CA1104274A - Separation of sulfides by selective oxidation - Google Patents
Separation of sulfides by selective oxidationInfo
- Publication number
- CA1104274A CA1104274A CA317,321A CA317321A CA1104274A CA 1104274 A CA1104274 A CA 1104274A CA 317321 A CA317321 A CA 317321A CA 1104274 A CA1104274 A CA 1104274A
- Authority
- CA
- Canada
- Prior art keywords
- flotation
- product
- pulp
- chalcopyrite
- concentrate
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Expired
Links
Classifications
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/02—Froth-flotation processes
- B03D1/06—Froth-flotation processes differential
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/002—Inorganic compounds
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/004—Organic compounds
- B03D1/012—Organic compounds containing sulfur
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/02—Froth-flotation processes
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/02—Collectors
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2203/00—Specified materials treated by the flotation agents; specified applications
- B03D2203/02—Ores
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- Chemical & Material Sciences (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
ABSTRACT OF THE DISCLOSURE
A bulk concentrate containing chalcopyrite, pentlan-dite and pyrrhotite is aerated under alkaline conditions, then subjected in turn to primary flotation, cyanide addition, condi-tioning, secondary flotation and cleaning flotation. As a result, a three-way separation is achieved to produce a copper concentrate, a high grade nickel concentrate and a third product containing most of the pyrrhotite.
A bulk concentrate containing chalcopyrite, pentlan-dite and pyrrhotite is aerated under alkaline conditions, then subjected in turn to primary flotation, cyanide addition, condi-tioning, secondary flotation and cleaning flotation. As a result, a three-way separation is achieved to produce a copper concentrate, a high grade nickel concentrate and a third product containing most of the pyrrhotite.
Description
1104Z7~
FIELD OF THE IN~ENTION
The present invention relates to ore beneficiation, and more specifically to a flotation process for separating sulfide minerals of copper, nickel and iron from one another.
BACKGROUND OF THE INVENTION
The treatment of sulfidic copper-nickel ores (which generally contain varying amounts of chalcopyrite and pentlan-dite) is complicated by the fact that a certain amount of pyrrhotite is usually present in addition to the sulfides of the non-ferrous metals. Typically such ores are subjected to a bulk flotation to reject a substantial amount of gangue therefrom, and thereafter the bulk concentrate obtained is treated to separate the sulfides of nickel and copper from one another. In effecting such a separation, it is desirable to be able to produce a copper-rich concentrate exhibiting a copper-to-nickel ratio of 30 or more, although in practice it is not always possible to achieve so good a separation.
At the same time it is desirable to carry out the separation is such a way that as much as possible of the copper present is recovered in the copper-rich concentrate. Until now the best known method for producing such a copper concentrate has been that described in copending Canadian patent application Serial No. 267,349, filed December 7, 1976 and assigned in common with the present invention. The process described therein entails a flotation carried out at slightly elevated temperature in the presence of lime and sodium cyanide.
Under those conditions, depression of both pentlandite and pyrrhotiteleads to production of a froth product which is rich in copper and a tailings product in which most of the pyrrhotite reports so that the nickel assay thereof is only modest.
~6' Unless otherwise specified, all percentages quoted in the present specification and claims are percentages by weight.
OBJECT OF THE INVENTION
It is an object of the invention to provide a process which enables a three-way separation to be obtained so as to produce a first product which contains most of the chalco-pyrite present in the ore feed, a second product which contains most of the pentlandite present in the feed, and a third pro-duct containing most of the pyrrhotite.
SUMMARY OF THE INVENTION
.
According to the invention, an improved process is provided wherein a sulfidic ore containing copper, nickel and iron is subjected to bulk flotation using a xanthate collector to obtain a bulk concentrate which contains chalcopyrite, pentlandite and pyrrhotite, and the bulk concentrate is sub-jected to further flotation treatment to separate the chalco-pyrite therefrom, and wherein the further flotation treatment comprises the steps of:
i) treating an aqueous pulp of the bulk con-centrate with lime to raise the pulp pH to a value of at least about 12.0;
ii) introducing an oxidizing gaseous stream into the pulp to reduce the level of residual xanthate therein to below a predetermined level;
iii) subjecting the pulp to a primary flota-tion whereby a first float product containing primarily chalcopyrite and pyrrhotite is separated from a first sink product which contains primarily pentlandite and constitutes a high grade nickel con-centrate;
iv) treating the first float product with a cyanide salt to depress pyrrhotite and conditioning 1104Z~;~4 the cyanided first float product for a period suffi-cient to ensure a subsequent rapid and substantially complete flotation of chalcopyrite;
v) subjecting the conditioned first float product to secondary flotation to separate a second float product containing primarily chalcopyrite from a second sink product which contains primarily pyrrho-tite and constitutes a low grade nickel concentrate;
vi) subjecting the second float product to cleaning flotation to separate a final float product which constitutes a high grade copper concentrate from a third sink product; and vii) recycling the third sink product to sub-ject it to the cyanide addition and conditioning of step (iv).
The success of the process of the invention stems from the surprising discovery that subjecting the aqueous pulp to gaseous oxidation, e.g., by simply aerating it, results in de-pressing pentlandite while activating pyrrhotite in the pri-mary flotation step carried out thereafter. The aeration must be carried out under conditions of high alkalinity, and preferably it should be continued until the xanthate level has been reduced to a value not greater than 10 x 10 6 molar.
The observed effect of pre-aeration of the pulp is surprising in view of the teaching of prior workers who have examined the effect of pre-aeration on the subsequent flotation of copper minerals. So far as we are aware, no prior investigation of pre-aeration has been concerned with the separation of chalcopyrite from pentlandite. However, pre-aeration has been suggested as a means for improving copper flotation, and as an aid in obtaining separation of chalcopyrite from pyrite.
Thus, for example, U. S. Patent No. 3,456,792 describes a 1104Z7~
process in which aeration of a pulp containing chalcopyrite and pyrite is relied upon to depress the pyrite and enables its separation. Pyrrhotite might be expected to behave in a similar manner to pyrite with respect to the effect of aeration thereon. Indeed such is the observation made in a paper entitled: "The Role of Oxygen in Xanthate Flotation of Galena, Pyrite and Chalcopyrite" by I. B. Klymowsky and P. Salman, CIM TRANSACTIONS, Vol. LXXIII, pp 147-152, 1970, where the authors state:
"Aeration preferentially depresses the pyrite and pyrrhotite minerals asso-ciated with chalcopyrite, and therefore there is an improvement in the grade."
Yet we have found that when applied to bulk concentrates of the type with which the present invention is concerned, i.e., containing chalcopyrite, pentlandite and pyrrhotite, aeration carried out under highly alkaline conditions depresses only the pentlandite present while leaving the pyrrhotite readily floatable with the chalcopyrite thereby enabling a high grade nickel concentrate to be produced by the primary flotation step. In fact we have found that aeration generally improves the flotation of chalcopyrite, particularly in high sulfide environments such as those present when practicing the present invention. Thus when, as is generally preferred, a thickener is used at the start of the flotation circuit, the oxygen demand of the pulp is such that dissolved oxygen is consumed and the redox potential is observed to become very negative. Under such redox conditions chalcopyrite does not float readily, yet after aeration of the thickened pulp, the chalcopyrite flotation was found to be rapid and substantially complete.
~104Z74 The bulk concentrate feed to the process of the inven-tion will typically be in the form of a pulp of 30-35~ solids density, and such a pulp density is suitable for performing the various flotation operations in the process of the invention. It is preferred, however, to thicken the pulp prior to the aeration treatment and thereafter thin it before subjecting it to flotation. Thus the initial pH adjustment can be carried out in a thickener which increases the pulp density to about 60-70% solids. The use of such a thickener, while by no means essential to operation of the process of the invention, is preferred for several reasons. Firstly the lime addition to the feed slurry causes some of the xanthate to be released and hence discarded with the water re-moved in the thickening operation. Moreover, the reduction in pulp volume enables smaller vessels to be used for aeration of the pulp.
The aeration oxidizes the xanthate which is present in the feed as a result of the bulk flotation to which it has been subjected. Whatever the precise role played by the oxi-dant gas may be, we have found that measurement of the -xanthate concentration in the pulp provides a reliable guide to the desired end point of the aeration. While the aeration can be accomplished by injecting pure oxygen into the slurry, it is by no means necessary to rely on pure oxygen and, for reasons of convenience and economy, air will be used in practice. We have attempted to achieve similar results by using chemical oxidants to oxidize the xanthate or by using charcoal to adsorb the xanthate, however we have been unable to obtain the desired three-way separation of chalcopyrite, pentlandite and pyrrhotite from one another without resorting to gaseous oxidation.
Subsequent to the primary flotation, the float pro-duct is essentially a chalcopyrite-pyrrhotite mixture. In order to depress the pyrrhotite, cyanide is added, preferably in the form of sodium cyanide, and in an amount corresponding to at least about 0.3 gram/kilogram, e.g., between 0.3 and 0.5 g/kg of bulk concentrate feed treated. After addition of the cyanide to the pulp, a conditioning treatment is needed to ensure that chalcopyrite is not also depressed by the cyanidation but is rapidly and completely floatable in the subsequent flotation. The conditioning can be carried out at ambient temperature and will generally involve an equivalent batch residence time of at least about 5 minutes. It is a particular advantage of the process of the invention that neither the various flotation operations, nor the treatments of the pulp therebetween require raising the pulp temperature above ambient.
The conditioned pulp is then subjected to the second-ary flotation operation whereby a sink product is obtained which contains most of the pyrrhotite in the feed. The float product is cleaned to yield a final float product which con-tains a high proportion of the chalcopyrite, with little of the pentlandite and pyrrhotite previously associated with it.
The cleaning will generally be effected in a multistage countercurrent operation, i.e., a flotation operation wherein the froth product of each stage is fed to the succeeding stage, while the sink product of each stage is recycled to the preceding stage.
In order to provide a clearer understanding of the invention, examples thereof will now be specifically described with reference to the accompanying drawings.
BRIEF DESCRIPTION OF THE DRAWINGS
In the accompanying drawings:
1104Z7~
Figure 1 schematically represents a flow sheet of a process for treating a bulk concentrate in accordance with the present invention; and Figure 2 is a graph depicting the effect of various reagent additions on the effectiveness of the process of Figure 1.
DETAILED DESCRIPTION OF THE EMBODIMENTS
A series of tests were carried out in the manner illu-strated in Figure 1. The feed was a freshly prepared mill product obtained by flotation using a xanthate collector. It was first thickened to a pulp density of 65% solids and sub-jected to a liming operation 11 during which lime was added in such amounts as to provide a lime titration of 0.7 to 0.9 gram/kilogram of solution. This ensured a pH in excess of 12.
The alkaline pulp was then subjected to oxidation 12 by blowing air through it for a period of two hours. At the end of this oxidation period the residual level of xanthate had decreased to less than 10 x 10 6 molar. During this aeration, consi-derable frothing tends to occur due to the presence of frothing agents in the pulp. We found, however, that stirring which is vigorous enough to produce a vortex provided adequate froth control during the oxidation process.
The oxidized pulp was then thinned to a pulp density of 30-35% solids and subjected to primary flotation 13. The feed rate to this flotation operation corresponded to about `
10 kilograms of solids per minute, and an equivalent batch residence time of about 10 minutes. The float product from the primary flotation was combined with the sink product from cleaning flotation 17 and the resulting pulp was subjected to cyanide addition followed by conditioning. The cyaniding 14 comprised adding to the pulp 0.6 gram of sodium cyanide per . ' , ~ , ~ .
kilogram of bulk concentrate feed. The conditioning operation 15 comprised holding at room temperature for an equivalent batch residence time of at least 5 minutes.
The conditioned pulp was then fed to the secondary flotation 16 wherein most of the pyrrhotite was rejected as a sink product. This flotation was carried out with an equiva-lent batch residence time of about 12 minutes. The float product from operation 16 was mixed with the sink product from recleaning operation 18 and fed to a cleaning flotation 17 which had an equivalent batch residence time of 8 minutes. The float product from this operation was fed to recleaning flotation 18 which had an equivalent batch residence time of 5 minutes and produced a high grade copper concentrate as its float product.
Table 1 shows results obtained when the above described process was used with a feed which assayed:
Cu : 11.5%
Ni : 11.9%
Fe : 35.2%
S : 32.9%
Assay (%) Distribution (%) Product Cu Ni Fe Weight Cu Ni Fe Feed 11.511.9 35.2 100 100 100 100 Cu concen-trate 32.10.33 31.2 34.2 95.6 0.9 30.3 High Grade 0.21 30.2 28.6 31.2 0.6 79.3 25.4 trate Low Grade 1.286.8 45.1 34.7 3.9 19.8 44.4 Ntratoencen-.
1104Z7~
Using a feed concentrate which assayed:
Cu : 12.6%
Ni : 9.9%
Fe : 37.1%
S : 33.6%
the results in Table 2 were obtained.
. .
Assay (~) Distribution (%) Product Ni Fe Weight Cu Ni ¦ Fe Feed 12.69.937.1 100 100 100 100 Cu concen- 30.00.4432.0 40.396.31.7 34.8 High Grade Ni 0.2929.5 30.3 27.90.5 83.2 22.8 concentrate Law Grade Ni 1.374.7 49.5 31.83.1 15.1 42.4 concentrate .
To provide a numerical factor for evaluating the suc-cess of the process, we have used the measured data to calculate a Recovery and Separation Factor (RSF) which reflects both the extent of copper recovery in the copper concentrate and the grade of the latter. The RSF is defined as follows:
RSF = A C 4B
where: A is the weight of copper in the copper concentrate;
B is the weight of nickel in the copper concentrate; and C is the weight of copper in the feed.
An ideal copper-nickel separation process would exhibit an RSF of unity. The RSF values calculated from the results of Tables 1 and 2 are 0. 916 and 0. 906 respectively. Such high RSF values were unattainable using any prior known separation process. The RSF values are not, however, the only criteria ~104Z74 by which the process should be judged. An equally significant, and heretofore unattainable result is the production of a nickel concentrate assaying about 30% by weight of nickel and containing about 80% of all the nickel in the bulk concentrate feed. In fact the results of Tables 1 and 2 show an almost complete three-way separation of the minerals: chalcopyrite, pentlandite and pyrrhotite from one another. Thus from the results of Table 1 and the determined sulfur assays of the various products, the distribution of the minerals was calcu-lated to be as shown in Table 3.
. ....
Distribution (%) Product __ Chalcopyrite Pentlandite Pyrrhotite Feed 100 100 100 Cu Concentrate95.5 O.9 5.8 High Grade Ni Con- 0.6 79.3 6.0 centrate Low Grade Ni Con- 3.9 19.3 88.2 centrate .
In order to determine the preferred process conditions described above, the use of various amounts of lime and cyanide was investigated. In each case the RSF value was calculated from the measured results and from the data a mathematical model was developed to relate the RSF values to the lime and cyanide additions. Figure 2 shows a series of curves derived from the mathematical model to represent the profiles of RSF
values 0.84, 0.86, 0.88, 0.90 and 0.92. Also shown in Figure 2 are the individual data points representing the empirically determined RSF values indicated. It can be seen from these curves that an optimum RSF value of 0.92 can be achieved with a lime addition of 1.05 g/kg and a sodium cyanide addition ~04Z7~
of 0.37 g/kg. For practical reasons, including ease of fil-tration of the nickel concentrate produced, it is undesirable to use quite so much lime. Accordingly, if a preferred lime addition of about 0.9 g/kg is used, a cyanide addition of about 0.38 g/kg will be needed to ensure an RSF of at least 0.90.
The recovery of nickel in the high grade concentrate and the grade of the latter do not appear to be substantially affected by the amounts of lime or cyanide added.
The excellent results in Tables 1 and 2 were obtained by using a process which involved an aeration carried out con-tinuously in three tanks with a total mean residence time of
FIELD OF THE IN~ENTION
The present invention relates to ore beneficiation, and more specifically to a flotation process for separating sulfide minerals of copper, nickel and iron from one another.
BACKGROUND OF THE INVENTION
The treatment of sulfidic copper-nickel ores (which generally contain varying amounts of chalcopyrite and pentlan-dite) is complicated by the fact that a certain amount of pyrrhotite is usually present in addition to the sulfides of the non-ferrous metals. Typically such ores are subjected to a bulk flotation to reject a substantial amount of gangue therefrom, and thereafter the bulk concentrate obtained is treated to separate the sulfides of nickel and copper from one another. In effecting such a separation, it is desirable to be able to produce a copper-rich concentrate exhibiting a copper-to-nickel ratio of 30 or more, although in practice it is not always possible to achieve so good a separation.
At the same time it is desirable to carry out the separation is such a way that as much as possible of the copper present is recovered in the copper-rich concentrate. Until now the best known method for producing such a copper concentrate has been that described in copending Canadian patent application Serial No. 267,349, filed December 7, 1976 and assigned in common with the present invention. The process described therein entails a flotation carried out at slightly elevated temperature in the presence of lime and sodium cyanide.
Under those conditions, depression of both pentlandite and pyrrhotiteleads to production of a froth product which is rich in copper and a tailings product in which most of the pyrrhotite reports so that the nickel assay thereof is only modest.
~6' Unless otherwise specified, all percentages quoted in the present specification and claims are percentages by weight.
OBJECT OF THE INVENTION
It is an object of the invention to provide a process which enables a three-way separation to be obtained so as to produce a first product which contains most of the chalco-pyrite present in the ore feed, a second product which contains most of the pentlandite present in the feed, and a third pro-duct containing most of the pyrrhotite.
SUMMARY OF THE INVENTION
.
According to the invention, an improved process is provided wherein a sulfidic ore containing copper, nickel and iron is subjected to bulk flotation using a xanthate collector to obtain a bulk concentrate which contains chalcopyrite, pentlandite and pyrrhotite, and the bulk concentrate is sub-jected to further flotation treatment to separate the chalco-pyrite therefrom, and wherein the further flotation treatment comprises the steps of:
i) treating an aqueous pulp of the bulk con-centrate with lime to raise the pulp pH to a value of at least about 12.0;
ii) introducing an oxidizing gaseous stream into the pulp to reduce the level of residual xanthate therein to below a predetermined level;
iii) subjecting the pulp to a primary flota-tion whereby a first float product containing primarily chalcopyrite and pyrrhotite is separated from a first sink product which contains primarily pentlandite and constitutes a high grade nickel con-centrate;
iv) treating the first float product with a cyanide salt to depress pyrrhotite and conditioning 1104Z~;~4 the cyanided first float product for a period suffi-cient to ensure a subsequent rapid and substantially complete flotation of chalcopyrite;
v) subjecting the conditioned first float product to secondary flotation to separate a second float product containing primarily chalcopyrite from a second sink product which contains primarily pyrrho-tite and constitutes a low grade nickel concentrate;
vi) subjecting the second float product to cleaning flotation to separate a final float product which constitutes a high grade copper concentrate from a third sink product; and vii) recycling the third sink product to sub-ject it to the cyanide addition and conditioning of step (iv).
The success of the process of the invention stems from the surprising discovery that subjecting the aqueous pulp to gaseous oxidation, e.g., by simply aerating it, results in de-pressing pentlandite while activating pyrrhotite in the pri-mary flotation step carried out thereafter. The aeration must be carried out under conditions of high alkalinity, and preferably it should be continued until the xanthate level has been reduced to a value not greater than 10 x 10 6 molar.
The observed effect of pre-aeration of the pulp is surprising in view of the teaching of prior workers who have examined the effect of pre-aeration on the subsequent flotation of copper minerals. So far as we are aware, no prior investigation of pre-aeration has been concerned with the separation of chalcopyrite from pentlandite. However, pre-aeration has been suggested as a means for improving copper flotation, and as an aid in obtaining separation of chalcopyrite from pyrite.
Thus, for example, U. S. Patent No. 3,456,792 describes a 1104Z7~
process in which aeration of a pulp containing chalcopyrite and pyrite is relied upon to depress the pyrite and enables its separation. Pyrrhotite might be expected to behave in a similar manner to pyrite with respect to the effect of aeration thereon. Indeed such is the observation made in a paper entitled: "The Role of Oxygen in Xanthate Flotation of Galena, Pyrite and Chalcopyrite" by I. B. Klymowsky and P. Salman, CIM TRANSACTIONS, Vol. LXXIII, pp 147-152, 1970, where the authors state:
"Aeration preferentially depresses the pyrite and pyrrhotite minerals asso-ciated with chalcopyrite, and therefore there is an improvement in the grade."
Yet we have found that when applied to bulk concentrates of the type with which the present invention is concerned, i.e., containing chalcopyrite, pentlandite and pyrrhotite, aeration carried out under highly alkaline conditions depresses only the pentlandite present while leaving the pyrrhotite readily floatable with the chalcopyrite thereby enabling a high grade nickel concentrate to be produced by the primary flotation step. In fact we have found that aeration generally improves the flotation of chalcopyrite, particularly in high sulfide environments such as those present when practicing the present invention. Thus when, as is generally preferred, a thickener is used at the start of the flotation circuit, the oxygen demand of the pulp is such that dissolved oxygen is consumed and the redox potential is observed to become very negative. Under such redox conditions chalcopyrite does not float readily, yet after aeration of the thickened pulp, the chalcopyrite flotation was found to be rapid and substantially complete.
~104Z74 The bulk concentrate feed to the process of the inven-tion will typically be in the form of a pulp of 30-35~ solids density, and such a pulp density is suitable for performing the various flotation operations in the process of the invention. It is preferred, however, to thicken the pulp prior to the aeration treatment and thereafter thin it before subjecting it to flotation. Thus the initial pH adjustment can be carried out in a thickener which increases the pulp density to about 60-70% solids. The use of such a thickener, while by no means essential to operation of the process of the invention, is preferred for several reasons. Firstly the lime addition to the feed slurry causes some of the xanthate to be released and hence discarded with the water re-moved in the thickening operation. Moreover, the reduction in pulp volume enables smaller vessels to be used for aeration of the pulp.
The aeration oxidizes the xanthate which is present in the feed as a result of the bulk flotation to which it has been subjected. Whatever the precise role played by the oxi-dant gas may be, we have found that measurement of the -xanthate concentration in the pulp provides a reliable guide to the desired end point of the aeration. While the aeration can be accomplished by injecting pure oxygen into the slurry, it is by no means necessary to rely on pure oxygen and, for reasons of convenience and economy, air will be used in practice. We have attempted to achieve similar results by using chemical oxidants to oxidize the xanthate or by using charcoal to adsorb the xanthate, however we have been unable to obtain the desired three-way separation of chalcopyrite, pentlandite and pyrrhotite from one another without resorting to gaseous oxidation.
Subsequent to the primary flotation, the float pro-duct is essentially a chalcopyrite-pyrrhotite mixture. In order to depress the pyrrhotite, cyanide is added, preferably in the form of sodium cyanide, and in an amount corresponding to at least about 0.3 gram/kilogram, e.g., between 0.3 and 0.5 g/kg of bulk concentrate feed treated. After addition of the cyanide to the pulp, a conditioning treatment is needed to ensure that chalcopyrite is not also depressed by the cyanidation but is rapidly and completely floatable in the subsequent flotation. The conditioning can be carried out at ambient temperature and will generally involve an equivalent batch residence time of at least about 5 minutes. It is a particular advantage of the process of the invention that neither the various flotation operations, nor the treatments of the pulp therebetween require raising the pulp temperature above ambient.
The conditioned pulp is then subjected to the second-ary flotation operation whereby a sink product is obtained which contains most of the pyrrhotite in the feed. The float product is cleaned to yield a final float product which con-tains a high proportion of the chalcopyrite, with little of the pentlandite and pyrrhotite previously associated with it.
The cleaning will generally be effected in a multistage countercurrent operation, i.e., a flotation operation wherein the froth product of each stage is fed to the succeeding stage, while the sink product of each stage is recycled to the preceding stage.
In order to provide a clearer understanding of the invention, examples thereof will now be specifically described with reference to the accompanying drawings.
BRIEF DESCRIPTION OF THE DRAWINGS
In the accompanying drawings:
1104Z7~
Figure 1 schematically represents a flow sheet of a process for treating a bulk concentrate in accordance with the present invention; and Figure 2 is a graph depicting the effect of various reagent additions on the effectiveness of the process of Figure 1.
DETAILED DESCRIPTION OF THE EMBODIMENTS
A series of tests were carried out in the manner illu-strated in Figure 1. The feed was a freshly prepared mill product obtained by flotation using a xanthate collector. It was first thickened to a pulp density of 65% solids and sub-jected to a liming operation 11 during which lime was added in such amounts as to provide a lime titration of 0.7 to 0.9 gram/kilogram of solution. This ensured a pH in excess of 12.
The alkaline pulp was then subjected to oxidation 12 by blowing air through it for a period of two hours. At the end of this oxidation period the residual level of xanthate had decreased to less than 10 x 10 6 molar. During this aeration, consi-derable frothing tends to occur due to the presence of frothing agents in the pulp. We found, however, that stirring which is vigorous enough to produce a vortex provided adequate froth control during the oxidation process.
The oxidized pulp was then thinned to a pulp density of 30-35% solids and subjected to primary flotation 13. The feed rate to this flotation operation corresponded to about `
10 kilograms of solids per minute, and an equivalent batch residence time of about 10 minutes. The float product from the primary flotation was combined with the sink product from cleaning flotation 17 and the resulting pulp was subjected to cyanide addition followed by conditioning. The cyaniding 14 comprised adding to the pulp 0.6 gram of sodium cyanide per . ' , ~ , ~ .
kilogram of bulk concentrate feed. The conditioning operation 15 comprised holding at room temperature for an equivalent batch residence time of at least 5 minutes.
The conditioned pulp was then fed to the secondary flotation 16 wherein most of the pyrrhotite was rejected as a sink product. This flotation was carried out with an equiva-lent batch residence time of about 12 minutes. The float product from operation 16 was mixed with the sink product from recleaning operation 18 and fed to a cleaning flotation 17 which had an equivalent batch residence time of 8 minutes. The float product from this operation was fed to recleaning flotation 18 which had an equivalent batch residence time of 5 minutes and produced a high grade copper concentrate as its float product.
Table 1 shows results obtained when the above described process was used with a feed which assayed:
Cu : 11.5%
Ni : 11.9%
Fe : 35.2%
S : 32.9%
Assay (%) Distribution (%) Product Cu Ni Fe Weight Cu Ni Fe Feed 11.511.9 35.2 100 100 100 100 Cu concen-trate 32.10.33 31.2 34.2 95.6 0.9 30.3 High Grade 0.21 30.2 28.6 31.2 0.6 79.3 25.4 trate Low Grade 1.286.8 45.1 34.7 3.9 19.8 44.4 Ntratoencen-.
1104Z7~
Using a feed concentrate which assayed:
Cu : 12.6%
Ni : 9.9%
Fe : 37.1%
S : 33.6%
the results in Table 2 were obtained.
. .
Assay (~) Distribution (%) Product Ni Fe Weight Cu Ni ¦ Fe Feed 12.69.937.1 100 100 100 100 Cu concen- 30.00.4432.0 40.396.31.7 34.8 High Grade Ni 0.2929.5 30.3 27.90.5 83.2 22.8 concentrate Law Grade Ni 1.374.7 49.5 31.83.1 15.1 42.4 concentrate .
To provide a numerical factor for evaluating the suc-cess of the process, we have used the measured data to calculate a Recovery and Separation Factor (RSF) which reflects both the extent of copper recovery in the copper concentrate and the grade of the latter. The RSF is defined as follows:
RSF = A C 4B
where: A is the weight of copper in the copper concentrate;
B is the weight of nickel in the copper concentrate; and C is the weight of copper in the feed.
An ideal copper-nickel separation process would exhibit an RSF of unity. The RSF values calculated from the results of Tables 1 and 2 are 0. 916 and 0. 906 respectively. Such high RSF values were unattainable using any prior known separation process. The RSF values are not, however, the only criteria ~104Z74 by which the process should be judged. An equally significant, and heretofore unattainable result is the production of a nickel concentrate assaying about 30% by weight of nickel and containing about 80% of all the nickel in the bulk concentrate feed. In fact the results of Tables 1 and 2 show an almost complete three-way separation of the minerals: chalcopyrite, pentlandite and pyrrhotite from one another. Thus from the results of Table 1 and the determined sulfur assays of the various products, the distribution of the minerals was calcu-lated to be as shown in Table 3.
. ....
Distribution (%) Product __ Chalcopyrite Pentlandite Pyrrhotite Feed 100 100 100 Cu Concentrate95.5 O.9 5.8 High Grade Ni Con- 0.6 79.3 6.0 centrate Low Grade Ni Con- 3.9 19.3 88.2 centrate .
In order to determine the preferred process conditions described above, the use of various amounts of lime and cyanide was investigated. In each case the RSF value was calculated from the measured results and from the data a mathematical model was developed to relate the RSF values to the lime and cyanide additions. Figure 2 shows a series of curves derived from the mathematical model to represent the profiles of RSF
values 0.84, 0.86, 0.88, 0.90 and 0.92. Also shown in Figure 2 are the individual data points representing the empirically determined RSF values indicated. It can be seen from these curves that an optimum RSF value of 0.92 can be achieved with a lime addition of 1.05 g/kg and a sodium cyanide addition ~04Z7~
of 0.37 g/kg. For practical reasons, including ease of fil-tration of the nickel concentrate produced, it is undesirable to use quite so much lime. Accordingly, if a preferred lime addition of about 0.9 g/kg is used, a cyanide addition of about 0.38 g/kg will be needed to ensure an RSF of at least 0.90.
The recovery of nickel in the high grade concentrate and the grade of the latter do not appear to be substantially affected by the amounts of lime or cyanide added.
The excellent results in Tables 1 and 2 were obtained by using a process which involved an aeration carried out con-tinuously in three tanks with a total mean residence time of
2 hours. An attempt to achieve a similar separation by relying on chemical oxidation gave the results shown in Table 4. In this case the procedure was similar to that described above except that the two-hour aeration was replaced by the addition of 1-2 grams of sodium hypochlorite per kilogram of bulk con-centrate feed to be treated.
Assay (%) Distribution (%) Prcduet Cu Ni Fe Weight ~i Fe Feed 13.911.4 35.2 100 100 100 100 Cu coneentrate24.6O.87 35.5 55.898.8 4.3 56.3 High Grade Ni0.4326.7 32.7 18.8 0.644.0 17.5 concentrate IL~W Grade Ni0.3323.2 36.2 25.4 0.651.7 26.1 eoncentrate The results of Table 4 show an RSF value of 0.844 whieh, though inferior to the results of Tables 1 and 2, never-theless represents an adequate eopper-niekel separation. How-ever, the failure of this comparative test is clearly evidenced by the fact that less than half the nickel content is recovered 1104Z~4 in the so-called high grade nickel concentrate, and the latter's grade is little better than that of the so-called low grade nickel concentrate.
The results of further attempts to find a substitute for aeration are shown in Table 5 below. Four tests are report-ed in which essentially identical feed concentrate was used.
Test A was in accordance with the invention and in fact carried out in the manner described above with reference to Tables 1-3.
Tests B, C and D were carried out in almost the identical manner except that instead of aeration, an addition of chemical oxidant or adsorbant was relied upon to remove the unwanted xanthate.
In each case the hypochlorite, peroxide or charcoal was added in an amount corresponding to 1 g/kg of bulk concentrate.
. -Test ATest B Test C Test D
(using(using (using (using air) NaOCl) H202)charcoal) Cu 11.5 11.3 11.111.6 Feed Assay Ni 8.95 8.91 8.898.84 (%) Fe 37.9 37.6 38.638.3 .
Cu 25.4 21.8 23.930.7 Assay Ni 0.7090.927 0.755 0.490 (%) Fe 33.2 35.8 35.030.1 Cbpper wt 43 4 49.4 44 035.0 Concen- Distri- Cu 95 9 94 894 7 93.0 trate bution Ni 3.40 5 10 3.701.90 (%) Fe 38.0 47.0 39.827.5 .
Cu 0.12 0.21 0.210.15 Assay Ni 29.8 25.5 25.525.4 (%) Fe 29.6 32.0 33.033.1 High-Grade . . wt 24.7 18.7 24.924.0 Ni Dlstrl- Cu 0.30 0.30 0.500.30 Concen- bution Ni 82.3 53.4 71.468.9 trate Fe 19.3 15.9 21.320.7 Cu 1.40 1.72 1.721.90 Assay Ni 4.01 11.6 7.086.28 L ~ (%) Fe 50.7 43.7 48.248.4 Grade Ni CDncen- Di tr' wt 31.9 32.031.2 41.0 trate s 1- Cu 3.90 4.90 4.806.70 bution Ni 14.3 41.524.8 29.1 (%) Fe 42.6 37.1 39.051.7 RSF 0.85 0.79 0.830.87 ~10427~
While the RSF values show that acceptable copper-nickel separation was attained with hydrogen peroxide or char-coal instead of air, the superiority of aeration is clearly shown by the data for the high grade nickel concentrate. It is seen that aeration produced a 29.8% nickel assay in this concen-trate compared with 25.4% and 25.5% in the other tests. More-over, the distribution of nickel in this high grade concentrate was significantly lower, at 53.4-71.4%, in the comparative tests than the value of 82.3% obtained when aeration was used.
Thus, it will be seen that only where the appropriate amounts of lime and cyanide are used and pre-aeration is resorted can the following valuable results be achieved:
i) a high grade copper concentrate;
ii) a high recovery of copper in the copper concentrate;
iii) a high grade nickel concentrate; and iv) a high recovery of nickel in the high grade nickel concentrate.
While the present invention has been described with reference to preferred embodiments thereof, various modifications may be made to those embodiments without departing from the scope of the invention which is defined by the appended claims. - -
Assay (%) Distribution (%) Prcduet Cu Ni Fe Weight ~i Fe Feed 13.911.4 35.2 100 100 100 100 Cu coneentrate24.6O.87 35.5 55.898.8 4.3 56.3 High Grade Ni0.4326.7 32.7 18.8 0.644.0 17.5 concentrate IL~W Grade Ni0.3323.2 36.2 25.4 0.651.7 26.1 eoncentrate The results of Table 4 show an RSF value of 0.844 whieh, though inferior to the results of Tables 1 and 2, never-theless represents an adequate eopper-niekel separation. How-ever, the failure of this comparative test is clearly evidenced by the fact that less than half the nickel content is recovered 1104Z~4 in the so-called high grade nickel concentrate, and the latter's grade is little better than that of the so-called low grade nickel concentrate.
The results of further attempts to find a substitute for aeration are shown in Table 5 below. Four tests are report-ed in which essentially identical feed concentrate was used.
Test A was in accordance with the invention and in fact carried out in the manner described above with reference to Tables 1-3.
Tests B, C and D were carried out in almost the identical manner except that instead of aeration, an addition of chemical oxidant or adsorbant was relied upon to remove the unwanted xanthate.
In each case the hypochlorite, peroxide or charcoal was added in an amount corresponding to 1 g/kg of bulk concentrate.
. -Test ATest B Test C Test D
(using(using (using (using air) NaOCl) H202)charcoal) Cu 11.5 11.3 11.111.6 Feed Assay Ni 8.95 8.91 8.898.84 (%) Fe 37.9 37.6 38.638.3 .
Cu 25.4 21.8 23.930.7 Assay Ni 0.7090.927 0.755 0.490 (%) Fe 33.2 35.8 35.030.1 Cbpper wt 43 4 49.4 44 035.0 Concen- Distri- Cu 95 9 94 894 7 93.0 trate bution Ni 3.40 5 10 3.701.90 (%) Fe 38.0 47.0 39.827.5 .
Cu 0.12 0.21 0.210.15 Assay Ni 29.8 25.5 25.525.4 (%) Fe 29.6 32.0 33.033.1 High-Grade . . wt 24.7 18.7 24.924.0 Ni Dlstrl- Cu 0.30 0.30 0.500.30 Concen- bution Ni 82.3 53.4 71.468.9 trate Fe 19.3 15.9 21.320.7 Cu 1.40 1.72 1.721.90 Assay Ni 4.01 11.6 7.086.28 L ~ (%) Fe 50.7 43.7 48.248.4 Grade Ni CDncen- Di tr' wt 31.9 32.031.2 41.0 trate s 1- Cu 3.90 4.90 4.806.70 bution Ni 14.3 41.524.8 29.1 (%) Fe 42.6 37.1 39.051.7 RSF 0.85 0.79 0.830.87 ~10427~
While the RSF values show that acceptable copper-nickel separation was attained with hydrogen peroxide or char-coal instead of air, the superiority of aeration is clearly shown by the data for the high grade nickel concentrate. It is seen that aeration produced a 29.8% nickel assay in this concen-trate compared with 25.4% and 25.5% in the other tests. More-over, the distribution of nickel in this high grade concentrate was significantly lower, at 53.4-71.4%, in the comparative tests than the value of 82.3% obtained when aeration was used.
Thus, it will be seen that only where the appropriate amounts of lime and cyanide are used and pre-aeration is resorted can the following valuable results be achieved:
i) a high grade copper concentrate;
ii) a high recovery of copper in the copper concentrate;
iii) a high grade nickel concentrate; and iv) a high recovery of nickel in the high grade nickel concentrate.
While the present invention has been described with reference to preferred embodiments thereof, various modifications may be made to those embodiments without departing from the scope of the invention which is defined by the appended claims. - -
Claims (6)
1. In a process wherein a sulfidic ore containing copper, nickel and iron is subjected to bulk flotation using a xanthate collector to obtain a bulk concentrate which contains chalcopyrite, pentlandite and pyrrhotite, and the bulk concen-trate is subjected to further flotation treatment to separate the chalcopyrite therefrom, the improvement wherein the further flotation treatment comprises the steps of:
i) treating an aqueous pulp of the bulk con-centrate with lime to raise the pulp pH to a value of at least about 12.0;
ii) introducing an oxidizing gaseous stream into the pulp to reduce the level of residual xanthate therein to below a predetermined level;
iii) subjecting the pulp to a primary flotation whereby a first float product containing primarily chalcopyrite and pyrrhotite is separated from a first sink product which contains primarily pentlandite and constitutes a high grade of nickel concentrate;
iv) treating the first float product with a cyanide salt to depress pyrrhotite and conditioning the cyanided first float product for a period suffi-cient to ensure a subsequent rapid and substantially complete flotation of chalcopyrite;
v) subjecting the conditioned first float product to secondary flotation to separate a second float product containing primarily chalcopyrite from a second sink product which contains primarily pyrrhotite and constitutes a low grade nickel concen-trate.
vi) subjecting the second float product to cleaning flotation to separate therefrom a final float product which constitutes a high grade copper concentrate and a third sink product; and vii) recycling the third sink product to sub-ject it to the cyanide addition and conditioning of step (iv).
i) treating an aqueous pulp of the bulk con-centrate with lime to raise the pulp pH to a value of at least about 12.0;
ii) introducing an oxidizing gaseous stream into the pulp to reduce the level of residual xanthate therein to below a predetermined level;
iii) subjecting the pulp to a primary flotation whereby a first float product containing primarily chalcopyrite and pyrrhotite is separated from a first sink product which contains primarily pentlandite and constitutes a high grade of nickel concentrate;
iv) treating the first float product with a cyanide salt to depress pyrrhotite and conditioning the cyanided first float product for a period suffi-cient to ensure a subsequent rapid and substantially complete flotation of chalcopyrite;
v) subjecting the conditioned first float product to secondary flotation to separate a second float product containing primarily chalcopyrite from a second sink product which contains primarily pyrrhotite and constitutes a low grade nickel concen-trate.
vi) subjecting the second float product to cleaning flotation to separate therefrom a final float product which constitutes a high grade copper concentrate and a third sink product; and vii) recycling the third sink product to sub-ject it to the cyanide addition and conditioning of step (iv).
2. A process as claimed in claim 1 wherein step (ii) comprises aerating the pulp for a period sufficient to ensure that residual xanthate concentration therein does not exceed about 10 x 10 6 molar.
3. A process as claimed in claim 1 wherein the cyanide addition of step (iv) comprises adding sodium cyanide to the mix-ture of first float product and third sink product in an amount corresponding to at least 0.3 grams per kilogram of bulk concen-trate.
4. A process as claimed in claim 3 wherein the con-ditioning of step (iv) comprises maintaining the cyanided mixture at ambient temperature for at least about 5 minutes.
5. A process as claimed in claim 1 wherein the clean-ing flotation of step (vi) comprises at least two stages of flotation.
6. A process as claimed in claim 5 wherein the two stages consist of a cleaning stage to which the second float product is fed, and a recleaning stage to which froth from the cleaning stage is fed; the third sink product being comprised of tailings from the cleaning stage, and the final float product being comprised of froth from the recleaning stage.
Priority Applications (4)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA317,321A CA1104274A (en) | 1978-12-04 | 1978-12-04 | Separation of sulfides by selective oxidation |
| AU51652/79A AU526210B2 (en) | 1978-12-04 | 1979-10-10 | Sulphide ore flotation |
| ZA00795520A ZA795520B (en) | 1978-12-04 | 1979-10-16 | Flotation of sulphide ore |
| FI793432A FI64297C (en) | 1978-12-04 | 1979-11-01 | FOERFARANDE FOER SEPARATION AV KOPPARKIS PENTLANDIT OCH MAGNETKIS GENOM FLOTATION |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA317,321A CA1104274A (en) | 1978-12-04 | 1978-12-04 | Separation of sulfides by selective oxidation |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1104274A true CA1104274A (en) | 1981-06-30 |
Family
ID=4113066
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA317,321A Expired CA1104274A (en) | 1978-12-04 | 1978-12-04 | Separation of sulfides by selective oxidation |
Country Status (4)
| Country | Link |
|---|---|
| AU (1) | AU526210B2 (en) |
| CA (1) | CA1104274A (en) |
| FI (1) | FI64297C (en) |
| ZA (1) | ZA795520B (en) |
Cited By (5)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| WO1993004783A1 (en) * | 1991-08-28 | 1993-03-18 | Commonwealth Scientific And Industrial Research Organisation | Processing of ores |
| US5295585A (en) * | 1990-12-13 | 1994-03-22 | Cyprus Mineral Company | Method for achieving enhanced copper-containing mineral concentrate grade by oxidation and flotation |
| US5992640A (en) * | 1994-11-16 | 1999-11-30 | Boc Gases Australia Limited | Precious metals recovery from ores |
| WO2009121147A1 (en) * | 2008-04-04 | 2009-10-08 | Bhp Billiton Ssm Development Pty Ltd | Odour control |
| CN115921118A (en) * | 2022-10-11 | 2023-04-07 | 昆明理工大学 | Novel composite inhibitor for separation of pyrite and chalcopyrite and beneficiation method |
Families Citing this family (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CA1238430A (en) * | 1984-12-19 | 1988-06-21 | Gordon E. Agar | Flotation separation of pentlandite from pyrrhotite using sulfur dioxide-air conditioning |
| US6582494B1 (en) * | 1999-09-09 | 2003-06-24 | Billiton Intellectual Property, B.V. | Recovery of nickel and copper from sulphide concentrates by bioleaching |
-
1978
- 1978-12-04 CA CA317,321A patent/CA1104274A/en not_active Expired
-
1979
- 1979-10-10 AU AU51652/79A patent/AU526210B2/en not_active Ceased
- 1979-10-16 ZA ZA00795520A patent/ZA795520B/en unknown
- 1979-11-01 FI FI793432A patent/FI64297C/en not_active IP Right Cessation
Cited By (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US5295585A (en) * | 1990-12-13 | 1994-03-22 | Cyprus Mineral Company | Method for achieving enhanced copper-containing mineral concentrate grade by oxidation and flotation |
| WO1993004783A1 (en) * | 1991-08-28 | 1993-03-18 | Commonwealth Scientific And Industrial Research Organisation | Processing of ores |
| US5992640A (en) * | 1994-11-16 | 1999-11-30 | Boc Gases Australia Limited | Precious metals recovery from ores |
| WO2009121147A1 (en) * | 2008-04-04 | 2009-10-08 | Bhp Billiton Ssm Development Pty Ltd | Odour control |
| EA018117B1 (en) * | 2008-04-04 | 2013-05-30 | БиЭйчПи БИЛЛИТОН ЭсЭсЭм ДИВЕЛОПМЕНТ ПТИ ЛТД. | Odour control |
| US8734757B2 (en) | 2008-04-04 | 2014-05-27 | Bhp Billiton Ssm Development Pty Ltd. | Odor control |
| CN115921118A (en) * | 2022-10-11 | 2023-04-07 | 昆明理工大学 | Novel composite inhibitor for separation of pyrite and chalcopyrite and beneficiation method |
| CN115921118B (en) * | 2022-10-11 | 2024-04-05 | 昆明理工大学 | Novel composite inhibitor for separating pyrite from chalcopyrite and beneficiation method |
Also Published As
| Publication number | Publication date |
|---|---|
| ZA795520B (en) | 1981-02-25 |
| FI64297B (en) | 1983-07-29 |
| FI793432A (en) | 1980-06-05 |
| AU526210B2 (en) | 1982-12-23 |
| AU5165279A (en) | 1980-06-12 |
| FI64297C (en) | 1983-11-10 |
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