CA1156380A - Selective flotation of nickel sulfide ores - Google Patents
Selective flotation of nickel sulfide oresInfo
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- CA1156380A CA1156380A CA000373065A CA373065A CA1156380A CA 1156380 A CA1156380 A CA 1156380A CA 000373065 A CA000373065 A CA 000373065A CA 373065 A CA373065 A CA 373065A CA 1156380 A CA1156380 A CA 1156380A
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- pulp
- nickel
- flotation
- accordance
- pentlandite
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Abstract
A process for beneficiating a nickel sulfide ore in a selective manner comprising wet grinding the ore to liberation of minerals in the presence of sodium carbonate, treating the ground pulp at a pH of about 10 with sodium cyanide to depress the mineral content, conditioning the pulp, raising the redox potential of the pulp in a controlled manner, introducing therein a strong collector, floating a pentlandite concentrate, and thereafter floating a chalcopyrite concentrate.
Description
lls~3~n SELECTIVE FLOTATION OF NICKEL SULFIDE ORES
The invention is directed to beneficiation of complex nickel sulfide ores wherein high recovery ~ of nickel in a rich concentrate is achieved together with improved rejection of pyrrhotite.
In the Sudbury district of Canada, as well as in other parts of the world, nickel is found in a complex, finely disseminated ore with other valuable metals including copper, metals of the platinum group, etc. The principal mineralization consists of pentlandite, chalcopyrite and pyrrhotite, with the pyrrhotite (which itself contains only a minor portion of nickel in solid solution) being present in amounts far greater than the pentlandite. For example, the ratio of pyrrhotite to pentlandite may be approximately 5:1. The ores and the mill process-ing thereof have been under study for many years and much has been discovered as a result. For example, it is known that fine free pentlandite, a valuable mineral, is prone to oxidation and will not float 3~ n
The invention is directed to beneficiation of complex nickel sulfide ores wherein high recovery ~ of nickel in a rich concentrate is achieved together with improved rejection of pyrrhotite.
In the Sudbury district of Canada, as well as in other parts of the world, nickel is found in a complex, finely disseminated ore with other valuable metals including copper, metals of the platinum group, etc. The principal mineralization consists of pentlandite, chalcopyrite and pyrrhotite, with the pyrrhotite (which itself contains only a minor portion of nickel in solid solution) being present in amounts far greater than the pentlandite. For example, the ratio of pyrrhotite to pentlandite may be approximately 5:1. The ores and the mill process-ing thereof have been under study for many years and much has been discovered as a result. For example, it is known that fine free pentlandite, a valuable mineral, is prone to oxidation and will not float 3~ n
-2- PC-2115 once oxidized.
The processes which have been adopted for beneficiation of these complex ores involving the separation of the values into a nickel stream, a copper stream, a pyrrhotite stream and rejected gangue or rock, are effectively compromises in which overall recovery of desired metal values and degree of concentration are balanced. Thus, in flotation of the ores the practice of forming a bulk nickel-copper concentrate with rejection of pyrrhotite and gangue followed by further flotation to provide separation of nickel and copper has been adopted.
In effecting nickel-copper separation, chalcopyrite is floated preferentially to pentlandite. It is necessary to float a large proportion of pyrrhotite to recover slow floating fine pentlandite and middling particles. Thus, substantial quantities of pyrrhotite remain with the nickel concentrate with the result that the nickel concentrate analyzes approximately 11% of nickel. In contrast, the copper concentrate resulting will contain approximately 30~ of copper.
The nickel concentrate still contains copper and the copper c~ncentrate still contains nickel, factors which complicate further processing in the smelter.
The pyrrhotite concentrate rejected from the circuit should be as low as possible in nickel.
, ., 115~3S~0
The processes which have been adopted for beneficiation of these complex ores involving the separation of the values into a nickel stream, a copper stream, a pyrrhotite stream and rejected gangue or rock, are effectively compromises in which overall recovery of desired metal values and degree of concentration are balanced. Thus, in flotation of the ores the practice of forming a bulk nickel-copper concentrate with rejection of pyrrhotite and gangue followed by further flotation to provide separation of nickel and copper has been adopted.
In effecting nickel-copper separation, chalcopyrite is floated preferentially to pentlandite. It is necessary to float a large proportion of pyrrhotite to recover slow floating fine pentlandite and middling particles. Thus, substantial quantities of pyrrhotite remain with the nickel concentrate with the result that the nickel concentrate analyzes approximately 11% of nickel. In contrast, the copper concentrate resulting will contain approximately 30~ of copper.
The nickel concentrate still contains copper and the copper c~ncentrate still contains nickel, factors which complicate further processing in the smelter.
The pyrrhotite concentrate rejected from the circuit should be as low as possible in nickel.
, ., 115~3S~0
-3- PC-2115 The fact that, in the interest of maximizing nickel recovery, the nic~el concentrate presently obtained is of relatively low grade, means that the total content of sulfur in the nickel concentrate fed to the smelter is higher than is wanted in terms of operating cost, nickel throughput, losses of nickel, losses of cobalt and emissions of sulfur dioxide.
For example, raising concentrate grade from about 11% to about 17~ nickel would cut sulfur dioxide emissions approximately 25%, a highly desirable result, and other economies can be achieved. While provision of means for treating nickel sulfide ore to provide a nickel concentrate of improved grade has long been recognized as an objective, no practical means for doing so without encountering severe economic handicaps has heretofore been developed.
Flotation is one of the most useful mineral dressing techniques available to the enqineer for concentrating valuable minerals contained in ores.
Many specific flotation techniques have been developed for the treatment of specific ores and a wide selec-tion of chemical agents has been employed to provide a variety of beneficial results in applying such techniques. For example, S. Power Warren noted in 2~ 1933 a depressing action of cyanide ion on pyrrhotite (Trans. Can. Inst. of Min. and Metal 1933, 186) but ,.~
115~3~)
For example, raising concentrate grade from about 11% to about 17~ nickel would cut sulfur dioxide emissions approximately 25%, a highly desirable result, and other economies can be achieved. While provision of means for treating nickel sulfide ore to provide a nickel concentrate of improved grade has long been recognized as an objective, no practical means for doing so without encountering severe economic handicaps has heretofore been developed.
Flotation is one of the most useful mineral dressing techniques available to the enqineer for concentrating valuable minerals contained in ores.
Many specific flotation techniques have been developed for the treatment of specific ores and a wide selec-tion of chemical agents has been employed to provide a variety of beneficial results in applying such techniques. For example, S. Power Warren noted in 2~ 1933 a depressing action of cyanide ion on pyrrhotite (Trans. Can. Inst. of Min. and Metal 1933, 186) but ,.~
115~3~)
-4- PC-211~
he did not exceed a sodium cyanide addition of 0.2 pounds/ton presumably due to the fear of depressing pentlandite. He was also able to float pentlandite ahead of chalcopyrite but this was claimed to be due to the action of sodium pyrophosphate. Sodium cyanide in lime solution has also been used in an apparently unsuccessful attempt to selectively depress pyrrhotite from pentlandite in Australian ores ~Eltham and Tilyard, Australian Inst. of Min.
and Metal Conference, Western Australia, May 1973).
Soviet literature has reported that cyanide is an unselective depressant in the flotation of copper-nickel ores (Chem. Abs. 54, 10701 f) and that the use of cationic collectors for floating chalcopyrite and lS pentlandite from pyrrhotite has given 13% nickel concentrates (Chem. Abs. 78, 149907 u). Sodium cyanide can be used in a circuit adjusted to a pH of about 12 with lime as a depressant for pentlandite in copper-nickel separation. Sodium carbonate is known to enhance the separation of pentlandite from pyrrho-tite when treatin~ certain Manito~a nickel sulfide ores. Despite the diverse and frequently divergent teachings in the art, there was, prior to the present invention, no way of separating pentlandite from pyrrhotite ~y flotation in such a manner as to obtain a high recovery of pentlandite.
In general terms, the invention contemplates a method for the recovery of metal values contained in 115~3~0 nickel sulfide ores. The process selectively beneficiates a nickel sulfide ore pulp, The process comprising grinding the pulp, adjusting the pulp to about 10 pH units with sodium carbonate or potassium carbonate, treating the Ph adjusted pulp at a temperature of about 6C to about 35C with sodium cyanide or potassium cyanide in an amount to depress the sulfide mineral content of the pulp, establishing the redox potential of the pulp equal to or below a predetermined value, conditioning the pulp, and adding a xanthate collector to the pulp in amounts directly proportional to the com-position of the ore pulp to initiate flotation of pentlanditecaused by aeration of the pulp.
If desired, the remaining pyrrhotite can be floated with rejection of gangue. If regrinding is necessary to liberate miner-als from each other, it is preferable that such grinding be done in the presence of sodium carbonate to pH 10. During the operation the pulp temperature should be at least 6C and preferably not more than 30C, as otherwise the cyanlde ion is rapidly consumed so that the pulp redox potential may rise to the point where flotation of chalcopyrite commences before pentlandite recovery is complete.
Satisfactory operation proceeds at about 20C.
It is preferred to control the sodium cyanide addition to the sodium carbonate treated pulp and the ensuing selective flotation of mineral values contained in the feed through the use of redox potential. Redox potential in the sense used herein de-notes the voltage measurement obtained with a high impedance volt-meter using a metallic working electrode, such ',C
3~0 as a gold electrode, directly in contact with the pulp and a reference electrode, such as a saturated calomel electrode, with the normal porous plug and saturated potassium chloride in the electrode compart-ment. The term redox potential does not imply thatthis is a thermodynamic reversible potential of the pulp. Rather it is a mixed potential derived from both solution species and from instantaneous contact of the working electrode surface with mineral particles.
In more preferred aspects, the invention may be described as a sequential three-way separation by flotation of pentlandite, chalcopyrite and pyrrhotite in that order from an ore or concentrate containing these minerals. Sodium cyanide addition to the aqueous pulp of finely ground minerals at approximately pH 10 due to the presence of sodium carbonate preferably is controlled to provide a redox potential on the order of about -390 mV (gold vs. saturated calomel electrode). The pulp is then conditioned for a period of time, for example, about twenty minutes during which time the solids are held in suspension by brisk stirring with aeration. During the condi-tioning period the redox potential of the pulp rises to about -330 mV. At this point, add~tion of a sufficient amount of a strong sulfhydryl collector such as potassium amyl xanthate in an amount propor-tioned to the nic~el grade o~ the pulp initiates pentlandite flotation. Flotation of pentlandite then , ~ , 1 15~3~t) continues to completion, possibly with further additions of collector. Chalcopyrite and pyrrhotite remain depressed while the pulp redox potential remains below about -245 mV. The potential may be raised to about -240 mV to initiate the chalcopyrite float by, for example, adding a metal salt which reacts with cyanide ion, e.g., nickel sulfate.
Chalcopyrite is then floated without additional collector. Pyrrhotite flotation will be dependent upon the amount of xanthate previously added but usually does not commence until the redox potential rises to about -200 mV.
Substantial amounts of collector are required as compared to conventional flotation processes and the amount is related to pulp nickel grade. For example, a grade of 1.7% nickel requires about 0.25 grams of potassium amyl xanthate per kilogram of pulp solids while at a grade of 10% nickel 1.5 grams of collector per kilogram of pulp solids is needed.
Pulps treated in accordance with the invention generally will contain about 20% to about 45% solids, by weight.
Some examples will now be given.
A low grade concentrate assaying 1.18% copper, 1.75~ nickel and 48.1% pyrrhotite was ground to 13 wt. % on 38 um in sodium carbonate solution ~added to .~...~
I 1 5B3~0 pH 10.5). This concentrate was of a nature such that existing techniques of beneficiation by flotation are ineffective to provide improvement in grade without loss of metal values. The ground pulp was diluted to 30% solids and the pulp pH readjusted to 10 with sodium carbonate. Sodium cyanide in the amount of 1.4 g/kg was added and the pulp then conditioned for 10 minutes in a 3 liter flotation cell. The pulp was aerated at this point for 10 minutes with 1 litre/
minute air to bring the redox potential to -330 mV.
A three minute float was performed prior to xanthate addition to remove hydrophobic rock, yielding a product which was included in the nickel concentrate.
Potassium amyl xanthate in the amount of 0.16 grams per kg of pulp solids was then added. The pulp was conditioned for one minute then floated for 17 minutes, with a further 0.08 grams/kilogram addition of collector after 7 minutes of flotation to produce a nickel concentrate. The potential rose to -240 mV
during flotation of the nickel concentrate. The copper concentrate was taken off in the following 6 minutes flotation. The material balance shown in Table I was obtained.
~P
115~3~0 TABLE I
Ass~Y(%) Distribueion(~) Cu Ni Pn Po wt Cu Ni Pn Po Nickel 4.310.9 29.8 24.1 10.2 38.7 64.8 83.7 4.8 Conc.
Copper 13.6 2.2 5.4 37.6 4.047.95.2 5.9 2.9 Conc.
Pyrrhotite/0.1? 0.60 0.44 55.1 B5.8 13.3 30.0 10.5 92.3 Rock ~Tailing1 He~d Gr~de 1.12 1.72 3.6~ 51.
(Cnlc.) NO~E: Pentlandite (Pn) and Pyrrhotite lPo) are c~lculated assuminq O.B~ nickel in pyrrhotite and 35.9~ nickel in pentlandi~e.
The data of Table 1 show that over 92~ of the pyrrhotite contained in the initial low grade concen-trate was rejected and that the rejected nickel was that associated with pyrrhotite. The high rejection of pyrrhotite meant that the sulfur load on the smelter to recover desired metal values from this particular concentrate was greatly reduced.
In contrast to the results of Example 1 it is found that when the insufficient sodium cyanide is added initially, the redox potential of the pulp is insufficiently depressed with the result that chal-copyrite floats immediately and selectivity is not achieved in flotation. However, a large excess of sodium cyanide cause6 almost complete depression of all sulfides.
:
1 15B3~0 An excess of sodium cyanide results in slowly floating pentlandite and increases the amount of nickel reporting in the copper concentrate. A low S grade concentrate similar to that of Example 1 was similarly treated using an addition of 1.8 g/kg of sodium cyanide and 0.24 g/kg of potassium amyl xanthate collector with the results set fGrth in the following Table 2.
~ABLE 2 AssaY(~) Distribution~) Cu Nl Pn P~ Wt Cu Ni Pn P~
Nickel 4.13 12.8 35.1 22.1 10.5 33.5 60.1 71.1 5.0 Conc.
Copper 13.9 6.68 17.9 30.7 4.2 45.1 12.6 14.5 2.8 Cor~c .
Pyrrhotite/ 0.32 0.71 0.88 49.8 85.3 21.S 27.3 14.5 92.1 1 ~ Rock '~ (T~ ng) He~d Grade 1.29 2.23 5.18 46.1 (C~lc.) In a test similar to that of Example 1, but in which no sodium carbonate was employed, the pH was 7.9 after grinding in a tap water containing about 50 ppm calcium ion and 9.6 after sodium cyanide addition.
The results showed increased losses of pentlandite to the tailings. However, high grade concentrates were obtained. In another test similar to Example 1 but without sodium carbonate in the grind and using an industrial water containing 500 ppm of calcium ion, more than 26~ of the pentlandite failed to float.
:
lls~3~n Tests similar to Example 1 but using, respectively, an insufficient amount of collector, 0.18 grams per kilogram, and using an excess amount of collector, i.e., 0.30 grams/kilogram, showed, respectively, very slow flotation of pentlandite with slightly higher pentlandite losses in the tails while an excess of collector gave the rather unexpected result of reducing nickel recovery with relatively little effect on the pyrrhotite rejection or tailings. The results are tabulated, respectively, in Table 3 and 4 following.
~A6LE 3 ~0.18 qmjkq collector) _ As-ay(~ Distribution(4) Conc. 2.85 11.9 32.5 27.7 B.7 21.B60.37B.3 4.5 Copper 11.7 1.56 3.14 53.46.3 64.75.7 5.5 6.4 Conc.
Pyrrhotite/ 0.18 0.690.69 55.4 B5.0 13.5 34.0 16.2 B9.1 Rock ~ailing) Head Grade 1.14 1.723.61 52.8 ~Calc.) (0.30 gm~kg collector) Assav~) Dictribution(~) Cu Ni Pn Po _ Cu Ni Pn Po Nickel 4.40 8.05 21.825.5 10.7 37.2 3B.8 45.1 6.1 Conc.
Copper 7.13 9.46 25.343.5 8.9S0.0 37.9 43.5 8.6 Conc.
Pyrrhotite/ 0.20 0.650.73 47.B B0.4 12.B 23.4 11.5 85.4 Rock (~ailing) _~ Head Gr-de 1.27 2.235.1B 45.1 ~C-lc.) :
, . ~
lls~3~n It will be appreciated by those skilled in the art that other reagents may be used for purposes known in flotation as long as the combination of sodium carbonate and sodium cyanide is employed as taught herein. For example, known frothers such as MIBC, DOW
SA12~3, etc. may be employed. Potassium cyanide can be used in place of sodium cyanide. Likewise, potassium carbonate can replace sodium carbonate.
However, no substitute has been identified for cyanide ion.
It is also to be understood that the invention is applicable to the treatment of high grade material to produce a nickel concentrate grading 28% nickel, a copper concentrate grading 30% copper and a tailing to the scavenger circuits grading 1% nickel and 1%
copper. When concentrates of such grade are fed to the smelter, the sulfur load thereon is materially lowered as compared to current practice leading to substantial reduction in sulfur dioxide emission and other economics.
he did not exceed a sodium cyanide addition of 0.2 pounds/ton presumably due to the fear of depressing pentlandite. He was also able to float pentlandite ahead of chalcopyrite but this was claimed to be due to the action of sodium pyrophosphate. Sodium cyanide in lime solution has also been used in an apparently unsuccessful attempt to selectively depress pyrrhotite from pentlandite in Australian ores ~Eltham and Tilyard, Australian Inst. of Min.
and Metal Conference, Western Australia, May 1973).
Soviet literature has reported that cyanide is an unselective depressant in the flotation of copper-nickel ores (Chem. Abs. 54, 10701 f) and that the use of cationic collectors for floating chalcopyrite and lS pentlandite from pyrrhotite has given 13% nickel concentrates (Chem. Abs. 78, 149907 u). Sodium cyanide can be used in a circuit adjusted to a pH of about 12 with lime as a depressant for pentlandite in copper-nickel separation. Sodium carbonate is known to enhance the separation of pentlandite from pyrrho-tite when treatin~ certain Manito~a nickel sulfide ores. Despite the diverse and frequently divergent teachings in the art, there was, prior to the present invention, no way of separating pentlandite from pyrrhotite ~y flotation in such a manner as to obtain a high recovery of pentlandite.
In general terms, the invention contemplates a method for the recovery of metal values contained in 115~3~0 nickel sulfide ores. The process selectively beneficiates a nickel sulfide ore pulp, The process comprising grinding the pulp, adjusting the pulp to about 10 pH units with sodium carbonate or potassium carbonate, treating the Ph adjusted pulp at a temperature of about 6C to about 35C with sodium cyanide or potassium cyanide in an amount to depress the sulfide mineral content of the pulp, establishing the redox potential of the pulp equal to or below a predetermined value, conditioning the pulp, and adding a xanthate collector to the pulp in amounts directly proportional to the com-position of the ore pulp to initiate flotation of pentlanditecaused by aeration of the pulp.
If desired, the remaining pyrrhotite can be floated with rejection of gangue. If regrinding is necessary to liberate miner-als from each other, it is preferable that such grinding be done in the presence of sodium carbonate to pH 10. During the operation the pulp temperature should be at least 6C and preferably not more than 30C, as otherwise the cyanlde ion is rapidly consumed so that the pulp redox potential may rise to the point where flotation of chalcopyrite commences before pentlandite recovery is complete.
Satisfactory operation proceeds at about 20C.
It is preferred to control the sodium cyanide addition to the sodium carbonate treated pulp and the ensuing selective flotation of mineral values contained in the feed through the use of redox potential. Redox potential in the sense used herein de-notes the voltage measurement obtained with a high impedance volt-meter using a metallic working electrode, such ',C
3~0 as a gold electrode, directly in contact with the pulp and a reference electrode, such as a saturated calomel electrode, with the normal porous plug and saturated potassium chloride in the electrode compart-ment. The term redox potential does not imply thatthis is a thermodynamic reversible potential of the pulp. Rather it is a mixed potential derived from both solution species and from instantaneous contact of the working electrode surface with mineral particles.
In more preferred aspects, the invention may be described as a sequential three-way separation by flotation of pentlandite, chalcopyrite and pyrrhotite in that order from an ore or concentrate containing these minerals. Sodium cyanide addition to the aqueous pulp of finely ground minerals at approximately pH 10 due to the presence of sodium carbonate preferably is controlled to provide a redox potential on the order of about -390 mV (gold vs. saturated calomel electrode). The pulp is then conditioned for a period of time, for example, about twenty minutes during which time the solids are held in suspension by brisk stirring with aeration. During the condi-tioning period the redox potential of the pulp rises to about -330 mV. At this point, add~tion of a sufficient amount of a strong sulfhydryl collector such as potassium amyl xanthate in an amount propor-tioned to the nic~el grade o~ the pulp initiates pentlandite flotation. Flotation of pentlandite then , ~ , 1 15~3~t) continues to completion, possibly with further additions of collector. Chalcopyrite and pyrrhotite remain depressed while the pulp redox potential remains below about -245 mV. The potential may be raised to about -240 mV to initiate the chalcopyrite float by, for example, adding a metal salt which reacts with cyanide ion, e.g., nickel sulfate.
Chalcopyrite is then floated without additional collector. Pyrrhotite flotation will be dependent upon the amount of xanthate previously added but usually does not commence until the redox potential rises to about -200 mV.
Substantial amounts of collector are required as compared to conventional flotation processes and the amount is related to pulp nickel grade. For example, a grade of 1.7% nickel requires about 0.25 grams of potassium amyl xanthate per kilogram of pulp solids while at a grade of 10% nickel 1.5 grams of collector per kilogram of pulp solids is needed.
Pulps treated in accordance with the invention generally will contain about 20% to about 45% solids, by weight.
Some examples will now be given.
A low grade concentrate assaying 1.18% copper, 1.75~ nickel and 48.1% pyrrhotite was ground to 13 wt. % on 38 um in sodium carbonate solution ~added to .~...~
I 1 5B3~0 pH 10.5). This concentrate was of a nature such that existing techniques of beneficiation by flotation are ineffective to provide improvement in grade without loss of metal values. The ground pulp was diluted to 30% solids and the pulp pH readjusted to 10 with sodium carbonate. Sodium cyanide in the amount of 1.4 g/kg was added and the pulp then conditioned for 10 minutes in a 3 liter flotation cell. The pulp was aerated at this point for 10 minutes with 1 litre/
minute air to bring the redox potential to -330 mV.
A three minute float was performed prior to xanthate addition to remove hydrophobic rock, yielding a product which was included in the nickel concentrate.
Potassium amyl xanthate in the amount of 0.16 grams per kg of pulp solids was then added. The pulp was conditioned for one minute then floated for 17 minutes, with a further 0.08 grams/kilogram addition of collector after 7 minutes of flotation to produce a nickel concentrate. The potential rose to -240 mV
during flotation of the nickel concentrate. The copper concentrate was taken off in the following 6 minutes flotation. The material balance shown in Table I was obtained.
~P
115~3~0 TABLE I
Ass~Y(%) Distribueion(~) Cu Ni Pn Po wt Cu Ni Pn Po Nickel 4.310.9 29.8 24.1 10.2 38.7 64.8 83.7 4.8 Conc.
Copper 13.6 2.2 5.4 37.6 4.047.95.2 5.9 2.9 Conc.
Pyrrhotite/0.1? 0.60 0.44 55.1 B5.8 13.3 30.0 10.5 92.3 Rock ~Tailing1 He~d Gr~de 1.12 1.72 3.6~ 51.
(Cnlc.) NO~E: Pentlandite (Pn) and Pyrrhotite lPo) are c~lculated assuminq O.B~ nickel in pyrrhotite and 35.9~ nickel in pentlandi~e.
The data of Table 1 show that over 92~ of the pyrrhotite contained in the initial low grade concen-trate was rejected and that the rejected nickel was that associated with pyrrhotite. The high rejection of pyrrhotite meant that the sulfur load on the smelter to recover desired metal values from this particular concentrate was greatly reduced.
In contrast to the results of Example 1 it is found that when the insufficient sodium cyanide is added initially, the redox potential of the pulp is insufficiently depressed with the result that chal-copyrite floats immediately and selectivity is not achieved in flotation. However, a large excess of sodium cyanide cause6 almost complete depression of all sulfides.
:
1 15B3~0 An excess of sodium cyanide results in slowly floating pentlandite and increases the amount of nickel reporting in the copper concentrate. A low S grade concentrate similar to that of Example 1 was similarly treated using an addition of 1.8 g/kg of sodium cyanide and 0.24 g/kg of potassium amyl xanthate collector with the results set fGrth in the following Table 2.
~ABLE 2 AssaY(~) Distribution~) Cu Nl Pn P~ Wt Cu Ni Pn P~
Nickel 4.13 12.8 35.1 22.1 10.5 33.5 60.1 71.1 5.0 Conc.
Copper 13.9 6.68 17.9 30.7 4.2 45.1 12.6 14.5 2.8 Cor~c .
Pyrrhotite/ 0.32 0.71 0.88 49.8 85.3 21.S 27.3 14.5 92.1 1 ~ Rock '~ (T~ ng) He~d Grade 1.29 2.23 5.18 46.1 (C~lc.) In a test similar to that of Example 1, but in which no sodium carbonate was employed, the pH was 7.9 after grinding in a tap water containing about 50 ppm calcium ion and 9.6 after sodium cyanide addition.
The results showed increased losses of pentlandite to the tailings. However, high grade concentrates were obtained. In another test similar to Example 1 but without sodium carbonate in the grind and using an industrial water containing 500 ppm of calcium ion, more than 26~ of the pentlandite failed to float.
:
lls~3~n Tests similar to Example 1 but using, respectively, an insufficient amount of collector, 0.18 grams per kilogram, and using an excess amount of collector, i.e., 0.30 grams/kilogram, showed, respectively, very slow flotation of pentlandite with slightly higher pentlandite losses in the tails while an excess of collector gave the rather unexpected result of reducing nickel recovery with relatively little effect on the pyrrhotite rejection or tailings. The results are tabulated, respectively, in Table 3 and 4 following.
~A6LE 3 ~0.18 qmjkq collector) _ As-ay(~ Distribution(4) Conc. 2.85 11.9 32.5 27.7 B.7 21.B60.37B.3 4.5 Copper 11.7 1.56 3.14 53.46.3 64.75.7 5.5 6.4 Conc.
Pyrrhotite/ 0.18 0.690.69 55.4 B5.0 13.5 34.0 16.2 B9.1 Rock ~ailing) Head Grade 1.14 1.723.61 52.8 ~Calc.) (0.30 gm~kg collector) Assav~) Dictribution(~) Cu Ni Pn Po _ Cu Ni Pn Po Nickel 4.40 8.05 21.825.5 10.7 37.2 3B.8 45.1 6.1 Conc.
Copper 7.13 9.46 25.343.5 8.9S0.0 37.9 43.5 8.6 Conc.
Pyrrhotite/ 0.20 0.650.73 47.B B0.4 12.B 23.4 11.5 85.4 Rock (~ailing) _~ Head Gr-de 1.27 2.235.1B 45.1 ~C-lc.) :
, . ~
lls~3~n It will be appreciated by those skilled in the art that other reagents may be used for purposes known in flotation as long as the combination of sodium carbonate and sodium cyanide is employed as taught herein. For example, known frothers such as MIBC, DOW
SA12~3, etc. may be employed. Potassium cyanide can be used in place of sodium cyanide. Likewise, potassium carbonate can replace sodium carbonate.
However, no substitute has been identified for cyanide ion.
It is also to be understood that the invention is applicable to the treatment of high grade material to produce a nickel concentrate grading 28% nickel, a copper concentrate grading 30% copper and a tailing to the scavenger circuits grading 1% nickel and 1%
copper. When concentrates of such grade are fed to the smelter, the sulfur load thereon is materially lowered as compared to current practice leading to substantial reduction in sulfur dioxide emission and other economics.
Claims (11)
1. A flotation process for selectively beneficiating a nickel sulfide ore pulp, the process comprising grinding the pulp, adjusting the pulp to about 10 pH units with sodium carbonate or potassium carbonate, treating the pH adjusted pulp at a temperature of about 6°C to about 35°C with sodium cyanite or potassium cyanide in an amount to depress the sulfide mineral content of the pulp, establishing the redox potential of the pulp equal to or below a predetermined value, conditioning the pulp, and adding a xanthate collector to the pulp in amounts directly proportional to the composition of the ore pulp to initiate flotation of pentlandite caused by aeration of the pulp.
2. A process in accordance with claim 1 wherein said pulp contains pentlandite, chalcopyrite, pyrrhotite and gangue.
3. A process according to claims 1 or 2 wherein the pulp pH
adjustment is accomplished by adding sodium carbonate or potassium carbonate during grinding of the pulp.
adjustment is accomplished by adding sodium carbonate or potassium carbonate during grinding of the pulp.
4. A process in accordance with claim 1 wherein the predetermined value of the redox potential after cyanide ion addition is about -390 mV using a gold vs. saturated calomel electrode.
5. A process according to claim 4 wherein said pulp is con-ditioned with aeration to a redox potential of about -330mV, at which point a strong sulfhydryl collector is added and pentlandite is floated to provide a high grade nickel concentrate.
6. A process in accordance with claim 5 wherein the redox potential of the pulp after pentlandite flotation is adjusted to about -240 mV
at which point chalcopyrite is removed by flotation.
at which point chalcopyrite is removed by flotation.
7. A process in accordance with claim 2 wherein the ratio of pyrrhotite to pentalandite is approximately 5:1.
8. A process in accordance with claims 1 or 2 wherein the pulp temperature is about 20°C.
9. A process in accordance with claim 5 wherein the sulfhydryl collector is amyl xanthate in an amount proportioned to the nickel grade of the pulp.
10. A process in accordance with claims 1 or 2 wherein the pulp contains about 20% to about 45% solids, by weight.
11. A process in accordance with claim 6 wherein after chal-copyrite flotation the pulp redox potential is raised to about -200mV and a pyrrhotite concentrate is then taken off.
Applications Claiming Priority (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| GB8009501 | 1980-03-21 | ||
| GB80/09501 | 1980-03-21 |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1156380A true CA1156380A (en) | 1983-11-01 |
Family
ID=10512261
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA000373065A Expired CA1156380A (en) | 1980-03-21 | 1981-03-16 | Selective flotation of nickel sulfide ores |
Country Status (6)
| Country | Link |
|---|---|
| JP (1) | JPS571456A (en) |
| AU (1) | AU536735B2 (en) |
| BR (1) | BR8101620A (en) |
| CA (1) | CA1156380A (en) |
| FI (1) | FI66544C (en) |
| ZA (1) | ZA811606B (en) |
Cited By (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| AU578327B2 (en) * | 1984-12-19 | 1988-10-20 | Inco Limited | Flotation separation of pentlandite from pyrrhotite using sulfur dioxide-air conditioning |
| US5795466A (en) * | 1995-06-08 | 1998-08-18 | Falconbridge Limited | Process for improved separation of sulphide minerals or middlings associated with pyrrhotite |
-
1981
- 1981-03-11 ZA ZA00811606A patent/ZA811606B/en unknown
- 1981-03-12 AU AU68322/81A patent/AU536735B2/en not_active Ceased
- 1981-03-16 CA CA000373065A patent/CA1156380A/en not_active Expired
- 1981-03-19 BR BR8101620A patent/BR8101620A/en unknown
- 1981-03-19 FI FI810851A patent/FI66544C/en not_active IP Right Cessation
- 1981-03-20 JP JP4156181A patent/JPS571456A/en active Pending
Cited By (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| AU578327B2 (en) * | 1984-12-19 | 1988-10-20 | Inco Limited | Flotation separation of pentlandite from pyrrhotite using sulfur dioxide-air conditioning |
| US5795466A (en) * | 1995-06-08 | 1998-08-18 | Falconbridge Limited | Process for improved separation of sulphide minerals or middlings associated with pyrrhotite |
Also Published As
| Publication number | Publication date |
|---|---|
| ZA811606B (en) | 1982-03-31 |
| AU6832281A (en) | 1981-09-24 |
| FI66544B (en) | 1984-07-31 |
| FI66544C (en) | 1984-11-12 |
| BR8101620A (en) | 1981-09-22 |
| JPS571456A (en) | 1982-01-06 |
| AU536735B2 (en) | 1984-05-24 |
| FI810851L (en) | 1981-09-22 |
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