CN1187221A - Hydrometallurgical process for extraction of copper from sulphidic concentrates - Google Patents

Abstract

There is provided a novel hydrometallurgical process for the extraction of copper from sulphidic concentrates involving an oxidizing pressure leach using dilute sulphuric acid and a carbonaceous additive. The leaching step is carried out preferably at temperatures above the melting point of sulphur but below about 200 DEG C.

Description

Hydrometallurgical process for extracting copper from sulphidic concentrates

The technical field to which the invention belongs

The present invention relates to a novel hydrometallurgical process for the extraction of copper from copper bearing sulphide concentrates, particularly refractory sulphides such as chalcopyrite.

Background of the invention

Chalcopyrite, a sulphide copper ore, is the most important copper source from an economic point of view. To date, the smelting process remains the only successful industrial process for extracting copper from chalcopyrite. Modern smelting processes extract the valuable components from the concentrated sulphide, extracting most of the sulphur as sulphuric acid, while the rest is discharged to the surroundings as a fume of sulphur dioxide. Sulphur dioxide is harmful to the environment and therefore environmentally not allowed to be emitted, and in addition, the market for sulphuric acid is extremely limited in many regions. Due to the high cost of the smelting process, economical, low-pollution alternative methods of treating chalcopyrite are continually sought.

In recent years, hydrometallurgical processes, in particular processes for extracting copper from oxide ores, have achieved industrial application. Such processes are based on sulfuric acid leaching of oxide ores, extraction of copper from leachate containing impurities with a solvent, and electrolytic deposition of metallic copper from the extract. The process not only has lower cost than most smelting processes, but also has the quality of the electro-deposited copper which can be completely compared with the quality of electrolytic refined copper produced by smelting and refining. However, despite extensive and intensive research efforts, a viable hydrometallurgical process for the treatment of chalcopyrite has not yet been developed.

Theoretically, the preferred hydrometallurgical process for economically treating chalcopyrite concentrates should include a sulphuric acid leach step. The process can directly produce the high copper sulfate solution which does not contain impurities such As Se, Fe, As, Sb and the like, has low acidity and low iron content. Preferably the process also converts sulphide to elemental sulphur. Valuable metal by-products such as silver, gold, etc. should be recoverable from the leach residue at low cost.

However, direct leaching of chalcopyrite in sulphuric acid solutions presents several problems. Generally, at temperatures below the melting point of sulfur (about 118 ℃), the rate of copper dissolution is very low and is not economically significant; at temperatures above the melting point of sulfur, the elemental sulfur deposits and seals unreacted sulfide particles, rendering them passivated and thus economically ineffective for copper extraction; at higher temperatures, i.e. above 200 ℃, leaching of copper is rapid and complete, but the sulphides are oxidised to sulphates rather than elemental sulphur. This reaction consumes an economically unacceptable amount of oxygen while producing high concentrations of sulfuric acid, and the neutralization of these sulfuric acids consumes a significant amount of capital.

Such high temperature direct oxidation leaching processes are considered to be an alternative to smelting processes in certain specific applications. For example, a hydrometallurgical process for the complete pressure oxidation of copper-containing concentrates is described in a paper entitled "high pressure leaching of copper-containing concentrates" by j.a. king et al (proceedings of the 95 copper conference, the 95Cobre international conference, volume iii, electrorefining and hydrometallurgy of copper, editions by w.c. cooper et al, society of CIM metallurgy). The process comprises high pressure leaching of the concentrate at 200-. The oxidizing solution is neutralized with limestone and then treated/electrolyzed with an extraction solvent to extract copper therefrom. Since the process consumes a large amount of oxygen and limestone, it cannot compete with the smelting process from an economic point of view.

Over the last three decades, a number of leaching systems for the extraction of copper from chalcopyrite and leachates such as ferric chloride and ammonium amino sulphate have been investigated extensively and many processes have been proposed.

However, in the amine leaching process of chalcopyrite, the sulfide sulphur concerned is oxidised to harmful sulphate and must be removed from solution in a subsequent process step in the form of ammonium sulphate or calcium sulphate to maintain the sulphate balance.

Although amino solution leaching processes have been used industrially to extract copper from secondary or enriched sulphide ores such as chalcocite and chalcopyrite, (which leave chalcopyrite and bornite in the residue), such processes have not yet been achieved industrially to fully extract copper from chalcopyrite.

In U.S. patent 4,039,406 to Stanley r.w. et al, a novel process is disclosed which overcomes the difficulties of direct pressure leaching of chalcopyrite. The process not only uses chloride medium to obtain the fast reaction kinetics of chalcopyrite, but also uses copper sulfate-copper chloride solution to rapidly extract copper from the sulfide medium at 135-145 ℃ and 200psi oxygen pressure. As with other leaching processes developed to the laboratory and commissioning stages, these processes are toxic and corrosive due to the use of chloride media and are therefore difficult to access for industrial use.

The current state of the copper industry technology is reviewed comprehensively by a book (third edition) of "copper extraction metallurgy" published by Pergamon press, which is written by a.k.biswas and w.g.davenport, and the progress of the hydrometallurgy is described in detail, and is not repeated here.

In the following, the low temperature leaching of chalcopyrite will be reviewed in detail, i.e. the reaction between 25-150 ℃ in an oxidizing atmosphere in a sulphuric acid or ferric sulphate solution as follows:

the advantage is that the sulphide is oxidised to elemental sulphur instead of sulphate, thus greatly reducing the oxygen consumption and the subsequent neutralisation process. But this advantage is offset by the low extraction rate and low extraction efficiency (50-70%). An upper temperature limit of about 150 c has been found to be beneficial in forming elemental sulfur.

It is believed that during low temperature (i.e. below 200 ℃) oxidative leaching of the sulphate media, a passivation layer forms on the surface of the chalcopyrite particles which reduces and/or completely prevents copper extraction within industrially allowable time periods. It has been postulated that the passivation layer may include elemental sulphur and an intermediate sulphur-rich phase formed by partial oxidation of chalcopyrite.

Attempts have been made to resemble the mechanism of the zinc extraction process to that of the low temperature sulphuric acid oxidation leaching reaction of chalcopyrite.

It is well known that zinc can be extracted by reacting a zinc-bearing sulphide concentrate with sulphuric acid in the presence of oxygen at elevated temperature and pressure, the elemental sulphur formed in the leaching reaction being present as finely divided sulphur droplets at a temperature above the melting point of sulphur. As the reaction proceeds, the number of sulphur droplets increases, coating or covering the unreacted sulphide particles, again preventing further oxidation reactions. Additives may be added to the process to prevent or greatly slow the covering of unleached particles by liquid sulphur, thus allowing the reaction to continue until the zinc extraction is over 95% (typical values) or even over 98% (preferred values), as exemplified in us 3867,268 and 4,004,991. Furthermore, the additive also helps to form a finely separated leach residue with good handling characteristics. Without additives, the typical zinc extraction is only 50-70% and the liquid sulfur can agglomerate, resulting in the formation of coarse, difficult to handle particles that can clog pipes and vessels.

The additives used in the process must be suitable for the oxidation of sulphides and must not introduce impurities into the zinc containing process stream. Some surfactants used for this purpose include organic compounds such as lignin derivatives; especially calcium and sodium lignosulphonate, tannin compounds, especially bark and heartwood extracts such as quebracho, hemlock, sequoia extracts; o-phenylenediamine; and alkylaryl sulfonates, particularly sodium alkylphenyl sulfonates. Calcium lignosulfonate and quebracho extract have been used industrially. The state of the art of the use of soluble surfactants in zinc pressure leaching processes is summarized in a paper recently published by Owusu et al in hydrometallurgy 38, (1995) pp.315-324 entitled "Effect of surfactants on the dissolution of zinc and iron during Zinc oxide leaching".

However, r.p. hackl et al in a paper entitled "effect of sulfur dispersing surfactants on chalcopyrite oxygen pressure leaching" clearly indicate that the sulfur dispersing surfactants used in zinc extraction processes cannot be used directly for pressure leaching of copper sulphide concentrates, especially chalcopyrite, as published in the proceedings of the international conference on Cobre95 Cu95 edited by w.c. cooper et al, and in volume iii electro-refining and hydrometallurgy of copper. The authors investigated the feasibility of using a liquid sulfur dispersing surfactant at 125-. Tests have shown that although the addition of 50kg/t o-phenylenediamine (OPD) gives a copper extraction of 80% after 6 hours, most surfactants decompose too quickly and are not effective.

It should also be noted that copper is generally believed to exhibit a catalytic effect in the decomposition of the above-described sulfur dispersing surfactant.

Thus, despite much research into the industrially feasible hydrometallurgical process for the extraction of copper from chalcopyrite, pyrometallurgical processes are still the current state of the art.

Summary of the invention

The main object of the present invention is to provide an industrially feasible hydrometallurgical process for the extraction of copper from copper sulphide concentrates, in particular iron-containing copper sulphide concentrates, more particularly chalcopyrite.

According to the process of the present invention for extracting copper from a copper sulphide concentrate, the finely divided concentrate is first dispersed in an aqueous solution of sulphuric acid to a highly dispersed state, thus forming a slurry. The sulfuric acid concentration is adjusted to bring the leach solution to the desired, experimentally determined copper, iron and sulfuric acid concentrations. An effective amount of finely divided particulate carbon is provided and mixed with the slurry as a preferred mode of addition. The carbonaceous material should be compatible with the sulfuric acid leach solution and be effective to prevent the formation of the alleged deactivation of incompletely leached sulfide particles under the reaction conditions of the oxidative leach step. The slurry and carbon material are reacted with a free oxygen-containing gas with agitation in an autoclave at a temperature preferably in the range of 135 ℃ and 175 ℃, thereby substantially completely extracting copper from the concentrate as soluble sulfate. The sulphide sulphur is simultaneously converted to elemental sulphur. The leachate containing the amount of dissolved copper is separated from the solid residue.

The process of the invention is particularly suitable for the treatment of the refractory mineral chalcopyrite, which previously has been treated industrially by smelting processes.

The advantage of using this process to extract copper from chalcopyrite is that copper extraction rates of up to 95% or more can be achieved. At the same time, leaching at a relatively low temperature, with the addition of a specific additive, allows the reaction to be controlled and the sulphide sulphur to be converted to elemental sulphur, thus reducing the costs of neutralisation and purification in the subsequent steps. It is also important that the process be industrially advantageous to use low reaction temperatures and inexpensive additives.

The invention is therefore based on the provision of a novel additive, namely a carbonaceous material, coal being the preferred material, which additive is capable of eliminating the passivation of the complete leaching of sulphide particles at reaction temperatures and under conditions favourable to the formation of elements.

Description of the drawings

The invention will be better understood with reference to the following detailed description and the accompanying drawings.

Figure 1 is a general flow diagram for the extraction of copper from a copper sulphide concentrate comprising a single pressure leaching step.

Description of the preferred embodiments

The invention is described in more detail by means of a single leaching step and with reference to the accompanying drawings. However, those skilled in the art will readily appreciate that the process of the present invention can be extended to multi-step leaching processes, as well as co-current and counter-current leaching processes.

The process is applicable to copper sulphide concentrates, the most important of which includes chalcopyrite (CuFe)₂S), chalcocite (Cu)₂S), bornite (Cu)₂FeS₄) And blue copper ore (CuS) and less abundant copper-sulfur-arsenic ore (Cu)₃AsS₄) And tetrahedrite (Cu)₁₂Sb₄S₁₃). The process is particularly suitable for treating high and low grade concentrates from chalcopyrite by froth flotation. Such concentrates often contain small amounts of valuable silver, gold and other precious metals in addition to copper, iron and sulfide. For best results, the sulfidic material is preferably produced as a finely divided, finely particulate material. The particle size has a great influence on the reaction speed, the copper extraction rate and the residence time in the pressure leaching stage. To maximize the benefits of the present invention, the particle size of the feedstock is preferably about 90% less than 10 μm, but may also range from 90% less than 44 μm to 90% less than 25 μm. The supplied material may also require ball mill wet milling or other suitable equipment for size reduction measures to achieve the desired particle size.

Optionally, the ground concentrate is leached in dilute sulfuric acid under low temperature air pressure conditions to easily remove soluble impurities such as magnesium, manganese, chlorides and fluorides, followed by liquid/solid separation.

The process of the present invention includes providing an additive suitable for aqueous sulfuric acid solutions and which is effective in inhibiting the deactivation of theotherwise recognized incompletely leached sulfide particles under the reaction conditions of the process. The additive may comprise any suitable carbon material, but the most preferred material is a coal. The preferred additive is a low carbon content coal which does not introduce impurities into the leach solution which could interfere with subsequent processing. Such coals may include candelilla, peat, lignite, subbituminous, bituminous, semi-anthracite, and anthracite. Coal with low or medium carbon content is preferred over high grade or high carbon content coal. The total carbon content of the coal is preferably between 40 and 85%. Generally, it has been found that high carbon content coals (e.g., coals characterized by aromatic carbons such as anthracite) are less effective than low carbon content coals having an aliphatic carbon content of about 20% to about 80%, preferably about 25% to about 55% aliphatic carbon content. The coal must be finely divided to have a particle size of less than about 60 μm. Alternatively, the coal may be mixed with the concentrate for milling. If coal fines are provided, then grinding is not required and is applied directly. Conventionally, in any given situation, a minimum amount of coal is added that is effective to increase the copper extraction. In many cases the amount of coal added is about 3 to 50kg per ton of copper-containing concentrate, typically 10kg is sufficient. It is obvious to one skilled in the art that the amount of coal added can also be modified depending on the nature of the coal, i.e. grade and particle size, and the carbon content of the copper-containing concentrate. Under the conditions described in the present invention, it was found that approximately less than 50% of the coal chemically disassociated.

According to the background, the ASTM standard grades from lignite to anthracite according to the natural series with fixed carbon and heating value calculated on a mineralfree basis as classification criteria according to the degree of metamorphism or gradual change. The european classification includes: (a) an international classification for hard coal types and (b) an international classification for brown coal, as described in the seventh edition "Van Nostrand scientific encyclopedia" published by Considine, first volume pages 662 to 663. Generally, coal can be defined as a combustible, carbonaceous deposit mineral formed from the compression of plant parts.

The leaching reaction is carried out in a pressurised air vessel (e.g. an autoclave) equipped with agitation means at a temperature above the melting point of sulphur, i.e. above 120 c, but below 220 c, preferably in the range 135 c to 175 c, with about 150 c being the most preferred temperature. However, the reaction can still be carried out as low as 90 ℃.

The total pressure at which the leaching reaction is carried out is equal to the autogenous vapor pressure at the oxidation reaction temperature plus the residual pressure of the oxidizing gas. The oxidizing gas is preferably oxygen, but air and oxygen-enriched air may be used. The reaction proceeds well at oxygen residual pressures above about 100 KPa. However, as the residual oxygen pressure increases, the reaction rate increases. Therefore, the residual oxygen pressure is preferably about 400 to 750 KPa. The upper limit of the oxygen pressure is the highest pressure that can be applied by the autoclave used. Since it is generally economical to avoid the use of high pressure evaporators, the upper limit of the residual oxygen or residual air pressure is typically about 3000 Kpa.

In any given instance, the density of the leach slurry supplied to the leach (i.e. the relative amounts of sulphide and liquor supplied to the leaching process) is generally determined by the copper content of the sulphide and the desired copper, iron and acid concentrations in the final leach solution, etc.

The amount of sulphuric acid supplied in the slurry synthesis process, which is determined by the required copper, iron and acid concentrations in the final leach solution and the amount of sulphuric acid added, is determined experimentally, first by determining the yield of elemental sulphur and the amount of refractory sulphides (i.e. the most refractory, most typically pyrite). The remaining sulfur of the feed solids is oxidized to sulfate, i.e., copper sulfate, iron sulfate, and other ferrous sulfate salts. Thus, the stoichiometric amount required to effect the conversion reaction is added together with a sufficient amount of acid remaining in the final leach solution. Most of the iron is precipitated as sulfates of basic iron and iron hydroxide.

Water, ferrous sulfate, ferric sulfate, or mixtures thereof may be used as an in situ generated sulfuric acid source in place of sulfuric acid. Again, the required amounts will be determined analytically to obtain the desired concentrations of copper, iron and free acid in the final leach solution.

After the leaching step, the product leach solution containing dissolved copper is separated from the solid residue. The copper in the product liquor can be recovered by electrowinning and then acid neutralisation and removal of iron or optionally purification of the solution by solvent extraction.

The following non-limiting examples provide several embodiments for the extraction of copper from chalcopyrite.

Example 1

This example illustrates the effect of additives on the extraction of copper from a single concentrate. An industrially supplied copper sulphide concentrate containing 28% Cu, 29% Fe and 32% S, in which the copper is present substantially entirely in the form of chalcopyrite, in which no iron is found, and in which the iron is present predominantly in the form of pyrite, is subjected to a 4 litre laboratory autoclave at 150 ℃ and 750KPa oxygen partial pressure in a laboratory autoclave containing 30g/L Cu and 120-150 g/LH₂SO₄Is leached in the acidic copper sulfate solution for 6 hours. In the first test, no sulphur dispersant was added, in the second test, lignosulphonic acid was added to the autoclave feed in an amount of 2kg per ton of concentrate, and in the third test, lignosulphonic acid was added per ton of concentrate25kg of sub-bituminous coal having a carbon content of 5% and a particle size of 100% less than 63 μm was added to the autoclave feed and this coal was designated as coal A. The results of the three tests are summarized in table 1 below. The extraction rate of copper without additive was 49%, that with calcium lignosulfonate 71%, and that with coal 98%. In the coal addition test, 69.4% of the sulfide sulfur in the feed was converted to elemental sulfur, 27.4% was sulfate-forming and 3.2% was unreacted.

TABLE I

Test of	1	2	3
				Copper-containing concentrate	A
Analysis of the concentrate composition% Cu Fe Ni Si S Zn	27.8 28.8 ＜0.1 2.41 32.5 ＜0.1
				Concentrate particle size D90，μm	13
Acid: cu + Fe molar ratio	1.66	1.66	0.67
				Temperature, C Oxygen pressure, kPa	150 750
Additive material Addition rate, kg/t	Is free of 0	Wood pulp	Coal A 25
				Copper extraction rate% 30 minutes 60 minutes 120 minutes is 240 points of 360 minutes	46.4 46.8 40.0 42.8 49.3	55.8 62.5 67.2 71.8 70.8	50.4 64.7 83.3 96.7 98.4
Final extraction rate% Cu Fe S	49.3 59.3	70.8 70.6	98.4 26.8 16.3
				Final unreacted liquid, g/L	46.0 20.3 94.0	50.8 23.8 83.5	79.0 19.1 23.6

Final sulfur conversion,% To elemental sulfur To sulfate sulfur Unreacted

69.4 27.4 3.2

Example 2

The effect of concentrate particle size on copper extraction is shown in table 2 below. When 90% of the feed was able to pass through a 7 μm sieve, the copper extraction was greater than 95% after 2 hours of leaching in test 3.

TABLE II

Test of	1	2	3
				Copper-containing concentrate	B
Analysis of the concentrate composition% Cu Fe Ni Si S Zn	23.4 29.0 0.2 1.09 34.1 5.9
				Concentrate particle size D90，μm	28	16	7
Acid: cu + Fe molar ratio	0.70	0.69	0.69
				Temperature, C Oxygen pressure, kPa	150 750
Coal kind Coal addition rate, kg/t	A 25
				Copper extraction rate% 30Is divided into 60 minutes 120 minutes is 240 points of 360 minutes	20.0 36.1 63.3 82.9 90.6	44.6 64.9 83.1 94.7 97.9	62.6 83.4 95.2 98.2
Final extract of rice, milling Cu Fe S	90.6 8.6 -4.5	97.9 12.4 0.6	98.2 38.3 15.4

Final solution, g/L Cu Fe H2SO4	75.6 13.1 11.7	77.9 13.2 14.6	76.2 27.1 20.4
				Final sulfur conversion,% To elemental sulfur To sulfate sulfur Unreacted		70.3 27.2 2.5	65.3 33.1 1.6

Example 3

The effect of the addition of coal a on copper extraction is shown in table 3 below. When the amount of coal added was 5kg/t concentrate or more, the copper extraction rate was 98% or more.

TABLE III

Test of	1	2	3	4
					Copper-containing concentrate	B
Analysis of the concentrate composition% Cu Fe Ni Si S Zn	23.4 29.0 0.2 1.09 34.1 5.9
					Concentrate particle size D90，μm	7
Acid: cu + Fe molar ratio	0.69		0.86
					Temperature, C Oxygen pressure, kPa	150 750
Coal kind Coal addition rate%	A
					2	5	10	30
	Copper extraction rate% 30 minutes 60 minutes 120 minutes 240 minutes 360 minutes	78.3 87.5 89.0	50.4 75.8 93.2 98.4	62.5 79.9 94.3 98.7					68.9 84.5 94.3 97.5

Final extraction rate% Cu Fe S	89.0 46.7 31.7	98.4 29.8 8.8	98.7 68.3 25.0	97.5 59.8 25.0
					Final solution, g/L Cu Fe H2SO4	67.0 27.1 20.9	73.9 23.9 19.7	69.1 33.4 20.5	63.6 27.8 21.4
Final sulfur conversion,% To elemental sulfur To sulfate sulfur Unreacted	49.1 43.5 7.4	69.6 28.5 1.9	68.1 31.9	63.1 35.9 1.0

Example 4

This example provides data on the effect of type of material in the coal on copper extraction. The coal additives are divided into several examples of sub-bituminous coal and bituminous coal, and the carbon content is as follows:

total carbon content of coal type

A 59％

B 68％

C 70％

D 56％

E 84％

F84% table 4 provides the test results. When the additive contains one third or more of aliphatic carbon, the extraction rate of copper is more than 96%.

TABLE IV

Test of	1	2	3	4	5	6
							Copper-containing concentrate	B
Analysis of the concentrate composition% Cu Fe Ni Si S Zn	23.4 29.0 0.2 1.09 34.1 5.9

Concentrate particle size D90，μm	7	16	7	7	16	16
							Acid: cu + Fe molar ratio	0.69	0.69	0.86	0.86	0.69	0.69
Temperature, C Oxygen pressure, kPa	150 750
							Coal kind	A	B	C	D	E	F
Carbon distribution of coal Aliphatic carbon,% of Aromatic carbon,% of	50.4 49.6	44.8 55.2	43.2 56.8	37.2 62.8	23.0 77.0	16.6 83.4
							Coal addition rate, kg/t	25
Copper extraction rate% 30 minutes 60 minutes 120 minutes is 240 points of 360 minutes	62.6 83.4 95.2 98.2	41.9 57.1 79.3 93.4 96.6	67.3 84.4 96.1 98.5	68.4 93.9 94.9 98.9	30.1 40.5 45.4 51.4 54.1	28.7 36.3 52.7 54.5 57.2
							Final extraction rate% Cu Fe S	98.2 38.3 15.4	96.6 4.7 -3.8	98.5 76.8 26.7	98.9 79.4	54.1 49.0	57.2 55.5
Final solution, g/L Cu Fe H2SO4	76.2 27.1 20.4	76.0 13.2 15.3	67.7 37.2 22.2	67.9 36.8 19.0	57.2 29.4 33.1	58.7 3.3 31.2

Example 5

The extraction of copper and other metals from various types of concentrates with coal as an additive was investigated and the results are listed in table 5. The copper extraction rate can reach 98% in each case, and more than 60% of the sulphide sulphur in the feed is converted to elemental sulphur.

It is evident from the above examples that the process is not only effective for dispersed copper-containing concentrates, but can also be generalized to bulk concentrates of chalcopyrite mixed with other minerals.

TABLE V

Test of	1*	2**	3	4	5
						Copper-containing concentrate	A	B	C	D	E
Analysis of the concentrate composition% Cu Fe Pb Ni Si S Zn	27.8 28.8 0.01 2.41 32.5 0.02	23.4 29.0 0.46 0.15 1.09 34.1 59.0	24.6 28.0 0.08 0.01 4.42 30.8 0.07	13.8 31.6 0.05 2.48 5.23 26.5 0.07	8.67 33.2 3.64 0.10 1.47 31.4 9.79
						Mineralogical calculation of% CuFeS2 FeS2 PbS NiS ZnS	80.3 7.2 1.6	67.6 9.7 6.2 0.5 0.2 8.8	71.1 8.4 3.9 0.1 0.1	39.9 0.0 30.7 0.1 3.8 0.1	25.0 9.0 32.3 4.2 0.2 14.6
Particle size D% of concentrate, mum	13	7	16	25	9
						Acid: cu + Fe molar ratio	0.67	0.69	0.69	0.78	0.84
Temperature, C Oxygen pressure, kPa	150 750
						Coal kind Coal addition rate, kg/t	A
25	25	50	25	25
					Copper extraction rate% 30 minutes 60 minutes 120 minutes is 240 points of 360 minutes		50.4 64.7 83.3 96.7 98.4	62.6 83.4 95.2 98.2	64.1 77.9 90.7 96.5 98.1	72.1 87.1 95.0 98.3	61.8 89.5 97.0 97.9
Final extraction rate% Cu Fe Ni S Zn	98.4 26.8 16.3	98.2 38.3 15.4 99.9	98.1 29.1 18.7	98.3 22.6 98.8 16.5		97.9 6.5 -11.5 98.8

Final solution, g/L Cu Fe Ni H2SO4 Zn	79.0 19.1 23.6	76.2 27.1 20.4 11.6	77.7 15.9 25.8	27.4 16.1 5.0 29.5 0.2	17.4 5.3 21.1 19.7
						Final sulfur conversion,% To elemental sulfur To sulfate sulfur Unreacted	69.4 27.4 3.2	65.3 33.1 1.6	69.5 30.0 0.5	64.2 15.0 20.8	80.4 15.8 3.8

^*TABLE I, run 3^**TABLE II run 3

It will of course be appreciated that modifications may be made to the embodiments of the invention described above without departing from the scope and ambit of the invention as defined by the claims.

Claims (17)

1. A hydrometallurgical process for the extraction of copper from a copper sulphide concentrate, the process including:

dispersing a finely divided copper sulphide concentrate in an aqueous sulphuric acid solution to form a slurry, the sulphuric acid concentration being adjusted to be effective to provide a predetermined concentration of copper, iron and acid in the final leach solution;

providing an effective amount of a finely divided particulate carbonaceous material which is compatible with the acidic sulfate leach solution and which is effective to inhibit deactivation of incompletely leached sulfide particles under the reaction conditions of a subsequent oxidation reaction step;

reacting said slurry and carbonaceous material in a pressure vessel at a temperature and with agitation with a free oxygen-containing gas effective to substantially completely extract copper from said sulfide in the form of soluble copper sulfate and simultaneously convert a substantial portion of the sulfide sulfur corresponding to the amount of copper to elemental sulfur; and

the product leach solution containing the amount of dissolved copper is separated from the solid residue.

2. A hydrometallurgical process for the extraction of copper from a chalcopyrite-containing sulphide feed material, the process including:

dispersing finely divided feed materials in an aqueous sulfuric acid solution to form a slurry, said sulfuric acid concentration being adjusted to be effective to provide predetermined copper, iron and acid concentrations in the final leach solution;

providing an effective amount of finely divided particulate carbonaceous material which is compatible with the acidic sulfate leach solution and which is effective to inhibit deactivation of incompletely leached sulfide particles under the reaction conditions of a subsequent oxidation reaction step;

reacting said slurry and carbonaceous material in a pressure vessel at a temperature and with agitation with a free oxygen-containing gas effective to substantially completely extract copper from said sulfide in the form of soluble copper sulfate and simultaneously convert a substantial portion of the sulfide sulfur corresponding to the amount of copper to elemental sulfur; and

the product leach solution containing the amount of dissolved copper is separated from the solid residue.

3. A hydrometallurgical process for the extraction of copper from a chalcopyrite-containing sulphide feed material, the process including:

dispersing finely divided feed materials in an aqueous solution, an aqueous ferrous metal solution, an aqueous ferric sulphate solution or a mixture thereof to form a slurry and thereby produce sulphuric acid in situ in a subsequent oxidation reaction process and adjusting the sulphuric acid concentration to effectively provide predetermined copper, iron and acid concentrations in the final leach solution;

providing an effective amount of finely divided particulate carbonaceous material which is compatible with the acidic sulfate leach solution and which is effective to inhibit deactivation of incompletely leached sulfide particles under the reaction conditions of a subsequent oxidation reaction step;

reacting said slurry and carbonaceous material in a pressure vessel at a temperature and with agitation with a free oxygen-containing gas effective to substantially completely extract copper from said sulfides as soluble copper sulfate and simultaneously convert a substantial portion of the sulfide sulfur associated with the copper content to elemental sulfur; and

the product leach solution containing the amount of dissolved copper is separated from the solid residue.

4. The process of claim 1, 2 or 3 wherein the residual oxygen pressure of the oxygen is in the range of about 100kpa to about 3000 kpa.

5. The process of claim 1, 2 or 3 wherein the oxygen residual oxygen pressure is in the range of about 400kpa to about 750 kpa.

6. The process according to claim 4 or 5, wherein the finely divided particulate carbonaceous material is one selected from the group consisting of candelilla, peat, lignite, subbituminous coal, bituminous coal, semi-anthracite coal and anthracite coal.

7. The process of claim 4, 5 or 6 wherein the reaction temperature is in the range of about 90 ℃ to about 220 ℃.

8. The process as claimed in claim 4, 5 or 6, wherein the reaction temperature is in the range of about 120-180 ℃.

9. The process as claimed in claim 4, 5 or 6, wherein the reaction temperature is in the range of about 135-175 ℃.

10. A process according to claim 4, 5, 6, 7, 8 or 9, wherein the coal is added in an amount ranging from 3 to 50kg per ton of copper sulphide concentrate.

11. The process according to claim 9, wherein the amount of coal added to the slurry is 10kg per ton of copper sulphide concentrate.

12. The process of claim 10 wherein the coal is a low or medium carbon grade coal having a total carbon content in the range of about 40 to 85%.

13. The process of claim 10 or 12 wherein the coal has an aliphatic carbon content of less than about 20 to 80%.

14. The process of claim 10 or 12 wherein the coal has an aliphatic carbon content of about 25 to about 55%.

15. The process according to claim 1, 2, 3 or 13, wherein the particle size of the concentrate is about 90% less than 44 μm or less.

16. The process of claim 1, 2, 3 or 13 wherein said concentrate has a particle size of about 90% less than 25 μm or less.

17. The process according to claim 1, 2, 3 or 14, wherein said concentrate has a particle size of about 90% less than 10 μm.

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