Note: Descriptions are shown in the official language in which they were submitted.
lls~3~n
-1- PC-2115
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
-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
-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~)
-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
-- 5
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
-6- PC-2115
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)
-7- PC-2115
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.
EXAMPLE 1
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
-8- PC-2115
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
-9- PC-2115
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
-10- PC-2115
EXAMPLE 2
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
-11- PC-2115
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.)
SABLE 4
(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
-12- PC-2115
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.
