Note : Les descriptions sont présentées dans la langue officielle dans laquelle elles ont été soumises.
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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.
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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
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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.
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~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.
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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:
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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
. ' , ~ ,
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1104274
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%
TABLE 1
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.
TABLE 2
. .
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.
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
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~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.
TABLE 4
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.
TABLE 5
. -
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
-12-
~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. - -
