U.S. patent application number 15/101241 was filed with the patent office on 2016-10-20 for copper processing method.
The applicant listed for this patent is THE UNIVERSITY OF QUEENSLAND. Invention is credited to William HAWKER, Peter HAYES, Evgueni JAK, James VAUGHAN.
Application Number | 20160304988 15/101241 |
Document ID | / |
Family ID | 53272648 |
Filed Date | 2016-10-20 |
United States Patent
Application |
20160304988 |
Kind Code |
A1 |
VAUGHAN; James ; et
al. |
October 20, 2016 |
COPPER PROCESSING METHOD
Abstract
A method of processing a copper-containing source material is
provided whereby an aqueous acidic leach solution of the
copper-containing source material is formed and then contacted with
a pH increasing agent to thereby cause the precipitation of a
copper-containing intermediate. The copper-containing intermediate
can then be collected and exposed to a high temperature treatment,
such as would be encountered in smelter or converter
operations.
Inventors: |
VAUGHAN; James; (St Lucia,
Queensland, AU) ; HAWKER; William; (St Lucia,
Queensland, AU) ; HAYES; Peter; (St Lucia,
Queensland, AU) ; JAK; Evgueni; (St Lucia,
Queensland, AU) |
|
Applicant: |
Name |
City |
State |
Country |
Type |
THE UNIVERSITY OF QUEENSLAND |
Queensland |
|
AU |
|
|
Family ID: |
53272648 |
Appl. No.: |
15/101241 |
Filed: |
December 3, 2014 |
PCT Filed: |
December 3, 2014 |
PCT NO: |
PCT/AU2014/050393 |
371 Date: |
June 2, 2016 |
Current U.S.
Class: |
1/1 |
Current CPC
Class: |
C22B 15/0071 20130101;
Y02P 10/236 20151101; C25C 1/12 20130101; C22B 15/0069 20130101;
Y02P 10/20 20151101; C22B 1/02 20130101; C22B 3/10 20130101; C22B
3/44 20130101; C22B 3/08 20130101; C22B 15/0073 20130101; C22B
1/248 20130101; C22B 15/0015 20130101; C22B 3/065 20130101; C22B
15/0052 20130101; C22B 15/0089 20130101; C22B 15/0086 20130101 |
International
Class: |
C22B 15/00 20060101
C22B015/00; C22B 3/10 20060101 C22B003/10; C22B 1/248 20060101
C22B001/248; C22B 3/06 20060101 C22B003/06; C22B 3/44 20060101
C22B003/44; C25C 1/12 20060101 C25C001/12; C22B 1/02 20060101
C22B001/02; C22B 3/08 20060101 C22B003/08 |
Foreign Application Data
Date |
Code |
Application Number |
Dec 3, 2013 |
AU |
2013904688 |
Jan 30, 2014 |
AU |
2014900280 |
Claims
1. A method of processing a copper-containing source material
including the steps of: (a) providing an aqueous acidic leach
solution of the copper-containing source material; (b) introducing
a pH increasing agent into the acidic leach solution to cause the
precipitation of a copper-containing intermediate; and (c)
collecting the copper-containing intermediate and exposing it to a
high temperature treatment, optionally transporting the
copper-containing intermediate to another location prior to its
exposure to the high temperature treatment to thereby process the
copper-containing source material.
2. The method of claim 1 wherein the copper-containing source
material is a copper-containing ore, copper smelter slag,
copper-containing tailings, copper concentrate or a
copper-containing process intermediate or waste product.
3. The method of claim 2 wherein the copper-containing ore is
selected from a copper sulphide or copper oxide ore.
4. The method of claim 1 wherein the aqueous acidic leach solution
is formed from an acid selected from the group consisting of
sulphuric, nitric and hydrochloric acids.
5. The method of claim 1 wherein the pH increasing agent is
selected from the group consisting of calcium oxide, calcium
carbonate, calcium hydroxide, calcium ferrite, magnesium oxide,
magnesium carbonate, magnesium hydroxide, sodium carbonate, sodium
hydroxide, dolomite and materials containing one or more of these
compounds.
6. (canceled)
7. The method of claim 1 wherein the pH of the acidic leach
solution of the copper source material is raised by the pH
increasing agent to between about 4.0 to about 9.0.
8. The method of claim 1 wherein the high temperature treatment is
at a temperature of between about 1000.degree. C. to about
1400.degree. C.
9. The method of claim 1 wherein the high temperature treatment is
carried out in a smelter or converter.
10. The method of claim 9 wherein when the high temperature
treatment is performed in a converter then an oxygen-containing gas
and a sulphide compound are introduced into the heating
environment.
11. The method of claim 1 further including a step, prior to step
(b), of introducing an effective amount of a pH increasing agent
into the acidic leach solution to cause the precipitation of an
iron-containing compound without any substantial precipitation of
the copper containing-intermediate.
12. The method of claim 11 wherein the pH increasing agent will
raise the pH of the acidic leach solution to be between about 1.5
to 3.5 to cause the precipitation of an iron-containing compound
without any substantial precipitation of the copper
containing-intermediate.
13. The method of claim 1 further including the step of exposing
the copper-containing intermediate to a separation step prior to
the high temperature treatment.
14. The method of claim 13 wherein, when the pH increasing agent is
lime and/or limestone, the separation step is the separation of
precipitated gypsum from the copper-containing intermediate based
upon a difference in the particle sizes thereof.
15. The method of claim 1 further including the step, prior to step
(c), of heating the copper-containing intermediate to a temperature
between 200.degree. C. to 800.degree. C. to provide a more
concentrated copper-containing intermediate to introduce to step
(c).
16. The method of claim 1 wherein the copper-containing source
material is a copper sulphide-containing source material and the
method further includes the step (a-i) of exposing the copper
sulphide-containing source material to an oxidative roast, prior to
its exposure to the leach solution, to form a calcined
copper-containing source material.
17. The method of claim 16 wherein the oxidative roast is performed
at a temperature of between about 500.degree. C. to about
950.degree. C.
18. The method of claim 16 further including the step (a-ii) of
contacting the calcined copper-containing source material with a
leach solution to form the aqueous acidic leach solution of step
(a).
19. The method of claim 18 wherein the aqueous acidic leach
solution after substantial dissolution of the calcined
copper-containing source material has a pH of between about 2 to
about 5.
20. A method of processing a copper sulphide-containing source
material including the steps of: (a-i) exposing the copper
sulphide-containing source material to an oxidative roast to form a
calcined copper-containing source material; (a-ii) contacting the
calcined copper containing source material with a leach solution to
form an acidic copper-containing leach solution; (b) introducing a
pH increasing agent into the acidic copper-containing leach
solution to cause the precipitation of a copper-containing
intermediate; and (c) collecting the copper-containing intermediate
and exposing it to a high temperature treatment, to thereby process
the copper sulphide-containing source material.
21. The method of claim 20 further including the step (aa-i) of
contacting the copper sulphide-containing source material under
non-oxidising conditions, prior to the oxidative roast, with an
acidic or alkaline solution to leach out select impurities.
22. (canceled)
Description
FIELD OF THE INVENTION
[0001] The present invention relates to a method of processing a
copper-containing ore or other source material. Particularly, the
present invention relates to a method of processing an ore or
source material to recover copper, or a suitable copper compound,
therefrom.
BACKGROUND TO THE INVENTION
[0002] Any reference to background art herein is not to be
construed as an admission that such art constitutes common general
knowledge in Australia or elsewhere.
[0003] Copper is a highly valuable commodity with increased demands
being placed on production due, in part, to the rapid growth of
economies such as India and China. At the same time the
identification of new high grade ore deposits is becoming
increasingly challenging.
[0004] The majority of copper is extracted from its ores in one of
two ways. Most copper sulphide ores are concentrated by flotation
and treated using a pyrometallurgical route while other copper
ores, for example copper oxide ores and some lower grade sulphide
ores, are treated using a hydrometallurgical route. These two
methods have some inherent advantages and particular disadvantages
that become more profound as lower grade ores are treated.
[0005] If there is significant high grade sulphide minerals in an
ore body that can be readily liberated, then the most economic and
efficient route is to concentrate the minerals by grinding and
flotation, then treat the concentrate by smelting. This
pyrometallurgical process makes use of the highly exothermic
sulphur oxidation reaction to heat the concentrate minerals to high
temperatures which favours the reduction of copper to its metallic
state.
[0006] The high temperature component of this process typically
involves a two stage process of smelting and converting of the
copper; both involve the introduction of an oxygen-containing gas.
In some instances a direct-to-copper' smelting operation may be
employed using, for example, a flash furnace smelter. In the
smelting and converting processes the sulphides are oxidised by the
oxygen. The sulphur oxidation reaction also produces poisonous
sulphur dioxide gas which must be captured to avoid its release to
the environment. The sulphur dioxide gas can be used to produce
commercially valuable sulphuric acid.
[0007] The extent of the sulphide oxidation is controlled through
the two stages of the process such that in the smelting stage a
higher grade copper sulphide liquid, commonly referred to as matte,
is produced. The converting stage then produces a metallic copper
product, still containing some impurities, commonly referred to as
"blister copper". Silicon and/or calcium oxides may be added in
each stage to produce a separate liquid oxide phase referred to
commonly as slag. Significant proportions of the chemically bound
iron and other impurity elements are removed to this slag phase.
The composition of the slag is critical in ensuring the control of
chemical partitioning of metal species between gas, slag and
relatively dense copper rich phases, the proportion of solid oxides
present in the slag and hence the physico-chemical properties of
the slag itself. The molten slag and high grade matte separate due
to density differences. Some copper is lost to the slag phase by
oxidation to copper oxide or entrainment of matte or blister
copper. Direct, single-stage smelting of sulphide concentrate to
blister copper and slag can also be carried out, and this route is
employed commercially for concentrates having relatively low iron
concentrations.
[0008] The presence of calcium oxide in the slag phase has been
shown to be beneficial to these slag practices for a number of
reasons including that it improves the slag's ability to absorb
impurities such as arsenic, bismuth and antimony. The matte and
slag are tapped separately in the smelting stage and the matte is
transferred to the converting stage. In the converting stage the
slag and "blister" copper are tapped separately. A typical flow
sheet for this process is shown in FIG. 1.
[0009] Importantly, in both the smelting and copper converting
stages the chemical reactions are exothermic and provide heat to
sustain the processes at temperature. The heat balance in these
process units can be challenging to control and places limitations
on the process.
[0010] While most converting operations are still carried out in
batch converters, continuous converting of copper mattes is now
also undertaken using silica and/or calcium oxide as flux. The
silica and/or calcium oxide reacts with magnetite, molten copper
and oxygen to form molten slag. Both silica-rich and calcium-rich
(calcium ferrite) slags can be used in copper converting processes.
The partitioning of impurities such as arsenic, bismuth and
antimony between gas, metal and slag is dependent on process
variables, such as, slag composition, oxygen partial pressure and
temperature of the system, in addition to the reactor design and
operating practice. In general the presence of calcium oxide in
slag is beneficial in removing impurity elements from the metal
phase.
[0011] One of the major energy consuming steps in this concentrate
smelting route is in the electrical power used for the grinding of
the ores. A decrease in the ore grade, for example from 2 to 1 wt %
copper, at least doubles the energy required to produce copper
metal as twice as much ore must be treated in order to produce the
same amount of copper. The pyrometallurgical approach is thus
economically limited in the grade of ore it can process. This is
becoming increasingly problematic as the copper content of newly
discovered ores is steadily decreasing.
[0012] In addition, the arsenic content in many copper ores is
increasing. The detrimental effects of release of this element into
the environment are well recognised and it is essential that this
problem is also addressed. Arsenic is currently a problem in
conventional sulphide pyrometallurgy since the process conditions
in the first step, the smelting step, result in relatively reducing
conditions and the arsenic partitions preferentially to the gas
phase as species such as arsenic trioxide or arsenic trisulphide.
This creates significant problems with gas cleaning and disposal of
arsenic-containing fume.
[0013] The pyrometallurgical process is principally applicable to
sulphide ores. If the predominant copper minerals in an ore are not
sulphides, the ore is difficult to concentrate by physical means
and is unsuited to pyrometallurgical processing as the cost of
heating the host rock is prohibitive in terms of energy and cost.
Further, if certain impurities (such as arsenic) in a sulphide ore
and resulting concentrate surpass a critical concentration then
that ore cannot be treated using the pyrometallurgical processes.
Instead, processing by hydrometallurgical techniques becomes the
most economical method of extracting the copper.
[0014] The conventional hydrometallurgical process for extracting
copper is a leaching, solvent extraction and electrowinning
process. This hydrometallurgical process typically consists of
three closed loop circuits, as shown in FIG. 2. The first is a
sulphuric acid leach where the copper is dissolved along with
certain impurity elements, such as iron. The pregnant leach
solution is separated from the ore leach residue and contacted with
an organic chelating reagent.
[0015] The organic phase comprises the second of the closed loop
circuits and acts as a cation exchanger by releasing protons while
selectively bonding copper ions, particularly in preference to
ferric iron. The organic and aqueous phases are not miscible and
are physically separated by density difference. The relatively
dense mildly acidic aqueous phase, also known as the raffinate, is
recycled back to the leaching stage. The copper loaded organic is
then contacted with a highly acidic solution recycled from the
copper electrowinning stage which forms the third of the closed
loop circuits. The high acid concentration in this solution
reverses the cation exchange reaction, stripping copper off the
organic chelating reagent and into the aqueous phase while loading
the organic with protons. The organic phase loaded with protons can
be recycled back to contact with the leach solution. High purity
metallic copper is electrowon from the copper loaded aqueous
phase.
[0016] Electrowinning copper from the aqueous cupric sulphate
solution to metallic copper requires about 2 kWh/kg-cu. At the
cathode, copper from solution is reduced to copper metal while at
the anode, water is oxidised to produce oxygen gas and protons,
regenerating the acid solution required for stripping the copper
from the organic phase in the solvent extraction process.
[0017] The advantages of this process are well known. It is well
understood and relatively easy to operate, produces high purity
cathodes, can be used to treat quite dilute acid leach solutions
and is suitable for both small and large operations. However, a
handful of key issues can make the process unsuitable for
particular ore bodies. The capital costs for the solvent extraction
and electrowinning circuits are relatively high so the use of the
hydrometallurgical process route on some smaller or short lifetime
resources can be uneconomical. Furthermore, the electrowinning step
is energy intensive and requires a significant source of electrical
power. As the cost of energy generation increases through increased
demand and taxation, the cost of the required electrical power
becomes even more prohibitive, especially in remote areas where the
required infrastructure is not already installed.
[0018] A less obvious but much more technically challenging issue
for hydrometallurgical processing is the huge dependence of the
process on the proton and sulphate balance. The fact that the
process regenerates sulphuric acid in the electrowinning section is
often purported as a major advantage but it can also become a major
problem in the situation where the leach circuit is also generating
acid. This is often the case in the leaching of sulphides and
results in the need to neutralise a portion of the acid generated.
The required bleed to a neutralisation step results in extra
reagent costs, introduces a potential avenue of valuable copper
loss and the resulting residue requires environmentally sound
storage and disposal. These issues with the hydrometallurgical
process has resulted in ore bodies which contain a copper oxide cap
over a more substantial sulphide deposit having the oxide cap
removed and discarded or stored rather than being processed and the
copper value realised.
[0019] As mentioned previously, it is of concern that the mean
grade of newly discovered copper ore deposits is decreasing. Most
of the world's copper reserves are now in the form of copper
sulphide ores. As the concentration of copper in these reserves
decreases and the concentrations of impurity elements increase the
ores and concentrates are becoming increasingly difficult to treat
using existing industrial process routes and technologies. The
impacts of these trends are in the form of decreasing productivity,
increasing energy consumption and costs, and increasing capital
investment and operating costs required to avoid adverse
environmental impacts.
[0020] It is more difficult to effectively extract copper from such
complex ores and concentrates. For example, these complex ores
often have relatively high levels of iron which must be removed at
some stage. This is typically performed in the smelter where the
iron is converted to iron oxide and becomes incorporated in the
slag. The greater the iron content of the copper ore or concentrate
then the greater the levels of flux and energy required to deal
with it. Importantly, a percentage of the available copper will
always partition into the slag with the iron and greater quantities
of slag resulting from the higher levels of iron will inevitably
result in a greater loss of copper to slag.
[0021] It is clear that, using conventional pyre- and
hydro-metallurgy routes, as mean ore grades decrease the capital
and operating costs of copper production will increase along with
the electrical energy requirements and also the greenhouse gas
impact, if the energy used is produced from fossil fuels. In
addition, arsenic is becoming even more of a problem for copper
processing as its concentration in many copper ores is increasing
while, at the other end, stricter limits are being placed on its
environmental release. Similar issues exist for other impurity
elements such as lead, bismuth and a range of radioactive elements.
Due to these multiple constraints, it is becoming increasingly
important to find new ways to efficiently treat lower grade copper
ores to make use of existing are bodies whilst minimising the
environmental impact of processing
OBJECT OF THE INVENTION
[0022] It is an aim of this invention to provide for a method of
processing copper-containing source materials which overcomes or
ameliorates one or more of the disadvantages or problems described
above, or which at least provides a useful alternative.
[0023] Other preferred objects of the present invention will become
apparent from the following description.
SUMMARY OF INVENTION
[0024] According to a first aspect of the invention, there is
provided a method of processing a copper-containing source material
including the steps of: [0025] (a) providing an aqueous acidic
leach solution of the copper source material; [0026] (b)
introducing a pH increasing agent into the acidic leach solution to
cause the precipitation of a copper-containing intermediate; and
[0027] (c) collecting the copper-containing intermediate and
exposing it to a high temperature treatment,
[0028] to thereby process the copper-containing source
material.
[0029] In certain embodiments of the first aspect, there is
provided a method of processing a copper sulphide-containing source
material including the steps of: [0030] (a-i) exposing the copper
sulphide-containing source material to an oxidative roast to form a
calcined copper-containing source material; [0031] (a-ii)
contacting the calcined copper containing source material with a
leach solution to form an acidic copper-containing leach solution;
[0032] (b) introducing a pH increasing agent into the acidic
copper-containing leach solution to cause the precipitation of a
copper-containing intermediate; and [0033] (c) collecting the
copper-containing intermediate and exposing it to a high
temperature treatment,
[0034] to thereby process the copper sulphide-containing source
material.
[0035] The various features and embodiments of the present
invention, referred to in individual sections above apply, as
appropriate, to other sections, mutatis mutandis. Consequently
features specified in one section may be combined with features
specified in other sections as appropriate.
[0036] Further features and advantages of the present invention
will become apparent from the following detailed description.
BRIEF DESCRIPTION OF THE DRAWINGS
[0037] In order that the invention may be readily understood and
put into practical effect, preferred embodiments will now be
described by way of example with reference to the accompanying
figures wherein:
[0038] FIG. 1 is a representation of the typical steps involved in
the pyrometallurgical processing of copper-containing ore;
[0039] FIG. 2 is a representation of the typical steps involved in
the hydrometallurgical processing of copper-containing ore;
[0040] FIG. 3 is a representation of the steps which may be
involved in the present method of processing of a copper source
material (some steps are optional);
[0041] FIG. 4 is a representation of the steps which may be
involved in the present method when processing a copper sulphide
source material and employing an oxidative roast step (some steps
are optional);
[0042] FIG. 5 is a graphical representation of the selective
precipitation of iron and copper from an aqueous solution as a
function of pH;
[0043] FIG. 6 is a graphical representation of the precipitation of
a copper-containing intermediate from aqueous solution upon
addition of lime/limestone; and
[0044] FIG. 7 is a graphical representation showing the thermal
decomposition of a copper-containing sample with thermogravimetric
and differential scanning calorimetry results shown.
DETAILED DESCRIPTION OF THE DRAWINGS
[0045] The present invention is predicated, at least in part, on
the finding that a copper concentrate can be precipitated from an
aqueous acidic leach solution which is suitable for feeding
directly into a smelting or converting step to thermally decompose
to provide a copper product. The use of lime and/or limestone has
been found to be particularly useful in precipitating a copper
concentrate which can be advantageously integrated into a
smelting/converting operation. Such an approach allows the
integration of early steps of the hydrometallurgical approach with
downstream steps in the pyrometallurgical approach to enable the
efficient processing of a wider range of copper ore types and
grades thereof. Additionally, it has been identified that copper
sulphide ores or concentrates can be advantageously converted to
forms which will provide an ideal solution from which the copper
concentrate can be precipitated by exposing the ore or concentrate
to an oxidative roast prior to leaching. Not only does this
roasting step maximise the amount of copper in a form optimally
suited for dissolution and subsequent precipitation but it also
allows for impurity elements, such as iron, to be converted to
forms which will not leach along with the copper thereby providing
a purification or concentration of the copper prior to its
precipitation. In addition, during oxidative roasting impurities
may be removed and separated from the concentrate by partitioning
into the gas phase as gaseous species or as fine particulates that
are removed in the off gas stream. Further details and advantages
of the present process are described herein.
[0046] In this patent specification, adjectives such as first and
second, left and right, front and back, top and bottom, etc., are
used solely to define one element or method step from another
element or method step without necessarily requiring a specific
relative position or sequence that is described by the
adjectives.
[0047] Unless defined otherwise, all technical and scientific terms
used herein have the same meaning as would be commonly understood
by those of ordinary skill in the art to which this invention
belongs.
[0048] In a first aspect of the invention, there is provided a
method of processing a copper-containing source material including
the steps of: [0049] (a) providing an aqueous acidic leach solution
of the copper source material; [0050] (b) introducing a pH
increasing agent into the acidic leach solution to cause the
precipitation of a copper-containing intermediate; and [0051] (c)
collecting the copper-containing intermediate and exposing it to a
high temperature treatment,
[0052] to thereby process the copper-containing source
material.
[0053] A representative flow sheet for one embodiment of the
present method is shown in FIG. 3. It will be appreciated that not
all of the steps shown are strictly required, rather FIG. 3
highlights how the present method can be used to integrate
pyrometallurgical and hydrometallurgical flow sheets to take
advantage of the main benefits of each approach.
[0054] Briefly, the input, in this case a copper ore, is exposed to
an acidic leach, in this case sulphuric acid, to provide a leach
solution which is then exposed to a slight and controlled pH
increase or precipitation step. In the flow sheet in FIG. 3
limestone is used as the precipitation or pH increasing agent. This
impurity precipitation step is an optional step, as indicated by
the hatched shading in the relevant box in FIG. 3, but provides
advantages in removal of significant portions of impurities such as
iron and arsenic from the leach solution prior to precipitation of
the actual copper-containing intermediate. When processing copper
sulphide and oxide ores the presence of significant quantities of
iron are inevitable and so this impurity precipitation step may be
desirable.
[0055] After removal of the impurity precipitate, additional pH
increasing agent, again limestone in the embodiment shown in FIG.
3, is added to raise the pH sufficiently to begin precipitation of
the copper-containing intermediate which is subsequently collected
and the remaining leach solution sent to tailings.
[0056] The copper-containing intermediate may then be subjected to
an optional physical separation step, again the fact of this step
being optional is indicated by hatched shading in FIG. 3. In the
flow sheet in FIG. 3, wherein sulphuric acid is the leaching agent
and limestone is the precipitation agent, gypsum
(CaSO.sub.4.2H.sub.2O) will inevitably precipitate with the
copper-containing intermediate. This actually presents certain
advantages which will be discussed further below but in some
instances the quantity of gypsum may be undesirably high such that
the energy requirements in the smelter/converter are increased
unnecessarily. In such instances a substantial portion of the
gypsum can be separated from the copper-containing intermediate
based on the larger size of the gypsum particles. As is shown in
FIG. 3 some of the gypsum solid can be recycled to the leach
solution prior to precipitation to potentially encourage the growth
of larger gypsum crystals and thereby maintain the effectiveness of
the separation step.
[0057] The precipitated copper-containing intermediate may then be
exposed to an optional heating step, indicated by hatched shading
in FIG. 3, largely designed to remove moisture and decompose
limestone and gypsum as well as certain of the entrained copper
compounds into forms more suited to the smelter/converter
operation. To be clear, the smelter/converter could also achieve
these aims but it may be more desirable, in terms of energy
requirements and, particularly, suitability of copper compounds to
be added to the converting stage, to employ this initial slightly
more moderate heating step.
[0058] The copper-containing intermediate, which may have been
altered in composition due to the heating step, can then be
introduced directly to the smelter or converter. It is preferred,
in many instances, that it be directly introduced to the converter
rather than the smelting step. It is a distinct advantage of the
present method that the copper-containing intermediate is suitable
for direct introduction into the converter without the need for
smelting since this enables the copper throughput or productivity
of the converter to be increased without adverse impact on the
preceding smelting operation. The copper-containing intermediate
contains oxygen bonded with the copper and the additional oxygen
thereby provided allows a reduction in the tonnage oxygen gas that
must be injected into the converter to attain the oxidising
conditions and so assists in reducing capital and operating costs
in oxygen production and increasing the potential copper production
rate from a particular converter operation. When lime or limestone
is used as the pH increasing agent then the accompanying gypsum
that is precipitated with the copper-containing intermediate, and
unreacted lime and limestone, acts as flux and so reduces the
required amount of calcium oxide that needs to be introduced into
the converter to control the slag chemistry.
[0059] Further, unlike in the smelting stage, the copper converter
stage typically operates with more oxidising conditions meaning
that arsenic is more readily dissolved and stabilised in the molten
slag. Incorporation of the arsenic in the stable slag phase then
avoids the problem of arsenic release into the environment, which
is taken advantage of by the present process. Finally, converters
tend to produce excess heat. The exothermic reaction of these
phases with the copper matte in the converter may be balanced with
the enthalpy requirements for heating and decomposition of the
copper-containing intermediate compounds, thereby utilising the
excess heat to boost copper production rates.
[0060] The converter has received matte in the usual manner from
the smelter and so the copper-containing intermediate is simply
introduced to supplement this material and the two are processed in
the converter together. The matte has been produced by the normal
steps in the pyrometallurgical processing route which are shown in
FIG. 3 for the sake of clarity. The capture of sulphur gases and
their use in the production of sulphuric acid is shown. It is a
further advantage of the present method that the integration of the
hydrometallurgical and pyrometallurgical processes allows the
sulphuric acid produced on site to be fed back into the leaching
stage.
[0061] The materials and requirements in each step of the present
method will now be addressed in more detail.
[0062] The copper source material may be selected from the group
consisting of copper-containing ore, copper smelter slag,
copper-containing tailings or tailings sediment, such as those from
a copper concentrator, a copper-containing process intermediate,
waste water, galvanic waste, copper containing acidic leach
solution and waste product from another process.
[0063] The copper-containing ores may be selected from the group
consisting of copper oxide and sulphide ores, copper-gold deposits
and mixed metal deposits which may also include elements such as
nickel, cobalt, zinc and manganese.
[0064] When nickel, cobalt, zinc or manganese are present in the
source material they are advantageously dealt with by the present
method through the precipitation step. Particularly, they will
precipitate at a higher pH than the copper and so, after the copper
precipitation is complete, the remaining leach solution containing
these metals would be suitable for further treatment and,
potentially, realisation of the value of these metals. This could
be important for processing of a polymetallic waste stream or
deposit.
[0065] Preferably, the copper-containing ore is a copper sulphide
or copper oxide ore.
[0066] It is an advantage of the present method that such a wide
range of copper-containing materials can be used as inputs. The
limitations of hydrometallurgical and pyrometallurgical approaches
in this regard are significant and have already been discussed. As
a major cost of the pyrometallurgical approach is the energy
consumption in grinding of the ores the present method allows the
more straightforward dissolution of the copper compounds from the
source material via acid leaching. The key precipitation step then
provides a suitable input for the pyrometallurgical
smelting/converting operations to thereby avoid the need for a
relatively expensive electrowinning approach to copper
recovery.
[0067] The method may include the step of exposing the copper
source material to an acid to form the aqueous acidic leach
solution. That is, the method may include the actual leaching of
the copper source material to form the acidic leach solution. This
may not be necessary in every case as the source material may
already be an obtained acidic leach input or a waste or recycled
solution of copper.
[0068] Preferably, the acidic leach solution has a pH of less than
about 4.0, more preferably less than about 3.0, even more
preferably less than about 2.0 and yet still more preferably
between about pH 0.0 to 2.0. Typical sulphuric acid leach solutions
currently used in hydrometallurgical processing have pH values of
less than 1.0. Such existing acidic leach solutions would be
suitable for use in the present method although values in the pH
range of 3 to 4 are preferred in practice. The final pH of the
leach solution will depend upon how difficult it is to get all of
the copper in solution. If this is challenging or if it is
acceptable to co-dissolve iron and other metal impurities then
lower pH values will present. If the leach is to be selective for
copper over iron then the pH will be higher, for example between pH
2.0 to 4.0.
[0069] Suitably, the acidic leach solution is formed using an acid
selected from the group consisting of hydrochloric, nitric and
sulphuric acids. Sulphuric acid may be preferred to produce a
solution of dissolved copper sulphates which are particularly
suited to subsequent precipitation and use in the heat treatment
stage of the present method as the sulphate can be used to
regenerate sulphuric acid.
[0070] The pH increasing agent, which may also be referred to as a
precipitating agent, may be any basic compound or any material
containing such a compound.
[0071] The pH increasing agent may be an alkali metal or alkali
earth metal carbonate, oxide, hydroxide or compounds or
associations thereof.
[0072] The pH increasing agent may be selected from the group
consisting of calcium oxide, calcium carbonate, calcium hydroxide,
calcium ferrite, magnesium oxide (magnesia), magnesium carbonate,
magnesium hydroxide, sodium carbonate, sodium hydroxide, dolomite
(CaMg(CO.sub.3).sub.2) and other minerals containing one or more of
these compounds.
[0073] Preferably, the pH increasing agent is lime (calcium oxide)
and/or limestone (calcium carbonate).
[0074] Lime and limestone are advantageously low cost pH increasing
agents which also have the added benefit of removing most of the
sulphate from the tailings solution, assuming a sulphuric acid
leach. It may also be possible to employ calcium containing smelter
or converter slags as the precipitation reagent in the impurity or
copper precipitation stages as the calcium content of the slag will
be in the form of calcium oxide which should be able to react to
raise the pH of the leach solution. The slag will contain iron and
so considerations should be given to the quantity of this metal
which is being introduced as an impurity, especially in the copper
precipitation stage.
[0075] In one embodiment, the pH increasing agent is not sodium
hydroxide. Although in some instances sodium hydroxide may be
appropriate it has the disadvantage of having a cost per mole of
neutralising value of approximately three to ten times greater than
the equivalent molar neutralising value of lime or limestone.
Importantly, the present inventors postulate that the precipitation
of copper with lime or limestone may result in a higher achievable
recovery of copper than can be attained with sodium hydroxide. The
key factor that differentiates the precipitation of copper with
limestone versus sodium hydroxide is the rate of the reaction. The
reaction of copper with sodium hydroxide occurs very quickly with
all copper removed from solution within minutes whereas the
reaction with limestone takes at least an order of magnitude
longer. Whereas sodium hydroxide would be completely dissolved in
water, or would dissolve rapidly if introduced in solid form, the
dissolution rate of limestone is significantly slower and this is
believed to assist in controlling the rate of the copper
precipitation. The slower dissolution rate of the limestone is
believed to result in a more crystalline copper product.
[0076] A molar ratio of between about 0.5:1 to about 5:1 of pH
increasing agent to copper in solution may be required to
precipitate out substantially all of the copper as the copper
containing-intermediate. Preferably, the ratio of pH increasing
agent to copper in solution is between about 0.55:1 to about 3:1,
more preferably between about 0.60:1 to about 2:1, even more
preferably between about 0.65:1 to about 1.5:1.
[0077] In certain embodiments, the ratio of pH increasing agent to
copper in solution is between about 0.50:1 to about 3:1, more
preferably between about 0.50:1 to about 2:1, even more preferably
between about 0.50:1 to about 1.5:1, inclusive of between about
between about 0.55:1 to about 1:1.
[0078] In certain embodiments, the ratio of pH increasing agent to
copper in solution is between about 0.60:1 to about 3:1, more
preferably between about 0.60:1 to about 21, even more preferably
between about 0.60:1 to about 1.5:1, inclusive of between about
between about 0.65:1 to about 1:1.
[0079] In certain embodiments, the ratio of pH increasing agent to
copper in solution is between about 0.70:1 to about 3:1, more
preferably between about 0.70:1 to about 2:1, even more preferably
between about 0.70:1 to about 1.5:1, inclusive of between about
between about 0.75:1 to about 1:1.
[0080] The pH of the acidic leach solution of the copper source
material will be raised by the pH increasing agent to greater than
about pH 4.0. It has been found experimentally that significant
amounts of copper can be precipitated from the leach solution at pH
values above 4.0. Higher values will precipitate further copper up
to a point after which increases in pH provide no additional gains.
Preferably, the pH is increased to between about 4.0 to about 10.0,
more preferably between about 4.0 to about 9.0, even more
preferably between about 4.0 to about 8.0, still yet more
preferably between about 4.0 to about 7.0.
[0081] Increasing the leach solution pH means there is a higher
concentration of hydroxide ions in solution. The increased
hydroxide concentration provides conditions where copper oxide,
hydroxide and hydroxy compounds are stable causing them to
precipitate out to give the copper-containing intermediate. The
exact chemical form of the intermediate copper solid will depend on
solution conditions, for example, if the solution is high in
sulphate a basic copper sulphate may form, if the solution is high
in carbonate then a basic copper carbonate may form. In each of
these cases the main mechanism of reaction is the pH
adjustment.
[0082] Examples of the copper solid that could be produced by pH
adjustment, depending on the various solution phase parameters, are
copper oxide (CuO), copper hydroxide (Cu(OH).sub.2); basic copper
sulphates such as, or analogous, to Brochantite
(CuSO.sub.4.3Cu(OH).sub.2), Posnjakite
(CuSO.sub.4.3Cu(OH).sub.2.H2O), Wroewolfeite or Langite
(CuSO.sub.4.3Cu(OH).sub.2.2H.sub.2O), Vernadskite
(CuSO.sub.4.3Cu(OH).sub.2.4H.sub.2O), Antlerite
(CuSO.sub.4.2Cu(OH).sub.2), Antlerite hydrate
(CuSO.sub.4.2Cu(OH).sub.2.H.sub.2O), Antlerite dihydrate or
Kamarezite (CuSO.sub.4.2Cu(OH).sub.2.2H.sub.2O) and Dolerophanite
(CuSO.sub.4.CuO); basic copper carbonates such as Azurite
(2CuCO.sub.3.Cu(OH).sub.2) and Malachite (CuCO.sub.3.Cu(OH).sub.2).
From a nitrate solution basic copper nitrate
(Cu.sub.2(NO.sub.3)(OH).sub.3) may be precipitated. The basic
copper nitrate has a few mineral names including Gerhardtite and
Rouaite. From a chloride solution the copper
containing-intermediate precipitated out may be a basic copper
chloride (Cu.sub.2Cl(OH).sub.3), Belloite (CuClOH) or cuprite
(Cu.sub.2O). Basic copper chloride has been assigned a number of
mineral names including Atacamite, Paratacamite, Botallackite or
Clinoatacamite. It may also be possible to precipitate the copper
as a copper carbonate (CuCO.sub.3).
[0083] In one embodiment, the precipitated copper-containing
intermediate may comprise a compound selected from the group
consisting of copper oxides, hydroxides, sulphates, nitrates,
chlorides and carbonates or compounds containing a combination of
these, That is, the precipitated copper-containing intermediate may
comprise, for example, copper hydroxide or copper sulphate or mixed
compounds such as a basic copper sulphate or basic copper
carbonate.
[0084] The high temperature treatment may occur in a smelter,
direct-to-copper smelter or converter. As discussed previously
there are distinct advantages in producing a copper-containing
intermediate which can be introduced directly into the converter.
For example, direct introduction of the copper-containing
intermediate to the converter means available converter capacity is
taken advantage of, it is a distinctly down stream step of the
pyrometallurgical approach thereby minimising energy requirements
and the converter requirements for the addition of both oxygen and
calcium oxide flux can be reduced due to the nature of the
copper-containing intermediate produced and the accompanying
precipitated solids.
[0085] In one embodiment, the high temperature treatment occurs at
a temperature of at least 200.degree. C., preferably at least
300.degree. C., more preferably at least 1000.degree. C. This
includes temperature treatments of between about 200.degree. C. to
about 1500.degree. C., about 300.degree. C. to about 1500.degree.
C., about 400.degree. C. to about 1500.degree. C. about 500.degree.
C. to about 1500.degree. C., about 600.degree. C. to about
1500.degree. C., about 700.degree. C. to about 1500.degree. C.,
about 800.degree. C. to about 1500.degree. C., about 900.degree. C.
to about 1500.degree. C.,
[0086] In certain embodiments, the high temperature treatment is at
a temperature of between about 1000.degree. C. to about
1500.degree. C., such as may be achieved in a converter, preferably
about 1000.degree. C. to about 1400.degree. C., more preferably
between about 1200.degree. C. to about 1400.degree. C. These broad
ranges are inclusive of sub-ranges of about 1100.degree. C. to
about 1500.degree. C., about 1100.degree. C. to about 1400.degree.
C., about 1100.degree. C. to about 1350.degree. C., about
1200.degree. C. to about 1500.degree. C., about 1200.degree. C. to
about 1450.degree. C. and about 1250.degree. C. to about
1400.degree. C. Temperatures of about 1250.degree. C. 1300.degree.
C. and 1350.degree. C. may be preferred.
[0087] It will be appreciated that there may be a range of
temperatures at which the smelter or converter is operated
depending on the nature of the input materials but the ranges
presented above will generally result in the desired impure solid
or liquid copper, for example blister copper. Values of greater
than 1500.degree. C. and up to 1800.degree. C. may be acceptable
but are not generally required.
[0088] Both the smelter and converter operations may have
introduced oxygen but the converter generally operates under
relatively more strongly oxidising conditions than the smelter.
[0089] If the high temperature treatment is not performed in a
converter then it may be preferred to maintain a reduced oxygen
content in the heating environment.
[0090] If the high temperature treatment is performed in a
converter then it may be preferred to introduce an
oxygen-containing gas and a sulphide compound into the heating
environment.
[0091] In one embodiment, the method includes the step, prior to
step (b), of introducing an effective amount of a pH increasing
agent into the acidic leach solution to cause the precipitation of
an iron-containing compound prior to the substantial precipitation
of the copper containing-intermediate. The pH increasing agent may
be as described for the copper precipitation.
[0092] As discussed, copper sulphide and oxide ores contain
significant amounts of iron-containing compounds. These are
typically dealt with in the pyrometallurgical approach by the
introduction of flux and heating to separate out the iron in a slag
oxide phase. It is an advantage of the present process that a large
proportion of the iron impurity can be removed from the acidic
solution by a simple, and relatively selective over copper,
precipitation step. This lowers the amount of iron oxides which
need to be dealt with in a more energy intensive fashion in the
smelter and converter stages.
[0093] In some embodiments, arsenic can also be removed with the
iron in this precipitation step.
[0094] The step of precipitating the iron-containing compound may
involve increasing the pH of the leach solution to be between about
1.5 to about 4.0, preferably between about 1.5 to about 3.5, for
example 1.5 to 2.5. These ranges are inclusive of between about 2.0
to about 4.0, preferably between about 2.0 to about 3.0, for
example about 2.0 to about 2.5.
[0095] The method may include the step of collecting the
iron-containing precipitate to separate it from the processing
stream prior to further increasing the pH of the leach solution to
be between about pH 3.0 to about 10.0, preferably about 4.0 to
about 10.0, more preferably between about 4.0 to about 9.0, even
more preferably between about 4.0 to about 8.0, still yet more
preferably between about 4.0 to about 7.0, to thereby precipitate
out the copper-containing intermediate. These ranges are inclusive
of about 4.5 to about 10.0, preferably between about 4.5 to about
9.0, more preferably between about 4.5 to about 7.0.
[0096] The method may further include the step of exposing the
copper-containing intermediate to a separation step. This step is
optional but may serve to remove unwanted precipitated minerals
which may otherwise become an energy expense in the later smelter
or converter operation. The precipitation conditions for the
unwanted mineral precipitate may be controlled so as to encourage
particle growth to form larger particles or possibly crystals which
are more easily separated from the smaller copper-containing
intermediate particles. The leach solution may be seeded with
crystals of the mineral to encourage such growth.
[0097] In one embodiment, the separation step is a physical
separation based upon particle size or settling out of particles
i.e. mass and/or density differences.
[0098] Suitably, the separation may be effected by size screening
or sieving, hydrocycloning and the like.
[0099] In an embodiment wherein the pH increasing agent is lime
and/or limestone then the separation step will include the
separation of precipitated gypsum from the copper-containing
intermediate. Unreacted lime or limestone may also be separated
from the copper-containing intermediate in this step. It may not be
necessary or even desirable to remove all of the gypsum precipitate
and unreacted pH increasing agent. In fact unreacted lime or
limestone and gypsum in the precipitated copper-containing
intermediate may react in a heat treatment step to give a further
source of lime. This will be beneficial in converters that are
using a slag that includes lime as it will decrease the fluxing
requirements.
[0100] The method may further include the step of heating the
copper-containing intermediate to a temperature between 25.degree.
C. to less than 1000.degree. C. or between 200.degree. C. to
800.degree. C., including 25.degree. C. to 250.degree. C. or
25.degree. C. to 200.degree. C., prior to its exposure to higher
temperatures.
[0101] Depending on the temperature used in this temperature
treatment step, the heating will initially evaporate any associated
moisture in the copper-containing intermediate. It will then, at
higher temperatures, begin to decompose copper hydroxide and
sulphate portions and eventually leave a copper oxide. If heated
further the copper oxide will decompose to eventually form copper
metal although this is achieved in the more high temperature
converter step.
[0102] By way of example, during this temperature treatment step
moisture associated with the copper-containing intermediate could
be driven off by exposing the precipitated solid to a dry gas,
whether it is cold, warm or hot or just placing it in a heated
environment. In one embodiment this could be achieved by exposing
the copper-containing intermediate to a warm dry gas (such as
smelter off gas at 25-200.degree. C.) to remove a proportion of the
moisture relatively quickly. Subsequently, or alternatively, the
copper-containing intermediate could be exposed to hot gases (such
as smelter off gas at >200.degree. C.) which would also cause
the solid to decompose to copper oxide (as described in tables 4
and 5), or possibly even as far as to copper metal.
[0103] At any stage during this decomposition in the moderate
heating step it would be possible to stop externally heating the
solid and instead add it to the converter where it will continue to
decompose to eventually form copper metal.
[0104] It will therefore be appreciated that to decompose the
precipitated copper-containing intermediate it will be necessary to
expose it to a high temperature treatment of at least 200.degree.
C. However, as discussed above, it may be advantageous to first
expose the copper-containing intermediate to a lower temperature
heating step which will remove associated moisture but will not
result, to a significant degree, in decomposition to a preferred
end product, such as copper oxide. Higher temperature treatments,
for example between 200.degree. C. to 800.degree. C. or so, will
result in decomposition of the copper-containing intermediate
towards compounds, such as copper oxide, as described in table 5.
However, introduction to a converter to reach temperatures of
around 1200.degree. C. to 1300.degree. C., for example, will result
in decomposition of the copper-containing intermediate to give a
preferred decomposed copper end product, such as an impure copper
metal solid or liquid including liquid blister copper. It will be
appreciated that all heat treatments which result in decomposition
of the copper-containing intermediate are considered to be within
the scope of the high temperature treatment of step (c) of the
present method while those high temperature treatments, such as
would be experienced in a converter, which result in a copper metal
comprising product in solid or liquid form are preferred.
[0105] The method of the present invention provides certain
additional advantages when processing a copper sulphide-containing
source material. Copper sulphide minerals are not readily dissolved
in sulphuric acid at room temperature. The sulphur in the copper
sulphide minerals must be oxidized for the copper to be extracted
by leaching into solution. This solid-liquid oxidation reaction is
slow and can be complicated by the formation of a passive
sulphur-rich layer on the surface of the particles that limits the
rate and extent of the reaction. Fast and efficient leaching of
copper sulphide minerals requires some combination of fine
grinding, elevated temperatures and pressures, surfactants, the
presence of chlorides, catalytic bacteria or minerals. The leaching
behaviour of copper sulphides also varies significantly depending
on the specific type of copper mineral present. Thus, while the
process described in FIG. 3 is applicable to copper sulphide ores
it can be challenging to obtain the initial acidic leach solution
ready for precipitation of a copper-containing intermediate via a
pH increase. This can be overcome by exposing the copper sulphide
source material, or a concentrate thereof, to an oxidative
pyrometallurgical roast step.
[0106] Therefore in one embodiment, wherein the copper source
material is a copper sulphide-containing source material, the
method may further include the step (a-i) of exposing the copper
sulphide-containing source material to an oxidative roast, prior to
its exposure to the leach solution, to form a calcined
copper-containing source material.
[0107] The oxidative roast causes the conversion, at suitable
temperature, of copper and iron sulphides, and other sulphides
depending on the content of the ore, to sulphates and oxides.
Particularly, if the conditions for the roast are appropriate then
the vast majority of the iron sulphides can be converted to oxides
while the copper sulphides can be converted to sulphate and/or
oxide forms. In forming copper sulphate, the copper sulphide may
react with oxygen to form copper oxide and SO.sub.2 or SO.sub.3
gas. The SO.sub.2 or SO.sub.3 gas in turn can react with the copper
oxide to form copper sulphate. The chemical thermodynamic stability
of copper and iron compounds differ from each other making it
possible to prepare different combinations of compounds dependent
on process conditions within the roasting reactor. For example,
partial oxidation of the sulphide compounds may result in the
formation of iron sulphate. In other process conditions iron can
form the compounds Fe.sub.3O.sub.4 and Fe.sub.2O.sub.3. The
formation of these copper-free iron oxide compounds is extremely
advantageous as, while copper sulphate is water soluble and copper
oxide can be dissolved in mild acid, iron oxides require stronger
acid to dissolve them. This allows for a selective leach step where
the copper compounds can substantially all be dissolved in a mildly
acidic solution while the iron compounds are left behind in the
solid leach residue thereby simplifying the removal of an otherwise
challenging impurity.
[0108] The oxidative roast may occur at a temperature of from about
500.degree. C. to about 950.degree. C. This range is inclusive of
the oxidative roast being carried out at a temperature of from
500.degree. C. to 900.degree. C., 500.degree. C. to 850.degree. C.,
500.degree. C. to 800.degree. C., 500.degree. C. to 750.degree. C.,
550.degree. C. to 950.degree. C., 550.degree. C. to 900.degree. C.,
550.degree. C. to 850.degree. C., 550.degree. C. to 800.degree. C.,
550.degree. C. to 750.degree. C., 600.degree. C. to 950.degree. C.,
600.degree. C. to 900.degree. C., 600.degree. C. to 850.degree. C.,
600.degree. C. to 800.degree. C., 600.degree. C. to 750.degree. C.,
650.degree. C. to 950.degree. C., 650.degree. C. to 900.degree. C.,
650.degree. C. to 850.degree. C., 650.degree. C. to 800.degree. C.,
650.degree. C. to 750.degree. C., 700.degree. C. to 950.degree. C.,
700.degree. C. to 900.degree. C., 700.degree. C. to 850.degree. C.
and 700.degree. C. to 800.degree. C.
[0109] It will be appreciated, however, by those skilled in the art
of roasting and copper processing generally that the optimal
roasting conditions are a combination of factors including
temperature, partial pressure of sulphur dioxide and partial
pressure of oxygen. The equilibrium relationships between these
factors have been studied and predominance diagrams linking the
variables are available and known to those working in this field to
thereby indicate what particular temperature is most appropriate
for roasting depending on the gas pressures present.
[0110] In reviewing such predominance diagrams and deciding on the
roasting conditions it is optimal to aim for conditions which will
ensure the vast majority of the iron sulphides are converted to
oxides which do not contain copper while the copper sulphides are
converted to copper sulphate and/or copper oxide. It is desirable,
although not essential, to maximise the amount of copper sulphate
formed relative to the amount of copper oxide as copper sulphate is
highly soluble and is ideal for subsequent precipitation by pH
increase.
[0111] The oxidative roast may be performed in the presence of air,
oxygen-enriched air or other suitable oxygen-containing gas. So
long as a suitable amount of oxygen is available for the conversion
of the sulphides then any gaseous atmosphere may be appropriate.
The roast may be performed using equipment, for example a fluidised
bed roaster, which is currently available and known in the
field.
[0112] The method may further include the step (a-ii) of contacting
the calcined copper-containing source material with a leach
solution. This will provide the acidic leach solution which will
subsequently be exposed to a pH increase, as already described, to
provide a copper-containing intermediate.
[0113] The leach solution is an aqueous leach solution which may be
acidic or neutral. The pH of the leach solution to which the
calcined copper-containing source material is exposed is preferably
mildly acidic however, if substantially all of the copper sulphide
in the source material has been converted to copper sulphate rather
than copper oxide during the roasting process then water is all
that will be required to dissolve the copper. In most instances it
would be expected that at least some amount of copper oxide will be
present in the calcined copper-containing source material and so a
mildly acidic leach solution would be preferred. Note that even in
the instance wherein the leach solution is just water, i.e. pH
approximately 7, due to all copper being in the sulphate form, an
acidic leach solution containing the copper would still be formed
due to the freeing of the sulphate anion on dissolution resulting
in a pH drop for the solution. For example, the solution may drop
from pH 7 to about pH 4 to 5 upon dissolution of a substantial
amount of copper sulphate.
[0114] Therefore, in one embodiment, the leach solution which is to
be used to contact the calcined copper-containing source material
may have a pH of from about 2.0 to about 7.0 inclusive of 2.0 to
6.5, 2.0 to 6.0, 2.0 to 5.5, 2.0 to 5.0, 2.0 to 4.5, 2.0 to 4.0,
2.5 to 6.5, 2.5 to 6.0, 2.5 to 5.5, 2.5 to 5.0, 2.5 to 4.5 and 2.5
to 4.0.
[0115] Preferably, the leach solution is an acidic leach solution.
The nature of the acid may be as previously described.
[0116] The leach solution after substantial dissolution of the
calcined copper-containing source material will have a pH of
between about 2 to about 5, preferably about 2.0 to about 4.5, for
example 2.0 to 4.0.
[0117] As discussed previously, at this pH all of the copper
sulphate and oxide within the calcined copper-containing source
material will be dissolved but the iron oxide will require more
strongly acidic conditions to dissolve and so remains in the solid
state. This provides for a relatively dense and compact iron
residue which can be easily removed and disposed of to landfill, if
necessary. If this approach is employed then the optional pH
increase of the acidic leach solution to precipitate out iron,
discussed in relation to FIG. 3 as an optional step, would not be
necessary.
[0118] When an oxidative roast is used the remaining steps, such as
the precipitation of a copper-containing intermediate and ensuing
steps already described, both optional and otherwise, are as
already discussed.
[0119] Therefore, in certain embodiments of the first aspect, there
is provided a method of processing a copper sulphide-containing
source material including the steps of: [0120] (a-i) exposing the
copper sulphide-containing source material to an oxidative roast to
form a calcined copper-containing source material; [0121] (a-ii)
contacting the calcined copper containing source material with a
leach solution to form an acidic copper-containing leach solution;
[0122] (b) introducing a pH increasing agent into the acidic
copper-containing leach solution to cause the precipitation of a
copper-containing intermediate; and [0123] (c) collecting the
copper-containing intermediate and exposing it to a high
temperature treatment, to thereby process the copper
sulphide-containing source material.
[0124] This process is set out in a representative flow sheet as
shown in FIG. 4. Again, it will be appreciated that not all of the
steps shown are strictly required and the flow sheet is exemplary
only. The flow sheet of FIG. 4 is substantially identical to that
in FIG. 3 except for it being limited to the processing of a copper
sulphide source material and including roasting and pre leach steps
prior to formation of the acidic leach solution containing the
copper compounds. It can also be seen that, as for the process
exemplified in FIG. 3, the embodiment set out in FIG. 4 allows for
integration of hydro- and pyrometallurgical pathways.
[0125] The embodiment of FIG. 4 provides for considerable
advantages in operation. Particularly, the success and flexibility
of the manner in which impurities can be dealt with is extremely
beneficial. The easy removal of iron via the roast and selective
leach has been discussed. However, complex copper sulphide ores
contain significant amounts of lead, arsenic, bismuth as well as
uranium and other radioactive metals. Lead and bismuth will be
dealt with in much the same manner as the iron in that the
compounds, be it oxides or sulphates, produced by the roasting
process are either not soluble in aqueous acid at all or are less
soluble than the equivalent copper compounds and so will not be
dissolved in the mild acidic leach. Dealing with these compounds at
this stage is much simpler and more cost effective than doing so in
a smelting operation. Elements such as nickel and cobalt, behaving
in a chemically similar manner to copper, follow the route taken by
the copper species through the process and so can also be recovered
from the original sulphide ore or source material.
[0126] The presence of radioactive metals can make an ore or
concentrate challenging to deal with. It is common to carry out
some initial concentrating of ores at or near the mine site before
they are transported to a central smelting plant. However, if the
amount of radioactive materials is above certain levels then, due
to government transport regulations, they cannot be transported and
so must be further processed on site. Radioactive metals forming
insoluble or sparingly soluble oxides during the roast will remain
with the leach residue along with the iron. For those radioactive
metals, such as uranium, which are found in copper sulphide ores in
an acid soluble form they can be removed by an acidic or alkaline
`pre-leach` prior to roasting. This allows for their easy removal
with minimal loss of copper which, at that point, is still in the
sulphide form which is not highly acid soluble.
[0127] Further, during the oxidative roast the arsenic present may
be converted into gaseous species such as arsenic sulphide or
arsenic oxide. During oxidative roasting fine particulate material
containing impurity elements may also be formed and these may
become entrained in the gas stream. Roasters have gas extraction
and collection systems that enable the gas and fine particulates to
be separated from the solid concentrates. During typical
pyrometallurgical processing the ideal scenario is that, in the
smelter, arsenic partitions to the slag and, due to the nature of
slag, it remains there in an environmentally stable form.
Unfortunately, the practical reality is different and a significant
quantity of the arsenic goes into the matte (molten sulphide). The
matte is passed on to the converter where the arsenic partitions
between the gas phase, the slag and the copper metal. Due to the
converting conditions, more arsenic is typically to be found in the
converter slag than smelter slag but converter slag will be
typically recycled back to the smelter. The end result of this is
that, if there is more arsenic in the feed material, then more will
find its way into the copper metal. Since there are practical
limits to the amount of arsenic which can be allowed in anode
copper the amount of arsenic that can be introduced in the feed to
the process is also limited. As discussed, due to the conditions in
the roasting step employed in certain embodiments of the present
invention a much greater proportion of arsenic originally in the
concentrate will partition to the gas phase forming gaseous species
or fine particulates. This means the majority of the arsenic can be
treated in the off gas stream from the roaster with the advantage
that very little or no arsenic will actually follow the treated
copper concentrate that is subsequently sent to the copper
converter thereby reducing the arsenic in the blister copper metal
product.
[0128] The roasting off gas will also contain sulphur dioxide
and/or sulphur trioxide which can advantageously be used to
generate sulphuric acid for use in the leaching stage. It may also
be desirable to use excess heat from the roaster to enhance the
leaching step.
[0129] Finally, the present approach is favourable in terms of the
energy demands for processing of the source material. Since iron,
and a range of other impurities, can be removed early in the flow
sheet with a simple leaching step energy is not required in the
smelting operation to address them and a more pure copper
concentrate is being treated pyrometallurgically. If the
copper-containing intermediate is to be transported for
smelting/converting operations then further energy savings are made
in terms of the quantity of material being transported. The
sulphide oxidation reaction in the roasting step is highly
exothermic and with appropriate control of process parameters the
roasting step is auto-thermal or requires reduced fuel input
compared to a non reactive roast. Thus a number of significant
advantages are obtained for a relatively low energy input.
[0130] Finally, the present approach is favourable in terms of the
energy demands for processing of the source material. Since iron,
and a range of other impurities, can be removed early in the flow
sheet with a simple leaching step energy is not required in the
smelting operation to address them and a more pure copper
concentrate is being treated pyrometallurgically. If the
copper-containing intermediate is to be transported for
smelting/converting operations then further energy savings are made
in terms of the quantity of material being transported. While the
roasting step does itself require some energy to kick start the
reactions it is then more or less auto-thermal as the sulphide
oxidation reaction is highly exothermic. Thus a number of
significant advantages are obtained for a relatively low energy
input.
[0131] It has not been previously appreciated in the art that
roasting of a copper sulphide ore or concentrate would provide for
a copper-containing leach solution which would contain suitable
copper compounds (Le, oxide and sulphate species) for a selective
precipitation operation by a simple pH increase. That is, the
oxidative roast provides for an ideal copper-containing leach
solution for easy and effective subsequent selective precipitation
of a copper concentrate by the addition of a pH increasing agent,
as described previously.
[0132] It will be appreciated that the various options provided for
by the roast-leach-precipitate-heat treat approach together provide
for a previously unavailable level of flexibility to deal with
impurities including at the (i) roasting, (ii) selective mild acid
leach; (iii) selective precipitation; and (iv) smelting/converting
stages. Given the increasing complexity of ores it is crucial to
have this level of adaptability as well as provide for integration
with existing smelting/converter operations.
[0133] This particular embodiment may further include the step
(aa-i) of contacting the copper sulphide-containing source
material, prior to the oxidative roast, with an acidic or alkaline
solution to leach out certain impurities. Uranium, for example, may
be dissolved in its oxide form into either acid or alkaline
solutions. This pre-leach step should be performed under
non-oxidising conditions so as to not convert the copper sulphides
present into potentially soluble sulphate or oxide forms.
[0134] After step (aa-i) the copper sulphide-containing source
material can be separated from the acidic or alkaline
impurity-containing leach solution. It will then be ready for
introduction to the roasting step.
[0135] The copper sulphide-containing source material may be a
copper sulphide-containing ore, copper sulphide-containing
concentrate or copper sulphide tailings. If the material is a
concentrate then it may be obtained in the usual way, for example,
by grinding and flotation operations.
[0136] A second aspect of the invention results in a concentrated
copper product when produced by the method of the first aspect.
[0137] The concentrated copper product may be substantially pure
copper metal.
EXPERIMENTAL
Oxidative Roast
[0138] A copper concentrate was produced from a copper sulphide ore
using, a lab scale flotation cell. The concentrate contained mostly
chalcopyrite with a small amount of silica and pyrite. Separate
samples of the concentrate were heated to 600.degree. C.,
750.degree. C. and 900.degree. C. in a tube furnace. An atmosphere
of sulphur dioxide and air was enforced. The ratio of the sulphur
dioxide and to air and therefore oxygen in the furnace was
controlled by adjusting, the flow rate of these gases into the
furnace. A single flow rate set point for each gas was chosen based
on Factsage modelling. All three experiments were carried out at
the same flow rates and therefore the same atmospheric conditions.
The flow conditions were 400 mL-air/min and 25 mL-SO.sub.2/min
which equates to an enforced atmosphere of about 0.06 atm SO.sub.2,
0.94 atm air which is equivalent to 0.20 atm O.sub.2. All solids
formed were characterised by Powder XRD. All solids contained some
silica.
TABLE-US-00001 TABLE 1 Products of oxidative roast Temperature
Species Identified in Calcined Solid 600.degree. C. Fe.sub.2O.sub.3
CuSO.sub.4 750.degree. C. Fe.sub.2O.sub.3 CuSO.sub.4 Small amounts
of CuO.cndot.CuSO.sub.4 and CuO.cndot.Fe.sub.2O.sub.3 900.degree.
C. Fe.sub.2O.sub.3 CuO.cndot.Fe.sub.2O.sub.3
[0139] At 600.degree. C. it is apparent that the solids are mostly
iron oxide and copper sulphate which represents a solid well suited
to leaching, as previously described. At 750.degree. C., again,
iron oxide and copper sulphate make up the majority of the calcined
solid along with a small amount of copper oxide. Finally, at
900.degree. C. the majority of the copper is copper oxide and iron
is also in the oxide form. As has been discussed, the present
method can be used to selectively leach copper from a copper
sulphate and/or copper oxide-containing calcined solid and so all
three temperatures tested for the roast have proved useful.
Iron Precipitation
[0140] A batch experiment was carried out to show the potential for
selectively precipitating the majority of iron from the leach
solution without removing significant amounts of copper. In this
experiment an initial solution containing approximately 6.5 g-Fe/L
as ferric sulphate and 3.3 g-Cu/L as cupric sulphate was prepared.
Solid limestone was dosed into the reactor every 30 minutes.
Solution samples were taken one minute prior to the next dose of
limestone. The results are indicated in FIG. 5.
[0141] The results show that iron began to precipitate at some
point above pH 1.6 with effectively all of it removed from solution
by pH 3.4. Meanwhile the copper began to precipitate from solution
at some point between pH 2.3 and 3.4. This experiment shows that
the majority of ferric iron can be selectively precipitated from a
leach solution containing both ferric and cupric sulphates. By
precipitating at a pH between about 2.3 and 3.4, more than 90% of
the ferric can be removed from solution with less than 10% copper
loss. The graph in FIG. 5 indicates that a pH of just under 3.0 may
be optimal in providing close to 95% ferric iron removal with
minimal copper precipitation.
[0142] These results show that a pre-treatment of the pregnant
leach solution, when employing the present inventive method, to
raise the pH to a predetermined level can result in selective
removal of the ferric iron by precipitation. After removal of the
iron a further pH increase will result in precipitation of a much
purer copper-containing intermediate which can be carried on to the
heat treatment step (smelting or converting).
Copper Precipitation Experiments
[0143] Lime or limestone solids were slaked in 200 mL of either
de-ionised water or gypsum saturated water for 20 minutes before
800 mL of a copper sulphate containing solution (the synthetic
leach solution) was added. The addition of the copper sulphate
containing solution to the alkaline slurry resulted in a reaction
whereby the lime or limestone dissolved and the copper
precipitated. The resulting solution pH was measured throughout and
samples of the slurry were filtered and analysed by ICP for copper,
sulphur and calcium concentrations. After the desired reaction time
the slurry was filtered and the final solids were dried. A portion
of the solids was dissolved in acid and sent for ICP analysis to
determine its composition.
[0144] The results of a precipitation experiment using lime and
limestone is presented in the graph shown in FIG. 6. In order to
precipitate all of the copper in solution as a basic copper
sulphate (CuSO.sub.4.3Cu(OH).sub.2), a precipitation reagent to
copper sulphate mole ratio of 0.75:1 is required. In experiments
labelled CuB8 and CuB10 an excess of the precipitation reagent was
added so the mole ratio of reagent to copper in the reactor was
1:1. This is equivalent to a 33% excess of reagent. In the
experiment labelled CuB15, the limestone to copper mole ratio added
to the reactor was 0.5:1. This is equivalent to adding only 66% of
the limestone required to precipitate all of the copper.
[0145] The results in FIG. 6 indicate that both lime and limestone,
when present in sufficient amounts, are effective at precipitating
substantially all of the copper out of solution. It can be seen
that the ratio of pH increasing agent to copper sulphate plays a
role in the precipitation in that the experiment where insufficient
limestone was present did not result in complete precipitation. The
other two results indicate that lime has a more immediate effect on
the precipitation of copper-containing intermediates although
limestone was also very effective.
[0146] Table 2 indicates the results of the ICP analysis showing
the composition of the copper-containing intermediate which was
obtained for each precipitation experiment.
TABLE-US-00002 TABLE 2 Composition of copper-containing
intermediate precipitated from three experiments. Experiment
Details Intermediate Solid Composition wt % Molar ratios Ca Cu S
CuB8 13% 22% 10% 1 CaCO.sub.3:Cu 1 CuB10 15% 20% 16% 1 CaO:Cu, 1
CuB15 13% 25% 18% 1 CaCO.sub.3:Cu 1
Size Separation of Gypsum from Precipitated Copper
[0147] When precipitating copper with lime or limestone in the
sulphate system, the precipitated product will always be
contaminated by gypsum (calcium sulphate dihydrate). One method of
removing a portion of the gypsum from the solid product is to
control the precipitation conditions such that the gypsum, which
tends to form long needle like particles, grows to form large
particles and the copper solid remains small. This differential in
size will allow a physical size or density separation process to be
carried out which will remove a portion of the gypsum from the
copper product. In practice the size separation may be carried out
by sieving, hydrocycloning or a number of other size or settling
based physical separation methods. Recycling of a portion of the
solid would also be useful to ensure that the gypsum particles
always grow larger than the copper particles.
[0148] An experiment was performed wherein copper was precipitated
with limestone, as described above, and left to stand for two days
to allow the gypsum to continue to crystallise out and grow large
needles. A portion of the solids were then taken and screened
through a 53 micron sieve. The screening was carried out with a
calcium sulphate saturated wash solution in order to avoid
dissolving any of the crystallised gypsum. The different screen
portions were sampled and analysed for their copper, calcium and
sulphur compositions.
TABLE-US-00003 TABLE 3 Composition of solids before and after
screening. Solid Composition wt % Sample Ca Cu S Initial solids 15%
19% 15% Oversize solids 24% 7% 15% Undersize solids 12% 19% 10%
TABLE-US-00004 TABLE 4 Recovery of calcium, copper and sulphur,
quoted in wt % of the solid composition, in size fractions after
screening at 53 micrometres. Total Ca Cu S Mass Recovery to 55% 11%
56% 38% oversize Recovery to 45% 89% 44% 62% undersize
[0149] These experimental results indicate that over half of the
calcium can be removed from the copper by screening at 53 micron.
Removal of more than half of the calcium resulted in a loss of just
11% of the copper. Another way of looking at this is that the
initial solids had a Cu:Ca weight ratio of 1.34:1. The oversize
solids had a Cu:Ca weight ratio of 0.27:1 and the undersize solids
had a Cu:Ca ratio of weight 2.64:1.
[0150] In practice the amount of copper in the oversize material
can be minimised so the oversize material can either be recycled to
the iron precipitation step or leach to recover any remaining
copper or, if significant amounts of copper are not expected,
simply disposed of. The undersize material may undergo a further
size separation such that an even higher grade copper portion can
be extracted while a lower grade copper portion can be recycled to
the copper precipitation step in order to seed the gypsum
crystallisation. If a second size separation is not carried out
then instead a split of the undersize portion would most likely be
used to recycle some seed to the copper precipitation.
Heat Treatment
[0151] The thermal decomposition of a brochantite
(CuSO.sub.4.3Cu(OH).sub.2) sample was investigated to demonstrate
the decomposition of a copper-containing intermediate compounds.
The thermogravimetric and differential scanning calorimetry results
are shown in FIG. 7. The sample used in this experiment was dried
in air at room temperature prior to the thermogravimetric analysis
so the amount of moisture associated with the sample is less than
would typically be associated with the precipitated product. As
such the mass loss should only include the losses caused by
decomposition of the solid. The decomposition of the basic copper
sulphate (BCS) is described by the equations in table 5, below.
TABLE-US-00005 TABLE 5 Reactions describing the decomposition of a
basic copper sulphate solid at increasing temperatures carried out
under an analytical grade nitrogen gas atmosphere. Theoretical step
mass loss Overall Temp. Reaction (from BCS) mass loss range
.degree. C. Dehydroxylation reaction 12.0% 12.0% 25-450
CuSO.sub.4.cndot.3Cu(OH).sub.2 (s) .fwdarw. CuSO.sub.4.cndot.3CuO
(am s) + 3H.sub.2O (g) Recrystallization of doleroph- 0% 12.0%
490-520 anite CuSO.sub.4.cndot.3CuO (s) .fwdarw.
CuSO.sub.4.cndot.CuO (s) + 2CuO (s) Desulphurisation reaction 17.7%
29.7% 575-750 CuSO.sub.4.cndot.CuO (s) .fwdarw.2CuO (s) + SO.sub.3
(g) Reduction to copper (I) oxide 7.1% 36.7% 800-910 4CuO (s)
.fwdarw. 2Cu.sub.2O (s) + O.sub.2 (g) Reduction to copper metal
7.1% 43.7% 2Cu.sub.2O (s) .fwdarw. 4Cu (s) + O.sub.2 (g)
[0152] Initially any associated water and water of crystallisation
will evaporate. This occurs before and at the same time as the
copper hydroxide portion of the solid decomposes to copper oxide
and steam, up to about 450.degree. C. As noted previously, the
amount of moisture associated with the sample was minimised prior
to this experiment by drying the sample in air at room temperature
so the initial loss of associated water and water of
crystallisation is minimal in this experiment. At about 500.degree.
C. some of the copper oxide and copper sulphate recrystallises to
form dolerophanite. This is an exothermic reaction, hence the spike
in the heat flow trace shown in FIG. 7. The copper sulphate portion
of the solid then decomposes to copper oxide and sulphur trioxide
gas (SO.sub.3). The remaining solid from 750.degree. C. onwards is
copper oxide (CuO). At about 800.degree. C. the copper oxide
decomposes to monovalent copper oxide (Cu.sub.2O) and then to
copper metal at even higher temperature. The exact temperatures to
which the copper oxides are stable before decomposing to copper
metal will depend on the oxygen partial pressure in the gas
phase.
[0153] To confirm this decomposition route, larger samples of the
brochantite were heated in a furnace to a specific temperature. The
resulting solids were then analysed by X-ray Powder Diffraction.
The results of this XRD analysis are shown in table 6, below.
TABLE-US-00006 TABLE 6 The composition of a copper-containing
material after heating under a dry argon gas atmosphere. Mass Loss
Temperature .degree. C. wt % Minerals Mineral Formula 25 0.0%
Brochantite CuSO.sub.4.cndot.3Cu(OH).sub.2 450 13.4% Tenorite CuO
Antlerite CuSO.sub.4.cndot.2Cu(OH).sub.2 Dolerophanite
CuSO.sub.4.cndot.CuO 750 30.8% Tenorite CuO
[0154] The initial sample testing proved that the solid was
brochantite. At 450.degree. C. there is some copper hydroxide
remaining and some copper oxide and dolerophanite formed. It is
interesting to note that the presence of antlerite indicates that
the copper hydroxide and sulphates are recrystallising just as the
copper oxide and sulphates are to form the dolerophanite. By
750.degree. C. all of the hydroxides and sulphates have been
removed and only copper oxide remains indicating that the
copper-containing intermediate obtainable by the method of this
invention can be fed into a heat treatment step to successfully
obtain a copper-containing product suitable for conversion to a
final copper product for commercial applications.
[0155] In providing a point of integration between the
hydrometallurgical and pyrometallurgical routes the present method
opens up a much wider range of source materials which can be
processed such that they can ultimately be fed into a converter
operation. This presents distinct operational advantages not only
in providing the ability to better process previously
under-utilised or ignored copper sources but also in reducing costs
of operation and environmental impact of the processing.
[0156] A key component of the present method is the realisation
that, not only could a copper-containing intermediate be
selectively precipitated out of an acidic leach solution with the
addition of low cost reagents such as lime and limestone but,
importantly, that such a precipitate would actually be well suited
to introduction into a converter operation for a final
pyrometallurgical step to produce blister copper or a similar
useful end product.
[0157] A number of advantages are realised along the way including
those in the ease of impurity reduction. Particularly, (i) lead
will not leach to any significant extent in a sulphate solution;
(ii) arsenic can be precipitated along with iron prior to copper
precipitation; (iii) nickel, cobalt and zinc precipitate at
noticeably higher pH values than the copper; (iv) precipitation
with lime or limestone will produce a relatively clean tailings
solution as the copper, calcium and sulphate will all be mostly
removed; and (v) the precipitation process should not be
significantly affected by process water salinity.
[0158] Further advantages include that no capital is required for
solvent extraction and electrowinning facilities as the formation
of a precipitate which can be fed into a smelter/converter obviates
the need for these expensive circuits of the hydrometallurgical
approach. Lower comparative ongoing operational expenses would be
likely due to the negation of the need for solvent extraction
reagent and the lower usage of electricity.
[0159] Further, and as mentioned previously, the precipitated
copper-containing intermediate, due to its composition including
significant amounts of oxygen, provides an extra source of oxygen
in the converter thereby lowering the cost of oxygen injection and
potentially increasing the oxygen injection rate into the process.
Unreacted lime or limestone and gypsum in the precipitated copper
product will react and provide a source of lime. This will be
beneficial in converters that are using a slag that includes lime
as it will decrease the fluxing requirements. The sulphate
contained in gypsum and any sulphate associated with the
precipitated copper solid can also be used to regenerate sulphuric
acid.
[0160] Significant amounts of the world's copper supply come from
countries such as Chile and Peru which do not necessarily have
adequately developed infrastructure for hydrometallurgical copper
processing or which are coming under increasing pressure from the
cost of electrowinning. A distinct advantage of the present method
is that the precipitated copper-containing intermediate could be
transported from a remote location to an existing copper smelter
and would be suitable for charging directly into the smelter or
converter.
[0161] Stabilisation of arsenic to the slag phase is an important
environmental advantage of the present process. Unlike in the
smelting stage, the copper converter stage typically operates with
more oxidising conditions meaning that arsenic is more readily
dissolved and stabilised in the molten slag. Incorporation of the
arsenic in the stable slag phase then avoids the problem of arsenic
release into the environment and so the present method allowing, as
it does, direct introduction of the copper-containing intermediate
into a converter can take advantage of this fact. Additionally,
there is also the option of removing arsenic with iron during the
impurity precipitation stage which would avoid having to deal with
it in pyrometallurgical processes completely.
[0162] The present inventive method represents an approach whereby
the cost and environmental impacts of copper processing are able to
be significantly reduced from those presently seen. The processing
steps are simple, low cost operations chosen to achieve the
extraction and separation required while also minimising the
environmental impact of any residues produced.
[0163] The above description of various embodiments of the present
invention is provided for purposes of description to one of
ordinary skill in the related art. It is not intended to be
exhaustive or to limit the invention to a single disclosed
embodiment. As mentioned above, numerous alternatives and
variations to the present invention will be apparent to those
skilled in the art of the above teaching. Accordingly, while some
alternative embodiments have been discussed specifically, other
embodiments will be apparent or relatively easily developed by
those of ordinary skill in the art. Accordingly, this patent
specification is intended to embrace all alternatives,
modifications and variations of the present invention that have
been discussed herein, and other embodiments that fall within the
spirit and scope of the above described invention.
[0164] In the claims which follow and in the preceding description
of the invention, except where the context clearly requires
otherwise due to express language or necessary implication, the
word "comprise", or variations thereof including "comprises" or
"comprising", is used in an inclusive sense, that is, to specify
the presence of the stated integers but without precluding the
presence or addition of further integers in one or more embodiments
of the invention.
* * * * *