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An evaluation of pretreatments to increase gold recovery from a refractory ore containing arsenopyrite and pyrrhotite
0892--6875(99)00145-4
Mnera1.v Engincermng, Vol. 13, No. 1, pp. l-18, 2000 0 1999 Elsevier Science Ltd All rights reserved 0892-6875/00/$ - see front matter
AN EVALUATION OF PRETREATMENTS TO INCREASE GOLD RECOVERY FROM A REFRACTORY ORE CONTAINING ARSENOPYRITE AND PYRRHOTITE
M. N. LEHMANN”, S. O’LEARY?, A. J. Parker Cooperative
Research Center for Hydrometalhngy.
5 School of Applied Chemistry,
Curtin University
and J. G. DUNN@” E-mail: [email protected]
of Technology,
Perth, Western Australia
7 Mineral and Processing Laboratories of the Department of Energy, Waterford, (Received 5 July 1999; accepted 10 September 1999)
Western Australia
ABSTRACT The effect of three ore pretreatments on gold recoveries from an ore in which the gold existed in solid solution in an arsenopyrite matrix, as well as being encapsulated in a pyrrhotite phase, was examined. The three pretreatments evaluated included fine grinding, pyrolysis-roast (Pyrox), and an oxidative pressure acid leach (PAL). It was found that untreated ores did not yield high gold recoveries even when a fine grind (~45~) was employed, while oxidative treatments led to a signtjicant improvement of recoveries. The Pyrox pretreatment was found to decompose the sulfide matrix of the ore to a greater extent than a 3 hour PAL treatmemt. However, the PAL process produced a slightly better gold recovery of 79% compared to the Pyrox recovery of 75%, and both of these were greater than the 55-60% recoveries achieved with the untreated ore. Furthermore, the depletion of dissolved oxygen levels observed during cyanide leaching were signtj?cantly reduced by the two oxidative treatments, compared to the untreated ores. PAL was found to benefit over the Pyrox pretreatment by afording a solid product, which eflectively “locked” the arsenic component of the mineral from being recovered in solution. The PAL process required more lime for neutralization of acid produced from the decomposition reaction. 0 1999 Elsevier Science Ltd. All rights reserved. Keywords Gold ores; sulfide ores: roasting; grinding;
pressure leaching
INTRODUCTION Gold deposits can be broadly classed into two categories, primary and secondary. The primary deposits are sulfides, mainly pyrite and arsenopyrite, and more rarely other gold bearing sulfides such as pyrrhotite. Direct cyanide leaching of the secondary deposits give good gold recoveries in excess of 95%, but direct leaching of the primary deposits give poor recoveries of 5-70%, depending on the mineralogical composition of the deposit and the form of the gold in the host. The difficult-to-treat concentrates are described as “refractory”. There are a number of reasons why ores are refractory. It has been demonstrated that gold can occur in sulfide matrices in a range of sizes, from so-called “invisible” or solid solution gold (Cabri et al., 1989), to particles of 50-100 pm. Although the matrix offers particulate gold against contact with cyanide, dissolved oxygen and leaching
some protection to the reagents, fine grinding
* Presently employed with AMIRA gold project, Murdoch University, Murdoch, Australia ** Present address: Dept. of Chemistry, University of Toledo, Ohio, USA
2
M.N.Lehmannetal.
followed by cyanide leaching can sometimes produce acceptable recoveries (Espiell et al., 1986). However, gold in solid solution can only be effectively recovered if the matrix is altered by some chemical or biological pretreatment. Furthermore, sulfides (including arsenopyrite) will oxidatively decompose during cyanide leaching, and utilise reagents normally required for the dissolution of gold. For example, if arsenopyrite is leached directly in aerated cyanide solutions containing lime, arsenites may be formed according to the reaction in Eq. 1 (Nagy et al., 1966). This reaction consumes oxygen as well as lime (Eq. l), and contributes to reduced gold recoveries. 4FeAsS + 4Ca(OH)* + 1102 (s) +
4FeS04 (aq)+ 4CaHAs03 (Q + HZ0
Any pyrite or pyrrhotite in the ore would also be expected cyanide and oxygen (Eqs. 2-4).
&%+I
cs)
+ CN- +
nFeSe, + SCN- (aq)
4FeS,,, + 302 +6H20 + 4FeS,,, + 902 + 10&O +
4Fe(OHb
to consume
(1) substantial
quantities
of both
(2)
($)+ 4S”,,,
4Fe(OH),(,) + 4SO,2-,,,
(3) + 8H+,,,,
(4)
The solution to these problems is the implementation of an oxidative pretreatment which will completely break down the sulfide matrix and release the gold from solid solution (Demopoulos and Papangelakis, 1989; Dry and Coetzee, 1986; Dunn et al., 1989; Kontopoulos and Stefanakis, 1990; Linge, 1992; Udupa et al., 1990; Weir and Berezowsky, 1986; Weir et al., 1996). A prior concentration step such as flotation would also be required to minimise the amount of material treated, and thus make the added treatment cost effective (Swash, 1988). The aim of this investigation was to undertake a comparative evaluation of pretreatments on the recoveries obtained after cyanide leaching of a well characterised natural ore body, where the majority of the gold is located as invisible gold in an arsenopyrite matrix. The pretreatments that were evaluated included fine grinding the ore, a two stage pyrolysis/oxidative roast treatment (Pyrox) (Avraamides et al., 1991; Dry and Coetzee, 1986; Dunn, 1997; Dunn and Chamberlain, 1997; Dunn et al., 1989; Dunn et al., 1995b; Ferreira et al., 1989; Graham et al., 1992), and an acidic oxidative pressure acid leach (PAL) (Berezowsky and Weir, 1983; Dziurdzak et al., 1989; Papangelakis and Demopoulis, 1990a; Papangelakis and Demopoulis, 1990b; Robinson and Williams, 1992; Thomas, 1991; Weir and Berezowsky, 1986; Weir et al., 1996)
EXPERIMENTAL Samples Core samples of a gold ore were assessed for overall general mineralogy using powder X-ray diffraction (XRD). Scanning electron microscopy (SEM) and electron probe microanalysis (EPMA) were employed assess the native gold and/or telluride content, and solid solution gold in arsenopyrite and other sulfides.
to
The ore samples were made from a riffle split blend of five crushed core samples which were mixed to yield a sample with a theoretical gold head grade of 7.5 ppm. The blended sample was milled to the three size fractions of <45 urn, ~75 pm, and cl25 urn. The <75 urn fraction was divided into three quantities, two of which were oxidised prior to cyanide leaching. During storage the ground samples were sealed in air proof plastic bags and refrigerated.
Evaluation of pretreatments to increase gold recovery from refractory ore
Oxidative
3
pretreatments
The pyrolysis/roasting (Pyrox) pretreatment was carried out by heating about 5Og of ~75 um’ore under nitrogen in a furnace at 700°C for a period of 30 minutes, followed by heating in air at the same temperature for 3 hours. The acid pressure leach process was carried out as follows. 400g of the ore (~75 pm) was combined with 2.0 L of H2SO4 (10 g/L) to yield a mixture having a pulp density of 16.7%. The mix was placed in a 3.2 litre capacity autoclave and agitated at 400 rpm. The vessel was sealed and the head space purged with about 300 kPa nitrogen, leaving a residual 25 kPa of nitrogen gas overpressure. The sample temperature was then ramped to 180°C. On reaching the target temperature, 1500 kPa of oxygen gas was introduced to give a total vessel pressure of 2600 kPa. At this point the reaction timer was started. The experiment was conducted over three hours and intermediate samples were taken at 0.5, 1.0, 1.5, 2.0, 2.5, and 3.0 hours (testing for iron and sulfur in solution/solid phase and free acid in solution). The solutions were analysed for acidity (ie. free H2S04), iron by atomic absorption spectroscopy (AAS) and total sulfur (AAS). At the end of the experiment the system was quickly cooled to 30°C. The residual slurry was removed from the autoclave and washed. This involved 12 stages of a counter-current decantation, using approximately 240 L of tap water. The residue was then filtered and dried. Cyanide
leaching
A day before cyanide leaching, pretreated and untreated ore samples (50g) were transferred to 2 litre capacity plastic bottles fitted with screw cap lids. Each lid was perforated with a hole to allow air to enter. Deionised water (150 ml) was added to the solids, and the pH of the solution was adjusted to 10.5 by the addition of calcium hydroxide (lime). Additionally a sample of iron(II1) oxide obtained from Sigma chemicals (50g) was treated in a similar manner as a control for the experiment. The bottles were placed on a bottle roller and rotated at 60 rpm for 24 hours. Prior to actual leaching, and if needed, the pH of the prepared pulps was readjusted to 10.5. Leaching commenced after the addition of 50 mL of 6000 ppm sodium cyanide solution, which afforded a 20% w/v pulp with an initial cyanide concentration of 1500 ppm. The bottles were immediately placed on a bottle roller and rotated at 60 rpm for a period of 64 hours. During this time the pH of the solution was monitored intermittently with a Hanna HI1230 pH electrode and, where needed, maintained at a pH of 10.5 with the addition of lime. The mass of lime added to the solutions was recorded. The cyanide concentration of the pulp was monitored with an Ionode I/CN/Hg ion selective electrode connected to a TPS digital pH meter referenced to a Philips Rl 1 glass silver/saturated silver electrode, and readjusted to maintain the cyanide concentration at 1500 ppm. The oxygen concentration of the pulps was monitored utilising an Activon Model 40 1 oxygen meter. Sampling of the dissolved species present in the leach liquor was carried out by intermittently removing the plastic bottles from the roller and allowing the solids to settle, after which a sample of the supematant (3 ml) was removed with an auto pipette, filtered, and its weight recorded. The solution was then acidified with the addition of 2M HCl (10 ml) and oxidised with 20% v/v hydrogen peroxide (5 ml). The solution volume was reduced on a hotplate and made up to 25 ml in a volumetric flask. Leached gold, iron and sulfur were determined by ICP-OES (optical emission spectroscopy). The head grade of gold, arsenic, iron and sulfur in the solids before and after both pretreatments and cyanide leaching was determined by fire assay (Au) and ICP-OES (As, Fe & S) on an aqua-regia digest. Additionally, a sample of filtered liquor obtained after leaching the untreated ore sample (~75 urn), was analysed by ion chromatography in order to determine the ions present in solution.
4
M. N. Lehmann et al.
RESULTS
AND DISCUSSION
Mineralogy The sample was first checked for graphite, which in fine-grained form is responsible for preg-robbing. Optical microscopy of the untreated ore samples showed no sign of carbonaceous material, which is normally detectable by its anisotropic interaction to polarised light. It seems certain, therefore, that pregrobbing is not a cause of refractory behaviour in the ore. XRD analysis of the cl25 pm size fractions revealed that quartz was the most significant phase present in The next most abundant phases were clinochlore each of the five core samples. (Mg,Fe,A1)6(Si,A1)4010(OH)8);chlorite (Mg,Fe)S(Al,Si)S010(OH)8;and muscovite (KA12(Si3Al)0,0(OH.F)2. Quantities of hexagonal pyrrhotite (FeS), and arsenopyrite (FeAsS), were also found. These were less abundant than any of the silicate materials. Pyrite and galena, as well as gold, were not detected at the 1% detection level of the technique. An extensive particle to particle analysis of the ~125 urn ground ore samples, utilising SEM and EPMA, revealed that the ore contained significant quantities of both arsenopyrite and pyrrhotite. The maximum possible pyrrhotite contents were calculated to be 11% and 10% respectively. The trace results using the electron microprobe were derived from 6 minute counts at a current of 450 nA. Pyrite was also found to be present, as was the occasional grain of chalcopyrite (CuFe&) and galena (PbS), and an iron arsenide, possibly FeAs*. The gangue minerals were largely undifferentiated by brightness in the SEM, and were characterised by their qualitative energy-dispersive x-ray spectra. The major minerals found included quartz, a potassium aluminium silicate with varying amounts of iron, an iron aluminium silicate, and an iron manganese calcium aluminium silicate. Additionally grains of titanium were found mainly present as sphene, with very occasional grains of ilmenite. Siderite was the only carbonate definitely identified, but in general carbonates were elusive. Gold was only observed
in large particles
which had avoided
fine crushing,
and was consistent
with a
native gold content of about 5 & 10 ppm. Despite a long search, gold or tellurides were not observed. Arsenopyrite particles were analysed for trace gold, with the result that some arsenopyrite contained up to 250 ppm of gold, but most contained only a few ppm. The distribution was too irregular to allow quantification for the limited number of analyses conducted, but is a characteristic observation for solid solution gold in arsenopyrite. Measurements on several particles of pyrrhotite indicated that some of the gold present was most likely encapsulated
in this phase, and had a gold content of up to 5 & 10 ppm.
PAL pretreatment The conditions
for pressure
oxidation
were selected on the basis of prior studies on ores having
similar
sulfide matrices. Pressure oxidation was carried out for a period of 3 hours at 180 ‘C with an oxygen overpressure of 1500 kPa. Table 1 shows the mass transfer of sulfur and iron from the solid sulfide matrix to solution as the reaction occurs. Figure 1 shows that over a three-hour period, a 75% decomposition of the sulfide was achieved. Just over two thirds of this maximum breakdown occurred in the first half hour, and then the rate slowed considerably. If this second slower rate was maintained, it would take about a further 3 hours to decompose the sample completely. Figure 2 shows that the oxidation produces free acid. This is favourable in terms of increasing the rate of sulfide matrix breakdown. However, an effective residue wash is required to limit the amount of lime required for neutralisation to the pH value needed for cyanide leaching.
Evaluation of pretreatments to increase gold recovery from refractory ore
5
TABLE 1 Solution and solid breakdown occurring during oxidative pressure acid leach Time
@ours) 0 0.5 1 1.5 2 2.5 3
Solid
Free Acid G/L 10 19.6 23.0 25.7 26.9 27.3 27.8
Fe % 16.2 10.7 11.1 10.3 10.0 10.6 10.3
aaL.......,.....,
0
“.
Solution s % Fe % 0.33 0.00 1.03 2.22 1.80 1.27 1.67 1.35 1.65 1.40 1.62 1.45 1.62 1.46
S% 8.79 3.82 5.38 3.36 4.15 2.33 1.48 .‘.“.““‘.“‘>“
.
”
0.5
“,
““. 1
,.,
” 1.5
% Sulfide Decomposition 0 40 55 62 66 72 75
”
2
”
a..
2.5
”
”
”
3
”
3.5
Fig. 1 Total sulfide matrix breakdown as calculated from sulfur leached into solution.
Total Free Acid
Time (hours)
Fig.2 Acid concentrations measured during oxidative pressure acid leach.
M. N. Lehmann et al.
6
Pyrox pretreatment Figure 3 shows the thermogram pyrolysis.
of a small sample (24 mg) of the <75 urn ore sample obtained
'\ :
24 -
during
\ '1 \‘;\ \
235
'\
g
\
:: g
\
23.
\ i \! :\
_
225
22
0
~'~'~"~ 100
200
300
400
I"""""'! 500 600
Temperature
Fig.3 Mass loss of an untreated atmosphere).
ore sample
(175
700
800
900
1000
(Celcius)
pm) during
pyrolysis
(Heating
rate, 20’C/min,
N2
Decomposition, as monitored by weight loss, commenced at 500 “C, and a maximum weight loss is reached at temperatures greater than 720 “C. This supports the assumption that a maximum breakdown of the sulfide matrix would be expected under the conditions employed in the Pyrox pretreatment. The chemical analysis results shown in Table 2 indicate that significant quantities of both arsenic (ca. 96%) and sulfur (ca. 96%) were lost from the ore after the Pyrox treatment. TABLE 2 Elemental
Before Pyrox After Pyrox
compositions
Arsenic (kg/t) 49 2
before and after Pyrox treatment Sulfur (kg/t) 74 3
Gold (g/t) 6.61 7.82
These results suggest that the Pyrox pretreatment affords a greater decomposition of the sulfide matrix compared to the three hour oxidative pressure acid leach. The rise in the gold concentration after treatment is expected on the basis that the samples undergo weight-loss after the Pyrox treatment, while gold is conserved in the sample. Gold recoveries
by cyanide leaching
Table 3 shows the gold recoveries achieved those pretreated by Pyrox and PAL.
for the untreated
ores for a range of size fractions,
Figure 4 shows the leaching profile of gold during the period of cyanidation.
as well as
Evaluation
of pretreatments
TABLE 3 Gold recoveries
to increase gold recovery
of pretreated
and untreated
from refractory
ore
ores obtained from cyanidation
100
T
7
60
-----__-.._
----__.________~
-2
5 i?
$ 8
75pm
T
.-.--
PAL
1
___
45 pm Untreated .--------_-__.___ 75 blrn Untreated
~_~ ____ _.~___ _~__~
6o
125 pm Untreate
2 s
s
40
20
0 10
20
30
40
50
60
70
Time (hours)
Fig.4 The leaching profile of gold from pretreated and untreated
ores during cyanidation.
It can be concluded that gold recoveries are significantly improved with a pretreatment. The solution analysis of the gold indicated that after a period of 10 hours 79% and 75% of the gold was found to be leached from both the PAL and Pyrox pretreated ores respectively, as compared to 60%, 57%, and 55% of the untreated ores, listed in increasing particle size. The fact that recoveries from a range of particle sizes are not only low but similar (55-60%) suggests that fine grinding alone would not significantly improve recoveries, and is consistent with the gold being trapped in a solid solution form. In batch leaching the rate of leaching is described by Eq. 5 (Nicol, 1988);
8
--
M. N. Lehmann et al.
4Aul dt
=
kl{[Au]-[Au]e}2
Equation 5 can be integrated function of time; 1
[Au]- [Au1e
=kl.t
+
(5) to give Eq. 6, which describes the concentration
of undissolved
gold as a
1
(6)
[Au1o - [Au1e
where [Au] is the concentration of gold in the ore at any time t, [Au]~ and [Au]~ are the initial and equilibrium gold concentrations in the solid respectively, while k, is the leaching rate constant. From Eq. 6, the rate constant for the cyanidation of the ~45 urn untreated ore can be obtained by plotting the data of Table 4 as shown in Figure 5. TABLE 4 Data for the determination
of leaching rate constant kl for the ~45 pm untreated
*For the above calculations
ore
[AulBt 61hrs.= [Au]~ = 2.25 g/t
Time (hours) Fig.5 Plot of l/( [Au] - [Au]~ ) vs time for the determination 18.5 hours of cyanidation of the ~45 urn untreated ore.
of the leaching rate constant
k, for the first
Evaluation of pretreatments to increase gold recovery from refractory ore
Table 5 lists the leaching rate constants obtained from the initial stages equilibrium concentrations of the gold at the completion of the leach time. TABLE 5 Leaching rate constants stages of cyanidation
(k,) for treated
Ore Blend Untreated
ore ~45 pm Untreated ore ~75 pm Untreated ore ~125 urn Pyrox treated ore ~75 pm PAL treated ore <75 urn
and untreated
9
of leaching,
ores obtained
utilising
during
Leaching rate constant (tg-‘h-l) 0.10
0.98
0.28 0.21 0.53
0.96 0.97 0.90
0.32
0.96
the
the initial
R2
It is apparent that over extended time the PAL treatment yields slightly higher gold recoveries than the Pyrox treatment. These are reflected in the final recoveries of 85% and 75% for the PAL and Pyrox pretreatments respectively, as determined by fire assay. However, as displayed by Table 5, gold is leached out of the Pyrox treated ore more quickly compared to the PAL ore, as indicated by the greater leaching rate constant of the former. Figure 4 shows that the Pyrox treated ore exhibits a slight reduction of the gold in solution at extended leach times, which may be due to the reprecipitation of gold. The leaching rate constants for the untreated ores are significantly less than the treated ores. The <45 pm ore was found to have a leaching rate constant less than for the <75 and ~125 pm ores, and although unexpected might be accounted for by the small particles forming a thicker impervious iron oxide coating during milling. Since sulfur in the cyanide leach liquor originates from the sulfide component of the ore, a measure of its concentration gives an indication of the leach rate of the sulfide matrix component in untreated ores. In treated ores it gives evidence as to the level of completion of the oxidative pretreatment. The importance of the amount of sulfur in an ore that can undergo dissolution is dependent on the nature of the sulfur species that will be dissolved. In general, sulfur species other than sulfate can react with cyanide in a variety of ways and act as cyanicides. For example, cyanide readily reacts non-oxidatively with elemental sulfur to form thiocyanate (Eq. 7) as do a range of sulfoxy ions such as thiosulfate (Eq. 8) and tetrathionate ions (Eq. 9). Under oxidative conditions sulfide ions released from non-oxidative mineral sulfide dissolution can be oxidised to thiosulphate ions (Eq. lo), to then form thiocyanate ions via Eq. 8. S + CN- +
SCN-
(7)
S2032- + CN- +
SCN- + S032-
(8)
S4062- + CN- +
2SCN- + 2S032-
(9)
S*- + O2 +H20 +
S20s2- + 20H-
(10)
Although gold can form stable complexes with thiocyanate, at the pH employed for cyanidation, not complex with thiocyanate to any appreciable extent (Marsden and House, 1992).
gold does
Table 6 presents the quantity of sulfur dissolved for the pretreated and untreated ores, while Figure 6 shows the rate of sulfur dissolution into the cyanide liquor.
IO
M. N. Lehmann et al.
TABLE 6 Dissolved
sulfur levels of pretreated
00”“‘~‘~“‘~“‘~‘~‘~~~““‘~‘~~~“~ 0 10 20
30
and untreated
40
ores obtained from cyanidatiou
50
60
70
Time (hours)
Fig.6 The extent of sulfur dissolution
from pretreated and untreated
ores during cyanidation.
Each material (untreated and treated) was conditioned in deionised water for 24 hours, and the pH of the solution adjusted to 10.5 prior to the addition of cyanide solution. In the first 24 hours of cyanide leaching, all but the Pyrox treated ore show increases in the concentration of sulfur in the leach liquor. The Pyrox sulfur level in solution was relatively high to start with, indicating that water soluble sulfur compounds were present in the roasted product. This may have been entrained S02, or compounds such as calcium sulfate formed during the roasting step. The sulfur level was relatively constant over the leach period. Furthermore, Table 6 shows that almost all the sulfur in the matrix is leached. Both solution and residue analysis showed that the ~75 pm leached a greater amount of sulfur than the ~45 pm or the cl25 pm. This anomalous observation may be due to the fact that the ore is comprised of a variety of minerals, each with differing reactivities. Consequently, at certain particle sizes the dissolution of certain fractions of the ore may be favoured over others. For example pyrrhotite readily oxidises in air to form protective coatings (Pratt, et al.,) which may either be thicker or more uniform at certain particle sizes. However, the overall
Evaluation of pretreatments to increase gold recovery from refractory ore
11
ncrease in dissolved sulfur levels indicates that the sulfidic components of the ore are oxidising under the conditions of cyanide leaching. Figure 7 shows the dissolved oxygen (DO) levels during cyanide leaching. During the initial stages, both treated and untreated ores showed a significant decrease in the levels of DO. The most extreme decreases were experienced by the ~45 and ~75 urn untreated ores, and these results are consistent with the sulfide minerals consuming oxygen during oxidation (Dunn et al., 1995a). Table 7 shows the relative concentrations of soluble ions found after leaching the ~75 urn ore for the duration of the test under normal bottle roll conditions, and additionally when oxygen gas is sparged through the solution, instead of air, to promote further oxidation.
1, 75 pm Pyrox 2. 75 pm PAL 3. 4. 5. 6.
cl
10
20
125 pm Untreated Control 75 pm Untreated 45 pm Untreated
30
._
40
_^
50
^^
BU
IO
Time (hours)
Fig.7 Dissolved oxygen levels during cyanidation of pretreated and untreated ores.
TABLE 7 Concentration untreated
of major ions found in solution ore for a period of 64 hours
after cyanidation
of the minus
75 pm
12
M. N. Lehmann et al.
The major soluble product of mineral sulfide oxidation during cyanidation is the thiocyanate ion. As expected, the yield of this product rises with increasing oxidation of the mineral. In contrast, the dissolution of product ions which are cyanicides, such as copper and nickel, are inhibited when the oxidation potential of the solution is increased and is due to the oxidation of these ions to insoluble hydroxy carbonates such as malachite (Eq. 11). A reduction in the concentration of these ions would be expected if the CIP plant is sparged with oxygen gas rather than air (Muir et al., 1988). This is a positive outcome, as complexes of copper and nickel are known to catalyse the oxidation of cyanide.
2Cu(CN),-,,,,
+ ; O2 + 9H20 +
Cu(OH)&uC03
(s) + 3CO,‘-(,,, + 4NH‘,+ (aq)
(11)
The concentration of the ferro- and ferricyanide complexes in solution are also noticeably less significant than the concentration of thiocyanate ions. If iron hydroxide/oxides are initially formed from oxidative pathways, then the kinetics of subsequent dissolution in solutions containing cyanide ions (Eq. 12) must be slow. This would explain the absence of a large amount of the former, even though many reports have et al., 1984; postulated on the thermodynamic stability of these iron cyanide complexes (Osseo-Asare Wang and Forssberg, 1990). Fe(OH)3 cs)+ 6CN- caq)+
Fe(CN)b3- caq)+ 3OH. caq)
(12)
Figures 8 and 9 show the cumulative cyanide additions made to the untreated and treated ores, respectively, in order to maintain the cyanide concentration at 1500 ppm. In the initial 8 hours of leach there was little difference in the amount of cyanide consumed by any of the samples. After this time differentiation in cyanide consumption became evident. Notably, the Pyrox treated ore consumed significant quantities of cyanide with respect to the PAL and the untreated ores, and may be related to the elevated levels of dissolved iron that were observed in the leach, as shown in Figure 10. Interestingly, during the initial stages of cyanide leaching of the Pyrox treated sample, iron is readily solubilised and reprecipitated. Such behaviour is not observed with the PAL or untreated ores.
0.4""""'."'."""',""'."."" 0 10 20
30
40
50
60
7(
Time (hours)
Fig.8 Cyanide consumed during the cyanidation 1500 ppm (kg of NaCN / tonne ore).
of untreated
ores to maintain
the cyanide concentration
at
Evaluation of pretreatments to increase gold recovery from refractory ore
0
10
20
50 Tilt!
60
70
(houf$
Fig.9 Cyanide consumed during the cyanidation of treated ores to maintain the cyanide concentration at 1500 ppm (kg of NaCN / tonne ore). .
The dissolution of arsenic from the untreated ores is most likely due to the formation of soluble arsenates arising from reactions similar to Eq. 13, and also may arise from resolubilisation of arsenites formed from the reaction described by Eq. 1. FeAsS (sj+ 702 + 2H20 -_) 4HAs02 (aqj+ 4Fe+’caqj+ 4SOa2- (aqj
(13)
Ferrous ions produced from oxidation of iron sulfides may be further oxidised with oxygen to yield ferric ions (Eq. 14) (Hiskey and Schlitt, 1981). The ferric ions produced can combine with arsenate to form ferric arsenate, which precipitates as crystalline scorodite (Eq. 15). Furthermore, it is this reaction which is thought to contribute largely to the free acid produced during pressure oxidation of FeAsS as shown in Figure 2 (Bailey and Peters, 1976; Papangelakis and Demopoulis, 199Oa). + o2 4Fe2+(,,)
+
Fe3+(,,) + H&Q
4H’,,,, + 4Fe3+(,) + 2H20 caq)
+ 2H20 + FeAs04.2HzO cs)+ 3H’,,,
(14) (15)
Consequently, the observation that the iron concentration for the PAL ore initially starts low and remains comparatively low throughout the leach, is most likely a result of it being bound in scorodite. This also explains why the PAL treated ore had the highest head grade of iron (Table 8), and that throughout the leach little iron was dissolved (Figure 10). Table 9 shows the low arsenic recoveries obtained fi-om cyanide leaching of the PAL treated ores as compared to those untreated, even though the initial head grades are the same. This further confirms that after PAL, the arsenic is “locked” into the scorodite matrix, which is desirable as it is nonreactive. Although the stability of this complex has been questioned, provided that the molar ratio of ironzsenic is four, the ferric arsenate is stable enough for landfill disposal (Haines, 1986). The Pyrox treated ore is also observed to yield a low arsenic recovery, which is not unexpected given the low residual arsenic levels remaining after the pyrolysis reaction.
14
M. N. Lehmann et al.
TABLE 8 Dissolved
iron levels of pretreated
and untreated
ores obtained from cyanidation
50
40
45 pm Untreated f-
30
h 75 pm PAL
/1
20
IO 125 pm Untreatf
f
1 30
40
50
60
70
Time (hours)
Fig. 10 The extent of iron dissolution
TABLE 9 Dissolved
from pretreated and untreated
arsenic levels of pretreated
and untreated
ores during cyanidation.
ores obtained from cyanidation
Evaluation
of pretreatments
to increase gold recovery from refractory
ore
15
The PAL treated ore showed a continual increase in the amount of sulfur released into the leach liquor. However, the percentage of sulfur recovered in the tail was much higher than for the Pyrox treated ore but lower than for the untreated ores. The oxygen consumption during cyanidation was also comparatively low for the PAL treated ore, but the consumption of lime over the leach period was high with respect to the untreated ores but lower than the Pyrox treated ore. It would be expected that any sulfide present in the ore would be oxidised in the PAL process to a soluble sulfate species. It is curious that much of the residual sulfur in the head material was retained after cyanidation, and yet a relatively high cyanide consumption was recorded, compared to the untreated ores. One possibility is that the consumption of cyanide is due to either an insoluble polysultide species, or elemental sulfur retained in the PAL product. This is in agreement with the observation of Dunn et al. (1996) who postulated that a major product of oxidation must be elemental sulfur, in order to stoichiometrically balance the oxidation of arsenopyrite in aqueous solutions. It would be expected that if it was merely a sulfur layer that the rate of dissolution would be quite rapid. However, this is not the case, and hence the sulfur must also be integrated into either the iron oxide or scorodite product layers. Figure 11 shows that the PAL treated ore also had the highest consumption of lime during the leach, which is consistent with an iron oxide product layer undergoing reactions similar to those described by Eqs. 16 and 17. Furthermore, an initially large lime addition might have been required to neutralise any residual acid in the PAL product, in a reaction described by Eq. 18. FeO,,, + Hz0 -_) FeOOH,,, + H+ + 1em
(16)
FeO(,, + OH- +
(17)
30H-,,,,
FeOOHe, + 1e-
+ H3As04 (aq)+
As0d3- (aq)+ 3H20
(18)
Time (hours)
Fig. 11 Lime consumed during the cyanidation Ca(OH)2 / tonne ore).
of ores to maintain
the slurry pH at a value of 10.5
(kg
The solubilisation and reprecipitation of iron for the Pyrox treated ore in Figure 10 is also observed in the control material hematite and can be thought of as being due to a combination of hydration and dehydration reactions which might follow a pathway as described by Eq. 19. Fez03 cs)+ 3H20 --+ 2Fe(OH)3
(sacurated sown.) +
2FeOOH
cs) + 2&O
(19)
16
M. N. Lehmann et al.
CONCLUSION This study has shown that in order to obtain appreciable gold recoveries from a gold ore, where the gold is located as “invisible” gold in a sulfide matrix composed mainly of arsenopyrite and pyrrhotite, an oxidative pretreatment is required prior to cyanide leaching. Fine grinding (~45 pm) of this ore type did not lead to a significant improvement in the gold recovery. A cyanide leach of untreated ores yielded soluble arsenic ions. As a consequence, if this treatment was to be adopted, the liquor from the leach would require further treatment to remove these ions. The oxygen consumption of the PAL and Pyrox treated ores was significantly lower during leaching than for the untreated ores, while the leaching rates of the ores was significantly increased with pretreatment. Of the oxidative pretreatments examined, the PAL treatment yielded a gold recovery of 79% after 10 hours which is slightly better than the 75% achieved after the Pyrox process, and appreciably better than the 55-60% recovered without treatment. The Pyrox treated ore was observed to have the faster leach rate and decomposed the sulfide matrix of the ore to a greater extent than the PAL treatment. PAL, however, afforded a solid product which effectively “locked” the arsenic component of the mineral from being recovered in solution. As a consequence, lime consumption during cyanide leaching was significantly higher for the PAL treated ores than for Pyrox or untreated ores. The choice of which oxidative pretreatment to employ commercially will have to be balanced by the fact that soluble arsenic will be evolved during leaching of the Pyrox treated ore. While this is not a problem with the PAL treated ores, substantially more lime will be needed in order to leach the latter.
ACKNOWLEDGMENTS The authors wish to thank Dr J. Graham for conducting the mineralogical assessment of the ore used in this study. This research has been supported by the Australian Government’s Cooperative Research Centres Program, through the A J Parker CRC for Hydrometallugry.
REFERENCES Avraamides, J., Nguyen, G., Dunn, J. and Graham, J., ‘Roasting and Cyanide Leaching of a Gold Bearing Pyritic Concentrate from Kalgoorlie, Western Australia’, In Proc. of World Gold 91, Cairns, Queensland, 1991, pp 45-53. Bailey, L.K. and Peters, E., Decomposition of Pyrite in Acids by Pressure’Leaching and Anodization: The case for an electrochemical mechanism., Canadian Metallurgical Quarterly, 1976, 15(4), 333-344. Berezowsky, R.M.G.S. and Weir, D.R., ‘Pressure Oxidation Pretreatment of Refractory Gold’, In Proc. of Practical Hydrometallurgy 83: Annual Symposium on Uranium and Precious Metals, Fort Saskatchewan, Alberta Canada, 1983, pp 101-104. Cabri, L.J., Chryssoulis, S.L., Villiers, d.J.P.R., Laflamme, J.H.G. and Buseck, P.R., The Nature of Invisible Gold, The Canadian Mineralogist, 1989, 27(3), 353-362. Demopoulos, G.P. and Papangelakis, V.G., Recent Advances in Refractory Gold Processing, CIM Bulletin,, 1989,82(931), 85-91. Dry, M.J. and Coetzee, C.F.B., ‘The Recovery of Gold from Refractory Ores’, In Proc. oflnternational Conference on Gold, Johannesburg, S. Africa, 1986, pp 259-274. Dunn, J.G., The Oxidation of Sulfide Minerals, Thermochimica Acta, 1997,300(1-2) 127-139. DUM, J.G. and Chamberlain, A.C., The Recovery of Gold From Refractory Arsenopyrite Concentrates By Minerals Engineering, 1997, 10(9), 919-928. Pyrolysis-Oxidation, Dunn, J.G., Femandez, P.G., Hughes, H.C. and Linge, H.G., ‘Aqueous Oxidation of Arsenopyrite in Acid’, In Proc. of A&MM Annual Conference, Perth-Kalgoorlie, W. Australia, 1996, pp 2 17-220. Dunn, J.G., Graham, J. and Linge, H.G., The Treatment of Refractory Gold, Chemistry in Australia, 1989, 401-404. DUM, J.G., Hollbach, D.J. and Ibrado, A.S., ‘The Dissolution of Pyrrhotite in Cyanide Solution’, In Proc. of Rand01 Gold Forum, Perth, W. Australia, 1995a, pp 145-148.
Evaluation of pretreatments to increase gold recovery from refractory ore
17
Dunn, J.G., lbrado, A.S. and Graham, J., Pyrolysis of Arsenopyrite for Gold Recovery by Cyanidation, Minerals Engineering, 1995b, 8(4/5), 459-471. Dziurdzak, G., Kyle, J.H. and Dunne, R.C., ‘The Pressure Aqueous Pre-Oxidation of Refractory Gold Ore from the Golden Mile, Kalgoorlie, Western Australia’, In Proc. of World Gold 89: Forum on Technoloa and Practices, Reno, Nevada, 1989, pp 3 15-32 1. Espiell, F., Rota, A., Cruells, M. and Nunez, C., Gold and Silver Recovery by Cyanidation of Arsenopyrite Ore, Hydrometallurgv,, 1986, 16, 141-151. Ferreira, P.M., Knott-Craig, K. and Boydell, D.W., Roasting of Refractory Au Concentrates at Consort. The new two-stage fluidized-bed roasting plant has negligible adverse impact on the environment, Engineering and Mining Journal, 1989, ,26-30. Graham, J., Fletcher, A., Walker, G. and Aylmore, M., ‘A Pyrolysis Approach to Refractory Gold Processing’, In Proc. of The International Conference on Extractive Metallqy of Gold and Base Metals, Kalgoorlie, W. Australia, 1992, pp 441-443. Haines, A.K., ‘Factors Influencing the Choice of Technology for the Recovery of Gold from Refractory Arsenical Ores’, In Proc. of GOLD 100. International Conference on Gold, Johannesburg, S. Africa, 1986, pp 227-233. Hiskey, J.B. and Schlitt, W.J., ‘Aqueous Oxidation of Pyrite’, In Proc. of Interfacing Technologies in Solution Mining: The 2nd SME-SPE International Solution Mining Symposium, Denver, Colorado, 198 1, pp 55-74. Kontopoulos, A. and Stefanakis, M., ‘Process Options for Refractory Sulfide Gold Ores: Technical, Environmental and Economic Aspects’, In Proc. of EPD Congress, 1990, pp 393-4 12. Linge, H.G., ‘New Technology for Refractory Gold Ores’, In Proc. of The International Conference on Extractive Metallurgy of Gold and Base Metals, Kalgoorlie, W. Australia, 1992, pp 339-343. Marsden, J. and House, l.,Z’he Chemistry of Gold Extraction, 1992, Chapter 6, Leaching, Ellis Hoi-wood Limited, London. Muir, D.M., Aziz, M. and Hoeker, W., ‘Cyanide Losses Under CIP Conditions and Effect of Carbon on Cyanide Oxidation’, In Proc. ofRando1 Gold, Perth, W. Australia, 1988, pp 253-257. Nagy, I., Mrkusic, P. and McCulloch, H.W., Chemical Treatment of Refractory Gold Ores, 1966, Report No. 38, National Institute for Metallurgy, South Africa, 36 pp. Nicol, M.J., 1988, Notes accompanying SIMCIL, software for ‘The Simulation of Gold Leach, Carbon-inleach and Carbon-in-pulp Absorption Circuits’. Osseo-Asare, K., Xue, T. and Ciminelli, V.S.T., ‘Solution Chemistry of Cyanide Leaching Systems’, In Proc. of Precious Metals: Mining, Extraction and Processing, LOS Angeles, U.S.A., 1984, pp 173-197. Papangelakis, V.G. and Demopoulis, G.P., Acid Pressure Oxidation of Arsenopyrite: Part 1, Reaction Chemistry, Canadian Metallurgical Quarterly, 1990a, 29(l), l- 12. Papangelakis, V.G. and Demopoulis, G.P., Acid Pressure Oxidation of Arsenopyrite: Part 2, Reaction Kinetics, Canadian Metallurgical Quarterly, 1990b, 29(l), 13-20. Robinson, J. and Williams, C., ‘Gold Recovery from a Golden Spec Stibnite Concentrate by Aqueous Pressure Techniques’, In Proc. of The International Conference on Extractive Metallurgy of Gold and Base Metals, Kalgoorlie, W. Australia, 1992, pp 333-337. Swash, P.M., A Mineralogical Investigation of Refractory Gold Ores and Their Beneficiation, with Special Reference to Arsenical Ores, Journal of the South African Institute of Mining and Metallurgy, 1988, 88(5), 173-l 80. Thomas, K.G., ‘Barrick Gold Autoclaving Processes’, In Proc. of World Gold 91, Cairns, Queensland, 1991, pp 89-97. Udupa, A.R., Kawatra, S.K. and Prasad, M.S., Developments in Gold Leaching: A Literature Survey, Mineral Processing and Extractive Metallurgy Review, 1990, 7, 115-135. Wang, X. and Forssberg, K.S.E., The Chemistry of Cyanide-Metal Complexes in Relation to Hydrometallurgical Processes of Precious Metals, Mineral Processing and Extractive Metallurgy Review, 1990,6,81-125. Weir, D.R. and Berezowsky, R.M.G.S., ‘Refractory Gold: The Role of Pressure Oxidation’, In Proc. of GOLD 100. International Conference on Gold, Johannesburg, S. Africa, 1986, pp 275-285. Weir, D.R., King, J.A. and Robinson, P.C., Preconcentration and Pressure Oxidation of Porgera Refractory Gold Ore, Minerals and Metallurgical Processing, 1996,3(4), 20 l-2 15.
18
M. N. Lehmann et nl
Correspondence on papers published in Minerals Engineering is invited, preferably by e-mail to [email protected], or by Fax to -t-44-(0)1326-3 18352
Mnera1.v Engincermng, Vol. 13, No. 1, pp. l-18, 2000 0 1999 Elsevier Science Ltd All rights reserved 0892-6875/00/$ - see front matter
AN EVALUATION OF PRETREATMENTS TO INCREASE GOLD RECOVERY FROM A REFRACTORY ORE CONTAINING ARSENOPYRITE AND PYRRHOTITE
M. N. LEHMANN”, S. O’LEARY?, A. J. Parker Cooperative
Research Center for Hydrometalhngy.
5 School of Applied Chemistry,
Curtin University
and J. G. DUNN@” E-mail: [email protected]
of Technology,
Perth, Western Australia
7 Mineral and Processing Laboratories of the Department of Energy, Waterford, (Received 5 July 1999; accepted 10 September 1999)
Western Australia
ABSTRACT The effect of three ore pretreatments on gold recoveries from an ore in which the gold existed in solid solution in an arsenopyrite matrix, as well as being encapsulated in a pyrrhotite phase, was examined. The three pretreatments evaluated included fine grinding, pyrolysis-roast (Pyrox), and an oxidative pressure acid leach (PAL). It was found that untreated ores did not yield high gold recoveries even when a fine grind (~45~) was employed, while oxidative treatments led to a signtjicant improvement of recoveries. The Pyrox pretreatment was found to decompose the sulfide matrix of the ore to a greater extent than a 3 hour PAL treatmemt. However, the PAL process produced a slightly better gold recovery of 79% compared to the Pyrox recovery of 75%, and both of these were greater than the 55-60% recoveries achieved with the untreated ore. Furthermore, the depletion of dissolved oxygen levels observed during cyanide leaching were signtj?cantly reduced by the two oxidative treatments, compared to the untreated ores. PAL was found to benefit over the Pyrox pretreatment by afording a solid product, which eflectively “locked” the arsenic component of the mineral from being recovered in solution. The PAL process required more lime for neutralization of acid produced from the decomposition reaction. 0 1999 Elsevier Science Ltd. All rights reserved. Keywords Gold ores; sulfide ores: roasting; grinding;
pressure leaching
INTRODUCTION Gold deposits can be broadly classed into two categories, primary and secondary. The primary deposits are sulfides, mainly pyrite and arsenopyrite, and more rarely other gold bearing sulfides such as pyrrhotite. Direct cyanide leaching of the secondary deposits give good gold recoveries in excess of 95%, but direct leaching of the primary deposits give poor recoveries of 5-70%, depending on the mineralogical composition of the deposit and the form of the gold in the host. The difficult-to-treat concentrates are described as “refractory”. There are a number of reasons why ores are refractory. It has been demonstrated that gold can occur in sulfide matrices in a range of sizes, from so-called “invisible” or solid solution gold (Cabri et al., 1989), to particles of 50-100 pm. Although the matrix offers particulate gold against contact with cyanide, dissolved oxygen and leaching
some protection to the reagents, fine grinding
* Presently employed with AMIRA gold project, Murdoch University, Murdoch, Australia ** Present address: Dept. of Chemistry, University of Toledo, Ohio, USA
2
M.N.Lehmannetal.
followed by cyanide leaching can sometimes produce acceptable recoveries (Espiell et al., 1986). However, gold in solid solution can only be effectively recovered if the matrix is altered by some chemical or biological pretreatment. Furthermore, sulfides (including arsenopyrite) will oxidatively decompose during cyanide leaching, and utilise reagents normally required for the dissolution of gold. For example, if arsenopyrite is leached directly in aerated cyanide solutions containing lime, arsenites may be formed according to the reaction in Eq. 1 (Nagy et al., 1966). This reaction consumes oxygen as well as lime (Eq. l), and contributes to reduced gold recoveries. 4FeAsS + 4Ca(OH)* + 1102 (s) +
4FeS04 (aq)+ 4CaHAs03 (Q + HZ0
Any pyrite or pyrrhotite in the ore would also be expected cyanide and oxygen (Eqs. 2-4).
&%+I
cs)
+ CN- +
nFeSe, + SCN- (aq)
4FeS,,, + 302 +6H20 + 4FeS,,, + 902 + 10&O +
4Fe(OHb
to consume
(1) substantial
quantities
of both
(2)
($)+ 4S”,,,
4Fe(OH),(,) + 4SO,2-,,,
(3) + 8H+,,,,
(4)
The solution to these problems is the implementation of an oxidative pretreatment which will completely break down the sulfide matrix and release the gold from solid solution (Demopoulos and Papangelakis, 1989; Dry and Coetzee, 1986; Dunn et al., 1989; Kontopoulos and Stefanakis, 1990; Linge, 1992; Udupa et al., 1990; Weir and Berezowsky, 1986; Weir et al., 1996). A prior concentration step such as flotation would also be required to minimise the amount of material treated, and thus make the added treatment cost effective (Swash, 1988). The aim of this investigation was to undertake a comparative evaluation of pretreatments on the recoveries obtained after cyanide leaching of a well characterised natural ore body, where the majority of the gold is located as invisible gold in an arsenopyrite matrix. The pretreatments that were evaluated included fine grinding the ore, a two stage pyrolysis/oxidative roast treatment (Pyrox) (Avraamides et al., 1991; Dry and Coetzee, 1986; Dunn, 1997; Dunn and Chamberlain, 1997; Dunn et al., 1989; Dunn et al., 1995b; Ferreira et al., 1989; Graham et al., 1992), and an acidic oxidative pressure acid leach (PAL) (Berezowsky and Weir, 1983; Dziurdzak et al., 1989; Papangelakis and Demopoulis, 1990a; Papangelakis and Demopoulis, 1990b; Robinson and Williams, 1992; Thomas, 1991; Weir and Berezowsky, 1986; Weir et al., 1996)
EXPERIMENTAL Samples Core samples of a gold ore were assessed for overall general mineralogy using powder X-ray diffraction (XRD). Scanning electron microscopy (SEM) and electron probe microanalysis (EPMA) were employed assess the native gold and/or telluride content, and solid solution gold in arsenopyrite and other sulfides.
to
The ore samples were made from a riffle split blend of five crushed core samples which were mixed to yield a sample with a theoretical gold head grade of 7.5 ppm. The blended sample was milled to the three size fractions of <45 urn, ~75 pm, and cl25 urn. The <75 urn fraction was divided into three quantities, two of which were oxidised prior to cyanide leaching. During storage the ground samples were sealed in air proof plastic bags and refrigerated.
Evaluation of pretreatments to increase gold recovery from refractory ore
Oxidative
3
pretreatments
The pyrolysis/roasting (Pyrox) pretreatment was carried out by heating about 5Og of ~75 um’ore under nitrogen in a furnace at 700°C for a period of 30 minutes, followed by heating in air at the same temperature for 3 hours. The acid pressure leach process was carried out as follows. 400g of the ore (~75 pm) was combined with 2.0 L of H2SO4 (10 g/L) to yield a mixture having a pulp density of 16.7%. The mix was placed in a 3.2 litre capacity autoclave and agitated at 400 rpm. The vessel was sealed and the head space purged with about 300 kPa nitrogen, leaving a residual 25 kPa of nitrogen gas overpressure. The sample temperature was then ramped to 180°C. On reaching the target temperature, 1500 kPa of oxygen gas was introduced to give a total vessel pressure of 2600 kPa. At this point the reaction timer was started. The experiment was conducted over three hours and intermediate samples were taken at 0.5, 1.0, 1.5, 2.0, 2.5, and 3.0 hours (testing for iron and sulfur in solution/solid phase and free acid in solution). The solutions were analysed for acidity (ie. free H2S04), iron by atomic absorption spectroscopy (AAS) and total sulfur (AAS). At the end of the experiment the system was quickly cooled to 30°C. The residual slurry was removed from the autoclave and washed. This involved 12 stages of a counter-current decantation, using approximately 240 L of tap water. The residue was then filtered and dried. Cyanide
leaching
A day before cyanide leaching, pretreated and untreated ore samples (50g) were transferred to 2 litre capacity plastic bottles fitted with screw cap lids. Each lid was perforated with a hole to allow air to enter. Deionised water (150 ml) was added to the solids, and the pH of the solution was adjusted to 10.5 by the addition of calcium hydroxide (lime). Additionally a sample of iron(II1) oxide obtained from Sigma chemicals (50g) was treated in a similar manner as a control for the experiment. The bottles were placed on a bottle roller and rotated at 60 rpm for 24 hours. Prior to actual leaching, and if needed, the pH of the prepared pulps was readjusted to 10.5. Leaching commenced after the addition of 50 mL of 6000 ppm sodium cyanide solution, which afforded a 20% w/v pulp with an initial cyanide concentration of 1500 ppm. The bottles were immediately placed on a bottle roller and rotated at 60 rpm for a period of 64 hours. During this time the pH of the solution was monitored intermittently with a Hanna HI1230 pH electrode and, where needed, maintained at a pH of 10.5 with the addition of lime. The mass of lime added to the solutions was recorded. The cyanide concentration of the pulp was monitored with an Ionode I/CN/Hg ion selective electrode connected to a TPS digital pH meter referenced to a Philips Rl 1 glass silver/saturated silver electrode, and readjusted to maintain the cyanide concentration at 1500 ppm. The oxygen concentration of the pulps was monitored utilising an Activon Model 40 1 oxygen meter. Sampling of the dissolved species present in the leach liquor was carried out by intermittently removing the plastic bottles from the roller and allowing the solids to settle, after which a sample of the supematant (3 ml) was removed with an auto pipette, filtered, and its weight recorded. The solution was then acidified with the addition of 2M HCl (10 ml) and oxidised with 20% v/v hydrogen peroxide (5 ml). The solution volume was reduced on a hotplate and made up to 25 ml in a volumetric flask. Leached gold, iron and sulfur were determined by ICP-OES (optical emission spectroscopy). The head grade of gold, arsenic, iron and sulfur in the solids before and after both pretreatments and cyanide leaching was determined by fire assay (Au) and ICP-OES (As, Fe & S) on an aqua-regia digest. Additionally, a sample of filtered liquor obtained after leaching the untreated ore sample (~75 urn), was analysed by ion chromatography in order to determine the ions present in solution.
4
M. N. Lehmann et al.
RESULTS
AND DISCUSSION
Mineralogy The sample was first checked for graphite, which in fine-grained form is responsible for preg-robbing. Optical microscopy of the untreated ore samples showed no sign of carbonaceous material, which is normally detectable by its anisotropic interaction to polarised light. It seems certain, therefore, that pregrobbing is not a cause of refractory behaviour in the ore. XRD analysis of the cl25 pm size fractions revealed that quartz was the most significant phase present in The next most abundant phases were clinochlore each of the five core samples. (Mg,Fe,A1)6(Si,A1)4010(OH)8);chlorite (Mg,Fe)S(Al,Si)S010(OH)8;and muscovite (KA12(Si3Al)0,0(OH.F)2. Quantities of hexagonal pyrrhotite (FeS), and arsenopyrite (FeAsS), were also found. These were less abundant than any of the silicate materials. Pyrite and galena, as well as gold, were not detected at the 1% detection level of the technique. An extensive particle to particle analysis of the ~125 urn ground ore samples, utilising SEM and EPMA, revealed that the ore contained significant quantities of both arsenopyrite and pyrrhotite. The maximum possible pyrrhotite contents were calculated to be 11% and 10% respectively. The trace results using the electron microprobe were derived from 6 minute counts at a current of 450 nA. Pyrite was also found to be present, as was the occasional grain of chalcopyrite (CuFe&) and galena (PbS), and an iron arsenide, possibly FeAs*. The gangue minerals were largely undifferentiated by brightness in the SEM, and were characterised by their qualitative energy-dispersive x-ray spectra. The major minerals found included quartz, a potassium aluminium silicate with varying amounts of iron, an iron aluminium silicate, and an iron manganese calcium aluminium silicate. Additionally grains of titanium were found mainly present as sphene, with very occasional grains of ilmenite. Siderite was the only carbonate definitely identified, but in general carbonates were elusive. Gold was only observed
in large particles
which had avoided
fine crushing,
and was consistent
with a
native gold content of about 5 & 10 ppm. Despite a long search, gold or tellurides were not observed. Arsenopyrite particles were analysed for trace gold, with the result that some arsenopyrite contained up to 250 ppm of gold, but most contained only a few ppm. The distribution was too irregular to allow quantification for the limited number of analyses conducted, but is a characteristic observation for solid solution gold in arsenopyrite. Measurements on several particles of pyrrhotite indicated that some of the gold present was most likely encapsulated
in this phase, and had a gold content of up to 5 & 10 ppm.
PAL pretreatment The conditions
for pressure
oxidation
were selected on the basis of prior studies on ores having
similar
sulfide matrices. Pressure oxidation was carried out for a period of 3 hours at 180 ‘C with an oxygen overpressure of 1500 kPa. Table 1 shows the mass transfer of sulfur and iron from the solid sulfide matrix to solution as the reaction occurs. Figure 1 shows that over a three-hour period, a 75% decomposition of the sulfide was achieved. Just over two thirds of this maximum breakdown occurred in the first half hour, and then the rate slowed considerably. If this second slower rate was maintained, it would take about a further 3 hours to decompose the sample completely. Figure 2 shows that the oxidation produces free acid. This is favourable in terms of increasing the rate of sulfide matrix breakdown. However, an effective residue wash is required to limit the amount of lime required for neutralisation to the pH value needed for cyanide leaching.
Evaluation of pretreatments to increase gold recovery from refractory ore
5
TABLE 1 Solution and solid breakdown occurring during oxidative pressure acid leach Time
@ours) 0 0.5 1 1.5 2 2.5 3
Solid
Free Acid G/L 10 19.6 23.0 25.7 26.9 27.3 27.8
Fe % 16.2 10.7 11.1 10.3 10.0 10.6 10.3
aaL.......,.....,
0
“.
Solution s % Fe % 0.33 0.00 1.03 2.22 1.80 1.27 1.67 1.35 1.65 1.40 1.62 1.45 1.62 1.46
S% 8.79 3.82 5.38 3.36 4.15 2.33 1.48 .‘.“.““‘.“‘>“
.
”
0.5
“,
““. 1
,.,
” 1.5
% Sulfide Decomposition 0 40 55 62 66 72 75
”
2
”
a..
2.5
”
”
”
3
”
3.5
Fig. 1 Total sulfide matrix breakdown as calculated from sulfur leached into solution.
Total Free Acid
Time (hours)
Fig.2 Acid concentrations measured during oxidative pressure acid leach.
M. N. Lehmann et al.
6
Pyrox pretreatment Figure 3 shows the thermogram pyrolysis.
of a small sample (24 mg) of the <75 urn ore sample obtained
'\ :
24 -
during
\ '1 \‘;\ \
235
'\
g
\
:: g
\
23.
\ i \! :\
_
225
22
0
~'~'~"~ 100
200
300
400
I"""""'! 500 600
Temperature
Fig.3 Mass loss of an untreated atmosphere).
ore sample
(175
700
800
900
1000
(Celcius)
pm) during
pyrolysis
(Heating
rate, 20’C/min,
N2
Decomposition, as monitored by weight loss, commenced at 500 “C, and a maximum weight loss is reached at temperatures greater than 720 “C. This supports the assumption that a maximum breakdown of the sulfide matrix would be expected under the conditions employed in the Pyrox pretreatment. The chemical analysis results shown in Table 2 indicate that significant quantities of both arsenic (ca. 96%) and sulfur (ca. 96%) were lost from the ore after the Pyrox treatment. TABLE 2 Elemental
Before Pyrox After Pyrox
compositions
Arsenic (kg/t) 49 2
before and after Pyrox treatment Sulfur (kg/t) 74 3
Gold (g/t) 6.61 7.82
These results suggest that the Pyrox pretreatment affords a greater decomposition of the sulfide matrix compared to the three hour oxidative pressure acid leach. The rise in the gold concentration after treatment is expected on the basis that the samples undergo weight-loss after the Pyrox treatment, while gold is conserved in the sample. Gold recoveries
by cyanide leaching
Table 3 shows the gold recoveries achieved those pretreated by Pyrox and PAL.
for the untreated
ores for a range of size fractions,
Figure 4 shows the leaching profile of gold during the period of cyanidation.
as well as
Evaluation
of pretreatments
TABLE 3 Gold recoveries
to increase gold recovery
of pretreated
and untreated
from refractory
ore
ores obtained from cyanidation
100
T
7
60
-----__-.._
----__.________~
-2
5 i?
$ 8
75pm
T
.-.--
PAL
1
___
45 pm Untreated .--------_-__.___ 75 blrn Untreated
~_~ ____ _.~___ _~__~
6o
125 pm Untreate
2 s
s
40
20
0 10
20
30
40
50
60
70
Time (hours)
Fig.4 The leaching profile of gold from pretreated and untreated
ores during cyanidation.
It can be concluded that gold recoveries are significantly improved with a pretreatment. The solution analysis of the gold indicated that after a period of 10 hours 79% and 75% of the gold was found to be leached from both the PAL and Pyrox pretreated ores respectively, as compared to 60%, 57%, and 55% of the untreated ores, listed in increasing particle size. The fact that recoveries from a range of particle sizes are not only low but similar (55-60%) suggests that fine grinding alone would not significantly improve recoveries, and is consistent with the gold being trapped in a solid solution form. In batch leaching the rate of leaching is described by Eq. 5 (Nicol, 1988);
8
--
M. N. Lehmann et al.
4Aul dt
=
kl{[Au]-[Au]e}2
Equation 5 can be integrated function of time; 1
[Au]- [Au1e
=kl.t
+
(5) to give Eq. 6, which describes the concentration
of undissolved
gold as a
1
(6)
[Au1o - [Au1e
where [Au] is the concentration of gold in the ore at any time t, [Au]~ and [Au]~ are the initial and equilibrium gold concentrations in the solid respectively, while k, is the leaching rate constant. From Eq. 6, the rate constant for the cyanidation of the ~45 urn untreated ore can be obtained by plotting the data of Table 4 as shown in Figure 5. TABLE 4 Data for the determination
of leaching rate constant kl for the ~45 pm untreated
*For the above calculations
ore
[AulBt 61hrs.= [Au]~ = 2.25 g/t
Time (hours) Fig.5 Plot of l/( [Au] - [Au]~ ) vs time for the determination 18.5 hours of cyanidation of the ~45 urn untreated ore.
of the leaching rate constant
k, for the first
Evaluation of pretreatments to increase gold recovery from refractory ore
Table 5 lists the leaching rate constants obtained from the initial stages equilibrium concentrations of the gold at the completion of the leach time. TABLE 5 Leaching rate constants stages of cyanidation
(k,) for treated
Ore Blend Untreated
ore ~45 pm Untreated ore ~75 pm Untreated ore ~125 urn Pyrox treated ore ~75 pm PAL treated ore <75 urn
and untreated
9
of leaching,
ores obtained
utilising
during
Leaching rate constant (tg-‘h-l) 0.10
0.98
0.28 0.21 0.53
0.96 0.97 0.90
0.32
0.96
the
the initial
R2
It is apparent that over extended time the PAL treatment yields slightly higher gold recoveries than the Pyrox treatment. These are reflected in the final recoveries of 85% and 75% for the PAL and Pyrox pretreatments respectively, as determined by fire assay. However, as displayed by Table 5, gold is leached out of the Pyrox treated ore more quickly compared to the PAL ore, as indicated by the greater leaching rate constant of the former. Figure 4 shows that the Pyrox treated ore exhibits a slight reduction of the gold in solution at extended leach times, which may be due to the reprecipitation of gold. The leaching rate constants for the untreated ores are significantly less than the treated ores. The <45 pm ore was found to have a leaching rate constant less than for the <75 and ~125 pm ores, and although unexpected might be accounted for by the small particles forming a thicker impervious iron oxide coating during milling. Since sulfur in the cyanide leach liquor originates from the sulfide component of the ore, a measure of its concentration gives an indication of the leach rate of the sulfide matrix component in untreated ores. In treated ores it gives evidence as to the level of completion of the oxidative pretreatment. The importance of the amount of sulfur in an ore that can undergo dissolution is dependent on the nature of the sulfur species that will be dissolved. In general, sulfur species other than sulfate can react with cyanide in a variety of ways and act as cyanicides. For example, cyanide readily reacts non-oxidatively with elemental sulfur to form thiocyanate (Eq. 7) as do a range of sulfoxy ions such as thiosulfate (Eq. 8) and tetrathionate ions (Eq. 9). Under oxidative conditions sulfide ions released from non-oxidative mineral sulfide dissolution can be oxidised to thiosulphate ions (Eq. lo), to then form thiocyanate ions via Eq. 8. S + CN- +
SCN-
(7)
S2032- + CN- +
SCN- + S032-
(8)
S4062- + CN- +
2SCN- + 2S032-
(9)
S*- + O2 +H20 +
S20s2- + 20H-
(10)
Although gold can form stable complexes with thiocyanate, at the pH employed for cyanidation, not complex with thiocyanate to any appreciable extent (Marsden and House, 1992).
gold does
Table 6 presents the quantity of sulfur dissolved for the pretreated and untreated ores, while Figure 6 shows the rate of sulfur dissolution into the cyanide liquor.
IO
M. N. Lehmann et al.
TABLE 6 Dissolved
sulfur levels of pretreated
00”“‘~‘~“‘~“‘~‘~‘~~~““‘~‘~~~“~ 0 10 20
30
and untreated
40
ores obtained from cyanidatiou
50
60
70
Time (hours)
Fig.6 The extent of sulfur dissolution
from pretreated and untreated
ores during cyanidation.
Each material (untreated and treated) was conditioned in deionised water for 24 hours, and the pH of the solution adjusted to 10.5 prior to the addition of cyanide solution. In the first 24 hours of cyanide leaching, all but the Pyrox treated ore show increases in the concentration of sulfur in the leach liquor. The Pyrox sulfur level in solution was relatively high to start with, indicating that water soluble sulfur compounds were present in the roasted product. This may have been entrained S02, or compounds such as calcium sulfate formed during the roasting step. The sulfur level was relatively constant over the leach period. Furthermore, Table 6 shows that almost all the sulfur in the matrix is leached. Both solution and residue analysis showed that the ~75 pm leached a greater amount of sulfur than the ~45 pm or the cl25 pm. This anomalous observation may be due to the fact that the ore is comprised of a variety of minerals, each with differing reactivities. Consequently, at certain particle sizes the dissolution of certain fractions of the ore may be favoured over others. For example pyrrhotite readily oxidises in air to form protective coatings (Pratt, et al.,) which may either be thicker or more uniform at certain particle sizes. However, the overall
Evaluation of pretreatments to increase gold recovery from refractory ore
11
ncrease in dissolved sulfur levels indicates that the sulfidic components of the ore are oxidising under the conditions of cyanide leaching. Figure 7 shows the dissolved oxygen (DO) levels during cyanide leaching. During the initial stages, both treated and untreated ores showed a significant decrease in the levels of DO. The most extreme decreases were experienced by the ~45 and ~75 urn untreated ores, and these results are consistent with the sulfide minerals consuming oxygen during oxidation (Dunn et al., 1995a). Table 7 shows the relative concentrations of soluble ions found after leaching the ~75 urn ore for the duration of the test under normal bottle roll conditions, and additionally when oxygen gas is sparged through the solution, instead of air, to promote further oxidation.
1, 75 pm Pyrox 2. 75 pm PAL 3. 4. 5. 6.
cl
10
20
125 pm Untreated Control 75 pm Untreated 45 pm Untreated
30
._
40
_^
50
^^
BU
IO
Time (hours)
Fig.7 Dissolved oxygen levels during cyanidation of pretreated and untreated ores.
TABLE 7 Concentration untreated
of major ions found in solution ore for a period of 64 hours
after cyanidation
of the minus
75 pm
12
M. N. Lehmann et al.
The major soluble product of mineral sulfide oxidation during cyanidation is the thiocyanate ion. As expected, the yield of this product rises with increasing oxidation of the mineral. In contrast, the dissolution of product ions which are cyanicides, such as copper and nickel, are inhibited when the oxidation potential of the solution is increased and is due to the oxidation of these ions to insoluble hydroxy carbonates such as malachite (Eq. 11). A reduction in the concentration of these ions would be expected if the CIP plant is sparged with oxygen gas rather than air (Muir et al., 1988). This is a positive outcome, as complexes of copper and nickel are known to catalyse the oxidation of cyanide.
2Cu(CN),-,,,,
+ ; O2 + 9H20 +
Cu(OH)&uC03
(s) + 3CO,‘-(,,, + 4NH‘,+ (aq)
(11)
The concentration of the ferro- and ferricyanide complexes in solution are also noticeably less significant than the concentration of thiocyanate ions. If iron hydroxide/oxides are initially formed from oxidative pathways, then the kinetics of subsequent dissolution in solutions containing cyanide ions (Eq. 12) must be slow. This would explain the absence of a large amount of the former, even though many reports have et al., 1984; postulated on the thermodynamic stability of these iron cyanide complexes (Osseo-Asare Wang and Forssberg, 1990). Fe(OH)3 cs)+ 6CN- caq)+
Fe(CN)b3- caq)+ 3OH. caq)
(12)
Figures 8 and 9 show the cumulative cyanide additions made to the untreated and treated ores, respectively, in order to maintain the cyanide concentration at 1500 ppm. In the initial 8 hours of leach there was little difference in the amount of cyanide consumed by any of the samples. After this time differentiation in cyanide consumption became evident. Notably, the Pyrox treated ore consumed significant quantities of cyanide with respect to the PAL and the untreated ores, and may be related to the elevated levels of dissolved iron that were observed in the leach, as shown in Figure 10. Interestingly, during the initial stages of cyanide leaching of the Pyrox treated sample, iron is readily solubilised and reprecipitated. Such behaviour is not observed with the PAL or untreated ores.
0.4""""'."'."""',""'."."" 0 10 20
30
40
50
60
7(
Time (hours)
Fig.8 Cyanide consumed during the cyanidation 1500 ppm (kg of NaCN / tonne ore).
of untreated
ores to maintain
the cyanide concentration
at
Evaluation of pretreatments to increase gold recovery from refractory ore
0
10
20
50 Tilt!
60
70
(houf$
Fig.9 Cyanide consumed during the cyanidation of treated ores to maintain the cyanide concentration at 1500 ppm (kg of NaCN / tonne ore). .
The dissolution of arsenic from the untreated ores is most likely due to the formation of soluble arsenates arising from reactions similar to Eq. 13, and also may arise from resolubilisation of arsenites formed from the reaction described by Eq. 1. FeAsS (sj+ 702 + 2H20 -_) 4HAs02 (aqj+ 4Fe+’caqj+ 4SOa2- (aqj
(13)
Ferrous ions produced from oxidation of iron sulfides may be further oxidised with oxygen to yield ferric ions (Eq. 14) (Hiskey and Schlitt, 1981). The ferric ions produced can combine with arsenate to form ferric arsenate, which precipitates as crystalline scorodite (Eq. 15). Furthermore, it is this reaction which is thought to contribute largely to the free acid produced during pressure oxidation of FeAsS as shown in Figure 2 (Bailey and Peters, 1976; Papangelakis and Demopoulis, 199Oa). + o2 4Fe2+(,,)
+
Fe3+(,,) + H&Q
4H’,,,, + 4Fe3+(,) + 2H20 caq)
+ 2H20 + FeAs04.2HzO cs)+ 3H’,,,
(14) (15)
Consequently, the observation that the iron concentration for the PAL ore initially starts low and remains comparatively low throughout the leach, is most likely a result of it being bound in scorodite. This also explains why the PAL treated ore had the highest head grade of iron (Table 8), and that throughout the leach little iron was dissolved (Figure 10). Table 9 shows the low arsenic recoveries obtained fi-om cyanide leaching of the PAL treated ores as compared to those untreated, even though the initial head grades are the same. This further confirms that after PAL, the arsenic is “locked” into the scorodite matrix, which is desirable as it is nonreactive. Although the stability of this complex has been questioned, provided that the molar ratio of ironzsenic is four, the ferric arsenate is stable enough for landfill disposal (Haines, 1986). The Pyrox treated ore is also observed to yield a low arsenic recovery, which is not unexpected given the low residual arsenic levels remaining after the pyrolysis reaction.
14
M. N. Lehmann et al.
TABLE 8 Dissolved
iron levels of pretreated
and untreated
ores obtained from cyanidation
50
40
45 pm Untreated f-
30
h 75 pm PAL
/1
20
IO 125 pm Untreatf
f
1 30
40
50
60
70
Time (hours)
Fig. 10 The extent of iron dissolution
TABLE 9 Dissolved
from pretreated and untreated
arsenic levels of pretreated
and untreated
ores during cyanidation.
ores obtained from cyanidation
Evaluation
of pretreatments
to increase gold recovery from refractory
ore
15
The PAL treated ore showed a continual increase in the amount of sulfur released into the leach liquor. However, the percentage of sulfur recovered in the tail was much higher than for the Pyrox treated ore but lower than for the untreated ores. The oxygen consumption during cyanidation was also comparatively low for the PAL treated ore, but the consumption of lime over the leach period was high with respect to the untreated ores but lower than the Pyrox treated ore. It would be expected that any sulfide present in the ore would be oxidised in the PAL process to a soluble sulfate species. It is curious that much of the residual sulfur in the head material was retained after cyanidation, and yet a relatively high cyanide consumption was recorded, compared to the untreated ores. One possibility is that the consumption of cyanide is due to either an insoluble polysultide species, or elemental sulfur retained in the PAL product. This is in agreement with the observation of Dunn et al. (1996) who postulated that a major product of oxidation must be elemental sulfur, in order to stoichiometrically balance the oxidation of arsenopyrite in aqueous solutions. It would be expected that if it was merely a sulfur layer that the rate of dissolution would be quite rapid. However, this is not the case, and hence the sulfur must also be integrated into either the iron oxide or scorodite product layers. Figure 11 shows that the PAL treated ore also had the highest consumption of lime during the leach, which is consistent with an iron oxide product layer undergoing reactions similar to those described by Eqs. 16 and 17. Furthermore, an initially large lime addition might have been required to neutralise any residual acid in the PAL product, in a reaction described by Eq. 18. FeO,,, + Hz0 -_) FeOOH,,, + H+ + 1em
(16)
FeO(,, + OH- +
(17)
30H-,,,,
FeOOHe, + 1e-
+ H3As04 (aq)+
As0d3- (aq)+ 3H20
(18)
Time (hours)
Fig. 11 Lime consumed during the cyanidation Ca(OH)2 / tonne ore).
of ores to maintain
the slurry pH at a value of 10.5
(kg
The solubilisation and reprecipitation of iron for the Pyrox treated ore in Figure 10 is also observed in the control material hematite and can be thought of as being due to a combination of hydration and dehydration reactions which might follow a pathway as described by Eq. 19. Fez03 cs)+ 3H20 --+ 2Fe(OH)3
(sacurated sown.) +
2FeOOH
cs) + 2&O
(19)
16
M. N. Lehmann et al.
CONCLUSION This study has shown that in order to obtain appreciable gold recoveries from a gold ore, where the gold is located as “invisible” gold in a sulfide matrix composed mainly of arsenopyrite and pyrrhotite, an oxidative pretreatment is required prior to cyanide leaching. Fine grinding (~45 pm) of this ore type did not lead to a significant improvement in the gold recovery. A cyanide leach of untreated ores yielded soluble arsenic ions. As a consequence, if this treatment was to be adopted, the liquor from the leach would require further treatment to remove these ions. The oxygen consumption of the PAL and Pyrox treated ores was significantly lower during leaching than for the untreated ores, while the leaching rates of the ores was significantly increased with pretreatment. Of the oxidative pretreatments examined, the PAL treatment yielded a gold recovery of 79% after 10 hours which is slightly better than the 75% achieved after the Pyrox process, and appreciably better than the 55-60% recovered without treatment. The Pyrox treated ore was observed to have the faster leach rate and decomposed the sulfide matrix of the ore to a greater extent than the PAL treatment. PAL, however, afforded a solid product which effectively “locked” the arsenic component of the mineral from being recovered in solution. As a consequence, lime consumption during cyanide leaching was significantly higher for the PAL treated ores than for Pyrox or untreated ores. The choice of which oxidative pretreatment to employ commercially will have to be balanced by the fact that soluble arsenic will be evolved during leaching of the Pyrox treated ore. While this is not a problem with the PAL treated ores, substantially more lime will be needed in order to leach the latter.
ACKNOWLEDGMENTS The authors wish to thank Dr J. Graham for conducting the mineralogical assessment of the ore used in this study. This research has been supported by the Australian Government’s Cooperative Research Centres Program, through the A J Parker CRC for Hydrometallugry.
REFERENCES Avraamides, J., Nguyen, G., Dunn, J. and Graham, J., ‘Roasting and Cyanide Leaching of a Gold Bearing Pyritic Concentrate from Kalgoorlie, Western Australia’, In Proc. of World Gold 91, Cairns, Queensland, 1991, pp 45-53. Bailey, L.K. and Peters, E., Decomposition of Pyrite in Acids by Pressure’Leaching and Anodization: The case for an electrochemical mechanism., Canadian Metallurgical Quarterly, 1976, 15(4), 333-344. Berezowsky, R.M.G.S. and Weir, D.R., ‘Pressure Oxidation Pretreatment of Refractory Gold’, In Proc. of Practical Hydrometallurgy 83: Annual Symposium on Uranium and Precious Metals, Fort Saskatchewan, Alberta Canada, 1983, pp 101-104. Cabri, L.J., Chryssoulis, S.L., Villiers, d.J.P.R., Laflamme, J.H.G. and Buseck, P.R., The Nature of Invisible Gold, The Canadian Mineralogist, 1989, 27(3), 353-362. Demopoulos, G.P. and Papangelakis, V.G., Recent Advances in Refractory Gold Processing, CIM Bulletin,, 1989,82(931), 85-91. Dry, M.J. and Coetzee, C.F.B., ‘The Recovery of Gold from Refractory Ores’, In Proc. oflnternational Conference on Gold, Johannesburg, S. Africa, 1986, pp 259-274. Dunn, J.G., The Oxidation of Sulfide Minerals, Thermochimica Acta, 1997,300(1-2) 127-139. DUM, J.G. and Chamberlain, A.C., The Recovery of Gold From Refractory Arsenopyrite Concentrates By Minerals Engineering, 1997, 10(9), 919-928. Pyrolysis-Oxidation, Dunn, J.G., Femandez, P.G., Hughes, H.C. and Linge, H.G., ‘Aqueous Oxidation of Arsenopyrite in Acid’, In Proc. of A&MM Annual Conference, Perth-Kalgoorlie, W. Australia, 1996, pp 2 17-220. Dunn, J.G., Graham, J. and Linge, H.G., The Treatment of Refractory Gold, Chemistry in Australia, 1989, 401-404. DUM, J.G., Hollbach, D.J. and Ibrado, A.S., ‘The Dissolution of Pyrrhotite in Cyanide Solution’, In Proc. of Rand01 Gold Forum, Perth, W. Australia, 1995a, pp 145-148.
Evaluation of pretreatments to increase gold recovery from refractory ore
17
Dunn, J.G., lbrado, A.S. and Graham, J., Pyrolysis of Arsenopyrite for Gold Recovery by Cyanidation, Minerals Engineering, 1995b, 8(4/5), 459-471. Dziurdzak, G., Kyle, J.H. and Dunne, R.C., ‘The Pressure Aqueous Pre-Oxidation of Refractory Gold Ore from the Golden Mile, Kalgoorlie, Western Australia’, In Proc. of World Gold 89: Forum on Technoloa and Practices, Reno, Nevada, 1989, pp 3 15-32 1. Espiell, F., Rota, A., Cruells, M. and Nunez, C., Gold and Silver Recovery by Cyanidation of Arsenopyrite Ore, Hydrometallurgv,, 1986, 16, 141-151. Ferreira, P.M., Knott-Craig, K. and Boydell, D.W., Roasting of Refractory Au Concentrates at Consort. The new two-stage fluidized-bed roasting plant has negligible adverse impact on the environment, Engineering and Mining Journal, 1989, ,26-30. Graham, J., Fletcher, A., Walker, G. and Aylmore, M., ‘A Pyrolysis Approach to Refractory Gold Processing’, In Proc. of The International Conference on Extractive Metallqy of Gold and Base Metals, Kalgoorlie, W. Australia, 1992, pp 441-443. Haines, A.K., ‘Factors Influencing the Choice of Technology for the Recovery of Gold from Refractory Arsenical Ores’, In Proc. of GOLD 100. International Conference on Gold, Johannesburg, S. Africa, 1986, pp 227-233. Hiskey, J.B. and Schlitt, W.J., ‘Aqueous Oxidation of Pyrite’, In Proc. of Interfacing Technologies in Solution Mining: The 2nd SME-SPE International Solution Mining Symposium, Denver, Colorado, 198 1, pp 55-74. Kontopoulos, A. and Stefanakis, M., ‘Process Options for Refractory Sulfide Gold Ores: Technical, Environmental and Economic Aspects’, In Proc. of EPD Congress, 1990, pp 393-4 12. Linge, H.G., ‘New Technology for Refractory Gold Ores’, In Proc. of The International Conference on Extractive Metallurgy of Gold and Base Metals, Kalgoorlie, W. Australia, 1992, pp 339-343. Marsden, J. and House, l.,Z’he Chemistry of Gold Extraction, 1992, Chapter 6, Leaching, Ellis Hoi-wood Limited, London. Muir, D.M., Aziz, M. and Hoeker, W., ‘Cyanide Losses Under CIP Conditions and Effect of Carbon on Cyanide Oxidation’, In Proc. ofRando1 Gold, Perth, W. Australia, 1988, pp 253-257. Nagy, I., Mrkusic, P. and McCulloch, H.W., Chemical Treatment of Refractory Gold Ores, 1966, Report No. 38, National Institute for Metallurgy, South Africa, 36 pp. Nicol, M.J., 1988, Notes accompanying SIMCIL, software for ‘The Simulation of Gold Leach, Carbon-inleach and Carbon-in-pulp Absorption Circuits’. Osseo-Asare, K., Xue, T. and Ciminelli, V.S.T., ‘Solution Chemistry of Cyanide Leaching Systems’, In Proc. of Precious Metals: Mining, Extraction and Processing, LOS Angeles, U.S.A., 1984, pp 173-197. Papangelakis, V.G. and Demopoulis, G.P., Acid Pressure Oxidation of Arsenopyrite: Part 1, Reaction Chemistry, Canadian Metallurgical Quarterly, 1990a, 29(l), l- 12. Papangelakis, V.G. and Demopoulis, G.P., Acid Pressure Oxidation of Arsenopyrite: Part 2, Reaction Kinetics, Canadian Metallurgical Quarterly, 1990b, 29(l), 13-20. Robinson, J. and Williams, C., ‘Gold Recovery from a Golden Spec Stibnite Concentrate by Aqueous Pressure Techniques’, In Proc. of The International Conference on Extractive Metallurgy of Gold and Base Metals, Kalgoorlie, W. Australia, 1992, pp 333-337. Swash, P.M., A Mineralogical Investigation of Refractory Gold Ores and Their Beneficiation, with Special Reference to Arsenical Ores, Journal of the South African Institute of Mining and Metallurgy, 1988, 88(5), 173-l 80. Thomas, K.G., ‘Barrick Gold Autoclaving Processes’, In Proc. of World Gold 91, Cairns, Queensland, 1991, pp 89-97. Udupa, A.R., Kawatra, S.K. and Prasad, M.S., Developments in Gold Leaching: A Literature Survey, Mineral Processing and Extractive Metallurgy Review, 1990, 7, 115-135. Wang, X. and Forssberg, K.S.E., The Chemistry of Cyanide-Metal Complexes in Relation to Hydrometallurgical Processes of Precious Metals, Mineral Processing and Extractive Metallurgy Review, 1990,6,81-125. Weir, D.R. and Berezowsky, R.M.G.S., ‘Refractory Gold: The Role of Pressure Oxidation’, In Proc. of GOLD 100. International Conference on Gold, Johannesburg, S. Africa, 1986, pp 275-285. Weir, D.R., King, J.A. and Robinson, P.C., Preconcentration and Pressure Oxidation of Porgera Refractory Gold Ore, Minerals and Metallurgical Processing, 1996,3(4), 20 l-2 15.
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M. N. Lehmann et nl
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