Patent Application
20160304988
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.
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.
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.
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.
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.
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.
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.
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
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.
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.
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.
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.
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.
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
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.
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.
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.
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.
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.
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.
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
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.
Other preferred objects of the present invention will become apparent from the following description.
According to a first aspect of the invention, there is provided a method of processing a copper-containing source material including the steps of:
to thereby process the copper-containing source material.
In certain embodiments of the first aspect, there is provided a method of processing a copper sulphide-containing source material including the steps of:
to thereby process the copper sulphide-containing source material.
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.
Further features and advantages of the present invention will become apparent from the following detailed description.
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:
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.
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.
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.
In a first aspect of the invention, there is provided a method of processing a copper-containing source material including the steps of:
to thereby process the copper-containing source material.
A representative flow sheet for one embodiment of the present method is shown in
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
After removal of the impurity precipitate, additional pH increasing agent, again limestone in the embodiment shown in
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
The precipitated copper-containing intermediate may then be exposed to an optional heating step, indicated by hatched shading in
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.
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.
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
The materials and requirements in each step of the present method will now be addressed in more detail.
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.
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.
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.
Preferably, the copper-containing ore is a copper sulphide or copper oxide ore.
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.
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.
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.
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.
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.
The pH increasing agent may be an alkali metal or alkali earth metal carbonate, oxide, hydroxide or compounds or associations thereof.
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(CO3)2) and other minerals containing one or more of these compounds.
Preferably, the pH increasing agent is lime (calcium oxide) and/or limestone (calcium carbonate).
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.
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.
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.
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.
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.
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.
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.
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.
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)2); basic copper sulphates such as, or analogous, to Brochantite (CuSO4.3Cu(OH)2), Posnjakite (CuSO4.3Cu(OH)2.H2O), Wroewolfeite or Langite (CuSO4.3Cu(OH)2.2H2O), Vernadskite (CuSO4.3Cu(OH)2.4H2O), Antlerite (CuSO4.2Cu(OH)2), Antlerite hydrate (CuSO4.2Cu(OH)2.H2O), Antlerite dihydrate or Kamarezite (CuSO4.2Cu(OH)2.2H2O) and Dolerophanite (CuSO4.CuO); basic copper carbonates such as Azurite (2CuCO3.Cu(OH)2) and Malachite (CuCO3.Cu(OH)2). From a nitrate solution basic copper nitrate (Cu2(NO3)(OH)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 (Cu2Cl(OH)3), Belloite (CuClOH) or cuprite (Cu2O). 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 (CuCO3).
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.
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.
In one embodiment, the high temperature treatment occurs at a temperature of at least 200° C., preferably at least 300° C., more preferably at least 1000° C. This includes temperature treatments of between about 200° C. to about 1500° C., about 300° C. to about 1500° C., about 400° C. to about 1500° C. about 500° C. to about 1500° C., about 600° C. to about 1500° C., about 700° C. to about 1500° C., about 800° C. to about 1500° C., about 900° C. to about 1500° C.,
In certain embodiments, the high temperature treatment is at a temperature of between about 1000° C. to about 1500° C., such as may be achieved in a converter, preferably about 1000° C. to about 1400° C., more preferably between about 1200° C. to about 1400° C. These broad ranges are inclusive of sub-ranges of about 1100° C. to about 1500° C., about 1100° C. to about 1400° C., about 1100° C. to about 1350° C., about 1200° C. to about 1500° C., about 1200° C. to about 1450° C. and about 1250° C. to about 1400° C. Temperatures of about 1250° C. 1300° C. and 1350° C. may be preferred.
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° C. and up to 1800° C. may be acceptable but are not generally required.
Both the smelter and converter operations may have introduced oxygen but the converter generally operates under relatively more strongly oxidising conditions than the smelter.
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.
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.
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.
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.
In some embodiments, arsenic can also be removed with the iron in this precipitation step.
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.
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.
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.
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.
Suitably, the separation may be effected by size screening or sieving, hydrocycloning and the like.
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.
The method may further include the step of heating the copper-containing intermediate to a temperature between 25° C. to less than 1000° C. or between 200° C. to 800° C., including 25° C. to 250° C. or 25° C. to 200° C., prior to its exposure to higher temperatures.
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.
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° 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° 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.
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.
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° 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° C. to 800° 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° C. to 1300° 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.
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
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.
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 SO2 or SO3 gas. The SO2 or SO3 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 Fe3O4 and Fe2O3. 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.
The oxidative roast may occur at a temperature of from about 500° C. to about 950° C. This range is inclusive of the oxidative roast being carried out at a temperature of from 500° C. to 900° C., 500° C. to 850° C., 500° C. to 800° C., 500° C. to 750° C., 550° C. to 950° C., 550° C. to 900° C., 550° C. to 850° C., 550° C. to 800° C., 550° C. to 750° C., 600° C. to 950° C., 600° C. to 900° C., 600° C. to 850° C., 600° C. to 800° C., 600° C. to 750° C., 650° C. to 950° C., 650° C. to 900° C., 650° C. to 850° C., 650° C. to 800° C., 650° C. to 750° C., 700° C. to 950° C., 700° C. to 900° C., 700° C. to 850° C. and 700° C. to 800° C.
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.
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.
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.
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.
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.
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.
Preferably, the leach solution is an acidic leach solution. The nature of the acid may be as previously described.
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.
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
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.
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:
This process is set out in a representative flow sheet as shown in
The embodiment of
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.
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.
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.
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.
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.
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.
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.
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.
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.
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.
A second aspect of the invention results in a concentrated copper product when produced by the method of the first aspect.
The concentrated copper product may be substantially pure copper metal.
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° C., 750° C. and 900° 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-SO2/min which equates to an enforced atmosphere of about 0.06 atm SO2, 0.94 atm air which is equivalent to 0.20 atm O2. All solids formed were characterised by Powder XRD. All solids contained some silica.
At 600° 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° 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° 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.
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
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
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).
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.
The results of a precipitation experiment using lime and limestone is presented in the graph shown in
The results in
Table 2 indicates the results of the ICP analysis showing the composition of the copper-containing intermediate which was obtained for each precipitation experiment.
Size Separation of Gypsum from Precipitated Copper
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.
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.
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.
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.
The thermal decomposition of a brochantite (CuSO4.3Cu(OH)2) sample was investigated to demonstrate the decomposition of a copper-containing intermediate compounds. The thermogravimetric and differential scanning calorimetry results are shown in
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° 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° 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
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.
The initial sample testing proved that the solid was brochantite. At 450° 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° 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.
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.
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.
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.
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.
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.
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.
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.
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.
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.
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.
| Number | Date | Country | Kind |
|---|---|---|---|
| 2013904688 | Dec 2013 | AU | national |
| 2014900280 | Jan 2014 | AU | national |
| Filing Document | Filing Date | Country | Kind |
|---|---|---|---|
| PCT/AU2014/050393 | 12/3/2014 | WO | 00 |
