Valuable metals, such as cobalt, can be recovered from slags by treating these waste materials with a carbonaceous reducing agent in a DC-arc furnace. This technology is applicable to the recovery of nickel, cobalt, copper, and platinum group metals from furnace or converter slags emanating from copper or nickel smelters treating sulphide concentrates. Pilot-plant testwork at Mintek has demonstrated recoveries of 98% for nickel and over 80% for cobalt, at power levels of up to 600 kW. Results of computer simulations and experimental tests are presented.
INTRODUCTION
Slags produced by non-ferrous smelters usually contain significant
quantities of valuable metals, such as cobalt, nickel, and copper,
present both in an entrained metallic or sulphide form, and in
a dissolved oxidized form. These slags can emanate from either
furnaces (e.g. six-in-line, reverberatory, or flash) or from converters;
they may be internal recycle streams or final waste products.
Large quantities of valuable metals are present in the huge dumps
that have built up over many years of operation of nickel and
copper smelters. Slags are sometimes treated by slow cooling,
milling, and flotation. This approach is satisfactory when the
metals in the slag are in either the sulphide or metallic form,
but is not suitable for the recovery of oxidized metals. Cobalt
in particular, and nickel to a lesser extent, are found in an
oxidized form, and for these slags, treatment in an electric furnace
operated under reducing conditions is necessary.
Conventional slag cleaning furnaces (typically AC submerged-arc)
rely largely on a gravity settling mechanism, whereby the entrained
sulphide and metallic droplets are simultaneously collected.
Sometimes, a quantity of matte is added to the slag, to enhance
the coalescence of entrained matte droplets. However, conditions
are not usually sufficiently reducing to recover much of the cobalt.
Cobalt recoveries may be as low as 20 per cent. The most effective
means for the recovery of metals involves the addition of a reductant
(such as carbon) to capture some of the metals present in an oxidized
form. Processes have previously been described in which carbothermic
reduction has been carried out in electric furnaces (1-3).
The aim of a slag cleaning process is to maximize the recovery
of valuable metals (such as Co, Ni, and Cu) in an alloy with the
lowest possible iron content. The amount of metallic iron produced
should be kept to a minimum, as the more iron present in the resulting
matte or alloy, the greater the cost of the subsequent hydrometallurgical
separation of the valuable metals, and the resulting disposal
of the iron residues. Because of the similarities in the reduction
behaviour of cobalt and iron, some loss of cobalt is inevitable
while separating the iron from the nickel and copper.
Mintek has been working on the recovery of cobalt, and the associated
valuable metals, from slags, since 1988, using DC-arc furnace
technology to effect selective carbothermic reduction of the oxides
of cobalt, nickel, copper, (and zinc, where present), while retaining
the maximum possible quantity of iron as oxide in the slag.
Flowsheets for slag cleaning
Many non-ferrous smelters employ a process whereby the concentrates are fed to a furnace which produces a matte (for further treatment) and a slag (which is dumped). The furnace matte is treated in a converter (often of the Peirce-Smith type) to remove most of the remaining iron and sulphur. This resulting 'white matte' or alloy is then treated hydrometallurgically. The converter slag is usually recycled to the furnace. Because of the highly oxidizing conditions in the converter, much of the cobalt is oxidized. The turbulent conditions cause entrainment of valuable metals as well. Of all the streams in a flowsheet of this sort, the converter slag is richest in cobalt. As shown in Figure 1, it is possible to divert this liquid stream of converter slag for slag cleaning, allowing most of the valuable metals to be reclaimed. The impoverished slag can still be recycled to the furnace (with fairly minimal disruption to the existing process, and with the benefit of reduced quantities of magnetite which otherwise builds up in the furnace), or can be dumped (breaking the recycle entirely, necessitating some changes to the operation of the furnace).
A second possibility is to leave the converter slag recycle stream alone, and focus on the treatment of the furnace slag which is the point at which the waste materials finally leave the process. It is also possible to treat material from existing slag dumps at the same time, as the dumped material is usually similar to the furnace slag currently being produced. This option is shown in Figure 2. It is, of course, also possible to use a hybrid of these approaches.
A wide range of slags are amenable to slag cleaning. These slags
differ according to the ores which have been processed, as well
as according to the type of process, and whether the slags have
arisen from furnaces or from converters. Most of the slags of
interest are rich in iron oxide and silica, and many have a bulk
composition approximating that of fayalite. Fortunately, similar
principles apply to the treatment of all of these slags, although
the actual results will differ according to the composition of
the slag. For the sake of illustration, two representative slags
are shown in Table I, differing primarily in the level of cobalt
contained in the slag.
Note that, for the purposes of these calculations, all of the
Co, Cu, and Ni in the feed slag was assumed to be in the oxidized
form, i.e. present as CoO (1.272% or 0.254%), Cu2O
(0.563%), and NiO (2.546%).
When carbon is added to the slag, the various metallic elements reduce to different extents, at a given level of carbon addition. This behaviour allows a reasonable degree of separation to take place during smelting. The intention in this part of the process is to separate the valuable non-ferrous metals from the iron and the gangue constituents present in the slag. Figure 3 illustrates the differences in reducing behaviour between nickel, cobalt, and iron. The desirable area of operation is clearly somewhere in the region where the recovery of cobalt is high, and the recovery of iron to the alloy is still reasonably low. Note that, in actual practice, there is less than 100 per cent carbon utilization, and the carbon addition would need to be somewhat higher than that shown here, because of burn-off of some of the reductant.
The calculations for Figures 3, 4, and 5 were carried out using
Mintek's Pyrosim computer software (4) for the calculation of
steady-state mass and energy balances for pyrometallurgical processes.
These simulations were based on the assumption of chemical equilibrium
between gas, slag, and alloy. The equilibrium composition was
calculated using free-energy minimization, together with the Ideal
Mixing of Complex Components solution modelling technique. An
entrainment of 2 per cent of the matte in the resulting slag was
assumed, as this was found to agree well with experimental results
from DC-arc furnace testwork. Of course, there is a spread of
experimental data around the simulated values, as the model does
not take into account subtle differences in the mode of operation
of the furnace. It should be noted that a good mineralogical
description of the slag feed is required in order to provide an
accurate estimate of the energy requirements of the process.
(Pyrosim can also be used in non-predictive mode, using the empirical
Pyrobal model, which allows the user to specify, for example,
the actual experimentally obtained percentages of cobalt and nickel
in the slag, or the distribution of iron between metal and slag.
This is very useful for generating a completely consistent mass
balance from incomplete or conflicting experimental data, where
at least some information is known to be accurate.)
The equilibrium behaviour shown in Figure 3 is not very sensitive
to temperature, and curves plotted for temperatures 100°C
colder or hotter are virtually indistinguishable from those presented.
Clearly, the temperature has effects other than on the chemical
thermodynamics of the system. It has been reported elsewhere
(6) that a 100°C increase in temperature may decrease the
cobalt and nickel solubility in slag by as much as three times,
at a constant partial pressure of oxygen, and at a temperature
around 1300°C. However, the solubility is even more strongly
affected by the reducing nature of conditions in the furnace (i.e.
the partial pressure of oxygen in the system).
The most striking feature of this separation process is the variation of cobalt recovery according to the iron content in the alloy. This behaviour is shown in Figure 4. It can be seen that the recovery of cobalt (in percentage terms) is highest (albeit not by very much) in the case of the slag with the highest initial cobalt content. This is in line with experimental findings (2,7) that recoveries are dependent on initial slag composition, with higher grades leading to better recoveries. If we accept the evidence presented elsewhere (7) that metallic Fe is the effective (intermediary) reductant in the process, it may be more correct to say that the recovery of valuable metals is related to a combined function of the iron and non-ferrous metal contents of the initial slag.
However, as shown in Figure 5, the residual cobalt content in the cleaned slag, as a function of the iron content of the alloy, is very sensitive to the initial cobalt content.
Alloy melting point
The major constituents of the alloys produced melt at rather high temperatures, as shown in Table II. The melting point of the alloy increases sharply with iron content, and melting temperatures in the range from 1300 to 1420°C have been reported (2). This is one of the important determinants of a suitable operating temperature for the furnace. An operating temperature of 1500 to 1550°C has been selected for the process, in accordance with that used for a similar process (2).
The experimental work on the slag-cleaning process began on the
laboratory scale, and was extended to pilot scale on Mintek's
DC transferred-arc furnaces (8). In addition to numerous 100 kVA
(60 kW) supporting batch tests, five campaigns (of 50 to
100 hours each) have been carried out at the 200 kVA (150 kW)
scale, and a campaign (treating about 20 t of slag) has also been
undertaken on the 3.2 MVA (600 kW) furnace. Future
work is planned for the 5.6 MVA (1-3 MW) furnace facility.
The DC-arc furnace has a single electrode positioned above the molten bath; the molten metal in the furnace forms part (the anode) of the electrical circuit. The furnace comprises a refractory-lined cylindrical steel shell, and a water-cooled roof lined with an alumina refractory. The outer side walls of the furnace are spray-cooled with water, to protect the refractories, and to promote the formation of a freeze lining within the vessel. The roof contains the central entry port for the graphite electrode and up to three equi-spaced side feed ports. The return electrode, or anode, consists of multiple steel rods built into the hearth refractories and connected at their lower end to a steel plate which, via radially extending arms, is linked to the furnace shell, and further to the anode cable. A schematic diagram of this arrangement is shown in Figure 6.
Molten slag may be fed directly to the 5.6 MVA furnace from
a pre-melting furnace. The solid-feed system for the 3.2 MVA
furnace comprises a batching plant and a final controlled feeding
system. The batching plant consists of feed hoppers mounted on
load cells, vibratory feeders positioned under the hoppers, an
enclosed belt conveyor, a bucket elevator, and a pneumatically
actuated flap valve to direct the feed to one of two final feed
hoppers. The final feeding system is made up of separate centre
and side feeding arrangements. The centre feeder uses a screw
feeder discharging into a telescopic pipe attached to the hollow
graphite electrode. The side feeders are vibratory, and discharge
into feed chutes.
The gas-cleaning system consists of a water-cooled off-gas pipe,
a refractory-lined combustion chamber, water-cooled ducting, a
forced-draft gas cooler, a reverse-pulse bag filter, a fan, and
a stack. The condensed fume and dust, which accumulates in the
lower conical section of the bag plant, is discharged via a rotary
valve into a collecting drum. This dust would, of course, be
recycled back to the furnace in an industrial situation.
Copper converter slag - 100 kVA
During 1988 and 1989, tests were carried out using a converter slag having a composition of Co: 0.46%, Cu: 3.25%, Fetotal: 52%, and Ni: 0.42%. A 100 kVA furnace (operating at 20 to 25 kW) was used. In some tests, solid slag and coal were fed together, while in others the slag was first melted then the coal added afterwards. Coal (at 53% fixed carbon) additions varied between 4, 6, 8, 10, and 12 per cent of the mass of the slag. The mode of addition of the coal, and the reduction periods of the tests were also varied (around 30 minutes). The results of the tests are shown in Table III and Figure 7.
The use of fine coal (less than 1.5 mm) did not seem to have
an effect on the degree of reduction. A representative composition
of the alloy produced was Co: 6%, Cu: 32%, Fe: 53%, and Ni: 6%.
Nickel-copper converter slag 'A'- 100 kVA and 200 kVA
Using converter slag from a nickel-copper plant, tests were carried out, in 1990, on a 100 kVA furnace (operated at 30 kW) and on a 200 kVA furnace (operated at 85 kW), at low additions of reductant (in order to minimize the reduction of iron). The slag had a composition of Co: 0.45%, Cu: 3%, Fetotal: 47%, Ni: 3.5%, and S: 3%. These tests examined four different methods of operation, in an attempt to optimize the selective reduction of the slag. These methods included smelting of composite pellets of milled slag and graphite, adding selected quantities of crushed coal to already molten slag, co-feeding crushed cold slag and coal, and pneumatic injection of pulverized coal into the molten slag. On the 100 kVA furnace, the alloy produced typically comprised Co: 1.7%, Cu: 14%, Fe: 48%, Ni: 16%, and S: 14%. On the 200 kVA furnace, an alloy of Co: 2%, Cu: 15%, Fe: 44%, Ni: 22%, and S: 10% was produced. At this level of reduction, 91% of the iron was retained in the slag phase, while only 50% of the cobalt, 63% of the copper, and 83% of the nickel were recovered to the alloy. It was found that injection of pulverized coal greatly improved the reduction, and hence the recovery, of nickel and cobalt oxides from the slag. The results of these tests are shown in Figure 8, where the scatter of results needs to be seen in the context of the various methods employed. Under good conditions, at 7% carbon addition, on the 100 kVA furnace, calculated recoveries of Co: 81%, Cu: 78%, and Ni: 97% were obtained, while retaining 80% of the iron in the slag.
Nickel-copper converter slag 'B' - 200 kVA
This slag was generated in a plant utilizing a conventional six-in-line furnace and Peirce-Smith converter configuration. The composition of the bulk slag was Co: 1.25%, Cu: 1.0%, Fetotal: 49%, Ni: 3.6%, and SiO2: 30%. During testwork carried out in 1993, the furnace was operated at power levels of 100 to 170 kW and tapping temperatures of 1400 to 1500°C. The alloys produced comprised Co: 4.5 to 5.5%, Cu: 5.5 to 8.5%, Ni: 25 to 35%, Fe: 35 to 50%, and S: 8 to 10%. From the starting level of 1.25% cobalt in the feed slag, it was possible to produce a discard slag with typical values of 0.22 to 0.29% cobalt. Results from the campaign are summarized in Figure 9.
Co-feeding of the coal and sequential feeding (addition of coal
to a molten slag bath) gave similar recoveries to the alloy of
74 and 80% respectively. While cobalt recoveries were similar,
sequential feeding (the preferred route for an industrial plant
where molten slag is available) produces a better grade of alloy,
with an iron content in the region of 10 to 15% lower.
The most abundant phase present in the solid converter slag is
Fe2SiO4 (olivine), followed by Fe3O4
(spinel). The cobalt is associated primarily with the olivine,
whereas the nickel is distributed between the olivine and the
spinel. Copper was present only in entrained sulphides. The
analysis of highly reduced slags has shown that it is possible
to remove virtually all of the cobalt and nickel from the olivine
in the slag.
Nickel-copper converter slag 'B' - 3.2 MVA
Large-scale testwork on converter slag was conducted, during 1994,
on a 3.2 MVA DC-arc furnace operating at a power level of
600 kW. The sequential feeding of reductant was used as
the preferred mode of feeding. Operating temperatures were in
the region of 1300 to 1600°C, and neither the
temperature of the bath before the reduction period nor the tapping
temperature seemed to have a pronounced effect on the cobalt recovery.
The average electrode consumption during the campaign was 2.6 kg/MWh,
while the dust loss was low, at 1% of the mass of the feed. The
alloy produced comprised Co: 7.8%, Cu: 3.8%, Ni: 26.4%,
Fe: 56.9%, and S: 2.1%. The cobalt levels achieved
in the discard slag were between 0.15 and 0.33%.
A coal addition of 9% was required to achieve cobalt recoveries
of at least 80%. Increasing the coal addition did not significantly
increase the recovery of cobalt. Increasing the batch mass of
slag from 500 kg to 1000 kg and increasing the reduction period
by 75% resulted in increases in cobalt recovery from 71 to 86%,
and from 70 to 82%, for coal additions of 9 and 11% respectively.
The main factor affecting cobalt recovery appears to be the time
allowed for the reduction to take place. At this scale of operation,
a duration of two hours was required to achieve cobalt recoveries
greater than 80%.
Nickel-copper furnace and converter slag -200 kVA
During 1995, a campaign was undertaken on the 200 kVA furnace
with the intention of combining furnace slag with the converter
slag previously treated. When treating furnace slag containing
0.22% Co on its own, the maximum cobalt recovery that could be
obtained was 66%, with a cobalt value of 0.08% in the discard
slag. As in previous testwork, the cobalt in the discard slag
when treating converter slag on its own was still in the region
of 0.22%. However, by combining increasing amounts of furnace
slag with converter slag, values approaching 0.08% Co in the discard
slag could still be achieved. This resulted in cobalt recoveries
approaching 90% being attained, while the average recovery was
in the region of 85%.
The fume produced was of the order of 1% of the mass of the slag
fed. The electrode consumption was 2 kg/MWh.
Copper reverberatory furnace slag - 200 kVA
During 1995, a campaign was conducted on Mintek's 200 kVA DC-arc furnace to recover cobalt from copper reverberatory furnace slag (initially containing Co:0.77% and Cu:1.3%). Alloys containing Co: 6-7%, Cu: 9-11%, Fe: 77-78%, and S: 2-3% were produced, leaving slags containing Co: 0.08-0.16% and Cu: 0.18-0.29%. Coal additions of 4 per cent of the mass of the slag fed, and residence times of one to two hours, gave cobalt recoveries ranging from 77 to 91%. Tapping temperatures of 1490-1560°C were attained. As shown in Figure 9, the cobalt level in the slag (and therefore the cobalt recovery) varied according to the residence time in the furnace. A retention time of 2 hours was required to achieve cobalt recoveries in the region of 90%.
During this campaign, the dust loss was 0.7 per cent of the mass
of feed slag. The electrode consumption was 1.0 kg/MWh.
The slag resistivity was calculated to be 1.1 cm. Measurements
were also made of the arc characteristics. The melting point
of the alloy produced in the furnace was 1340°C, as determined
by differential thermal analysis.
Some of the alloy was upgraded by blowing with oxygen in a top-blown
rotary converter (TBRC), preferentially oxidizing the iron, and
thereby concentrating the cobalt and copper in the alloy. The
iron content in the alloy was lowered from 76 to 25%, with the
effect that the alloy was concentrated up to 30% cobalt and 40%
copper. The sulphur level in the resulting alloy was 0.8%. This
means that 75% of the cobalt remained in the alloy, while 90%
of the iron was removed to the slag. The slag from the TBRC would,
of course, be recycled back to the DC-arc furnace, in order to
prevent the blown cobalt from being lost.
In the case of copper reverberatory furnace slag, mineralogical
studies showed that the cobalt is present as CoO. Copper in the
slag is mainly attributed to the presence of copper-rich sulphides.
The cobalt oxide, and, to a lesser extent, the copper oxide associated
with the silicate / oxide phases, is reduced by Fe from the alloy
to form metallic Co (and Cu), resulting in the formation of FeO
in the slag. Given that this reaction occurs between the metal
bath and the overlying slag, the exchange of Co and Cu with Fe
will take place only at the slag/metal interface. Improved recoveries
of valuable metals can be achieved by allowing greater quantities
of slag to come into contact with the alloy (by mild stirring,
for example), and increasing the length of the contact time between
slag and metal. The CoO in the slag is associated primarily with
Fe2SiO4, and analysis by scanning electron
microscopy showed some Fe2SiO4 particles
with no detectable Co or Cu, thus demonstrating that it is, in
principle, possible to remove all the Co and Cu from this phase.
The low concentrations of Co found in the sulphide phases of the
alloy indicate that the possible reaction of CoO + FeS CoS +
FeO is not significant for the cobalt in this system. (This reaction
might, of course, be more significant in systems where the sulphur
content of the alloy approaches that of a matte.)
Copper sulphide and copper metal (with its inherent cobalt content),
because of their higher densities and immiscibility with the slag,
settle from the molten slag into the metal bath. Improving recoveries
via this mechanism would involve decreasing the viscosity of the
slag (by increased temperature or flux additions), and / or increasing
the settling time to allow smaller droplets to fall.
The question might well be asked as to why a DC-arc furnace should
perform better than an AC slag-resistance furnace for this type
of process. There are a number of reasons for this.
Metallurgical flexibility because of independent power supply
Since DC-arc furnaces operate under open arc conditions with the
electrode positioned above the bath, they do not rely very much
on the resistivity of the slag in order to supply energy to the
furnace bath. This renders the energy supply nearly independent
of slag composition, which allows the slag chemistry to be optimized
for the best recovery of valuable metals (instead of for the required
electrical characteristics). When this type of process is operated
in an AC slag-resistance furnace, the degree of reduction cannot
readily be controlled, because the electrodes are immersed in
the slag, and alloys with high iron contents (relative to the
amount of cobalt reduced) result (2,3).
Temperature control
As the current flowing in a DC-arc furnace has to travel through the entire depth of the liquid bath (as opposed to merely between the electrodes of an AC furnace), the temperature distribution is more likely to be relatively even. This is very important for determining factors such as slag viscosity (which is very temperature-sensitive) and density, which play a significant role in allowing the efficient settling of the droplets of alloy. Furthermore, it is possible to achieve a specified power input, almost without regard to the temperature of the slag. This is in marked contrast to the case of an AC furnace, where, as the slag gets hotter, the conductivity increases, thereby limiting the amount of energy which can be dissipated in the slag by the mechanism of resistive heating. This fact limits the temperature that can be obtained in a submerged-electrode or slag-resistance AC furnace. Furthermore, the high iron oxide content, of iron-silicate converter slags in particular, results in high electrical conductivity (9) which does not permit effective energy generation in the melt when using a slag-resistance furnace.
Stable operation
The inherent stability of a DC arc offers the potential for improved
operational control. In an AC furnace, conditions under one electrode
affect the currents in the other two electrodes. In the slag-cleaning
process considered here, the furnace would be operated with a
layer of electrically conductive coke or coal covering the slag.
As the electrodes of the AC furnace penetrate this layer, there
exists the possibility of the current flowing between the electrodes
on the surface of the slag. This inter-electrode conduction not
only reduces the energy dissipated in the slag layer, but would
also result in difficulty in controlling electrode penetration
of the slag.
Because of the difficulty of controlling the position of AC electrodes,
there are often large imbalances in the power distribution. This
results in hot zones, which reduce the freeze-line of solidified
slag near the dissipating electrodes, exposing the refractories
to slag attack. With a closely controlled single electrode, the
freeze-lining on a DC-arc furnace can be better maintained, thus
significantly reducing the slag attack on the refractories.
Electrode consumption and maintenance
AC furnaces used for slag cleaning are usually resistively heated
three-phase three-electrode furnaces, where the electrodes are
in contact with the slag. Unavoidably, the electrodes are attacked
by the highly aggressive slag. Furthermore, electrodes carrying
AC current suffer from the 'skin effect', where most of the electrode
current is forced to a narrow outer band of the electrode, thereby
reducing the electrode's current-carrying capacity. As a consequence
of this, the electrodes have to be larger, thereby exposing an
even greater surface to the reactive slag. The DC-arc furnace
used for slag cleaning operates under open-arc conditions, with
the electrode tip removed from contact with the aggressive slag.
DC electrode current also eliminates the 'skin effect', allowing
a greater current loading (A/cm2) of the electrode.
This phenomenon, added to the fact that there is only one instead
of three electrodes, results in a much smaller electrode erosion
area. These design considerations result in a significantly lower
electrode consumption, anticipated to be around 1 to 3 kg/MWh.
Since the electrodes do not react with the slag, the reductant
addition and consumption can be controlled accurately.
Because of the interactive nature of the electrodes in a three-phase
furnace, a loss of power on one electrode will inevitably result
is a loss of power in the other two. The problem is particularly
prevalent when Söderberg-type electrodes are used, because
such electrodes are prone to breakages, and also need to be baked
in at low current (and hence low power) for extended periods of
time after an upset. Losses of around 1% of total plant production
are typically due to this type of problem.
AC furnaces with Söderberg-type electrodes are cumbersome
to maintain, since electrode casings continually need to be welded
to the three electrode columns, and the paste level in the electrodes
monitored and maintained. On a DC furnace, one electrode section
simply needs to be screwed on to the column at regular intervals
(although this renders the furnace unavailable during the short
power-off down-time).
Structural benefits
Because it has only one electrode to support, the superstructure
of a DC-arc furnace is a lot simpler and cheaper than that required
for a three-electrode AC furnace. The transformer and rectifier
can be placed away from the superstructure, in a convenient location.
Because there is only one electrode, there is room on the roof
to accommodate the suitable positioning of feed chutes to allow
the feeding of coke (or coal) directly onto the slag entering
through the melt inlet. This facilitates good mixing between
the slag and reductant. It is simpler to perform maintenance
above a DC furnace electrode, since there is less electrode clutter
above the roof. The gas seal on a DC furnace is better, as only
one electrode seal, instead of three, is used. There is also
usually less electrode movement, as the arc is more stable, and
the electrode is not interacting with others.
Electrical power supply
The new generation of DC power supplies specifically addresses
the problems of harmonics and flicker associated with AC power
supplies. For example, the Robicon power supply uses a 24-pulse
rectification system with an output chopped at a frequency of
1 kHz, and conforms to the IEEE 519 specification for
both voltage and current distortion. This eliminates the need
for costly harmonic filters, or for flicker-reducing static VAR
compensators. An added benefit of the new-generation DC power
supply is the fact that it is designed to provide low-reactance
power at a power factor of 0.95 or better throughout the required
voltage range, thereby reducing the transformer size and MVA demand,
and eliminating the cost of any external power factor correction.
In some DC-arc furnaces operating at high voltages (and hence
with long arcs) or high process temperatures, problems have been
experienced with uncontrolled or 'stray' arcing between the electrode
and the roof or refractories, resulting in damage to the roof
(and water leaks in the case of a water-cooled roof). With a
suitable roof design, and by maximizing the current / voltage
capability of the power supply, the arc length can be reduced,
and stray arcing minimized.
There are many aspects to the recovery process, and many of these
can be tailored specifically to the particular material being
treated.
Upgrading of cobalt-rich iron alloys
It is possible to upgrade the alloys produced in the DC-arc reduction process by the selective oxidation of iron (2,10). Figure 10 shows the flow of material streams in a process that allows the upgrading of cobalt-rich iron alloys. If the DC-arc furnace is run under strongly reducing conditions (to ensure the greatest possible recovery of valuable metals), the alloy will contain a large quantity of iron. However, if this alloy is subjected to a blowing stage (using air or oxygen), much of the iron can be eliminated into a slag phase. Typically, a silica flux is added in order to combine with the resulting iron oxide to form a fayalitic slag (nominally 2FeO.SiO2). This blowing stage can be carried out in a converter of some kind, or even in a closed furnace utilising a submerged top lance. Tests at Mintek have been done using a top-blown rotary converter (TBRC). The slag from the upgrading step would be recycled back to the DC-arc furnace at industrial scale, in order to recover the portion of valuable metals that have been re-oxidized. It is possible to arrange conditions in the two process units such that an alloy containing only 30% iron is produced, even though a cobalt recovery of 80% is achieved.
Objections to having a converter in the process usually revolve
around the loss of sulphur to the atmosphere. Most modern non-ferrous
smelters are moving away from Peirce-Smith converters for just
that reason. However, there is a vast difference between the
conversion of a high-sulphur matte and the blowing of an iron-rich
alloy. The small quantity of sulphur present in the alloy (a
small fraction of a per cent of the incoming sulphur from the
concentrate to the smelter) is moderately tightly bound to the
metallic elements, and is not too readily given up to the gas
phase. Furthermore, the gas stream is only a waste product from
a secondary stream much smaller than the main stream, and the
very small quantity of sulphur in the gas could be scrubbed.
If this is still seen to be a problem, a stationary enclosed vessel,
such as those used by the new generation of submerged-lance systems,
should be investigated for their suitability to the task.
Kinetics / mass transfer
The gas injection of solid reductant has been found to increase
reaction and mass transfer rates. Stirring (using nitrogen injection,
for example) can also be employed in order to improve the mass
transfer particularly between slag and reductant.
Lime addition
The addition of CaO (up to a certain point) decreases the cobalt
and copper content of slags (9). The activity coefficients of
the valuable metal oxides are increased by the addition of CaO.
In addition, the viscosity of the slag decreases as the slag
becomes more basic (e.g. by the addition of CaO or MgO). This
is important for slag-metal reaction kinetics, and for the settling
behaviour of metal droplets. The liquidus temperature of the
slag also decreases (to a minimum value) with increasing content
of CaO.
It has been found (2) that the kinetics of reduction (for copper
converter slag) are enhanced by the addition of lime, and good
recoveries could be obtained in a one hour reduction period.
It should also be mentioned that the addition of silica and magnesia
can raise the liquidus temperature, thereby reducing the degree
of attack on the furnace refractory lining.
Recovery of platinum group metals
The platinum group metals (PGMs) are often found together with
nickel, copper, and cobalt sulphide deposits. Even in small quantities,
these can be economically significant. The PGMs follow the nickel,
copper, and iron through the pyrometallurgical process, and can
be extracted from the hydrometallurgical leach residues for further
processing.
Water atomization of alloys
Cobalt-rich iron alloys are virtually unbreakable, which poses
a problem of delivery of the alloy to the downstream process units.
It is common practice in a number of slag-cleaning processes
to add sulphur (in the form of pyrite, concentrate, or matte)
to the alloy, in order to make it sufficiently brittle to be able
to be successfully milled after granulation. Apart from the inconvenience
and expense of having to add this material to the furnace, this
sulphur needs to be removed during subsequent hydrometallurgical
processing.
Water atomization, involving the 'smashing' of a stream of molten
alloy with a high-pressure stream of water, can directly produce
fine particles of alloy with a mean diameter of less than 100 µm
(even as small as 40 µm). The design of the atomizing
system is simplified by not having any tight constraints on the
range of particle sizes and shapes of the particles. Small-scale
experiments have been carried out successfully on 5 kg batches
of alloy. This technology is commercially available up to industrial
scale, and appears to be very cost-effective when compared to
the option of granulation and milling. This step introduces another
level of flexibility into the process; one can now optimize the
metallurgy to maximize recovery of the valuable metals, without
needing to be too concerned with the physical properties of the
alloy.
Very little material handling is required after atomization.
A simple screen to separate oversize materials (for recycling
to the DC-arc furnace) is about all that is needed. The alloy
particles can be pumped as a slurry, then de-watered. Drying
is not necessary, as the alloy will next be subjected to a wet
process.
Downstream hydrometallurgical processing
The alloy produced in the reduction process can be leached with
spent electrolyte from a copper or nickel electrowinning process.
Mintek has developed a leaching process that solubilizes the
nickel, copper, and cobalt, while rejecting the iron and any sulphur
into the solid residue (by decreasing the pH appropriately).
The resulting solution can be processed further to separate the
cobalt, nickel, and copper from each other via conventional technology.
Mintek has also developed direct solvent extraction of cobalt
from cobalt-bearing nickel solutions. Nickel and cobalt have
been successfully electrowon from the raffinate and strip liquor
respectively.
The following furnace design specifications can be provided, once
testwork has been carried out on the slag of interest.
DC-arc furnace technology has been successfully applied to the
recovery of cobalt, nickel, and copper from non-ferrous furnace
and converter slags. Pilot-plant testwork at Mintek has demonstrated
recoveries of 98% for nickel and over 80% for cobalt, at power
levels of up to 600 kW.
Mintek expects to participate in the commissioning of the first
commercial units during 1998, and is currently involved with several
engineering companies on final feasibility studies. Mintek intends
to offer this patented technology (11) for implementation world-wide,
and to play an active role in the design, supply, and commissioning
of suitable DC-arc furnaces to meet clients' needs.
This paper is published by permission of Mintek. Contributions
to this work were made by a number of individuals at Mintek.
Particular mention should be made of the pioneering DC-arc smelting
testwork of the late L.B. (Bruce) McRae. Thanks are also
due to T.R. (Tom) Curr for many useful suggestions,
to Dr A.S.E. (Arno) Kleyenstuber, Dr J. (Johan)
Nell, A.D. (Alan) McKenzie, and S.D. (Steve) McCullough for mineralogical analyses, to
Dr M.J. (Mike) Dry for a description of the downstream hydrometallurgical
process, and to E. (Elana) Engelbrecht for the schematic diagram
of the furnace. Helpful discussions with consulting metallurgists
and other personnel from the industry are gratefully acknowledged.
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Copyright © 1996-2001 Mintek, rtjones@global.co.za, GlenD@mintek.co.za
19 June 2001