CA2418689A1

CA2418689A1 - Gold and silver recovery from polymetallic sulfides by treatment with halogens

Info

Publication number: CA2418689A1
Application number: CA002418689A
Authority: CA
Inventors: Jean-Marc Lalancette
Original assignee: Nichromet Extraction Inc
Current assignee: Individual
Priority date: 2003-02-11
Filing date: 2003-02-11
Publication date: 2004-08-11

Abstract

A method for the extraction from a polymetallic sulfide ore of precious metals consisting of gold and silver as well as base metals selected from the group consisting of copper zinc, lead, nickel, cobalt, mercury and tin is disclosed. The method comprises the steps of oxidizing the ore using lean air at temperatures ranging from about 400 to about 600° to produce an oxidized ore and beaching the oxidized ore using a brine solution containing free chlorine, chlorine combined with water, and a catalytic amount of a bromide salt producing a solution of metal chlorides and a barren solid.

Description

TITLE OF THE INVENTION
GOLD AND SILVER RECOVERY FROM POLYMETALLIC
SULFIDES BY TREATMENT WITH HALOGENS
FIELD OF THE INVENTION
[0001] The present invention relates to gold and silver recovery from polymetallic sulfides by treatment with halogens.
BACKGROUND OF THE INVENTION

[0002) The use of chemical agents, particularly halides, for the recovery of gold and silver is quite ancient. It was noted very early that the adjunction of sodium chloride to mercury improved the performances of the amalgamation process. This discovery translated into the Patio or ~azo processes, which were implemented on an empirical basis from the early 1600's in Central and South America more than 150 years before the discovery of elemental chlorine by Scheele in 1774. The Patio method involved the digestion of a finely divided gold ore with mercury and sodium chloride, in the presence of air and moisture over a three month period. The values were then collected by further leaching with mercury, followed by amalgam distillation (T.
Egleston, The IVletallurgy of Silver, Gold and Mercury in the United States, Vol.
1, p. 261, John Wiley, 1887).

[0003] Later, in the late 1700s, chloridizing roasting followed by barrel amalgamation was developed in Central Europe as an improved method for gaining access to precious metals from sulfide ores. This process called upon a high temperature treatment of the goldlsilver ores in the presence of sodium chloride, air and steam, in order to transform the precious metal sulfides into their corresponding chlorides. The gold and silver was then recovered either by amalgamation or cementation on pure copper (T. Varley et al, IJ.S. Bureau of Mines, Bulletin N° 211, 1923). However, it was discovered that the high temperature chloridizing of gold or silver ores resulted in very important losses of values by volatilization. In some cases these losses reached 80 °!o or more of the precious metal content (S.B. Christy, Transaction of the American Institute of Mining Engineering, Vol. 17, p. 3, 1888).

[0004] It appeared that the presence of pyrites or iron sulfides contributed significantly to the volatilization of gold and silver during the high temperature chtoridization with NaCI (S. Croasdale, The Engineering and Mining Journal, August 29, 1903, p. 312). It was finally established that the mechanism explaining these losses involves the formation of a mixed chloride of gold and iron (AuCl3 ~ FeCt3), which is highly volatile at the chloridization temperatures (J. A.
Eisele et al. U.S. Bureau of Mines, Report N° 7489).

[0005] Elemental chlorine dissolved in water, as introduced by Plattner around 1850, constituted an alternative to high temperature chloridization.

However, this process was characterized by low efficiency in addition to collecting the gold by amalgamation.

[0006) The general characteristics of the various processes involving chlorine, either as elemental chlorine or as chlorides, either at ambient temperatures or at high temperatures, were not attractive. The yields obtained with these processes were generally iow (often below h0 %) and the values were collected as amalgams or as cemented products on copper or iron. In addition, complex procedures were involved in order to obtain the precious metals in a pure form. The environmental impacts of such operations, where large amounts of sulfur are disposed with the tailings, would have been completely unacceptable by current standards.

[0007) The advent of cyanide extraction in 1916, tem~inated the extraction of gold by various forms of chloridation. The cyanide process calls upon the action of a cyanide salt such as sodium cyanide on gold in the presence of oxygen, to give a soluble gold salt (Eq. I):
2 Au + 4 NaCN + 1l2 02 + HZO -~ 2 NajAu(CN)2] + 2 Na~H

[0008) The gold can then be recovered from the cyanide complex by the action of excess zinc (Eq. II):
2 Na[Au(CN)2) + Zn,~cegs, --~ Na2[Zn(CN)~, + 2 Au [0009) Under the best circumstances, gold recovery can be as high as 98 %. This process calls for a contact time of one to three days at near ambient temperature in the presence of air.
[0010 In some instances the cyanide process performs very poorly.
Ores refractory to cyanide extraction can be grouped under the general term of polymetailic ores. In such ores, one fnds small amounts of base metals such as copper or zinc, typically 0.1 % Cu or 0.3% Zn. Such srnall amounts qualify the ore as of very low grade for the production of copper or zinc. If such a polymetallic ore body contains some gold (for example, 4 gTT Au or Ag or a mixture of both}, the cyanide extraction process does not perfiorm well. The poor performance is due to the base metals, either copper or zinc, (as well as silver), having a much stronger ability to form complexes with cyanide than gold. In fact, this inherent property is used to recover gold from a pregnant solution by zinc treatment following cyanide extraction (see Eq. Il). The base metals will consume all the cyanide present and the gold extraction will only begin after all the available base metals, as well as silver, have been dissolved. Because of the excessive consumption of relatively costly cyanide, this process for recovering gold is uneconomical.
[0011] There exist other types of gold ores refractory to cyanide extraction, namely arsenopyrites and carbonated ores. With certain types of arsenopyrites, the gold exists as a solid solution in the crystal lattice. In order to gain access to gold, the structure of the crystal lattice (sulfide} must be completely broken up.

[0012] In order to break up the crystal latkice, and to gain access to gold contained therein, a reagent tike cyanide cannot be used and a more aggressive approach is required. Vllith carbonated gold ores, the carbon carried by the ore acts as an adsorbent for the cyanide complexed gold, and will cause it to reprecipitate it in situ. Another approach is thus required in order to extract the gold content of these types of ores.
[0013] Polymetallic ores constitute complex mixtures of sulfides. The tailings discarded as a result of gold and silver extraction using the cyanide process, as well as by other methods, still contain very substantial amounts of sulfur.. This sulfur is prone to bio-oxidation (Thiobacilius fer~-ooxidans), and the resulting drainage is quite acidic and toxic due to its metallic content.
[0014] The spent cyanide solutions, kept in ,large ponds following gold recovery, represents a substantial environmental hazard and has recently created major disasters in Guyana and Central Europe, thus restricting the use of the cyanide process in many areas.
[0015] In the last twenty years, chloridation has been reconsidered as a process for extracting base metals such as copper, nickel or silver. The lntec base Metal process (J. Moyes and F. Houllis, Chloride Metallurgy 2002, Vol.
II, p.
577, Canadian Institute of Mining, Metallurgy and Petroleum) constitutes a typical example. This process calls for the digestion at 85°C, over a period ranging from 12 to 14 hours, of the sulfides of copper or zinc in a concentrated brine solution (250 g/l NaCI) comprising a cupric mixed halide (BrCl2)Cu prepared electrolytically. The mixture is aerated and the copper is collected as cuprous chloride. The cuprous chloride is decomposed at the cathode to elemental copper by electrolysis upon regeneration of the mixed halide of copper (Eq. III):
2 CuFeS2 + 5 BrCl2 -+ 2 Cu+~ + ~ Fe+3 + 4 S° + 5 Br + 10 CI' [0016] The above described chloridation process was reported as also extracting gold, if present. However, the requirement of recycling copper so as to have the cupric/cuprous system needed to oxidize iron and sulfur, makes this approach very cumbersome when the main concern is gold recovery rather than copper recovery. Further, the electroiytical oxidation of sulfur via the cupric salt, which is regenerated by electrolysis, is a very costly process rendering the treatment of a gold ore not having a very large gold content uneconomical.
Finally, the presence of elemental sulfur in the tailings is a potential source of acid drainage.
[0017] Another chloridation process called Platsol, was reported as being very efficient for the recovery of base and precious metals from sulfide ores (C.J. Ferron et al, Chloride MetaNurgy 2002, Vol. I, p. 11, Canadian Institute of Mining, Metallurgy and Petroleum). This process involves a pressure oxidation in the presence of oxygen and sulfuric acid, in an autoclave at a temperature above 200°C. The implementation of such a technique is very capital-incentive, calling for titanium autoclaves and a source of pure oxygen. The operation of this equipment is also prone to problems due to scaling ofi the reactor, complicating heat transfer. The sulfur resulting from the operation is in an innocuous form, i.e.
a hydrated iron sulfate (jarosite). The high capital and operating costs render this approach unattractive for polymetailic sulfides having a modest gold content.
(00'18] Other techniques such as the Plint process (fbid, Vol. I, p. 29) or, the Ito process (Ibid, Vol. I, p. 69), are techniques used for the recovery of gold and silver from sulfides, by oxidation with ferric chloride in concentrated brine. The ferrous chloride is re-oxidized to ferric chloride by chlorine alone or by exposure to air and hydrochloric acid (Eq. IV):
2 PbS ~ Ag2S ~ 3 Sb2S3 + 24 FeCl3 -~ 24 FeCl2 + 2 PbCiz + 2 AgCI + 6 SbCl3 +
12 S°
[0019] In these processes, sulfur is again oxidized electrochemically via the oxidation of ferrous chloride by chlorine or HCI. As explained previously, such an approach is costly for the recovery of gold or silver from sulfide ores, because of the electrochemistry involved. Elemental sulfur is again discarded with the tailings, generating a potential source of acid drainage.
[0020] There thus remains a need for an improved process for the recovery of gold and silver from polymetailic ores.
[0021 ] The present invention seeks to meet these and other needs.
[0021] The present invention refers to a number of documents, the content of which is herein incorporated by reference in their entirety.

[0022] The present invention relates to a process for the recovery of gold and silver from polymetallic ores, characterized by Bow operational and cost investments.
[0023] The present invention relates to a process for the recovery of gold and silver from polymetallic ores, characterized by being carried out at atmospheric pressure and at low temperatures prior to leaching.
[0024] The present invention also relates to a process for the recovery of gold and silver from polymetailic ores, characterized by producing tailings devoid of elemental sulfur, sulfides, or soluble sulfates and by fast reaction rates allowing for high rates of treatment.
[0025] In addition, the present invention relates to a process for the recovery of precious metals such as gold and silver, as well as base metals such as copper, nickel, cobalt, zinc, tin and lead, in addition to relating to the safe removal of sulfur, arsenic and mercury as well as to the disposal of iron, chromium, aluminum and titanium in an inert and insoluble form.
[0026] Further scope and applicability will become apparent from the detailed description given hereinafter. It should be understood however, that this detailed description, while indicating preferred embodiments of the invention, is given by way of illustration only, since various changes and modifications within the spirit and scope will become apparent to those skilled in the art.
DRIEF DESCRIPTION OF TFiE DRAU111NGS
[0027] In the appended drawings:
[0028] Figure 'I is a block diagram illustrating the various steps of the process of the present invention;
[0029) Figure 2 is a block diagram illustrating the various steps of the sulfur removal aspect of the process of the present, invention;
[0030] Figure 3 is a block diagram illustrating the various steps of the gold and silver recovery aspect of the process of the present invention;
and [0031] Figure 4 is a block diagram illustrating the various steps of the base metal recovery aspect of the process of the present invention; and [0032] Figure 5 is a schematic illustration of an electrolytic cell used in the process of the present invention.
DETAILED DESCRIPTION OE THE INVENTION
[0033] In a broad sense, the present invention relates to a new process for the recovery of precious metals such as gold and silver, as well as base metals such as copper, nickel, cobalt, zinc, tin, mercury, and lead. The present invention also relates to the safe removal of sulfur, arsenic and mercury as well as to the disposal of iron, chromium, aiuminurr~ and titanium in an inert and insoluble form. This is achieved at considerably lower cost than with the current chloridation or cyanide processes, by avoiding sulfur oxidation by electrochemical means. The process of the present invention is very time efficient, of the order of a few hours, and is carried out at atmospheric pressure and at near ambient temperatures. The process allows for the separation of the precious metals as well as the base metals from the common metals, while recycling the reagents and releasing only inert waste materials into the environment.
[0034) Gold and silver, along with base metals such as copper, zinc, lead, tin, nickel, cobalt and mercury can be recovered from polymetallic sulfides in yields well above 90 % by the process of 'the present invention comprising the following steps:
(0035] (a) Oxidizing the polymetallic ore using lean air having about 10% 02, at a temperature ranging from about 400 to about 600°C reducing the sulfur content of the ore to about 0.5 % S (as sulfide) or less. 'The resulting S02 is fixed by calcium carbonate as calcium sulfite, which auto-oxidizes to calcium sulfate dihydrate (gypsum). This results in the elimination of sulfur in a manner compatible with environmental regulations.
(0036) (b) slurring the sulfur free ore with a near-saturated (275 to 300 g/l} aqueous solution of sodium chloride (brine), and adding a solution of chlorine/HCl/hypochlorous acid such that the precious metals and the base metals are chlorinated and dissolved in the strongly complexing brine milieu.
The chloridation reaction is significantly accelerated by the presence of a catalytic amount, less than one percent of the halides present in the brine, of bromide ion.
The chlorinelHCl/hypochlorous acid solution, containing a catalytic amount of bromine, is generated by circulating a portion of the brine solution used to slurry the oxidized ore through the anodic compartment of an electrolytic cell, at a rate sufficient to dissolve the chlorine in the brine solution. Following the slurring operation, the ore is maintained in suspension in the acidic halogenated brine at a temperature ranging from about 35-45°C by slow stirring, without aeration, for a period of 2-3 hours for most ores, and up to 5 hours for exceptionally refractory ores. After separating the barren solid followed by washing with brine, the combined filtrate and rinsings are circulated over activated carbon for gold and silver recovery.
[0037] ~c) treating the soiution deprived of precious metals with a sodium hydroxide solution raising the pH to about 3 to 4.. The sodium hydroxide required to achieve this partial neutralization is produced by circulating the initial brine solution through the cathodic compartment of the electrolytic cell. The caustic sodium hydroxide solution is generated concomitantly at the cathode, in a stoEChiometric ratio, with the chlorinelhydrochloric acid/hypochlorous acid solution produced at the anode of the electrolytic cell. Raising the pH to about 3-4 induces the precipitation of iron, aluminum, chromium and titanium as insoluble oxides of these metals, in various hydrated forms. These oxides are filtered and washed with brine. Raising the pH of the resulting filtrate to values above 8, induces the precipitation ofi the base metals such as copper, zinc, Bead, tin, nickel and cobalt as a base metal concentrate.
[0038] Any arsenic, often present in significant amounts in polymetallic sulfides, is eliminated along with the sterile solids after the digestion as ferric arsenate, an insoluble and inert arsenic salt. Mercury, if present, is largely recovered with the flue dusts after oxidation, and any remaining traces of this metal are Iixiviated by the chlorinated brine, and recovered as an amalgam upon the removal of the base metals.
[0039] The brine solution, following the removal of the metals, is recirculated for further leaching. The sterile solids are rinsed with water and the rinsings concentrated by evaporation, using waste heat from the sulfide oxidation step. The concentrated rinsings, along with the brine solution, are then recycled so as to prevent salt losses or salt release into the environment.
Sulfur removal tFir~ure 1) [0040] The gold and/or silver containing ore, additionally comprising small amounts (0.1-2.0%) of base metals such as Cu, Zn, Pb, Sn, Ni, and Cu, is a sulfide or complex sulfide. The ore may further incorporate one or more other common metals such as iron, aluminum, titanium, chromium, as well as elements such as arsenic, antimony or bismuth. Mercury is occasionally also present in the ore.

[0041 The ore is reduced to a particle size of less than about 140 mesh by standard methods known in the art. The sulfur content of the ore, which can be as high as 15 %, is reduced to about 0.5% or less (as sulfides) by controlled oxidation in a reactor or kiln (2). The reactor or kiln provides for a control of the oxygen content in the reaction chamber. A relatively low oxidation temperature, typically ranging from about 400 to about 600°C, is very advantageous since it prevents any sintering of the material and generates a solid product having a large surface area and having good reactivity. This treatment is much preferred to standard roasting where temperatures as high as 1200°C have beeri observed. Such high reaction temperatures induce much sintering and volatilization. Standard roasting involves the free burying of the sulfides in the presence of excess air.
[0042] The contr~I of the low oxidation temperatures is achieved by recycling part of the lean air back to the reactor. This allows for the oxygen content in the reactor to be maintained at values not exceeding 10°/~
02. It is important to prevent any sodium chloride present in the ore from being oxidized. It is well known that sodium chloride contaminations as low as 0.01 percent, can induce significant volatilization of gold and silver.
[0043] The gas stream from the oxidation reactor is cooled in a settling chamber, allowing for the collection of volatile oxides or products generated during the oxidative treatment such as arsenic oxide, traces of zinc oxide, and metallic mercury if present in the starting ore. Dusts carried mechanically from the fines in the reactor are also collected in the settling chamber. The amount of solids collected is generally small, less than one percent of the weight of the ore treated. The solids thus collected can be recovered arid used for recuperation of values such as As203 or mercury, or they can be safely disposed of in sealed containers. The gas at the exit of the settling chamber, essentially composed of SO2 and lean air, is partly redirected back to the oxidation reactor for oxygen level control, and partly directed to a S02 scrubbing unit. The SO2 is adsorbed using a finely divided limestone slurry (200 mesh), allowing for the transformation of essentially all of the S02 (about 98%) into calcium sulfite, which auto-oxidizes to calcium sulfate dehydrate or gypsum. Gypsum is a very stable and inert product representing a definitive solution for the safe disposal of sulfur. It can be used as a building material in the production of Portland cement or as land fill. The water, following the dewatering of the gypsum, is recirculated back to a water thank.
Since gypsum is a dehydrate, there is a net consumption of water in the scrubbing process. The gases freed of SO2, are vented through a flue duct.
[0044] In the first step of the process therefore, the ore was made more reactive towards leaching, and essentially ail of the sulfur initially present has been disposed of in a safe and environmentally compatible manner. The present approach constitutes an economically attractive alternative to the presently available methods. The cost of electrochemically oxidizing 1 % of sulfur in one metric ton of sulfide ore is BUS 4.71 per unit percent of S2- per ton with a KWh at $US 0.09 per kilowatt and with an efficiency of 80 °/~. The cost of oxidizing the sulfide content of an ore containing 10°l0 S2- to elemental sulfur, using an electrochemically-produced reagent such as chlorine, would be in the best case scenario $US 47.10 per ton of ore for power only. The controlled oxidation of the sulfur content using lean air can be done at 10 % or less of that cost; and transforms the sulfur into a safe and environmentally disposable form.
The electrochemical oxidation process leaves elemental sulfur in the tailings generating a potential source of acid drainage.
Goldlsilver recovery (Figure 2) [0045] The recovery of gold and silver from the oxidized ore is achieved by leaching with a reagent derived from elemental halogens. The ..
halogens (Br2, CI2) have significantly different behaviors towards gold.
Bromine can readily dissolve gold at room temperature, even in the absence of water ~Kruss and Schmidt, Berichte der Deutschen Chemichen Gesellschaft, 20, 2634, 1887). Gold, on the other hand, is inert to dry chlorine at room temperature, and the attack of this gas on gold requires the presence of water and slight heating (Voigt and Biltz, Z. anorg. Chem., 133, 277, 1924). Even though bromine is the more reactive reagent with gold, chlorine is more electronegative (VILM.
Latimer, the Oxidation State of the Elements, pp. 56 and 62, Prentice Hall, 1952):
Ci- ~ Ci2 (-1.359 V);
Br -~ Br2 (-1.07 V].
[0046] It is possible to take advantage of this reactivity difference to accelerate gold leaching from the oxidized ore, if a catalytic amount of a bromide is introduced into the leaching solution. The leaching solution is a brine solution having a high concentration of sodium chloride, i.e. from 275 to 300 g/I of NaCI.
Lower salt concentrations yielded lower percentages of silver recovery, when silver was associated with gold in the oxidized ore. A portion of the concentrated brine solution, also containing a trace (~-3 gll) of NaBr, is circulated in the anodic compartment of an electrolytic cell, at an appropriate rate, so as to dissolve the halogen liberated at the anode. As mentioned above, the bromide ion will be reduced first, followed by some chloride ions so as to give a mixture of halogens dissolved in the brine solution. The brine solution containing dissolved CI2 and Br2 is mixed with fresh brine from a brine tank to provide a volume of liquid necessary to form a 20% slurry with the oxidized ore in a reactor kept at 35-45°C. The slurry is slowly stirred in order to prevent settling of the ore. The reacting mass was not aerated since aeration was neither improving the reaction rate nor the reaction yield, instead it resulted in the loss of dissolved halogens. Due to the trace amounts of bromine in the system, the gold leaching process is believed to involve the initial formation of gold tribromide (Eq. V):
2 Au + 3 Br2 --~ 2 Au,Br~
~0047~ The gold tribromide is then believed to be transformed, because of the stronger oxidizing capacity of C12, into gold trichloride with the concomitant regeneration of elemental bromine (Eq. VI):
2AuBr3+3012--~2AuCl3+3Br2 [0048] A similar type of reaction is obtained for silver, the high concentration of sodium chloride allowing the solubilization of the silver halides by complexation.
[0049] In the course of the leaching reaction, the other ions are similarly solubilized, and exist at their maximum valency; copper as cupric chloride, iron as ferric chloride, tin as stannic chloride, and arsenic as arsenate (As+5). Particularly with arsenic, the strong oxidizing environment leads to the precipitation of all the arsenic as an insoluble and inert ferric arsenate (Eq. VII):
Fe3+ + AsO4 3 -~ FeAs04 [0050] The pH of the reaction mixture drops below 0.1 as the leaching reaction proceeds. This strong acidification is an indication of the reaction of chlorine with water (Eq. VIII):
Ha0 + C12 ~ HCI + HOCI
[0051] The presence of hypochlorous acid could account for the observed chloridation of gold by chlorine in the presence of water. A similar equation can be written to describe the behavior of bromine, which is in equilibrium with hydrobromic acid and hypobromous acid. The solubilized species can therefore be seen as a mixture of chlorides and hypochlorides, which eventually end up as chlorides when the hypochlorous ion decomposes with the concomitant evolution of nascent oxygen (Eq. IX):
t-IOCI ~ HCI + 11202 [0052] The production of nascent oxygen accounts in part for the very strong oxidizing capability of the system without aeration of any sort.
[0053] The duration of the leaching, at 35-~.5°C in the reactor, ranges from 2 to 3 hours. With exceedingly refractory ores it is necessary to extend the contact time to about 5 hours. Following the leaching, the slurry is filtered or centrifuged, producing a pregnant solution and a waste or sterile solid.
[0054] The sterile solid was first rinsed with brine in order to recover any held-up values in the cake, followed by washing . with water to recover any salt. The so-obtained tailings contain arsenic as an. iron arsenate, and are free of sulfur and of soluble base metals. The pregnant solution is circulated over carbon to collect the gold and silver. Following the recovery of gold and silver from the carbon by known methods, these precious metals are obtained by electrowinning or other standard techniques. The goldlsilver-free solution is then recovered to be further treated so as to collect the base metals.
Recovery of base metals (Figure 3) [0055] The base metals to be obtained from the leaching of gold-bearing polymetallic sulfiides are of two categories. The first category contains metals of relatively high commercial value, often obtained by pyrometallurgical operations. This category contains metals such as nickel, cobalt, copper, zinc, lead, tin and mercury. The second category contains metals of low economic value, and comprises predominantly iron with considerably smaller amounts of aluminum, titanium, chromium and traces of the p-bloc elements.
[0056] In order to isolate these two types of base metals, sodium hydroxide is generated in the cathodic compartment of the electrolytic cell.
The sodium hydroxide solution is accumulated in a caustic tank and is then used to raise the pN of the previously produced barren solution, devoid of gold and silver, leaving the carbon columns, from below 1 to about 3-4. At a pH ranging from about 3-4, any iron existing as Fe~3 is instantaneously precipitated by hydrolysis as a hydrated iron oxide. Titanium, aluminum and chromium react similarly within this pH range. The hydrated oxides are removed by filtration. The solids are rinsed with brine in order to recuperate any base metals of values held up in the solid cake, followed by washing with water to remove any traces of salt. The salt-free mixture of oxides is then discarded as an insoluble and inert material of little or no commercial value.
[0057] The solution obtained from the filtration and the brine rinsings contains the base metals of value. IVlercury if present, is separated by amalgamation on pure copper. The ply of the mercury-free solution, pH between 3-4., is further raised using an additional portion of the sodium hydroxide solution to values above 8, causing all of the base metals {Ni, Co, Cu, Zn, Pb, Sn) to precipitate as oxides or hydrated oxides. The oxides are removed from the mixture by filtration and are rinsed with water to remove any traces of salt, to provide a concentrate of metals having significant commercial value. The brine, being free of metals, is recycled back to the fresh brine reservoir. The rinsings are concentrated by evaporation so as to give a brine solution of appropriate concentration, and which is also recycled back to the fresh brine reservoir.
[0058) The implementation of the process of the present invention, using a large variety of gold-bearing polymetallic sulfdes, provides for the recovery of gold and silver in high yields, essentially always above 90 °/~ and frequently above 95 %. The process of the present invention also provides for the recovery in high yields of the base metals of commercial value, frequently above 95 %.
(0059) ~f all the base metals of little commercial value, iron is generally the predominant one. Following the oxidation of the sulfides at 400-600°C, the resulting iron oxide is quite inert and no more than about 20-25 °/~ of this iron is leached, thus significantly decreasing the power consumption of the process. In fact, for a KVllh costing US$ 0.09, and with an efficiency at the electrolytic cell of 80 %, each percent of iron in the ore would cost US$ 1.00 for power to take care of, and each percent of base metals such as copper or zinc in -the ore would cost US$ 2.36 for power to extract. Thus, for an ore having 1 °/~
copper and 8 % iron, the value of recovered copper (US$ 16.50 at US$ 0.75/Ib for copper) covers all the electrolytical power costs (US$ 10.36) plus a fair reserve and no power imputations have to be made against the gold and silver values recovered.

[0060) Using the process of the present invention, polymetallic ores containing gold and/or silver which do not qualify for base metals extraction either because of a low base metal content, problems of enrichments by flotation or other restrictions, can be treated economically from the return generated by the base metals in order to collect the precious metals. Consequently, the process of the present invention provides for an attractive alternative to the currently available technologies, allowing the treatment of ores or tailings previously not attractive, at a profit.
[OOf1) The recycling of the brine solution, and the disposal of sulfurs arsenic and metal oxides as stabte and inert solids, reduces the environmental impacts of the operation to a minimum. Furthermore, the implementation of the process of the present invention at near-ambient temperatures and atmospheric pressure, reduces the investment per unit weight of ore to very competitive values. Finally, the low temperature oxidation of sulfur being an exothermic process, the energy consumption at that level is minimal and much lower than the corresponding electrochemical oxidation of sulfide to elemental sulfur.
[0062) The process of the present invention was implemented using a variety of polymetallic ores and tailings containing gold and silver.
Example 1 [0063) A Canadian ore sample (90 g) from the Sudbury (Ontario) area containing 4.5 g!T Au, 8 gIT Ag, 0.1 % As, 7.5 % S, 5.5 % Fe, 0.1 % Ni, 0.008 Co and 0.5 % Cu was reduced to a particle size of about 140 mesh and heated at 585-600°C in an atmosphere composed of N2 (50%) and 50 % air, over a period of two hours in a VycorT"" tube heated externally in a LindbergT"" furnace.
The temperature was measured inside the mass being oxidized. The external heating was reduced when the oxidation began at around 400°C.
[0064, A small amount of a white deposit, arsenic oxide, could be observed at the discharge side of the VycorT"" tube. The color of the oxidized material changed from black to brown and the weight loss during the process was 12 %.
[0065 A sample of the oxidized material (25.0 g) was placed in a three-necked one liter flask, along with 500 g of water, 1508 of sodium chloride and 1.2 g of sodium bromide. The suspension was stirred magnetically and the flask was closed so as to exclude air from entering the system.
[0066 The slurry was extracted from the flask through one of the necks using a peristaltic pump, and was subsequently circulated through the anodic compartment of an electrolytic cell, operating with a brine solution having the same concentration as the brine solution in the flask (anode of graphite, operation at 2.5 V). The anodic fluid was returned to the flask after dissolving chlorine. The cell was operated on and off in such a manner as to maintain a slight reddish coloration in the flask indicative of the presence of free bromine.

(0067] The reaction flask was maintained at 40°C for a period of 2.5 hours after which it was filtered on a Buchner funnel. The solid was rinsed three times with a brine solution containing 300 g/l NaCI. The mixed filtrate and rinsings were very acid, having a pH below 1Ø The acidic filtrate and rinsings were then treated with 30 g of carbon (NoritT"" R~3515 so as to collect gold and silver.
The sterile solid was then rinsed with water to completely remove any traces of brine {negative test to AgN03), dried at 110°C {16.8 g) and submitted to elemental analysis. The elemental analysis indicated that 96% of the gold and 94% of the silver initially present in the oxidized material, were leached out and then adsorbed on the carbon.
[0068] The barren solution following contacting with carbon was combined with the aqueous rinsings and was submitted to elemental analysis.
The solution was found to be essentially free of gold and silver, and contained 99% of the extracted iron, 98 % of the nickel and copper and 91 % of the cobalt present in the starting oxidized ore sample. Adjusting the pH at 3.5 with sodium hydroxide, resulted in the precipitation of the iron. Further raising the pH
to 8.5 precipitated the nickel, cobalt and copper. The brine, being essentially free of metals, is available for further use It was noted by elemental analysis that the bromine content in the brine did not change during the process, taking into account the dilution induced by the rincings. Further is was f~und that the gold and silver content, following treatment, was below 0.05 gIT and 0.16 g/T
respectively, while 23% of the iron was extracted.

[0069) The process was repeated using several types of polymetallic sulfides containing gold, silver or both, along with base metals of value. All the operational parameters were the same as in Example 1, except for the duration of the digestion. Those results are reported in Table !.

i ~Y ~ O ~ N

N ChO aO Oa O O

a a cn C~ O7 ~ O?
L

d ~ O
O O O ~

O O

o N tn N c0 a0 1~.. cp Q o m o 0 0 0 ~ o O ~Yc0 cDtC~~ t0 I~. M
O~ O ~ CO

~ 05 O

C

~ O

L

O O O O tn tty O tC'y tf>

D ~ N M M N M tp C~ M

L
~
~

v~ O
,~1 - t() M M tf~ O
O ~

r. M
~ Cn h-r COI~ d' CD r U

L M ci~W M

_ M
O r- '-~_ 1 t(7 O
T' a ~

Z

O

C O ~'-O M ~ r ~

tn ~ M
~ ~-r m ~ 00 0 _ O ch ,ON ~ M_ N
' 0 ' U 0 Q o o co ~ 'n N

O

N o N tt~ M a0 tc~ C~~6 o O O

v O h.
' O y 1 M O ~, M ~ M
U Q > N N
M

M M Lt> t1? N

O

a9 C

.U .U

X ~ ? c E E > ~~ f.
~ O

C9 ~ t~ m U o . , n ' ~ ~ ~
~

c c a L

c * ~ m cLo . cB

(JD ~ ~ r..~ O * U O
* ~ U

O td .~J .C N C (8 U cn ~ 'O tOp O fB N
Z
.

N U ~ ~ ~
n ~ ~ ~ o c 7C p N M V'tn CO t~ 00 ~ p7 w-[0071, Although the present invention has been described hereinabove by way of preferred embodiments thereof, it can be modified, without departing from the spirit and nature of the subject irwention as defined in the appended claims.

Claims

1. A method for the extraction from a polymetallic sulfide ore of precious metals consisting of gold and silver as well as base metals selected from the group consisting of copper zinc, lead, nickel, cobalt, mercury and tin, the method comprising:

(a) oxidizing the ore using lean air at temperatures ranging from about 400 to about 600°C to produce an oxidized ore;

(b) leaching the oxidized ore using a brine solution containing free chlorine, chlorine combined with water, and a catalytic amount of a bromide salt producing a solution of metal chlorides and a barren solid;

(c) filtering the barren solid producing a filtered solution, and contacting the filtered solution with carbon producing a first solution and carbon loaded with silver and gold;

(d) removing a first set of base metals from the first solution by raising the pH, producing a second solution;

(e) removing a second set of base metals by raising the pH of the second solution.