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A NON-CYANIDE ROUTE FOR PROCESSING OF REFRACTORY GOLD ORES AND

CONCENTRATES

Bryn Harris and Carl White Neomet Technologies Inc. Dorval, Quebec, Canada

bryn@neomet.com

ABSTRACT

Despite several attempts to supersede its use, cyanide has been the reagent of choice for the recovery of gold and silver since it was first used in the mid 1800s. However, an increasing number of jurisdictions have banned its use, and relatively recent accidents, such as that into the Danube, are increasing environmental pressures to finding an alternative for cyanide. Chloride solutions are, and have always been, the basis of gold and platinum-group metals refining, but the volumes involved for refining are very small compared to those generated by large mining operations. The reasons that such a process has never been implemented at mine sites is partly due to the non-discriminatory nature of chloride acid attack compared to cyanidation, but mostly because there has never previously been a viable method, both technically and economically, for the recovery and recycle of hydrochloric acid, with the simultaneous control of iron, from such solutions. This paper describes a chloride-based flowsheet, wherein the key unit process is the breakthrough technology for the recovery and recycle of chloride ion. Results of a continuous miniplant run on a low-grade refractory gold ore are presented. It is shown that the chloride-based flowsheet is competitive with the conventional cyanide-based flowsheets, and that it has a number of other advantages.

Paper Presented at ALTA 2011

May 23-27, Perth, WA

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INTRODUCTION

Cyanidation has been the principal method for the recovery of gold and silver from ores and concentrates since it was first introduced in the mid-1800s, replacing the equally toxic mercury amalgamation method. However, despite creating a voluntary code of use for cyanidation [1], the industry is coming under increasing pressure to replace cyanidation, so that incidents such as the August, 1995 Omai Gold Mine [2,3] and the January, 2000 Baia Mare [4] cyanide spills in Guyana and Romania, respectively, no longer capture the news headlines. Some jurisdictions have already banned its use, and it is only a matter of time before others do so. Hence, it seems appropriate to accept that this is inevitable, and to develop an alternative to cyanidation.

Neomet Technologies Inc. (Neomet) has developed a chloride-based process for the recovery of not only gold and silver, but also associated value metals from refractory and carbonaceous gold ores, concentrates and intermediate materials. The process, therefore, does not require the use of cyanide to recover gold, nor does it require pre-treatment methods, such as pressure oxidation or ultra-fine grinding, to liberate gold from the refractory matrix. It is also possible to recover gold and silver separately or together, and allows for the recovery of other value metals such as copper, nickel, cobalt, rare earths, minor rare metals such scandium, gallium and indium, and the platinum group metals.

This paper reviews the background to the process, and describes some initial laboratory results on a variety of ores and a continuous, 1-L scale, miniplant run, treating 0.4 kg of low-grade ore per hour.

BASIC PRINCIPLES OF THE NEOMET GOLD PROCESS

Background

As noted above, the process is based on chloride chemistry, and there is considerable information in the literature to show that chloride systems are not novel, and are accepted practice.

Commercial Chloride Practice

It has long been known that high concentration chloride solutions can effectively leach gold from ores and concentrates, e.g. as practiced from the copper leach residues generated by the short-lived (6 years commercial operation) CLEAR Process commercialized by Duval Corporation in the 1970s [5]. It is also known that gold can be effectively and selectively recovered from dilute solutions using the ion exchange resin Amberlite XAD-7 [6,7,8 ,9]. Chloride solutions are, and have always been, the basis of gold and platinum-group metals refining, but the volumes involved for refining are very small compared to those generated by large mining operations. The reasons that such a process has never been implemented at mine sites is partly due to the non-discriminatory nature of acid chloride attack compared to cyanidation, but mostly because there has never previously been a viable method, both technically and economically, for the recovery and recycle of hydrochloric acid, with the simultaneous control of iron, from such solutions. Neomet has developed such a method, which forms the core of the Neomet Gold Process.

Chloride-based atmospheric leach processes, whilst not as common as sulphate-based ones, are not unique. Noranda operated the Brenda Leach Process, employing a high-temperature (105-110oC), high-strength chloride (CaCl2+NaCl+HCl) atmospheric leach of copper-molybdenum concentrates until the mine shut down in the 1990s [10]. The process essentially leached out all of the copper, lead and calcium from molybdenum concentrates, and the only real factor of concern was to maintain the lagging on the piping in winter to prevent the high strength brine from cooling and crystallizing [11].

Xstrata (formerly Falconbridge) has operated a chloride process at its nickel-cobalt refinery in Kristiansand, Norway, for over forty years [12,13,14]. There are also other operations, such as those operated by Jinchuan [15], Sumitomo [16] and SLN [17] in a mixed chloride-sulphate medium, which clearly demonstrate that chloride-based circuits are commercially feasible, and that any material-

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handling issues associated with a chloride environment can be and are being overcome. Furthermore, Vale (Inco) [18] has been using mixed chloride-sulphate electrolytes in nickel electrorefining operations for decades at its Port Colborne and Thompson operations.

In Quebec, Noranda’s Magnola plant [19] operated successfully for three years until price pressure on magnesium caused it to shut down (along with Dow’s plant in Texas and Norsk Hydro, also in Quebec). The chloride leach circuit operated at 100-105oC (atmospheric conditions) and always performed up to expectations. More recently, Rio Tinto Fer et Titane (QIT) has operated a high temperature (145-165°C) chloride leach plant (known as UGS) at Sorel for upgrading its titania slag [20,21] since the early 2000s. Finally, in January of 2011, Anglo American announced the building of a chloride-based pilot plant for treating nickel laterites in Brazil [[22,23].

Currently Operating Refractory Gold Extraction Processes

The modern industry standard for treating refractory gold ores is pressure oxidation (POX) at 220°C in an autoclave, followed by conventional cyanidation (CIL or CIP) of the leach residue. This technique was pioneered initially by Homestake, and then especially by Barrick at its Goldstrike operations in Nevada, in the 1980s. Prior to the development of the pressure leach, roasting was the standard, but this has fallen into disrepute because any arsenic contained in the ore was not fixed, but collected as a mixture of trisulphide and trioxide. However, even today, roasting remains the method of choice for carbonaceous ores [24]. With roasting, arsenic elimination ranged from 80-95%. Operations which have employed roasting include the Dickenson (now Goldcorp) [25] and Campbell mines [26] of the Red Lake camp in Ontario, Giant Yellowknife Mines [27, 28, 29] and the New Consort Mine [30] amongst others.

There are two obvious environmental problems associated with the treatment of such ores, namely sulphur and arsenic. Current process options have been roasting, pressure leaching or biological leaching [31,32,33,34]. Roasting generates sulphur dioxide, for which the only really effective recovery process is a sulphuric acid plant. Arsenic is converted to arsenic trioxide and recovered in a baghouse, and options for its disposal are limited. Placer Dome (now Barrick) was able to sell the arsenic trioxide from the operation of its Campbell mine roaster for a time, but was eventually forced to store the material on site [32-34]. The Nerco Con Mine in the Northwest Territories of Canada was also forced to store its arsenic trioxide, this time in permafrost, but eventually found it necessary to treat it to meet environmental regulations [35]. A further problem with the roasting option and conversion of sulphur dioxide to acid is the weak acid bleed from the acid plant. Such bleeds usually contain significant levels of toxic impurity elements, so that a treatment plant, as well as a disposal area is required [32].

The leaching (bacterial, pressure) options convert the bulk of the sulphide sulphur directly to either the element or to sulphuric acid, and subsequently to gypsum, and the arsenic to an iron-arsenic residue. Sulphur is not really a direct environmental problem with the leach option, and the formation of iron-arsenic residues are considered to be the most environmentally acceptable method of arsenic disposal [36,37,38]. Pressure leaching is generally regarded as giving the most stable form for arsenic disposal [36], and data from the pressure leach installation at the Campbell mine initially supported this assumption. However, more recent data [38] suggest that the disposal environment is equally as important, and in the wrong milieu, the arsenic minerals can break down again. The data from the residues generated by bacterial leaching are inconclusive at best.

From an environmental point of view, of the processes currently in operation, it would seem that the pressure leach option is the most favourable for treating arsenical refractory gold ores. Noranda concluded that there was relatively little difference from an economic viewpoint between the various options mentioned above [31]. Certainly, practical commercial experience suggests that pressure leaching is generally the process of choice, having been chosen by Placer Dome [32-34], Gencor at São Bento [39] and Barrick at Goldstrike [40]. Barrick continues to favour autoclave processing, with the world’s biggest vessels now being installed in the Dominican Republic at Pueblo Viejo [41], albeit at a cost of over $3 billion. However, there are no reported data of pressure leaching where there are unusually high levels potassium and sodium, and what the impact of their presence is on residue

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stability. Aluminium is also known to be a problem in laterite pressure leaching, where alunite formation has caused considerable scaling problems [42,43].

Newer Developments

Newer developments are generally variations of previously-tried methodologies, namely the bacterial leaching approach to arsenic, but which has not demonstrably addressed any of the previous concerns over the stability of the arsenic residues produced, and thiosulphate or thiourea leaching, neither of which offer any improvement as a replacement for cyanide, and which are inferior both economically and environmentally.

The only really new entrants are those being marketed by Eco Refractory Solutions, a joint venture between Globex Mining Enterprises and Drinkard Metalor [44], and Nichromet in a joint venture with Metanor Resources Inc. [45,46]. Very few details have been released on the nitric process, other than that high gold recoveries (95-98%) by CIL cyanidation have been achieved (head grades of the gold have not been disclosed, so quoting such recovery values can be misleading) following a pre-treatment (undefined) under atmospheric conditions to liberate gold and fix the arsenic in an environmentally-friendly manner. It is known that the pre-treatment involves nitric acid leaching, which is certainly effective for destroying arsenopyrite and forming scorodite. However, high consumption of nitric acid appears to be involved, and unless a cheap and effective method for regenerating the nitric acid is employed, then costs will be high. Further, nitrates are highly soluble, and unless close to 100% washing efficiency of the leach residues is achieved, then it is inevitable that nitrates will leak into the environment.

The process described in the Nichromet patent [45] is chloride/bromide-based, and would appear to be very similar to the Intec Process which has used for a number of years the concept of chloride/bromide leaching solutions [47]. The Nichromet Process requires a pre-treatment roasting step to eliminate the sulphur, presumably as SO2, which would require collecting and scrubbing. Any arsenic in the feed would report to the off-gas and have to be dealt with. The roasted ore is leached in sodium or potassium chloride using bromide as a catalyst, and the leach liquor is regenerated by electrolysis, again similarly to what Intec does. Gold and silver recoveries are quoted as being about 80%, using activated carbon. It is not said how the carbon is stripped, but presumably it is by cyanide. Impurities such as iron, aluminium, etc., are removed by neutralization with caustic generated during the electrolysis step. It is not said how these are filtered, but experience suggests that it is very difficult, since caustic-generated precipitates are usually very slimy and difficult to handle.

Neomet Gold Flowsheet

The Neomet Gold Process is able to treat a variety of feeds, ranging from ore to intermediates. It was developed as offshoots of the CCR Gold Refinery in Montreal [48,49] combined with a more generalised chloride-based metals extraction process, the core component of which is acid regeneration and hematite precipitation (termed “Hydrolytic Distillation”). A schematic of the flowsheet is shown in Figure 1. Both the specific acid regeneration and the more general gold process have been filed as PCT (Patent Co-Operation Treaty) applications [50,51]. Pyrohydrolysis, which has been the standard method of hydrochloric acid recovery for the past sixty years, particularly in the steel pickling and magnesia production industries, is both energy-intensive and non-selective, as well as recovering a relatively low-concentration hydrochloric acid (≤18% HCl). The Neomet Process recovers the acid as either gas, or very concentrated (30-35%) HCl, depending on the feed chloride concentration, as well as a very pure, low-volume hematite or hematite/alumina suitable for sale, or at the very least as a benign material for disposal, and requires much less energy than the pyrohydrolysis approach.

Leaching is carried out under atmospheric conditions at 105-110°C in recycled hydrochloric acid, with some recycle of the leach filtrate to increase both the gold and chloride concentration. For those feeds where there is a significant pyrrhotite or other reactive sulphide component, the leach is divided into two stages; a primary (reducing) leach to destroy all of the reactive sulphide, followed by a secondary (oxidising) leach to dissolve the gold. The oxidising leach is carried out to ensure a high (>750 mV)

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redox potential (ORP) of the final solution. This can be achieved by a variety of oxidants such as a small amount of nitric acid to act as an oxygen transfer agent, hydrogen peroxide or chlorine in association with a high concentration of cupric or ferric chloride. The choice of reagent depends very much on the characteristics of the material being treated.

Secondary Leach

Acid Recovery

Gold Recovery

Primary Leach

Gold Ore/Tailings

SL

S L

GOLD

Fe, Al Oxides

ResidueTo Disposal

HCl

Silver Recovery SILVER

S L

So, EnergyH2S

Bleed for Impurity Control

Ga, Sc Recovery Ga, Sc

Figure 1. Schematic of the Neomet Gold Process

During the oxidising leach, any arsenic minerals, such as arsenopyrite or enargite, are broken down. Arsenic concentration in the leach solution first increases, then decreases to a limiting value as scorodite (FeAsO4•2H2O) is formed. Scorodite is one of the most stable of the iron arsenates, and is acknowledged by the US EPA as being one of the BDAT (Best Developed Available Technology) options for the disposal of arsenic into the environment.

Gold is recovered from the leach liquor by passing the solution through the ion exchange resin. Once the resin is loaded, it can be eluted by very dilute hydrochloric acid to give a concentrated (>20 g/L) solution of gold chloride. Pure gold is recovered from this solution by any of the classical methods, although ferrous chloride (often formed in the primary leach) is preferred in this flowsheet. Any silver present in the leach liquor, together with any gold not recovered by the resin (there is normally none) can be recovered by adsorption on activated carbon.

An advantage of the Neomet Gold Process is that any other minor valuable metals present in the feed, such as scandium or gallium, can easily be recovered from the leach solution by a method analogous to that for gold. There are specific ion exchange resins for most of these metals, and they can be employed to recover the metals of interest after gold and silver recovery.

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Following gold recovery, the solution is then injected into an inert matrix bath at 180-190°C. The matrix compound remains fluid at such temperatures, and is the key component of the Neomet Acid Regeneration Process. Reactions are virtually instantaneous at these temperatures, with the iron hydrolyzing to form hematite and releasing concentrated HCl gas to be condensed or recycled directly to the leaching stage. Any arsenic remaining from the leaching stage is also transformed into scorodite in this stage.

Some ores contain other metals such as Al, Mg and base metals, especially copper, which dissolve to a lesser or greater degree in the leaching stages. These initially dissolve into the matrix solution as the iron reacts, and can be removed/recovered in a separate step by treating a bleed of the matrix.

Comparison of Neomet with Other Gold Processes

Although primarily conceived as an alternative to the use of cyanidation for refractory and carbonaceous ores and concentrates, the Neomet Gold Process has a number of advantages over competing processes, as summarised below in Table 2. The Neomet Process can treat either ore or concentrate, does not require fine grinding, and is able to treat both refractory and carbonaceous ores with high gold recovery, produces fine gold and silver directly without need for additional refining, and most importantly, does not require the use of cyanide.

EXAMPLES USING THE NEOMET GOLD PROCESS

Laboratory Tests

Preliminary batch laboratory tests have been carried out on a number of different feeds to demonstrate the flexibility of the process. Tests were conducted at 110°C, with the addition of an appropriate amount of HCl/tonne feed, the amount being determined by its chemical composition. Table 1.shows the result of leaching a complex, refractory arsenopyrite gold ore in two stages. The first, reducing leach stage was operated for two hours, care being taken to ensure that all of the pyrrhotite in the feed was destroyed (determined by no more H2S gas being evolved). The oxidation leach was then carried out, also for two hours, the end point being determined when no more arsenopyrite was visible floating on the surface of the slurry, and the ORP of the leach liquor had attained a value of 930 mV.

Table 1. Chemical Analyses of Batch Laboratory Leach Test on Refractory Whole Ore

Chemical Analysis (g/t or %)

Au Ag Sc Ga Fe As S g/t g/t g/t g/t % % %

Feed 1.41 0.14 12.8 16.2 3.68 2.74 0.74 Residue 0.01 <0.01 4.5 9.2 0.14 2.05 0.01

The results indicated that substantially all of the gold was leached, and that the arsenic was mostly in the residue at the conclusion of the test, which is where it should have been, since it first leaches, along with the iron, and then re-precipitates as scorodite. Furthermore, the data also indicated that all of the sulphur was eliminated. Of further interest for this particular feed is the fact that much of the scandium and gallium dissolved. These minor metals have significant value, and are relatively easily recovered from chloride leach solutions. Their recovery adds significantly to the economic evaluation of the Neomet Gold Process in that a credit of as much as $50/oz Au can be applied, depending on the price of the minor metal.

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Table 2. Comparison of Refractory Gold Processes

Operation Pressure

Oxidation (POX) Bacterial Leaching Roasting Albion/Activox Nitric Acid

Neomet Chloride

Feed Concentrate Concentrate Ore or Concentrate Concentrate, ultra fine grind

Concentrate Ore or Concentrate

Arsenic Type II mineral, ferrihydrite finish, stable

Ferrihydrite, stable only if sufficient Fe in solution at time of precipitation

Arsenic trioxide, separate treatment plant required plus separate source of iron to make stable residue

Ferrihydrite, stable only if sufficient Fe in solution at time of precipitation

Scorodite, ferrihydrite finish, stable

Scorodite, no finish necessary

Residue disposal

Dedicated area, non-reducing

Dedicated area, non-reducing

Dedicated area, non-reducing

Dedicated area, non-reducing

Dedicated area, non-reducing

Dedicated area, non-reducing

Effluent issues

None if no K or Na dissolves

Bacteria cultures leaking into water table

Weak acid from scrubbers or acid plant

None if no K or Na dissolves

Nitrates leaking into environment

Chlorides into environment. However, wash efficiency so high, that such is minimal

Alkali metals

Potential to form KOH and NaOH on lime neutralization, which will slowly decompose As minerals. Potential to form jarosites, consuming acid and iron.

Potential to form KOH and NaOH on lime neutralization, which will slowly decompose As solids. Potential to form jarosites, consuming acid and iron.

No issues Potential to form KOH and NaOH on lime neutralization, which will slowly decompose As solids. Potential to form jarosites, consuming acid

Potential to form KOH and NaOH on lime neutralization, which will slowly decompose As minerals. Potential to form jarosites, consuming iron, but helping control sulphate balance.

K and Na separated, chloride recovered. Not mixed with As solids, therefore no impact other than cost to control.

Aluminium Forms alunites, consumes acid, scaling issues

Will dissolve, consumes acid and lime

No issues Will dissolve, consumes acid and lime

Will dissolve, consumes acid and lime

Dissolves, has to be separated

Gold Cyanide in separate leach circuit,

Cyanide in separate leach circuit,

Cyanide in separate leach

Cyanide in separate leach

Cyanide in separate leach circuit, requires

Direct from chloride leach as

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recovery requires cyanide destruction.

requires cyanide destruction.

circuit, requires cyanide destruction.

circuit, requires cyanide destruction.

cyanide destruction. pure gold.

Silver recovery

Along with gold as doré metal

Along with gold as doré metal

Along with gold as doré metal

Along with gold as doré metal

Along with gold as doré metal

Separately as pure silver or together with gold

By-products

None None None None None Can recover any PGMs, Ga, Sc, etc.

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Table 3 shows the results of leaching a so-called “double refractory” ore containing not only arsenopyrite but also a significant (>1%) organic carbon content. Conventional pressure oxidation was ineffective, as it did not eliminate the preg-robbing ability of the carbon during subsequent cyanidation. The ore was leached in HCl at 100°C with chlorine gas addition.

Table 3. Chemical Analyses of Laboratory Leach Test on Double Refractory Whole Ore

Solids Chemical Analysis (g/t or %) Solution Analysis, mg/L

Au Fe As S Au Fe As g/t % % %

Feed 8.20 2.01 0.92 1.45 - - - 3 h 2.21 0.0243 0.0061 na 1.27 3353 1889 5.5 h nd 0.0121 0.0037 na 1.49 4617 2721

The results showed that all of the gold was recovered into the leach solution. No attempt was made during this test to force the re-precipitation of arsenic or to optimize the amount of chlorine added, since it was meant to determine whether oxidative chloride leaching could be effective in leaching gold from a refractory, carbonaceous ore. In this case, however, with the relatively low iron content, there would be no value in trying to produce a pure hematite product for sale, since the tonnage is so low, and it makes more sense to concentrate the arsenic into the hematite solids rather than the leach residue.

Miniplant Operation

The ore in Table 1 was also used as the feed and was ground to be 100% passing 100 mesh (150 microns). An overall view of the miniplant is shown in Figure 2. Two 8-hour day tanks were used as the primary (reducing) leach, and a series of five 1-L capacity reactors, each with a nominal residence time of one hour was used as the oxidizing leach. In this circuit, nitric acid was initially used as the oxidant, but later on, this was changed to chlorine. The use of chlorine is much easier, and serves the dual purpose of not only being the oxidant, but also of providing chloride make-up. The fifth reactor in the series served as a holding tank to act as a feed to the filtration system. The reacted slurry was then fed to a vacuum filter, with filtrate being pumped to the ion exchange gold recovery system. The cake was washed with 5% HCl and then twice with hot water. Vendor testing of the leach slurries was carried out during this campaign, and showed excellent filtration rates of >300 kg dry solids/m2/hr for both leach slurries.

Figure 2. Overview of Miniplant

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Temperature was maintained at 100-105°C in the reactors, which were jacketed in an oil bath to give uniform heating of the slurry. ORP was measured periodically in both the slurry and freeboard. With such an oxidizing environment, it is not possible to leave conventional ORP probes permanently immersed, as the reference platinum or gold electrode will dissolve over time.

The filtrate from the second leach was passed through the ion exchange circuit to recover the gold. No attempt was made to recover the silver in this particular circuit, since it was present in very low levels. After ion exchange, the liquor proceeded to the acid regeneration circuit for recovery of the acid and hematite. The principles and operation of this circuit have been described elsewhere [52].

Leaching Performance

Samples were taken at 2-hr intervals for analysis, and a daily composite also made. The following results are from the daily composites, showing the overall performance of the circuit. The elements of particular interest in this campaign were gold and arsenic.

Gold

Figure 3 shows the daily gold residue profiles throughout the campaign. In general, the profiles show most of the gold was leached with high recovery, even from such a low head grade.

Figure 3. Gold Residue Profile

The data suggest that, as far as gold is concerned, three reactors were more than sufficient to effect maximum gold recovery.

Arsenic

The solution profiles for arsenic are shown in Figure 4 and the solids in Figure 5. These clearly indicate that, as expected, the arsenic concentration first increased as the arsenopyrite was oxidised and broken down (reactor 2), and then decreased as the mineral scorodite was formed and precipitated. The final solution concentration of 100-200 mg/L is consistent with all arsenic mineral (as opposed to amorphous ferrihydrite) precipitation processes. However, in all other processes, the arsenic has to be polished with the ferrihydrite process in order to reduce its concentration to <1 mg/L, which requires an iron to arsenic molar ratio of at least 4, and neutralization to at least pH5 to be effective. In the Neomet flowsheet, this is not necessary, as the

0.00

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1 2 3 4 5

Au

Co

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Day1 Day2 Day4 Day7 Day8 Day9 Day10

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arsenic is effectively eliminated as scorodite or Type II mineral (there is a small amount of sulphate present due to the oxidation of pyrite and arsenopyrite) during the acid regeneration process.

Figure 4. Solution Arsenic Profiles

Figure 5. Solids Arsenic Profiles

However, unlike for gold, a longer residence time appears to be necessary for the complete set of arsenic reactions to occur, and to reduce the terminal arsenic concentration to 100-200 mg/L.

Gold Recovery

Gold can be recovered directly from the leach filtrate by ion exchange with Amberlite XAD-71. A composite solution was prepared from a miniplant run, and the loading results are shown below in

1 Formerly produced by Rohm & Haas, but now Dow.

0

50

100

150

200

250

300

350

400

1 2 3 4 5

As, m

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Day2 Day4 Day7 Day8 Day9 Day10 Day11

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Table 4. The data clearly show that the gold was loaded onto the resin, and that the resin was not fully loaded during this test (no breakthrough seen). The gold was quantitatively recovered from the solution, despite there being high levels of ferric iron. No iron, or any of the other major background elements such as aluminium, calcium, potassium or magnesium were loaded onto the resin.

Table 4. Ion Exchange of Recovery of Gold from Actual Leach Solution – Loading

Bed Volume Passed Gold Concentration, mg/LFeed solution 12.5 Displacement nd Bed Volume 1 nd Bed Volume 2 nd Bed Volume 3 nd Bed Volume 4 nd Bed Volume 5 nd Bed Volume 6 nd Bed Volume 7 nd Bed Volume 8 nd Bed Volume 9 nd Bed Volume 10 nd Bed Volume 11 nd Bed Volume 12 nd Bed Volume 13 nd Bed Volume 14 nd nd = not detected

Figure 6 shows a typical stripping profile from a fully loaded resin. The resin was initially loaded from a leach solution with a gold concentration of 200 mg/L, generated by leaching a different type of gold feedstock, one containing in excess of 2 ozs/tonne, since the gold recovered during the miniplant run was insufficient to achieve full loading. The resin was stripped with very dilute hydrochloric acid, and as can be seen, a concentrated strip solution assaying 23 g/L gold was obtained, a concentration factor of 100 in a single pass. In practice, it is possible to build up the gold concentration by repeated strip contacts, such that solutions with a concentration of >100 g/L can be attained.

Figure 6. Gold Stripping Profile (Resin loaded from 200 mg/L solution)

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10

15

20

25

1 3 5 7 9 11 13 15 17 19 21

Go

ld C

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, g/L

Bed Volumes

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Acid Regeneration and Iron Precipitation

The ore tested during the miniplant run was quite complex, and because of this, the leach solution contained significant quantities of aluminium, magnesium, potassium and calcium in addition to iron. This offered an opportunity, not previously tested, to determine the behaviour of all of these elements together. Testing was divided into two runs as a precaution in case modifications to the conditions needed to be made.

Figure 7 shows data from the first regeneration test at 190°C, where filtrate from the miniplant run was injected into a solvent matrix (500 mL in a 1-L reactor) at a rate of 6 mL/minute. It can be seen that a steady production of acid was achieved, demonstrating that the hydrolysis reaction occurs more or less instantaneously consistent with the feed rate. The total acid equivalent (in) was calculated from all the potential hydrolysable ions in the feed, but not all of them reacted, hence the divergence.

Figure 7. Acid Regeneration Test at 190°C - Acid In and Out

Figure 8 shows the data from the same test presented in a different way. Here it can be seen that the concentration of the produced HCl was consistent at around 230-240 g/L (6.5M HCl). Primary and secondary refer to two collection points, the secondary being a back-up to collect any acid not recovered by the primary condenser. The ORP values of 900-1000 mV are typical of acid regenerated in this manner, and are due to a small amount of chlorine being generated.

Table 5 shows the analyses of the final matrix solution, the composition of the solids produced, and the distribution of the elements between solids and solution. It is clear that the solids were comprised of primarily of iron and aluminium, with virtually all of the iron reporting to the solids. Potassium and calcium all reported entirely to the matrix solution, as was anticipated. At 190°C, minimal magnesium was found in the solids, and that present is likely due to inefficient washing.

The results from the second test were virtually identical, again with constant production of acid at approximately 230-240 g/L (6.5M HCl). Table 6 shows the analyses of final matrix solution, solids, and their deportment form the second test. Most of the elements in the second test deported much as they had done in the first test. It is clear from both tests that any arsenic in the feed reports to the solids. Under these highly oxidising conditions, the arsenic compound formed is likely to be scorodite, although its concentration was too small to be confirmed by XRD. No arsenic was detected in the acid, although it had been thought that some might be present since arsenic chlorides are quite volatile. The only noticeable difference in the two tests was that more iron reported to the matrix solution in the second test than in the first. However, this may be an anomaly, since iron remained at that concentration throughout the test, and did not increase unlike potassium, magnesium, calcium and aluminium, and may therefore have been due to the initial addition temperature being too low.

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9:30 10:30 11:30 12:30 13:30 14:30 15:30 16:30

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(g

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H+ In (g)

H+(Fe) In (g)

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Figure 8. Acid Regeneration Test at 190°C - Acid Concentration and ORP

Table 5. Matrix and Solids Analyses from First Acid Regeneration Test at 190°C

Element Feed, Final Matrix Solids Analysis, Distribution, % g/L (mg/L) g/L (mg/L) % Matrix Solids

Al 19.0 5.41 17.0 10.2 89.8 As (180) (2.4) 0.16 0.5 99.5 Ca 6.05 13.7 0.10 98.3 1.7 Fe 34.2 (45) 34.1 0.5 99.5 K 10.2 32.6 nd 100 0 Mg 13.7 35.4 0.08 99.3 0.7 Mn (456) 1.34 nd 100 0

Table 6. Matrix and Solids Analyses from Second Acid Regeneration Test at 190°C

Element Feed, Final Matrix Solids Analysis, Distribution, % g/L (mg/L) g/L (mg/L) % Matrix Solids

Al 18.5 2.02 17.0 9 90 As (183) (1.84) 0.16 2 98 Ca 5.61 3.08 0.10 93 7 Fe 34.7 5.85 34.1 19 81 K 10.3 12.7 nd 100 0 Mg 13.8 12.7 0.08 100 0 Mn (465) (389) nd 99 1

DISCUSSION AND CONCLUSIONS

The data presented in this paper have demonstrated that the Neomet Gold Process works on both refractory and carbonaceous ores, and is a technically viable alternative option to the currently-practiced pressure oxidation and roasting. Environmentally, it offers a unique alternative to the use of cyanide leaching, eliminating the need for cyanide-containing effluent treatment . The elimination of pyrrhotite also reduces or eliminates acid generation potential of tailings. High gold recovery was seen in the leaching step, with the ability to fix arsenic in both or either the leach and acid regeneration residues. Efficient and effective gold recovery directly from the leach solution

700

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0

50

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09:30 10:30 11:30 12:30 13:30 14:30 15:30 16:30

OR

P (

mV

)

[H+

] (g

/L)

Time of day

[H+]Cond (g/l) [H+]Prim (g/l)[H+]Sec (g/l) ORPCond (mV)ORPPrim (mV) ORPSec (mV)

15

was also seen with the ion exchange circuit, and recovery of gold from the resin in a more concentrated and pure solution demonstrated.

The key novel and unique acid regeneration step was shown to be effective, as it has been in laterite leaching [52]. A particular observation is that for ores containing a lot of gangue minerals that are easily leached, high levels of potassium, aluminium and magnesium report to the leach liquor. These elements are all problematic in any hydrometallurgical circuit, and if they can be largely removed initially by a simple physical separation process, be it for example flotation or dense medium without compromising gold recovery, then it is clearly a preferred method of operation since it would significantly reduce both capital and operating costs.

It is beyond the scope of this paper to give any breakdown of costs, but preliminary estimates of the costs for the processing plant alone, for a concentrate assaying 1 oz/tonne at 500 t/day, is of the order of $100-150 million (including mill and flotation plant), depending on the make-up of the ore and whether a single or 2-stage leach circuit is required. For whole ore, the costs are around $300 million, but again very much depend upon the make-up of the ore. Those for an ore as in Table 1 containing a significant pyrrhotite content and highly reactive gangue minerals, would be appreciably higher than for the carbonaceous ore of Table 3 which had neither of those.

Projected processing costs of $35-50/oz Au (before by-product credits) were estimated for the concentrate case, again depending on the chemical make-up of the feed. Furthermore, the Neomet Process, unlike cyanidation, offers the potential for recovering additional value metals, such as scandium, indium and gallium, as these elements leach from the ore in a chloride medium, and thus can be recovered as an additional revenue stream.

The preferred oxidant is chlorine gas, since the oxidation requirement is usually more or less similar to the chloride make-up required from the loss in the filter cakes. Per unit of chloride, chlorine gas is cheaper than hydrochloric acid (in eastern Canada), and thus there would be savings in this respect. It should be noted that, as mentioned earlier, hydrochloric acid/chlorine leaching of sulphide minerals is known practice, and has been carried out at Kristiansand in Norway for over forty years [12,13,14].

REFERENCES 1. International Cyanide Management Institute, International Cyanide Management Code for the

Gold Mining Industry, 2010, http://www.cyanidecode.org/cyanide_use.php

2. Dam Review Team, Preliminary report on technical causation: Omai Tailings Dam Failure, Report Submitted to Guyana Geology and Mines Commission, November 16, 1995, URL: http://antenna.nl~wise/uranium/mdgr.html

3. Omai Gold Mine Resumes Full Operation, Caribbean Daylight, April 1, 1996.

4. Wikipedia, 2000 Baia Mare Cyanide Spill http://en.wikipedia.org/wiki/2000_Baia_Mare_cyanide_spill

5. A.T. Wilson, An Economical Method for the Recovery of Gold from the Sulphur Containing Residue of a Hydrometallurgical Process, in Complex Sulphides, Processing of Ores, Concentrates and By-Products (A.D. Zunkel, R.S. Boorman, A.E. Morris and R.J. Wesely, Editors), TMS, 1985, pp. 143-148.

6. Bryn G. Harris et al., Recovery of Gold from Acidic Solutions, US Patent 5,028,260 issued on July 2, 1991.

7. G.B. Harris, New Developments in Precious Metals Refining: An Ammonia-Free Process for Platinum and Palladium, Presented at Precious Metals 1994, Proceedings of the Eighteenth International Precious Metals Conference, Vancouver, British Columbia, Canada, June, 1994, (J.G. Peacey, Editor), IPMI, Allentown, PA, p. 259

8. A. Filcenco-Olteanu, et al., Statistical model of gold recovery from chloride media using Amberlite XAD-7, Chem. Bull. "POLITEHNICA" Univ. (Timisoara), Volume 55(69), 1, 2010, pp. 64-67.

16

9. R. Radulescu, et al., New Hydrometallurgical Process for Gold Recovery, Chem. Bull.

"POLITEHNICA" Univ. (Timisoara), Volume 53(67), 1-2, 2008, pp. 135-139.

10. P.H. Jennings, R.W. Stanley and H.L. Ames, Development of a Process for Purifying Molybdenite Concentrates,. in Proceedings of Second International Symposium on Hydrometallurgy, (D.J.I. Evans, Editor), AIME, New York, 1973, p. 868.

11. R.W. Stanley, Noranda Research Centre, Personal Communication, April, 1985.

12. P.G. Thornhill, E. Wigstol and G. Van Weert, The Falconbridge Matte Leach Process, Journal of Metals, 23(7), 1971, p. 13.

13. E.O. Stensholt, H. Zachariasen and J.H. Lund, The Falconbridge Chlorine Leach Process, in Nickel Metallurgy, Volume I - Extraction and Refining of Nickel, (E. Ozberk and H. Marcusson, Editors), CIM Montreal, 1986, p. 442.

14. E.O. Stensholt, H. Zachariasen, J.H. Lund and P.G. Thornhill, Recent Improvements in the Falconbridge Nickel Refinery, in Extractive Metallurgy of Nickel and Cobalt, TMS-AIME, Warrendale, 1988, p. 403.

15. Yang Yuhua and Meng Xianxuan, Operating Practice and Technical Developments in Nickel Refining and Cobalt Recovery at Jinchuan Non-Ferrous Metal Company, in Electrometallurgical Plant Practice, (P.L. Claessens and G.B. Harris, Editors), Proceedings of the 20th Annual CIM Hydrometallurgical Meeting, Pergamon Press, 1990, p.253.

16. M. Fujimori, N. Ono, S. Itasako and I. Fukui, Solvent Extraction in Sumitomo's Cobalt Refining Process, Paper presented at Cobalt 80, 10th Annual CIM Hydrometallurgical Meeting, Edmonton, Alberta, October 26-28, 1980.

17. J.M. Demarthe, L. Gandon and M. Goujet, Hydrometallurgical Method For Treating Nickel Mattes, Canadian Patent 1,126,201, to Société Métallurgique le Nickel-SLN, 22 June, 1982.

18. J.R. Boldt and P.ueneau, Winning of Nickel, Van Nostrand Company, 1967.

19. P. Ficara, et al., Magnola: A Novel Commercial Process For The Primary Production Of Magnesium, Light Metals 1996, (M. Avedesian, R. Guilbault and D. Ksinisk, Editors), CIM, Montreal, 1996, p. 183.

20. C. Coscia and J.A. Kozinski, Pyrohydrolysis of an Al-Fe-Mg-Cl Solution – Characteristics of the Transformation Process, in Chloride Metallurgy 2002 Volume 2 (E. Peek and G. van Weert, Editors), Proceeding of the 32nd Annual CIM Hydrometallurgical Conference, CIM, Montreal, 2002, p. 683.

21. M-C. Patoine, A. De Mori, K. Borowiec and C. Coscia, Pyrohydrolysis of a Calcium and Magnesium Bearing FeCl2 Leach Liquor, in Chloride Metallurgy 2002 Volume 2 (E. Peek and G. van Weert, Editors), Proceeding of the 32nd Annual CIM Hydrometallurgical Conference, CIM, Montreal, 2002, p. 699.

22. M. Pelser, J.D.T. Steyl and J. Smit, Development of the Anglo Research Nickel (ARNi) Process for the Treatment of Laterite Ores, Proceedings of Hydrometallurgy of Nickel & Cobalt 2009 (J.J. Budac, R. Fraser, I. Mihaylov, V.G. Papangelakis and D.J. Robinson, Editors), 48th CIM Annual Conference of Metallurgists, 39th Annual Hydrometallurgy Meeting, Sudbury, Ontario, Canada, August 23-26, 2009, p. 409.

23. Anglo American plc., Annual Report 2010, p. 39.

24. Ivanhoe Mines, Altynalmas’ Kyzyl Gold Project Estimated Average Annual Gold Production of 368,000 Ounces Confirmed by Independent Pre-Feasibility Study, June 30, 2010.

25. K.P. Wright, Fluid Bed Roasting Practice in the Red Lake Camp, CIM Bulletin, August 1961, p. 595.

26. J.O. Burckle, Arsenic Emissions and Control: Gold Roasting Operating, Environmental International 6, 1981, p. 443.

27. E.O. Foster, The Collection and Recovery of Gold from Roaster Exit Gases at Giant Yellowknife Mines Ltd., CIM Bulletin, June 1963, p. 469.

28. L. White, Giant and Con Underpin Yellowknife Economy, E&MJ 183(2), February 1982, p. 79.

17

29. L. Connel, et al., Roasting Process at Giant Yellowknife Mine, Paper presented at 20th CIM

Conference of Metallurgists, August 1981.

30. D.W. Penman, Metallurgical Aspects of the Treatment of Refractory Ores from Barberton, in Impurity Control and Disposal, Proceedings of the 15th Annual CIM Hydrometallurgy Meeting, Vancouver, August 1985, p. 8-1.

31. G.D. Hallett, W.F. Luinstra and R.W. Stanley, The Economics of Treating an Arsenopyrite Refractory Gold Concentrate, in Precious Metals 1990, (Donald Corrigan, Editor), Proceedings of the 14th IPMI Conference, San Diego, CA, June 1990, p. 105.

32. J. Frostiak, R. Raudsepp and R. Stauffer, The Application of Pressure Oxidation at the Campbell Red Lake Mine, Presented at Randol Forum '90, Squaw Valley, CA, September 13-15, 1990.

33. K. Norris, Process Improvements at Placer Dome Campbell Mine, in Proceedings - 24th Annual Meeting of the Canadian Mineral Processors, Ottawa, Ontario, January, 1992, Paper 22.

34. J. Frostiak and B. Haugrud, Start-up and Operation of Placer Dome's Campbell Mine Gold Pressure Oxidation Plant, Mining Engineering, August 1992, p. 991.

35. J. Geldart, R. Williamson and P. Maltby, in Hydrometallurgy, Theory and Practice, Proceedings of the Ernest Peters International Symposium, Hydrometallurgy 30(1-3), p. 29.

36. G.B. Harris and E. Krause, The Disposal of Arsenic from Metallurgical Processes: Its Status Regarding Ferric Arsenate, Paper presented at Paul Queneau International Symposium on the Extractive Metallurgy of Copper, Cobalt and Nickel, Denver, CO, February 1993.

37. G.B. Harris, The removal and stabilization of arsenic from aqueous process solutions: past, present and future, Proceedings of Minor Elements 2000: Processing & Environmental Aspects of As, Sb, Se, Te, Bi, SME 2000 Annual Meeting & Exhibit, February 28 - March 1, 2000, Salt Lake City, Utah, USA, p. 3.

38. Bryn Harris, The removal of arsenic from process solutions: theory and industrial practice, in Proceedings of Hydrometallurgy 2003 - International Symposium in Honor of Professor Ian Ritchie, (C.A. Young, A. Alfantazi, C. Anderson, A. James, D. Dreisinger and B. Harris, Editors), Vancouver, CIM/TMS/SME, August 2003, p. 1889.

39. T.M. Carvalho, , et al., Start-up of the Sherritt Pressure Oxidation Process at São Bento, in Precious Metals 1989, (Bryn Harris, Editor), Proceedings of 13th IPMI Conference, Montreal, Quebec, June 1989, p. 319.

40. A.W. Bolland, H.J.H. Pieterse and R.H. Colquhoun, American Barrick Goldstrike operation in Elko, Nevada - a summary of the operation and future expansions, in Proceedings - 24th Annual Meeting of the Canadian Mineral Processors, Ottawa, Ontario, January, 1992, Paper 4.

41. Barrick Gold Corporation, Pueblo Viejo Brings Innovative, Environmentally-Responsible Mining Technology to the Dominican Republic, Beyond Borders, September 8, 2010.

42. P. Perdikis and V.G. Papangelakis, Scale Formation in a Batch Reactor Under Pressure Acid Leaching of a Limonitic Laterite, Canadian Metallurgical Quarterly 37(5), 1998, pp. 429-440.

43. V.G. Papangelakis, B.C. Blakey and J. Kambossos, Behaviour of Aluminum During Direct Acid Leaching of Limonitic Laterites, in Impurity Control And Disposal In Hydrometallurgical Processes (B. Harris and E. Krause, Editors), 24th Annual CIM Hydrometallurgical Meeting, Toronto, Canada, August 21-24, 1994, CIM, pp. 315-325.

44. Globex Mining Enterprises Inc., Gold Recovery Technology Proven Successful Again,

http://www.marketwire.com/mw/release.do?id=1379554, January 12, 2011.

45. Jean-Marc Lalancette, Gold and Silver Recovery from Polymetallic Sulphides by Treatment With Halogens, US Patent 7,537,741 (to Nichromet Extraction Inc.), May 26, 2009.

46. Metanor Resources Inc., Metanor Preliminary Discussions with Nichromet, January 13, 2011.

47. Intec Limited, Australia, http://intec.com.au/

18

48. G.B. Harris and R.W. Stanley, Hydrometallurgical Treatment of Silver Refinery Anode Slimes,

in Precious Metals 1986 (U.V. Rao, Editor), Proceedings of the 10th International Precious Metals Institute Conference, Lake Tahoe, Nevada, June 1986, IPMI, Allentown PA, 1986, p. 239.

49. A. Bilodeau, G.B. Harris, C.A. MacDonald, K. Hooper and R.W. Stanley, Silver Refinery Anode Slimes Treatment at the CCR Division of Noranda Inc., in The Electrorefining and Winning of Copper (J.E. Hoffman, R.G. Bautista, V.A. Ettel, V. Kudryk, R.J. Wesely, Editors), TMS-AIME 1987, p. 527.

50. Bryn Harris and Carl White, Process for the Recovery of Metals and Hydrochloric Acid, PCT International Application No. PCT/CA2011/000141, February 4, 2011 (Priority Date February 18, 2010).

51. Bryn Harris and Carl White, Process for the Recovery of Gold from an Ore, PCT International Application No. PCT/CA2011/000143, February 4, 2011 (Priority Date February 4, 2010).

52. Bryn Harris and Carl White, Recent Developments in the Chloride Processing of Nickel Laterites, ALTA Ni-Co-Cu 2011, Perth, WA, May 23-27, 2011.