US20190017146A1

US20190017146A1 - Single stage pressure leach hydrometallurgical method for upgrade of copper concentrates

Info

Publication number: US20190017146A1
Application number: US16/032,726
Authority: US
Inventors: Grenvil Marquis Dunn
Original assignee: ORWAY MINERAL CONSULTANTS (WA) Pty Ltd
Current assignee: ORWAY MINERAL CONSULTANTS (WA) Pty Ltd
Priority date: 2017-07-12
Filing date: 2018-07-11
Publication date: 2019-01-17

Abstract

A hydrometallurgical method 80 for the removal at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) from a copper concentrate to produce an upgraded copper concentrate 92, the method comprising the step of: subjecting the copper concentrate to an acidic leaching process (NONOX leach 120) in a multi-compartment autoclave using a copper sulfate containing lixiviant under lowered electrochemical conditions, to implement a controlled removal of between 20% and 99% of at least one of the impurities to produce the upgraded copper concentrate. The NONOX leach 120 is conducted to produce the majority of required soluble copper sulfate from a partial oxidation of the feed copper concentrate in a first compartment 200 of the multi-compartment autoclave under particular conditions.

Description

FIELD OF THE INVENTION

The present invention relates to a single stage method for the upgrade of copper concentrates by the removal of impurities. The method relates particularly although not exclusively to the removal of at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) from a copper flotation concentrate. The method relates typically but not exclusively to the treatment of primary copper sulphide flotation concentrates without pre-treatment or post-treatment. The method also relates to the efficient use of soluble copper in the non-oxidative leaching step for upgrade of copper concentrates and so minimize copper losses.

BACKGROUND TO THE INVENTION

The dominant copper-containing minerals in most copper sulfide deposits are chalcopyrite, cubanite and bornite. Chalcocite, covelite and in some cases enargite or tennantite are also present. The gangue mineral sulfides sometimes have pyrite and pyrrhotite present, many of these along with lesser quantities of host or gangue minerals report to the final flotation concentrate.
High-grade, copper sulfide concentrates, (typically greater than about 25% Cu weight/weight), are commonly treated by pyrometallurgical routes, whereas hydrometallurgical routes are generally favoured for the lower-grade, or impurity bearing copper concentrates. The economically and technically most favourable processing route can also be influenced by the concentration of minor amounts of valuable metals such as cobalt and nickel, or payable precious metals such as silver, gold, palladium and platinum, as well as contamination by radioactive elements such as uranium, thorium, radium, lead, bismuth or polonium and deleterious metals such as arsenic, present in the feed material. Hydrometallurgical processing routes are generally more energy consuming than smelting, because the heat of combustion of the concentrates is not efficiently utilized.
The three dominant pyrometallurgical routes for high-grade, copper sulfide concentrates are;

- a) smelting to a matte followed by converting to blister copper,
- b) direct to blister smelting and
- c) oxidative roasting.

The efficiency of the smelting technology is determined by, amongst other things, the Cu/S ratio and the concentration of slag forming components, especially iron, magnesium and silica. Conventional smelting processes are generally not applicable to lower grade copper concentrates. Not all of the copper content of the original feed is recovered as blister copper, with the remaining copper reporting to the slag and to the smelter dusts or fumes recovered from the smelter off-gases.
Roasting of copper concentrates requires the conversion of the copper content to a water-soluble or sulfate form, which is recovered from the roaster calcine by leaching, followed by solvent extraction and electrowinning. Roasting is often inefficient because copper-containing insoluble ferrite phases can form during the roasting stage and lock some copper and valuable by-products such as cobalt.
Many hydrometallurgical processes have been described for treating copper-containing concentrates, for example;
Burkin A. R, Chemical Hydrometallurgy, 1952-1994, Trans. Inst Min. Metall., 103, 1994, C169-C176.
Dreisinger, D, Copper leaching from primary sulfides: Options for biological and chemical extraction of copper, HYDROMETALLURGY, 2006, 83, 10-20.
Few of the proposed processes have attained full-scale commercial development, and most give little or no attention to removal of impurity or penalty elements, including radionuclides, or disposal of these elements by environmentally benign methods. Hydrometallurgical processes for copper concentrates struggle to compete economically against pyrometallurgical steps such as smelting, for reasons including:

- a) effective removal of impurity or penalty elements,
- b) cost of power,
- c) environmentally acceptable disposal of residues, and
- d) difficulty in precious metal recovery.

Economic performance of the smelting routes in particular is improved if copper concentrates can be upgraded in their copper content or deleterious impurities can be removed before being fed to the smelting furnaces. But the source or production of copper sulfate solution for ‘metathesis’ reactions, and deportment of the impurity, radioactive or value elements, is not considered in most treatment routes, and nor is the disposal of residues and effluents.
Various means of hydrometallurgical upgrading of the copper content of a copper concentrate have been proposed, including ‘metathesis’ leaching which displaces the iron content of the concentrate with copper. The so-called ‘metathesis’ process, in which the chalcopyrite component of the concentrate is reacted with a copper sulfate solution to produce low-iron copper sulfide (e.g. digenite) and an acidic, ferrous sulfate solution, can be represented.
3CuFeS₂+6CuSO₄+4H₂O→5Cu_1.8S+3FeSO₄+4H₂SO₄ (1)
A similar reaction also occurs for any bornite present in the copper concentrate.
3Cu₅FeS₄+6CuSO₄+4H₂O→5Cu_1.8S+6Cu₂S+3FeSO₄+4H₂SO₄ (2)
Similar reactions occur when cobalt is recovered from a blend of cobaltite and chalcopyrite minerals or carrollite.
The descriptions of processes which involve metathesis reactions do not generally include the source of copper sulfate, the deportment of impurity, valuable and radioactive elements, or the treatment and disposal of residues and effluents.
One or both of these above reactions (1) and (2) are referred to directly or indirectly in U.S. Pat. Nos. 2,568,963, 2,662,009, 2,744,172 and 4,024,218, Canadian Patent No. 1,258,181, South African Patent No. 2007/01337, and WIPO Patent Publication No. WO 2004/106561. All of these patents propose to forward the upgraded copper sulfide concentrate, which typically contains above 50% Cu, to either a smelter or treatment by other means.
The flowsheets in these patent specifications contain several deficiencies, such as identifying an economic source of copper sulfate solution, incomplete separation of iron and copper in solution, the requirement of additional flotation steps, economic recovery of precious metals from the residues, or have difficulty removing other impurities such as radionuclides, including uranium and its decay elements, and the final destination or treatment route of residues and effluents which can be problematic.
More recently PCT/AU2014/000268 and PCT/AU2014/000269 aim to at least partially overcome some of these deficiencies. This prior art addresses more importantly the removal of uranium, along with the other radionuclides which are its decay elements, that would otherwise limit or penalise the processing of the concentrate in an off-shore or remote smelter, or prohibit or restrict the international trade of copper concentrate across international borders. Thus these inventions have successfully lowered the level of uranium and other radionuclides in a radioactive or ‘dirty’ copper concentrate to allow the concentrate to be smelted within the limits of national or international regulations. These recently published flowsheets have included pre- and/or post-treatment of the concentrates, which necessarily complicates the number, and type of process steps, as well as the control of the chemistry in each of the process steps. Such complexity adds to the operating and capital costs of implementing the flowsheet for the treatment of copper concentrates compared to the methods described by the present invention.
The present invention was developed with a view to providing a simplified hydrometallurgical removal step of impurities from copper concentrates and so enhance the economics of the treatment process. Overall capital and operating cost components of the total processing of concentrates are minimised, as well as ensuring residues may be disposed by means acceptable to regulatory authorities.
Some copper concentrators can be constrained to ensure that an acceptable smeltable grade concentrate is produced. Such constraints can inter alia result in copper metal losses to the tails. Removing these constraints in the concentration step and maximising copper yield can result in a low grade concentrate that is either not attractive to smelters or may have an associated treatment penalty. The method of the invention is also intended for the upgrade of an otherwise low grade copper concentrate to render it smeltable without penalty for the recovery of its value metals notably copper and any precious metals. Additionally some concentrators are remote in their geographic location and an upgrade step will assist in reducing not only transport cost but also the treatment charges levied by a smelter. Additionally some copper operations need to suppress pyrite in flotation to meet saleable copper concentrate grade which can result in an acid generating tail. With this process the restriction of grade may be reduced or removed allowing more sulphides to report the concentrate which may move the designation of the tailings from being one that is acid generating to non-acid generating.
References to prior art in this specification are provided for illustrative purposes only and are not to be taken as an admission that such prior art is part of the common general knowledge in Australia or elsewhere.

SUMMARY OF THE INVENTION

According to one aspect of the invention there is provided a hydrometallurgical method for the removal at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) from a copper concentrate to produce an upgraded copper concentrate, the method comprising the step of:
subjecting the copper concentrate to an acidic leaching process (NONOX leach) in a multi-compartment autoclave using a copper sulfate containing lixiviant under lowered electrochemical conditions, to implement a controlled removal of between 20 and 99% of at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) to produce the upgraded copper concentrate, with reduced soluble copper losses in the discharge stream, and wherein the NONOX leach is conducted:

- to produce the majority of required soluble copper sulfate from a partial oxidation of the feed copper concentrate in a first compartment of the multi-compartment autoclave;
- to pass the partially oxidised slurry from the first compartment into the second and ensuing compartments of the multi-compartment autoclave;
- such that a division between the first compartment and the second and ensuing compartments does not permit transfer of unused oxidant gas from the first compartment to the second and ensuing compartments;
- to provide a non-oxidising atmosphere in the second and ensuing compartments such that a majority of the soluble copper in the partially oxidised slurry from the first compartment is uploaded onto the partially leached solids to provide the upgraded concentrate;
- employing elevated temperature and elevated pressure to suppress boiling in the leaching process; and
- utilising acid generated from the hydrolysis of ferric iron in the first compartment to re-dissolve the ferric iron in the non-oxidising conditions of the second and ensuing compartments.

Preferably the leaching process (NONOX leach) is conducted at an electrochemical potential of greater than 150 mV (Ag/AgCl 3.8M KCl). More preferably the leaching process (NONOX leach) is conducted at an electrochemical potential in the range of between about 175 mV and 450 mV (Ag/AgCl 3.8M KCl).
Typically the electrochemical potential of the NONOX leach is controlled by the presence of, or the addition to the NONOX leach of, one or more of cupric sulfate, ferric ion, air, oxygen, pyrolusite, or hematite.
The copper sulfate containing lixiviant typically comprises one or more of the following solutes: sulfuric acid, sodium sulfate, copper sulfate, the sulfate containing lixiviant being obtained from within the NONOX leach, or from a source external to the NONOX leach. Additionally the lixiviant may contain some chloride anions.
Preferably the NONOX leach is conducted between about 100° C. and 240° C. More preferably the NONOX leach is conducted between about 160° C. and 240° C.
In the hydrometallurgical method of the invention, copper-iron-sulfides are typically transformed in the NONOX leach, and more usually the NONOX leach is conducted under conditions to transform 50% to 99% of the copper-iron-sulfides to iron-depleted copper sulfides and covellite and to a sulfur-depleted variant (digenite/chalcocite). Similarly pyrite and pyrrhotite can be substantially altered under NONOX conditions to iron depleted copper-iron-sulfides.
Preferably the NONOX leach is conducted at a pressure in the range of 500-3500 kPa.
Typically the retention time of the NONOX leach is between about 0.5 and 8 hours.
The partial oxidation of the feed copper concentrate in the first compartment of the multi-compartment autoclave preferably employs oxygen, oxygen enriched air or air as the oxidant.
The first compartment of the multi-compartment autoclave is preferably divided from the other compartments by a divider construction that restricts the free flow of the autoclave vapour phase from the first compartment to other compartments, i.e. the divider construction forms the division between the first compartment and the second and ensuing compartments so that the transfer of unused oxidant gas from the first compartment to the second and ensuing compartments is minimised or not permitted.
Furthermore a counter current vapour flow within the autoclave is usually present such that the second and remaining compartments have a vapour phase that is essentially free of oxygen.
The method may further comprise feeding into the NONOX leach a copper sulfate slurry from a crystalliser or a copper sulfate leach liquor from an external pressure oxidation reactor (PROX reactor), or copper sulfate solution from an external source, or a copper sulfate leach slurry direct from a PROX reactor. The copper sulfate slurry typically comprises copper sulfate, iron sulfate and sulfuric acid. Additionally the copper sulfate can be added to more than one compartment in the NONOX leach reactor. Typically the copper aqueous concentration levels within the discharge liquor from the NONOX leach reactor are less than 5 g/L and more typically less than 1 g/L.
Throughout the specification, unless the context requires otherwise, the word “comprise” or variations such as “comprises” or “comprising”, will be understood to imply the inclusion of a stated integer or group of integers but not the exclusion of any other integer or group of integers. Likewise the word “preferably” or variations such as “preferred”, will be understood to imply that a stated integer or group of integers is desirable but not essential to the working of the invention.

BRIEF DESCRIPTION OF THE DRAWINGS

The nature of the invention will be better understood from the following detailed description of preferred embodiments of the invention, given by way of example only, with reference to the accompanying drawings, in which:

FIG. 1 is a flow diagram of a hydrometallurgical method for the upgrade of copper concentrates; and

FIG. 2 is a diagram showing the components of the NONOX autoclave.

DETAILED DESCRIPTION OF PREFERRED EMBODIMENTS

FIG. 1 is a single stage pressure leach process flowsheet for a hydrometallurgical method [80] for the removal of at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) from a copper concentrate [92].
Copper concentrate [92] is produced by known methods of treating sulfide ore [91] such as milling and flotation [90]. Tailings [93] which contain the gangue minerals may be disposed directly to impoundment [169] or utilized for neutralization of effluent streams, or treatment of discharge liquors [180].
The copper flotation concentrate [92] containing primary and secondary copper sulfides, iron sulfides, and at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) is repulped [100] with process liquors that may contain sodium, copper, iron, sulfates [94]. Typically the copper sulfide minerals include one or more of: chalcopyrite, covelite, bornite, chalcocite, cubanite, enargite, tennantite, tetrahedrite. Typically the iron sulfides may contain pyrite, arsenopyrite, pyrrhotite and the like. Typically the primary uranium minerals may contain brannerite, betafite, davidite, coffinite and uraninite. Generally the arsenic and base metals are associated with the sulfides, but may be distributed with other minerals depending on the prior treatment of the ore and concentrate.
The repulped concentrate slurry [101] is advanced to a first compartment [200] of a multi-compartment NONOX autoclave or leach reactor [120], as shown in FIG. 2.
Air, oxygen or oxygen enriched air [114] is fed to the first compartment [200] of the multi-compartment NONOX autoclave [120] where an oxidative process releases copper and iron into the aqueous fraction of the slurry. Some of the solubilised iron may precipitate as iron oxide or iron-hydroxy-sulfate.
Acid generated from the hydrolysis of ferric iron in the first compartment is utilised to re-dissolve the ferric iron in the non-oxidising conditions of the second and ensuing compartments.
The oxidation process raises the temperature in the first compartment [200] to the preferred temperature at which copper and iron are solubilised.
One or more of the impurity metals dissolve with the copper and iron.
Some steam [121] may be added to the first compartment [200] to support the required temperature of the oxidative process.
Preferably the oxidative process is conducted at a temperature of between 120 and 240° C. to dissolve sufficient copper before it passes into the second [200] and further compartments [204] and [206] of the multi-compartment autoclave [120].
The partially leached slurry passes from the first compartment [200] to the second compartment [202] of the same vessel [120] through a divider construction or dividing wall [210] that restricts movement of any vapour phase oxidant with the partially leached slurry as can be seen in FIG. 2.
The conditions in the second [202] and ensuing compartments of the NONOX autoclave [120] are selected to encourage the metathetic upload of the solubilised copper from the first compartment.
Preferably the conditions in the second [202] and ensuing compartments of a multi-compartment autoclave head space support a counter-current vapour flow with respect to the slurry flow, within the autoclave [120] such that the oxygen containing atmosphere is confined within the first compartment [200].
The upgraded copper concentrate is typically a concentrate which is upgraded in terms of copper metal and which contains reduced levels of at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities).
Optionally stream [115] can be a copper sulfate crystal slurry. Optionally the copper sulfate addition [123] to the NONOX reactor [120] can be in two or more locations within the autoclave. Preferably the copper concentration in the NONOX reactor aqueous phase is maintained in excess of 0.2 g/L.
Steam [121] may be added to the autoclave to maintain the reactor at the target temperature.
Temperatures below 250° C. and typically below 220° C. are employed in the NONOX leach. Typically the NONOX leach is conducted between about 120° C. and 250° C., and more typically between about 200° C. and 220° C.
Nominally anaerobic leach conditions are maintained in the second and ensuing compartments of the NONOX reactor. Small quantities of oxygen, air or an appropriate oxidant may be required to influence the overall chemistry within the reactor or to control the electro-chemical potential of the reactor above typically 150 mV (Ag/AgCl; 3.8M KCl), and more typically in the range of between about 175 mV and 450 mV (Ag/AgCl 3.8 M KCl). The electrochemical potential is chosen with a lower limit of just above that at which elemental copper could form and precipitate, and be lost from the aqueous phase.
Conditions are employed in the autoclave [120] such that between 20% and 99% of at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) are removed from the feed concentrate and report to the exit stream [122] and a near stoichiometric amount of copper is precipitated in the concentrate.
Soluble copper is lowered to concentrations between 0.2-5.0 g/L in the exit stream [122] from the NONOX leach reactor.
The other components of the aqueous fraction of NONOX discharge [122] will vary with the mineral assemblage in the feed concentrate slurry [101]. The solute in the aqueous phase [122] will include at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) and sulfuric acid.
Gold and silver report quantitatively to the upgraded copper concentrate exiting in stream [122].
An overall NONOX leach retention time of between 0.5 and 8 hours is required and typically this could be between 0.5 to 3 hours.
Stream [122] is preferably cooled in a flash tank [125] where steam [123] and slurry [124] are separated and discharged.
The flash steam [123] may be scrubbed and then used for preheat duties. Excess flash steam can be released to atmosphere or employed elsewhere in the flowsheet for heating.
The cooled slurry [124] may be further cooled and then thickened in a decanter [130]. Flocculant [131] and recycle filtrate [133] and internal solution recycles can be employed to aid slurry thickening. The thickener overflow [132] can be further clarified before optionally neutralising it [180] with tailings [93] or similar alkali before disposal in the Tailings Storage Facility [169].
The thickened slurry [134] is filtered [135] and the upgraded copper concentrate is then washed with water [136] before it is repulped in clean water [139] and re-filtered and transferred as a washed upgraded copper concentrate [137].

EXAMPLE 1

A copper concentrate containing 28% Copper, 32% Iron, 35% total Sulfur, consisting of chalcopyrite, pyrite, and some bornite was tested, employing the flowsheet in FIG. 1. The silver concentration was 20 g/t.
The concentrate at 10% solids slurry density was subjected to a partial oxidation leach employing oxygen at 201° C., and the NONOX leach was conducted at about 2400 kPa. The oxidation process was complete within 18 minutes of retention time and the aqueous composition was:


	g/L

	Cu	23.4
	Fe	1.7
	S	20.6
	Ag	Below Detection
	Eh	438 mV (Ag/AgCl)

On completion of the partial oxidation step the slurry was advanced to the equivalent of the second compartment of the NONOX leach operating at 220° C. and the solubilised copper from partial oxidation leach was uploaded onto the remaining unleached concentrate. The final barren liquor composition after a further 50 minutes of leaching was:


	g/L

	Cu	0.7
	Fe	32.6
	S	29.3
	Ag	0.0001
	Eh	265 mV (Ag/AgCl)

The upgraded copper concentrate assayed:


	%

	Cu	51.2
	Fe	14.0
	S(Sulfide)	27.8
	Ag	38 g/t

The upgraded concentrate mass was 56% of the feed concentrate mass.

EXAMPLE 2

The same copper concentrate in Example 1 was tested employing the flowsheet in FIG. 1 but in this case air was employed as the oxidant. The NONOX leach was conducted at about 2400 kPa. The partial oxidation step required 90 minutes of retention time and the composition of liquor after partial oxidation was:


	g/L

	Cu	21.0
	Fe	33.3
	S	21.0
	Ag	Below Detection
	Eh	380 mV (Ag/AgCl)

On completion of the partial oxidation step the slurry was advanced to the equivalent of the second compartment of the NONOX leach operating at 220° C. where the solubilised copper from the partial oxidation leach was uploaded on to the remaining unleached concentrate.
The final barren liquor composition after a further 120 minutes of leaching was:


	g/L

	Cu	0.8
	Fe	36.7
	S	67.0
	Ag	0.0001
	Eh	259 mV (Ag/AgCl)

The final upgraded copper concentrate assayed:


	%

	Cu	60.8
	Fe	10.0
	S	22.2
	Ag	39 g/t

The upgraded concentrate mass was 48% of the feed concentrate mass.
Now that preferred embodiments of the hydrometallurgical method for the removal of impurities from a copper concentrate have been described in detail, it will be apparent that the described embodiments provide a number of advantages over the prior art, including the following:
(i) Single stage pressure leach flowsheet with less process steps, and hence a simplified operation.
(ii) The simplified operation results in lower capital and operating costs.
(iii) The mass of the concentrate entering the NONOX leach (feed concentrate mass) is significantly reduced once the concentrate exits the process (upgraded concentrate mass), resulting in ease of handling and lower transport costs.
It will be readily apparent to persons skilled in the relevant arts that various modifications and improvements may be made to the foregoing embodiments, in addition to those already described, without departing from the basic inventive concepts of the present invention. Therefore, it will be appreciated that the scope of the invention is not limited to the specific embodiments described and is to be determined from the appended claims.

Claims

1. A hydrometallurgical method for the removal at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) from a copper concentrate to produce an upgraded copper concentrate, the method comprising the step of:

subjecting the copper concentrate to an acidic leaching process (NONOX leach) in a multi-compartment autoclave using a copper sulfate containing lixiviant under lowered electrochemical conditions, to implement a controlled removal of between 20 and 99% of at least one of iron, aluminum, arsenic, sulfur, nickel, cobalt, zinc and cadmium (impurities) to produce the upgraded copper concentrate, with reduced soluble copper losses in the discharge stream, and wherein the NONOX leach is conducted:

to produce the majority of required soluble copper sulfate from a partial oxidation of the feed copper concentrate in a first compartment of the multi-compartment autoclave;

to pass the partially oxidised slurry from the first compartment into the second and ensuing compartments of the multi-compartment autoclave;

such that a division between the first compartment and the second and ensuing compartments does not permit transfer of unused oxidant gas from the first compartment to the second and ensuing compartments;

to provide a non-oxidising atmosphere in the second and ensuing compartments such that a majority of the soluble copper in the partially oxidised slurry from the first compartment is uploaded onto the partially leached solids to provide the upgraded concentrate;

employing elevated temperature and elevated pressure to suppress boiling in the leaching process; and

utilising acid generated from the hydrolysis of ferric iron in the first compartment to re-dissolve the ferric iron in the non-oxidising conditions of the second and ensuing compartments.

2. A hydrometallurgical method as defined in claim 1, wherein the leaching process (NONOX leach) is conducted at an electrochemical potential of greater than 150 mV (Ag/AgCl 3.8M KCl).

3. A hydrometallurgical method as defined in claim 2, wherein the leaching process (NONOX leach) is conducted at an electrochemical potential in the range of between about 175 mV and 450 mV (Ag/AgCl 3.8M KCl).

4. A hydrometallurgical method as defined in claim 1, wherein the electrochemical potential of the NONOX leach is controlled by the presence of, or the addition to the NONOX leach of, one or more of cupric sulfate, ferric ion, air, oxygen, pyrolusite, or hematite.

5. A hydrometallurgical method as defined in claim 1, wherein the copper sulfate containing lixiviant typically comprises one or more of the following solutes: sulfuric acid, sodium sulfate, copper sulfate, the copper sulfate containing lixiviant being obtained from within the NONOX leach, or from a source external to the NONOX leach.

6. A hydrometallurgical method as defined in claim 5, wherein the copper sulfate containing lixiviant also contains some chloride anions.

7. A hydrometallurgical method as defined in claim 1, wherein the NONOX leach is conducted between about 100° C. and 240° C.

8. A hydrometallurgical method as defined in claim 7, wherein the NONOX leach is conducted between about 160° C. and 240° C.

9. A hydrometallurgical method as defined in claim 1, wherein copper-iron-sulfides are transformed in the NONOX leach, and the NONOX leach is conducted under conditions to transform 50% to 99% of the copper-iron-sulfides to iron-depleted copper sulfides and covellite and to a sulfur-depleted variant (digenite/chalcocite).

10. A hydrometallurgical method as defined in claim 9, wherein pyrite and pyrrhotite are substantially altered under NONOX conditions to iron depleted copper-iron-sulfides.

11. A hydrometallurgical method as defined in claim 1, wherein the NONOX leach is conducted at a pressure in the range of 500-3500 kPa.

12. A hydrometallurgical method as defined in claim 11, wherein the retention time of the NONOX leach is between about 0.5 and 8 hours.

13. A hydrometallurgical method as defined in claim 1, wherein the partial oxidation of the feed copper concentrate in the first compartment of the multi-compartment autoclave employs oxygen, oxygen enriched air or air as the oxidant.

14. A hydrometallurgical method as defined in claim 1, wherein the first compartment of the multi-compartment autoclave is preferably divided from the other compartments by a divider construction that restricts the free flow of the autoclave vapour phase from the first compartment to other compartments, i.e. the divider construction forms the division between the first compartment and the second and ensuing compartments so that the transfer of unused oxidant gas from the first compartment to the second and ensuing compartments is minimised or not permitted.

15. A hydrometallurgical method as defined in claim 14, wherein a counter current vapour flow within the autoclave is usually present such that the second and remaining compartments have a vapour phase that is essentially free of oxygen.

16. A hydrometallurgical method as defined in claim 1, further comprising feeding into the NONOX leach a copper sulfate slurry from a crystalliser or a copper sulfate leach liquor from an external pressure oxidation reactor (PROX reactor), or copper sulfate solution from an external source, or a copper sulfate leach slurry direct from a PROX reactor.

17. A hydrometallurgical method as defined in claim 16, wherein the copper sulfate slurry typically comprises copper sulfate, iron sulfate and sulfuric acid.

18. A hydrometallurgical method as defined in claim 1, wherein the copper sulfate can be added to more than one compartment in the NONOX leach reactor.

19. A hydrometallurgical method as defined in claim 1, wherein the copper aqueous concentration levels within the discharge liquor from the NONOX leach reactor are less than 5 g/L and more typically less than 1 g/L.