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NORTH RIVER NAMIB LEAD / ZINC
PROJECT
FEASIBILITY STUDY
Volume 1
Section: 3 Process Development
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TABLE OF CONTENTS
1 PROCESS ENGINEERING ................................................................................ 4
1.1 Process Design Criteria............................................................................... 4
1.2 Mintek Testwork .......................................................................................... 5
1.3 Head Analyses ............................................................................................ 6
1.4 Mineralogy .................................................................................................. 6
1.5 Crushing Work Index and Abrasion Index ................................................... 7
1.6 Bond Ball and Rod Work Indices ................................................................. 7
1.7 Grindmill Testwork ...................................................................................... 8
1.8 Flotation Testwork ....................................................................................... 8
1.9 Flotation of Tailings ..................................................................................... 8
1.10 Flotation of Ore ......................................................................................... 10
1.11 Sedimentation and Filtration Testwork ...................................................... 12
1.12 Mass Flow Testwork ................................................................................. 13
2 PROCESS DESCRIPTION .............................................................................. 13
2.1 ROM Receipt and Crushing ...................................................................... 13
2.2 Tailings Reprocessing ............................................................................... 17
2.3 Milling ........................................................................................................ 19
2.4 Floatation of Ore ....................................................................................... 21
2.5 Flotation of Tailings ................................................................................... 23
2.6 Reagents .................................................................................................. 23
2.7 Sampling ................................................................................................... 24
2.8 Concentrate Dewatering............................................................................ 25
2.9 Tailings Disposal ....................................................................................... 26
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2.10 Process Water .......................................................................................... 27
2.11 Compressed Air ........................................................................................ 28
2.12 Emergency Power ..................................................................................... 28
2.13 Operating and Control Philosophy ............................................................. 28
3 CAPEX AND OPEX ......................................................................................... 28
3.1 Capex ....................................................................................................... 28
3.2 Opex ......................................................................................................... 30
4 CONCLUSIONS AND RECOMMENDATIONS ................................................ 30
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1 PROCESS ENGINEERING
The purpose of the process section of the final report is to:
Compile all process deliverables and process related information,
including process design criteria and mass balance
Describe how the process design and testwork programme were
developed
Describe the process flow and utilities requirements
Explain the mechanical equipment selection for tailings reprocessing
and fresh ore
Explain the operational and control philosophy for tailings reprocessing
and fresh ore
Detail the material and resource requirements for input into capex and
opex calculations
Review the capex and opex estimates
Make recommendations for the implementation of the project.
1.1 Process Design Criteria
The Process Design Criteria were baselined from various past studies and
from similar operations. A conservative view has been taken on which
values to use for Bond Work Indices, flotation residence times and other
parameters.
In order to improve the accuracy of the final cost estimate, Mintek was
contracted to undertake a testwork programme which has run in parallel with
the project. A copy of the Mintek testwork proposal and scope of work is in
the report appendices.
The Mintek testwork outputs were fed back into the process design criteria,
project capex and opex where possible. The latest testwork informed the
design of the crushing and milling circuit. However the flotation testwork was
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not completed in time to update the design of the flotation circuit and
downstream equipment, for which layouts, civil and structural costs are
based on the Rev A design criteria e.g. lead and zinc rougher flotation
residence times, concentrate filtration rates and tailings settling rates.
The process design allows for two distinct plant feeds, namely zinc-rich
tailings from the existing tailings dam and fresh ore from the underground
mine. North River Resources (NRR) requested the option of a tailings
retreatment facility, comprised of a closed circuit ball mill, followed by zinc
flotation. Reprocessing of tails could proceed in parallel with mine
development and the construction of a front-end crushing circuit as well as a
lead flotation circuit upstream of the zinc flotation circuit. Alternatively
tailings could be retreated at a later date, such as prior to mine closure. The
process design allows for easy switching between tailings and fresh ore,
although the optimum ball charge is different.
The processing of tailings is important both economically and strategically
for NRR. However, the economics of fresh ore processing are more
attractive than for tailings reprocessing. Therefore the process design is
based on fresh ore, and tailings will be reprocessed at whatever rate the
capacity of the overall circuit will economically allow. The mill will have the
capacity to mill 250 000 tpa of crusher product to a target grind of 80%
passing 75 microns. The capacity of the mill for tailings will be higher due to
their finer size distribution. However, the retention time of the flotation plant
will be reduced at higher tonnages, with consequent losses of recovery.
The design criteria value of 300 000 tpa of tailings was chosen to represent
a reasonable trade-off between milling throughput and flotation recovery.
1.2 Mintek Testwork
The aim of the testwork was to confirm or update the ore characteristics and
key process design criteria:
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Mineralogical analyses and head assays for tailings and fresh ore
samples
Crushing work index and abrasion index for ore and waste
Bond rod and ball mill indices for ore
Grindmill testwork to establish the specific energy for ball milling
tailings
Flotation circuit configuration for ore and tailings
Grades and recoveries for ore and tailings
Mass pulls for ore and tailings
Flotation kinetics for ore and tailings
Reagent consumptions for ore and tailings
Sedimentation rate and flocculent consumption for final tailings
Filtration rates for lead and zinc concentrates.
1.3 Head Analyses
The tailings sample assays 2.98% zinc, slightly higher than the value used
for the PDC. However the ore sample assays 3.91% zinc and 1.94% lead,
which are somewhat lower than the PDC values. There is 23.4% iron in the
ore sample, rising to 25.8% in the tailings sample. Calcium measures 12%
in the tailings and 15.3% in the ore. Total sulphur is 15.4% in the tailings and
13.4% in the ore.
1.4 Mineralogy
The mineralogical reports show that, apart from being depleted in galena,
the tailings have other noticeable differences compared to fresh ore, i.e.
Slightly lower in sphalerite (4.89% vs. 6.13%)
Much higher in Fe-Oxides (11.96% vs. 2.85%)
Lower in calcite (18.82% vs 30.57%)
Much higher in sulphates (5.43% vs. 1.03%).
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The high sulphates combined with the low calcite means that the tailings
slurry samples have a low natural pH of around 6 and there appears to be
some buffering in solution. This means a lot of alkali is required to reach
target pH’s of 10-11, making some commonly-used flotation recipes
uneconomical.
Oxidation of sphalerite is not apparent from microscopic analysis, but
surface oxidation may explain why the sphalerite in the tailings sample is
apparently not activated by copper sulphate.
Electron microprobe analysis shows that the galena is composed of lead
and sulphur only. However, the sphalerite contains on average 10.48% iron
in the tailings and 11.03% iron in the ore. These represent minimum values
that are achievable in the concentrate if the sphalerite were quantitatively
recovered.
1.5 Crushing Work Index and Abrasion Index
The crushing work index of 10.87 ± 1.89 kWh/t for the ore sample is
indicative of soft ore, verging on very soft. The waste rock was very soft with
a CWI of 8.69 ± 1.63 kWh/t. The Bond Abrasion Indices for ore and waste
were both less than 0.1, where less than 0.2 is indicative of low abrasivity.
These results indicate that single stage impact crushing of ROM ore with a
maximum size of 400 mm to a ball mill feed size of 15 mm is achievable
without causing excessive equipment wear.
1.6 Bond Ball and Rod Work Indices
The Bond Ball Work Index of 10.9 kWh/t at a closed screen size of 106 µm
and Rod Work Index of 10.7 kWh/t are also indicative of soft ore. Despite
the low nominal tonnage of 32 tph a motor power of 450 kW is still required
to achieve the fine target grind of 80% -75 µm.
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1.7 Grindmill Testwork
The grinding of tailings is not accurately predicted by Bond’s Law. Batch
milling tests are therefore used to derive breakage and selection functions,
which are then used to simulate the continuous circuit and output the
specific energy requirements.
The Mintek Grindmill testwork gives a specific energy of 3.72 kWh/t for
grinding of tailings to 80% -75µm using a ball top size of 30 mm. This rises
to 6.46 kWh/t with a ball top size of 70 mm. This means that there is the
potential to mill tailings at a much higher rate than fresh ore. In practice the
processing of tailings is likely be limited by the recovery versus retention
time relationship in the zinc flotation circuit, which has been designed for 20
minutes rougher retention time for a fresh ore feed of slightly less than 30
tph and 30% solids by mass.
The higher energy requirement for milling tails with large balls means that
switching the plant feed from ore to tailings would result in high electrical
consumption if the ball charge was not switched to smaller balls. Switching
from tailings to fresh ore would be easier as it would only require larger balls
to be added to the mill.
1.8 Flotation Testwork
The majority of the Mintek testwork was to confirm flotation process design
criteria. These include the effect of grind, rougher flotation rates, number of
cleaning stages, reagent addition points and reagent dosages, and finally
locked cycle cleaner tests. The flotation of tailings and fresh ore are
discussed separately.
1.9 Flotation of Tailings
The tailings dam sample contained over 32% fines in the -38 micron
fraction. The first scouting test used a polypropylene glycol methyl ether as
frother, copper sulphate as activator and PAX, a strong collector, at the
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natural pH of the slurry –made with fresh water from site. Sodium silicate
was used to prevent coating of ore particles with slimes and increase
selectivity, but a low recovery (51.7%) was obtained at a high rougher mass
pull (22.9%).
The second scouting test used lime instead of sodium silicate with the aim of
depressing pyrite. This approach was recommended by the plant foreman of
a tailings retreatment plant which ran between 1990 and 1992. The recovery
dropped to only 34%.
The tailings dam is now at least 22 years old and there is evidence of
oxidation of iron minerals such as pyrrhotite to sulphate. For this reason, 5
kg/t of lime was required to raise the pH to 10.5 and 8.5 kg/t was required to
raise the pH to 10.8. Therefore some alternative to the use of lime as a
pyrite depressant was indicated. Addition of Cytec 7261 iron depressant
improved the recovery but did not improve the selectivity.
The use of sodium silicate was revisited, this time conditioning at 5 kg/t and
70% solids. A zinc recovery of 74% was achieved without milling, increasing
to nearly 90% when milling to 80% passing 75µ. However, mass pulls were
high, zinc grades were low and iron grades were high.
By reducing the reagent dosages and mass pull, better selectivity could be
obtained, but at the expense of recovery. The best quality concentrate
contained 46% zinc and 19% iron at a recovery of 35%. Striking a better
compromise between grade and recovery has proved to be a difficult
challenge, and the extra effort put into this testwork contributed to delays in
the overall Mintek testwork program.
Aeration of the pulp was tried at a late stage of the testwork with the aim of
preferentially oxidising the iron sulphides and depressing them. The results
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were promising, but the detailed assays are not available to date. Therefore
an economical process for tailings cannot be ruled out at this stage.
The municipal water supplied for the testwork was originally at pH 8, but
over the prolonged testwork period, the pH climbed to 10.7 and there was
some precipitation of solids. This effect may prove beneficial in treating
tailings material which has a low natural pH of around 6.
1.10 Flotation of Ore
Initial results with ore showed faster than expected lead flotation kinetics,
with a recovery of over 90% lead achievable at less than 5% mass pull.
Initially some zinc reported to the lead concentrate, but this was reduced to
an acceptable level by reduction of the collector dosage in the lead float.
The flotation of ore follows a more conventional route than for tailings. A
small amount of hydrated lime (200 g/t) is used to raise the pH in the mill
above 8.5 to prevent formation of hydrogen cyanide. 300 g/t of zinc sulphate
and 80 g/t of sodium cyanide are then added to the mill to depress zinc in
the lead float. 365 g/t more lime, 30 g/t of XP200 frother and 15 g/t of
sodium ethyl xanthate are then added to the conditioner. The lead
concentrate is then subjected to cleaning and re-cleaning stages.
A flotation grade of 62.5% lead, 1.0% zinc and 4.1% iron was achieved at a
recovery of 83%, which is in line with the process design criteria. The
laboratory rougher flotation retention time was 3 minutes.
Zinc flotation kinetics were also fast, but mass pulls were initially too high,
indicating that reagent dosages for the zinc float were also high.
Flotation of zinc starts with the addition of 1.9 kg/t of lime to conditioner 1 to
raise the pH to 11.5 and the zinc is then activated with 500 g/t of copper
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sulphate. 15 g/t more XP200 frother is added to conditioner 2, with 60 g/t of
SEX and 125 g/t of Aero 7261 iron depressant.
A flotation grade of 53.0% zinc and 13.4% iron was achieved at a recovery
of 83%, which is in line with the process design criteria. A recovery of 73%
silver at a grade of 170 g/t was also achieved (telephonically reported). The
laboratory rougher flotation retention time was only I minute, however, and
this process would be very difficult to control in practice. For this reason
Mintek recommends the use of the Cytec iron depressant, which slows the
kinetics. Therefore the baseline conditions for locked cycle tests are still to
be determined, based on the results with depressant. Aeration was a
variable which was incorporated at a late stage in the testwork, and is also
showing promising results. However, Mintek remains confident of the grades
and recoveries achievable for lead, zinc and silver.
There was some evidence of oxidation of the ore sample reported by
Mintek. Fresh ore should require less lime to raise the pH. Because of the
recycling of reagents, the locked-cycle tests should give reduced reagent
consumptions that will be a better prediction of continuous plant operation.
The full-scale retention times must be adjusted to compensate for the faster
kinetics associated with a small laboratory batch cell. However this requires
specialised testwork and modelling, which are beyond the scope of this
study. Instead the following scale-up factors can be used as a basis for
comparison of the PDC with the testwork:
Flotation Stage Scale-Up Factor
Conditioner 100% (i.e. none)
Rougher 200%
Cleaner 250%
Recleaner 300%
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The practical implications are that the conditioner and rougher cell volumes
are considerably larger than required, the cleaner cells are slightly larger
and the recleaner cells are slightly smaller, based on the PDC. However,
additional recleaner capacity was included in the design to allow for higher
than expected froth carrying and lip loading capacities associated with
pockets of high grade ore.
Concentrate mass pulls obtained from testwork are considerably lower than
allowed for in the PDC. The rougher mass pull for lead is a quarter of the
design value, while the rougher mass pull for zinc is half the value in the
PDC. This means that higher than expected head grades and recoveries
should not overload the concentrate sumps. Again, locked-cycle tests will
give the best prediction of the plant retention times and mass balance.
The testwork results therefore confirm the conservative process design
basis used for the flotation plant, and indicate that capex can be reduced
here.
1.11 Sedimentation and Filtration Testwork
Sedimentation tests were planned for the flotation tailings and sedimentation
and filtration tests were planned for the concentrates. This testwork included
optimisation of flocculent addition for the tailings thickener.
The testwork on filtration rates was scheduled to follow the locked-cycle
testwork, which will produce the most representative samples of
concentrate. Flotation testwork delays means that suitable samples have not
been tested to date. There is a concern that, at the fine grind employed, it
may not be possible to produce a filter cake with the target moisture content
of 9% at the specified tonnage. This problem could be exacerbated by poor
blending of ROM ore, resulting in higher than expected concentrate
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production. However, given the low rainfall and high evaporation in the area,
NRR is confident that producing dry concentrates will not be a problem.
Sedimentation testwork on tailings will also follow the locked-cycle testwork.
1.12 Mass Flow Testwork
Mintek was responsible for all the testwork apart from the mass flow
testwork, which was undertaken by Green Technical, an independent
consultant, using the Tenova Bateman laboratory facility.
The 0-4 mm size fraction is critical in determining the bulk flow properties of
solids. The fine ore bin and feeder design for crusher product / ball mill feed
was therefore based on bulk solids flow testwork undertaken by
Greentechnical (Pty) Ltd. The large opening at the bottom of the bin is
dictated by the need to avoid bridging and rat-holing from occurring. This in
turn requires an oversized belt feeder to feed the mill. A cheaper vibrating
feeder of the required size for the silo opening would not be able to control
the low feed rate.
2 PROCESS DESCRIPTION
2.1 ROM Receipt and Crushing
OVERSIZE
SCREENIMPACT
CRUSHERUNDERSIZE TO MILLING
ROM STOCKPILE
ROM BINFRESH ORE ONLY
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MOBILEIMPACT
CRUSHER
ROM STOCKPILE
ROM BINFRESH ORE & TAILS UNDERSIZE TO MILLING
The ore body is not homogeneous, but consists of marble host rock
interspersed with zones of massive galena and sphalerite. For this reason,
some blending should be undertaken to ensure a relatively constant feed
grade to the concentrator plant, even though double handling of the ore will
have a cost implication. Multiple stockpiles will increase the amount of
handling involved and will require extra space, and so a single blending
stockpile will be used.
ROM ore with a top size of 400 mm will be brought to the tip in 15 t trucks.
The ROM tip will have a capacity of 2000 t of ROM ore. Trucks will be able
to drive onto the tip and spread their loads laterally to facilitate blending. A
front end loader will profile the ramps on the tip and will transfer ore from the
tip into the ROM feed bin, further blending the ore so as to reduce
fluctuations in the lead and zinc head grade. The ROM feed bin will be able
to accommodate 50 t of ore. The bin will be equipped with a static grizzly
with a 400 x 400 mm aperture size. The small scale mining method is not
expected to generate much oversize material. Any oversize material will be
pushed off the grizzly and broken with a hammer in a demarcated area next
to the bin where no pedestrian traffic will be allowed. Broken material will be
retrieved by front end loader as required.
Several options were considered for the crushing circuit. The most
conservative option included a primary jaw crusher and secondary cone
crusher in closed circuit with a vibrating screen. However, there was early
evidence that the ore was relatively soft and of low abrasiveness, which was
later confirmed by Mintek testwork. This allowed consideration of impact
crushers which could take the ROM ore from a top size of 400 mm to a top
size of 12-15 mm in a single crushing stage.
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The use of impact crushers means that extra attention should be paid to the
accompanying high noise levels and dust generation.
Consideration was given to a mobile impact crusher, which has the
advantage of minimal installation cost. This option is the most flexible as it
lends itself to the treatment of both tailings and fresh ore, while minimising
redundant equipment once the tailings resource has been exhausted.
A modular skid mounted impact crushing circuit was also considered since
the ground conditions are particularly favourable. However, the extra
simplicity and reduced cost of operation and maintenance of a modular skid-
mounted plant means that this option is recommended if only fresh ore is
treated.
The modular skid-mounted concept also allows for reduced construction
time and reduced civil costs when compared with more permanent crushing
installations. This approach was used wherever it was practical to do so.
The modular crushing circuit is supplied as a complete skid-mounted
package by Metso. It comprises ROM bin (26 m3) with fixed grizzly (400 x
400 mm aperture), vibrating grizzly feeder for pre-screening of -38 mm fines
from the crusher feed , impact crusher (110 kW), single deck closed circuit
vibrating screen (4 m L x 1.5 m W), plus all conveyors and transfer chutes. A
belt magnet and metal detector are included. In order to cut costs, Metso
have used narrow conveyors (500 mm) and steep conveyor lifts of 16-18
degrees. When these lifts were queried, Metso supplied the diagram below,
which shows that the lifts are safely within Metso’s design envelope.
However, there is still a concern about the belt angle at the loading point,
which we would recommend not being more than five degrees, and this
should be revisited at the implementation stage.
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If an economical recipe for tailings reprocessing is forthcoming from
testwork, then a ball mill feed must be produced from the hard aggregated
tailings. These tailings are likely to contain some tramp metal and rocks. A
mobile impact crusher would be ideal for this application, designed for
building rubble containing structural steel and concrete. This would make a
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skid-mounted crushing plant redundant, since the mobile crusher could
alternate between tailings and fresh ore.
The crushed fresh ore will be stored in a 300 ton fine ore bin. This will allow
just less than 10 hours of milling time in the event of crushing circuit
downtime. It should be noted that the availability of the mobile crusher is
likely to be less than the fixed crusher circuit. Because there is a 300 ton
fine ore bin, the mill availability can be maintained for fresh ore. However,
for tailings reprocessing the fine ore bin will be bypassed and the mill
availability will be dependent on the mobile crusher availability.
2.2 Tailings Reprocessing
CYCLONEMECHANICAL
RECLAIMOVERFLOW TO
ZINC FLOTATIONBALL MILL
MOBILEIMPACT CRUSHER
UNDERFLOW
The initial project strategy involves initial reprocessing of approximately
700kt of existing tailings through the ball mill and zinc flotation circuit while
the underground mine is developed. The crushing circuit and lead flotation
circuit would be constructed in time to receive fresh ore from underground.
However the strategy is flexible, and it may be economically preferable to
defer tailings reprocessing to the end of end of the project.
The tailings dam is comprised of a solid crust of agglomerated material at
the surface which gradually transforms into a slimy liquid at depth.
Mechanical reclamation of the surface tails to ball mill feed size could be
achieved by impact crushing, but the major unknown is the location and
extent of the wet material which would require specialised hydraulic
monitoring.
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The uncertainty in the depth and location of the slimes means that even
mechanical reclamation may require experience in tailings dam
management. The risk increases for hydraulic reclamation, where tailings
dam management experience is essential, and where the investment in
equipment cannot be justified without knowing the amount of slimes to be
treated.
Furthermore the companies which specialise in hydraulic mining of tails
(Fraser Alexander & Steffanuti Stocks) do not design systems but prefer the
build/own/operate/maintain business model.
For these reasons, and to guarantee no equipment redundancy, only the
mechanical reprocessing of tails will be allowed for initially. It is likely that
the residual moisture in the dump has dropped and that, once wet areas are
exposed, this moisture will quickly evaporate. Therefore the need for
hydraulic reclamation can be re-evaluated at a later date.
The mechanically reprocessed tails will be discharged by front end loader
into the feed hopper of a mobile impact crusher (capacity 2.6 m3). The
mobile crusher has a maximum capacity of 160 tph (depending on the ore)
and is equipped with a feeder, pre-screen for fines removal, impact crusher,
closed circuit screen and recirculating belt to return oversize to the crusher
feeder. The screen and pre-screen undersize will combine and be
discharged onto a transfer conveyor. The transfer conveyor will discharge
via a transfer chute onto the mill feed conveyor. The reprocessed tails will
bypass the fine ore bin in order to avoid re-compaction.
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2.3 Milling
BALL MILL CYCLONE
UNDERFLOW
Lime, NaCN & ZnSO4
FINE ORE BINCYCLONE OVERFLOW TO LEAD FLOTATION
CRUSHER PRODUCT
A conventional overflow ball mill was selected because it is a simple and
robust design in comparison to a grate discharge mill. It is also preferred for
fine grinding because an overflow mill has a longer retention time than the
same size grate discharge mill. A rubber lining was preferred, as it is quicker
and easier to install, giving a lower installed cost. Other advantages are
reduced maintenance and reduced noise levels. Reagent additions are for
fresh ore only.
The nominal throughput for fresh ore is 250 000 tpa, which at 90%
availability equates to 32 tph. The target grind is 80% -75 microns although
the flotation testwork has revisited coarser and finer grinds for both tailings
and fresh ore. In order to generate a narrow particle size distribution for the
flotation feed, the ball mill is in closed circuit with a hydrocyclone and a
circulating load of 250% was chosen as a suitable compromise between
efficient milling and the size of the mill discharge pumps.
The mill discharge pumps and other main slurry pumps in the milling and
flotation areas will employ expeller seals in order to avoid having a separate
gland service pumping system. The mill discharge, lead tailings, zinc tailings
and tailings underflow pumps will be variable speed and will have standby
capacity. Tech-Taylor valves are specified on the discharge lines to allow
rapid changeover or starting of a standby pump.
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A 400 mm diameter cyclone was chosen because it allows for flexible
operation i.e. higher volumetric throughputs to allow tailings to be treated at
a higher tonnage than fresh ore. Fresh ore will be fed using the conventional
configuration, while tailings can be treated using a reverse configuration, i.e.
feeding the mill discharge sump rather than the feed hopper.
It will be possible to mill tailings at a substantially increased rate compared
with fresh ore. Based on a specific energy of 3.72kWh/t, it would be possible
to mill at over 100 tph. If the standby slurry pumps are started on the main
process lines the increased pressure drop means increasing the flow rate to
match the mill’s capacity for tailings will not be possible. However a
substantial increase in throughput would be achievable, particularly if the
zinc flotation were run at higher feed % solids and/or lower retention time. A
nominal capacity of 300 000 tpa has been chosen as a reasonable
compromise between throughput and recovery.
Running the plant at higher throughputs of tailings without sacrificing
retention time would be possible if the lead and zinc circuits were run in
parallel for zinc recovery, but this would require minor modifications to the
existing plant design and early construction of the lead flotation circuit. The
optimum ball filling and ball size distribution will be different for fresh ore and
tailings, with a lower ball filling and smaller balls required for tailings.
Therefore it is preferable not to switch from tailings feed to ore and back
again.
The top-up costs for forged steel balls were quoted at N$12589/t delivered.
The equivalent high-chrome balls are more expensive at N$16785/t, but
have a longer wear life. If the high chrome balls wear at 75% of the rate of
the steel balls, then the effective cost is the same. Claims have been made
that reductions in media consumption of 60% are possible, with no adverse
effect on metallurgy, i.e. 40% of the wear rate of steel balls (Ref. “The
Application of High-Chrome Grinding Media at MMG Century Mine for
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Improved Grinding Media Consumption and Metallurgy Performance”,
Greet, Obeng, Kinal and de Bosscher, Metallurgical Plant Design and
Operating Strategies (MetPlant 2013), 15 - 17 July 2013, Perth WA). This
translates to potential savings of nearly 50% in opex. However, steel balls
were chosen for the opex calculations, as these were used by Mintek for the
testwork.
2.4 Floatation of Ore
LEAD ROUGHERS
LEAD CLEANERS
LEAD RECLEANERS
LEAD CONDITIONER
CYCLONE OVERFLOW
ZINC CONDITIONER
1
ZINC CONDITIONER
2
ZINC ROUGHERS
ZINC CLEANERS
ZINC RECLEANERS
LIME & CuSO4
XP200 & SEX & LIME
XP200 & SEX & AERO 7261
TAILINGS THICKENER
ZINC CONC FILTER
LEAD CONC FILTER
The skid-mounted modular philosophy was also discussed with flotation cell
vendors. However, none of the proposals included this as an option. For in-
house design, the traditional concrete “mushrooms” are considered more
cost-effective than designing steel skid-mounted structures from scratch.
Based on historical testwork, which indicates the lead and zinc sulphides are
fast-floating and that iron is slower floating, a rougher/cleaner/recleaner
configuration (no scavenger cells) was chosen for both lead and zinc
flotation, subject to confirmation by the current Mintek testwork. The
reagents chosen have been in common use for many years. The exception
is Aero 7261, a newer product which has proven effective in depressing
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pyrrhotite. XP200 is a polypropylene glycol methyl ether frother. A cheaper
frother could be tested at a later stage.
Flotation cells will be individually stepped tank cells with forced air
introduced via low pressure blowers (part of vendor package). The air flow
and the cell levels will be automatically controlled. This level of automation is
considered essential for concentrate product quality control.
The retention times chosen for the flotation circuit were conservative. This
means that the cell volumes, froth carrying capacity and cell lip loading will
also be conservative. However, the nature of the ore body is such that high
grade pockets of sphalerite and galena are likely to come through to the
float even though there is some blending of ore at the ROM tip. A safety
factor is therefore prudent, although too much retention time can result in
the unwanted recovery of slower-floating iron sulphides.
Lip loading and froth carrying constraints will be most tested in the final
cleaning stage. For this reason the circuit design has a 200% safety factor
for zinc and a 250% safety factor for lead on the recleaner cell residence
times. The different values are dictated by the standard flotation cell sizes,
and not because a higher safety factor was required for lead.
Combined sump pumps were specified for concentrate sumps. These have
the disadvantage of lower mechanical efficiency, but the advantage of
increased simplicity and flexibility over the traditional centrifugal pump /
sump arrangement.
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2.5 Flotation of Tailings
ZINC CONDITIONER 2
ZINC ROUGHERS
ZINC CLEANERS
ZINC RECLEANERS
TAILINGS THICKENER
UNDERFLOW TO TSF
OVERFLOW TO PROCESS WATER TANK
ZINC CONC FILTER
ZINC FILTER CAKE
O/F TO PROCESS WATER TANK
Na2SiO3 & CuSO4 & SEX
LIME & CuSO4 & SEX
CYCLONE OVERFLOW
ZINC CONDITIONER 1
The reprocessing of tailings should ideally be accommodated within the
existing flowsheet. If attritioning and aeration are required, as seems likely,
then this could be accomplished by earlier construction of the lead flotation
circuit, provided that a lead removal step is also not required because of
smelter penalty charges.
2.6 Reagents
Reagents will be stored in a dedicated storage area. They will be made up
on a daily basis and stored in tanks using a skid mounted concept. Most of
the reagents will be delivered in solid form. Each solid reagent will require
two agitated tanks: one for mixing and one for dosing. When the dosing tank
is empty, it will become the mixing tank for the next batch of reagent.
Relatively high reagent concentrations were chosen so as to reduce the
consumption of fresh water for dissolution, i.e. 10% for xanthate and
sulphate, 20% for cyanide and 30% slurry for hydrated lime. The liquid
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reagents that do not require mixing will be pumped directly from the
containers that they are delivered in.
The reagents selected and the dosage rates were based on historical
testwork, particularly the work undertaken by Kupfermelt Metal Processing.
The reagent selection, dosage points and dosage rates will be updated for
opex purposes, and the capex implications will be discussed.
Each reagent will be delivered by a dedicated peristaltic pump to a particular
point in the process using flexible stainless steel tubing. Dosage rates can
be varied using variable speed drives. Flow rates will be measured and
controlled manually.
Reagent spillage will be pumped to the final tailings sump, except for lime
which is contained within a dedicated sump and can be pumped back into
the mixing or dosing tank.
2.7 Sampling
Sampling of lead flotation feed (or zinc flotation feed if tailings are
processed) will be by means of a two-stage cross-cut sampler. Sampling of
lead flotation tails and zinc flotation tails are by single stage cross cut
samplers of the same size as the flotation feed primary sampler. Therefore
there may be the opportunity to save costs and reduce the height of the
cyclone structure to accommodate a single stage sampler, if a slightly larger
sample is collected.
Sampling of lead and zinc flotation concentrates will also be carried out by
single stage cross cut samplers
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2.8 Concentrate Dewatering
LEAD CONC FILTER
LEAD FILTER CAKE
FILTRATE TO PROCESS WATER
TANK
LEAD FILTER FEED TANK
LEAD RECLEANER
CONC
ZINC CONC FILTER
ZINC FILTER CAKE
FILTRATE TO ZINC PROCESS WATER TANK
ZINC FILTER FEED TANK
ZINC RECLEANER
CONC
The process design was initially based on thickening followed by filtration.
Filter presses were chosen because of their simplicity and cheapness. Filter
presses with manual discharge of cake and cloth washing were considered
the most appropriate design for the duty. However, the filter press sizes for
thickener underflow feed were only slightly smaller than if they were fed with
thickener feed. Therefore concentrates will be filtered only.
The lowest capex option was chosen. (Delkor FAST Filter Presses). This is
a new design with a smaller footprint, but this option should only be chosen
for the construction phase if there are at least a few suitable local reference
installations, as there are some reservations about the closing mechanism,
which doesn’t use the typical hydraulic ram.
Confirmatory testwork will be required to confirm throughput and moisture
content. The final moisture specified may necessitate the use of more
expensive filter options, such as membrane squeezing, or the use of more
expensive Larox-type vertical filter presses, which are employed in many
concentrator plants. However, air-drying is an obvious option in a desert
environment to mitigate process risk.
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There is a concern that pockets of high-grade ore may reach the plant and
overload the concentrate filters. This could be overcome by increasing the
capacity of the filter feed tanks. A cheaper alternative would be to build a
bund wall around the filter feed tanks, and allow them to overflow during
periods of high concentrate production. The products could then be
manually returned to process when capacity allows.
Lead concentrate filtrate will be recycled as process water. Zinc concentrate
filtrate will only be recycled within the zinc section, to avoid activation of zinc
in the lead float.
2.9 Tailings Disposal
TAILINGS THICKENER
U/F TO TSF
O/F TO PROCESS WATER TANK
CATCHMENT DAM RETURN
ZINC ROUGHER TAILINGS
Flotation tailings thickening will be carried out in a high-rate thickener. A
high-rate thickener will have a lower capex than a larger conventional
thickener, although there is an opex penalty because of increased flocculent
consumption. A skid-mounted manually operated flocculent plant has been
included. A hydrocyclone is often specified upstream of the thickener, with
the cyclone overflow being thickened and the thickener and cyclone
underflow being combined and pumped to tails. This level of complexity is
not justified for the scale of operation, given the relatively small potential
saving in thickener size.
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Where centrifugal pumps are specified for thickener underflow, vendors
usually prefer a gland seal pump or expensive mechanical seals. For this
reason, seal water will be taken from the high pressure potable water line.
The thickener underflow will be pumped to a final tailings sump, from where
it will be pumped to the tailings dam.
Tailings thickener overflow will be recycled as process water.
2.10 Process Water
RAW WATER MAKE-UP
PROCESS WATER TANK
THICKENER OVERFLOW
BALL MILL, LEAD FLOTATION, ETC.
The water requirement for the project will be potable quality and therefore
expensive. The tailings dam design indicates a recovery of 49-59% of the
water that is pumped to tailings. Concentrate products will also be
dewatered in filter presses and the water recycled.
The recycling of copper sulphate from the zinc concentrate filtrate to the
lead circuit must be avoided, since otherwise zinc will be activated and will
report to the lead concentrate. The most conservative option is to have
separate tailings thickeners after lead flotation and zinc flotation, and to
keep the water streams separate, but this adds substantially to capex.
What has proven to work on a large scale operation is to recycle the water in
the zinc concentrate stream to the zinc circuit only, since the copper
sulphate is mainly contained in the zinc concentrate stream, and this is the
preferred design.
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2.11 Compressed Air
PROCESS AIR COMPRESSOR
FILTER PRESSESMILL GREASECONTROL VALVES
BACK-UP COMPRESSED AIR FROM MINING
AMBIENT AIR
The major users of compressed air will be the filter presses. Additional
compressed air is required for application of grease to the mill girth gear, as
well as for the control valves in the flotation circuit.
A single compressor has been sized for this duty, with three air receivers -
one air receiver for each of the two filter presses and an additional air
receiver for the remaining plant. The plant air will be dried by a refrigerator
drier.
A back-up compressed air supply line at 7 bar will be taken from the mining
compressor station.
2.12 Emergency Power
Electricity will be supplied by NamPower. An emergency generator will
supply power to the thickener rake and filter feed tank agitators in the event
of a power outage. Emergency lighting will be battery-powered.
2.13 Operating and Control Philosophy
Refer to Volume 2 – Process in the report appendicies.
3 CAPEX AND OPEX
3.1 Capex
The limited mine life means that the project is highly capex-sensitive. Capital
cost cutting measures include the following:
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Design factor of 10% for transfer equipment (pumps & conveyors) only
Standby equipment limited to the main process flow and process water
pumps
Common discharge line for running / standby pumps
HDPE piping more economical than rubber-lined steel (although
increased support required).
Single stage impact crushing to reduce ROM ore to ball mill feed size
Skid-mounted crushing circuit to reduce construction time and civil
costs
High-rate thickener instead of conventional thickener for tailings
Filtration of tailings without intermediate thickening
Flotation cells limited to a minimum of three in series
Reduction in fine ore bin size by increasing crushing plant operating
hours
Reduction in reagent tank sizes by daily make-up.
Skid-mounting of reagent tanks.
Reduction in capex inevitably implies increased opex, and a fit-for-purpose
philosophy should ensure that opex penalties are not excessive. For
instance, the opex of a single stage crushing plant would be prohibitive if the
ore had a high crushing work index or was abrasive, but the ore is soft and
non-abrasive. The relative capital costs obtained from a single supplier are
summarised below.
Crushing Plant
Description
Equipment Cost
(Uninstalled)
2-Stage Jaw / Cone N$ 10.192 m
2-Stage Jaw / Impact N$ 7.343 m
Single-Stage Impact N$ 5.849
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For instance the capex for the German-built Rubble Master RM80 Go mobile
crusher of N$ 4.371 m is less than half the cost of a Metso mobile crusher
with similar specs (closed-circuit screen for proper ball mill feed size control)
at N$ 9.456 m. The Metso machine features a bigger crusher which is also
bigger than the fixed impact crusher specified by Metso for the same duty.
Although the Metso mobile crusher was not favourably priced, the Metso
fixed crushing plant was the cheapest offering that was technically
acceptable (incorporating feed to the fine ore bin). It should be noted that
Metso mobile equipment is sold through a third party (Barloworld
Equipment).
Although capex reduction was a priority, there are opportunities to make
further reductions e.g. by using a contractor or purchasing second-hand
equipment, provided the risk is acceptable.
3.2 Opex
Refer to Section 7 of this report.
4 CONCLUSIONS AND RECOMMENDATIONS
The Mintek testwork on tailings was exhaustive, and delayed completion of
the testwork on fresh ore. There were indications at the time of writing that
the zinc can be extracted profitably, but the margin would always be low at
best, especially when compared with the economics for fresh ore.
The fresh ore has excellent milling and flotation characteristics, and high
recoveries and grades are achievable. Controlling the flotation process has
proven to be a challenge because of the faster than expected flotation
kinetics, so determining the optimal flotation cell configuration, optimum
residence times and the right reagent dosages will be critical to the
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profitability of the project. Locked cycle testwork will be the best method
(after pilot plant testwork) of finalising the flotation plant design, and will
produce the most representative samples to complete the flitration and
sedimentation tests, and to update the process process design if required,
hopefully with cost savings.