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FLOTATION OF COARSE PARTICLES OF COPPER SULPHIDES

By: Jorge Ganoza, Metallurgist.


Introduction

It’s well known that the behavior of the particle size in the froth phase is important in determining flotation performance. The flotation rate constant is influenced by the probability that a coarse or fine particle survives the cleaning action of the froth zone and reports to the froth product[1]. It is generally accepted that a froth of good stability is important for the achievement of good grade and high recovery of copper sulphides because many operations try to increase the tonnage, but the stability of the operation is not evaluated properly when the feed size is coarser than the design.

In the present investigation, a flotation testing program was conducted to provide a better understanding of the behavior of the coarse particles of copper sulphides in rougher and cleaning flotation. Parameters examined included the effect of the particle size on rougher flotation performance, the recovery of copper and gold, and the level of regrinding a coarse rougher copper concentrate ahead of the cleaning stages. A locked cycle test was performed to estimate the recovery of copper and gold in a continuous circuit.

At primary grind size of approximately 250µm K80, primary and secondary copper sulphides were greatly liberated. There was very little locking between copper sulphides and other sulphide minerals in the sample. A small fraction of copper minerals was reported as non-sulphide minerals. The ore represented by the composite tested is, therefore, simple and should respond most favorably to standard flotation treatment protocols with mechanical cells. In rougher flotation test work at 250µm K80, copper recoveries averaged closed to 90% and gold recovery near 73%. A flotation feed size coarser than 250µm K80, reduced copper recoveries by about 10%. In cleaner tests, concentrate grades reported an average of 42% copper. 

The technology of coarse particle flotation increases the upper limit of the particle size of flotation feed size and can reduce energy consumption, which is especially suitable for the disposal of tailings by the presence of coarse particles[2].

Particle Attachment

The flotation process is a very effective method to separate valuable minerals from the gangue during treatment in mineral processing operations. The process consists of a series of consecutive sub-processes that include bubble-particle collision, attachment and stability of a bubble-particle. After bubble-particle collision, the particle is attached to the bubble surface and a bubble-particle aggregate is formed in the process. The aggregate is transported to the froth phase. The bubble-particle attachment and detachment are critical sub-processes for successful flotation. The bubble-particle collision is the principal sub-process of flotation that has a significant effect on the flotation and recovery[3].

The attachment efficiency (Ea) and detachment efficiency (Ed) quantify the sub-processes of attachment and detachment, respectively. The bubble–particle attachment and detachment sub-processes have been relatively unexplored because they are complex processes and are controlled by the surface chemistry and physicochemical aspects of particles and air bubbles. The detachment process is controlled by the hydrodynamic of flotation cells. In general, Ea increases when the particle size is smaller and the contact angle is larger. In the same way, Ea decreases with increasing the particle size and bubble size, but increases with increasing particle contact angle[3]. This point is important at the moment of performing the flotation of coarse particles. In addition, Ed increases by increasing the turbulence. The collision efficiency increases by increasing the bubble size.

Before a bubble–particle attachment can take place, a particle has to collide with a bubble, taking a distance of separation at which, the surface forces start to operate. The determination of the bubble-particle collision implies the evaluation of the forces that motivate a particle to deviate from the fluid streamlines near the bubble surface and collide with a bubble. The determination of bubble–particle collision involves the evaluation of forces that cause a particle to deviate in its trajectory from fluid streamlines near the bubble surface and collide with a bubble. Figure 1 shows four bubble–particle collision mechanisms, inertia (a), gravity (b), interception (c), and Brownian diffusion (d). The coarse lines represent the movement of the particles and the fine lines the movement of the fluid. The mechanism of inertial collision is more probable for coarse and dense particles that are not able to follow fluid streamlines and tend to move along a straight trajectory. If the density of particles is greater than that of the surrounding fluid, the particles could have a certain settling velocity and then, their trajectory deviates from fluid streamlines. This deviation may cause particles to collide with the bubble surface. The collision of particles with the surface of the bubble for interception is due to a flow that transports particles along the fluid streamlines. The particles come in contact with the surface of the bubble because their finite size. The bubble-particle collision by Brownian diffusion is significant for very tiny particles that moves randomly in the fluid [4]. The bubble-particle collision can occur by the individual mechanisms mentioned or could be the result of two or more of these mechanisms.

Limit of Particle Size 

The degree of hydrophobicity can be expressed by the contact angle, which is the angle in the triphasic line of contact between the mineral, the aqueous phase and the air bubble. It is accepted that the higher the contact angle of a mineral surface, the more easily it becomes wetted with the air, therefore more hydrophobic. The hydrophobicity or contact angle of the particles depends on the type and distribution of species present on the mineral surface. Recovery decreases with increasing particle size due to detachment, and decreases with a small particle size due to inefficient collision[5].

There is an upper size limit for floatable particles. The balance of forces acting on the particle and the bubble will determine the stability of the assemblage. Coarse particles, either of a single type or composite, can detach from the surface of the bubble[6]. After attaching, two conditions are necessary for flotation: stability and floatability of the aggregate,

The predominant forces are gravitational and capillary forces. The maximum liftable particle size is different from the maximum floatable particle size, the first is obtained under static conditions while the second is influenced by a dynamic state. Some results with quartz particles indicate that a higher contact angle is required to lift large particles. For example, a 3.4 mm particle can only be lifted by a 1.8 mm bubble, at a constant upward velocity of 20 µm/s, and if the contact angle of advancing water with the particles is at least 80º. The size of the particles that can rise decreases as the diameter of the bubble decreases. The upward velocity is also important and the size of the particle that can be lifted decreases as the upward velocity increases with the acceleration[5]. Figure 2 shows the influence of the contact angle for a bubble to transport the maximum quartz particle size respect a bubble size.

Although high agitation/turbulence in the flotation cell increases the frequency of particle and bubble collision, and therefore the possibility of flotation and recovery, too much turbulence is detrimental as coarser particles can be detached from the bubbles when the force of detachment that can be simply represented by the centrifugal force and gravity become larger than the bubble-particle adhesion force, which is considered proportional to the hydrophobicity of the particle or the contact angle[7]. Considering that the centrifugal force and gravity. are proportional to the mass of the particle, the detachment force increases with the diameter of the particles and the density of the particles, and consequently the possibility of flotation and recovery will decrease in the flotation cell. After attachment, two conditions are necessary for flotation: stability and floatability of the bubble-particle aggregate.

Figure 3 shows the recovery curves versus particle size for a particular copper ore along a bank of flotation cells[7]. These curves present a maximum recovery at approximately 100 µm and a decrease in recovery for smaller and larger particles, attributed to the low bubble-particle collision efficiency and high detachment of the collected minerals. The shape of these curves is common to all mechanical flotation cells, regardless of feed size and types of minerals floated, although maximum particle size and recovery can be increased or decreased by mineralogical and density factors of the minerals floated. Figure 4 shows results of increased bubble-particle collision/attachment of different types of minerals at 30 seconds and 10 minutes of flotation.

Copper Sulphide Liberation

The degree of liberation among the certain minerals must be considered as an important factor that affects the flotation performance. The liberation among valuable minerals and gangue strongly has an impact on flotation results in coarse particle region, and decrease with particles increase in case of porphyry cooper ores[8]. The liberation is usually obtained via crushing, and grinding operations of valuable mineral grains from those considered waste. These operations fracture mineral aggregates, inducing or increasing the liberation of valuable mineral grains, resulting in two types of fracture corresponding to two types of liberation. If the interface between grains is strong, the fracture occurs across the grain, with liberation induced only by size reduction. In the other case, if the interface between grains is weak, the fracture will be intergranular, with liberation increased by detachment[9].

The liberation of the ore is dependent on the mineralogical characteristics. In this way, the metallurgical performance of a flotation circuit depends on the texture of the ore and the particle size obtained in the grinding circuit[10]. Figure 5 displays a theoretical grade-recovery curve influenced by liberated and locked particles with gangue minerals. The mineralogical studied are an important source of information to determine the presence of copper sulphides, gangue minerals and degree of liberation. 

Considering a practical point of view, it is necessary to relate the size of the particles to the degree of liberation and secondly to relate the size and liberation to the flotation process. The generation of meaningful quantitative data for prediction of the grind size requirements to achieve desired mineral separation from liberation data is overlooked in some metallurgical studies and can be an elusive goal. For example, the random fracture pattern assumption, that considers that assays in all size classes after breakage are the same is a weak point of some liberation studies[11]. It is necessary to indicate that the direct measurement of mineral liberation in ground ores is a complex issue because polished section used for microscopical analysis introduce stereological errors since a locked particle may be observed as free if the section happens to pass through only one of the mineral presents[11]. This kind of situation is typical of copper sulphides. The experience to read and make a good interpretation is important to estimate the metallurgical performance.

A stream of liberated particles with the value and gangue minerals in different classes of particles will separate sharply, whereas a stream of composite particles will not be separated properly[12]. Figure 6 represents the products from rock fragmentation. The fragments of a complex ore can be liberated particles, binary composites, or tertiary composites.

A mineralogical study by QemScan (Quantitative Evaluation of Minerals by Scanning Electron Microscopy)[13] generates a good estimation of the level of liberation after grinding. Liberation classes are defined as Free: A mineral with ≥95% surface area; Liberated: A mineral with ≥80% but <95% surface area; Middling: A mineral with ≥50% but <80% surface area; Sub-Middling: A mineral with ≥20% but <50% surface area and Locked: A mineral with <20% surface area[13]. Figure 7 is an example of the level of liberation by size of copper sulphides.

It is important to indicate that a complete liberation does not guarantee a high recovery. The liberated minerals have to be concentrated in an optimal size range and appropriate flotation parameters (e.g., slurry density, K80, pH, dosage of collector and frother). There is an interaction between mineral liberation and particle size[14]. Figure 8 shows different recoveries of chalcopyrite-bearing particles of different sizes and liberation after 8 minutes. In this case coarse particles of chalcopyrite have a low flotation because the optimum flotation parameters were not identified properly.

The limit of liberation must be analyzed and selected considering the mineralogy of the ore and the results of the flotation process. If the study is not well performed, some losses of valuable mineral and dilution of the concentrate grade will have an impact on the recovery process. Coarse particles of economic minerals and gangue can be reported in the rougher concentrate and may increase the recovery, but reduce the rougher concentrate grade. Try to reject these particles is an alternative, but the recovery is affected, although the overall concentrate grade could remain almost stable[12]. Figure 9 illustrates the causes of low recovery and concentrate grade.

By determining the appropriate grind and possible size of liberation, we can know if there might be any mechanical entrainment issues of gangue minerals in the froth. There are several factors that influence mechanical entrainment, one of them being particle size. This is usually evaluated using a recovery curve at different sizes, and comparing the results with the average particle size evaluated. In the analysis, the particles reported by mechanical entrainment show a greater displacement when the particle size is reduced[15]. The flotation at a coarse size can partially reduce the presence of gangue in the concentrate, but particles with very little liberation can also be reported in the froth, which can complicate the design and/or operation of the regrind and cleaning circuit.

The presence of large particles is an indication that high-grade concentrates can be produced if the valuable particles have a appropriate degree of liberation and it is possible to produce a good concentrate with high separation efficiency if the degree of liberation of the particles is closed to 80% or more[16]. Other point to consider is the slurry density because the K80 values of the concentrates obtained from the flotation test with a slurry density of 30% w/w are greater than the K80 values of the concentrates obtained from the flotation test with a slurry density of 25% w/w. In both flotation tests, as time passes, the K80 value decreases[16]. Figure 10 shows a downward trend in the particle size of chalcopyrite from the first concentrate to the final concentrate. Figure 11 displays the rougher concentrate collected after 20 seconds.

Experimental Work

A copper ore sample was used to perform the metallurgical testing program. The composite was crushed to -10 mesh and thoroughly homogenized. The composite was split into 2-kilogram sub-samples and stored in a freezer until required for testing program. A representative portion was taken for head assay analysis and mineralogical study.

The ore is essentially a high-grade copper, assaying 2.07% Cu, 0.008% Pb, 0.006% Zn, 10.3% Fe, 0.006%Mo, 20g/t Ag and 0.34g/t Au. Acid soluble copper in the sample was as high as 0.23%, which indicates that levels of copper oxides would be a restriction to get very high copper recovery by flotation. This is a sample of low oxide, 11% of total copper[17], but it may have an impact on copper recovery.

The copper minerals detected included the copper sulphides bornite, chalcocite, and chalcopyrite. No native copper was detected in the analyzed sample; however, this is usually a nuggety occurrence. The copper silicate mineral detected was chrysocolla (only a small fraction). The higher levels of bornite and chalcocite would be expected to result in high copper concentrate grades. Table 1 shows the mineral abundance.

Pyrite was also detected. At pyrite to copper sulphide mineral ratio of 0.028 for the composite, it is not necessary to consider a great effort to prevent pyrite flotation into the copper concentrate.

The flotation testing program was performed using 2-kilogram ore charges and laboratory rod mill was used to grind the copper ore. The ground pulp was transferred into a rectangular rougher flotation cell made of stainless steel, and the cell was mounted on a Denver Laboratory Model D-12 Flotation Machine. After the remaining volume of the cell had been filled with tap water, flotation reagents were added into the cell and the pulp was conditioned for a given period of time at 1,800 rpm prior to introducing air into the pulp to start the flotation. The Denver Laboratory Model D-12 was used because the flotation parameters can be scaled up to get reliable information. The laboratory flotation machine is mechanically agitated, and it can simulate the large-scale models commercially available. Introduction of air is normally via a hollow standpipe surrounding the impeller shaft. In the standard operation, the cell works as self-aspirated, the action of the impeller draws air down the standpipe, and the air rate being controlled by the speed of the impeller. The cell Denver D12 is versatile and can be modified to work as forced-air injection. In both operating modes, the air stream is sheared into fine bubbles by the impeller[18]

In this test program, batch rougher and cleaner tests were conducted primarily to evaluate rougher circuit performance and reagent dosages The flotation testing program on the copper composite tested a primary grind of 150, 200, 250 300 and 350µm K80 in rougher flotation using PAX (Potassium Amyl Xanthate) as collector. F-549 and MIBC were added as frother. The addition of two frothers may help to increase the collision probability and a coarse particle can be collected and attached to the bubble[19]. Lime was added to regulate the pH at 10.5. The copper rougher concentrate was reground and cleaned to produce a copper concentrate. Results of the locked cycle tests were used to estimate the metallurgical performance of the composite

Results Mineralogy

Results from mineralogical study were useful to get information on the mineralogical association on the composite sample ground to 150µm K80. The results of this analysis brought out the following points:

About 35% to 65% of the copper sulphides (liberated and binary forms) are liberated with most of the remainder locked in binary form with pyrite, non-sulphide gangue minerals. The binary association data for the copper sulfides indicated a preferential association between copper sulfides. Bornite was associated preferentially with chalcocite and chalcopyrite. Chalcocite was strongly associated with bornite and chalcopyrite, and chalcopyrite was preferentially associated with bornite and chalcocite. The strong association of the copper sulfides supports the existence of a coarse grain size for these associated minerals and the flotation of the copper sulfide minerals as a single collective concentrate. That is, it is not necessary to liberate the copper sulfides from each other, they can be reported in the copper concentrate. The most common contaminants are likely to be silicates and carbonates. 

At a sizing of 150um K80, the liberation of copper sulphides was within the typical range for a practical copper flotation. Try to consider a finer flotation feed size will not improve the liberation and metallurgical performance. It is necessary to mention that a finer size could favor the production of slimes and some of them could smear over the economic minerals of copper. The presence or smeared and tarnished particles affect the recovery of economic sulphides[20]. However, it is possible to consider a coarser flotation feed size because the main copper sulphides detected in the composite are bornite and chalcopyrite, 47% and 33%, respectively, represents 80%, and the remaining fraction represented by chalcocite is 20%. It means that the recovery of chalcopyrite and bornite has a significant impact on the metallurgical performance. Coarse particles can be composites of copper-bearing minerals and gangue minerals[21]. The distribution of copper sulphide minerals is shown in Table 2.

Interlocking between the pyrite and the copper sulphides was essentially non-existent for the composite because about 10% of the copper sulphides were locked together with pyrite in binary forms.

Gangue minerals were almost totally liberated (96.5%) from the copper sulphide minerals at the 150µm K80 grind sizing. This would imply that the ores could be readily processed at much coarser flotation feed sizings. For instance, if the ore is ground to 250 µm K80, and depending on the size of sulphide grains, the copper sulphides could be liberated, fully or partially, then that specific coarser size would represent liberation at coarse fragment size[22].

Effect of Primary Grind

The effect of primary grind on metallurgical performance was investigated over a range of nominal primary grind size of 150µm to 350µm K80. The metallurgical performance was assessed using rougher kinetic tests. The addition of flotation reagents was similar in the rougher tests and the flotation feed size was the main variable. In general, the rate of flotation of the copper sulphide minerals was good. This would imply that better metallurgical response might be expected from ores with similar mineralogy. Figure 12 displays the comparative copper grade-recovery data for the composite. 

Copper recoveries to the rougher concentrate averaged 88%, and the solids mass pulls averaging 9.5%. Gold recovery to the rougher concentrate averaged 72%. A primary grind size of 250µm K80 was determined as optimum for the copper composite. At this size, the copper recovery was 90% and the gold recovery 74%. A similar flotation feed size has been adopted by some copper operations[23].

Copper recovery was lower when the primary grind size was increased to 300 and 350µm K80. At this point, the copper recovery was reduced by 9% and the mass pull trend to be much higher, approximately 2.6% more. The copper performance did not change significantly when the primary grind size was reduced to 150 and 200µm K80, but the mass pull was higher. The copper recovery was slightly higher, but not drastically different than the results obtained at 250µm K80. Figure 13 displays the variability of the mass respect to the copper recovery. It is important to indicate that a high degree of liberation is not always necessary and may be undesirable in some cases. In this way, flotation requires at least a surface of the valuable mineral to be exposed to get a good response[24].

The best results for gold recovery, where a high recovery of 74% was achieved at a mass pull of 9%, were found at 250µm K80. The good result suggest that gold bearing minerals liberates somewhat preferentially in grinding. At finer size, the gold recovery is not sensitive. The gold recovery to the copper rougher concentrate is sensitive to primary grind sizing between 300 to 350µm K80. Figure 14 displays the comparative gold grade-recovery data for the composite.

Copper Cleaner Tests

The results from the rougher tests were important to select the primary grind size, rougher flotation time, and determine the potential misplacement of gangue minerals in the copper rougher flotation circuit. It was noted that composite does not need a large flotation time to get acceptable recoveries of copper in the rougher concentrate, but it was necessary to evaluate the effect of regrinding the copper rougher concentrate. The objective of performing cleaner tests was to evaluate the grade-recovery relationship, across the cleaning circuit, following the regrinding stage.

Cleaner tests produced a saleable grade copper concentrate. Dependent upon the relative abundances of copper sulphides, concentrate grades ranged from 43% to 49% copper content. See Figure 15. It was not necessary to consider a first cleaner-scavenger stage because copper sulphides showed a good floatability.

In the batch cleaner tests, at a regrind product sizing of 36µm K80, a concentrate grade of 49% copper at 80% recovery to the final concentrate was produced. If the regrind product is too fine, the copper recovery is low. This copper sample required some regrinding of the rougher concentrate in order to ensure the production of high-grade copper concentrate. Acceptable metallurgical performance in the cleaner stage was achieved at an average regrind product sizing of 50µm K80 at around 90% copper recovery. Figure 16 displays a sample of copper concentrate and rougher tailings. 

Gold recoveries to the final concentrate averaged approximately 59% at copper recoveries to the final concentrate of 83%. When the rougher concentrate was not reground, the gold recovery was affected, being only 43%. Gold recovery appeared to trend linearly with copper recovery, although some variability was noted in the results from regrinding tests. The gold grade in the concentrate averaged 4.3 g/t at concentrate grades averaging 46% copper.

Effect of Regrinding the Rougher Concentrate

One of the key variables investigated by performing batch cleaner tests was the effect of regrinding the rougher concentrate ahead of the cleaner flotation stages. This part of the study was confined to work on the copper rougher concentrate produced at 250µm K80. Batch cleaner tests were selected to investigate the effect of regrinding on the copper concentrate grade. 

A series of batch cleaner tests were performed, investigating regrind discharge sizings within the range 30 to 80µm K80 for the copper rougher concentrate. Following regrinding, high-grade copper concentrates were produced by two stages. In general, regrind size showed a predictable effect on metallurgical response over the size range of 50 to 60µm K80. Finer regrinds did not have a significant effect on the copper grade at a lower recovery of copper. A summary of the results can be found in Figure 17.  

For the copper concentrate, there was a linear relationship between copper concentrate grade and the regrind discharge sizing. For the composite, as the regrind size was reduced, concentrate grade is higher; the change is significantly respect to the test without regrinding the rougher concentrate. Considering the results from batch cleaner tests, the appropriate regrind size of the copper rougher concentrate is around 50µm K80. Figure 18 shows the effect on copper and gold recovery.

Regrinding Energy Requirement

The grindability tests for fine materials indicate the specific energy needed to regrind the rougher concentrate from its original size to any specified size or sizes. The objective in grindability tests for fine materials is not to determine the Bond Work Index because the relevance of the Index to materials such as sand or flotation middling products has not been established and the use of the Index might lead to make wrong estimations. The Levin test allows direct estimation of the energy needed for the grinding to a specific size[25, 26].  

In general, the rougher concentrate target regrind size was 50µm K80. The relative copper rougher flotation performance heavily influenced the regrind mill operation. Copper rougher concentrates from batch flotation tests at 250 µm K80 were used to conduct standard Levin plot tests. This data can be used to engineer and specify full-scale equipment. The summary of the tests is shown in Table 3. As shown in the table, the power requirement for the copper rougher concentrate is around 11 kWh/t.

Locked Cycle Tests

Results form batch cleaner tests were useful to plan the Locked Cycle Test (LCT) on the copper composite. LCT was performed to evaluate the equilibrium metallurgical performance that might be anticipated from continuous flotation operation.  

Locked cycle test at 250µm K80 was performed for the copper sample. A basic reagent scheme was utilized for the test: Lime was used to regulate the pH. PAX was used as collector for the copper sulphides and gold bearing minerals in the rougher and cleaner stages. F549 and MIBC were used as frothers. The copper rougher concentrate was reground ahead of the cleaner stages at around 50µm K80.

Based on the last two cycles, results indicates that copper, and gold recovery in the copper cleaner concentrate 89% and 73%, respectively. By combining the results, the copper concentrate grade for copper was on average 42.6% copper, and the gold grade was on average 3.9 g/t gold. The iron misplacement was 16.2%. Results from LCT are shown in Table 4. The calculation of the balance was made considering the Cycle-by-Cycle technique, which is relatively straightforward. The major stipulation is that, like the combined-products method of calculation, the feed to each cycle must be the same weight and must be true samples, it means that they must be identical in mineral composition[27, 28]. The approach to steady state has been examined by plotting the distribution of elements and total weight as a function of cycle number in Figure 19.

It is important that the locked cycle test reached steady state. A complementary analysis is to examine the distribution of the concentrate weight cycle by cycle and the circulating load. The weight recovery to the final concentrate appears to have reached an equilibrium value between 4.6 and 4.8 % and this was reached in the third cycle. At the same time the circulating load varied from 1.24% in the first cycle to the final value of 1.30%. These circulating loads are quite acceptable and considering the levels in the operating plant these circulating loads are remarkably low. 

There will be gold and silver credits in selling the copper concentrate. Buyer can pay for 97.50% of the final gold content, subject to a minimum deduction of 1 g/t. In the case of silver, the buyer can pay for 90.0% of the final silver content, subject to a minimum deduction of 30 g/t[29, 30].

Conclusions

1. The sample provided for study was of variable copper mineralization. The dominant sulphides minerals, comprised of a range of copper sulphides minerals (bornite, chalcopyrite and chalcocite) together with gold bearing minerals and pyrite. A small content of copper oxide mineral was detected.

2. Copper sulphides liberation, at primary grind sizings of 250µm K80 was acceptable, and was sufficient to ensure good copper recoveries to the rougher concentrate. Projected mineral liberation levels at the coarse size were also satisfactory indicating that rougher flotation circuit performance may be acceptable at this sizing. The majority of un-liberated copper sulphides occurred as simple binary structures with non-sulphide gangue. These interlocked particles still had, on average, enough copper sulphides to anticipate recovery to the rougher concentrate.

3. Based on the locked cycle test performed on the sample, the composite displayed favorable metallurgical performance in the rougher circuit. Utilizing a single rougher flotation feed sizing of about 250µm K80, and a simple xanthate-based reagent scheme and the addition of two frothers, acceptable flotation recoveries of the copper sulphide minerals and gold bearing minerals were recorded into the copper rougher concentrate.

4. Regrinding of copper rougher concentrates is necessary to improve concentrate grade and achieve good recoveries of copper and gold. Following regrinding of the copper rougher concentrate to about 50µm K80, two stages of cleaning was necessary to get the production of gold enriched copper concentrate.

5. Marketable grade copper concentrate containing appreciable gold and silver contents can be produced from the copper sample tested in the laboratory, and using essentially simple processing parameters. 

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