View
2.895
Download
4
Embed Size (px)
DESCRIPTION
Expositor: Esther Rodríguez // Auditorio del Minem // 10-01-2013
Citation preview
PT Nº 001
Enero 2013
Geometalúrgia, Procesamiento de Oro
Sostenible al Medio Ambiente y diseño de
Planta
Esther Rodriguez
B.Eng. (Chem), M.Env.Eng. Sc., PhD Minerals Sc.
I. The Geo-Metallurgical Study – what this means and why is important.
II. Stages for the Development of a Gold Plant – from the concept to construction.
III. The Design of an Environmental Sustainable Gold Plant.
Geo-Metallurgy is the study of the drivers for the
metallurgical response of an ore deposit based
mainly on its mineralogy.
In other words; what is required to investigate to
recover profitable a valuable metal from a specific
type of ore.
Generally, the main topics to cover are
The several type of mineral species: valuable metal and gangue.
The size of grains.
The strength and abrasion of the ore to be processed.
The liberation of a valuable metal respect to the ore size.
The comminution test work.
The flotation and hydrometallurgy test works.
The selection of the flow sheet.
Process simulation for optimisation of plant performance.
Models to characterize the ore variability throughout the mine-life are not considered.
Source: Metso Handbook
Some common minerals found in mineral processing:
Native Gold, Au: > 75% of Au; S.G. 16 – 19.3; deep yellow
Electrum, (Au,Ag): 45 to 75 % of Au; S.G. 13 – 16; pale yellow
Calaverite, AuTe2: 39 to 43 % of Au; S.G. 9.2; white or creamy yellow
Sylvanite, (Au,Ag)2Te4: 24 to 30 % of Au; S.G. 8.2; creamy white
Iron Sulphides: pyrite (FeS2), arsenopyrite (Fe,As,S), pyrrhotite (Fe1-
0.8S)
Copper Sulphides: chalcocite (Cu2S), chalcopyrite (CuFeS2)
Other Metallic Sulphides: sphalerite (ZnS), galena (PbS)
Quartz: SiO2
Silicates: feldspars XAl(1-2)Si(3-2)O8
Clays: kaolinite Al2(Si2O5)(OH)4
Carbonates: calcite (CaCO3), siderite (FeCO3)
Iron Oxides: goethite (FeO.OH), hematite (Fe2O3)
SEM image of gold mineral grains: native gold associated with pyrite.
Grain # 7: liberated gold - 7 x 4 µm.
Grain # 8: locked gold – 23 x 2 µm – 88% Au.
The knowledge of the gold association and degree of liberation are important
to decide what techniques of extraction should be considered.
The aim of the mineralogical investigations is to characterize the gold mineral
composition, gold grain size, its liberation and association.
Currently it is available automatic equipment that identify and quantify the
different mineral phases in an ore sample and also determine the degree of
liberation.
QEM*SEM (Quantitative Evaluation of Materials using Scanning Electron
Microscope) is one of these modern equipment. It is composed of a computer
controlled SEM fitted with backscattered electron (BSE) and X-ray detector.
Type of gold ore bodies / deposits in relation to weathering.
Lateritic deposits: soft rocks and consists generally of SiO2 and iron oxides. Gold particles are less enclosed.
Supergene deposits: deeper down in the weathered profile, usually near the water table. They form in arid areas where groundwater is very saline that leach the weathered rock. Soluble gold chloride complexes are re-precipitated at the water table by Fe2+.
Primary ore deposits: unweathered ore. Mainly igneous rocks are found in gold deposits (granite, basalt, andesite).
Rock oxidation by weathering:
FeS2(s) + 15/4O2(g) + 7/2H2O(l) → 2SO42-
(aq) + Fe(OH)3(s) + 4H+(aq)
Feldspar Kaolinite
Source: Gold Plant Operators Course, Western Australia School of Mines
Source: Gold Plant Operators Course, Western Australia School of Mines
The crushability/grindability is an important ore characteristic that
quantifies the difficulty to break down the ore by impact, abrasion and/or
attrition forces. It is known as the “Work Index, Wi”.
Wi a key ore parameter that affect directly the energy requirements in the
comminution circuit.
More used Wi are: crushing work index (CWi), rod work index (RWi), ball
work index (BWi).
Generally, the RWi define an ore as:
Soft (less competent): 7 to 9 kWh/t
Medium: 9 to 14 kWh/t
Hard: 14 to 20 kWh/t
Very Hard (more competent): > 20 kWh/t
Each rock and mineral has their specific range of Wi (measured value).
There is a relationship between the area being extracted, the ore
mineralogy and the ore competence. Data below are test results of samples
from the same gold deposit.
Orebody Zone Ore Type RWi BWi
Lateritic Oxide 5 46
Supergene Sulphide + Oxide 14 74
Rock/Mineral BWi, kWh/t
Andesite 20
Basalt 19
Granite 17
Quartz 15
Limestone 14
Hematite 14
Magnetite 11
Pyrrhotite ore 11
Pyrite ore 10
Clay 7
The BWi is defined as the energy required to reduce the ore from infinitive
size to a P80 of 100µm.
The test involves a series of consecutive batch grinds in a standard ball mill
fed with 10 kg of minus 3.5 mm ore sample. After each grind, the
discharge is screened to remove the undersize that is replenished with an
equal mass of a new feed. Testing continues until the weight of the
undersize becomes constant.
The Wi (rod and ball) is used with the Bond’s Third Theory of
Comminution to calculate the specific energy required for a mill (E, kWh/t)
to reduce a tonne of feed of which 80% passes size F80 microns to a
product of which 80% passes P80 microns.
E = Wi (10/√P – 10/√F)
F = size at which 80% of the feed passes, µm
P = size at which 80% of the product passes, µm
The power required for the mill at the pinion (P, kW) is calculated as
followed:
P = E.T
T = new feed throughput, t/h
The energy requirement calculated with the Bond method work well for ball
mills charged with rod mill product and for diameter up to 4.9 m.
The Bond method also requires that the feed and product size distributions to
be approximately parallel lines when plotted cumulative percentage vs. Log
size.
Larger diameter ball mills require correction factor to the Bond equation.
The estimation of the energy requirement for AG and SAG mills needs a
larger set of correction factors and/or a pilot scale test.
The Abrasion Index quantifies the capacity of the ore to wear
equipment/tools when motion is applied.
Ai is a key ore parameter because it allows the estimation of the wearing
rates.
Correlations have been developed to estimate wearing rates in kg of metal
wearing/ kWh used in comminution circuit. They are:
Crusher liners (jaw, gyratory, cone): (Ai + 0.22)/ 24.3
Roll shell of the roll crusher: 0.45(0.1Ai)0.667
Balls of the wet ball mill: 0.159(Ai – 0.015)0.33
Liners of the wet ball mill: 0.0118(Ai – 0.015)0.3
The abrasion test is carried out in a rotating drum in which dry ore is placed
over a hardened steel paddle rotating concentrically but in opposite direction
for 1 h. Every 15 minutes, the ore is removed and replaced. The Ai is
determined from the weight loss of the paddle.
Each rock and mineral has their specific range of Ai.
The operational cost increases when processing an ore of bigger Ai.
Rock/Mineral Average, Ai
Granite 0.40 – 0.55
Basalt 0.20 - 0.45
Quartz 0.69 - 0.75
Feldspar 0.19
Clay 0.04
Limestone 0.001 - 0.05
Hematite 0.37 - 0.50
Magnetite 0.20 - 0.48
Apart of the Wi and Ai, additional testworks are sometimes required for a
complete characterisation of the ore and definition of the comminution
circuit. They are briefly described below.
Unconfined Compressive Strength (UCS): the max. compressive stress that
a sample can withstand before failure is measured. It is used for crusher
selection.
JK Tech Drop Weight Test: It assesses the impact breakage characteristics
of an ore five sized fractions over the range of 13 to 60mm. A steel weight
falls under gravity to crush a single particle. The test generates the rock
breakage parameters “A” and “b” Axb indicates the strength of the ore.
Higher value indicates lower strength. A and b are used in model
calculations for ball, SAG, AG mills and crushers.
As shown in this figure, bigger particles have lower strength, for soft and
hard ores.
SAG Mill Comminution Test (SMC): this is a shorter version of the JK
drop weight test, and only one size fraction is tested. It also generates the
“A” and “b” parameters for use in simulation, though with less accuracy.
Once the ore mineralogy and the comminution parameters have been
obtained, the crushing and milling circuits can be determined.
This involves the selection of the type of crushers and mills that have the
capacity to produce the size and degree of liberation required downstream
the process at the minimum operational cost (energy, balls consumption).
For example, for a medium competent and non-abrasive ore, and a
product size of P80 < 120 µm, it is suggested:
Primary Crusher + Single Stage SAG Mill
Three Stage Crushing + a Ball Mill
For a medium competent and abrasive ore, and a product size of P80 < 120 µm, it is suggested:
Primary Crusher + SAG-Ball Mills
Primary Crusher + AG Mill with Recycled Crusher
Very common to use the Bruno Simulation programme from Metso.
Data input required: Wi, Ai, SG, feed particle size distribution, feed
throughput
Crusher itself, classification and handling equipment are part of the
circuit.
The model performs:
The flowsheet for the whole circuit
Mass Balance of the circuit, including screen oversize recycles.
Product particle size distribution
Leach testwork (bottle roll) are required to obtain the following process
information:
The reagents consumption rates
The residence time
The gold recovery by cyanidation at different feed particle size,
different CN rate, oxygen or air sparging.
Conventional Leaching: typical process for an oxide ore.
Intense Leaching: typical process for a sulphide ore.
Gold Leaching (see figure).
METSIM, HSC-SIM, IDEAS, SYSCAD
Scoping Study
(conceptual)
Pre-Feasibility
Study
Defined Feasibility
Study
Detailed Engineering
Construction
DFS is also called bankable feasibility study. It is used to get loan for
construction.
Detailed Engineering involves the detailed design of each of the process,
mechanical, electrical, instrumentation and piping equipment/items.
It is possible to attract investors for a project only with the scoping study.
All of these studies are executed to reduce design risks of the process plant
technical failure and estimate the capital and operational costs (CAPEX,
OPEX).
SCOPING STUDY PFS DFS
AIM
Provide a general overview
and several alternatives for
processing
Evaluate alternatives and
chose one option for
processing or abandon the
project
Provide a comprehensive
technical evaluation to
justify project economic
viability and decrease
investment risk.
COST ESTIMATE
ACCURACY ± 50 % ± 30 % ± 10 - 15 %
STUDY PERIOD 2 to 5 months 9 to 12 months ~ 18 months
GEOLOGY
Completed wide spaced
drilling to define minimum
amount of resources and
limits of mineralisation.
Availability of an ore body
block model. Knowledge of
ore grade variability.
Completed close spaced
drilling. Availability of
detailed ore body block
model. Mineral resource
identified to indicated or
measured status.
SCOPING STUDY PFS DFS
PROCESS
FLOWSHEETS
One flow sheet for
each alternative
Flow sheets for the whole process plant
considering main equipment, services and
reagents.
Details to PFS flowsheets
are added and updated.
MASS BALANCE Basic mass balance
Develop of mass balance in Excel with
average flows to size main equipment.
Raw and process water balance.
Detailed mass balance of
flows & metals for a range
of throughputs with Excel
or a simulation program.
P&ID NA
Selection of main instruments and process
control equipment. Basic P&IDs for whole
process plant.
P&IDs are completed.
Development of the process
control philosophy
DESIGN
CRITERIA NA Design characteristics of main equipment
Detailed design of main
equipment, ore and water
characteristics.
PLANT LAYOUT
Basic concept for a
green-field.
Identification of
available area for
brown-field projects.
General layout of equipment,
tanks and buildings on scale. Completion of plant layout.
Parameter Units Value Source
Gold solids head grade - design Au g/t 0.60 Calc
Average Gold Lab Recovery
(at P80 = 9.3mm and P50 = 3.5mm)
Au % 60 Calc
Gold Recovery - Design Au % 60 Calc
Gold tail solids grade - Design Au g/t 0.24 Calc
Heap pad tonnes irrigated dry t / a 1,723,600 Calc
Bulk Density t/m3 1.6 Calc
Heap pad volume irrigated m3 10,772,700 Calc
Heap pad height m 10 Client
Heap pad area irrigated m2 1,077,260 Calc
No. of Leach cycles # 3 Client
Total Leach Time days 297 Calc
No. of Leach cycles # 3 Client
Total Leach Time days 297 Calc
Solution Specific Gravity t/m3 1.02 M/B
Leach Cycles days 98 Agreed
Solution application rate - Design L/m2/hr 10 Client
t solution / t
ore 1.28 Calc
Total Solution application rate - design m3 / a 65,269,900 Calc
t solution / t
ore 3.80 Calc
Water losses by evaporation % 15 Client
Water losses by retained moisture in cell m3/t 0.05 Assumed
Gold recovered - Design g / a 6,406,760 Calc
Total Solution flowing out of heap under
primary leach (incl. water losses retained in cell) m3 / h 2,380 Calc
Pregnant Solution Tenor - average Au g/m3 0.3 Calc
SCOPING STUDY PFS DFS
SAMPLING Use grab sample but not
essential
Use fresh drill-core samples or RC
chips from identified holes. Obtain
typical samples of mayor ore types.
Obtain extensive drill-core
samples from each ore type.
CRUSHING WORK INDEX NA A typical value is acceptable. Each ore type should be
measured.
ROD AND BALL WORK
INDEX
A typical value for the ore
type is acceptable. Use composite of ore types.
Each ore type should be
measured.
"A" AND 'b" NA A typical value is acceptable. Each ore type should be
measured.
ABRASIVITY INDEX, Ai NA A typical value is acceptable. Each ore type should be
measured.
RECOVERY Use typical value or
previous
data if available.
Establish range of recovery
Metal recovery should be
confirmed with test work to
acceptable level of accuracy.
RESIDUAL EFFLUENT NA Carry out tailing solution assays.
Carry out tailing solution assays
for all waste streams. Carry out
Detox test work.
SCOPING STUDY PFS DFS
MINING
METHODOLOGY
Various mining
techniques for either
open-cut or underground
mining are listed
One mining technique is
chosen. Ore grade and
dilutions are estimated. The
mine production capacity is
estimated.
The sequential schedule for
mining areas, with production
tonnage and grades, is
produced in detail.
MINE
PRODUCTION
The production is
estimated
based on experience of
similar properties.
The production schedule for
the year for tonnes and grades
is estimated. Production
equipment is chosen.
The tonnes and grades to be
mined by year are finalised.
Risk assessment is carried
out.
SCOPING STUDY PFS DFS
EQUIPMENT LIST ~ 50% completed ~ 80% completed ~ 90% completed
EQUIPMENT SIZE Type, size and capacity of
main equipment are stated.
Type, size and capacity of
most equipment are stated.
Type, size and capacity of
all equipment are stated.
DATA SHEETS NA Datasheets for main equipment
are developed
Datasheets for most
equipment
are developed
BUDGET PRICING
Budget pricing for some
mechanical
equipment is obtained from
vendors.
Budget pricing for main
mechanical
equipment is obtained from
vendors.
Budget pricing for most
mechanical
equipment is obtained from
vendors.
Surveys and studies required for the environmental scoping document.
Desktop fauna assessment and survey
Flora and vegetation survey
Air quality and green - house assessment
Acid mine drainage study
Surface hydrology
Groundwater modelling and assessment
Site water balance
Tailings design and management
The EIS needs to be evaluated and in Australia is performed by EPA (Environmental
Protection Authority. If the proposal is complex and high level of environmental
concern, the document is assessed through the PER procedure (public environmental
review). This process is outlined below (Western Australia Government).
The main options based on extraction methods are:
Gravity separation + Melting of Concentrate
Gravity separation + Amalgam. of Conc. (moving to Cyan.)
Gravity separation + Agitated leaching of concentrate
(intensive)
Heap Leaching
Agitated Leaching
Flotation + Regrind + Agitated leaching of Concentrate
(intensive)
Flotation + Oxidation + Agitated leaching (intensive)
The possibility of gravity separation as a pre-treatment before leaching
should be evaluated as it produces a less tonnage with a higher grade
fraction to be treated downstream.
Gravity separation has lower CAPEX and OPEX than flotation.
Gravity concentrates contains mainly coarse gold (>50µm to 2cm) and high
concentration of dense sulphide + oxide minerals.
The most used lixiviant for Au is still the cyanide ion (CN-). Other lixiviants
such as chloride, thiocyanate, thiosulphate and ammonia are not used at
commercial scales.
Intensive cyanidation improves the leaching kinetics of sulphides, by
increasing CN-, O2, and sometimes T and P.
Amalgamation is still commonly being used for the treatment of gravity
concentrates but rarely for ROM ores due to health and environmental
reasons. However, it is rapidly being superseded by intensive cyanide
leaching.
Oxidative pre-treatment techniques: roasting, pressure and biological
oxidations.
Source: The chemistry of gold extraction, Marsden and House.
PROCESSING METHOD SOUTH
AFRICA USA AUSTRALIA CHINA RUSSIA PERU
Heap Leaching 0 70.1 1.3 0.2 5.5 114.1
Vat Leaching 0 0 0.5 0 0 0
Agitated Cyanide Leaching 293.6 36.1 123.1 4.8 48.9 0
Agit.Cya. Leaching + Grav,Conc. 13.8 20.2 28.9 0 0 0
Gravity Concentration only 0 0 0 59 52.6 0
Flotation only 0 9.6 20.8 9.3 0 0
Gravity Concent. + Flotation 0 2.7 8.9 4.3 0 0
Flotation + Agit.Cyanide Leach. 1 0 0 22.8 4.7 0
Whole Ore Oxid.Pretreat. + CN
Leach. 0 107 0 0 0 0
Float. + Oxi.Pretreat. + CN
Leaching 3.3 7.5 39.6 3.0 3.1 0
Unaccounted for 31 8.8 35.3 113.9 66.8 59.1
TOTAL, t/y of gold 342.7 262 258.4 217.3 181.6 173.2
2004 gold world production. The “unaccounted for” segment refers to small,
old and remote located operations.
It was not quantified amalgamation processes, as not reliable data are
available.
34% of the Peruvian gold production comes from “unaccounted for”
segment; which probably means from small mines. It is unknown with the
info available if these small miners operate with Hg or CN.
66% of the Peruvian gold production is produced by heap leaching. This
production comes only from 2 mines: Yanacocha (94 t/y, first in the world,
Newmont) and Pierina (20.1 t/y, 20th in the world, Barrick).
53% of the Chinese gold production is generated by small miners.
27% of the gold from China is produced by gravity concentration followed
by smelting. 29% is for Russia.
Gold is present as fine particles accessible to solution at a coarse rock size
Gold ore grade = 1-2g/t
Gold recovery = 70%
Type of lixiviation = Heap Leaching of ROM ore
Drip irrigation flow = 5 to 10L/h/m2
Cyanide consumption = approx. 0.03 kg of NaCN/t (very low)
Lime consumption = approx. 0.8kg of CaO/t
Leach solution pH = 10.5 to 11
Pregnant solution cyanide concentration = 50mg/L of WAD (weak acid
dissociable)
Source: The chemistry of gold extraction, Marsden and House
Gold is present as ultrafine free gold particles (<0.1µm) disseminated within
a porous host rock. The leach occurs at relatively coarse crush size
Gold ore grade = 2.0-2.2g/t
Gold recovery = 80%
Type of lixiviation = Heap Leaching of crushed ore
Ore size to leach = P80 of 38mm
Cyanide consumption = approx. 0.22 kg of NaCN/t
Lime consumption = approx. 0.42kg of CaO/t
Leach solution pH = > 10
Pregnant solution gold concentration = 2 to 3g/t
The most common industrial method for cyanide detoxification applied the
principle of CN- oxidation to less toxic cyanide species. They are:
Sulphur dioxide and oxygen treatment
Hydrogen peroxide process
Alkaline chlorination process
Natural degradation also reduces the CN- concentration but at slow kinetics.
Some of the mechanisms responsible for this include volatilisation as HCN,
the adsorption of cyanide by minerals in the soil (e.g., 0.5 mg CN-/ g of
carbonaceous material), photo-decomposition and biological degradation.
Basically, cyanide is oxidised to cyanate (OCN-).
Free and weakly complexed metal cyanides (zinc, nickel and copper) are oxidized
to cyanate by sulphur dioxide (or Na2S2O5) and oxygen in the presence of soluble
copper as catalyst:
CN- + SO2 + O2 + H2O Cu+2 OCN- + SO4-2 + 2H+
M(CN)4-2 + 4SO2 + 4O2 + 4H2O Cu+2 4OCN- + 8H+ + 4SO4
-2 + M+2
The optimal pH range is from 6 to 10. The process performance decrease outside
this range.
Lime should be added to react with the acid generated.
The free metal is removed by the ferrous cyanide and hydroxide ions:
2M+2 + Fe(CN)6-4 M2Fe(CN)6 (solid)
M+2 + 2OH- M(OH)2 (solid)
Lime should be added to react with the acid generated.
The process treats slurry effluents containing > 200 mg/L total
CN.
Required 3 to 4 kg SO2/ kg of CN + 1 to 2 L of air/min/L
solution + Cu2+ at ~ 15% of initial WAD cyanide.
Discharge CN, Fe, Ni, Zn of < 1 mg/L is achieved after 1 – 2 h.
If treated high level of CN, the reagents costs could be high.
Also sulphate increases.
The free metal is removed by the ferrous cyanide and hydroxide ions:
Hydrogen peroxide oxidises free cyanides and Cu, Zn, Ni complex cyanides to
cyanate as follow:
CN- + H2O2 CNO- + H2O
2Cu(CN)32- + 7H2O2 + 2OH- 6CNO- + 2Cu(OH)2 + 6H2O
It has been demonstrated that solutions containing 500 mg/L WAD cyanide can
be reduced to < 2 mg/L in 1 – 2 h by adding 75 – 125 mg of H2O2/L solution.
Effluents of < 0.1 mg CN/L can be produced at higher H2O2 dosage.
Estimated hydrogen peroxide consumption: ~ 3 kg H2O2/ kg of CN.
Caro’s acid (H2SO5) is been used as a source of hydrogen peroxide because
reduce cyanide within minutes (< 50 mg CN/L).
H2SO5 + CN- H2SO4 + CNO-
Caro’s acid must be prepared in situ (hydrogen peroxide + sulphuric acid) and
use immediately to be effective.
Source: The chemistry of gold extraction, Marsden and House.
FEED,
mg/L
EFFLUENT,
mg/L
USEPA DRINKING
WATER,
mg/L
Cyanide 280 3 0.20
Arsenic 0.2 < 0.05 0.01
Copper 4.5 < 1.00 1.30
Iron 16.0 < 0.015 0.30
Zinc 157.0 < 1.0 5.0
Selenium 5.0 4.0 0.05
The activate range is the hypochlorite ion (OCl-), that is formed when Cl2 gas is
disolved in water:
Free cyanide reacts rapidly with hypochlorite to form cyanogens chlorite that is
hydrolised to cyanate:
CN- + H2O + ClO- = CNCl(g) + 2OH-
CNCl + 2OH- = CNO- + Cl- + H2O
Free cyanide reacts rapidly with hypochlorite to form cyanogens chlorite, which
is hydrolized to cyanate and chlorite ions.
Effective oxidation can be achieved in 10 to 15 min at pH 10 to 11.
Require consumptions of approximately 8 to 24 kg of Cl2/ kg CN.
The reduction of toxic cyanide concentrations to 0.1 mg/L WAD cyanide can
usually be achieved at the residual of Cl2 concentrations between 10 -15 mg/L;
range that is very toxic to fish.
10 ppm of HCN is the maximum level for a daily exposure of 8 h (USAEPA).
Though, the Occupational and Health Department said that less than 5 of HCN/
m3 of air (4.7 ppm) should not be exceeded at any part of the working
environment as this gas is very toxic chemical.
HCN is formed and evolved from the leaching solutions in a reaction that is
very fast
HCN(aq) = H+(aq) + CN-
(aq) Log K(25 C) = - 9.22
The HCN molecules in solution can readily volatilise to the atmosphere due to
the high vapour pressure, even from stagnant solutions. The volatilasation
increases with temperature and agitation degree.
The amount of HCN changes with pH and salinity of solution.
The concentration of HCN decreases with increasing pH; and the equilibrium
shifts completely to the left and only cyanide ion exists above pH 11.
In gold process operations, the typical pH in the leach circuits is between 9.5
and 11.5 and between 7 and 9 in tailing ponds.
Greater loss of gaseous HCN is obtained in strongly saline solutions at the same
pH. At the pH of 9.5, about 25% of the total cyanide is as HCN at an ionic
strength less than 1M whilst this increases to about 85% at an ionic strength of
5M.
The control of HCN volatilization is by the operational conditions other than
design considerations. Try to keep the pH in the CIL tanks close to 10 is
recommended.
Significant amount of dust is generated when ore have high content of < 0.5 mm
particles; and specially at transfer points.
The installation of sprays for dust control at the stockpiles and bins is
common/essential. Recycled water will be advantageous to use but possible
salts precipitation should be evaluated.
Enclosure of the transferred equipment is another design alternative.
Baghouse Centrifugal Collector