US20140076735A1

US20140076735A1 - Electrorecovery of gold and silver from leaching solutions by simultaneous cathodic and anodic deposits

Info

Publication number: US20140076735A1
Application number: US13/993,247
Authority: US
Inventors: Gretchen Terri Lapidus Lavine; Alejandro Rafael Alonso Gómez; Ricardo Benavidez Perez; Carlos Lara Valenzuela; Javier Alejandro Silva Alonso
Original assignee: Universidad Autonoma Metropolitana (UAM); Servicios Administrativos Penoles SA de CV
Current assignee: Universidad Autonoma Metropolitana (UAM); Servicios Administrativos Penoles SA de CV
Priority date: 2010-12-13
Filing date: 2011-12-09
Publication date: 2014-03-20
Also published as: RU2013132451A; AU2011341844A2; EP2653590A2; BR112013014874A2; WO2012081952A4; MX2010013717A; PE20140494A1; WO2012081952A2; CA2821421A1; AU2011341844A1; WO2012081952A3; CN103380234A; JP2014505788A; CO6801793A2

Abstract

This relates to mining and mineral or materials treatment industries that deal with gold and silver. Specifically, it is related to the process to recover gold and silver from thiosulfate or thiourea solutions, with an electrolysis that occurs simultaneously on both the anode and cathode. Increased velocity and greatly reduced energy consumption are obtained in relation to those found in conventional electrolytic cells.

Description

CROSS-REFERENCE TO RELATED APPLICATIONS

This application is a 371 U.S. National Stage of International Application No PCT/MX2011/000151, filed Dec. 9, 2011, which claims priority to Mexican Patent Application Serial No. MX/a/2010/013717, filed Dec. 13, 2010. The disclosures of the above applications are incorporated by reference herein.

FIELD OF THE INVENTION

The present invention is related to the mining industry for treatment of minerals and materials which contain gold and silver. Specifically, it is related to a process to recover gold and silver, from leaching solutions with a simultaneous anodic and cathodic electrodeposition process, after which the poor solution is recycled back to the leaching stage.

BACKGROUND

The recovery of gold and silver from their minerals has been performed by various methods; among the most employed are pyrometallurgical treatments, in which upon the addition of a considerable amount of energy, part of the mineral is oxidized, in this manner liberating the precious metals. This great amount of energy is the principal inconvenience of the process, which in the end reflects on the operation costs.
On the other hand, the hydrometallurgical methods are characterized for their high selectivity and relatively low reagent and energy costs. Gold and silver has been obtained by one such method for over 100 years, using cyanide and oxygen as a complexing agent and an oxidant, respectively. Despite the high efficiency of this system, the treatment of complex minerals, as well as environmental restrictions, has encouraged research on other leaching systems that could compete with cyanide, without its disadvantages.
Thiosulfate, in the presence of copper, and the combination of thiourea with formamidine disulfide (Poisot-Diaz, M. E., Gonzalez, I. and Lapidus, G. T. (2008), “ Effect of Copper, Iron and Zinc Ions on the Selective Electrodeposition of Dorée from Acidic thiourea Solutions”, Hydrometallurgy 2008, Eds. C. A. Young, P. R. Taylor, C. G. Anderson y Y. Choi, Society for Mining, Metallurgy and Exploration, Inc. (SME), Littleton, Colo., U.S.A., ISBN: 978-0-87335-266-6, pp. 843-848 and Alonso-Gómez, A. R. and Lapidus, G. T. (2008), “Pretreatment for Refractory Gold and Silver Minerals before Leaching with Ammoniacal Copper Thiosulfate”, Hydrometallurgy 2008, Eds. C. A. Young, P. R. Taylor, C. G. Anderson y Y. Choi, Society for Mining, Metallurgy and Exploration, Inc. (SME), Littleton, Colo., U.S.A., ISBN: 978-0-87335-266-6, pp. 817-822.) are two chemical systems that leach gold and silver from minerals for which cyanidation has proved to be inefficient. In this same manner, it was shown possible to recover gold and silver metals in both systems using direct electrodeposition (A. Alonso. G. T. Lapidus and I. Gonzalez, A strategy to determine the potential interval for selective silver electrodeposition from ammoniacal thiosulfate solutions Hydrometallurgy, Volume 85, Issues 2-4, March 2007, Pages 144-153); However, this recovery was accomplished in geometrically complex reactors (F. C. Walsh, C. Ponce de Leon and C. T. Low, The rotating cylinder electrode (RCE) and its application to the electrodeposition of metals, Australian Journal of Chemistry, 58, (4), 246-262 and A. Alonso, G. T. Lapidus and I. González, Selective silver electroseparation from ammoniacal thiosulfate solutions using a rotating cylinder electrode reactor (RCE), Hydrometallurgy, Volume 92, Issues 3-4, June 2008, Pages 115-123), with an energy consumption that renders un-attractive from an economic and financial standpoint.
At this point, it is important to mention a characteristic of the thiourea and thiosulfate systems: both complexing agents can oxidize at potentials near the reduction potential of silver (FIGS. 1 and 2). The diagrams of both ligands with gold are similar. This originates the formation of a narrow potential region where Ag(I) and Au(I) ions are soluble and because of this, both the leaching as well as the electroseparation conditions should be controlled with precision. This could imply a great disadvantage with respect to other systems and has motivated the use of membrane reactors, in order to avoid contact of these solutions with the anode.

SUMMARY

One objective of the present invention is to provide a method to separate gold and silver from thiosulfate or thiourea solutions by simultaneous anodic and cathodic electrodeposition, increasing in this manner the velocity of the process. Another is to accomplish this with a minimum affectation of the solution composition, so that it may be recirculated back to the leaching stage. Yet another is to promote efficient energy use. Other objectives and advantages that apply the principles and are derived from the present invention may be apparent from the study of the following description and diagrams that are included here for illustrative and not limitative purposes.
The present invention is intended to solve the problem of gold and silver separation from thiosulfate and thiourea leaching solutions, providing an improvement over the traditional electrochemical reactors now in use. This improvement is characterizes by a novel process to simultaneously deposit metals in on the anode and cathode in a one compartment reactor, using a commercial copper sheet as the anode and a titanium sheet as the cathode.
The conditions which permit this technique to operate were chosen from the analysis of FIG. 1, where a region of the soluble complex Ag(S2O3)23- is observed within the metallic silver stability zone. When the potential is decreased below −110 mV, the Ag(I) species is reduced to Ag0, in a typical electrolytic process. However, the most interesting aspect of this diagram is when the potential is less negative than −50 mV, where part of the thiosulfate oxidizes, destabilizing the soluble complex and forming metallic silver. The present invention takes advantage of this phenomenon and has not been previously reported for this or other ligands.
The application of the simultaneous anodic-cathodic electrode-position of gold and silver allows more efficient use of the electrical energy in electrochemical reactors of simple geometry without a membrane; additionally, the separation process occurs in less time than that required in conventional electrochemical reactors. In order to better understand the characteristics of the invention, the following description is accompanied by diagrams and figures, which form an integral part of the same and are meant to be illustrative but not limitative and are described in the following section.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a Pourbaix-type diagram in which the predominance zones for the soluble species Ag(S2O3)23- (thiosulfate-silver complex) and metallic silver Ag0 are shown.

FIG. 2 is a Pourbaix-type diagram in which the predominance zones for the soluble species AgTu3+ (thiourea-silver complex) and metallic silver Ag0 are shown.

FIG. 3 shows a leaching-electrodeposition scheme for obtaining gold and silver which utilizes the present invention.

FIG. 4 is a diagram showing a recirculation system which includes the electrochemical reactor.

FIG. 5 is a schematic diagram of the electrochemical cell in which the simultaneous anodic and cathodic deposits are achieved.

FIG. 6 is a graphic representation of the change in silver concentration with leaching time.

FIG. 7 is a graphic representation of the change in silver concentration with electrolysis time where there is simultaneous anodic and cathodic electrodeposition.

FIG. 8 is a graph that compares the change in silver concentration for

leaches

1, 2 and 3 with the same solution.

FIG. 9 shows the comparison of the silver concentration during

electrolysis

1, 2 and 3 with the same solution.

FIG. 10 shows a comparison of XRD spectra for the anodic deposit obtained after the electrolysis and for pure metallic silver.

DETAILED DESCRIPTION

The simultaneous electrodeposition process, referred to in the present invention, is illustrated in FIG. 3. A thiosulfate or thiourea solution, rich in gold and silver ions, originating from the leaching stage (100) and after having been filtered (200), is introduced into the electrochemical reactor (300). Once the electrodeposition has finalized, the cathode (312, FIG. 5) and the anode (313, FIG. 5) are removed from the reactor and mechanically abraded to remove the gold and silver metals. The solution is then recirculated back to the leaching stage (301). The electrodeposition is performed in a recirculation scheme, illustrated in FIG. 4, in which the solution is charged to the reservoir (320) from which it is pumped (330) to the electrochemical reactor (310) and then returned by gravity to the reservoir.

EXAMPLES

Example 1

To better understand the invention, one of the many experiments is detailed as an example, which employs a system such as that schematized in FIGS. 3 to 5. A 60 cm2 (exposed geometrical area) titanium plate was used as the cathode and a copper plate with the same exposed area was the anode. As shown in FIG. 3, the first stage is gold and silver leaching from the mineral or concentrate, using a thiosulfate solution, in this case, whose composition is presented in Table 1. The pH was adjusted to 10.0 with NH4OH.

TABLE 1

Composition of the leaching solution

	Component	Composition (mol/L)

	(NH4)2S2O3	0.2
	CuSO4	0.05
	EDTA	0.025
	(NH4)2HPO4	0.1

The solutions were prepared with reagent grade chemicals using deionized water (1×1010 MΩcm−1). 500 mL of this solution was placed in contact with 3.75 g of a flotation concentrate, with a particle size less than 10 μm, containing 21 kg/ton of silver. After six hours in continuous agitation, the solution was separated from the solid by filtration and placed in a reactor such as that represented in FIGS. 4 and 5.
During the electrodeposition, a flow of 1.1 L/min was used with a cell voltage of 100 mV; with this voltage, the potential at the cathode was −260 mV versus the normal hydrogen electrode (NHE), which is adequate to obtain a selective silver deposit on this electrode.
FIG. 6 shows a graphic representation of the silver concentration with respect to the leaching time. A maximum value was attained in 120 minutes, after which time the concentration remained relatively constant.
The change in silver concentration during the electrolysis is shown in FIG. 7. Within the first 15 minutes a sharp descent is observed, which then gradually decreases to values below 10 mg/L. The current registered throughout the experiment was 0.01 A, which together with the cell voltage translates to 0.004 W-h. Considering that the deposited mass of silver was 0.065 g, the energy consumption was 0.062 W-h per g of deposited silver.
After finalizing the electrodeposition, the solution was recycled back to the leaching stage, where it was contacted with fresh unleached concentrate, under the same conditions as described previously. The entire procedure was repeated until three full cycles were completed.
FIG. 8 shows a graphic representation of the leaching results for all three cycles; an increase in the leaching velocity and the maximum silver concentration may be observed in the second and third leach, relative to the first, possibly due to the stabilization of the equilibria between the thiosulfate and the Cu(II) and Cu(I) ions. On the other hand, the second and third electrolyses (the dashed and dotted lines of FIG. 9) show similar tendencies to that of the first (solid line), only differentiable by the initial value, which depends on the previous leaching stage. In all three cases, the values reached below 10 mg/L in approximately 4 hours.
These results clearly show that the thiosulfate solution can be recirculated after the electrodeposition stage, back to the leaching stage, at least three times without reconditioning or make-up. Additionally, during the three electrolyses, the current maintained a constant value of 0.01 A, conserving the same energy expenditure as the first cycle. Anode consumption was negligible after three electrodeposition cycles.
Finally, it is important to mention that X-ray diffraction analysis of both the anodic and the cathodic deposits showed that they consisted exclusively of metallic silver. FIG. 10 compares the XRD spectra for the deposit obtained from the anode, at the end of the electrolysis and the corresponding spectra for pure metallic silver. As can be observed, the anodic deposit corresponds to metallic silver; indicating that oxidation of thiosulfate is forming only soluble species, such as tetrathionate, dithionate or even sulfate, and is not contaminating the deposit.

Claims

1. (canceled)

2. A method for recovering silver or gold from a mineral comprising performing electrolysis from a thiosulfate or thiourea leaching solution simultaneously depositing the silver or gold on anode and cathode surfaces by operating in potential zones that permit silver or gold reduction at the cathode and ligand oxidation at the anode.

3. The method according to claim 2, further comprising combining the leaching solution with the mineral in a leaching stage and circulating the leaching solution to an electrochemical reactor including the cathode and the anode.

4. The method according to claim 3, further comprising recirculating the leaching solution to the leaching stage after the silver or gold is deposited on the anode and cathode surfaces.

5. The method according to claim 3, further comprising filtering the leaching solution after the leaching stage, and prior to electrolysis.

6. The method according to claim 2, further comprising abrading the cathode and anode mechanically to remove the silver or gold metals.

7. The method according to claim 2, wherein the anode comprises copper and the cathode comprises titanium.

8. The method according to claim 2, wherein the potential zone comprises a cell voltage of 100 mV, wherein the potential at the cathode is −260 mV versus a normal hydrogen electrode.

9. A method for recovering a metal from a flotation concentrate comprising:

(a) transferring a leaching solution to the flotation concentrate in a leaching stage to generate a metal-leaching solution;

(b) filtering the metal-leaching solution;

(c) circulating the filtered metal-leaching solution to an electrochemical reactor comprising an electrochemical cell having an anode and a cathode; and

(d) performing electrolysis on the metal-leaching solution, the electrolysis resulting in a deposition of the metal onto the anode and cathode.

10. The method according to claim 9, wherein the metal is gold or silver and the leaching solution comprises thiosulfate or thiourea.

11. The method according to claim 10, wherein the metal is silver and the leaching solution comprises 0.2 M (NH₄)₂S₂O₃, 0.05 M CuSO₄, 0.025 M EDTA, and 0.1 M (NH₄)₂HPO₄.

12. The method according to claim 9, wherein the anode comprises a copper plate with a 60 cm²exposed geometrical area and the cathode comprises a titanium plate with a 60 cm²exposed geometrical area.

13. The method according to claim 9, wherein the flotation concentrate has a particle size of less than 10 μm.

14. The method according to claim 9, wherein electrolysis comprises a cell voltage of 100 mV, wherein the potential at the cathode is −260 mV versus a normal hydrogen electrode.

15. The method according to claim 9, wherein the leaching stage comprises at least 2 hours of continuous agitation.

16. The method according to claim 9, further comprising collecting the metal deposited on the anode and cathode by abrading the anode and cathode.

17. The method according to claim 9, further comprising recirculating the leaching solution to a subsequent leaching stage after performing electrolysis.

18. A method for recovering silver from a flotation concentrate comprising silver, comprising:

(a) transferring a leaching solution to the flotation concentrate in a leaching stage to generate a silver-leaching solution;

(b) filtering the silver-leaching solution;

(c) circulating the filtered silver-leaching solution to an electrochemical reactor comprising an electrochemical cell having a copper anode and a titanium cathode;

(d) performing electrolysis on the metal-leaching solution, the electrolysis resulting in a deposition of the metal onto the anode and cathode;

(e) recovering the silver from the electrochemical reactor by mechanically abrading the anode and cathode, wherein leaching solution remains in the reactor;

(f) recirculating the leaching solution to the leaching stage; and

(g) repeating steps (a) through (f) a plurality of times.

19. The method according to claim 18, wherein electrolysis comprises a cell voltage of 100 mV, wherein the potential at the cathode is −260 mV versus a normal hydrogen electrode.

20. The method according to claim 19, wherein the electrolysis further comprises a current of 0.01 A, which together with the cell voltage translates to 0.004 W-h.

21. The method according to claim 20, wherein the leaching solution comprises 0.2 M (NH₄)₂S₂O₃, 0.05 M CuSO₄, 0.025 M EDTA, and 0.1 M (NH₄)₂HPO₄.