RECOVERY OF GOLD FROM REFRACTORY AURIFEROUS IRON-CONTAINING SULPHIDIC MATERIAL

Abstract: In a process for the recovery of gold from refractory auriferous iron containing sulphidic material which compris­ es subjecting an aqueous feed slurry of the sulphidic material to a pressure oxidation step (14), and recovering the gold from the oxidized solids, e.g. by cyanidation (22), problems associated with elemental sulphur formation in the oxidation step are alleviated by adding inert mateial, prefer­ ably by recycle of oxidized solids, to the feed slurry to provide an initial pulp density in the range 30-60% by weight.

Description: This invention relates to the recovery of gold from refractory auriferous iron containing sulphidic material, for example ore or concentrate. It is known that the recovery of gold from refractory auriferous sulphidic material by cyanidation is improved if the material is first subjected to a pressure oxidation treatment to liberate gold from refractory material, see for example United States Patent No. 2,777,764 (Hedley et al) issued January 15, 1957. In the pressure oxidation treatment it is desirable to fully oxidize the sulphide sulphur to the sulphate form for effective liberation of the gold. The sulphidic minerals present are usually predominantly arsenopyrite and/or pyrite and may also include appreciable amounts of pyrrhotite as well as less amounts of base metal sulphides such as zinc, lead and copper sulphides. Elemental sulphur may be formed as an intermediate or primary oxidation product in the pressure oxidation treatment and, since the pressure oxidation treatment is usually carried out at temperatures of from about 120° to 250°C, more commonly from about 140° to about 200°C, the sulphur is present in a molten state. Molten sulphur has a strong tendency to wet and/or coat many of the sulphides, with resultant formation of agglomerates of sulphur and unreacted sulphides, and can consequently severely limit oxidation and gold liberation. This is especially the case in continuous operations in which the agglomerates may build up to the point where they remain in and build up in the reaction vessel. Also, the presence of elemental sulphur is detrimental to subsequent gold recovery by cyanidation, not only because of increased comsumption of cyanide but also because molten sulphur has an affinity to collect gold and hinder access of the cyanide solution to the gold. Although the prior art teaches use of various additives, such as lignosulphonates or quebracho in the pressure oxidation of sulphides to reduce problems caused by molten sulphur, see United States Patent No. 3,867,268 (Kawulka et al) issued February 18, 1975, it has been found that the use of such additives is not commercially desirable in the pressure oxidation of refractory auriferous sulphidic material which contains arsenopyrite, pyrite or pyrrhotite, because undesirably large quantities of additives are required with consequent expense. The use of higher reaction temperatures, i.e. above about 235°C, may to some extent overcome the problem by providing more rapid oxidation of elemental sulphur, but it is doubtful whether this would be effective in a continuous operation. In any event, use of such high temperatures is undesirable because of higher equipment costs. The use of reaction temperatures below the melting point of sulphur, i.e. below about 120°C, in the pressure oxidation treatment of refractory auriferous sulphidic material has been proposed, see for example Canadian Patent No. 1,080,481 (Wyslouzil) issued July 1, 1980. However, with such treatment, the sulphur content of arsenopyrite, pyrrhotite and many of the base metal sulphides is oxidized to elemental sulphur to an undesirable extent, and much of the pyrite tends to remain unreacted. It has been proposed to digest the oxidized solids in a caustic solution to dissolve and remove the elemental sulphur. This is also undesirable, not only because an additional step is involved, but also because the caustic solution reacts with ferric arsenate and sulphur-containing iron precipitates formed during the pressure oxidation treatment and disposal or treatment of the resultant solution presents additional problems because the resultant solution will usually contain polysulphides, arsenate, sulphate and possibly a variety of unsaturated sulphur compounds. The present invention seeks to provide a process for the recovery of gold from refractory auriferous iron containing sulphidic material in which the previously mentioned problems caused by the presence of molten sulphur during the pressure oxidation step are substantially reduced. The present invention is based on the discovery that the problem of sulphide wetting by molten sulphur and the attendant problem of agglomeration can be substantially overcome at pressure oxidation treatment temperatures above about 120°C, without resorting to excessively high temperatures or excessive amounts of additives, by recycling a proportion of the oxidised solids, as a relatively inert material, to the fresh feed of refractory auriferous iron containing sulphidic material in the form of ore or concentrate to provide a relatively high slurry pulp density at least in the initial stages of the treatment where elemental sulphur formation is more likely to occur, ie. in the initial compartments of a multi-compartment horizontal autoclave, the initial reactors or kettles of a series of reactors or the initial portion of a tubular or pipeline reactor. It has been found that the recycling of oxidised solids apparently promotes dispersion of elemental sulphur which is formed, thereby reducing the tendency for agglomeration, and also promotes suspension of any agglomerates which are formed, thereby allowing them to react more completely. It is of course known to recycle solids in various metal recovery processes, usually to improve recovery totals, but not simply as an inert diluent which is effectively the case in the present invetion. For example, US-A-4,304,644 discloses a non-ferrous metal recovery operation, primarily for Ni and Co recovery, and in which, following an oxidation leach, the Ni-Co leach liquor is separated from the solids residue, and treated with copper matte to precipitate a second, copper-containing residue which is recycled in order to enhance subsequent copper recovery. This, however, is recycle of a subsequent solids residue precipitated from the leach solution, and not the recycle of the original solids residue as in the case of the present invention. The recycle of oxidised solids to the fresh feed to form a feed slurry of relatively high pulp density in accordance with the invention is to be preferred over the use of fresh feed alone to provide a high pulp density, since the resultant high sulphur content (and probably also arsenic content) may result in the production of excessive heat in the pressure oxidation treatment. The present invention is also to be preferred over the production in a preliminary flotation step of low sulphur grade concentrates for use in the pressure oxidation treatment, since in such a flotation step the sulphidic material is in effect diluted with gangue. The relatively high amounts of gangue in such low sulphur grade concentrate may cause problems in the pressure oxidation treatment, when relatively high pulp density is used. For example, the original ore may contain relatively high levels of carbonates which, if present in the pressure oxidation treatment, generate carbon dioxide which requires considerable venting with attendant losses of oxygen. Also, the acid consuming content of many refractory gold ores may be in excess of the acid available from the oxidation of sulphur thereby necessitating the addition of acid to the system. In accordance with the invention, the feed slurry pulp density including both fresh solids and oxidised recycle solids, at least in an initial stage of the pressure oxidation treatment, is maintained at a relatively high value, i.e. from about 30 to about 60% solids by weight, and preferably from about 40 to about 55%. The oxidised solids, recycled subsequent to the pressure oxidation treatment can be recycled before or after liquid-solids separation, but usually the oxidised slurry will be subjected to a liquid-solids separation step and the solids are usually washed, for example in a countercurrent decantation thickener circuit, prior to recycling a proportion thereof and processing the remainder through a gold recovery operation, e.g. a cyanidation circuit. Although oxidized slurry direct from the pressure oxidation treatment may be recycled, it will usually be preferable to rcycle oxidized solids which have been subjected to liquid-solids separation and a wash stage, since such washed solids will be cooler than oxidized slurry directly from pressure oxidation treatment. However, if the acid consuming gangue content of the fresh feed is high (for example with relatively high carbonate content), it may be preferable to recycle oxidized slurry to maximize the amount of acid recycled and hence facilitate decomposition of the carbonates. The amount of solids recycled to obtain the relatively high pulp density will primarily depend upon the sulphur content of the feed solids but will be in the range of from about 0.5:1 to 10:1 by weight, preferably from about 2.5:1 to about 4:1, relative to the fresh feed. It has been found that such recycle of oxidized material to provide a high pulp density substantially reduces agglomeration, thereby facilitating continuous operation. It has also been found that completely oxidized residue efficiently disperses elemental sulphur, preventing its selective wetting of unreacted sulphidic materials and consequently their agglomeration. Also, the recycled oxidized material will contain acid which tends to decompose carbonates in the fresh feed. The resultant carbon dioxide is thus removed prior to the pressure oxidation treatment, thereby maximizing oxygen utilization. The recycled oxidized material also contains soluble iron and/or readily soluble iron, and it has been found that such iron promotes the oxidation reaction. The recycled oxidized material has also been found effective in batch operations by accelerating the oxidation and effecting more complete liberation of gold than if fresh feed is oxidized alone. Also, the recycle of solids provides, in effect, additional retention time for imcompletely reacted sulphides. The invention is particularly useful where a plurality of mineral types are being treated. For example, a refractory gold concentrate may contain pyrrhotite, pyrite and arsenopyrite, and a zinc concentrate may contain galena, sphalerite, marmatite and pyrite. Some of these minerals are more reactive than others, and further the most reactive minerals have a propensity for producing elemental sulphur as an intermediate reaction product. Embodiments of the invention will now be described, by way of example, with reference to the accompanying drawing which shows a flow diagram of gold recovery process. Referring to the drawing, fresh ground refractory auriferous iron containing sulphidic ore or concentrate is slurried to form an aqueous slurry which is fed to a blending step 12 to which washed oxidized solids from a subsequent pressure oxidation step (to be described in more detail later) is also fed to form an aqueous feed slurry with a relatively high pulp density of from about 30 to about 60% solids by weight, preferably from about 40 to 55%. The high pulp density slurry is then subjected to a pressure oxidation step 14 in a multi-compartment horizontal autoclave at a temperature of from about 120 to about 250°C under a total pressure of from about 350 to about 6000 kPa for a retention time sufficient to effect adequate oxidation of the sulphides to sulphates. Oxidized slurry from the pressure oxidation step 14 then proceeds to a washing step 16 where water is added to the slurry. The diluted slurry then passes to a liquid-solids separation step 18 comprising a thickener where used wash water is removed as thickener overflow. A portion of the oxidized solids in the thickener is then recycled to the blending step for mixing with incoming fresh feed slurry to form the feed slurry of relatively high pulp density for subsequent pressure oxidation. The weight ratio of recycled oxidized solids to fresh feed may be in the range of from about 0.5:1 to 10:1, preferably from about 2.5:1 to about 4:1. The remaining solids are passed to a neutralization step 20 where a neutralizing agent such as lime is added to raise the pH of the slurry to a value suitable for cyanidation, for example about 10.5. The neutralized slurry then proceeds to a cyanidation step 22 where gold is recovered. Alternatively, instead of oxidized solids from the thickener 18 being recycled to the blending step 12, the recycling of oxidized solids may be effected by recycling some of the oxidized slurry leaving the autoclave in the pressure oxidation step 14, as indicated by dotted line in the drawing. The results of various tests carried out in connection with the invention will now be described. EXAMPLE 1 Tests were carried out with a concentrate containing 33.4 g/t Au, 12.4% As, 33.3% Fe and 21.4% S. It was first found that conventional cyanidation extracted 30% Au, yielding a residue containing 23.3 g/t Au. EXAMPLE 2 Such concentrate was also subjected to batch pressure oxidation treatment in accordance with the prior art at a pulp density of 10% solids, 85 kg/t H₂SO₄ and 1750 kPa total pressure. Samples were taken at predetermined time intervals and amount of sulphur oxidation to sulphate was measured as well as gold extraction in subsequent cyanidation. The results are shown in Table 1. These results show increase of gold extraction with increased sulphur oxidation. EXAMPLE 3 Batch tests were then carried out on the same concentrate under slightly different conditions with different amounts of additives. The initial charge contained 2.2% by weight plus 100 mesh solids, 373 g dry solids per charge, and the pressure oxidation was carried out for 20 minutes at a pulp density of 13% solids, with 150 kg/t H₂SO₄, a temperature of 185°C and a total pressure of 1500 kPa. The results are shown in Table II. The results show the large amounts of additives needed to reduce agglomeration. EXAMPLE 4 Tests were carried out on the pressure oxidation of the concentrate with recycle of varying amounts of oxidized solids and various pulp densities. No additives were used. The fresh concentrate contained 21.4% S and 2.2% by weight of plus 100 mesh solids. Pressure oxidation was carried out at 185°C, 1500 kPa total pressure and 20 minute retention time. The initial pH of the blended slurry was in the range of 0.8 to 0.9. The recycled solids were 100% minus 100 mesh and typically contained about 11.5% As, 28.2% Fe, 11.9% SiO₂, 6.4% S (total), less than 0.1% S (elemental) and 6.34% S (sulphate). The results are shown in Table III. The results of these tests show that with adequate dilution of the sulphur content of the fresh feed by oxidized solids and with oxidation at increased solids content in the slurry, agglomeration can be substantially reduced. EXAMPLE 5 Batch tests were then carried out on the concentrate blended with acidic underflow slurry from a first wash stage thickener generated in a continuous oxidation run. The weight ratio of recycled oxidized solids to fresh concentrate was 4:1, the feed blend slurry contained 45% solids and had an initial pH of 1.2. The oxidation was conducted at 190°C at 1780 kPa total pressure. The results of the oxidation and of subsequent cyanide amenability are shown in Table IV. The results, when compared to those of Table I, clearly demonstrate the effectiveness of the invention, in that the degree of sulphur oxidation and the extraction of gold after 120 and 180 minute oxidation are markedly higher than in the oxidation of the concentrate alone. The same concentrate as before was then used in continuous test runs. EXAMPLE 6 In the first run, the pressure oxidation was conducted at 185°C under 1510 total pressure at a pulp density of 15% solids by weight. Lignosal and quebracho were added at levels of 1 and 2 kg/t concentrate respectively. During the run, severe agglomeration of the solids was experienced in the autoclave. By 24 h, about 15% of the solids had accumulated in the first two compartments, and the run was terminated. It was found by analysis that arsenopyrite and pyrite were predominant sulphides in the agglomerates. The minus 6.7 mm to plus 0.50 mm fractions contained 90.2 to 94.5 g/t Au compared with 33.4 g/t Au in the concentrate, indicating appreciable retention and upgrading of the gold in the agglomerate. Consequently, the oxidation thickener underflow solids contained only 16.3 g/t Au, and accounted for only 40% of the gold fed into the autoclave. EXAMPLE 7 The second continuous run was conducted with increased agitation in the first two autoclave compartments and at higher addition rates of quebracho (up to 7.5 kg/t) in an attempt to disperse and suspend the agglomerates. Nevertheless, the agglomeration problem persisted during the run, which was terminated after 44 h. Autoclave inspection after the run showed that about 15% of the feed was in the first two compartments, with an additional 13% accumulated in the third compartment. Oxidation thickener underflow solids contained only 11.5 to 19.4 g/t Au. EXAMPLE 8 A third continuous run was conducted with recycle of oxidized solids, the recycle ratio of oxidized solids to fresh concentrate being 3.5:1 to produce a blended slurry with a pulp density of 50% solids by weight. The run was continued for 57 h, and no significant agglomeration problem was encountered. Oxidation thickener underflow solids contained 28.5 to 30.7 g/t Au. The advantages of the invention are therefore clearly evident. Other examples and embodiments will be readily apparent to a person skilled in the art without departing from the scope of the invention herein disclosed.

Claims: A process for the recovery of gold from refractory auriferous iron-containing sulphidic material which comprises (i) forming the iron-containing sulphidic material into an aqueous slurry by grinding the sulphidic material and suspending the ground material in an aqueous suspension medium; (ii) feeding the aqueous slurry containing the ground sulphidic material to pressure oxidation step (14); (iii) subjecting the slurry in said pressure oxidation step to a pressurised oxidation treatment carried out at a temperature in the range 120° to 250°C and at a pressure in the range 360 to 6000 kPa; and (iv) subjecting the oxidized solids, with or without separation from the aqeuous suspension medium, to a gold recovery operation (20, 22) to recover the gold from the oxidised solids; characterised in that a proportion of the oxidised solids are recovered subsequent to the oxidation step (14) and prior to the gold recovery operation (20, 22) and recycled upstream of the oxidation step (14) to provide a feed slurry to the oxidation step having a pulp density, including both unoxidised and oxidised solids, in the range 30-60% by weight, and a weight ratio of oxidised solids to fresh solids in the range 0.5:1 to 10:1. A process according to claim 1, characterised in that the recycled solids, in the slurry fed to the oxidation step, has a pulp density, including both oxidised and unoxidised solids in the range 40 to 55% solids by weight. A process according to claim 1 or 2, characterised in that the weight ratio of recycled oxidised solids to fresh unoxidised solids in the slurry fed to the oxidation step is in the range 2.5:1 to 4:1. A process according to any one of claims 1 to 3, characterised in that said oxidized solids are recycled to the feed slurry by recycling oxidized slurry directly from the pressure oxidation step. A process according to to any one of claims 1 to 3, characterised in that said recycled oxidized solids are first of all separated from the oxidized slurry in a solids-liquid separation step (18). A process according to claim 5, characterised in that the oxidized solids from the pressure oxidation step are washed prior to or during the liquid-solids separation step (18), and before recycling a proportion of the washed solids as said recycle stream. A process according to any one of claims 1 to 6, characterised in that, in the gold recovery operation, the gold is recovered from that portion of the oxidized solids that is not recycled by neutralizing those solids in a neutralization step (20) and subsequent cyanidation of those neutralized solids in a cyanidation step (22).

Inventors: WEIR, DONALD R.
Applicants: SHERRITT GORDON LIMITED
Application year: 1985
Main classification: C22B3
IPC codes: C22B11