GB2086768A - Selective flotation of nickel sulphide ores - Google Patents

Abstract

A nickel concentrate is recovered from nickel sulphide ores or concentrates by froth flotation of an aqueous pulp of the ground ore or concentrate, preferably containing 20-45% solids, by weight, by the steps of adjusting the pulp pH to 9-10, adding cyanide ions in an amount sufficient to depress all the sulphide mineral content of the pulp, conditioning the pulp for the flotation of nickel sulphide, adding a strong sulphhydryl collector, e.g. potassium amyl xanthate, to the pulp in an amount sufficient to float the nickel sulphide, and floating and removing the nickel sulphide. Pentlandite, chalcopyrite and pyrrhotite present are successively floated by progressively increasing the redox potential of the pulp.

Description

SPECIFICATION Selective flotation of nickel sulphide ores The invention is directed to beneficiation of complex nickel sulphide 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 as the nickel sulphide mineral, 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 processing 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 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 concentrate 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.

The fact that, in the interest of maximizing nickel recovery, the nickel concentrate presently obtained is of relatively low grade, means that the total content of sulphur 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 sulphur dioxide. For example, raising concentrate grade from about 1 1% to 17% nickel would cut sulphur dioxide emissions approximately 25%, a highly desirable result, and other economies can be achieved. While provision of means for treating nickel sulphide 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 engineer for concentrating valuable minerals contained in ores. Many specific flotation techniques have been developed for the treatment of specific ores and a wide selection of chemical agents has been employed to provide a variety of beneficial results in applying such techniques. For example, S.

Power Warren noted in 1 933 a depressing action of cyanide ion on pyrrhotite (Trans. Can. Inst.

of Min. and Metal 1933, 186) but 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 pendlandite 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 cooper-nickel ores (Chem. Abs. 54, 10701 f) and that the use of cationic collectors for floating chalcopyrite and pentlandite from pyrrhotite has given 13% nickel concentrates (Chem. Abs 78, 1 49907 u).Sodium cyanide can be used in a circuit adjusted to a pH of about 1 2 with lime as a depressant for pentlandite in copper-nickel separation. Sodium carbonate is known to enhance the separation of pentlandite from pyrrhotite when treating certain Manitoba nickel sulphide 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 by flotation in such a manner as to obtain a high recovery of pentlandite.

According to the invention a process for the recovery of metal values from nickel sulphide ores or concentrates by froth flotation of an aqueous pulp of the ground ore or concentrate comprises the steps of adjusting the pulp pH to about 10, e.g. in the range 9-10, adding cyanide ions in an amount sufficient to depress all the sulphide mineral content of the pulp, conditioning the pulp for the flotation of nickel sulphide, adding a strong sulphhydryl collector to the pulp in an amount sufficient to float the nickel sulphide, and floating and removing the nickel sulphide to provide a nickel concentrate.

Flotation is effected in the conventional manner with aeration of the pulp.

The adjustment of the pulp pH to about 10 is preferably done by adding sodium carbonate during the grinding of the pulp. If the pH is too low, substantial amounts of nickel sulphide mineral will fail to float.

If, as will generally be the case, the pulp contains pentlandite, chalcopyrite, pyrrhotite and gangue, the pendlandite is first floated and removed, the chalcopyrite may then be floated, and if desired finally the pyrrhotite, with rejection of gangue. If regrinding is necessary to liberate minerals from each other, it is preferable that such grinding be done in the presence of sodium carbonate to pH 10.

The cyanide ions are preferably added as sodium cyanide, but if desired potassium cyanide may be used instead, and potassium carbonate may be used to adjust the pH instead of sodium carbonate.

It is important to add enough cyanide ions to depress all the sulphide minerals in the pulp: if too little is added, chalcopyrite floats immediately together with the pentlandite. A large excess of cyanide, however, is to be avoided, as this results in slowly floating pentlandite and increases the amount of nickel reporting in the copper concentrate.

It is preferred to control the cyanide addition to the 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 denotes the voltage measurement obtained with a high impedance voltmeter using a metallic working electrode, such 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 compartment. The term redox potential does not imply that this is the 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.

Thus for the sequential three-way separation by flotation of pendlandite, chalcopyrite and pyrrhotite in that order from an ore or concentrate containing these minerals the addition of cyanide to the aqueous pulp of finely ground minerals at approximately pH 10 is preferably controlled to provide a redox potential not less than 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 conditioning period the redox potential of the pulp rises to about - 330 mV. At this point, addition of a sufficient amount of a strong sulphhydryl collector such as potassium amyl xanthate in an amount porportioned to the nickel grade of the pulp initiates pentlandite flotation.

Flotation of pendlandite then continues to completion, further additions of collector being made if necessary.

Too little collector leads to very slow flotation of pentlandite, and too much can reduce the nickel recovery. 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.

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 sulphate.

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.

The pulp temperature during the process should be at least 6"C, but preferably it is not more than 30"C, as otherwise the cyanide 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 20"C.

The pulps treated advantageously contain from 20 to 45% solids, by weight.

By way of example, a low grade concentrate assaying 1.18% copper, 1.75% nickel and 48.1% pyrrhotite was ground to 1 3 wt.% on 38 yam in sodium carbonate solution (added to 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 litre 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 in the absence of collector to remove hydrophobic rock, yielding a product which was included in the nickel concentrate. Potassium amyl xanthate was then added in the amount of 0.1 6 grams per kg of pulp solids. The pulp was conditioned for one minute, then floated for 1 7 minutes, with a further 0.08 g/kg 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.

TABLE I Assay (%) Distribution (%) Cu Ni Pn Po Wt Cu Ni Pn Po Nickel 4.3 10.9 29.8 24.1 10.2 38.7 64.8 83.7 4.8 Conc.

Copper 13.6 2.2 5.4 37.6 4.0 47.9 5.2 5.9 2.9 Conc.

Pyrrhotite/ 0.17 0.60 0.44 55.1 85.8 13.3 30.0 10.5 92.3 Rock (Tailing) Head Grade 1.12 1.72 3.64 51.3 (Calc.) NOTE: Pentlandite (Pn) and Pyrrhotite (Po) are calculated assuming 0.8% nickel in pyrrhotite and 35.9% nickel in pentlandite.

The date of Table I show that over 92% of the pyrrhotite contained in the initial low grade concentrate was rejected and that the rejected nickel was that associated with pyrrhotite. The high rejection of pyrrhotite meant that the sulphur load on the smelter to recover desired metal values from this particular concentrate was greatly reduced.

The effect of an excess of sodium cyanide in slowing the flotation of the pentlandite and thus increasing the amount of nickel reporting in the copper concentrate is shown by the results in Table 2, which were obtained when the process just described was repeated with an addition of 1.8 g/kg of sodium cyanide and 0.24 g/kg of potassium amyl xanthate collector.

TABLE 2 Assay (%) Distribution (%) Cu Ni Pn Po Wt Cu Ni Pn Po 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 Conc.

Pyrrhotite/ 0.32 0.71 0.88 49.8 85.3 21.5 27.3 14.5 92.1 Rock (Tailing) Head Grade 1.29 2.23 5.18 46.1 (Calc.) The effect of pH is shown by the results of tests similar to the Example 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 the Example 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.

Tests similar to the Example but using, respectively an insufficient amount of collector, 0.1 8 g/kg and using an excess amount of collector, i.e. 0.30 g/kg, 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 Tables 3 and 4 following: TABLE 3 (0.18 g/kg collector Assay (%) Distribution (%) Cu Ni Pn Po Wt Cu Ni Pn Po Nickel 2.85 11.9 32.5 27.7 8.7 21.8 60.3 78.3 4.5 Conc.

Copper 11.7 1.56 3.14 53.4 6.3 64.7 5.7 5.5 6.4 Conc.

Pyrrhotite/ 0.18 0.69 0.69 55.4 85.0 13.5 34.0 16.2 89.1 Rock (Tailing) Head Grade 1.14 1.72 3.61 52.8 (Caic.) TABLE 4 (0.30 g/kg collector) Assay (%) Distribution (%) Cu Ni Pn Po Wt Cu Ni Pn Po Nickel 4.40 8.05 21.8 25.5 10.7 37.2 38.8 45.1 6.1 Conc.

Copper 7.13 9.46 25.3 43.5 8.9 50.0 37.9 43.5 8.6 Conc.

Pyrrhotite/ 0.20 0.65 0.73 47.8 80.4 12.8 23.4 11.5 85.4 Rock (Tailing) Head Grade 1.27 2.23 5.18 45.1 (Calc.) The process of the invention can be used to treat 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 sulphur load thereon is materially lowered as compared with current practice, leading to substantial reduction in sulphur dioxide emission and other economies.

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 or potassium carbonate and sodium or potassium cyanide is employed. For example, known frothers such as MIBC and DOW SA1263 may be employed. However, no substitute has been identified for cyanide ion.

Claims (11)

1. A process for the recovery of metal values from nickel sulphide ores or concentrates by froth flotation of an aqueous pulp of the ground ore or concentrate, characterised by the steps of adjusting the pulp pH to 9-10, adding cyanide ions in an amount sufficient to depress all the sulphide mineral content of the pulp, conditioning the pulp for the flotation of nickel sulphide, adding a strong sulphhydryl collector to the pulp in an amount sufficient to float the nickel sulphide, and floating and removing the nickel sulphide to provide a nickel concentrate.

2. A process according to claim 1 applied to a pulp that contains pentlandite, chalcopyrite, pyrrhotite and gangue.

3. A process according to claim 2, characterised in that the pulp pH is adjusted to 9-10 by adding sodium carbonate during grinding of the pulp.

4. A process according to claim 2 or claim 3, characterised in that the amount of cyanide added is such that the pulp redox potential (measured with the use of a gold v. saturated calomel electrode) is reduced to a value not lower than - 390 mV.

5. A process according to claim 4 characterised in that the pulp is conditioned for the flotation of pentlandite by agitation with aeration to raise the pulp redox potential to - 330 mV, a xanthate collector is added, and pentlandite is floated and removed.

6. A process according to claim 5, characterised in that after flotation of the pentlandite the pulp redox potential is raised to - 240 mV and chalcopyrite is floated and removed.

7. A process according to claim 6, characterised in that after flotation of the chalcopyrite the pulp redox potential is raised to - 200 mV and pyrrhotite is floated and removed.

8. A process according to any preceding claim, characterised in that the collector is potassium amyl xanthate.

9. A process according to any preceding claim, characterised in that the pulp temperature is about 20"C.

10. A process according to any preceding claim, characterised in that the pulp contains from 20 to 45% solids, by weight.

11. A process according to claim 1 substantially as hereinbefore described with reference to the Example.