey the gold and silver species to the

top (free-surface) of the electrolyte where it is concentrated, collected and removed. The

removal mechanisms in EC may involve oxidation, reduction, decomposition, deposition,

coagulation, absorption, adsorption, precipitation and flotation.

However, it is the reaction involving the metal ions that enhance the formation of the

coagulant. The metal cations react with the OH- ions produced at the cathode during the

evolution of hydrogen, to yield both soluble and insoluble hydroxides that will react with or

adsorb pollutants, respectively, from the solution and also contribute to coagulation by

neutralizing the negatively charged colloidal particles that may be present at neutral or

alkaline pH. This enables the particles to approach closely and agglomerate under the

influence of Van der Waals attractive forces. Depending on the pH range, the electrode

reactions that have been proposed to describe EC mechanisms for the production of H2(g),

OH� (cathode) and H+ and O2(g) (anode) are (Moreno et al., 2009):

i. pH <4

Anode:

Fe2+ + 2e- = Fe0 (25)

Noble Metals 84

Fig. 11. An illustration of the EC mechanism (arrow indicate the migration of ions, the H2

evolution and the formation of green rust).

Cathode:

2H++ 2 e-oH2(g)Ń (26)

ii. pH 4 <pH< 7:

Anode: as before, reactions (25) and (26). Furthermore, iron also undergoes hydrolysis:

Fe + 6H2O o Fe(H2O)4(OH)2 (aq) + 2H+ + 2e- (27)

Fe + 6H2O o Fe(H2O)3(OH)3 (aq) + 3H+ + 3e- (28)

Fe(III) hydroxide begins to precipitate as a floc with yellowish color.

Fe(H2O)3(OH)3 (aq) o Fe(H2O)3(OH)3(s) (29)

“Rust” can also be formed.

2Fe(H2O)3(OH)3 l Fe2O3 (H2O)6 (30)

Cathode:

2H++ 2 e-o H2(g)Ń (31)


Additional hydrogen evolution takes place, but [H+] now comes from iron hydrolysis.

iii. pH 6 <pH< 9:

Anode: reactions (25) and (26). Precipitation of Fe(III) hydroxide occurs according to Reaction (30) simultaneous with Fe(II) hydroxide precipitation whereby a dark-green floc is produced.

Fe(H2O)4(OH)2 (aq) o Fe(H2O)4(OH)2 (s) (32)

The minimum solubility of iron hydroxides, Fex(OH)y, occurs in the pH range of 7–8. EC flocs are formed due to the polymerization of iron oxyhydroxides. Formation of rust (dehydrated hydroxides) occurs according to the following:

2Fe(OH)3 ņ Fe2O3 + 3H2O (hematite, maghemite) (33)

Fe(OH)2 ņ FeO + H2O (34)

2Fe(OH)3 + Fe(OH)2 ņ Fe3O4 + 4H2O (magnetite) (35)

Fe(OH)3 ņ FeO(OH) + H2O (goethite, lepidocrocite) (36)

The species hematite, maghemite, rust magnetite, lepidocrocite and goethite have been identified as EC products by (Parga et al. 2005).

Cathode:

H2O + 2e- = H2(g) + OH- (37)

Overall reactions are:

Fe + 6H2O o Fe(H2O)4(OH)2 (s) + H2(g)Ń (38)

Fe + 6H2O oFe(H2O)3(OH)3 (s) +1 ½ H2(g)Ń (39)

The concentrations of the various species within the cell are not uniform; in addition the concentration of species such as iron and hydronium ion are also time dependent. Typically, the EC process employs bipolar electrodes (Parga et. al. 2005). It has been demonstrated that with this configuration where the electrodes are connected in series, and consequently low current-densities are present, iron (or aluminum) coagulant is produced more effectively, at higher rates and more economically compared to chemical coagulation (Parga et. al. 2005).

Also, in the electrocoagulation cell for the high voltage we produce a very strongly oxidizing environment around the anode and this is suitable for destroying strong cyanide solutions (greater than 1000 ppm) and is a direct oxidation of the cyanide ion at the anode to cyanate ion which is further decomposed to carbon dioxide and nitrogen, ammonium, and carbonate or oxalate ions according to the pH (Hwang et. al., 1987). The reactions are as follows:

In strong alkaline solution (pH=12):

CN- + 2OH- o CNO- + H2O + 2e (40)

Noble Metals 86

CNO- + 2OH- o CO2 +1/2N2 +H2O + 3e (41)

In neutral and weak alkaline solution (pH = 7.0-11.7):

2CN o C2N2 (42)

C2N2 + 2OH- o CNO- + CN- + H2O (43)

CNO- + 2H2O o NH+4 + CO2-3 (44)

In weak acidic solution (pH = 5.2-6.8):

C2N2 + 4H2O o C2O2-4 + 2 NH+4 (45)

3.2 Experimental details

The EC experiments were performed with a Fisher magnetic stirrer and a 400ml beaker size reactor equipped with two carbon steel electrodes (6 cm x 3 cm) that were 5 mm apart. As a source of current and voltage a universal AC/DC adaptor was used. pH was measured with a VWR scientific 8005 pH meter and electrodes were properly scrubbed and rinsed prior to each experiments to ensure a clean surface free from passive oxide layers . Gold and silver adsorption onto iron hydroxide species was investigated with pregnant cyanide solutions provided by Bacis S.A. de C.V mining group (13.25 mg L-1 Au, 1357 mg L-1 Ag, 200 mg L-1 free CN- and 1400 mg L-1 total CN- and pH of 8). Analysis were performed by ICP/Atomic Emission Spectrometry (Perkin Elmer 3100). The conductivity of pregnant solutions was adjusted by adding one gram of NaCl per liter (Fisher, 99.8% A.C.S. Certified, lot #995007).To identify and characterize the iron species in the solid products, formed during the EC process for the removal of gold and silver using iron electrodes, X-ray diffraction (XRD) (Phillips model X-PERT), FT-IR analysis were carried out by Thermonicolate FTIR spectrometer and Scanning Electron Microscope (SEM / EDX) (FEI Quanta 2000, Oxford Instruments) were used.

Analysis of Au and Ag were conducted to the Bacis solution, by AES. EC was run at 15 Volts (DC) and the corresponding current was of 0.1 A. EC was run for five minutes, and a sample was taken every minute in order to determinate the removal efficiency for Au and Ag. Solutions and solids from the EC process were separated by filtration through cellulose filter paper. The sludge from the EC was dried either in an oven or under vacuum at room temperature and characterized. The experimental set-up is presented in Figure 12. The current and voltage during the EC process were measured and recorded, using Cen-Tech multimeters. The pH values of the solution before and after EC were measured with a VWR scientific 8005 pH meter.

The resulting sludge of iron hydroxide gel precipitate with Au/Ag is filtered. Then, this rich sludge is treated in an acid leach step with sulfuric acid under oxidizing conditions caused by the addition of air. The conditions of this acid leach are such that the major portion of iron and copper are leached into solution with gold and silver remaining in the residue. The resultant residue from the filtration step is gold and silver, as well as little iron (10 % Au, 80 % Ag and 5 % Fe) suitable for further refining by conventional commercial method.


Fig. 12. Experimental set-up of the EC process for the removal of Ag and Au.

3.3 Results and discussion

With the optimal parameters for cell work of EC, they were tested for removal efficiency of gold and silver, the table 6 shows the results when making treatment of the solution containing gold and silver and also the removal efficiency of cyanide, as shown in Figure 13. A maximum in gold and silver recovery were achieved at 5 minutes of treatment and as the arithmetic average of five replications; achieved an efficiency of 99.24% Au to 99.93% Ag for both, with a standard deviation of 0.26 in gold and 0.06 in silver. Also, this studies shown a very efficient recovery in the range of 2 to 3 minutes for gold and 1 to 2 for silver, this occurs in the pH range from 9 to 11 approximately, which coincides with the production of the magnetic iron, Fe3O4, which has magnetic properties that accelerates the process of adsorption of metals, the adsorption rate is then physically, because it is caused by the magnetic forces of the magnetite into gold and silver these forces without altering their chemical composition. Also, it is likely that the electrocoagulation cell is oxidizing the Au and Ag cyanide complexes and converting them to a less-soluble form that is captured by the iron hydroxide gel. The high voltage in the EC cell around the anode destroys some of the cyanide.

About the same studies, Figure 14 shows graphically the evolution of pH during the operation time, there was an increase in pH of the solution which is attributed to the evolution of hydrogen at the cathode which is accompanied by alkalinization of the aqueous solution. The final effect is the oxidation suffered by the water coupled with the generation of hydroxyl ions generated during EC.

Noble Metals 88

EC residence time (min)

Au (mg L-1)

Recovery (%)

Ag (mg L-1)

Recovery (%)

pH Removal of

Cyanide (mg L-1)

Removal of Cyanide

(%) 0 13.25 0 1357.0 0 8.0 1400 0 1 12.50 5.66 1240.0 8.62 9.2 1050 25 2 10.50 20.37 219.5 83.82 9.5 870 38 3 1.00 92.45 9.0 99.33 10.7 750 46 4 0.50 96.22 7.0 99.48 11.2 400 71 5 0.10 99.24 0.9 99.93 11.5 210 85

Table 6. Recovery of gold and silver by EC.

Fig. 13. Gold and silver recoveries from Bacis cyanide solutions.

Fig. 14. Variation of pH vs. EC residence time and Cyanide removal vs time.


This results shown that in the EC cell with the iron electrodes decompose the content of cyanide in the pregnant reach solution of gold and silver from initial cyanide content of 1400 to 210 mg L-1. It was found that (Tamura et al., 1974) the anodic oxidation of cyanide is proportional to the alkalinity of the electrolyte and consistent with the following mechanism:

CN- + 2OH o CNO- + H2O + 2e- (46)

2CNO- + 4OH- o 2CO2 + N2 + 2H2O + 6e (47)

CNO- + 2H2O o NH4 + CO-3 (48)

3.3.1 Product characterization

X-ray Diffraction Analysis. Diffraction patterns of flocs collected from the experiment with gold and silver, (the sample were ground to a fine powder and loaded into a sample holder) were obtained with a diffracted X-PERT Phillips meters equipped with a vertical goniometer, with a range of analysis 2Ĳ 10 ° to 70 °. The source of X-rays has a copper anode, whose radiation is filtered with a graphite monochromator (ǌ = 1.541838Å) with scan rate of 0.02 ° and a duration of 10 seconds per count. The X-Ray Diffractometer is controlled by a Gateway 2000 computer, by PC-APD 2.0 with software for Windows.

Figure 15 shows the ray diffraction pattern of the flocs recovered from the sample of gold and silver, respectively 13.25 mg/L and 1357 mg/L, initial pH 8, 5 minutes of treatment, 0.1 amperes and 15 volts. The species identified were magnetite, lepidocrocite, goehtite, silver and copper hexacyanoferrate.

Fig. 15. X-ray diffractogram of solids obtained in the recovery of gold and silver. C: Cu2Fe (CN) 62H2O A: Silver, M: Magnetite and L: Lepidocrocite.

Fourier Transform Infrared Spectroscopy. FT-IR analysis were carried out by Thermonicolate FTIR spectrometer and OMNIC software using potassium bromide pellets (sample: KBr =1: 50). The spectra were usually recorded in the range of 4000-400 cm-1 with 2 cm-1 resolution. 64 scans were collected for each specimen. Figure 16 shows the FT-IR spectrum of the by-product. Infrared analysis of iron electrode by-product showed OH

Noble Metals 90

stretching at 3738 and 3447 cm-1, hydroxyl bending and J`(OH) water bending vibration or overtones of hydroxyl bending around 1637 cm-1 . Bands for lepidocrocite phase showed up at 1120, 1023, and 745 cm-1. Magnetite (Fe3O4 or Fe3-xO4) band at 575 cm-1and Fe-O vibration band is seen at 469 cm-1. For details of FT-IR analysis see Table 7.

XRD analyses also confirmed the presence of these species detected by FT-IR.

Fig. 16. FT-IR spectrum of iron electrode by-product.

Electrode Material

Type of Vibrations Vibration

wavenumbers (cm-1) Vibration Range

(cm-1)

Iron

OH stretching 3738 3689-3787 3447 3550-3000

Hydroxyl bending 1637 1572-1813 J`(OH) water bending 1637 1572-1813 Overtones of hydroxyl bending

1637 1572-1813

Magnetite (Fe3O4 or Fe3-

xO4) 575 526-840

Fe-O 469 416-510

Lepidocrocite 1120 1090-1245 1023 923-1057 745 730-790

Table 7. FT-IR vibrations and their corresponding wave numbers and region for the bands observed for the EC-byproduct

Scanning Electron Microscopy (SEM/EDAX). Figure 17 shows SEM images and EDAX of silver adsorbed on iron species. These SEM and EDAX results show that the surfaces of these iron oxide/oxyhydroxide particles were coated with a layer of silver. It is worth clarifying that, given the low concentration of gold it was impossible to locate any nanoparticle of it.


Transmission Mössbauer Spectroscopy. Figure 18 shows the spectrum obtained from the EC silver, gold and iron solid product from 1350 mg/L of silver and 13 mg/L of gold cyanide solutions at pH = 11.0, Mössbauer Spectra for each sample was obtained on a ±15 mm/s velocity scale, which allows observation of wide magnetic hyperfine spectra expected from iron oxide compounds. The spectrum consists of a doublet magnetic spectrum, which is probably due to fine particles of iron oxides (non-stoichiometric magnetite) or iron hydroxides (Lepidocrocite, Goethite, etc.).From the analysis of these techniques the in-situ generated small fine particles of iron- oxide/oxyhydroxides in the EC process are: non-stoichiometric magnetite, goethite and iron hydroxide oxide.

Fig. 17. Chemical composition of solid product as determined by EDX, which shows the presence of silver in the particle of iron.

Fig. 18. Mössbauer spectrum with silver and gold at pH=10.5, indicating the presence of magnetite species.

Noble Metals 92

4. Conclusions

This study has been very valuable in identifying that, gold and silver values are associated with argentopyrite, pyrite, sphalerite and chalcopyrite in the Bacís concentrate. The kinetics of the direct pressure oxidation/cyanidation was found to be strongly dependent on particle size, concentration of sodium cyanide, oxygen pressure, temperature and pH. Single stage direct pressure oxidation/cyanidation, has proven to be effective in treating argentopyrite refractory gold and silver concentrates from the Bacís mining and processing operations, for both gold and silver it was found that the precious metals recovery exceeded 95%.

Also in this process, because of the short leaching time the inventory of gold and silver is reduced. The relatively mild operating conditions of 80 qC and 80 psi oxygen pressure offer distinct advantages. For example, low cost materials of construction can be utilized for the autoclave. Finally the pressure oxidation cyanidation process is flexible and can accommodate gold and silver ore of different mineralogical composition and origin.

In addition it has been shown that Eletrocoagulation is an interesting process for the recovery of gold and silver from the cyanide leach solution that is yet to be fully realized. EC comprises complex chemical and physical processes involving many surface and interfacial phenomena. Also, the results of this study suggest that EC produces magnetic particles of magnetite and amorphous iron oxyhydroxides, and that this process can be used to remove gold and silver cyanide ions. The results of this study indicate that silver and gold can be successfully adsorbed on iron species produced by the Electrocoagulation process. So EC may be used to recover gold and silver from cyanide solutions.

The X-Ray Diffraction, FT-IR analysis and Scanning Electronic Microscopy techniques demonstrate that the formed species are of magnetic type, like lepidocrocite and magnetite, and amorphous iron oxyhydroxide which adsorbed the silver and gold particles on his surface due to the electrostatic attraction between both metals.

The 99.5% of gold and silver were removed in the experimental EC reactor, and it was achieved in 5 minutes or less with a current efficiency of 99.7%. Finally, the high voltage in the EC cell around the anode destroys some of the cyanide and this process can be accelerated in the presence of copper ions.

5. Acknowledgment

The author thanks the management of Bacís Mining Co., Williams Mining Co., CONACYT and DEGEST for the support and permission to publish this chapter and appreciation is extended to Professor Jan D. Miller of University of Utah and Gerard P. Martins of Colorado School of Mines for their interest in this research.

6. References

Adams, M.D. (2005). "Advances in gold ore processing” Elsevier Inc. 346-369, ISBN 9780444517302, Amsterdam Netherlands.


Aguayo, S., Valenzuela, J.L., Parga, J.R. & Lewis, R.G. (2007). “Continuous laboratory gold solvent extraction from cyanide solutions using LIX79 reagent”. Chem. Eng.

Technol., 30, 1532-1536. Anderson, C.G. & Nordwick, S.M. (1996). "Pretreatment using alkaline sulfide leaching and

nitrogen species catalyzed pressure oxidation on a refractory gold concentrate". EPD Congress the Mineral and Materials Society., 323-341.

Chander, S. & Briceño, A. (1987). "Kinetics of pyrite oxidation". Minerals and Metallurgical

Processing., 8, 170-176. Deitz, G.A. & Halpern, J. (1953). "Reaction of silver with aqueous solutions of cyanide and

oxygen". Journal of Metals., 5, 1109-1116. Emamjomeh, M.M. & Sivakumar, M. (2009). “Review of pollutants removed by

electrocoagulation and electrocoagulation /flotation processes”. Int. J. Environ.

Manage., 90, 1663-1679. Graham, J.S. & James, T.W. (1995). "Cyanide and other lixiviant leaching systems for gold

with some practical applications". Mineral Processing and Extractive Metallurgy

Review, 14, 193-247. Holt, P., Barton, G. & Mitchell, C. (2005). “The future of EC as a localized water treatment

technology”. Chemosphere 9, 13, 335-367. Hwang, J. Y., Wang, Y. Y. & Wan, C.C., (1987). “Electrolytic oxidation of cuprocyanide

electroplating wastewaters under different pH conditions”, J. Appl. Electrochem, 17, 684-694.

Marsden J. O. & House, C.L. (2006). “The chemistry of gold extraction”, Second edition, Society of Mining, Metallurgy and Exploration, 237-238, ISBN 9780873352406. Littleton, Colorado, USA.

Matteson, M. & Dobson, R, (1995). “Electrocoagulation and separation of aqueous suspensions of ultrafine particles”. Colloids and Surfaces A: Physicochemical and

Engineering Aspects, 104, 1, 101-109. Mollah, M., Morkovsky, P., Gomez, J., Parga, J.R., & Cocke, D. (2004). “Fundamentals,

present and future perspectives of electrocoagulation”. J. Hazard. Mat., B1, 14, 199-210.

Moreno, H., Cocke, D. L., Gomes, J. A. G., Morkovsky, P., Parga, J. R., Peterson, E. & Garcia, C. (2009). “Electrochemical reactions for electrocoagulation using iron electrodes”. Ind. Eng. Chem. Res., 48, 2275–2280.

Parga, J.R., Cocke, D.L., Valenzuela, J.L., Kesmez, M., Gomes, J.A.G. & Valverde, V. (2005). “As removal by EC technology in the Comarca Lagunera Mexico”. Journal of

Hazardous Materials, 124, 1-3, 247-254. Parga, J.R., Valenzuela, J.L & Cepeda, F. (2007). “Pressure cyanide leaching for precious

metals recovery”. Journal of Metals, 10, 43-47. Senanayake, G. (2008). “A Review of effects of silver, lead, sulfide and carbonaceous matter

on gold cyanidation and mechanistic interpretation”, Journal Science Direct, 2008, 90, 46-73.

Tamura, H., Arikado, T., Yoneyama, H. & Matsuda, Y. (1974). “Anodic oxidation of potassium cyanide on platinum electrode”, Electrochim.Acta, 19, 273.

Thomas, K.G, Burns, D. & Hill, A.R. (1992). "Autoclaving technology and barrick gold", (Paper prepared for the International Precious Metals Institute, Reno, Nevada), 11.

Noble Metals 94

Vik, E.A., Carlson, D.A., Eikun, A.S. & Gjessing, E.T. (1984). “Electrocoagulation of potable water”. Water Research, 18, 11, 1355-1360.

Wan, R.Y., Le Vier, M. & Miller, J.D. (1993). "Research and development activities for the recovery of gold from non-cyanide solutions". Hydrometallurgy fundamentals,

Technology & Innovations. (Society for Mining, Metallurgy & Exploration), 415-436.

Uso de Resina de Intercambio Aniónico para la Recuperación del Complejo Oro Chaparro

Información Tecnológica Vol. 23 Nº 2 - 2012 53

Uso de Resina de Intercambio Aniónico para la Recuperación del Complejo Oro Tiosulfato desde Soluciones Acuosas Miriam E. Chaparro(1), Jesús L. Valenzuela(2), Guillermo T. Munive(2) y José R. Parga(3)

(1) Instituto de Ingeniería, Universidad Autónoma de Baja California, Blvd. Benito Juárez y Calle a la Normal S/N, Col. Insurgente Este, Mexicali, B.C.-México. ([email protected]). (2) Departamento de Ingeniería Química y Metalurgia, Universidad de Sonora, Rosales y Blvd. Luis Encinas, Col. Centro, Hermosillo, Sonora-México. ([email protected]). (3) Departamento de Metalurgia y Materiales, Instituto Tecnológico de Saltillo, Blvd. Venustiano Carranza # 2400, Saltillo, Coahuila-México. ([email protected]).

Recibido Ago. 26, 2011; Aceptado Nov. 02, 2011; Versión Final recibida Dic. 26, 2011

Resumen

Se estudió la adsorción de oro utilizando la resina AuRIX®100 en medio tiosulfato de amonio, evaluando algunas variables que afectan la cinética del proceso tales como: temperatura, velocidad de agitación, pH, concentración de tiosulfato de amonio (NH4)2S2O3 y concentración de oro. El estudio se llevó a cabo en un reactor batch y una columna de intercambio iónico. Las condiciones de operación que presentaron mejores resultados de extracción de oro a 25ºC, fueron: pH=10.5, velocidad de agitación=500 rpm, [Au]=1 mg/l, [(NH4)2S2O3]=0.04 M, 5 gramos de resina. Los resultados indican que al aumentar [(NH4)2S2O3] favorece la adsorción en un 99% durante tres horas, siendo afectado notablemente por la presencia de amonio. Las condiciones con mejores resultados en la columna a 25ºC fueron, pH= 10.5 y [Au] =1 mg/l.

Palabras clave: adsorción oro, tiosulfato, resina AuRIX®100, cinética

Use of Anion Exchange Resin for the Recovery of the Complex Gold Thiosulfate from Aqueous SolutionsAbstract

The adsorption and elution of gold in thiosulfate-ammonia media were studied using the resin AuRIX®100, evaluating some variables that affect the kinetics of the process, such as: temperature, stirring speed, pH, thiosulfate concentration (NH4)2S2O3 and gold concentration. The study was carried out in a batch reactor and an ion exchange column. The operation conditions that presented better results of gold extraction at 25ºC were: pH=10.5, stirring speed=500 rpm, [Au]=1 mg/l, [(NH4)2S2O3]=0.04 M, and 5 grams of resin. The results indicate that by increasing [(NH4)2S2O3] favors the adsorption by 99% during three hours, being noticeably affected by the presence of ammonia. The conditions with results on the column at 25°C were, pH= 10.5, [Au]=1 mg/l.

Keywords: gold adsorption, thiosulfate, AuRIX®100 resin, kinetics

Información Tecnológica Vol. 23(2), 53-60 (2012) doi: 10.4067/S0718-07642012000200007

Nicolas Martinez


Información Tecnológica Vol. 23 Nº 2 - 201254

INTRODUCCIÓN

La mayoría de las plantas minero metalúrgicas utilizan el proceso de cianuración para la extracción de oro, pero el principal problema es su alta toxicidad, por lo que en recientes años, se han desarrollado métodos alternativos para ello, como es el caso del tiosulfato de amonio y tiourea, que son tan efectivos como el cianuro, pero que aun no han logrado aplicaciones industriales ya que presenta dificultades inherentes a la química de la solución y los métodos para su recuperación han limitado su progreso (Aylmore et al, 2001). Además, las disoluciones de lixiviación han sido realizadas por Adsorción en carbón activado, extracción por solventes, precipitación por el proceso Merrill-Crowe, electrodeposición y resinas de intercambio iónico. Estudios recientes (Chaparro, 2008), demostraron la viabilidad de la extracción del complejo oro tiosulfato, tratándolo con guanidina sobre una resina de intercambio aniónica. Por otro lado, la lixiviación de oro desde concentrados y/o minerales usando disoluciones de tiosulfato y amoniaco han sido estudiadas como alternativas al proceso de cianuración. Sin embargo, uno de los problemas asociados al uso del tiosulfato es la recuperación de oro desde las disoluciones obtenidas ya que son catalizadas por cobre y amoniaco, presentando un posible mecanismo de reacción, como el observado en las ecuaciones 1 y 2.

4NH+)OCu(S+)O Au(SO4S+)Cu(NH+Au 3-3

232-3

232-2

32243 → (1)

4OH+O8S+)4Cu(NH6NH+O2H+O+)OCu(S --232

+243322

-3232 → (2)

Con respecto al uso de resinas para la recuperación de oro desde disoluciones acuosas con tiosulfato y amoniaco, existen estudios en literatura (Molleman et al, 2002), donde utilizan resinas de base débil y la especie estable que se adsorbe es el complejo oro-tiosulfato, existiendo aniones en la disolución acuosa, como sulfito, tritionato y tetrationato que son adsorbidos por la resina, provocando una disminución en el complejo .)OAu(S -3

232 Del mismo modo, en investigaciones realizadas (Aguayo, 2007) sobre la extracción por solventes para Ag y Au en soluciones alcalinas provenientes de cianuración, se utilizó como extractante LIX 79 en base guanidina e igualmente que AuRIX®100 en soluciones de tiosulfato y como diluyente keroseno, determinando que el complejo oro-cianuro, se extrae a 10.5 < pH < 11.2. Al agregar tridecanol 5% en volumen, obteniendo como resultado una adecuada selectividad con 10% LIX 79. Lo anterior, fue generado a partir de licores de lixiviación en columna con minerales oxidados y sulfurados pero conteniendo Ag y pequeñas cantidades de Au, Cu y Zn. Sin embargo, es necesario realizar una exhaustiva investigación utilizando el complejo oro-tiosulfato. Así mismo, en investigaciones de adsorción de Au (Navarro et al, 2006), utilizando resina Amberlita IRA-410 en medio tiosulfato-amoniaco, se demostró que la adsorción de Au es rápida, debido a la presencia de tiosulfato esté desfavoreció la adsorción; sin embargo, este aspecto no queda muy claro. De tal manera, que estudios realizados para evaluar la eficiencia de la resina AuRIX®100 (Valenzuela et al, 2006), para la recuperación del complejo oro-cianuro a pH 10.7, se determino que al aumentar la relación sólido-líquido la cantidad de Au por unidad de masa de resina aumenta proporcionalmente. Durante la adsorción de Au en carbón activado, se ha discutido que el complejo oro-tiosulfato no es adsorbido eficientemente, pero no obstante a ello (Vargas et al, 2006), ha reportado que la presencia de tiosulfato, amoniaco e impurezas (Cu, Zn), son perjudiciales para la adsorción de Au ya que la Ea= 9.13 kJ/mol; por lo contrario, el complejo oro-cianuro si es adsorbido por el carbón activado. Por otro lado, se han evaluado materiales (Seob et al, 2010) para la adsorción de Au obteniendo 427.77 mg/g con resina AmberjetTM4400, 170.64 mg/g en carbón activado y 361.76 mg/g en medio biosorbente. Por otra parte (Munive et al, 2011), se realizaron estudios comparativos de lixiviación sobre un mineral refractario conteniendo cantidades de sulfuros, pirita y pirrotita, en presencia de cianuro y tiosulfato, donde este último permanece estable, pero con la finalidad de recuperar Au y Ag durante 48 hrs y ajustando el pH menor a 9.6 con solución de NaOH y Ca(OH)2. Resultados obtenidos (Breuer et al, 2002), donde el tiosulfato disuelve al cobre presente en el mineral, se muestra que favorece al proceso, ya que actúa como oxidante, provocando que la disolución de Au y Ag disminuya, atribuido a la inestabilidad del complejo cuprotetramina (Yen et al, 1998) a pH=9.5. De lo anterior, se puede observar que existe un gran interés por proponer lixiviantes alternativos al cianuro para la adsorción de Au, con compuestos



menos tóxicos y buscando técnicas viables que permitan obtener una máxima adsorción de metales precisos. En la Tabla 1. Se enlistan las condiciones de lixiviación de diferentes minerales. Observando que la resina AuRIX®100 contiene el componente activo guanidina, utilizada para soluciones de lixiviación del complejo oro-tiosulfato de amonio, constituida por una red de macro retículas de estireno y divinil-benceno que funcionalizada con grupos guanidina actúan como bases orgánicas fuertes.

Tabla 1. Comparación de parámetros de lixiviación en minerales y avances en procesos de adsorción de oro

Antecedentes Tipo de Mineral T (°C)

pH Ret (h) )M(

OS -232 )M(

NH3 % de Rec.

Yen et al, (1998) Au-Cu-0.36% Cu, Au: 7.2-7.9 g/ton

A 10 24 0.5 6 95-97

Molleman et al, (2002) Concentrado de Pirita 35 10-11 6 0.05 0.4 83 Navarro et al, (2002) Conc. de Au: 95 g/ton 25 9-10 10 0.3 - 94 Breuer et al, (2002) Cu(II), Au 30-50 8-9.4 - 0.2 0-0.6 -

Ficeriová et al, (2002) Conc. de CuPbZn 25 6-7 2 0.5 - 99 Zhang et al, (2002) 20 mg/l, Au: 2 g/ton 23-25 11 24 0.1 0.2 80-90

Valenzuela et al, (2006) Au: 0.32, Ag: 0.43,Cu: 6.2

25 8-10.7 3 - - 98

Navarro et al, (2006) Au: 8 mg/l 25 9-11 3 0-0.5 0-0.5 92-98 Vargas et al, (2006) Au: 10 mg/l, 2.2 g/ton 25 9.5 4 0-0.50 1-4 15-28 Jeffrey et al, (2010) 10 mg/l 25-60 9 24 - 5-100 nM 2.7 Seob et al, (2010) Au 25±2 1-6 2-3 - 0.1 84-96

Munive et al, (2011) Au: 3.15 g/ton 25 9.5 48 0.25 0.4 70-82

Por lo anteriormente expuesto, en este trabajo se propone estudiar el proceso de adsorción de Au, mediante pruebas batch y en columna, evaluando la eficiencia de la resina de intercambio aniónico AuRIX®100 para la adsorción del complejo oro-tiosulfato en medio amoniaco-tiosulfato, con el objetivo de obtener los parámetros cinéticos que intervienen en la reacción, evaluando el contenido de Au a un tiempo determinado, dando seguimiento a la reacción en las soluciones por absorción atómica con la finalidad de establecer un proceso como método alternativo para el uso de las resinas de intercambio iónico.

METODOLOGÍA

El método seguido para el desarrollo del presente estudio, consistente en la extracción de oro mediante el uso de guanidina, se utiliza como componente activo guanidina en la resina AuRIX®100, determinando además la influencia de las variables que intervienen en la reacción durante la extracción del metal.

Extracción de oro utilizando una resina AuRIX®100, pruebas batch

En un reactor batch Kettles marca Pyrex, capacidad 500 ml, con disolución acuosa de 250 ml, con parámetros experimentales: Tiempo de contacto= 3 h, temperatura= 25ºC, [Au]i= 1-8 mg/l, razón resina/disolución= 20000 mg/l, velocidad de agitación= 200-800 rpm. Evaluando el porcentaje de oro adsorbido en función del tiempo. Las variables se presentan en la Tabla 2. Agregando al reactor la disolución acuosa, [Au]i, [(NH4)2S2O3], ajustando la temperatura y pH, se agregó la resina a la disolución y se agitó el sistema, se sacaron muestras de la fase acuosa de 10 ml cada 30 min para análisis, se midió el pH, en 3 h se detuvo la agitación, se separaba la resina de la disolución acuosa y para su análisis por la técnica de absorción atómica en un espectrofotómetro Perkin Elmer 3110 y la carga de oro en la resina se determinó por balance de oro.

Extracción de oro utilizando una resina AuRIX®100, pruebas en columna

Las pruebas en columna de intercambio iónico dividida en cuatro secciones conectadas en serie separación sólido-líquido, con una suspensión acuosa y resina en cada etapa, cada sección tiene



una malla de alambre en la parte inferior y superior para retener la resina colocada en cada una de las celdas. Las dimensiones de las celdas son de 2.5 x 7.7 cm de diámetro y altura (2.54 x 7.52 cm), para pasar la solución conteniendo oro, a través de la columna se usó una bomba peristáltica en tres corridas experimentales, con 5 gramos de resina como fase fija a 25°C, con una velocidad de 0.50, 0.60, 0.70 rpm, con 7.43 ml/min, 15 ml/min, 23 ml/min, equivalente a 89-276 BVH, fue necesario saber la relación flujo de solución/volumen de resina y también fueron analizadas las isotermas de adsorción.

Disolución de oro y disolución eluyente

La disolución de oro se llevo a cabo utilizando agua destilada y deionizada y tiosulfato de amonio (NH4)2S2O3 con pureza 99% marca Aldrich, la solución de oro (1000±2 µg/ml), se ajusto el pH con NH4OH y en la elución con hidróxido de sodio [NaOH], sulfato de cobre pentahidratado [CuSO4·5H2O], perclorato de sodio [NaClO4·2H2O], sulfito de sodio [Na2SO3] y cloruro de sodio [NaCl].

Pruebas de desorción

Los experimentos de desorción se realizaron en un reactor de 500 ml de capacidad con agitación, donde se mantuvieron constantes las siguientes condiciones experimentales: pH= 9.5-11.5, concentración de -

4ClO , concentración de -23SO , concentración de -Cl , concentración de NaOH.

Caracterización de la resina AuRIX®100

La caracterización se llevó a cabo por análisis difracción de rayos-X, caracterización por análisis infrarrojo y microscopio electrónico de barrido (MEB).

Tabla 2: Variables y rangos de experimentación en la etapa de adsorción

Variable Rango de experimentación pH 9-10.50 Concentración de Aui 1-8 mg/l Concentración de (NH4)2S2O3 0.00674-0.04 M

RESULTADOS Y DISCUSIÓN

En las Fig. 1 a 3 se estudió el efecto de tiempo de contacto en la extracción de oro, se ve el efecto del cambio de concentraciones de 1 mg/l a 8 mg/l, al aumentar el pH se logró un aumento en porcentaje extracción, donde el pH mejor fue de 10.5, en 3 h. En la Fig. 4 se muestra el efecto de la velocidad de agitación en la extracción de oro en resina AuRIX®100.

Fig. 1: Extracción de oro en resina AuRIX®100. Condiciones: [(NH4)2S2O3]= 0.04 M, [Au]i=1 mg/l,

AuRIX®100= 5 g. y T= 25ºC

Fig. 2: Extracción de oro en resina AuRIX®100. Condiciones: [(NH4)2S2O3]= 0.04 M, [Au]i= 2 mg/l, AuRIX®100= 5 g, velocidad= 500 rpm y T= 25ºC



Fig. 3: Efecto de tiempo de contacto en la extracción de oro en resina AuRIX®100.

Condiciones: [(NH4)2S2O3]= 0.04 M, [Au]i= 8 mg/l, AuRIX®100= 5 g, pH= 9-10.5 y T= 25ºC

Fig. 4: Efecto de velocidad de agitación en la extracción de oro en resina AuRIX®100. Condiciones: [(NH4)2S2O3]= 0.04 M, [Au]i= 1 mg/l, AuRIX®100= 5 g,

pH= 10.5, velocidad= 200, 500, 800 rpm

En la Fig. 5 se muestra el efecto de pH en la extracción de oro en la resina. Se encontró que la mayor extracción de oro se realizó a pH= 10.5 y 120 minutos y en la Fig. 6 se observó el efecto de la [(NH4)2S2O3] en la adsorción de oro en la resina al trabajar a pH 10.5. Es evidente el efecto positivo que produce el contenido de tiosulfato en la disolución acuosa sobre el porcentaje de oro adsorbido, en 3 h. En presencia de tiosulfato, la adsorción de oro alcanzó un 99% y con 0.00674 M de tiosulfato el comportamiento no fue estable, la adsorción disminuyó a 98%, 0.03 M de tiosulfato la adsorción disminuyó a 98% y 0.04 M de tiosulfato la adsorción aumento a 99%.

Fig. 5: Efecto de pH en extracción de oro en resina AuRIX®100. Condiciones: [(NH4)2S2O3]= 0.04 M,

[Au]i= 1, 2, 8 mg/l, AuRIX®100= 5 g, pH= 9-10.5 y T= 25ºC

Fig. 6: Efecto de la concentración de tiosulfato en extracción de oro en resina AuRIX®100.

Condiciones: [(NH4)2S2O3]= 0.00674, 0.03, 0.04 M y [Au]i= 1 mg/l, AuRIX®100= 5 g, velocidad= 500 rpm y

T= 25ºC

Se realizaron tres corridas experimentales con 5 gramos de resina, con una velocidad de 0.50, 0.60, 0.70 rpm, con 7.43 ml/min, 15 ml/min, 23 ml/min, equivalente a 89, 180 y 276 BVH. Los resultados se muestran en las Fig. 7 a 10.

En la Fig. 7 Muestra el perfil de la velocidad de adsorción, en cada sección de la columna, observándose que el trabajo en la última es inapreciable, pero en la primer sección es más rápida, por lo que esta será la primera que llegue a saturación, se observó que con el paso del tiempo la concentración que sale de cada sección va en aumento, en el momento que la concentración que entre sea igual a la que sale.



Fig. 7: Perfil en la extracción de oro al utilizar resina AuRIX®100, en las secciones de la columna con

velocidad de 0.50, con flujo de 7.43 ml/min, equivalente a 89 BVH

Fig. 8: Perfil de la concentración de oro, en las secciones de la columna con velocidad de 0.60, con

flujo de 15 ml/min, equivalente a 180 BVH

Fig. 9: Comportamiento en la extracción de oro al utilizar resina AuRIX®100 en las secciones de la

columna a 0.70, con flujo de 23 ml/min, equivalente a 276 BVH

Fig. 10: Perfil de la concentración en la extracción de oro al utilizar resina AuRIX®100 de la primera

sección de la columna manejando diferentes flujos

Para 180 BVH. se compararon las secciones se observó que se incrementa la velocidad de adsorción en cada sección con respecto a la velocidad mostrada en la Fig. 7 a 89 BVH. Se observó que la concentración de salida en cada sección se incrementa un poco con respecto a la de salida mostrada en la Fig. 7. Al aumentar el flujo aumenta la concentración de salida de la solución, esto indica que a mayor flujo el tiempo para alcanzar la concentración de saturación de la resina es menor y en caso de manejar flujos mayores sería necesario una quinta sección en la columna, para lograr la recuperación total del ión intercambiado.

CONCLUSIONES

En base a los resultados obtenidos del estudio de adsorción de oro, se puede concluir que: Al aumentar la concentración de tiosulfato de amonio de 0.00674, 0.03, 0.04 Molar se favoreció la extracción de oro, debido a una mayor presencia de tiosulfato libre. Cuando aumenta el valor de



pH por encima de 11, los iones hidroxilo promueven la degradación de iones tiosulfato, causando la precipitación de oro.

Se encontró que la presencia de tiosulfato favorece la adsorción de oro. Con tiosulfato, se obtuvo en 0.04 M, se obtuvo un 99% de adsorción mientras que con 0.03 M de tiosulfato, la adsorción disminuyó al 98%, el proceso de adsorción fue más inestable y con 0.00674 M la adsorción disminuyó al 98%, pero las condiciones de adsorción fueron demasiado inestables.

La velocidad de agitación del sistema es una variable muy importante porque permite la suspensión de las partículas en movimiento y eso facilita la transferencia de masa. Las concentraciones de oro como afectan de 1, 2, y 8 mg/l, los porcentajes de recuperación respectivamente fueron similares de 99, 93 y 95.75% de recuperación de oro.

La resina AuRIX®100 tiene las ventajas de trabajar a pH altos típicos de las soluciones industriales (9-11.5), dando como resultado el pH mejor 10.5 y en pH 9 disminuye la disolución de oro con tiosulfato de amonio, lo cual es atribuido a la inestabilidad en la disolución de oro.

AGRADECIMIENTOS

Los autores expresan sus agradecimientos a la Universidad de Sonora por brindar su apoyo durante la investigación, al Departamento de Ingeniería Química y Metalurgia, al Laboratorio de Servicios de Metalurgia, CONACyT, empresa COGNIS, a CESUES, a M.C. Juan Arévalo Amezcua, por su valiosa colaboración en la elaboración del proyecto.

REFERENCIAS

Aguayo, S., J.L. Valenzuela, J.R. Parga y M. Cruz, Continuous Laboratory Gold Solvent Extraction from Cyanide Solutions using LIX 79 Reagent, Chemical Engineering & Technology 30(11), 1457-1599 (2007).

Aylmore, M.G. y D.M Muir, Thiosulfate leaching of gold-a review, Minerals Engineering 14, 135-174 (2001).

Breuer, P.L. y M.I. Jeffrey, Electrochemical study of gold leaching in thiosulphate solutions containing ammonia and copper, Hydrometallurgy 65, 145-157 (2002).

Chaparro, M., “Extracción del complejo oro tiosulfato utilizando guanidina en una resina de intercambio aniónico”, Tesis de Maestría, Dpto. de Ing. Química y Metalurgia, Univ. de Sonora, Sonora, México (2008).

Ficeriová, J., P. Baláž., E. Boldižarová y J. Stanislav, Thiosulfate leaching of gold from a mechanically activated CuPbZn concentrate, Hydrometallurgy 67(5), 37-43 (2002).

Jeffrey, M.I., D.M. Hewitt, X. Dai y S.D. Brunt, Ion exchange adsorption and elution for recovering gold thiosulfate from leach solutions, Hydrometallurgy 100, 136-143 (2010).

Molleman, E. y D. Dreisinger, The treatment of copper-gold ores by ammonium thiosulfate leaching, Hydrometallurgy 66, 1-21 (2002).

Munive, T.G., M. Encinas, A. Valenzuela, Valenzuela J.L. y J.R. Parga, Estudio comparativo de la lixiviación de un mineral refractario con cianuro de sodio y tiosulfato de sodio para la recuperación de oro y plata, GEOMIMET 291, 12-24 (2011). Navarro, P., C. Vargas, A. Villarroel y F.J. Alguacil, On the use of ammoniacal/ammonium thiosulphate for gold extraction from a concentrate, Hydrometallurgy 65, 37-42 (2002).



Navarro, P., C. Vargas, V. Reveco y J. Orellana, Recuperación desde un medio amoniaco-tiosulfato con resina de intercambio iónico Amberlita IRA-410, Revista de Metalurgia 42(5), 354-366 (2006).

Seob, I. y otros siete autores, Sequential process of sorption and incineration for recovery of gold from cyanide solutions: Comparison of ion exchange resin, activated carbon and biosorbent,Chemical Engineering Journal: 165, 440-446 (2010).

Valenzuela A., J.L. Valenzuela, S. Aguayo y J.R. Parga, Estudio de la resina AuRIX®100 de intercambio iónico para recuperar oro en las soluciones cianuradas, GEOMIMET (268), 26-38 (2006).

Vargas, C., P. Navarro, E. Araya, F. Pavez y F.J. Alguacil, Recuperación de oro a partir de disoluciones de amoniaco y tiosulfato utilizando carbón activado, Revista de Metalurgia 42 (3), 222-233 (2006).

Yen, W.T., M. Aghamirian, G. Deschenes y S. Theben, “Gold extraction from mild refractory ore using ammonium thiosulfate”. Proceedings of the International Symposium on Gold Recovery, Montreal May 3-6 CIM (1998).

Zhang, H. y D.B. Dreisinger, The adsorption of gold and copper onto ion-exchange resins from ammoniacal thiosulfate solutions, Hydrometallurgy 66, 67-76 (2002).

REVISTA DE METALURGIA, 48 (3)MAYO-JUNIO, 165-174, 2012

ISSN: 0034-8570eISSN: 1988-4222

doi: 10.3989/revmetalm.1102

165

1. INTRODUCCIÓN

En México existen muchas menas de baja ley, asícomo otras en las que el oro y la plata se encuentranocluidos o asociados a minerales de hierro, arsénico,manganeso y silicio, los cuales se clasifican comodepósitos refractarios auríferos y argentíferos. Lamayoría de las veces, la cianuración de estos mineralesrequiere largos periodos de lixiviación y, desafortu-nadamente, bajas recuperaciones de oro[1-5]. La tos-tación de minerales y concentrados refractarios deoro, antes de la lixiviación con cianuro, es un métodoalternativo que se ha empleado durante décadas. Elprincipal objetivo de la tostación es liberar el oroíntimamente asociado a los minerales sulfurosos y

telurosos, para aumentar la extracción del oro durantela lixiviación con cianuro. Esto se realiza oxidandolos sulfuros a óxidos o sulfatos, o reduciendo los sili-catos y óxidos a un estado metálico crudo, con el pro-pósito de cambiar los compuestos metálicos insolublesen otros que sean solubles en el agente lixiviante; asícomo para volatilizar ciertas impurezas solubles quepudieran contaminar la solución cargada. Tambiénes útil para volver porosos los compuestos metálicos,y hacerlos de esta forma más fácilmente accesibles aldisolvente de la lixiviación[6].

Como la tostación es una operación relativamentecostosa, solo puede aplicarse a materiales de alta leyque puedan justificar los altos costos de este trata-miento. Sin embargo, puede obtenerse el efecto de

(•) Trabajo recibido el día 13 de enero de 2011 y aceptado en su forma final el día 20 de febrero de 2012.* Departamento de Ingeniería Química y Metalurgia, Universidad de Sonora. Rosales y Blvd. Luis Encinas S/N, Hermosillo, Sonora, 83000,

México. E-mail:([email protected].

Tostación de un concentrado refractario de oro y plata(•)

J. H. Coronado*, M. A. Encinas*, J. C. Leyva*, J. L. Valenzuela*, A. Valenzuela* y G.T. Munive*

Resumen En el procesamiento de minerales de metales preciosos con altos contenidos de pirita, se obtienen concentradosrefractarios difíciles de procesar. En este estudio se lixivió un concentrado refractario de oro y plata con cianuro desodio, obteniéndose extracciones de 34 % para oro y 40 % para plata. Se utilizó el método de tostación para oxidarel concentrado, haciéndolo más susceptible a la cianuración, y extraer más eficientemente el oro y la plata. Lasvariables analizadas fueron: temperatura de tostación y tiempos de tostación y cianuración. Además, la calcina calientese agregó a la solución lixiviante a temperatura ambiente para analizar el efecto en el tamaño de partícula y recuperación.Los mejores resultados, aunque no del todo satisfactorios (50 % oro y 61 % plata), se obtuvieron mediante tostacióndel concentrado durante 4 h a temperaturas de 600 °C, seguido de una cianuración de 20 h. El consumo de cal paraelevar el pH a 11,3, se incrementó notablemente hasta 25 kg/m³.

Palabras clave Tostación; Minerales refractarios; Cianuración.

Roasting of refractory gold and silver concentrate

Abstract In processing of precious metal ores with high pyrite content, refractory concentrates are obtained, which aredifficult to process. A refractory gold and silver concentrate was leached with sodium cyanide. Results show lowextraction percentages, being 34 % of gold and 40 % of silver. A roasting method to oxidize the concentrate wasused, making it more susceptible to cyanidation, hence a more efficient way to extract precious metals. Thevariables include roasting temperature and roasting and cyanidation time. In addition, the hot calcine was addedto the leaching solution at room temperature to analyze the effect on particle size and recovery. The best results,although not entirely satisfactory (50 % of gold and 61 % of silver) were obtained by roasting the concentrate for4 h at 600 °C, followed by cyanidation for 20 h. The lime consumption to raise the pH to about 11.3 was increasedmarkedly to 25 kg/m³.

Keywords Roasting; Refractory minerals; Cyanidation.

J. H. CORONADO, M. A. ENCINAS, J. C. LEYVA, J. L. VALENZUELA, A. VALENZUELA Y G.T. MUNIVE

166 Rev. metal. 48 (3), MAYO-JUNIO, 165-174, 2012, ISSN: 0034-8570, eISSN: 1988-4222, doi: 10.3989/revmetalm.1102

una tostación oxidante o sulfatizante de mineralessulfurosos de baja ley, mediante su intemperizaciónnatural experimentada durante un periodo prolon-gado de tiempo.

La mineralización de los materiales refractariosde oro no se ha establecido detalladamente paratodos los depósitos; no obstante, en muchos depó-sitos, la resistencia a la cianuración directa se atri-buye al oro extremadamente fino encapsulado enpirita y ciertos minerales de la ganga, como cuarzo.La presencia de materia carbonosa, arcillas, ciani-cidas, películas pasivantes, teluros y otros muchosfactores pueden también propiciar la refractariedadde una mena[7]. En algunos depósitos se ha observadoque el oro incrustado en pirita se presenta en unaforma no metálica asociada a zonas ricas de arseno-piritas. Una explicación es que los átomos de orohan substituido al arsénico o a los posibles átomosde hierro, dando lugar a una adecuada substituciónintersticial. Metalúrgicamente esto podría ser sig-nificativo, y puede explicar el porqué en una oxi-dación se proyecta solamente la oxidación parcialde los sulfuros[8].

La química básica de una oxidación por tostaciónpara minerales concentrados piritosos y arsenopiri-tosos de oro, es relativamente sencilla; los sulfurosmetálicos son convertidos a óxidos metálicos y dió-xido de azufre (SO2), este último liberándose comogas. Las reacciones se desarrollan generalmente entre600 y 700 °C.

La tostación de la pirita (FeS2) en una atmósferaaltamente oxidante, produce la hematita (Fe2O3)según lo mostrado en la reacción (1):

4 FeS2 + 11 O2 ! 2 Fe2O3 + 8 SO2 (1)

Cuando la atmósfera de la tostación es menos oxi-dante, la magnetita (Fe3O4), será producida según lomostrado en la reacción (2):

3 FeS2 + 8 O2 ! Fe3O4 + 6 SO2 (2)

Para la extracción más favorable del oro, un pro-ducto que contiene aproximadamente 75 a 85 % dehematita, y el resto magnetita, es generalmente elmás deseable. Se puede formar durante la tostaciónun poco de sulfato de hierro (FeSO4), el cual es gene-ralmente indeseable, los ferrocianuros se forman fácil-mente con FeSO4

[9].La arsenopirita (FeAsS) es oxidada a dióxido de

azufre (SO2), hematita (Fe3O4) y trióxido de arsénico(As2O3).

La oxidación de la arsenopirita es algo más complejaque la de la pirita. La reacción (3) ocurrirá si la arse-nopirita se tuesta en una atmósfera altamente oxidante.

FeAsS + 3 O2 ! FeAsO4 + SO2 (3)

Esta es una reacción indeseable puesto que seforma el arseniato férrico (FeAsO4), que inhibirá laextracción del oro durante la lixiviación con cianuro.Las reacciones deseables son mostradas en las reac-ciones (4 y 5); la reacción (4) convierte el arseniatoa trióxido de arsénico volátil (As2O3), en una atmós-fera menos oxidante a una temperatura relativamentemás baja (500 °C y 80 % de O2).

La reacción (5) se desarrolla a una temperatura ya una atmósfera altamente oxidante.

12 Fe As S + 29O2 !6As2 O3 + 4Fe3O4 + 12SO2 (4)

4Fe3O4 + O2 ! 6Fe2 O3 (5)

Para lograr estas reacciones se requiere una tos-tación en 2 etapas. La etapa 1 debe ser a bajas tem-peraturas y poco aire para producir As2O3 y evitar laformación de FeAsO4. La etapa 2 debe ser a una tem-peratura alta y con exceso de aire para producir lahematita porosa (As2O3 es durante la tostación).

La reacción (6) muestra que el carbón orgánicoelemental (C) se oxida a dióxido de carbono (CO2).

C + O2 ! CO2 (6)

Esta reacción depende del tipo de carbón presentey de la temperatura de tostación. El carbón orgánicocon una temperatura de ignición bastante baja se oxi-dará fácilmente a CO2, mientras que los carbonos detipo grafíticos con altas temperaturas de ignición pue-den no reaccionar durante la tostación. Los carbo-natos de magnesio y calcio (MgCO3, CaCO3), se cal-cinan parcialmente a óxidos; sin embargo, las tem-peraturas de tostación son demasiado bajas para lacalcinación completa[6].

La reacción (7) muestra la descomposición de loscarbonatos de calcio y magnesio a óxidos de calcioy magnesio (CaO/MgO).

CaCO3 /MgCO3 ! CaO/MgO + CO2 (7)

TOSTACIÓN DE UN CONCENTRADO REFRACTARIO DE ORO Y PLATAROASTING OF REFRACTORY GOLD AND SILVER CONCENTRATE

Rev. metal. 48 (3), MAYO-JUNIO, 165-174, 2012, ISSN: 0034-8570, eISSN: 1988-4222, doi: 10.3989/revmetalm.1102 167

Esta reacción depende del tipo de carbonatos pre-sentes y de la temperatura. El carbonato de magnesiose descompone a una temperatura mucho más bajaque el carbonato de calcio (350 °C frente a 825 °C,dependiendo de los minerales). La descomposicióndel dióxido de azufre descomponiéndose a sulfato osulfito (SO3

=) es a menudo una reacción deseable,porque reduce la cantidad de SO2 que se debe lavaren los gases de salida. La desventaja de la formaciónde sulfato de calcio (CaSO4) es que pueden alterarselas características de la calcina e inhibir la disolucióndel oro.

La reacción (8) muestra la reacción entre CaO oMgO, generado en la reacción (7), y el SO2 durantela tostación.

CaO/MgO + SO2 + ½ O2 ! CaSO4/MgSO4 (8)

Esta reacción es generalmente deseable, sinembargo, con algunos minerales concentrados, la for-mación de sulfato de calcio inhibirá la extracción deloro durante la lixiviación con cianuro (el oro se encap-sula en el sulfato de calcio). Esta reacción se refierea la fijación del SO2. Una tostación ideal formará lahematita porosa; la temperatura excesiva de tostación(sobre tostación o sinterización del mineral), inhibirála formación de la hematita y reducirá seriamente laextracción del oro. La tostación generalmente es per-judicial para la plata en las extracciones por lixivia-ción; cuanto más alta es la temperatura de tostación,más baja es la extracción de plata. Las variables deproceso más importantes que deben ser consideradasen la tostación son: cantidad de alimentación, tem-peratura, tiempo de tostación, adición de oxígeno,oxidación del azufre y del carbón orgánico elemental,fijación del dióxido de azufre, entre otras[5, 10 y 11].

2. MÉTODO EXPERIMENTAL

El análisis químico del concentrado mineral utilizadoen este estudio se muestra en la tabla I. Las pruebasde tostación se llevaron a cabo utilizando un hornoeléctrico marca Lindberg/Blue M., modelo 55345.Para hacer los análisis del tamaño de partícula, se uti-lizó un analizador de partículas marca Coulter 100 Q.Para evaluar los resultados obtenidos en la tostacióndel concentrado se utilizó una lixiviación agitada concianuro de sodio (NaCN), en un reactor de vidrioPyrex de 100 ml con acceso para la toma de muestra,y un sistema de agitación marca STIR-PAK de 25-2.300 rpm, equipado con flecha y propela. Losreactivos utilizados en la lixiviación fueron cianuro

de sodio, hidróxido de calcio y nitrato de plata, todosellos de grado reactivo y disueltos en agua destilada.Para el análisis de la solución lixiviada, se utilizó unespectrofotómetro de absorción atómica, marca PerkinElmer modelo 3110 con lámpara de cátodo hueco auna longitud de onda de 242,8 nm para el oro; parala plata se aplicó una llama de oxígeno-acetileno yuna longitud de 328,1 nm.

En cada prueba experimental, primero se ajustó elhorno tostador a temperatura de trabajo, posterior-mente se pesaron 12 g de concentrado, los cuales secolocaron en un recipiente de porcelana; la muestrade concentrado se introdujo en el interior del hornotostador por un tiempo programado. Al finalizar laetapa de tostación, la muestra se retiró del horno,transfiriéndose a un reactor con el agente lixiviante,midiendo inmediatamente el valor de pH.Seguidamente se adicionó cal hasta alcanzar un valorde pH entre 10,5 y 11,0, para posteriormente agregarel cianuro de sodio, tomando este paso como el iniciode la etapa de lixiviación y manteniéndola duranteel intervalo de tiempo determinado. Se tomaron mues-tras de solución a intervalos de tiempo predetermi-nados para analizar las concentraciones de oro y plataen la solución. La cantidad de NaCN en la solución,se obtuvo por titulación con nitrato de plata.

3. RESULTADOS Y DISCUSIÓN

En la figura 1 a) se presenta el análisis de tamaño departícula del concentrado sin tostar. Podemos obser-var que la media del tamaño de partícula se encuentraalrededor de 23,56 micras, y la moda alcanza un valorde 23,81 micras.

Tabla I. Análisis químico del concentrado

Table I. Chemical analysis of concentrate

Elemento Composición (%)

Au 60(*)

Ag 10,200(*)

Fe 36,38Cu 0,50Zn 4,72As 0,15S 33,34

pH 6,05

(*) g/t.



Analizando de forma general, podemos concluirque el concentrado obtenido del proceso de flotacióncontiene partículas finas, con tamaños menores de33 micras. En la figura 1 b) se presenta el análisis detamaño de partícula obtenido del concentrado calci-nado por un tiempo de 1 h a una temperatura de500 oC, con el objetivo de observar el cambio detamaño de partículas al someterse al proceso de tos-tación. En esta figura se observa que existe un aumentoconsiderable en el tamaño de la partícula después delproceso de tostación, ya que el valor medio en estecaso resultó ser de 95,18 micras, con una moda de127,6 micras. Por lo anterior se concluye que existeun aumento en el tamaño de las partículas del con-centrado al estar expuesto a este tratamiento térmico,el cual posiblemente se deba a una expansión de laspartículas, generada por la salida de elementos en formade gases, los cuales tienen su punto de ebullición menora 500 oC. En la figura 1 c), se presenta el análisis detamaño de partícula del concentrado que fue sometidoa un proceso de tostación a una temperatura de 500 oCdurante una hora, y posteriormente la calcina calientese agregó inmediatamente a la solución lixiviante, quese encontraba a temperatura ambiente, para lograr conesto, un cambio rápido de temperatura en el sistema;con el objetivo de determinar si la partícula caliente,por efecto del cambio térmico sufre algún tipo de frag-mentación en su tamaño, lo cual podría ayudar a laliberación de los valores. Sin embargo, comparandoestos resultados con los de la figura 1 b), observamosdiferencias poco significativas en el tamaño de partí-cula; éstas se consideran no significativas para deter-

minar que exista algún cambio apreciable en el tamañode la partícula. La tabla II, presenta el análisis de losparámetros estadísticos antes mencionados.

Figura 1. Resultados del analizador de partículas del concentrado: 1 a) sin tostar; 1 b) con-centrado tostado, t = 1 h, T = 500 ºC; 1 c) concentrado tostado, posteriormente la calcina calientese agregó inmediatamente a la solución lixiviante, t = 1 h, T = 500 ºC.

Figure 1. Particle size analyses of the concentrate: 1 a) unroasted concentrate; b) roastedconcentrate, t = 1 h, T = 500 °C; 1 c) roasted concentrate, after roasting the hot calcine wasadded immediately to the leaching solution, t = 1 h, T = 500 °C.

Tabla II. Análisis estadístico del analizador de partículas

Table II. Statistical analysis of particle size analyzer

Volumen estadístico (aritmético) 100%

a b c

Media 23,56 95,14 106,3Mediana 17,38 78,14 94,47Moda 23,81 127,6 127,6D.S. 21,10 77,26 75,35C.V. 89,5(*) 81,2(*) 70,9(*)

% < a b c

10 2,498 10,29 17,4025 7,405 31,86 46,1250 17,38 78,14 94,4975 33,03 140,5 151,890 56,98 207,6 212,7

Cálculo de partículas de 0,375 a 948 µm.Tamaño en micras, (*) %



La figura 2 presenta los resultados obtenidos enla recuperación de oro y plata durante la lixiviacióndel concentrado, por un periodo de 25 h sin ningúntratamiento de tostación.

En esta figura se muestra que a las 2,5 h, el orotiene un porcentaje de recuperación del 24,16 %,posteriormente a las 5 h se incrementa hasta un28,89 %, alcanzando un máximo del 34,16 % a las25 h. La plata tiene un comportamiento ligeramentesuperior, con porcentajes de recuperación de 31,36 %en las primeras 2,5 h, hasta llegar a 40,19 % despuésde 25 h de lixiviación. El consumo de cianuro paraeste tipo de concentrado sin tostar alcanza los3,6 kg/m3. Se puede apreciar tanto para el oro comopara la plata, que las recuperaciones por el métodode cianuración de estos concentrados son demasiadobajas, por lo que se considera al concentrado comorefractario, siendo necesario utilizar un método com-plementario y/o alternativo para su tratamiento, conel fin de aumentar la recuperación de los valores deoro y plata.

En la figura 3 se observan los resultados obtenidosen las pérdidas de peso, azufre y arsénico del concen-trado calcinado a diferentes temperaturas, de los 350a 800 °C con un tiempo de tostación de 1 h.

A temperaturas menores de 350 °C existe unapérdida de peso no mayor del 0,94 %; al aumentar la

temperatura a 400 °C el incremento en la pérdidade peso aumenta notablemente alcanzando un13,42 %, siguiendo este aumento hasta un 18,68 %a una temperatura de 450 °C. A partir de esa tem-peratura, la velocidad de pérdida de peso disminuye,alcanzando un máximo de 19,6 % a una temperaturade 800 °C. Podemos pensar que a temperaturas alre-dedor de 450 °C la pérdida de peso se relaciona conla pérdida de azufre presente en el concentrado, entreotros elementos volátiles.

En el caso del azufre, a temperaturas menores de350 °C, la eliminación por volatilización del conte-nido total es bajo, con un 1,14 %. Entre 350 y 450 °Cla pérdida de azufre se incrementa rápidamente hastaun 76,3 % y a 500 °C alcanza un 84,55 %. A partirde esta última temperatura, la velocidad de pérdidase hace más lenta hasta alcanzar un 93,16 % a 700 °C.Observamos en la gráfica, que los resultados obtenidoscoinciden con los esperados, ya que el punto de ebu-llición del azufre (444,6 °C), está dentro de los inter-valos de temperaturas manejadas.

El porcentaje de pérdida de arsénico a 350 °C esaproximadamente del 12 %, al ir incrementando latemperatura, se aprecia que el porcentaje aumentade tal manera que a una temperatura de 700 °C lapérdida es del 40 %. Sin embargo, la ley del arsénicoen el concentrado es baja, 0,15 %; analizando estos

Figura 2. Recuperación de oro y plata y consumo de cianuro del con-centrado sin tratamiento de tostación. t = 25 h, peso de muestra = 12 g.

Figure 2. Recovery of gold and silver, and cyanide consumption from theconcentrate without a roasting threatment, t = 25 h, sample weight = 12 g.



resultados se puede pensar que el intervalo de tem-peratura de tostación no afecta de forma general enel contenido de arsénico, por lo que posiblementeéste se encuentra asociado con alguna especie queno permite su completa eliminación.

En la figura 4 se muestran los resultados del com-portamiento en la recuperación de oro y plata de con-centrado tostado y lixiviado durante 1 h. La recupe-ración del oro a temperaturas de tostación de 450 °Ces del 32,16 %; al aumentar la temperatura a 500 °Cexiste una disminución drástica en los porcentajesde recuperación del oro de 10,83 %, hasta llegar a6,58 % a 700 °C. Los porcentajes de lixiviación dela plata se encuentran alrededor del 10,68 % a 450°C, manteniendo valores casi similares hasta la tem-peratura de 700 °C. El consumo de cianuro de sodioes de alrededor de los 0,5 kg/m3 a 400 °C, generán-dose un aumento paulatino hasta llegar alrededor delos 0,8 kg/m3 a 700 °C.

Puede observarse, que tanto las recuperacionesde plata como las de oro son muy bajas; posiblementelos tiempos de tostación y lixiviación no son los ade-cuados. La calcina que se obtiene en la tostación esde color gris-negruzco, indicándonos que hay presen-cia de una mayor cantidad de magnetita, lo cual nofavorece a la lixiviación, coincidiendo con lo obser-vado en otros estudios, donde un color de calcina

café-chocolate indica una mayor presencia de hema-tita (≈ 75-80 %), siendo el resto de magnetita, dondese obtienen las mejores recuperaciones[9].

La figura 5 presenta los resultados obtenidos enla recuperación de oro y plata del concentrado tostadoy lixiviado con un tiempo de 4 h de tostación y 20 hde lixiviación respectivamente.

En esta figura puede observase que el oro se recu-pera aproximadamente en un 27 % a una temperaturade tostación entre 400 y 500 °C. Al aumentar la tem-peratura a 600 °C hay un aumento en la recuperaciónde oro casi del 50 %, y al tostar a 700 °C la recupe-ración disminuye drásticamente hasta un 20,66 %.Respecto a la plata, ésta sigue una tendencia parecidaal oro, pero con una mayor recuperación, así podemosobservar, que al tostar el concentrado a 400 °C, larecuperación es del 31,32 %, alcanzando un máximode 61,27 % a 600 °C. Al incrementar la temperaturaa 700 °C existe una caída en la recuperación de plataal 51 %. La pérdida de la recuperación del oro y plataa altas temperaturas, posiblemente se deba a la sin-terización de la partícula de concentrado.

El consumo de cianuro a 400 °C, es de 0,8 kg/tony disminuye hasta un 0,2 kg/m3 a 700 °C.

En la figura 6 se presentan los resultados obtenidosen la recuperación de oro y plata, cuando se agregóinmediatamente el concentrado tostado (calcina),

Figura 3. Efecto de la temperatura de tostación, en la pérdida de peso,azufre y arsénico del concentrado, t = 1 h, peso de muestra = 12 g,T = 500 ºC.

Figure 3. Effect of roasting temperature, in weight loss of sulfur and arsernicfrom the concentrate, t = 1 h, sample weight t = 12 g, T = 500 ºC.



Figura 4. Resultados en la recuperación de oro, plata y consumo de cianurodel concentrado tostado y lixiviado, t = 1 h, peso de muestra = 12 g.

Figure 4. Recovery of gold and silver, and cyanide consumption fromthe concentrate roasted and leached, t = 1 h, sample weight = 12 g.

Figura 5. Resultados en la recuperación de oro, plata y consumo decianuro del concentrado tostado y lixiviado, tostación t = 4 h, lixiviaciónt = 20 h, peso de muestra = 12 g.

Figure 5. Recovery of gold and silver, and cyanide consumption fromthe concentrate roasted and leached roasting treatment, t = 4 h, leachingt = 20 h, sample weight = 12 g.



desde el horno a la solución lixiviante, que se encon-traba a 40 °C, con el objetivo de observar el com-portamiento en los porcentajes de recuperación deloro y plata al efectuar un cambio brusco en la tem-peratura de la calcina caliente. La recuperación deloro a la temperatura de 400 °C se encuentra alrededordel 13 %; al aumentar la temperatura disminuye laextracción de tal manera que a 500 °C es aproxima-damente 7 %, hasta llegar a un 3 % en 700 °C. Porotro lado, la recuperación de plata se encuentra en11 % a 400 °C, empezando a subir a 15 % en 500 °C,para alcanzar un 26 % a 700 °C.

Estas recuperaciones de oro y plata son demasiadobajas, ya que la recuperación del oro cae hasta un3 %, y aunque la plata sube, en general es muy poco.Se observa que a diferencia de las pruebas anterioresel comportamiento es muy diferente, posiblementeel cambio de temperatura de forma rápida afecta a lasuperficie de la partícula, lo cual impide la penetra-ción de la solución lixiviante para la extracción delos valores. Existe un aumento de la temperatura enla solución de 2 a 3 °C, lo cual no provoca beneficioalguno para la etapa de lixiviación en este caso. Elconsumo de cianuro oscila en un promedio de0,9 kg/m3.

En la figura 7 se presentan los resultados de larecuperación de oro y plata cuando la calcina calientese pone en contacto con la solución lixiviante quese encuentra a una temperatura de 50 °C. Los valoresobtenidos son casi similares a los obtenidos en la grá-fica anterior (Fig. 6), ya que el porcentaje de recu-peración de oro se encuentra alrededor del 24 % auna temperatura de tostación de 400 °C, y disminuyehasta un 1,5 % a 700 °C. La recuperación de plata a400 °C es del 10 % y sufre un incremento hasta un32 % a 700 °C. Con lo anterior se constata que alagregar la calcina caliente directamente del horno ala solución lixiviante, no existe una mejora en larecuperación del oro, sino que la perjudica notable-mente, ya que las recuperaciones son demasiado bajas.Por otro lado, aunque la recuperación de plata seincrementa, estos resultados no son del todo satis-factorios. El consumo de cianuro se encuentra alre-dedor de 0,8 kg/m3.

En la figura 8 se presenta el consumo de cal paraelevar el pH del concentrado lixiviado en 24 h, alre-dedor de 1,3 kg/m3 para subir el pH a 11,4. Tambiénse presenta el consumo de cal de la lixiviación delconcentrado tostado a 500 °C. Al inicio el pH de lasolución oscila alrededor de 3,0; al agregar el equi-

Figura 6. Resultados en la recuperación de oro, plata y consumo decianuro del concentrado tostado y lixiviado, tostación t = 4 h, lixiviacióna una temperatura inicial de 40 °C, t = 20 h, peso de muestra = 12 g.

Figure 6. Recovery of gold and silver, and cyanide consumption of theconcentrate roasted and leached roasting treatment, t = 4 h, leachingat initial temperature of 40 °C, t = 20 h, sample weight = 12 g.



Figura 7. Resultados en la recuperación de oro, plata y consumo decianuro del concentrado tostado y lixiviado, tostación t = 4 h, lixiviacióna una temperatura inicial de 50 °C, t = 20 h, peso de muestra = 12 g.

Figure 7. Recovery of gold and silver, and cyanide consumption of theconcentrate roasting and leached roasting t = 4 h, leaching at initial temperature of 50 °C, t = 20 h, sample weight = 12 g.

Figura 8. Consumo de cal del concentrado lixiviado y tostado. Lixiviaciónt = 20 h, tostación t = 4 h, peso de muestra = 12 g.

Figure 8. Consumption of limestone from the concentrate leached androasted. Leaching t = 20 h, roasting t = 4 h, , sample weight = 12 g.



valente a 7 kg/m3 de cal el pH es de 3,2; al aumentarla cal a 17 kg/m3 el pH es de 6,1, siendo necesarioagregar otros 7 kg de cal (suman ya 23 kg/m³) parallegar a un pH mayor de 11. Si comparamos el con-sumo de cal para elevar el pH alrededor de 10,5 conrespecto a un mineral oxidado, éste requiere aproxi-madamente 1 kg/m3, por lo que concluimos que esdemasiada la cal necesaria para obtener el pH ade-cuado para la lixiviación.

CONCLUSIONES

— El concentrado presenta una ley de oro de 60 g/ty de plata de 10,200 kg/t, un 36,38 % de hierroy 33,34 % de azufre. El azufre se elimina hasta un60 % a una temperatura de tostación inicial de400 °C, hasta alcanzar un 93 % a una temperaturade 700 °C.

— Al tostar el concentrado se detectó un incre-mento en el tamaño de partícula, debido posi-blemente a la eliminación del azufre en forma degas. La lixiviación del concentrado por 24 h,alcanza un porcentaje de recuperación del orodel 34 % y del 40 % para la plata, confirmándosela clasificación de mineral como refractario.

— A una temperatura de tostación de 600 °C y des-pués de lixiviar con cianuro de sodio se obtieneun porcentaje de recuperación máxima del orodel 50 % y 61 % para la plata. Al aumentar latemperatura a 700 °C, las recuperaciones dismi-nuyen, posiblemente debido a una sinterizaciónde la partícula. Al someter la calcina caliente, ala solución lixiviante, existe un aumento de lasolución de 2 a 3 °C, resultando en porcentajesde recuperación de oro y plata muy bajos.

— El consumo de cal para elevar el pH a las condi-ciones de lixiviación con cianuro (pH > 10),oscila alrededor de 25 kg/m³. El color de la calcina

café-chocolate, nos indica que se efectúa una oxi-dación de los sulfuros a hematita (75-80 %) ymagnetita (20-25 %) lo cual favorece a la lixi-viación, coincidiendo con lo observado en otrosestudios[9], donde reportan buenas recuperacionesdel mineral de oro y plata. Se recomienda hacerotras pruebas con otros lixiviantes, para compararrecuperaciones, y/o utilizar otro método alterna-tivo en lugar de tostación.

REFERENCIAS

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[2] J. R. Parga y F. R. Carrillo. Rev. Metal. Madrid,32 (1996) 254-261.

[3] V. Arias-Arce, R. Coronado-Falcón, L. PuenteSantibáñez y D. Lovera-Dávila. Revista delInstituto de Investigación FIGMMG, 8 (2005)5-16.

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[5] J. H. Coronado, Tesis de Maestría, Departamentode Ingeniería Química y Metalurgia, Universidadde Sonora, 1997.

[6] J. R. Goode. TMS, EPD Congress, 1993, pp.291-317.

[7] J. C. Yannopoulus. The extractive metallurgyof gold, Van Nostrand Reinhold, 1991 pp. 79-98.

[8] D. C. Harris. Mineral. Deposita 25 (1990) 3-7. [9] R. B. Coleman. Proceedings of Gold. Symposum,

1990, pp. 381-388.[10] A. Vladimir, R. C. Falcon, L. Puente y D. Lovera.

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Advances in Chemical Engineering and Science, 2012, 2, 342-349 doi:10.4236/aces.2012.23040 Published Online July 2012 (http://www.SciRP.org/journal/aces)

Kinetic Aspects of Gold and Silver Recovery in Cementation with Zinc Power and Electrocoagulation

Iron Process

Gabriela V. Figueroa Martinez1, José R. Parga Torres1, Jesús L. Valenzuela García2, Guillermo C. Tiburcio Munive2, Gregorio González Zamarripa1

1Department of Metallurgy and Materials Science, Institute of Technology of Saltillo, Saltillo, México 2Department of Chemical Engineering and Metallurgy, University of Sonora, Hermosillo, México

Email: [email protected]

Received February 11, 2012; revised March 16, 2012; accepted March 29, 2012

ABSTRACT The Merrill-Crowe or Cementation process is used for concentration and purification of gold and silver from cyanide solutions. Among other available options for recovery of precious metals from cyanide solutions, Electrocoagulation (EC) is a very promising electrochemical technique for the recovery of this precious metals. In this work first, an intro-duction to the fundamentals of the Merrill Crowe and EC process are given, then Kinetic aspects conditions and results of the both process, for the removal of gold and silver from cyanide solutions are presented, and finally the characteri-zation of the solid products formed during the EC process with X-ray Diffraction and SEM are shown. Results suggests that The cementation of both gold and silver by suspended zinc particles conforms to well-behaved fist order kinetics and for the EC process the results show that is an excellent option to remove Au and Ag from cyanide solution by using iron electrodes. Finally, 99.5% of gold and silver were removed in the experimental EC reactor, and it was achieved in 5 minutes or less. Keywords: Cyanidation; Merrill Crowe; Electrocoagulation; Kinetics

1. Introduction Cyanide leaching processes have been used by the min-ing industry for over 150 years in the extraction of noble metals, the popularity of cyanidation is based mostly on the simplicity of the process. Elsner in Germany in 1846 studied the dissolution of gold in cyanide aqueous solu-tion and noted that atmospheric oxygen played an im-portant role during dissolution of gold [1]. Also, sodium cyanide in an alkaline solution is a strong solvent for gold and silver, most mill operators use it to dissolve fine gold particles with a practical maximum size is no greater than 50 microns. In most cyanidation operations, the gold particles require 24 to 72 hrs for complete dissolution in slurry or pulp of about 50 percent solids. Extremely large leaching reactors known as Pachuca tank in with the finely ground ore was agitated with the alkaline cyanide leaching agent and equipped with compressed air injec-tion in the pulp had been designed to dissolve the gold and silver. It is widely accepted that the gold cyanidation process can be represented by the classic Elsner’s Equa-tion (1). The mechanism and kinetics have been dis-cussed in several papers and reviews [1,2].

� �2 2 24Au 8CN O 2H O 4Au CN 4OH� �� (1)

Silver, similarly, dissolves readily in dilute cyanide solutions in the presence of oxygen. However, since silver in Mexicans ores occurs as argentite the cyanidation and sulphide oxidation reaction are as follows:

� � 22 2Ag S 4CN 2Ag CN S�� (2)

22 2

1S CN O H O CNS 2OH2

� � �� (3)

The cyanide concentration determines the rate of an-odic gold dissolution while the oxygen reduction rate is dependent on the concentration of dissolved oxygen. In the cyanidation process, free cyanide ions in solution can be provided only at pH > 9.3. The pH of the pulp can be increased with the additions of alkali hydroxides (NaOH, KOH, Ca(OH)2 etc.), known as proactive alkalis. Also, details of this electrochemical reaction have received considerable attention and under certain circumstances the reaction is limited by the coupled diffusion of CN– and O2 to the gold surface.

Lixiviation of undesirable base metals, such as copper,

Copyright © 2012 SciRes. ACES

G. V. F. MTZ ET AL. 343

iron and arsenic, reduces the efficiency of the process by consuming additional reagents, and necessitates further processing to remove these contaminants [3].

Actually, the two conventional processes for gold and silver recovery from cyanide leach solution are: the Mer- rill Crowe zinc dust cementation and the carbon adsorp-tion process. In the first process, the efficiency of the reaction is significantly improved by removing dissolved oxygen from the system prior to zinc addition, using a vacuum deaeration technique. This reaction of cementa-tion is sensitive to suspend solids in the leach solution and thus requires solution clarification before cementa-tion. In the activated carbon process, the precious metals are absorbed onto granules of carbon and after loading, they are then stripped of the loaded gold by a hot caustic- cyanide solution. Then, the cathodes from the carbon adsorption process or the precipitates from the Merrill Crowe process (metal displacements) are then melted in crucible furnaces along with fluxing materials such as borax, niter and silica. The resultant product from smelt-ing is Dore bullion of precious metals typically analyzing more than 95 percent of precious metals.

Each recovery method has advantages and disadvan-tages. Process selection depends on the specific condi-tions for particular operation and the facilities already available. Another new process is the Electrocoagulation (EC) technology and has recently been reviewed as an alternative for gold and silver recovery from alkaline cyanide solutions [3]. The EC process is a very promis-ing technique for the recovery of precious metals such as silver and gold: EC needs no chemical reagents, does not generate toxic materials requiring special disposal and this also make it an ecologically viable technique. Lit-erature reviewing showed that the potential of EC as an alternative to traditional treatment recovery of precious metals (silver and gold cyanide) has not yet been ex-ploited.

1.1. The Merrill Crowe Process or Cementation Technology

The term cementation comes from a Spanish word mean- ing “precipitation”. This term was first used in 1500 to describe a process for the recovery of copper from aque-ous solutions at Rio Tinto in Spain. However, cementa-tion of a noble metal from solution by means of more base, or electronegative, metal has been practiced since ancient times [4]. Near the beginning of the fourth cen-tury, Zosimos stated that, when iron is immersed in a solution of a copper salt, it acquires a coating of copper. In the sixteenth century, Paracelsus precipitated silver from silver nitrate by inserting a plate of copper in the solution, and he noted that the copper dissolved. Berg-man, in this “De Precipitates Metallics”, in the mid-

eighteenth century, observed that the metals precipitate one on another after a certain order of nobility: zinc, iron, lead, tin, copper, silver and gold.

Then, cementation reactions, variously know as metal- displacement reactions or contact-reduction reactions, are processes in which a metal ion in a solution is reduced to the elemental metallic state with the concurrent oxidation of a more electronegative metal placed into the same solution. This process is one of the most ancient, yet economical and efficient, hydrometallurgical processes known and has been used for the recovery of dissolved metal values from leach solutions, as well as for the puri-fication of electrolytes. Cementation has been exten-sively applied for the recovery of gold and silver from cyanide solutions, typically followed by the production of Dore metal, which is known as the Merrill-Crowe process. Although cementation of precious metals is widely used in the metallurgy industry, most of the arti-cles published on cementation of gold and silver until about 1980 dealt only with plant practice. Scientific de-tailed laboratory investigations of the cementation reac-tion have only been reported during the past three dec-ades. Studies of gold and silver have mostly been con-fined to the kinetics of gold and silver cementation on rotating discs or cylinders [5].

In spite of this research effort, the mechanistic details of the reactions taking place during cementation and the role of various species in solution in promoting or inhib-iting the recovery of precious metals have received little attention and are still somewhat unclear. For example, it is generally accepted that cementation is sensitive both to the alkalinity and to the free cyanide concentration in solution. However, suitable experimental data are not available to describe plant practice in any detail. Only a few fundamental studies have been carried out in the recent past and also, until now, no detailed research has been reported regarding the morphological influence of the deposit structure on the cementation rate.

1.2. Review of Cementation Phenomena

In cementation reactions the metal to be reduced from aqueous solutions is more noble, having a greater elec-tron affinity, than the precipitating metal. For example, the recovery or metallic silver by cementation with zinc dust is an electrochemical process involving the oxida-tion of zinc and the reduction of the silver cyanidation. The overall stoichiometry for the reaction in the case of silver is as follows:

� � � �2 42Ag CN Zn 2Ag Zn CN

� �� (4)

A similar reaction can be written for gold precipitation. Silver is deposited at cathodic surface sites while zinc dissolves from anodic sites, and electrons are conducted



between the two metallic phases as shown in Figure 1.

1.3. Cementations Kinetics Considerable research has been reported on the charac-teristic features and mechanistic details of cementation reactions and this literature has been reviewed in terms of electrochemical theory and transport phenomena on sev-eral occasions [6].

In most cases, the cementation reactions follow first- order reaction kinetics and generally are limited by diffu-sion of the noble metal ion through the mass-transfer boundary layer. In some cases the structure and mor-phology of the reaction product can have a significant effect on passivating the surface in other cases.

Most of the cementation reactions are found to be first- order diffusion processes [7] with respect to the noble metal ion, and the reaction velocity constant, k, for such a reaction may be computed from the general first-order rate equation:

dc kAC

dt V

� (5)

If k is not concentration dependent and the area is un-changing, Equation (5) may be integrated resulting in the integrated first-order expression,

o

C kAlog

C 2.3 �

t

V (6)

where C and Co are the noble metal concentrations at tine t and the initial concentration at time t = 0, respectively;

Figure 1. Electrochemical reaction of cementation of silver on zinc particle.

k is the reaction velocity constant (cm/sec); A is reaction surface area (cm2); and V is the solution volume rate (cm3). In Equation (5), it can be seen that the cementation rate is a function of the reaction area A. it has been shown by many investigators that in almost all cementa-tion systems the initial exposed geometric area of the precipitant metal can be used in the analysis of initial rate data. However, in a more general sense, the area term is not always that simple to evaluate due to the changing nature of the noble metal deposit which grows on the active metal during the course of the reaction.

Experimental investigations [7] indicate that most ce-mentation reactions are controlled by a mass transfer process, film diffusion, as indicated by the results pre-sented in Table 1.

Notice in Table 1, that for almost all the cementation systems the apparent activation energy is in the range 2 - 6 kcal/mole which would suggest that the cementation reactions are limited by mass transfer in the aqueous phase, with some exceptions (Pb2+/Fe and Pb2+/Cu). Also the reaction velocity constants are of the order 10–2 cm/sec which also supports the position that these ce-mentation reactions are mass transfer controlled.

1.4. The Electrocoagulation Technology Electrocoagulation (EC) has been known as an electro-chemical phenomenon since the last century. It has been employed previously for treating many types of waste-

Table 1. Data for selected cementation aqueous systems at 25˚C [7].

System ¨Eo, VActivation Energy,

Kcal/mole Reaction VelocityConstant, cm/sec

Ag+/Cu 0.46 2.0 - 5.0 2.5 - 6.0 × 10–2

Ag+/Cu (CN–) 1.83 3.7 - 5.8 1.5 × 10–2

Ag+/Fe (Cl–) 1.29 3.0 2.2 × 10–2

Ag+/Zn (CN–) 0.95 5.5 5.5 × 10–2

Ag+/Zn 1.56 2.0 - 6.0 2.6 - 5.2 × 10–2

As3+/Cd 0.65 3.1 ± 1.3 7 × 10–3

Au+/Zn(CN–) 0.61 3.1 1.7 × 10–2

Bi3+/Fe 0.76 4.5 - 7.6 2.9 × 10–2

Cd2+/Zn 0.36 4.0 - 4.7 0.54 - 1.1 × 10–2

Cu2+/Fe 0.75 3.1 - 5.1 0.6 - 0.9 × 10–2

Cu2+/In 0.83 2.3 5.9 × 10–2

Cu2+/Ni 0.57 2.7 - 3.7 (14.2 - 19.0) 0.25 - 1.0 × 10–2

Cu2+/Zn 1.10 3.1 1.6 - 2.1 × 10–2

Ni2+/Fe 0.21 7.0 1.4 × 10–4

Pb2+/Fe 0.31 12.0 -

Pb2+/Zn 0.64 - 0.64 × 10–2

Pb2+/Cu 0.49 9.5 - 7.4 0.36 - 2.3 × 10–2



water with varying degrees of success [8-10]. The EC process can be considered as an acelerated corrosion process was green rust (GR) is an intermediate product that is responsible for the removal of contaminants (sus-pended and dissolved solids, metals, organic compounds, etc). EC mechanisms may involve oxidation, reduction, decomposition, deposition, coagulation, absorption, ad-sorption, precipitation and flotation. EC operates on the principle that hidrolized cations produced electrolytically from iron and/or aluminum anodes enhance the coagula-tion of contaminants from an aqueous medium.

The sacrificial metal anodes are used to continuously produce polyvalent metal cations in the vicinity of the anode. These cations facilitate coagulation by neutraliz-ing the negatively charged particles that are carried to-ward the anodes by electrophoretic motion. Generally in the EC process bipolar electrodes are used [9-11]. It has been reported that cells with bipolar electrodes connected in series operating at relatively low current densities produced iron or aluminum coagulant more effectively. In the EC technique, the production of polyvalent cations from the oxidation of the sacrificial anodes (Al or Fe) and the production of electrolysis gases (O2 and/or H2) are directly proportional to the amount of current applied (Faraday’s law). The electrolysis gases enhance the flota-tion of the coagulant material. A schematic representa-tion of the EC process, using iron electrodes, is shown in Figure 2. As mentioned above, the gas bubbles produced by electrolysis carry the gold and silver along with the sludge to the top of the solution where it is collected and removed.

However, it is the reactions of the metal ions that en-hance the formation of the coagulant. The metal cations are hydrolyzed, releasing hydrogen ions that result in hydrogen evolution at the cathode, to yield both soluble and insoluble hydroxides that will react with or adsorb gold and silver from the cyanide solution and also con-tribute to coagulation by neutralizing the negatively charged colloidal particles that may be present at neutral

Figure 2. An illustration of the EC mechanism (arrow indi-cate the migration of ions, the H2 evolution and the forma-tion of green rust).

or alkaline pH. This enables the particles to approach closely and ag-

glomerate under the influence of Van der Waals attrac-tive forces. The chemical reactions that have been pro-posed to describe EC mechanism [12] when using iron electrodes are:

Fe l Fe+3 + 3e– (7)

� � � �2 122H OFe OH Fe OH H� � �� o � (8)

2H+ + 2e– o H2(g)Ĺ (9)

� � � � � �1

2Fe OH e Fe OH� �� o 2 aq (10)

� � � � � � 1q 322 a H OFe OH Fe OH H� �� o � (11)

� � � � � �1

3 3 aqFe OH Fe OH e� �o � (12)

Overall reaction

� � � � � �

� � � � � � � �

2 g

3 2

2

s

16Fe 12 x 12 x H2

H O

xFe OH 6 x Fe OH

� � o � n

� u � (13)

The pH of the medium usually rises as a result of this electrochemical process and the Green Rust formed [xFe(OH)3 × (6 – x)Fe(OH)2(s)] remains in the aqueous stream as a gelatinous suspension, which can remove the gold and silver from pregnant cyanide rich solutions, either by complexation or by electrostatic attraction of magnetic nanoparticles followed by coagulation and flo-tation. Generally, in the EC process, bipolar electrodes are used [8]. It has been reported that cells with bipolar electrodes, connected in series operating at relatively low current densities, produce iron or aluminum coagulant more effectively, more rapidly and more economically when compared to chemical coagulation.

2. Experimental Procedures 2.1. Suspended Zinc Particle Experiments The bath reactor experiments were done in four-necked, one liter glass reaction cells supported in a constant tem-perature bath. A condenser, stirrer, nitrogen dispersion tube, and sampling derive were placed into a reactor through the openings in the lid. The Teflon stirrer was attached through the center port by means of a Chesa-peak stirrer connection. In all studies purified nitrogen was passed through the solution, via a dispersion tube, before and during the experiment to maintain an oxygen- free environment. Tyler sieves were used to prepare monosize fractions of zinc in the range from 70 to 400 mesh.

The specific surface area of the zinc dust was meas-ured using two techniques, the BET technique (Quan-


G. V. F. MTZ ET AL.


346

tasorb Model 0S10, made by Quantachrome Corporation), surface area 255 cm2/gr, and the air permametry tech-nique (Permaran made by Outokumpu), surface area 216 cm2/gr. The stirring speed was adjusted manually (1000 RPM) and was checked with a stroboscope. The particu-late sample of zinc was washed with acetone and cleaned with acid to remove the surface oxide film before being introduced into the reaction flash. For most of the ex-periments, one gram of 270 × 400 mesh (d50 = 45 µm) particles of zinc were introduced through one of the cover ports, and the reaction was initiated. In order to follow the course of cementation reaction, solution ali-quots of 2 ml were taken periodically and analyzed for silver and gold with a DCP (Direct-Current Plasma) Spectrophotometer manufactured by Beckman Instru-ments.

2.2. Experimental Electrocoagulation Details The EC experiments were performed in a 400 ml beaker size reactor equipped with two carbon steel electrodes (6 cm × 3 cm) that were 5 mm apart. As a source of current and voltage a universal AC/DC adaptor was used. pH was measured with a VWR scientific 8005 pH meter. Gold and silver adsorption onto iron species was investi-gated with pregnant cyanide solutions provided by Bacis S. A. de C. V. mining group (13.25 ppm Au and 1357 ppm Ag and pH of 8). Analyses were performed by ICP/ Atomic Emission Spectrometry. The conductivity of pregnant solutions was adjusted by adding one gram of NaCl per liter (Fisher, 99.8% A. C. S. Certified, lot #995007).

To identify and characterize the iron species in the solid products, formed during the EC process for the re-moval of gold and silver using iron electrodes, X-ray diffraction (XRD) and Scanning Electron Microscope (SEM/EDX) were used. Analysis of Au and Ag were conducted to the Bacis solution, by AES. EC was run at 15 Volts (DC) and the corresponding current was of 0.1 Å. EC was run for five minutes, and a sample was taken every minute in order to determinate the removal effi-ciency for Au and Ag. Solutions and solids from the EC process were separated by filtration through cellulose filter paper. The sludge from the EC was dried either in an oven or under vacuum at room temperature and char-acterized.

3. Results and Discussion 3.1. Zinc Dust Particle Size Results As was shown by Equation (5), the rate of cementation reaction is expected to be a function of the surface area and the concentration of the reacting noble metal ion (gold or silver in the present case). To examine the tradi-tional first order rate expression, the cementation reac-tion kinetics were studied using monosize zinc particles. Two experiments were conducted at 10 ppm (5 × 10–5 M) gold and 100 ppm (9 × 10–4) silver in order to study the reaction kinetics at these concentrations. The equilibrium concentration profiles (on logarithmic scale), the results were found to conform to the first-order rate process for over 99% removal as shown in Figure 3 (Au/Zn(CN)– system) and (Ag/Zn(CN)– system). From these data it is evident that the cementation of both gold and silver by zinc conforms to well-behaved fist order kinetics. Also, from these figure, the micrograph of reacted zinc parti-cles can be seen. This micrograph reveals the nature of the silver and gold deposits. The effect of the gold and silver deposits on the rate of cementation reaction prod-uct showed that the gold and silver formed uniforms, apparently porous layer around the zinc particle. Table 2 lists the values of the reaction velocity constant for the cementation of gold (10 ppm) and silver (100 ppm) on zinc dust.

3.2. Electrocoagulation Results Running the EC process for gold and silver removal on iron electrodes, gave the results shown in Table 3 and in Figure 4, shows the equilibrium concentration profile of gold and silver removal efficiency and final pHs values vs. Time. The obtained results show that EC is an excel-lent option to remove Au and Ag from cyanide pregnant solution by using iron electrodes. Also, under these con-ditions, the results show that, when residence time in-creases from 1 to 3 minutes the silver and gold removal increase to 92% and 99% respectively, this variation oc-curs at pH values from 9 to 10.7, approximately; these values coincide with the production of magnetic iron, Fe3O4 (green rust). Also, it is apparent that the uptake percents of gold/silver increased by increasing the pH value.

The mechanism of the EC process for gold and silver Table 2. Comparison of experimental reaction velocity constants with calculated mass-transfer coefficients for suspended particles. NaCN = 10–2 M, pH = 10.5, Stirring speed = 1000 rpm, Au = 10 ppm (5 × 10–5 M), Ag = 100 ppm (9 × 10–4 M).

System Particle size, mesh Experimental reaction velocity constant, cm/sec Calculated mass-transfer coefficient

Au/Zn(CN)– 400 × 270 1.9 × 10–2 2.35 × 10–2

Ag/Zn(CN)– 400 × 270 1.62 × 10–2 3.2 × 10–2


Figure 3. First-order plots showing removal of gold and silver from cyanide solution.

Table 3. Recovery of gold and silver by EC.

EC residence time (min)

Au (mg·L–1)

Recovery (%)

Ag (mg·L–1)

Recovery (%) pH

0 13.25 0 1357.0 0 8.0

1 12.50 5.66 940.0 30.72 9.2

2 10.50 20.37 219.5 83.82 9.5

3 1.00 92.45 9.0 99.33 10.7

4 0.50 96.22 7.0 99.48 11.2

5 0.10 99.24 0.9 99.93 11.5

recovery is then physical and chemical adsorption, caused by the highly reactive species and adsorption properties of green rust. These processes do not alter their chemical composition. Figure 4, also shows graphically the pH during the EC the process for gold and silver removal, the pH increment in the solution from 8 to a pH final of 11 is attributed to the hydrogen evolution at the cathode which is then accompanied by alkalinization of the aqueous solution.

3.3. Product Characterization X-ray Diffraction Analysis. Diffraction patterns of flocs collected from the experiment with gold and silver, (the sample were ground to a fine powder and loaded into a

sample holder) were obtained with a diffracted X-PERT Phillips meters equipped with a vertical goniometer, with a range of analysis 2ș 10˚ to 70˚. The source of X-rays has a copper anode, whose radiation is filtered with a graphite monochromator (Ȝ = 1.541838 Å) with scan rate of 0.02˚ and a duration of 10 seconds per count. The X-Ray Diffractometer is controlled by a Gatawey 2000 computer, by PC-APD 2.0 with software for Windows.

Figure 5 shows the ray diffraction pattern of the flocs recovered from the sample of gold and silver, respec-tively 13.25 mg/L and 1357 mg/L, initial pH 8, 5 minutes of treatment, 0.1 amperes and 15 volts. The species iden-tified were magnetite, lepidocrocite, goehtite, silver and copper hexacyanoferrate. Scanning Electron Micros-copy (SEM/EDAX). Figure 6 shows SEM images and EDAX of silver adsorbed on iron species. These SEM and EDAX results show that the surfaces of these iron oxide/oxyhydroxide particles were coated with a layer of silver. It is worth clarifying that, given the low concen-tration of gold it was impossible to locate any nanoparti-cle of it.

4. Conclusion The cementation of both gold and silver by suspended inc particles conforms to well-behaved fist order kinet- z



Figure 4. Gold and silver recoveries from Bacis cyanide solutions and pH vs. EC residence time.

Figure 5. X-ray diffractogram of solids obtained in the re-covery of gold and silver C: Cu2Fe(CN)62H2O, A: Silver, M: Magnetite and L: Lepidocrocite.

ics. The effect of the gold and silver deposits on the rate of cementation reaction product showed that the gold and silver formed uniforms, apparently porous layer around the zinc particle. The experimental reaction velocity con-stants correspond to the expected magnitude of the limit-ing mass-transfer coefficients and support the hypothesis that the cementation reactions under these conditions are mass-transfer limited reactions with recoveries of gold and silver of 99.8%. The obtained results show that EC process is an excellent option to remove Au and Ag from cyanide pregnant solution by using iron electrodes. The X-Ray Diffraction, Scanning Electronic Microscopy, techniques demonstrate that the formed species are of mag-netic type, like lepidocrocite and magnetite, and amor-phous iron oxyhydroxide which adsorbed the silver and gold particles on his surface due to the electrostatic at-traction between both metals. The 99.5% of gold and silver were removed in the experimental EC reactor, and it was achieved in 5 minutes or less with a current effi-ciency of 99.7%.

(a)

(b)

Figure 6. Chemical composition of solid product as deter-mined by EDX, which shows the presence of silver in the particle of iron.



5. Acknowledgements The authors wish to acknowledge support of this project to the National Council of Science and Technology (CONACYT) and to Dirección General de Educación Superior Tecnológica (DGEST) from Mexico.

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GMPR #601483, VOL 33, ISS 5

RECOVERY OF SILVER AND GOLD FROM CYANIDESOLUTION BY MAGNETIC SPECIES FORMED IN THEELECTROCOAGULATION PROCESS

J. R. Parga, M. Rodrıguez, V. Vazquez, J. L. Valenzuela, andH. Moreno, H. Moreno

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Recovery of Silver and Gold From Cyanide Solution by Magnetic SpeciesFormed in the Electrocoagulation ProcessJ. R. Parga, M. Rodrıguez, V. Vazquez, J. L. Valenzuela, and H. Moreno

RECOVERY OF SILVER AND GOLD FROM CYANIDESOLUTION BY MAGNETIC SPECIES FORMED IN THEELECTROCOAGULATION PROCESS

J. R. Parga1, M. Rodrıguez1, V. Vazquez1, J. L. Valenzuela2,5and H. Moreno3

1Department of Metallurgy and Materials Science, Saltillo Institute ofTechnology, Saltillo, Coahuila, Mexico2Department of Chemical Engineering and Metallurgy, University of Sonora,Hermosillo, Sonora, Mexico

103Department of Chemical Engineering, La Laguna Institute of Technology,Torreon, Coahuila, Mexico

Cyanidation is the predominant process by which gold and silver are recovered from

their ores in metallurgical operations, and it is recognized that the Carbon in Pulp, the

Merrill–Crowe, the Ion Exchange, and Solvent Extraction processes are used for concen-15tration and purification of gold and silver from cyanide solutions. Among other available

options for recovery of precious metals from cyanide solutions, Electrocoagulation (EC)

is a very promising water and wastewater electrochemical technique that does not require

high concentrations of silver and gold in cyanide solutions to yield excellent results. In this

work, an introduction to the fundamentals of the EC process is given, followed by the con-20ditions and results of the EC test run for removal of precious metals from cyanide solutions,

and finally the characterization of the solid products formed during the EC process with

X-ray Diffraction, SEM, and Transmission Mossbauer Spectroscopy. Results suggest that

magnetite particles and amorphous iron oxyhydroxides are present (Lepidocrocite and

Gohetite). With the EC process, the achieved removal efficiency of silver and gold from25cyanide solutions, within 5min, exceeded 99%.

Keywords: electrocoagulation, gohetite, green rust, lepidocrocite

INTRODUCTION

Cyanidation processes are especially suitable for treating gold=silver-bearingsulfidic materials. Modern hydrometallurgy technology of precious metals is based

30on the application of cyanide leaching for the dissolution of gold and silver. Goldcyanidation are reported to involve the chemical reactions shown in Eqs. (1) and(2). Silver follows a similar process.

2Auþ 4NaCNþO2 þ 2H2O ! 2Na Au CNð Þ2! "

þ 2NaOHþH2O2 ð1Þ

Address correspondence to J. R. Parga, Department of Metallurgy and Materials Science, SaltilloInstitute of Technology, V. Carranza 2400, Saltillo, Coahuila 25000, Mexico. E-mail: [email protected]

Mineral Processing & Extractive Metall. Rev., 33: 1–11, 2012Copyright # Taylor & Francis Group, LLCISSN: 0882-7508 print=1547-7401 onlineDOI: 10.1080/08827508.2011.601483

3b2 Version Number : 7.51c/W (Jun 11 2001)File path : p:/Santype/Journals/TandF_Production/Gmpr/v33n5/GMPR601483/GMPR601483.3dDate and Time : 14/04/12 and 12:22

1

2Auþ 4NaCNþH2O2 ! 2Na Au CNð Þ2! "

þ 2NaOH ð2Þ35

The equation proposed by Elsner is stoichiometrically correct but does notdescribe the cathodic reactions associated with the dissolution. The stoichiometryof the process shows that 4mol of cyanide are needed for each mole of oxygen presentin solution. At room temperature and standard atmospheric pressure, approximately

408.2mg of oxygen are present in 1L of water. This corresponds to 0.27$ 10%3mol=L.Accordingly, the sodium cyanide concentration (molecular weight of NaCN¼ 49)should be equal to 4$ 0.27$ 10%3$ 49¼ 0.05 g=L or approximately 0.01%. Thiswas confirmed at room temperature by a very dilute solution of NaCN of0.01–0.5% for ores, and for concentrates rich in gold and silver of 0.5–5% (Parga,

45Valenzuela, and Cepeda 2007). Details of this electrochemical reaction have receivedconsiderable attention and under certain circumstances the reaction is limited by thecoupled diffusion of CN– and O2 to the gold surface. Lime or sodium hydroxide(caustic) is added to keep the system at an alkaline pH beteween 10 and 11. This pro-tective alkalinity is required to counteract the generation of acid during cyanidation,

50to prevent cyanide degradation, and the formation of the deadly HCN gas.The two conventional processes for gold and silver recovery from cyanide leach

solution are: The Carbon Adsorption Process and the Merrill–Crowe zinc dustcementation process. In the Carbon Adsorption Process, the precious metals areabsorbed onto granules of activated carbon. After loading, they are then stripped

55of the loaded gold by a hot caustic-cyanide solution. This solution is then fed to elec-trowinning cells where gold and silver are electrolytically deposited onto steel woolcathodes. In the Merrill–Crowe process, the product is filtered zinc dust precipitate.The cathodes from the carbon adsorption process or the precipitates from the Mer-rill–Crowe process (metal displacements) are then melted in crucible furnaces along

60with fluxing materials such as borax, niter, and silica. The resultant product fromsmelting is Dore bullion of precious metals typically analyzed with more than 97%of precious metals.

Each recovery method has advantages and disadvantages. Process selectiondepends on the specific conditions for a particular operation and the facilities already

65available. Traditionally, The Merrill–Crowe method was the preferred process formany years. In just the past 40 years, the carbon adsorption process has becomepopular for recovering gold from large volumes of low grade pregnant leach solu-tions that contain mainly gold.

The Ion Exchange Resins and Solvent Extraction process has recently been70reviewed as an alternative for gold and silver recovery from alkaline cyanide solu-

tions (Valenzuela et al. 2003; Aguayo et al. 2007). Another method that has beentried for the recovery of gold and silver uses adsorption on a chemically modifiedchitosan with magnetic properties (Donia, Atia, and Elwakeel 2007).

The alternative processes are compared in Table 1 along with those of the EC75(Parga et al. 2005; Moreno Casillas 2009).Q1 The EC process is seen to be a very prom-

ising technique for the recovery of precious metals such as silver and gold: EC needsno chemical reagents, and does not generate toxic materials requiring special dis-posal making it an ecologically viable technique. Literature review shows the poten-tial of EC as an alternative to traditional treatment recovery of precious metals

80(silver and gold cyanide) yet to be exploited.

2 J. R. PARGA ET AL.

ELECTROCOAGULATION CHARACTERISTIC

Electrocoagulation (EC) has been known as an electrochemical phenomenonfor more than 70 years. EC has been employed previously to treat many types ofwastewater with varying degrees of success (Mollah et al. 2004; Parga et al. 2005;

85Emamjomeh and Sivakumar 2009).The EC process can be considered as an accelerated corrosion process where

green rust (GR) is an intermediate product responsible for the removal of contami-nants (suspended and dissolved solids, metals, organic compounds, etc). EC mechan-isms may involve oxidation, reduction, decomposition, deposition, coagulation,

90absorption, adsorption, precipitation, and flotation. EC operates on the principlethat hidrolized cations produced electrolytically from iron and=or aluminum anodesenhance the coagulation of contaminants from an aqueous medium. The sacrificialmetal anodes are used to continuously produce polyvalent metal cations in the vicin-ity of the anode. These cations facilitate coagulation by neutralizing the negatively

95charged particles that are carried toward the anodes by electrophoretic motion.

Table 1 Advantages and disadvantages of methods for recovery of gold and silver (Mollah et al. 2004;Parga, Valenzuela, and Cepeda 2007; Emamjomeh and Sivakumar 2009)

Method Advantages Disadvantages

Merrill–Crowe . Lower capital and operatingcosts.

. Handles solutions containinghigh silver and gold content.

. It is highly efficient (99.5%).

. Also can treat high-gradesolutions produced by carbonelution.

. Alternative to electrowinning.

. The pregnant solution needsclarification and deoxygenating.

. Low concentrations of metalsincreases amount of zinc.

. Depends on the pH and concen-tration of the free cyanide.

. The precipitate contains cyanideslike copper and arsenic.

Adsorption withActivated Carbon

. Does not require pretreatment ofpregnant solution.

. Not dependent on the concen-tration of metals.

. Large specific surface.

. The pulp needs no clarification.

. Fouled carbon needs to beregenerated by heating.

. Large carbon inventory.

. The pregnant solution has to gothrough 5 or 6 columns.

. High operating costs.Ion Exchange Resins . Does not need: washing,

revitalization, or heat treatment.. High abrasion resistance in tanksof adsorption.

. High selectivity.

. High cost of the process.

. Lower loading capacity.

. Royalty payments.

. The resin must be regenerated inacid medium.

Solvent Extraction . High selectivity.. Does not need: washing, revita-lization, or heat treatment.

. High cost of the process.

. New technology.

. Stripping Difficult.Electrocoagulation . Low residence time (minutes).

. Does not use chemicals.

. Handles solutions containinglower or high silver and goldcontents.

. Energy costs per m3 of pregnantsolution are lower than conven-tional treatment systems.

. Sacrificial anode must bereplaced periodically.

. Precise initial pH control.

. New technology.

. The product is high in iron.

RECOVERY OF SILVER AND GOLD FROM CYANIDE SOLUTION 3

Bipolar electrodes are generally used in the EC process (Mollah et al. 2004;Emamjomeh and Sivakumar 2009; Zhao et al. 2010). It is reported that cells withbipolar electrodes connected in series operating at relatively low current densitiesproduced iron or aluminum coagulant more effectively. In the EC technique, the

100production of polyvalent cations from the oxidation of the sacrificial anodes (Alor Fe) and the production of electrolysis gases (O2 and=or H2) are directly pro-portional to the amount of current applied (Faraday’s law). The electrolysis gasesenhance the flotation of the coagulant material. A schematic representation of theEC process, using iron electrodes, is shown in Figure 1. Aforementioned, the gas

105bubbles produced by electrolysis carry the gold and silver along with the sludge tothe top of the solution where it is collected and removed.

It is the reactions of the metal ions that enhance the formation of the coagu-lant. The metal cations are hydrolyzed, releasing hydrogen ions that result in hydro-gen evolution at the cathode. This process yields both soluble and insoluble

110hydroxides which react with or adsorb gold and silver from the cyanide solutionand contribute to coagulation by neutralizing the negatively charged colloidal parti-cles if present in a neutral or alkaline pH. This enables the particles to closelyapproach and agglomerate under the influence of Van der Waals attractive forces.The EC mechanism and chemical reactions (Moreno et al. 2009) when using iron

115electrodes are:

Fe $ Feþ3 þ 3e% ð3Þ

Fe OHð Þþ2 þH2O ! Fe OHð Þþ12 þHþ ð4Þ

2Hþ þ 2e% ! H2ðgÞ " ð5Þ

Figure 1 An illustration of the EC mechanism to remove gold and silver (arrows indicate the migration ofelectrolysis gases O2 and H2).


Fe OHð Þþ12 þ e% ! Fe OHð Þ2 aqð Þ ð6Þ

Fe OHð Þ2ðaqÞ þH2O ! Fe OHð Þ%13 þHþ ð7Þ

Fe OHð Þ%13 ! Fe OHð Þ3 aqð Þ þ e% ð8Þ

Overall reaction

6Feþ 12þ xð ÞH2O ! 1=2 12% xð ÞH2ðgÞ " þ xFe OHð Þ3 ' 6%Xð ÞFe OHð Þ2ðsÞ ð9Þ125

The pH of the medium usually rises as a result of this electrochemical processand the GR formed [xFe(OH)3

' (6 – X) Fe(OH)2 (s)] remains in the aqueous streamas a gelatinous suspension, which can remove the gold and silver from pregnantcyanide rich solutions, either by complexation or by electrostatic attraction followed

130by coagulation and flotation.Formation of rust (dehydrated hydroxides) occurs after the following

processes:

2Fe OHð Þ3 $ Fe2O3 þ 3H2O hematite;maghemiteð Þ ð10Þ

Fe OHð Þ2 $ FeO þH2O ð11Þ

2Fe OHð Þ3 þ Fe OHð Þ2 $ Fe3O4 þ 4H2O magnetiteð Þ ð12Þ

Fe OHð Þ3 $ FeO OHð Þ þH2O goethite; lepidocrociteð Þ ð13Þ

Due to the high voltage in the EC cell, a very strongly oxidizing environment isproduced around the anode suitable for destroying strong cyanide solutions (greater

140than 1000 ppm). This represents a direct oxidation of the cyanide ion at the anode tocyanate ion which is further decomposed to carbon dioxide and nitrogen,ammonium, and carbonate or oxalate ions according to the pH (Hwang, Wang,and Wan 1987). The reactions are as follows:

In strong alkaline solution (pH¼ 12)

CN% þ 2OH% ! CNO% þH2Oþ 2e ð14Þ

CNO% þ 2OH% ! CO2 þ 1=2N2 þH2Oþ 3e ð15Þ

In neutral and weak alkaline solution (pH¼ 7.0–11.7)

2CN ! C2N2 ð16Þ

C2N2 þ 2OH% ! CNO% þ CN% þH2O ð17Þ

CNO% þ 2H2O ! NHþ4 þ CO2%

3 ð18Þ


In weak acidic solution (pH¼ 5.2–6.8)

C2N2 þ 4H2O ! C2O2%4 þ 2NHþ

4 ð19Þ155

Tamura et al. (1974) determined that the anodic oxidation of cyanide is pro-portional to the alkalinity of the electrolyte consistent with the following mechanism:

CN% þ 2OH ! CNO% þH2Oþ 2e% ð20Þ

2CNO% þ 4OH% ! 2CO2 þN2 þ 2H2Oþ 6e% ð21Þ

CNO% þ 2H2O ! NH4 þ CO%3 ð22Þ

EXPERIMENTAL DETAILS

EC experiments were performed with a Fisher magnetic stirrer and a 400mLbeaker size reactor equipped with two carbon steel electrodes (6$ 3 cm) 5mm apart.

165The electrodes were properly scrubbed and rinsed prior to each experiment trial runto ensure a clean surface free of passive oxide layers. A universal AC=DC adaptorwas used as a source of current and voltage. The pH was measured with a VWRscientific 8005 pH meter. Gold and silver adsorption onto iron hydroxide specieswas investigated with pregnant cyanide solutions provided by Bacis S.A. de C.V.

170mining group (13.25mg=L Au, 1357mg=L Ag, 200mg=L free CN–and 1400mg=Ltotal CN–and pH of 8). Analyses were performed with ICP=Atomic Emission Spec-trometry (Perkin Elmer 3100). The conductivity of pregnant solutions was adjustedby adding 1 g of NaCl per liter (Fisher, 99.8% A.C.S. Certified, lot #995007). Toidentify and characterize the iron species in the solid products, formed during the

175EC process for the removal of gold and silver using iron electrodes, X-ray diffraction(XRD Phillips model X-PERT) and Scanning Electron Microscope (SEM=EDX,FEI Quanta 2000, Oxford Instruments) were used.

Analyses of Au and Ag were conducted with the Bacis solution by AES. ECwas run at 15V (DC) and the corresponding current was of 0.1 A.

180EC was performed on five replicates for 5min each with a sample taken everyminute in order to determinate the removal efficiency for Au and Ag. Solutions andsolids (shown in Figure 2) from the EC process were separated by filtration throughcellulose filter paper. The sludge from the EC was dried either in an oven or undervacuum at room temperature and characterized.

185The resulting sludge of iron hydroxide gel precipitate with Au=Ag is filtered.This rich sludge is treated in an acid leach step with sulfuric acid under oxidizingconditions caused by the addition of air. The conditions of this acid leach are suchthat the major portion of iron is leached into solution with gold and silver remainingin the residue. The resultant residue from the filtration step is gold, silver, and some

190iron (10% Au, 80% Ag, and 5% Fe) suitable for further refining by conventionalcommercial method.



The EC process, gave the results shown in Table 2, values of concentration comefrom the arithmetic average of five replications. As shown in Figure 3, the maximum

195values for gold and silver recovery achieved within 5min of treatment are of 99.24%for Au and of 99.93% for Ag with a standard deviation of 0.26 for gold and 0.06 forsilver. This study showed very good recovery values within 2–3min for gold and1–2min for silver. These results occur with pH ranging from 9 to 11, which coincideswith the production of magnetite (Fe3O4). Magnetite’s magnetic properties accelerate

200the process of metal adsorption. The adsorption rate is physically determined by themagnetic forces of the magnetite into gold and silver without altering their chemicalcomposition. The high voltage in the EC cell around the anode destroys some of thecyanide. It is likely that the EC cell oxidize Au and Ag cyanide complexes, thus convert-ing them to a less soluble form that might be captured by the iron hydroxide gel.

205Figure 4 graphically displays the evolution of pH during the operation time.There is an increase in pH of the bulk solution, which is attributed to the hydrogenevolution at the cathode that accompanied the generation of hydroxyl ions during EC.

Figure 4 also shows cyanide concentration vs. time. In the EC cell, with ironelectrodes, cyanides decompose in the gold and silver pregnant solution, and the con-

210centration changes from the original cyanide content of 1400 to 210mg=L.

Figure 2 Picture of the EC process for the removal of Au and Ag, showing GR.

Table 2 Recovery of gold and silver by EC

EC residence time(min) Au (mg=L)

Recovery(%) Ag (mg=L)

Recovery(%) pH

Cyanide(mg=L)

Cyanideremoval

0 13.25 0 1357.0 0 8.0 1400 01 12.50 5.66 1240.0 8.62 9.2 1050 252 10.50 20.37 219.5 83.82 9.5 870 383 1.00 92.45 9.0 99.33 10.7 750 464 0.50 96.22 7.0 99.48 11.2 400 715 0.10 99.24 0.9 99.93 11.5 210 85


PRODUCT CHARACTERIZATION

X-Ray Diffraction Analysis

Diffraction patterns of experimentally collected flocs with gold and silver (sam-ples were ground to a fine powder and loaded into a sample holder) were obtained

215with a diffracted X-PERT Phillips meters equipped with a vertical goniometer with a2h analysis range of 10(–70(. The X-ray source is a copper anode, whose radiation isfiltered with a graphite monochromator (k¼ 1.541838 A) with scan rate of 0.02( anda duration of 10 s per count. The X-Ray Diffractometer is controlled using aGateway 2000 computer by PC-APD 2.0 with software for Windows.

220Figure 5 shows the ray diffraction pattern of flocs recovered from the sample ofgold and silver, respectively, 13.25mg=L and 1357mg=L, initial pH 8, 5min of treat-ment, 0.1 A and 15V. The species identified were magnetite, lepidocrocite, goehtite,silver, and copper hexacyanoferrate.

Figure 4 pH vs. EC residence time and Cyanide removal vs. time.

Figure 3 Gold and silver recoveries from Bacis cyanide solutions.


Scanning Electron Microscopy (SEM/EDAX)

225Figure 6 shows SEM images and EDAX of silver adsorbed on iron species.These SEM and EDAX results show that the surfaces of these iron oxide=oxyhydr-oxyhydroxide particles were coated with a layer of silver. Given the low concen-tration of gold, it was impossible to locate any nanoparticle of it.

Transmission Mossbauer Spectroscopy

230Figure 7 shows the spectrum obtained from the EC silver, gold, and iron solidproduct from 1350mg=L of silver and 13mg=L of gold cyanide solutions at

Figure 5 X-ray diffractogram of solids obtained in the recovery of gold and silver C: Cu2Fe (CN)62H2OA: Silver, M: Magnetite, and L: Lepidocrocite.

Figure 6 Chemical composition of solid product as determined by EDX, which shows the presence ofsilver in the particle of iron.


pH¼ 11.0. Mossbauer Spectra for each sample was obtained on a )15mm=s velocityscale, which allows observation of wide magnetic hyperfine spectra expected fromiron oxide compounds. The spectrum consists of a doublet magnetic spectrum, which

235is probably due to fine particles of iron oxides (nonstoichiometric magnetite) or ironhydroxides (Lepidocrocite, Goethite, etc.).

From the analysis of these techniques, the in situ generated small fine particlesof iron- oxide=oxyhydroxides in the EC process are: nonstoichiometric magnetite,goethite, and iron hydroxide oxide.

240CONCLUSIONS

The results of this study indicate that silver and gold can be successfullyadsorbed on iron species produced by the EC process. EC may be a viable methodto recover gold and silver from cyanide solutions.

The X-Ray Diffraction, Scanning Electronic Microscopy, and Transmission245Mossbauer Spectroscopy techniques demonstrate that the formed species are

magnetic (e.g., lepidocrocite and magnetite and amorphous iron oxyhydroxide),which adsorb the silver and gold particles on the surface by electrostaticattraction.

Yields of 99.93% gold and 99.24% silver were efficiently removed in the experi-250mental EC reactor within 5min and a current efficiency of 99.7%.

Aditionally, the high voltage in the EC cell around the anode destroys some ofthe cyanide.

Figure 7 Mossbauer spectrum with silver and gold at pH¼ 10.5, indicating the presence of magnetitespecies.


ACKNOWLEDGMENTS

The authors wish to acknowledge support of this project to the National Coun-255cil of Science and Technology (CONACYT) and to Direccion General de Educacion

Superior Tecnologica (DGEST) from Mexico.

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chemically modified chitosan with magnetic properties.’’Hydrometallurgy, 87, pp. 197–206.Emamjomeh, M. M. and Sivakumar, M., 2009, ‘‘Review of pollutants removed by electrocoa-

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Hwang, J. Y., Wang, Y. Y., and Wan, C. C., 1987, ‘‘Electrolytic oxidation of cuprocyanideelectroplating wastewaters under different pH conditions.’’ Journal of Applied Electro-chemistry, 17, pp. 684–694.

Mollah, M., Morkovsky, P., Gomez, J., Parga, J. R., and Cocke, D., 2004, ‘‘Fundamentals,270present and future perspectives of electrocoagulation.’’ Journal of Hazardous Materials,

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Moreno, H., Cocke, D. L., Gomes, J. A. G., Morkovsky, P., Parga, J. R., Peterson, E., andGarcia, C., 2009, ‘‘Electrochemical reactions for electrocoagulation using iron electro-des.’’ Industrial and Engineering Chemistry Research, 48, pp. 2275–2228.

275Parga, J. R., Casillas, H., Vazquez, V., and Valenzuela, J. L., 2009, ‘‘Cyanide detoxification ofmining wastewaters with TiO2 nanoparticles and its recovery by electrocoagulation.’’Chemical Engineering and Technology, 32, pp. 1901–1908.Q3

Parga, J. R., Cocke, D. L., Valenzuela, J. L., Moreno, H., and Weir, M., 2005, ‘‘Arsenicremoval via electrocoagulation from heavy metal contaminated groundwater in La

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