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