MXPA01006067A - High temperature heap bioleaching process - Google Patents

Abstract

A heap (20) is constructed with chalcopyrite bearing ore. The constructed heap (20) includes exposed sulfide mineral particles wherein at least 25 weight percent of which are chalcopyrite. The concentration of the exposed sulfide mineral particles in the heap (20) is such that the heap (20) includes at least 10 kg of exposed sulfide per tonne of solids in the heap (20). Furthermore, at least 50%of the total copper in the heap (20) is in the form of chalcopyrite. A substantial portion of the heap (20) is then heated to a temperature of at least 50°C. The heap is inoculated with a culture including at least one strain of thermophilic microorganisms. A process leach solution that includes sulfuric acid and ferric iron is applied to the heap (20) to bioleach the sulfide minerals in the heap (20). Bioleaching is carried out so that sufficient sulfide mineral particles in the heap (20) are biooxidized to oxidize at least 10 kg of sulfide sulfur per tonne of solids in the heap (20) and to cause the dissolution of at least 50%of the copper in the heap (20) into the process leach solution in a period of 210 days or less from completion of the heap (20). A pregnant process leach solution (40) that contains dissolved copper is collected from the heap (20) as it drains from the heap.

Description

PROCEDURE OF B1OUXIVIAC1ON IN HIGH TEMPERATURE PILE BACKGROUND OF THE INVENTION FIELD OF THE INVENTION The present invention relates to the extraction of copper from minerals and concentrates having chalcopyrite.

BACKGROUND OF THE INVENTION Chalcopyrite, a sulphide copper mineral, is economically the most important source of copper. Currently, casting technology remains the main technology for recovering copper from chalcopyrite. However, chalcopyrite casting has a number of disadvantages. These include emissions of sulfur dioxide gas that are unacceptable to the environment, large-scale production of sulfuric acid even though there is currently only a limited market for sulfuric acid in most areas, and it is expensive. As a result, alternative methods to recover copper from chalcopyrite that are less damaging to the environment and less expensive have been sought for a number of years.

A number of alternatives that have been investigated to recover copper from chalcopyrite and its minerals have included hydrometallurgical procedures. Hydrometallurgical processes have long been used to recover copper from oxide minerals. These processes typically involve leaching sulfuric acid from the oxide ore, separating copper from the impregnation leaching liquid by solvent extraction techniques, and recovering metallic copper from the liquid evolved by electrolytic extraction. These techniques have not only demonstrated an ability to recover copper to a competitive cost advantage over most smelting processes, but also the electrolytically extracted copper produced in said process is also currently completely competitive in terms of quality with electrorefined copper. produced by the known techniques of smelting and refining. However, currently, a commercially viable hydrometallurgical process for the recovery of copper from chalcopyrite has remained elusive despite intensive research efforts to develop such a process. The development of a hydrometallurgical process for the direct leaching of chalcopyrite by either chemical or biological means has been pursued continuously for more than twenty years. The direct leaching of chalcopyrite in sulfuric acid solution has a variety of problems. At temperatures below the sulfur melting point (approximately 118 ° C), the rate of dissolution of copper has, to date, been slow in a non-economic manner. At temperatures above the sulfur melting point the chalcopyrite becomes passive so it is believed to be a layer of elemental sulfur that forms on the unreacted sulfide particles. This again converts non-economic copper extraction through this procedure. Other leaching systems that have been studied over the years for the extraction of copper from chalcopyrite in the laboratory or on a pilot scale include systems that use concentrated solutions of ferric chloride or ammonia ammonia as leaching. Efforts to biobleach chalcopyrite on a commercial scale have not been successful to date either. Chalcopyrite is notoriously difficult to bioleave even though bioleaching is currently used as the primary production method to extract copper from other copper sulphide minerals such as chalcocite and covellite in various mineral extraction operations around the world. Processes of agitated tank and biooxidation in pile that have used mesophiles, such as Thiobacillus ferrooxidans, the most commonly used microorganism for biooxidating sulphide minerals, have also not been successful due to the slow kinetics of chalcopyrite leaching. Slow leaching kinetics and incomplete biooxidation of chalcopyrite are often attributed to the formation of an inhibition or passivation layer that forms on the surface of the chalcopyrite as it oxidizes. A number of different additives have been used in an attempt to increase copper dissolution from chalcopyrite, presumably upsetting the passivation layer. These additives include metal salts such as Ag2S04, Bi (NO3), graphite, and other sulfide minerals. Any biohydrometallurgical procedure to treat chalcopyrite, therefore, will have to face the problem of this surface layer. The studies of the problem have led to several theories that refer to the nature of the inhibitory layer. One theory is that a coating of jarosite is formed on the chalcopyrite surface as it is leached. Jarosite is formed in the presence of sulfate and ferric iron, in environments in which the pH increases above 1.8. However, high concentrations of constituent molecules of jarosite (sulfate, ferric iron, ammonium or potassium) will lead to formation of jarosite at a lower pH. The presence of jarosite in biolixed chalcopyrite analysis supports this theory. However, the experiments carried out by the inventors of the present that show a slow leaching even at a low concentration of constituent molecule and at low pH, as well as reports in the literature, contradict this theory. Another theory is that elemental sulfur produced during bioleaching forms a thick mantle that excludes bacteria and other chemical oxidants from the chalcopyrite surface. The detection of large amounts of sulfur in bioleaching chalcopyrite supports this theory. In addition, many electron microphotographs have shown a thick coating of sulfur on leached chalcopyrite. However, this theory does not adequately explain why other metal sulphides that also form sulfur when leached are not leached as slowly as chalcopyrite. A third theory proposes that inhibition is caused by the formation of an intermediate sulfide passivation layer. It is believed that this passivation layer is less reactive than the original chalcopyrite and can also inhibit the flow of electrons and oxidants to and from chalcopyrite. The exact nature of this layer of passivation is complex and is the subject of scientific debate. However, there is a good agreement between the data in the literature that the passivation layer is unstable at higher temperatures. For example, it has been found that temperatures above 60 ° C are high enough to minimize chalcopyrite passivation during leaching. Experiments with leaching at higher temperatures by chemical and biological means have shown accelerated chalcopyrite leaching. Chemical leaching done at more than 100 ° C, however, requires expensive pressure reactors. Biological leaching is limited to the temperature limits of microorganisms that are capable of oxidizing metal sulphides or oxidizing ferrous to ferric. Some examples of microorganisms capable of oxidizing ferrous, metal sulphides, and elemental sulfur in environments above 60 ° C include: Acidianus brierleyi, Acidian? S infernus, Metallosphaera sedula, Sulfolobus acidocaldarius, Sulfolobus BC, and Sulfolobus metallicus. However, there are also other extreme thermophiles that can grow and leach metal sulphides at temperatures above 60 ° C. Agitated tank procedures using thermophiles have resulted in faster chalcopyrite bioleaching than those using moderate mesophiles or thermophiles. In fact, several microorganisms have been used in stirred tank procedures to leach chalcopyrite concentrate in less than 10 days of leaching time. However, the high temperature that is required for rapid chalcopyrite leaching increases the limitations of mass transfer of oxygen and carbon dioxide in the system. This in turn has placed severe limitations on the density of pulp that can be used in those stirred tank processes due to the high oxygen requirements of the thermophiles and the oxidation reaction that occurs on the surface of the chalcopyrite during leaching. In this way, even though the bioleaching process can be completed in less than 10 days in a stirred tank procedure, the high capital and operating costs associated with the operation of a plant at low pulp densities needed to meet the oxygen requirements of the system they have avoided the commercial implementation of stirred tank bioleaching for chalcopyrite concentrates. If an effective pile-up bioleaching process for chalcopyrite could be developed, it would have the potential to operate at a lower cost than the bioleaching in concentrate tank or pressure leaching of the concentrate or chalcopyrite minerals. In this way, heap leaching of chalcopyrite would be the preferred low-cost procedure if a procedure could be developed to extract a high percentage of copper in a matter of months. The use of thermophiles in a heap leach procedure on a pilot scale is reported in Madsen, B. and Graves, R., Percolation Leaching of a Chalcopyrite-Bearing Ore at Ambient and Elevated Temperatures with Bacteria, 1983, Bureau of Mines. However, the procedure described in that document was unable to achieve satisfactory recoveries in a reasonable short period of time and therefore is not commercially viable. There have been other reports of bio-leaching procedures in batteries that reach temperatures above 60 ° C. However, these have also not been commercially viable to extract copper from chalcopyrite minerals. The faults of all reported bioleaching processes for chalcopyrite minerals is that all have generally taken more than a year to leach and recover less than 50% of the copper in the chalcopyrite. The reasons for this are not completely clear. However, the inventors of the present have determined that there are several factors that have acted together to avoid bioleaching by successful stack of chalcopyrite ore. The first is that batteries that have eventually reached a temperature of 60 ° C or higher have taken a longer time to accumulate enough heat to reach such high temperatures. As a result, once the temperature of 60 ° C is reached, the amount of sulfide ore particles exposed in the pile is insufficient to maintain the temperature to complete the leaching of copper. Additionally, in the case of larger ore particles, such as those of 2.5 cm, sufficient copper sulfide in the ore is not exposed to the leach solution to allow adequate recoveries. Finally, high temperatures can also increase the amount of ferrous ions that precipitate as jarosite, which can further slow leaching.

BRIEF DESCRIPTION OF THE INVENTION The present invention is directed to a high temperature bioleaching process for extracting copper from chalcopyrite minerals. More particularly, the present invention is directed to providing a high temperature bioleaching process for extracting at least 50% of the copper from a pile comprising chalcopyrite mineral in a period of about 210 days or less. It will be understood, as used herein, that chalcopyrite minerals refer to crushed chalcopyrite minerals and chalcopyrite concentrates. A method according to an aspect of the present invention for extracting copper from chalcopyrite ore comprises the steps of: a) constructing a pile comprising ore having chalcopyrite, the pile includes particles of sulfide ore exposed at least 25% by weight of which comprise chalcopyrite in which the concentration of sulfide mineral particles exposed in the pile is such that the pile contains at least 10 kg of sulfuric sulfur exposed per ton of solids in the pile, and in which at least 50% of the total copper in the pile is in the form of chalcopyrite; b) heating a substantial portion of the stack to a temperature of at least 50 ° C; c) inoculating the cell with a culture comprising at least one thermophilic microorganism capable of bio-oxidizing sulfide minerals at a temperature above 50 ° C; d) irrigating the stack with a leaching process solution comprising sulfuric acid and ferric iron; e) Biolixivir enough sulfide mineral particles in the cell to oxidize at least 10 kg of sulfur in the form of sulfur per ton of solids in the cell and cause the dissolution of at least 50% of the copper in the cell in the process solution of leaching in a period of 210 days or less from the completion of the heap; and f) collecting an impregnating leaching process solution containing dissolved copper as it is drained from said stack. Preferably, a substantial majority of the heat that is required to initially heat the stack to temperature and to maintain the stack at temperature is derived from the bioleaching of sulphide minerals contained within the stack. In another aspect of the present invention, a high temperature cell bioleaching process is provided to recover copper from ore having chalcopyrite. The method according to this aspect of the invention comprises the steps of: a) constructing a stack comprising ore having chalcopyrite, the stack includes exposed sulfide mineral particles, at least 25% by weight of which comprise chalcopyrite, in which the concentration of sulfide minerals exposed in the pile is such that the stack includes at least 10 kg of sulfur in the form of exposed sulfur per tonne of solids in the stack, and in which at least 50% of the total copper in the pile is in the form of chalcopyrite; b) heating at least 50% of the stack to a temperature of at least 60 ° C; c) maintaining at least 50% of the stack at a temperature of at least 60 ° C until at least 50% of the copper in the stack dissolves; d) inoculating the cell with a culture comprising at least one thermophilic microorganism capable of biolixing sulfide minerals at a temperature above 60 ° C; e) irrigate the pile with a leaching process solution at a rate of at least 72 liters / m2 / day; f) bio-lixivizing sulfide mineral particles in the cell, in which sufficient sulfur minerals are oxidized in a bioleaching period of 210 days or less to oxidize at least 10 kg of sulfur sulfur per ton of solids in the cell and cause the dissolution of at least 50% of the copper in the pile in the leaching process solution; g) collecting an impregnating leaching process solution that includes copper cations as drained from the cell during the bioleaching period; and h) recovering copper from the impregnating leaching process solution. The above aspects and other aspects, features and advantages of the present invention will be apparent to those skilled in the art from the following description of the preferred embodiments taken in conjunction with the appended figures.

BRIEF DESCRIPTION OF THE DRAWINGS Figure 1 is a schematic illustration of a process flow chart according to an embodiment of the present invention. Figures 2A-2D illustrate another method for practicing the present invention. Figure 3 is a graph illustrating the estimated percentage of copper and iron leached for Example 1 illustrating certain principles of the present invention. Figure 4 is a graph illustrating the estimated percentage of copper and iron leached for example 2 that illustrates certain principles of the present invention. Figure 5 is a graph illustrating the estimated percentage of copper and iron leached for comparative example 3. Figure 6 is a graph illustrating the estimated percentage of copper and iron leached for example 4 that illustrates certain principles of ia present invention. Figure 7 is a graph illustrating the estimated percentage of copper and iron leached for example 5 that illustrates certain principles of the present invention.

Figure 8 is a graph illustrating the estimated percentage of copper and iron leached for example 6 that illustrates certain principles of the present invention. Figure 9 is a graph illustrating the estimated percentage of copper and iron leached for example 7 that illustrates certain principles of the present invention.

DETAILED DESCRIPTION OF THE PREFERRED MODALITIES The present invention improves chalcopyrite pile bio-leaching by providing a method to accelerate the rate of leaching and increase the percentage of copper leached from the pile. The introduction of a suitable fuel value in the stack during the construction of the stack is an important aspect of the high temperature stack bioleaching process of the present invention. The fuel component may be in the form of chalcopyrite, pyrite, chalcocite, covelite and other sulfur minerals that generate a large amount of heat energy when biooxidated. The heat is generated by the exothermic oxidation reactions that occur during the biooxidation of these fuel values. A significant portion of the sulfur fuel material, therefore, must be exposed to air, water, and biooxidant microorganisms or ferric ions within the stack to ensure that an adequate amount of heat can be generated in a sufficiently short period of time to supply a substantial portion of the heat required to maintain the pile at a temperature above 50 ° C while the chalcopyrite in the pile is biolixed. If sufficient fuel values are not present, the battery can not be maintained at a temperature above 50 ° C while biooxidation of the chalcopyrite proceeds without providing substantial amounts of heat from an external source, which would make the procedure economically prohibitive . The procedure would be typically economical if the sulphide minerals exposed in the stack, ie those particles of sulfide ore that can be biooxidated in a period of 220 days or less, contain at least 10 kg of sulfur in the form of sulfur per tonne. of solids in the pile. In other words, the stack must contain at least 10 kg of sulfur in the form of exposed sulfur per ton of solids in the stack. This concentration of sulfur in the form of sulfur translates into a heat value of approximately 50,000 kcal / ton of solids with oxidation. This is based on the fact that the standard free energy change for the oxidation of pyrite by the reaction according to equation (1): FeS2 + 3.502 + H2O Fe2 + + 2SOI ~ + 2H + (1) is approximately 1440 KJ Additionally, although the standard free energy change for the various other sulfide minerals is different, due to the heat of formation of each mass of S0 ~ account for most of the change in standard free energy for all those reactions, it can be assumed that the Change in standard free energy for the oxidation reactions of the other sulfide minerals is approximately the same. One may suppose, therefore, that for each mole of oxidized S2, approximately 1440 KJ of energy will be released. Thus, if a pile contains 1% by weight of sulfur in exposed sulfide minerals, or 10 kg of sulfur in the form of exposed sulfur per ton of solids, it contains a potentially useful fuel value of approximately 50,000 kcal per ton of solids. in the pile. Obviously while the concentration of exposed sulfide minerals contained within the pile is higher, the value of potentially useful fuel in the stack will be greater and the heat that will be needed will be less than that supplied by an external source. If the concentration of sulfide minerals exposed in the pile is sufficiently high, the fuel component of the pile will be such that the heat generated with oxidation will be sufficient to heat the pile to a temperature above 50 ° C and to maintain the pile above 50 ° C while the biooxidation of the chalcopyrite proceeds. A high temperature battery bioleaching process for extracting copper from ore having chalcopyrite in accordance with the present invention is illustrated schematically in Figure 1. According to the method, a pile 20 is constructed with mineral having chalcopyrite. It is desirable that the stack 20 be at least 2.5 m high and at least 5 m wide so that the outer ends of the stack will help to insulate the interior portions of the stack. The stack 20 will typically have larger dimensions to make the process as economical as possible. For example, the stack will typically have a height of at least 3 m and a width of at least 10 m. The length of the stack 20 will typically depend on the limitations of the site on which the stack is built, but in general the stack will be substantially longer than wide. Although the dimensions mentioned above have been provided as guides, those skilled in the art will recognize that the dimensions of the stack 20 can vary significantly. Additionally, the stack does not have to be rectangular as illustrated in Figure 1, but may also be circular or any other desired shape or perhaps required by the limitations of the site at which the procedure will be carried out. When complete, the stack 20 will generally contain at least 4% by weight of water. Preferably the stack 20 will include 7% or more water by weight. However, the more water is contained within the stack 20, the greater the amount of heat required to heat the stack 20 to a temperature above 50 ° C where the active biooxidation of the chalcopyrite in the stack starts. For example, a pile containing 7% water by weight will consume approximately 3,500 kcal of heat to heat the water in the pile from 20 ° C to 70 ° C for each ton of solids in the pile. While a pile containing 10% water by weight will take approximately 5,000 kcal of heat to heat the water in the pile from 20 ° C to 70 ° C for every ton of solids in the pile. Moreover, the specific heat of the water is greater than that of the mineral. Therefore, it is desirable to maintain the water content of the initial stack at a level that does not exceed 15% water by weight of solids in the cell. The water can be added to the pile during the formation of the pile or following the completion of the pile while it is being conditioned in preparation for the bioleaching process. As noted above, the stack 20 must also include exposed sulfur mineral particles. The concentration of sulfide mineral particles exposed in the stack 20 should be such that the stack includes at least 10 kg of sulfur sulfide exposed per ton of solids in the stack. To improve the performance of the present process, however, the concentration of sulfide mineral particles is preferably such that the cell will contain at least 30 kg of sulfur in the form of exposed sulfide. With adequate stack design considerations, as will be discussed in more detail below, the sulfur concentration in the form of exposed sulfur can reach levels of 40 to 90 kg per ton of solids in the stack or even higher. In this way, even more preferably the concentration of sulfur in the form of exposed sulfide is at least about 45 kg per ton of solids. As used herein, exposed sulfide mineral particles will be understood to be those sulfide mineral particles that are exposed to air, water, and biooxidant microorganisms or ferric ions within the stack so that they can generally be biooxidized within of a period of 210 days or less. Sulfur sulfide in those exposed sulfide mineral particles is referred to as sulfur sulfide exposed for purposes of the present application to distinguish it from other sulfur sulfides that may be in the stack but due to its clogging, in reinforcing material for example , its fuel value is not available for the battery in a period of reasonable biooxidation of 210 days or less. Typically, most of the sulfide mineral particles exposed within the preferred stack designs of the present invention will be finely ground and will have a particle size of 250 μm or less, and preferably a particle size of less than 107 μm. However, exposed sulfur ore particles may also be present in larger ore particles that can be found in the pile. This is because some fraction of the sulfide ore particles contained within larger ore particles will typically reside on the surface, or sufficiently close to the surface, of the ore particles to allow access of the necessary components to occur. oxidation, that is, air, water, and biooxidant bacteria or ferric ions, within a period of 210 days or less. As will be appreciated by those skilled in the art, finer ore particles will typically have more sulfide ore particles exposed than coarser ore particles.

Although the sulfide mineral particles exposed in the piles of the present invention will typically include a variety of sulfide minerals, a minimum of about 25% by weight of the sulfide mineral particles exposed in the stack must be chalcopyrite. Preferably the chalcopyrite fraction of the sulfide mineral particles in the stack is in the range of 30 to 70% by weight. The remainder of the sulfide mineral particles within the stack 20 preferably comprises more readily bio-oxidizable sulfide minerals such as pyrite, arsenopyrite, chalcocite, and coveiite. These less recalcitrant sulfur minerals provide an important fuel component to the pile, which can be used to heat the stack to temperature and help keep the stack at temperature while biooxidation of the chalcopyrite is proceeding. The presence of these other sulfide minerals is also desirable because they increase the galvanic leaching of chalcopyrite. Thus, in constructing the piles of the present invention, it is desirable that at least a portion of the exposed sulfide mineral particles comprise one or more less recalcitrant sulfide minerals such as pyrite, arsenopyrite, covellite and chalcocite. However, the invention can be practiced with up to 100% of the sulfide minerals in the pile being chalcopyrite. Because other copper sulfide minerals such as chalcocite and covellite can be easily bio-lixivied using mesophiles such as Thiobacillus ferrooxidans, the present high temperature process is not so economically justified to process those copper sulphide minerals alone. Accordingly, at least 50% of the copper in the stack 20 should be in the form of chalcopyrite so that this mineral is the main source of copper in the pile. Preferably at least 80 to 90% of the copper in the stack is in the form of chalcopyrite to maximize the amount of this recalcitrant ore that is being processed in the stack and hence the economic benefit of practicing the present invention. The stack 20 can be produced using any of the techniques known in the art to produce batteries for leaching as long as the above parameters are satisfied for the completed stack. As an example, the pile can be built by piling ore extracted from the mine to form a pile. However, preferably, the ore is crushed to a particle size so that 90% passes 2.54 cm. Alternatively, the crushed ore may be agglomerated before being stacked to improve the flow of air and liquid within the stack as is known in the art. Additionally, a sulfide ore concentrate can be added to the stack to increase the value of potentially useful fuel in the stack. A preferred method for forming the stack 20 is described in the U.S.A. 5,766,930, which is incorporated herein by reference as if fully set forth herein. The patent of E.U.A. DO NOT. No. 5,766,930 describes the construction and operation of large surface bioreactors that are particularly well suited for practicing the present invention.

Accordingly, the stack 20 can be constructed by grinding a chalcopyrite mineral to be bioleacted, such as a mineral extracted from the mine containing chalcopyrite, at a particle size that is less than 2.54 cm, and preferably less than 12.7 mm The fraction of fine material, for example the fraction that is less than 3 mm, is then removed to ensure an adequate air flow within the final stack. The plurality of crushed coarse ore particles are then coated with a sulfide mineral concentrate having a particle size of less than 250 μm, and preferably less than 107 μm. The concentrate comprises chalcopyrite and, preferably, one or more less recalcitrant sulfide minerals such as pyrite, arsenopyrite, chalcocite and covelite. However, as described above, since all of the sulfide mineral particles in the concentrate are considered exposed sulfide mineral particles, at least 25% by weight of the sulfide mineral content in the concentrate must be chalcopyrite. The concentrate can be coated on the substrates using a variety of techniques, including the use of a rotating drum or a suspension sprinkler. The thickness of the concentrate coating on the coarse ore is preferably less than 1 mm to ensure that the microorganisms that are used in the bioleaching process have adequate access to all the sulfide mineral particles in the concentrate. Thicker coatings will increase the capacity of the pile bioreactor, but the speed at which the bioleaching process proceeds will probably be encouraged due to decreased access of the microorganisms that are used to the underlying sulfide mineral particles in the concentrate. To make full use of the capacity of the pile bioreactor while ensuring adequate access of microorganisms, the thickness of the concentrated coating should be greater than 0.5 mm and less than 1 mm. This will generally result in a concentrate loading of about 9 to 30% by weight. Typically the concentrate loading on coarse ore will be about 10 to 15% by weight of the coarse ore substrates. However, this is generally sufficient to create the heat to raise the temperature of the pile to the optimum temperature of the end thermophiles capable of oxidizing iron and biolixing chalcopyrite at temperatures above 50 ° C. When coating the coarse ore with approximately 10% by weight of a chalcopyrite concentrate, the coated mineral will typically include at least 30 kg of exposed sulfur sulfur per tonne of ore. It will be appreciated by those skilled in the art, therefore, that the coarse ore should be coated with as much concentrate as possible and the concentrate should include as many sulfide minerals as possible to maximize the amount of sulfide ore particles exposed in the stack completed. For example, if a typical concentrate containing at least 30% by weight of sulfur in the form of sulfide is coated in the coarse ore at a load of 10% by weight, the coated mineral concentrate will contain at least 3% sulfur in the form of exposed sulfur, which translates to a potential fuel value of approximately 150,000 Kcal / ton of ore. On the other hand, if 15% of the concentrate is coated on the coarse ore support then the stack will contain at least 4.5% sulfur in the form of exposed sulfur, which has a fuel value of approximately 255,000 Kcal / ton of ore. Finally, if 30% by weight of concentrate can be coated on coarse ore substrates, the stack will contain at least 9% sulfur in the form of exposed sulfur, a fuel value of approximately 450,000 Kcal / tonne of ore. The use of coarse mineral substrates uniformly having a maximum particle size of 2.5 cm, and preferably less than 12.7 mm, ensures adequate exposure of the chalcopyrite mineral in the coarse ore support material to the oxidizing solutions that they contain ferric and cupric ions and microorganisms capable of converting ferrous ions to ferric ions that aid in the leaching process. The use of coarse support ore that is smaller than 2.5 cm and larger than 3.0 mm also results in a stack design that allows adequate loading of fuel values in the form of a sulfide ore concentrate in the stack while it ensures an adequate flow of liquid and air inside the pile and exposure of the sulfide mineral concentrate to the oxidant environment of the pile. Thus, when the thick mineral particles coated with concentrate having the above characteristics are stacked to form a pile, they provide a very large surface area bioreactor which is very efficient in terms of bioleaching of the coated concentrate. Most of the sulfide minerals highly exposed in the concentrate will be biooxidated in general in 30 to 90 days. However, in the case of more recalcitrant mineral sulphides such as chalcopyrite the leaching can be much slower in comparison. The sulfide mineral particle concentrate can be produced from fine materials generated by crushing the chalcopyrite mineral to a size of less than 2.5 cm. Typically this will be the portion of the mineral that is less than 3.0 mm. The sulfide mineral particles in this fine material fraction can be concentrated from the rest of the fine materials by flotation or separation by gravity or by a variety of other methods recognized by those skilled in the art. The removal of the fraction of less than 3.0 mm from the ore is beneficial because if too many fine materials are present in the pile they can limit the flow of liquid and air inside the pile. The fine ore could also consume unacceptable amounts of acid and thus lead to higher pH levels than the pile and more precipitation of jarocite and ferric. In addition to using the chalcopyrite concentrate produced from the fine material fraction of the ore, the coarser ore particles can be coated with chalcopyrite concentrate produced from other copper-bearing minerals. It may also be beneficial to mix concentrates of other sulfide minerals with the chalcopyrite concentrate for the reasons described above.

The chalcopyrite concentrates made for the casting process are generally separated from the other sulfide minerals such as pyrite. The separation process can be a variety of methods recognized by those skilled in the mineral processing art. The general purpose of this separation is to achieve a high copper content for the economical casting of the concentrate. Concentrates that are high in pyrite and therefore low in copper are less economical to process by smelting. The separation procedure that is used to achieve high concentrations of copper, however, increases the cost of producing copper. Total copper recovery also decreases. This is because while the concentration of copper to be achieved in the concentrate is higher, necessarily more copper will be lost at the end of the separation process. An advantage of the high temperature stack bioleaching process of the present invention, therefore, is that the concentrate added to the stack does not need to be as high in percent copper as is required for economical casting. As described above, the presence of pyrite can accelerate the leaching of chalcopyrite through galvanic interactions, furthermore, the biooxidation of pyrite in the pile also provides a source of heat that can help raise and maintain the temperature of the pile. on a scale of 60 to 80 ° C, which in turn will promote the growth of extreme thermophiles and the faster leaching of chalcopyrite.

Therefore, larger copper recoveries in general can be made from a chalcopyrite ore body with the present invention while simultaneously saving costs by not having to produce such a high grade concentrate. Although the stack has been described above as being constructed using coarse ore particles as support, other materials can also be used as support for the concentrate in the present invention. For example, the thick support material can be selected from the group consisting of rock, septum, slag and plastic. If the support mineral is rock, as will be appreciated by those skilled in the art, a variety of rocks can be used for coarse support, including volcanic rock, sterile rock and crushed copper ore. An advantage of using thick chalcopyrite mineral particles as the support material is that the chalcopyrite contained within this support material can be at least partially biooxidized during the process. Additionally, the coarse support material can be recycled a number of times through the process by removing the biooxidized concentrate and coating it with fresh concentrate, resulting in even higher recoveries of copper from the coarse support. In addition, after the coarse ore support is processed through the process one or more times, it can be comminuted and the remaining sulfide minerals contained therein separated using techniques known in the art to form a sulfide ore concentrate. This concentrate can be combined with other concentrates for coating on coarse ore support material and processed according to the invention. Sterile rock, such as granite, that contains a small amount of carbonate may be beneficial in helping to suppress the amount of iron removed in the impregnating leaching liquid. As the carbonate mineral in the rock reacts with the acid in the leaching process solution, it causes the local pH to increase resulting in the precipitation of iron. As a result, the concentration of copper in the final impregnation leach liquid collected from the stack and sent to the solvent extraction plant for copper recovery may be able to increase. This is because the fact that solvent extraction plants can typically handle only a maximum concentration of about 5 g / l of iron in the impregnating leaching liquid before special treatments must be performed to remove selectively the iron. Thus, without the precipitation effect caused by the carbonate mineral in the supporting rock, the impregnating leaching liquid should have lower copper concentrations than would otherwise be possible to ensure that the iron concentration does not exceed the limits of the solvent extraction plant.

Another preferred stack design for practicing the present invention is described in the U.S.A. 5,431, 717, which is incorporated by reference herein. According to this patent, a pile can be constructed by removing all of the fine material from the chalcopyrite material, for example the fraction of the mineral that is less than 0.3 cm, and then adding to the pile a concentrate having chalcopyrite. This can be achieved by distributing the concentrate over the top of the pile so that it migrates down through the pile during the bioleaching or simply mixes it with the rest of the mineral during pile formation without producing a uniform coating of the concentrate on the coarse ore before the pile formation. To fully utilize the heat generated from the exothermic oxidation reactions that will occur during biooxidation, the stack must be constructed in such a way to contain as much heat as possible but also to allow temperature control so that the optimum temperature of the The biooxidant microorganisms do not overdo it. This can be achieved by covering the stack with an insulating barrier layer 22 to contain the heat and water vapor. The insulating barrier layer 22 can be an awning, plastic sheets, fiberglass insulation, a layer of crushed rock, or any of the other insulating barriers known in the art. In the case of operations in cold climates it may be preferred that the pile be constructed within an annex of isolated walls to help maintain the heat.

In addition to covering the stack, the flow of leaching process solution from the emitters 24 down through the stack will transport heat from the top of the stack to the bottom of the stack. The movement of air through the pile will transport heat through the ore. Therefore, if the flow of liquid and air can be controlled separately, the heat generated from the process can be moved out of the stack either through the top of the stack in the form of steam or through the bottom of the pile in the form of hot liquid. Alternatively, the heat of reaction can be contained within the balanced stack the flow of liquid and air. The stack preferably contains one or more temperature monitoring devices such as thermocouples so that the profile of the stack temperature can be monitored continuously. The placement of several thermocouples through the stack will be preferred to better control the temperature of the stack. After the stack is constructed, a substantial portion of the stack needs to be heated to a temperature of at least 50 ° C, preferably at least 60 ° C and still more preferably at least 70 ° C. While the temperature of the stack 20 is higher, the biooxidation of the chalcopyrite will proceed more quickly. By substantial it means that finally at least 50% of the stack must reach a temperature at or above the target temperature. Preferably at least 80% of the stack will reach a temperature above the target temperature to maximize the copper recovery of the stack and the recovery speed. The stack 20 should be heated to temperature as quickly as possible. This will help ensure that the sulfide minerals exposed enough remain in the pile once it reaches the temperature to supply most, if not all, of the heat needed to keep the battery at temperature through high biooxidation. Chalcopyrite temperature in the pile. Typically heating the stack to temperature within a period of 45 will be adequate to meet this goal. However, the stack 20 is preferably heated to a temperature within a period of 30 days or less to minimize the total time for the procedure to be carried out and to maximize the concentration of exposed fuel values remaining in the pile to bioleate the chalcopyrite in the pile. Since the amount of heat lost from the stack 20 depends on the time, increasing the stack 20 to room temperature as quickly as possible will also help to minimize the amount of heat lost from the stack during the bio-oxidation process. The stack can be heated to temperature by a variety of methods. In the case of heap leaching operations in a cold climate or when sulfide minerals are insufficiently exposed to be added to the pile, an external source of heat such as hot liquid, steam or hot air can be added to the pile to start the procedure or to maintain the optimum temperature. For example, the hot leaching process solution can be pumped from the reservoirs 26 to the top of the stack 22 through the leaching process solution supply lines 28 and 30. The leaching process solution is then distributed over the top of the stack 20 through the pressure transmitters 24. Other means for distributing the leaching process solution which are known in the art can also be used, including baffle corrugators, sprinklers, oscillating plates, and flood. The advantage of pressure transmitters is that the amount of water lost due to evaporation is minimized. Additionally, the portion of the supply line 30 running along the top of the stack can be buried to further reduce evaporation and improve the insulation of the supply line 30 in situations where the solution of the leaching process It can be heated. Alternatively, the stack 20 can also be heated by pumping steam or hot air through the steam or hot air supply line 32 to perforated distribution pipes 34 buried at the bottom of the stack. The supply line 32 and the perforated distribution pipes 34 can also supply ambient air for purposes of increasing the oxygen and nitrogen levels in the stack as well as for removing the heat from the stack 20 in case of overheating. The cell must be inoculated with a culture that includes at least one thermophilic microorganism capable of biolixing sulfide minerals at a temperature above 50 ° C, and preferably above 60 ° C. This may occur before or after the battery reaches the temperature, or at any time during the biololing process to increase the amount of thermophilic microorganisms in the cell. A solution of the leaching process is also applied to the pile during the bioleaching step, typically at a rate of at least 72 I / m2 / day. The solution of the leaching process helps maintain the proper conditions inside the pile to bioleave the sulfide minerals and take away the soluble bio-oxidation products. In particular, as copper sulfide ores are bio-oxidized, the copper in these minerals dissolves in the leaching process solution, forming an impregnating leaching process solution. Once a portion of the pile reaches at least 50 ° C, the thermophilic microorganisms in that portion of the pile will become active and quickly begin to bioleer the exposed chalcopyrite and other sulfide minerals in that region of the pile. This will produce additional heat which in turn will help to increase the temperature of the surrounding regions in the stack above 50 ° C until finally a substantial portion of the stack is above 50 ° C, and preferably above 60 ° C. ° C. The current amount of the stack that is heated above the desired temperature will depend on the speed at which the heat is fed into the stack through the oxidation of the sulfide minerals and through other additions of heat to the stack, and the speed at which heat is lost from the pile by means of convection and radiation. If the bioleaching is carried out in such a way that at least 10 Kg of the sulfur in the form of sulfur per ton of solids in the cell is oxidized within a period of 210 days or less from the completion of the cell, a significant fraction of the heat required to maintain the pile at temperature while the chalcopyrite is biolixed into the pile can be obtained from the exothermic oxidation reactions occurring within the pile. Additionally, by having sufficient sulfide ore particles exposed within the stack as described above, it is possible to bioleave at least 50% of the copper sulphide minerals in the stack and thereby cause at least 50% of the copper in the the pile dissolves in the leaching process solution within a period of 210 days from the completion of the stack. In appropriately designed piles, it will be possible to extract at least 70%, and preferably more than 80%, of the total copper in a period of 210 days or less. In fact if enough chalcopyrite in the pile is found in particles having a size of less than 250 μm, and preferably less than 107 μm, it will be possible to achieve recoveries of more than 80 or 90% in 90 to 100 days. The use of thermophilic chemolithotropic microorganisms that biooxid chalcopyrite and other sulfide minerals make it possible to operate the cell at temperatures above 60 ° C and accelerate the rate of biooxidation of chalcopyrite. These microorganisms are defined as those that live at temperatures of more than 60 ° C, which derive their energy from inorganic elements, such as iron and sulfur, and obtain their carbon from the fixation of carbon dioxide. These organisms, represented by genera such as Sulfolobus, Acidianus, and Metallosphaera, are actually Archaea, but are frequently referred to as bacteria in the literature. Because the thermophilic microorganisms are able to survive on mineral sulphides in high temperature environments, those microorganisms are ideally suited for the process of the present invention, which requires the use of high temperature cell leaching and can use high temperature designs. pile with high concentrations of sulfur minerals that result in large amounts of excess heat. In addition to biooxidating sulphide minerals, many thermophilic microorganisms also oxidize ferrous iron and elemental sulfur. By oxidizing elemental sulfur, which is thought to contribute to a darkening of the chalcopyrite surface during bioxidation, the use of thermophilics can improve the leaching rate of this mineral by minimizing the amount of sulfur deposited on the surface of the chalcopyrite. that is biolixiviando. By oxidizing ferrous iron to ferric iron, these microorganisms also help maintain a high redox potential within the pile and allow additional ferric leaching of the sulfide minerals in the pile. Some examples of thermophilic microorganisms capable of oxidizing ferrous, sulfur and sulfur minerals are Acidianus brierleyi, Acidianus infemus, Metallosphaera sedula, Sulfolobus acidocaldanus, Sulfolobus BC, and Sulfolobus metallicus. These thermophilic organisms are able to leach the chalcopyrite concentrate and the mineral substrates of the preferred cell design in a period of less than 90 days at a temperature of 60 to 80 ° C. Other extreme thermophilics which are known in the art and which can grow and leach copper sulfides as well as other sulfide minerals within this temperature scale can also be used to practice the present invention. The cell 20 is preferably inoculated with a mixed culture of thermophilic containing two or more thermophilic. Although these microorganisms all survive at high temperatures, at low pH, and can use mineral sulphides as energy sources, they differ in attributes such as optimum growth temperature, affinity for and ability to leach particular minerals, and tolerance of solution components ( for example, salts). Also, due to the conditions inside the pile they can vary enough in terms of temperature, salt concentrations, etc., certain thermophiles can survive in some regions while others survive in other regions. Thus, when inoculated with a mixed thermophilic culture, the most potent species will dominate within the particular bioleaching conditions present within the pile or a region within the pile, resulting in the best possible leaching. As noted above, the preferred stack design is one that is more than 3 meters high and 10 meters wide. Batteries of this size will help to maximize heat retention in most cells. This is because the outermost ends of the stack will act as a heat insulator for the rest of the stack. Depending on how well the battery is isolated and the outside environment, however, the outermost ends of the battery may not reach a temperature of more than 50 ° C. By inoculating the pile with a combination of mesophiles and moderate thermophilics, it will therefore help in the bioleaching of the colder regions of the pile. Even though the amount of copper extraction from those colder regions will be less than the possible extraction within the higher temperature regions of the pile, the total extraction of the whole stack will be improved. Therefore, in a preferred embodiment of the practice of the present invention, in addition to inoculating the cell with one or more thermophilic cells, the cell is also inoculated with one or more mesophilic microorganisms and / or one or more more moderate thermophilic microorganisms. In a preferred method of the practice of the present invention, a substantial portion of the heat required to initially heat the stack at temperature is derived from the bioleaching sulphide minerals contained within the stack. If the stack 20 contains sufficient readily exposed biodegradable sulfide minerals such as pyrite, arsenopyrite, chalcocite, and covellite, then the stack 20 can be heated, at least partially, using the fuel values of those exposed sulfur ores, easily biooxidated. This is achieved by inoculating the cell with a culture containing one or more mesophilic microorganisms capable of biooxidating sulfide minerals that are well known in the art. As biooxidation proceeds, the heat generated from the exothermic oxidation reactions of readily exposed biooxidated sulphide minerals will begin to heat the stack. A significant portion of the heat required to heat the stack 20 to the temperature of 50 ° C, or preferably 60 ° C, can be supplied by bio-oxidation if enough sulfide ore particles are bio-excreted to oxidize at least 10 kg of sulfur in the form of sulfur per ton of solids in the cell within a period of 45 days or less, and preferably 30 days or less. In this way, in order to heat the stack 20 using the heat released through the biooxidation of sulfide minerals, the stack 20 must be constructed to contain at least one easily exposed bio-sulfur mineral in amounts sufficient to supply the heat with at least 10 Kg of sulfur in the form of sulfur per ton of solids. If the heat released through biooxidation is to be used as the heat source to heat the stack 20 to temperature, then the concentration of chalcopyrite exposed in the stack is also preferably sufficient to supply at least 10 Kg of sulfur in the form of sulfur to the pile. This is so that once the pile is heated to temperature, there will be enough sulfur minerals remaining in the pile to help maintain the temperature above the target temperature for biooxidation of the chalcopyrite to proceed over a period of about 60 to 150 days. The stack 20 will typically be heated in one way by steps if the heat released through the biooxidation is to be used to heat the stack. First, the pile is inoculated with one or more mesophiles that are capable of biooxidating sulphide minerals. Then it is inoculated with a culture comprising one or more moderate thermophiles. Alternatively, these inoculations may occur simultaneously if desired. Mesophiles typically operate within a temperature range of 25 ° C to 40 ° C, while moderate thermophiles typically operate on a scale of 40 ° C to 55 ° C. In this way the mesophiles can be used to heat the battery to a temperature of 40 ° C. Once the pile reaches a temperature of 40 ° C, the mesophiles will become less active. However, if the battery has been inoculated with moderate thermophiles, those microorganisms will become active as the battery temperature reaches approximately 40 ° C. Moderate thermophiles can then continue to oxidize the sulfide mineral particles exposed in the pile until a substantial portion of the pile reaches a temperature of 50 ° C to 55 ° C where moderate thermophiles begin to become less active. At this point, however, the growth of extreme thermophiles is favored. As a result, the extreme thermophiles within the cell will become active and will begin to oxidize additional sulfur minerals in the cell, further increasing the temperature of the cell. At temperatures above 50 ° C and especially above 60 ° C, the rate of biooxidation of chalcopyrite in the pile will increase rapidly due to the fact that the passivation layer that inhibits bioleaching at lower temperatures tends to degrade at temperatures above 50 ° C. Any of the mesophilic or moderately thermophilic microorganisms that are known in the art to be capable of biooxidating sulfide minerals can be used in the present invention. Examples of mesophiles that can be used in the practice of the present invention include Thiobacillus ferrooxidans, Thiobacillus thiooxidans, Thiobacillus organoparus, Thiobacillus acidophilus, and Leptospirillum ferrooxidans. Examples of moderate thermophiles that can be used in the practice of the present invention include Sulfobacillus thermosulfidooxidans, Thiobacillus caldus, and Thiobacillus cuprinus. The pile is irrigated with a solution of leaching process (PLS) through the period of biooxidation. The leaching process solution typically includes sulfuric acid and iron in ferric and / or ferrous form. The leaching process solution may also contain nutrients to aid the growth of biooxidant microorganisms. However, the nutrients necessary for the growth of microorganisms and to metabolize the sulfide minerals in the pile may be present within the mineral that is being biolixed. Chalcopyrite leaching can consume acid and cause the pH of the pile to increase. The increased pH can lead to jarosite formation and ferric precipitation. To prevent this precipitation from becoming extensive and delaying the leaching process, the leaching process solution must be maintained below a pH of 1.5, especially once the pile reaches a temperature above 50 ° C and biooxidation of chalcopyrite begins to proceed quickly. To further minimize the precipitation of ferric and jarosite, the ferric concentration should also be maintained below 3 g / l, especially once the pile has reached a temperature above 50 ° C. Nutrient salts should also be kept low after the pile temperature has risen above 50 ° C, especially in potassium and ammonium sulfate, both of which can increase jarosite formation. The addition of a small amount of chloride (1 to 5 g / l as NaCl) can help keep the ferric in solution and improve the leaching of copper on iron. In this wayIt may be desirable to use a chloride medium to bioleave the cell, especially after the cell temperature has risen above 50 ° C. However, if a chloride medium is used for the process solution of leaching, thermophilic microorganisms that exhibit resistance to chloride must be selected.The flow rate at which the cell is irrigated with the leaching process solution will depend on a number of factors.Two of the main functions of the process solution of Leaching is to provide acid and remove copper that has been dissolved during the bioleaching process, as a result, peak point flows will typically occur at the beginning of the process to reduce the pH of the pile to an adequate level that is conductive to the bioleaching process. Once the solution of the pile is below a pH 2.0, preferably 1.8, the stack is man-conditioned it was adequate and the appropriate conditions should exist within the bioleaching pile. The rate of application of the leaching solution of the process will also tend to be higher once the cell reaches the optimum temperature for chalcopyrite biooxidation. As the pile temperature is raised to a temperature suitable for chalcopyrite biooxidation, the chalcopyrite in the pile will begin to rapidly biooxide. Because the biooxidation of chalcopyrite consumes acid, the leaching solution of the additional process will typically need to be added in order to maintain a low pH environment suitable for further biooxidation. The chalcopyrite biooxidation rate will tend to be the highest for a period shortly after the pile is raised to the optimum temperature. As a result, the rate of application of the process leaching solution will also tend to be high during the period when the chalcopyrite biooxidation proceeds rapidly in the heap. As the copper sulphide minerals in the pile are biooxidized, the copper will dissolve in the leaching solution of the process, thus forming a leaching solution of the pregnant process. In determining the proper application rate of the process leaching solution, therefore, it is also advisable to use a flow rate that ensures a copper concentration greater than 1 g / l, preferably greater than 2 g / l, and very preferably greater than 5 g / l. This is particularly true, once the biooxidation of chalcopyrite, which is the primary copper sulphide mineral in the pile, begins to proceed quickly. However, the flow rate of the leaching solution of the process must be adequate to ensure that the final ferric iron concentration in the leaching solution of the pregnant process is less than 5 g / l and preferably less than 3 g / l . Although concentrations up to 5 g / l of ferric can typically be handled in solvent extraction processes known in the art, concentrations greater than about 3 g / l will tend to result in excessive ferric iron and jarosite precipitation in the environment. of high temperature of the pile. Finally, the flow rate of the leaching solution of the process is preferably selected to meet the above goals with the lowest possible application rate. By keeping the flow rate of the process leaching solution at the lowest possible level to achieve the above goals, the amount of heat lost from the stack can be minimized, thus minimizing the potential need for heat application externally maintain the battery at the optimum temperature during the biooxidation of chalcopyrite. With the above goals in mind, the process leaching solution will typically be applied at a rate of at least 72 l / m2 / day, and preferably at a speed of at least 144 l / m2 / day. For piles having the preferred dimensions mentioned above, the leaching solution of the process will generally be applied at a rate of about 300 to 600 l / m2 / day. The application of the leaching solution of the process does not have to be continuous. The present invention can be carried out with irrigation followed by periods of drying or resting. Although during the drying period, or drying cycle as it is sometimes called in the art, no process leaching solution is applied, the cell is not allowed to dry completely during this period of rest. Rather, the pile will typically continue to produce drains during the entire drying or resting period. As the pregnant process leaching solution is drained from the cell, it will collect in the drain 35. From the drain 35, the leaching solution of the pregnant process can be drained by gravity or pumped to the tanks 26 via the tube 36. Preferably, the leaching solution of the pregnant process is transferred to tanks 24 as quickly as possible to minimize the heat losses of the pregnant process leaching solution. To further minimize the heat losses of the pregnant process leaching solution, the tanks 24 can be isolated. Once the concentration of copper in the leaching solution of the pregnant process reaches a desired level, the pregnant process leaching solution is sent to a solvent extraction plant 38 for recovery of the copper. The design, construction, and operation of solvent extraction plants are well known in the art and need not be described in more detail herein. The elemental copper 44 can be recovered from the pregnant removal solution 40 leaving the solvent extraction plant using an electrolytic extraction cell 42 as is well known in the art. After the copper is removed from the pregnant removal solution in the electrolytic extraction cell 42, the fresh removal solution 46 is recycled to the solvent extraction plant 38 for recharging. After the copper values in the leaching solution of the pregnant process have been removed in the solvent extraction plant 38, the leaching solution of the replenished process 48 can be recycled to the stack to pass through the stack again. . Because most solvent extraction plants are operated at a temperature below 50 ° C, the pregnant process leaching solution that is supplied to the solvent extraction plant typically needs to be cooled to a temperature suitable for the solvent extraction plant. On the other hand, the leaching solution of the cooled process is preferably heated to a temperature that is closest to the operating temperature of the stack before its reapplication to minimize the purges of heat in the system. Therefore, in a preferred method for practicing the present invention, the leaching solution of the cooled process 48 and the leaching solution of the pregnant process are passed through separate sides of a heat exchanger 50 before supplying the leaching solution from the collected procedure to the solvent extraction plant. In this way, the heat can be removed from the leaching solution of the pregnant process collected in preparation for its treatment in the solvent extraction plant 38 and can be transferred to the leaching solution of the cooled procedure 48 before its application to the pile, thus minimizing losses of system heat. After passing through the heat exchanger 50, the leaching solution of the process can be pumped to the top of the stack 20 through the supply line 30. The fresh water supply 52 can be used to compensate for the losses of water in the system that are due to evaporation . In addition to using solvent extraction to recover copper from the leaching solution of the pregnant process, other techniques known in the art can also be employed, including copper cementation and ion exchange. Ion exchange procedures offer an advantage due to the fact that they can be operated at higher temperatures than solvent extraction plants. As a result, less heat will be lost from the systems because the need to cool the leaching solution of the process before the copper recovery can be effectively eliminated. Copper cementation offers a similar advantage. However, the copper purity produced through copper cementation is not as high as that produced through solvent extraction followed by electrolytic refining. In addition, due to the fact that copper in solution is replaced with iron during the cementing process, the use of copper cementation would also require frequent treatments of the process leaching solution to remove excessive iron concentrations to avoid excessive precipitation. Figures 2A-2D schematically illustrate one way of practicing the present invention in a period to more effectively take advantage of the heat values produced through the oxidation of the sulfide minerals in the cell. Essentially an initial stack 20 is prepared as described above. After the stack 20 has reached an optimum temperature for the biooxidation of the chalcopyrite contained therein, approximately 60 to 70 ° C, and the oxidation of the chalcopyrite is proceeding rapidly therein, a second stack or step can be added. 54 at the top of the stack 20. The heat emitted from the currently active stack 20 will help to heat the stack 54 to a temperature at which the biooxidation of the chalcopyrite within the stack 54 can proceed rapidly. Again, once the biooxidation of the chalcopyrite at high temperature is proceeding rapidly in the stack 54 and a substantial portion of the stack 54 has reached a temperature of about 60 to 70 ° C, a third stack or step 56 can be constructed at the top of the stack 54. Again, the heat emitted from the active stack 56 will help to heat the stack 56 to a temperature at which the biooxidation of the chalcopyrite at high temperature in the stack can proceed. This procedure can be repeated again and again with as many stacks or steps as desired. Figure 2D, for example, illustrates a fourth stack or step 58 being constructed on top of the third stack 56. Another advantage of implementing the present invention in a series of stacked stacks or steps is that according to the fuel values of sulfide ore in the lower piles or steps are reduced through the bio-oxidation process, the temperature of these lower piles will begin to decrease. However, the heat from the upper piles or steps will help keep the lower piles at a temperature sufficient for the biooxidation of chalcopyrite at high temperature to continue for a longer period than would otherwise be possible. Also, even if the values of sulfide ore exposed in the lower piles are reduced to the point that, even with the additional heat being supplied by the upper piles, the pile can not maintain a temperature high enough for the thermophilic microorganisms to remain active, biooxidation can continue with mesophilic and thermophilic microorganisms. In addition, the high concentrations of ferric iron that are produced in the upper piles will also help in the continuous leaching of the copper sulphide minerals in the lower piles or steps. ThusBy practicing the invention in a series of piles or steps stacked as described above, it may be possible to achieve higher total copper recoveries of the ore. Having described the preferred embodiments of the invention, several aspects of the invention are further extended in the following examples. Said enlargements are intended to illustrate the invention described herein, and not to limit the invention to the examples set forth.

EXAMPLE 1 Samples of recalcitrant chalcopyrite ore and concentrate from the San Manuel Copper Mine in Arizona were used to evaluate the use of thermophilic microorganisms to bioleave chalcopyrite into a pile process. To stimulate the heap leaching process, a column test was performed. A total of 491.2 g of chalcopyrite feed concentrate from the smelter were applied as a coating on 5 kg of ore from the same San Manuel mine. Because the concentrate was melting grade, the sulfide mineral particles within the concentrate consisted almost entirely of chalcopyrite. Analysis of the smelter concentrate showed that it contained 28.5% copper and 27.5% iron and 33.6% sulfur as sulfur. Therefore, without considering the sulfide minerals exposed in the lump ore support, the ore coated with the concentrate consisted of at least 3% sulfur in the form of exposed sulfur.

The support rock in which the concentrate was coated was prepared by size separation of crushed San Manuel ore. The crushed ore of minus 19 mm was separated into a fraction of minus 3.2 mm, a fraction of 3.2 to 6.4 mm and a fraction of 6.4 mm to 12.7 mm and a fraction of more 12.7 mm. The fractions of 3.2 to 6.4 mm and 6.4 to 12.7 mm were used in equal weights (2.5 kg each) as support rock for the melt concentrate. Fractions of less than 3.2 mm and over 12.7 mm were not used in the test. The exclusion of the fraction of less 3.2 mm ensured that the pile had good air flow characteristics. Mineral analysis of 3.2 to 6.4 mm indicated that it contained 0.549% copper and 2.37% iron. Analysis of the ore from 6.4 to 12.7 mm contained 0.523% copper and 2.38% iron. The mixture of the two sizes of chalcopyrite ore was coated with the high grade copper concentrate by rotating the ore in a drum at approximately 30 rpm while spraying 10% sulfuric acid. After the support rock was moistened, the dry concentrate was separated on the rock support roll. More liquid was sprayed into the mixture until the mineral particles in pieces were coated with the concentrate. The final water content of the mineral particles in pieces coated with the concentrate was approximately 3% by weight. The mixture of the concentrate and mineral support in pieces was then placed in an 8.0 cm glass column to simulate a pile. The column was wrapped with electricity-resistant heating tape to isolate the column and help control the temperature. The temperature was monitored by a thermo pair attached with tape to the outside of the glass tube and a glass thermometer at the top of the mineral at the top of the column. Air and liquid were introduced in the upper part of the column. The air was heated by bubbling through the heated water and then through a stainless steel tube heated to the top of the column. The liquid was collected from the bottom of the column in a heated beaker. Air exiting from the bottom of the column was bubbled through the liquid in the heated beaker. This was done to keep any bacteria in the solution alive and active. The flow velocity of the liquid pumped to the top of the column was at least one liter per day. The pH of the solution was measured once per day and adjusted to a pH between 1.1 and 1.3 with sulfuric acid. The levels of copper and iron in solution were determined once or twice a week. The solution was removed from the system and replaced with a new solution containing the nutrient mixture. This was done to prevent the solution from raising its copper content too much and becoming toxic to the microorganisms. The liquid medium introduced to the top of the column contained 0.16 g / l, NH4CI, 0.326 g / l Mg Cl2 6H20, 0.1 g / l K2 HP04, 0.1 g / l KCl, plus 1 ml / l of a trace mineral solution listed in table 1 below. As will be appreciated by those skilled in the art, the concentration of nutrients in the liquid medium was lower than that typically found in the 9K salt generally used in connection with biooxidation. The lower concentration of nutrients as well as the chloride medium and low pH were used to minimize ferric iron precipitation such as jarosite and any concomitant plugging of the air or liquid flow channels that might otherwise result.

TABLE 1 Trace mineral solution g /? MnCl2 x 4 H2O 1.8 Na2B4O7 x 10 H2O 4.5 ZnSO4 x 7 H2O 0.22 CuCl2 x 2 H2O 0.05 Na2MoO4 x 2 H2O 0.03 VOSO4 x 2 H2O 0.03 CoSO4 0.01 The temperature of the column was first maintained at 35 ° C and inoculated on day three with 25 ml of Thiobacillus ferrooxidans, which were originally initiated with strains ATCC 19859, 14119, 23270, and 33020 of the American Type Culture Deposit in Rockville, Maryland. The concentration of bacteria was approximately 108 bacteria per ml. On day four the column temperature was increased to 40 ° C. On day five the temperature was increased to 45 ° C. On day seven the temperature was increased to 65 ° C and re-inoculated with 25 ml of a mixed culture of Thiobacillus ferrooxidans, Leptospirillum ferrooxidans, and cultures of Thiobacillus thiooxidans (strains ATCC 8085 and 15494) and a culture of moderate thermophiles isolated of a sample of ore from the Atlanta Mine of the Ramrod Gold Co. in Idaho. On day 14 the column temperature was increased to 70 ° C. The column was then inoculated with a previously frozen mixed culture of extreme thermophiles comprising Acidianus brierleyi (strain DMS 1651), and Sulfolobus acidocaldarius (strains ATCC 33909 and 49426). The DMS strains were obtained from the Deutsche Sammiung von Mikroorganismen deposit center in Braunschweig, Germany. As biooxidation within the column had not increased as much as desired, although the column was inoculated with a mixed culture of thermophiles on day 14, on day 40 a fresh culture of extreme thermophiles including Acidianus brierleyi (strains DSM 1651 and 6334) , Acidianus infemus, (strain DSM 3191) Metallosphaera sedula (strain ATCC 33909) Sulfolobus acidocaldarius (strain ATCC 49426) and Sulfolobus metallicus (strain DMS 6482) were added to the column. The temperature of the column and the heated container of circulating solution were maintained between 60 and 75 ° C for the remainder of the 93 day experiment. The progress of the experiment was monitored by the solubilization of copper. The total copper in the concentrate and in the mineral support was estimated by analyzing a divided sample of each size fraction of mineral support used and the chalcopyrite concentrate. The estimated percentages leached for both iron and copper as the experiment progressed are listed in table 2 and plotted against time in figure 3. The concentrations of both iron and copper in the solutions in circulation are also listed in the table 2.

TABLE 2 Chalcopyrite smelter concentrate in chalcopyrite mineral Days with Fe g / l Con. Of Cu g / l% of Fe leached% of Cu leached 3 0.456 0.508 0.66 1.13 6 2.16 0.68 0.89 2.78 12 4.856 0.56 4.89 2.46 17 4.032 0.444 8.91 3.08 24 1,772 0.224 11.81 3.79 31 1.3 0.452 15.32 6.59 1,164 0.34 17.31 7.30 39 0.716 0.24 18.57 7.94 45 0.684 0.976 20.70 14.51 47 0.643 1.732 21.78 20.48 51 2.26 5.84 23.60 27.53 54 1.924 5.1 25.21 34.12 58 1.86 4.764 26.94 40.80 60 1.352 3.48 27.81 44.23 61 0.864 2.376 28.26 46.33 65 0.876 2.64 39.52 52.33 69 0.912 2.832 31.01 59.39 72 0.669 1.86 32.22 64.52 76 0.62 1.512 34.03 71.23 79 0.652 1.232 35.72 76.09 81 0.812 1.268 36.62 78.22 82 0.464 0.736 37.12 79.44 83, 0.584 0.832 37.59 80.42 86 0.924 1.228 38.66 82.57 90 0.464 0.532 39.96 84.83 93 0.655 0.632 40.97 86.33 The rate of copper leaching showed a remarkable increase 16 days after inoculation with the first culture of extreme thermophiles. The second inoculation of extreme thermophiles increased the rate of copper leaching even more on day 47 of the experiment. The leaching rate did not decrease until after day 86 of the experiment when the estimated total leached copper was 82.6%. After another week the column was dismantled so that each fraction could be analyzed for the degree of copper leaching. The material in the column was separated into four size fractions. A fraction was less than 0.14 mm and weighed 224.6 g. This fraction of size was considered as the remaining chalcopyrite concentrate. Another size fraction was greater than 0.14 mm and less than 3.2 mm and weighed 340.8 g. This scale of material size was never put on the column and was believed to be the result of the decomposition of 2.5 kg of ore from 3.2 mm to 6.4 mm that was put into the column at the beginning. The amount of material remaining on the size scale from 3.2 mm to 6.4 mm was 2, 108 g. The remaining 6.4 mm to 12.7 weight was 2.304 g. The analysis of each size fraction was used to calculate the degree of chalcopyrite leaching for both the concentrate and the ore. The analysis of the 224.6 g of smelter concentrate showed 3.24% copper and 18.5% iron or a total of 7.28 g of copper and 41.6 g of iron. The original 491.2 g of copper concentrate was 28.5% copper and 27.5% iron and therefore contained 140.0 g copper and 135.1 g iron. The percentage leached calculated was 94.8% for copper and 69.2% for iron. The copper and iron leaching estimate of the ore from 3.2 to 6.4 mm used as support was based on 0.355% copper and 4.19% iron remaining in the size fraction of 0.14 mm to 3.2 mm and 0.305% copper and 2.08% iron remaining in the size fraction of 3.2 mm to 6.4 mm. The total remaining copper and iron was 7.64 g and 58.1 g, respectively. Therefore, the leached percentage calculated for the original 3.2 to 6.4 mm size fraction was 44.3% for copper and 1.8% for iron. The high level of iron remaining suggested that this size fraction contained some of the concentrated or precipitated iron contained. The largest fraction was 0.353% copper and 2.27% iron. The total remaining copper was 8.13 g and the total remaining iron was 52.3 g of the original 13.08 g of copper and 59.5 g of iron. The amount of copper and iron leached for this size fraction was 37.8% and 12.1%, respectively. The low level of iron elimination indicated that some of the iron had precipitated. The calculated total copper that was leached from the mixture of chalcopyrite mineral and concentrate was 86.2% and 40.1% for iron. This coincided very well with the degree of leaching estimated by analysis of the solution in circulation EXAMPLE 2 The experiment described in Example 1 was repeated using 486.8 g of the same melt-grade concentrate that was used in that experiment.

The support rock in which the concentrate was applied comprised 2.5 Kg of ore from 3.2 mm to 6.4 mm and 2.5 Kg of ore from 6.4 mm to 12.7 mm. The mixture of the two sizes of chalcopyrite ore was coated with the high grade concentrate by rotating the ore in a drum at approximately 30 rpm, while being sprayed with water. Water was used in this experiment to show that acid could then be added. The dried concentrate was spread over the plurality of moistened substrate mineral that was stirred, as was done in Example 1. The final water content of the coated ore particles was about 3% by weight. Also, as in Example 1, without considering the sulfide minerals exposed in the crude support, the concentrate-coated mineral contained approximately 3% sulfur in the form of exposed sulfide. The substrates coated with concentrate were placed on an 8.0 cm glass column. The column was wrapped with electricity-resistant heating tape to isolate the column and to help control its temperature. The temperature was monitored by a thermo pair attached with tape to the outside of the column and a glass thermometer was placed in the center of the mineral at the top of the column. An additional 100 g of the unlined mineral mixture from 3.2 mm to 12.7 mm was used to cover the ore material coated with the concentrate. This uncoated mineral formed a layer approximately 5 cm thick that covered the coated substrates and acted as an insulating layer to prevent heat loss in the upper part of the bed. Air and liquid were introduced into the upper part of the column as in Example 1. The air was heated by bubbling through heated water and a stainless steel tube heated as in Example 1. The liquid that came out of the column was contained in a heated beaker as described in example 1. During the first 3 days the column temperature was maintained at 35 ° C while two liters of 5% sulfuric acid were circulated through of the column at a flow rate of more than one liter per day. The high acid concentration rapidly adjusted the pH to below 1.0. On the third day the temperature was increased to 70 ° C and the same nutrient mixture as described in example 1 was used to replace the circulating acid solution. After approximately four hours the column was inoculated with the same culture of extreme thermophiles, namely Acidianus brierleyi (strains DSM 1651 and 6334), Acidianus infernus (strain DSM 3191), Sulfolobus acidocaldaríus (strain ATCC 49426), and Sulfolobus metallicus * (strain DSM 6482), which were used to inoculate the column of Example 1 on day 40. Seven days later (day 10 from the beginning) the column was inoculated with a culture of microorganisms recovered from the column disassembly in the example 1. The bacteria can be recovered after the biooxidized concentrate is washed from the substrate. The suspension of the withdrawn concentrate is allowed to stand overnight. The cloudy liquid can have high levels (107 or more bacteria per ml) of bacteria that can be used to inoculate directly or that can be centrifuged to form even higher concentrations of bacteria. About a quarter of the bacteria recovered in this manner from the column experiment in Example 1 were used to inoculate the repeat column in this example on day 10. After the initial treatment with 5% sulfuric acid, the pH of the solution The leaching of the procedure added to the top of the pile was maintained between a pH of 1.1 and 1.3. The pH of the spent solution was generally between 1.3 and 1.6. The levels of copper and iron in the solution were determined once or twice a week. The solution was removed from the system and replaced with a new solution containing the nutrient mixture. This was done for the same reason as in example 1, namely to keep the toxic level of copper low until the microorganisms had time to adapt to the high copper concentrations. The main difference between the experiment in Example 1 and this is the early use of high temperature (70 ° C) and early inoculation with a fresh culture of extreme thermophiles. The degree of leaching of copper and iron was estimated by determining the concentrations of copper and iron in solution, which are plotted against the time in Figure 4. The earliest leaching start, which in the example hereof is the result of the earliest inoculation, demontes the benefit of using extremely thermophilic microorganisms when leaching recalcitrant chalcopyrite. The material in this column was separated into fractions of size. The fraction smaller than material of 0.14 mm weighed 315.4 g and contained 2.74% copper and 12.7% iron by analysis. The original 486.8 g of copper concentrate contained 138.7 g of copper and 133.8 g of iron. The percentage leached calculated was 93.8% for copper and 70.0% for iron. The copper and iron leaching estimate for ore from 0.14 to 12.7 mm used as support was based on the 15.36 g of copper and 78.9 g of iron remaining in this size fraction. The percentage leached calculated was 43.8% copper and 34.9% iron.

EXAMPLE 3 A mixture of chalcopyrite concentrate from the San Manuel copper mine in Arizona was used to perform a control experiment at 35 ° C in a similar column test. A total of 391.8 g of smelter grade concentrate containing 28.8% copper and 27.3% iron was applied as a coating to a plurality of granite support rocks. The 3920 g of support rock that was coated had no copper ore and was between 6.4 and 12.7 mm in size. The sample of granite support rock had a small amount of carbonate and tended to cause some iron precipitation. The method of coating was similar to the method used in Examples 1 and 2 except that approximately 110 ml of solution containing bacteria was used to wet the support rock before applying the dry concentrate. Since the concentrate contained more than 30% sulfur in the form of sulfide, the support rock coated with the concentrate contained approximately 3% sulfur in the form of exposed sulfide. The mixture of concentrate and raw granite support rock was placed in a 7.6-cm plastic column. The column was wrapped with electricity-resistant heating tape to isolate the column and to help control the temperature. In this example, the temperature was maintained at 35 ° C throughout the experiment. Air and liquid were introduced in the upper part of the column. The liquid was collected from the bottom of the column and the pH was adjusted before reapplying it to the top of the column. The pH of the spent solution varied between 1.37 and 1.76 and the pH of the reapplied solution was between 1.2 and 1.5. The copper and iron levels in the solutions were determined at least once a week. The solution removed from the system was replaced with a new pH adjusting nutrient solution. The medium contained 1.0 g / l NH 4 SO; 0.2 g / l MgSO4 7H20 0.02 g / l K2HPO; 0.03 g / l KCl. This experiment did not use the chloride nutrient solution used in Examples 1 and 2. The nutrient solution used in Examples 1 and 2 was used to minimize the amount of iron precipitation at the highest leaching temperature of 70. ° C. The concentration of bacteria in the solution that was used to wet the lump ore was approximately 108 bacteria per ml and was from the same mixed culture used in the first inoculation of example 1. On day 44, 990 ml of 10.08 g / l of Ferric solution was added to the five liters circulating solution to increase the iron level to approximately 2 g / l. The iron level had been low (less than 1 g / l) during the first 44 days. After the addition of ferric iron the iron levels remained above 1 g / l until after day 85. The copper level of the solution remained above 1 g / l after the first 10 days. The control experiment was carried out for 100 days. The Eh exceeded 0.6 volts after 50 days. The high Eh indicated that bacterial growth and bioleaching were in progress during most of the 100-day experiment. However, only 20% of the copper was leached after 60 days, and only 25.2% after 100 days. The experiment stopped after 100 days. The material of the column was separated into a size fraction of minus 0.14 mm and a size fraction of plus 0.14 mm. Each fraction was analyzed to determine the remaining copper in the system. The weight of the granite support rock increased to 4087.2 g and had collected 0.928% copper. The weight of the concentrate had decreased to 218.6 g and the copper content was 19.4%. The total copper remaining in the column after 100 days of bioleaching was 80.34 g or 71.2% of the original 112.8 g. The copper analysis of the solution estimated that 25.3% had leached out of the column. This compares well with 28.8% calculated by final copper analysis.

The degree of leaching of copper and iron was estimated by determining the concentrations of copper and iron in solution. The estimated leaching degree for copper and iron are plotted against the time in Figure 5.

EXAMPLE 4 A concentrate of the fraction of less 3.2 mm of the San Manuel ore was made. The concentrate was made by grinding the ore to pass 0.107 mm. The sample of ground ore was floated to form a sulfide concentrate. The flotation was carried out in small batches of 500 g each in a laboratory Wemco flotation cell. Prior to flotation, the sample of ground ore was adjusted to a pulp density of 30%. Then the pH was adjusted between 7 and 9 with NaOH. Potassium amylxanthate was added as a collector at approximately 100 g / metric ton and mixed more than 5 minutes before 50 g / metric ton of Dowfroth D-200 was added and mixed for another 5 minutes. Finally, air was introduced to produce a sulfide concentrate containing 8.5% copper and 30.4% iron and 35.8% sulfur by weight. Therefore, this concentrate contained almost twice the amount by weight of pyrite as well as chalcopyrite. A plurality of coated substrates was made by coating 200 g of the sulfide concentrate in 2,000 g of granite rock of plus 6.2 mm minus 12.7 mm. The concentrate was added as a dry powder to the support rock moistened in a drum rotating about 30 rpm. The dried concentrate was spread on the stirring support rock. More liquid was sprayed into the mixture until the raw support rock was coated with the concentrate. The final water content of the lumpy mineral particles was approximately 3% by weight. Also, the mineral coated with concentrate contained approximately 3.2% sulfur in the form of exposed sulfur. The plurality of coated substrates were placed on a 5 cm glass column. The column was wrapped with electrical-resistant heating tape to isolate the column and help control the temperature control. The temperature was monitored by a thermo pair linked by appointment to the outside of the glass tube and by a glass thermometer placed in the upper central part of the column. Air and liquid were introduced in the upper part of the column. The air was passed through heated water before it entered the column. The liquid was collected from the bottom of the column but was not heated as was done in example 1. The flow velocity of the pumped liquid to the top of the column was at least 0.5 liters per day. The pH of the solution was measured once a day and adjusted to a pH between 1.1 and 1.3. The first solution used in this experiment was a chloride nutrient mixture different from that used in Example 1 above. In this experiment the first nutrient solution comprised 2.03 g / l NH4CI; 0.08 g / l KCl; 0.04 g / l K2HP04; 0.35 g / l MgCl-6H20. On day four it was replaced with a solution that was the same except that it also contained 2 g / l of ferric iron made with ferric sulfate. This solution was removed again and replaced on day seven. The new solution also contained 2 g / l of ferric iron. On day 29 it was changed again and replaced with the chloride nutrient solution containing 2 g / l of ferric iron. This solution was again circulated until day 63 when a liter of the two liters in the beaker was replaced with fresh chloride nutrient solution. Another liter was removed and replaced on day 65 with the chloride nutrient solution. No ferric iron was added to the solution on days 63 and 65. One liter of solution was replaced with fresh chloride nutrient solution on days 74, 77, 81 and 84 and 91. The column experiment was stopped after day 93 The biolixed material of the column was separated into a size fraction of minus 0.14 mm and a fraction of more 0.14 mm. Each fraction of size was analyzed for copper, iron, and sulfur. The temperature of the column was first maintained at 35 ° C and inoculated with the same bacterial culture as was used to initially inoculate the column in Example 1. On day seven the temperature was increased to 40 ° C and returned to Inoculate with the same mesophilic culture. The following day the temperature was increased to 45 ° C and the column was inoculated with moderate thermophiles isolated from a sample of ore from the Atlanta mine of Ramrod Gold in Idaho. On day 10 the temperature was increased to 60 ° C. On day 11, 25 ml (108 bacteria per ml) of the same culture of mesophilic bacteria were added to the beaker of exhausted non-heated solution. The beaker was re-inoculated with 25 ml of 108 bacteria per ml on days 13 and 15. On day 18, the column was inoculated with the same previously frozen mixed culture of extreme thermophiles as was used to inoculate the column in Example 1 on day 14. The beaker of spent solution was also inoculated on day 30 and day 45 with mesophilic bacteria. This column experiment was never inoculated with the fresh culture of extreme thermophiles used in Example 1 on day 40. The estimated percentage of leached copper and iron was plotted against the time in Figure 6. The use of granite rock as Support may have caused excessive iron precipitation, due to its carbonate content. The estimated percentage of iron leaching was not above zero until after day 60. This precipitation may have limited the degree of copper leaching in this experiment as well. However, a benefit of not having iron leaching during the process is that a purer pregnant leach solution is produced for the solvent extraction plant. The final copper analysis indicated that 82.5% of copper had been leached from the concentrate. The amount of copper remaining in the 276 g of the material of minus 0.14 mm was 0.916%. The weight of the concentrate had increased in precipitation and loss of support rock. During the experiment the 2,000 g of granite support rock lost 140.8 g.

The analysis showed that 28.7% of the iron was removed and that 45.2% of the sulfur in the form of sulfur was biooxidated. The microscopic analysis of the water used to wash the coated concentrate from the support rock showed a higher number (more than 107 microorganisms per ml) of extreme thermophiles.

EXAMPLE 5 Another column experiment was carried out at the same time as the experiment described in example 4. This experiment was the same with one exception. The difference was that 10 g of fine powder graphite was mixed with 200 g of flotation concentrate by volume. This was the same concentrate used in Example 4, and the column was prepared in the same manner as in Example 4. The inoculations and pH adjustments were also the same as in Example 4. The estimated percentage of copper and iron leached graph against time in figure 7 for this example. The column experiment was continued for 93 days. The results of this column experiment indicated 89.8% copper leaching by analysis of spent solution and 89.0% by analysis of the material removed from the column. Iron leaching was also low in this experiment and was also thought to be due to precipitation caused by the carbonate in the granite support. Analysis for iron and sulfur in the form of sulfur indicated 18.6% iron removal and 53.9% sulfur biooxidation. Added and graphite to improve the galvanic connection between the chalcopyrite and pyrite in the concentrate and mineral support.

EXAMPLE 6 The experiment described in Example 1 was repeated using 491.8 g of the same melt-grade concentrate that was used in that experiment. The support rock in which the concentrate was applied comprised 2.5 kg of ore in pieces of 3.2 mm to 6.4 mm and 2.5 kg of ore in pieces of 6.4 mm to 12.7 mm. The mixture of the two sizes of chalcopyrite ore was coated with the high grade concentrate by rotating the ore in a drum at approximately 30 rpm, while being sprayed with 10% H2SO4 as used in example 1. The dry concentrate was Spread over the moistened plurality that was stirred from pomo mineral substrates was done in Example 1. The coated substrates were placed on an 8.0-cm glass column. The column was wrapped with electricity-resistant heating tape to isolate the column and help control the temperature. The temperature was monitored by a thermo pair attached by tape to the outside of the column and a glass thermometer in the upper central part of the mineral in the column. An additional 100 g of uncoated ore from 6.4 mm to 12.7 mm was used to cover the coated ore material. This uncoated mineral formed a layer approximately 2 cm thick that covered the coated substrate and acted as an insulating layer to prevent heat loss in the upper part of the bed. Air and liquid were introduced into the upper part of the column as in Example 1. The air was heated by bubbling through heated water and a stainless steel tube heated as in Example 1. The liquid that came out of the column was contained in a heated beaker as described in example 1. From the first day the temperature of the column was maintained at 70 ° C while four liters of a solution having a pH of 1.0 were circulated through the column at a flow rate greater than one liter per day. The solution used in relation to this example was different from examples 1 and 2. These media used in this example comprised 0.2 g / l (NH4) 2S04, 0.4 g / l MgSO4-7H2O, 0.1 g / l K2HPO4, 0.1 g / l. l KCl. The high concentration of acid used to coat the ore support in pieces quickly adjusted the pH of the coated mineral to below 1.6. On the second day the column was inoculated with the same culture of extreme thermophiles (Acidianus brierleyi (strains DSM 1651 and 6334), Acidianus infernus (strain DSM 3191), Sulfolobus acidocaldaríus (strain ATCC 49426), and Sulfolobus metallicus (strain DSM 6482)) that was used to inoculate the column of example 2. This mixed culture of extreme thermophiles was recovered from the disassembly of the column in example 1. Bacteria can be recovered after the biooxidized concentrate is washed from the support material. The suspension of the concentrate removed is allowed to stand overnight. The cloudy liquid can have high levels (107 or more bacteria per ml) of bacteria that can be used to inoculate directly or that can be centrifuged to form higher concentrations of bacteria. The pH of the leaching solution of the process added to the top of the column was maintained between a pH of 1.1 and 1.3. The pH of the spent solution was generally between 1.3 and 1.6. The levels of copper and iron in solution were determined once or twice a week. The solution was removed from the system and replaced with a new solution containing the nutrient mixture. This was done for the same reason as in Example 1, that is, to keep below the toxic level of copper until the microorganisms had time to adapt to the high copper concentrations. The main difference between the experiments in examples 1 and 2 and the present is the use of sulfate media to which chloride had not been added. The degree of leaching of copper and iron was estimated by determining the concentrations of copper and iron in the spent solution, and plotted against the time in Figure 8.

EXAMPLE 7 Another concentrate of the fraction of less 3.2 mm of ore from San Manuel was elaborated. The same procedure as described in example 4 was used. This pyrite-chalcopyrite concentrate by volume was 7.3% copper, 27.4% iron and more than 30% sulfur in sulfide form. Accordingly, this concentrate contained approximately twice the amount by weight of pyrite as well as chalcopyrite. Unlike example 4, 443.1 g of this concentrate were applied as a coating on the chalcopyrite mineral comprising 2.5 kg of ore from 3.2 mm to 6.4 mm and 2.5 kg of ore from 6.4 mm to 12.7 mm. The mixture of the two sizes of chalcopyrite ore was coated with the low grade concentrate by spinning the ore in a drum at approximately 30 rpm, while being sprayed with 10% H2SO4 as used in this example 1. The dry concentrate was It was spread over the moistened plurality that was stirred from substrate mineral as was done in Example 1. The final water content of the coated ore particles was about 3% by weight. In addition, without considering the sulfide mineral particles exposed in the ore support material, the concentrate coated mineral contained approximately 2.5% sulfur in exposed sulphide form. Because the chalcopyrite concentrate was of lesser degree than in the previous examples, the amount of copper in the concentrate is approximately the same as the amount of copper that is in the mineral support rock (54.2% of the copper is in the coated concentrate and 45.8% is found in the mineral support rock). The coated substrates were placed on an 8.0-cm glass column. The column was wrapped with electricity-resistant heating tape to isolate the column and help control the temperature. The temperature was monitored by a thermo pair attached with tape to the outside of the column and a glass thermometer placed in the center of the mineral at the top of the column. Additional 100 g of uncoated ore from the 6.4 mm to 12.7 mm fraction was used to cover the coated ore material. This uncoated mineral formed an approximately 2 cm thick layer that covered the coated substrate and acted as an insulating layer to prevent heat loss in the upper part of the bed. Air and liquid were introduced into the upper part of the column as in Example 1. The air was heated by bubbling through heated water and a stainless steel tube heated as in Example 1. The liquid that came out of the column was contained in a heated beaker as described in example 1. From the first day the column temperature was maintained at 70 ° C while four liters of a solution having a pH of 1.0 were circulated through the column at a flow rate exceeding one liter per day. In this example, the solution was the same as in Examples 1 and 2. The media comprised 0.16 g / l NH4CI, 0.326 g / l MgCl-6H2O, 0.1 g / l K2HPO, and 0.1 g / l KCl. The high concentration of acid used to coat the concentrate in the mineral support rapidly adjusted the pH to below 1.8. On the second day the column was inoculated with the same culture of extreme thermophiles (Acidianus brierleyi, (strains DSM 1651 and 6334), Acidianus infernus, (strain DSM 3191), Sulfolobus acidocaldaríus (strain ATCC 49426), and Sulfolobus metallicus (strain DSM 6482)) which was used to inoculate the column of example 2. This mixed culture of extreme thermophiles was recovered from the dismantling of the column in example 1. The bacteria they can be recovered after the biooxidized concentrate is washed from the substrate. The suspension of the concentrate removed is allowed to stand overnight. The cloudy liquid can have high levels (107 or more batteries per ml) of bacteria that can be used to inoculate directly or that can be centrifuged to form higher concentrations of bacteria. The pH of the leaching solution of the process applied to the top of the column was maintained between a pH of 1.1 and 1.3. The pH of the solution was generally between 1.3 and 1.6. The levels of copper and iron in solution were determined once or twice a week. The solution was removed from the system and replaced with fresh solution containing the nutrient mixture. This was done for the same reason as in example 1, that is, keep below the toxic level of copper until the microorganisms had time to adapt to the high concentrations of copper. The main difference between the experiments in Examples 1 and 2 and the present is the use of a pyrite-chalcopyrite concentrate in low grade volume. The presence of pyrite can increase the chalcopyrite leaching rate through galvanic interaction. This example is also different from Examples 4 and 5 since the bioleaching was carried out at 70 ° C from the start and the cell was inoculated with the same culture of extreme thermophiles as used in Examples 1, 2, and 6. The degree of leaching of copper and iron was estimated by a determination of the concentrations of copper and iron in solution, and plotted against time in Figure 9. The leaching of copper decreased in speed after leaching the equivalent of the amount of copper which was estimated was contained within the concentrate in volume applied as a coating on the ore. This was 54.2% of the total copper in the column and was leached before day 40 of the experiment. The remaining copper, which was believed to be from copper support ore leached at a lower rate from day 40 onwards. Although the invention has been described with reference to preferred embodiments and specific examples, those skilled in the art will readily appreciate that many modifications and adaptations of the invention are possible without departing from the spirit and scope of the invention as claimed below.

Claims (33)

NOVELTY OF THE INVENTION CLAIMS

1. - A high temperature pile bioleaching process for extracting copper from ore containing chalcopyrite, the process comprising the steps of: a. constructing a pile comprising mineral containing chalcopyrite, said pile including sulfide mineral particles exposed at least 25% by weight of which comprise chalcopyrite, characterized in that the concentration of sulfide mineral particles exposed in said pile is such that said pile contains at least 10 kg of sulfur in the form of sulfide exposed per metric ton of solids in said cell, and characterized in that at least 50% of the total copper in said cell is in the form of chalcopyrite; b. heating a substantial portion of said stack to a temperature of at least 50 ° C; c. inoculating said cell with a culture comprising at least one thermophilic microorganism that biooxides sulfur minerals at a temperature above 50 ° C; d. irrigating said stack with a process leaching solution comprising sulfuric acid and ferric iron; and. biolixivir enough sulfide mineral particles in said cell to oxidize at least 10 Kg of sulfur in the form of sulfur per metric ton of solids in said cell and cause the dissolution of at least 50% of the copper in said cell in said cellulose leaching solution. procedure in a period of approximately 210 days or less of the term of said battery; and f. collecting the leaching solution of pregnant process containing dissolved copper as drained from said stack.

2. The method according to claim 1, further characterized in that said cell includes at least 30 kg of sulfur in the form of sulfur per metric ton of solids in said cell.

3. The process according to claim 2, which comprises bioleaching sufficient particles of sulfide ore in said cell to oxidize at least 30 kg of sulfur in the form of sulfur per metric ton of solids in said cell in a period of about 210. days or less.

4. The process according to claim 1 or 3, which comprises bioleaching sufficient particles of sulfide minerals in said cell to cause the dissolution of at least 70% of the copper in said cell in said process leaching solution in approximately 210 days or less.

5. The process according to claim 1 or 3, which comprises bioleaching sufficient particles of sulfide mineral in said pip to cause the dissolution of at least 80% of the copper in said pile in said process leaching solution in a period of time. of approximately 100 days or less.

6. The method according to claim 1, further characterized in that said substantial portion of said stack is heated to a temperature of at least 60 ° C.

7 -. 7 - The method according to claim 1, further characterized in that said substantial portion of said stack is heated to a temperature of at least 70 ° C.

8. The process according to claim 1, further characterized in that said culture comprises a mixed culture that includes a plurality of different thermophilic microorganisms that biooxidize sulphide minerals at a temperature above 60 ° C.

9. The process according to claim 8, further characterized in that said culture comprises a mixed culture that includes a plurality of different Archaea thermophilic that biooxid sulfide minerals at a temperature above 60 ° C.

10. The method according to claim 1, further characterized in that said stack is constructed by a method comprising the steps of: a. coating the surface of a plurality of crude substrates having a particle size greater than about 0.3 cm and less than about 2.54 cm with a sulfide mineral concentrate comprising chalcopyrite and having a particle size of less than about 250 μm; and b. make a plurality of said plurality of coated raw substrates to form said stack.

11. The method according to claim 10, further characterized in that said plurality of raw substrates comprise at least one material selected from the group consisting of rock, brick, slag, and plastic.

12. - The method according to claim 11, further characterized in that said plurality of raw substrates comprise rock selected from the group consisting of lava rock, sterile rock, and crushed copper ore.

13. The process according to claim 10, 11 or 12, further characterized in that the amount of concentrate applied as coating on said plurality of raw substrates is about 9% to 30% of the combined weight of said concentrate and said substrates raw.

14. The process according to claim 13, further characterized in that said concentrate further comprises at least one easily oxidizable sulfide mineral.

15. The process according to claim 14, further characterized in that said easily oxidizable sulfide mineral (at least one) comprises at least one sulfide mineral from the group consisting of pyrite, arsenopyrite, covelite, and chalcocite.

16. The process according to claim 14, further characterized in that the amount of sulfur in sulfide form of said easily oxidizable sulfide mineral (at least one) is at least about 10 Kg per metric ton of solids in the cell and the amount of sulfur in the form of sulfur of said chalcopyrite is at least about 10 Kg per metric ton of solids in the cell.

17. - The method according to claim 15, further characterized in that the amount of sulfur in sulfide form of said easily oxidizable sulfide mineral (at least one) is at least about 10 kg per metric ton of solids in the cell and the The amount of sulfur in the sulphide form of said chalcopyrite is at least about 10 kg per metric ton of solids in the cell.

18. The method according to claim 16, further characterized in that said substantial portion of said stack is heated to a temperature of 50 ° C in a period of about 45 days or less and at least a portion of the heat required to heat said The battery is supplied by bioleaching of at least 10 kg of sulfur in the form of sulfur per metric ton of solids in said cell.

19. The method according to claim 17, further characterized in that said substantial portion of said stack is heated to a temperature of 50 ° C in a period of about 45 days or less and at least a portion of heat required to heat said The battery is supplied by bioleaching of at least 10 kg of sulfur in the form of sulfur per metric ton of solids in said cell.

20. The process according to claim 1, further characterized in that said process leaching solution further comprises chloride ions.

21. The method according to claim 1, further comprising covering said stack with an insulating barrier layer.

22. The method according to claim 21, further characterized in that said insulating barrier layer is selected from the group consisting of an awning, a glass fiber insulation layer, a plastic sheet layer, and a layer of crushed rock .

23. A process of bioleaching in a high-temperature cell for the recovery of copper from ore containing chalcopyrite, the process comprising the steps of: a. constructing a pile comprising mineral containing chalcopyrite, said pile including particles of sulfide ore exposed at least 25% by weight of which comprise chalcopyrite, characterized in that the concentration of sulfide minerals exposed in said pile is such that said pile contains minus 10 kg of sulfur in the form of sulfide exposed per metric ton of solid in said cell, and characterized in that at least 50% of the total copper in said cell is in the form of chalcopyrite; b. heating at least 50% of said battery to a temperature of at least 60 ° C; c. maintain at least 50% of said battery at a temperature of at least 60 ° C until at least 50% of the copper in said battery has dissolved; d. inoculating said cell with a culture comprising at least one thermophilic microorganism that biolixes sulfide minerals at a temperature above 60 ° C; and. irrigating said pile with a process leaching solution at a rate of at least 72 liters / m2 / day; F. bioleaching sulfide mineral particles in said pile to thereby cause the dissolution of the sulfide mineral particles and generate heat, characterized in that sufficient sulfide minerals are oxidized in a bioleaching period of 210 days or less to oxidize at least 10 kg of sulfur in the form of sulfur per metric ton of solids in said cell and causing the dissolution of at least 50% of the copper in said cell in said process leaching solution; g. collecting a leaching solution of pregnant process that includes copper cations from said cell during said period of bioleaching; and h. recover copper from said pregnant process leaching solution.

24. The process according to claim 23, further characterized in that said leaching solution of pregnant process contains at least 2 g / l of copper.

25. The process according to claim 23, further characterized in that said leaching solution of pregnant process comprises at least 5 g / l of copper.

26. The process according to claim 23, further characterized in that the copper is recovered from said leaching process leaching solution by a process selected from the group consisting of solvent extraction, ion exchange, and copper cementation.

27. The process according to claim 23, further characterized in that the copper is recovered from said leaching solution of pregnant process by solvent extraction.

28. The method according to claim 23, further comprising covering said stack with an insulating barrier layer.

29. - The method according to claim 23, further characterized in that said cell is heated by flowing at least one heat source selected from the group consisting of steam, heated air, and heated aqueous solution through said cell.

30. The method according to claim 23, further characterized in that said cell is heated to 60 ° C with heat generated from biolixing a portion of said sulfide mineral particles in said cell with at least one microorganism selected from the group consisting of of moderate mesophiles and thermophiles. 31.- The procedure according to claim 29 or 30, further characterized in that at least 50% of said stack is heated to 60 ° C in a period of 30 days. 32. The method according to claim 23, further comprising the step of passing said collected pregnant process leaching solution and said process leaching solution through a heat exchanger to transfer heat of • said process leaching solution leaching to said process leaching solution. 33. The method according to claim 23, further characterized in that the heat generated by leaching is sufficient to maintain at least 50% of said battery at said temperature.