Note : Les descriptions sont présentées dans la langue officielle dans laquelle elles ont été soumises.
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"METHOD FOR THE RECOVERY OF BASE METALS FROM ORES"
Field of the Invention
The present invention relates to a method for the recovery of base metals
from sulphide and/or oxide ores. More particularly; though not exclusively,
the invention relates to a hydrometallurgical method for leaching of nickel
from a combination of nickel sulphide and nickel oxide ores.
Background to the Invention
Nickel sulphide ores have traditionally been treated via a pyrometallurgical
smelting process, in order to recover nickel as a high grade nickel matte.'The
nickel content of the matte can range from 60 to 80% nickel as a sulphide. In
Western Australia flash smelting and converting has been commercially
applied to produce a high grade nickel matte 70% nickel, from nickel sulphide
concentrates. The nickel sulphide concentrate is typically 12 to 18% nickel.
The high grade matte is subsequently refined utilising the Sherritt Gordon
process.
Hydrometallurgical processes such as leaching have historically not been
applied to nickel sulphide ores or concentrates, as .smelting is commercially
competitive when. compared to hydrometallurgical processes. Unlike the
Activox or Albion hydrometallurgical processes, smelting unlocks significant
energy credits that is converted to electrical energy and produces sulphuric
acid or sulphur, as by-products. This co-generation approach improves the
overall competitiveness of pyrometallurgical process when compared to
hydrometallurgical processes.
Furthermore hydrometallurgical processes such as the Activox or the Albion
Process typically require fine grinding P90 or minus 10 microns, which
consumes energy. The energy released via the leaching process is lost to
cooling towers or a simple flash system that does not capture any of the
energy released during leaching.
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It is well documented that hydrometallurgical treatments such as High
Pressure Acid Leach (HPAL) plants operating in Western Australia have
added sulphides as either "transition" sulphide ore or non-smeltable
concentrates that cannot be treated via a conventional concentrator or
smelter. However, these plants are limited in their ability to add sulphides
due
to the reducing nature of the sulphide ores or concentrates. The reducing
potential of the sulphides lowers the oxidation reduction potential (ORP) and
can result in damage to the HPAL titanium autoclave lining. Therefore these
plants are selective in the type of sulphide ore or concentrates added as
secondary feed and they are also significantly limited in the amount of
sulphide ore or concentrate that can be added to the HPAL process.
The treatment of sulphide ores via pyrometallurgical methods requires
significant capital expenditure for flash furnace, convertors, slag treatment,
as
well as the utilities required such as power generators and sulphuric acid
plants to capture sulphur dioxide emissions.
Some nickel sulphide ores also contain magnesium (Mg expressed as MgO)
and when treated through a concentrator produce a sulphide concentrate that
has a low iron (Fe) to MgO ratio. A low iron to MgO ratio impacts on flash
furnace slag chemistry. The slag becomes viscous and can be difficult to
remove from the furnace without increasing the slag operating temperature.
Some of these ores also contain arsenic (As) at levels that require careful
blending to manage the occupational health and safety aspects of arsenic
and its impact on human health, which makes these sulphide ores
undesirable for smelting.
The present invention aims to alleviate or at least partially alleviate some
of
the difficulties associated with the conventional pyrometallurgical and
hydrometallurgical processes for the treatment of nickel sulphide ores or
concentrates. However it will be understood that it is not limited in its
application to nickel sulphide ores or concentrates.
The previous discussion of the background to the invention is intended to
facilitate an understanding of the present invention only. The discussion is
not
an acknowledgement or admission that any of the material referred to is or
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was part of the common general knowledge as at the priority date of this
application. 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 present invention there is provided a
hydrometallurgical method for leaching a base metal in a combined pressure
acid leach, the method comprising the steps of:
combining a sulphide ore or concentrate with a laterite or other oxide ore and
mixing them together to form a slurry;
leaching the combined slurry in a pressure acid leach circuit; and,
providing an oxidant to the pressure acid leach circuit wherein the oxidant
allows for the conversion of substantially all of the sulphide to transition
through to sulphate.
Preferably the step of combining sulphide ore or concentrate with laterite or
other oxide ore and mixing them involves milling the ores together.
Preferably the method comprises the further step of leaching the combined
sulphide and oxide ore together with pregnant leach solution produced from
an atmospheric leach circuit, and without the requirement of additional
sulphuric acid.
Typically the base metal is selected from the group consisting of nickel,
cobalt, copper, lead and zinc. Preferably the sulphide ore or concentrate is a
nickel sulphide or concentrate and the laterite or other oxide ore is a nickel
laterite or other nickel oxide ore.
Preferably the method further comprises the steps of directing a nickel
laterite
or other nickel oxide ore to an atmospheric leach process to produce a
pregnant leach solution, and adding the pregnant leach solution to the
combined sulphide ore or concentrate and laterite or other nickel oxide ore
during the combining step.
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Preferably the pressure acid leach circuit comprises a series of pressure
Pachuca tanks.
The preferred oxidant is oxygen or air; ferric iron generated in a separate
atmospheric leach process, and/or manganese present in the ores as
manganese dioxide (Mn02) may also be used.
In a further preferred aspect of the invention, the method further comprises a
roasting step in which pyrite is roasted to form a metallurgical gas. The
metallurgical gas is preferably cooled and conditioned before passing to a wet
end of a sulphuric acid plant, the sulphuric acid plant typically being used
to
generate sulphuric acid for use in the method of the invention as required.
Preferably the pyrite is mined from the same ore deposit as the sulphide ore
used in the combining step of the invention. Preferably the metallurgical gas
comprises at least about 9 to 11 % sulphur dioxide (SO2).
According to another aspect of the present invention there is provided a
hydrometallurgical method for leaching nickel, the method comprising the
steps of:
combining nickel sulphide ore or concentrate with nickel laterite or other
oxide
ore and mixing them together to form a slurry;
leaching the combined slurry in a pressure acid leach circuit; and,
providing an oxidant to the pressure acid leach circuit wherein the oxidant
allows for the conversion of substantially all of the nickel sulphide to
transition
through to nickel sulphate.
Preferably the method further comprises the steps of directing a nickel
laterite
or other nickel oxide ore to an atmospheric leach process to produce a
pregnant leach solution, and adding the pregnant leach solution to the
combined sulphide ore or concentrate and laterite or other nickel oxide ore
during the combining and milling step.
The pregnant leach solution (PLS) from the atmospheric leach circuit
preferably has a ferric iron concentration within the range of 10 to 60 g/L.
Preferably, the PLS from the atmospheric leach circuit has a free acid
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concentration of less than 30 g/I. Preferably the PLS from the atmospheric
leach circuit has a nickel concentration of >4 g/I.
The nickel sulphide ore or concentrate preferably has a nickel concentration
within the range of about 1 to 10% Ni. Preferably, the nickel laterite or
oxide
ore has a nickel concentration within the range of 0.8 to 3%. More preferably,
the nickel laterite or oxide ore for the atmospheric leach is a saprolite or
smectite ore and the laterite or oxide ore required for the pressure acid
leach
is a limonite ore.
Preferably the pressure acid leach circuit comprises a series of pressure
Pachuca tanks.
Preferably, the free acid concentration achieved in the pressure acid leach is
maintained within the range of 30 to 80 g/l. Preferably, the temperature
within
the pressure Pachuca tanks is maintained between 160 to 260 C. More
preferably, the temperature within the pressure Pachuca tanks is maintained
at about 220 to 250 C. Preferably, the oxygen over pressure within the
pressure Pachuca tanks is maintained between 100 to 1000 kPag.
Typically, the ratio of nickel sulphide ore or concentrate: nickel laterite or
other nickel oxide ore in the combining step is about 3:7.
Advantageously the process of leaching the combined slurry approaches
autogenous leaching, releasing energy, generating sulphuric acid and
producing hematite and alunite as the predominant residue minerals.
According to another aspect of the present invention there is provided a
hydrometallurgical method for leaching nickel, the method comprising the
steps of:
combining nickel sulphide ore or concentrate with nickel laterite or other
nickel oxide ore in relative proportions selected to achieve optimum density
and milling them together to form a slurry;
leaching the combined slurry in a pressure acid leach circuit; and,
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providing air or oxygen to the pressure acid leach circuit wherein the air or
oxygen allows for the conversion of substantially all of the nickel sulphide
to
transition through to nickel sulphate.
Preferably the method further comprises the steps of directing a nickel
laterite
or other nickel oxide ore to an atmospheric leach process to produce a
pregnant leach solution, and adding the pregnant leach solution to the
combined sulphide ore or concentrate and laterite or other nickel oxide ore
during the combining and milling step.
Typically, the ratio of nickel sulphide ore or concentrate: nickel laterite or
other nickel oxide ore in the combining step is about 3:7.
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 several specific embodiments of the hydrometallurgical
method for leaching of a base metal according to the invention, given by way
of example only, with reference to the accompanying drawing in which:
Figure 1(a) and (b) is a schematic diagram of a process circuit of a preferred
method for leaching nickel in accordance with the present invention; and
Figure 2(a) and (b) is a schematic diagram of a process circuit of a further
preferred method for leaching nickel in accordance with the present invention.
Detailed Description of Preferred Embodiments
A preferred embodiment of the hydrometallurgical method for leaching of a
base metal according to the invention, as shown in schematic form in Figure
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1, relates to the leaching of nickel. The method preferably comprises the step
of directing a nickel laterite or other nickel oxide ore to an atmospheric
leach
process 10, which in the embodiment of Figure 1 is a first heap leach process
(not illustrated). The clarified pregnant, leach solution (PLS) from the first
heap
leach process is then directed to a milling circuit 12. The PLS is preferably
heated prior to injection into the milling circuit 12. The PLS may be derived
from any suitable atmospheric leach process and it not limited to heap
leaching. However in the event that a suitable source of PLS from an
atmospheric leach process is not available, water may be substituted for the
PLS that is directed to the milling circuit.
The method further comprises the step of combining nickel sulphide ore or
concentrate 14 with nickel laterite or other nickel oxide ore 16 and milling
the
combination in the milling circuit 12 with the clarified PLS from the first
heap
leach process 10 (and/or water as the case may be). The nickel sulphide ore
or concentrate 14 preferably has a nickel concentration within the range of
about 1 to 10% Ni. Preferably, the nickel laterite or other nickel oxide ore
16
should have a nickel concentration within the range of 0.8 to 3% Ni.
Typically,
the ratio of nickel sulphide ore (or concentrate): nickel laterite ore (or
other
nickel oxide ore) is about 3:7. More preferably, the nickel laterite or nickel
oxide ore for the atmospheric leach is a saprolite smectite ore and the
laterite
or oxide ore used for the combined leach is a limonite ore.
The viscosity of laterite ores is impacted by additives such as free acid or
total
dissolved solids. Limonites typically exhibit a reduction in viscosity when
solutions from a heap leach operation are slurried with limonite ores. That
is,
for a given weight percent, solids milling in PLS reduces the viscosity of the
pulp. However with saprolite or smectite ores slurrying in PLS will increase
the viscosity for a given weight percent solids. Adding sulphides to all
laterite
ores, whether limonite, saprolite or smectite, acts to reduce the viscosity
and
is considered innovative. By appropriate selection of the relative proportions
of both kinds of minerals in the combined ores, milling at optimum density can
be achieved. Therefore saprolite or smectite is the preferred ore for the
atmospheric leach, and limonite is the preferred ore for milling in
atmospheric
PLS due to the improvement in slurry density achieved.
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The milling is typically carried out using the PLS from the first heap leach
process instead of, or in addition to, water. The clarified PLS from the first
heap leach process preferably has a ferric iron concentration within the range
of 10 to 60 g/l. Preferably the PLS from the first heap leach process 10 has a
free acid concentration of less than 30g/l. Preferably the PLS from the first
heap leach process 10 has a nickel concentration of more than 4 g/l. Hence a
further significant benefit of adding the PLS to the milling circuit 12 is
that the
head grade of ore passing through the plant is doubled. This, together with
acid credits, greatly improves the economies of scale and efficiency of the
plant.
The milled combined ore from the milling circuit 12 is then subject to a
screening step in screening circuit 18. Oversize ore is directed from the
screening circuit 18 back to the first heap leach process 10. Undersize ore is
fed from the screening circuit 18 to a slurry tank 19, and the combined slurry
is then pumped by high pressure slurry pumps to a combined pressure acid
leach (CPAL) circuit 20. Wash from the screening circuit 18 is returned to the
milling circuit 12.
In the illustrated embodiment the combined pressure Pachuca acid leach
(CPPAL) circuit 20 comprises a series of pressure Pachuca tanks 22, in
which pressure leaching of the combined slurry occurs preferably without the
addition of sulphuric acid. Preferably the temperature within the pressure
Pachuca tanks is maintained between 160 to 260 C, and more preferably
between 220 to 250 C. The use of brick-lined Pachuca tanks 22 instead of
an autoclave, is much less expensive to maintain due to the high
maintenance cost of the titanium used in autoclaves. Typically, pressure
leaching in the CPPAL 20 occurs for about 90 minutes at a temperature of
about 250 C and a pressure of about 44 bar. The nickel and cobalt in the
combined slurry is converted to soluble sulphates. More particularly, the
nickel and cobalt in the form of sulphides is largely converted to the metal
sulphates via the usual chemical transition steps.
An oxidant is preferably injected into the Pachuca tanks 22. The preferred
oxidant is oxygen or air; ferric iron generated in a separate atmospheric
leach
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process, and/or manganese present in the ores as manganese dioxide
(Mn02) may also be used. Preferably, the oxygen over pressure within the
pressure Pachuca tanks is maintained between 100 to 1000 kPag. Without
oxygen or air in the CPPAL a significant amount of iron as ferrous remains in
solution "locking up" sulphuric acid. However with sufficient oxygen (or other
oxidant) most of the iron is converted to hematite.
The combined ore is preferably leached in the pressure Pachuca tanks 22
without adding acid and most of the energy required is generated by the
oxidation of the sulphide minerals. Advantageously the process within the
CPPAL 20 approaches autogenous leaching, releasing energy, generating
sulphuric acid and producing hematite and alunite as the predominant residue
minerals. As is well known in the art, sulphides produce acid, whereas
laterites and other oxide ores consume acid. Therefore with the correct
balance of the two kinds of ores in the combined slurry fed to the CPPAL 20
the acid released in the hydrolysis circuit as free acid can be matched with
the acid consumed. However if additional acid is required an acid plant 24 is
provided to direct high pressure acid to the Pachuca tanks 22. Steam
generated by the acid plant 24 is also injected into the Pachuca tanks 22 at
60 bar. The balance of the steam may be used for generating power.
The resulting leach slurry exiting from the pressure Pachuca tanks 22 is at
high temperatures (typically between 160 to 260 C). Some of this heat is
used to preheat the combined slurry as it is fed into the CPPAL 20 to further
improve the efficiency of the CPPAL 20. For this purpose a heater coil 26
uses some of the waste heat to generate steam, which is then fed back to a
preheat circuit 28 on the CPPAL feed to heat the combined slurry to about
176 C. The PLS solutions have been found to scale in a similar way that an
autoclave will scale in a typical HPAL operation. Increasing the temperature
beyond 180 C will initiate the precipitation of iron and aluminium
predominantly. In the proposed system the first stage preheater 28 is kept
below 180 C to prevent iron and aluminium in the PLS solution (added to the
milling circuit) from precipitating in the preheat circuit.
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The leach slurry exiting from the CPPAL 20 undergoes further cooling to
atmospheric temperature and pressure in a flash cooling circuit 30. Waste
heat from the flash cooling circuit 30 is used for preheating the PLS from the
first heap leach process 10. The leach slurry then undergoes solid/liquid
separation (thickening) in a counter current decantation (CCD) circuit 32 (see
Figure 1(b)). The CCD circuit 32 typically consists of five thickeners which
separate the slurry into two streams. The first stream consisting of the
pressure leach residue solids is eventually sent as underflow from the last
thickener to tailings. The second stream consisting of the clarified solution
(containing dissolved nickel and cobalt sulphates) is then preferably directed
to a second heap leach process 34 (not shown).
The clarified solution from the CCD circuit 32 still has a substantial volume
of
free acid (typically about 20 litres/tonne) available for further downstream
processing. The purpose of the second heap leach process 34 is to utilise
this free acid for further leaching of nickel (and cobalt) from a nickel
laterite or
other nickel oxide ore, rather than wasting it by removal in a neutralisation
circuit. However, if preferred, the free acid can be removed in a conventional
neutralisation circuit, without passing the solution through a second heap
leach.
The PLS from the second heap leach process 34 is then subject to iron
removal using calcrete in a conventional precipitation circuit 36 consisting
of.a
series of agitated tanks. Calcrete is not as efficient as limestone or
quicklime
for iron removal, however it is readily available locally. The calcrete 37 is
milled and mixed with water to produce calcrete slurry. The calcrete slurry is
used to neutralise the free acid in the PLS from the second heap leach
process, and to precipitate the ferric ions in solution as jarosite/geothite.
An iron free clarified solution is obtained by treating the solution from the
precipitation circuit 36 in a second CCD circuit 38 for solid/liquid
separation
(thickening). The thickener underflow solids are discharged with the pressure
leach residue solids from the first CCD circuit 32 to tailings. The iron free
clarified solution is then subjected to a direct solvent extraction process 40
and electrowinning process 42 in a conventional manner for the extraction
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and refining of nickel, cobalt, zinc and copper. Solvent extraction raffinate
and
a barren liquor bleed from the electrowinning process 42 is returned to the
first CCD circuit 32 as wash liquor.
The invention may also include the additional step of mining pyrite 13 from
the same deposit 15 as that from which the nickel sulphide 14 ore is mined.
As can be seen in Figure 2a, the iron sulphides of the pyrite 13 are roasted
in
a roasting step 17 and the resulting metallurgical gas 21 is cooled through a
waste heat boiler 23. The resulting steam 25 is used to drive a turbine.
thereby generating power which may be used in the process plant as
required. The metallurgical gas stream passes to a conditioning step 27
through a wet end of the acid plant 24 prior to making sulphuric acid which is
then used to leach nickel and cobalt in an atmospheric leach process. In this
way, the pyrite mined with sulphide ore is burned in an initial process step
to
ultimately generate sulphuric acid for use in the overall extraction process.
In
other respects, the invention shown in Figures 2a and 2b resembles that of
the embodiment of Figures la and 1b, and will not be described in further
detail here.
Typically the metallurgical gas 21 comprises at least about 9 to 11 % sulphur
dioxide (SO2). The gas 21 is prepared for the integration into a standard
sulphur burning sulphuric acid plant 24. As it is the product of a
metallurgical
process it passes through a heat recovery section or boiler 23, following
which is passes to a conditioning step 27. In the conditioning step, the gas
21
passes to a humidifying tower to remove volatiles and some water, it then
passes to mist precipitators which remove most of the water, following which
it passes to a drying tower for final water removal. The gas stream then
passes into the acid plant 24. Sulphur is burnt as required to maintain a
nominal gas strength.
The present invention is further illustrated by way of the following non-
limiting
examples:
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Example I
One combined pressure acid leach was conducted at 255 C at 30% (w/w)
sulphide ore. The approximate ratio of sulphide ore: laterite ore used was of
the order of 3:7, and the ores were milled together in accordance with the
invention. The compositions of the laterite and sulphide ores are provided in
Table 1.
Table I
Sample Unit Analysis
Ni Co Fe Fe(Ill) Mg S2- S FA
Sulphide % 1.25 0.009 15.0 - 4.1 10.5 0.8 -
Laterite % 1.2 0.09 17.0 - 5.7 - - -
Leach Sol g/L 5.5 0.40 47 45.0 25 - - 15
The results from the leach test are provided in Table 2.
The main observations from this test were:
= No acid addition to the CPPAL was required
= 700 kPag oxygen overpressure was maintained using oxygen
= Almost all of the iron precipitated as hematite
= Almost all the aluminium precipitated as alunite
Over 94% nickel and cobalt extraction was achieved for the
combined leach
= The nickel concentration in the PLS after 90 minutes was much
higher than for a typical HPAL circuit (10 g/L compared with 6 g/L).
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Table 2
Test I Ni (%) Co (%) FA (g/L) ORP (mV)
CPPAL 97% 96% 75 /I 881
The test conditions are provided in Table 3.
Table 3
Test No. % (w/w) Vol. % of Acid Addition, kg/t
Sulphide Ore Column PLS
1 30 100 Nil
Conclusions:
The CPPAL of a combined laterite or oxide ore with a sulphide ore or
concentrate when leached in pregnant solution from a separate atmospheric
leach process, with an overpressure of oxygen of 700 kPag, can successfully
extract nickel and cobalt without the addition of acid.
It is therefore envisaged that with appropriate blending of sulphide ore or
concentrate in combination with a laterite or oxide ore milled in pregnant
solution from an atmospheric leach circuit, the combined blend can be
successfully leached without the addition of acid in a pressure Pachuca with
oxygen over pressure.
It is further envisaged that saprolite or smectite is used for the atmospheric
leach and limonite is the preferred ore for milling in atmospheric leach
pregnant solution due to the improvement in slurry density achieved.
It is further envisaged that clarified solution from either a heap leach or
atmospheric leach can be applied to replace water for milling in the CPPAL
circuit.
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It is further envisaged that a direct synergistic solvent extraction and
electro-
winning process is applied for the extraction and refining of nickel, cobalt,
zinc
and copper.
It is also envisaged that the CPPAL process can be applied to a wide variety
of nickel, cobalt, copper and zinc bearing laterite and sulphide ores or
concentrates.
Furthermore it is also envisaged that some mines can be unconstrained by
the successful application of the above process as it is tolerant to arsenic
and
iron to magnesium chemistry.
Now that preferred embodiments of a hydrometallurgical method for leaching
nickel in a combined pressure acid leach have been described in detail, it
will
be apparent that the embodiments provide a number of advantages over the
prior art, including the following:
(i) The combined pressure acid leach (CPPAL) significantly
reduces the acid production requirement per tonne of ore
treated.
(ii) The CPPAL process approaches autogenous heating therefore
reducing external energy generation requirements.
(iii) The CPPAL process is undertaken in brick-lined Pachuca tanks
instead of an autoclave, which are less expensive to maintain
than titanium-lined autoclaves.
(iv) The CPPAL process can be applied to a wide variety of nickel,
cobalt, copper, lead and zinc bearing laterite and sulphide ores
or concentrates.
It will be readily apparent to persons skilled in the relevant art 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.
