the coal handbook: towards cleaner production || surface chemistry fundamentals in fine coal...

© Woodhead Publishing Limited, 2013

347

12 Surface chemistry fundamentals

in fine coal processing

J. S. LASKOWSKI, University of British Columbia, Canada

DOI : 10.1533/9780857097309.2.347

Abstract : It is argued that the wettability which is fundamental for fl otation also determines the properties of fi ne coal aqueous suspensions and thus controls not only fl otation but also fl otation products dewatering and handling either as dry products or as suspensions (e.g. coal-water slurries). Typical fi ne particle technology problems appear also in gravity separation methods in which fi ne magnetite aqueous suspensions are used as a medium. In this chapter an attempt is made to look at these various unit operations in some unifi ed way based on the fact the main aspects of these unit operations result from the fundamental fact that all these are aqueous suspensions of fi ne particles characterized by the same rheological phenomena.

Key words : fi ne coal, coal fl otation, settling, fi ltration, fl occulation, oil agglomeration, pelletization, fi ne coal handleability, coal-water slurries, magnetite dense media, rheology.

12.1 Surface properties of coal

Coal is an organic sedimentary rock whose composition changes with coal-

ifi cation. Since metamorphic development of coal, also referred to as coal-

ifi cation, is synonymous in chemical terms with progressive enrichment of

the coal substance in organically bound carbon, all coals, regardless of their

origin or type, can be arranged in an ascending order of carbon content

(Fig. 12.1). As this fi gure shows, coal is a highly cross-linked polymer con-

sisting of a number of stable fragments connected by relatively weak cross-

links. Coal also contains heteroatoms, such as oxygen (which appears in coal

in the form of phenolic, etheric, and carboxylic groups), nitrogen, and sulfur,

and their presence in coal structure strongly affects coal surface properties.

Coal surface properties, like the properties of any other solid, can be stud-

ied via wettability measurements. This involves measurement of contact

angle ( Θ ) with the use of liquid with known surface tension ( γ L ).

The work of adhesion of liquid to solid ( W SL ) is given by

W W WSLWW SWW LLW

SLWW AB( ) +WSWW LLWγ L (1 [12.1]

348 The coal handbook


where WSLWW LW and WSLWW AB stand for the Lifshits–van der Waals contribution to

the work of adhesion and the acid-base interactions energy contribution,

respectively (please note that in older publications the term WSLWW d , the disper-

sion forces’ contribution, was used instead of WSLWW LW as is common today).

In order to evaluate the dispersion forces’ contribution to the wettabil-

ity of coals, Gutierrez-Rodriquez et al . (1984) used methylene iodide and

showed that the values of the contact angle measured with this compound

do not depend on coal rank, or on its oxidation. These contact angle values

for various coals were in the range of 28 ° ± 9 ° irrespective of the experimen-

tal technique (captive-bubble or sessile-drop).

As shown by Fowkes (1964)

W WSLWW dSLWW LW d

Ld=WSLWW LW 2 γ γs

ddLdd [12.2]

For water γ Lγγ dγγ ≈ 22 mJ/m 2 . Methylene iodide, as saturated hydrocarbons, is

a useful reference liquid because its intermolecular attraction is entirely due

to London dispersion forces. For methylene iodide γ Lγγ dγγ = γ L = 50.8 mJ/m 2 , and

for methylene iodide wetting coal surface, one can obtain:

W WSLWW SLWW d dLd

L=WSLWW d 2 dLd = Lγ γss

ddLdd γ L ( c+1 os )Θ [12.3]

Coal rank%Cdaf

Peat

Lignite60

70

87

91

Sub-bituminous

High-volatile bituminous

Medium-volatile bituminous

Low-volatile bituminous

Semi-anthracite

Anthracite

Graphite

12.1 Variation in coal structure and carbon content with coal rank.

Surface chemistry fundamentals in fi ne coal processing 349


This gives for coal γ sγ dγγ ≈ 44 mJ/m 2 .

For coal interacting with water, if it is assumed that coal is a homogenous

hydrocarbon matrix that is unoxidized, is mineral matter free, and interacts

with water only via dispersion forces:

W WSLWW SLWW d =WSLWW d ( )γ θL ( ))1+ [12.4]

and thus

cosθγ

γ γγ

= − +1 1+ = −2W γγ dγγ Lγ dγγSLWW d

L Lγγ γγ [12.5]

Putting for water γ Lγγ = 72 mJ/m 2 one can derive the contact angle on such a

coal surface would have been about 98 ° . Any smooth coal surface having a

water contact angle of less than 98 ° contains, therefore, various hydrophilic

areas (polar functional groups, inorganic impurities, etc.) on the hydropho-

bic hydrocarbon matrix (Laskowski, 1994, 2001).

12.1.1 Effect of coal rank on wettability

In the 1940s, Brady and Gauger (1940) observed that the contact angle

values measured on Pennsylvania bituminous coals were larger than on

anthracite, while North Dakota lignites were very hydrophilic. The results

of comprehensive wettability studies on coal from the Donbass Basin

(Ukraine) were published by Elyashevich (1941), while further details were

provided by Horsley and Smith (1951) in the 1950s. The analysis of the wet-

tability of coals as a function of coal rank was offered by Klassen in his

coal fl otation monograph (Klassen, 1963) in which he used Elyaschevich’s

data. This relationship is shown in Fig. 12.2 using more recent data of coal

analysis for oxygen content of Gutierrez-Rodriquez et al . (1984) and Bloom

et al . (1957). As Fig. 12.2 shows, low-rank coals that possess a lot of oxygen

are quite hydrophilic, while low-volatile matter bituminous coals are the

most hydrophobic of all. Comparison of the contact angle values shown in

Fig. 12.2 with the calculated value for pure coal organic matrix (about 98 ° )

indicates that while the contact angles measured on bituminous coals are

not that different from this calculated value, the difference increases with

decreasing coal rank. This is an obvious effect of increasing oxygen content

in coal with decreasing rank (also shown in Fig. 12.2). The contact angle

measured on bituminous coal is smaller than the calculated values for pure

coal organic matrix because coal always contains some hydrophilic inor-

ganic matter (ash).



Coal is a very heterogeneous solid. Figure 12.3 is a schematic represen-

tation of coal surface. Coal can be depicted as a hydrocarbon matrix that

contains various functional groups (Fuersteau et al ., 1982). The composi-

tion of the matrix varies with the coalifi cation (Fig. 12.1). Coal also contains

mineral matter and is porous. As Fig. 12.1 shows, with increasing coalifi ca-

tion degree hydrocarbons building coal become more aromatic. Rosenbaum

and Fuerstenau (1984) assumed that coal may be modeled as composite

material, the non-wettable portions of which are made up of paraffi ns and

aromatic hydrocarbons, and whose wettable portions are represented by

functional groups and mineral matter. To calculate the contact angle on such

a composite surface they used the Cassie-Baxter equation and assumed that

65 70 75 80 85

Carbon (%, daf)

90

Oxygen

Water-captivebubble

Water-sessiledrops

95 10000

10

20

30

40

50

60

70

80

5

10

Oxy

gen

(%)

Con

tact

ang

le (

°)

15

20

25

30

12.2 Relationship between coal rank and wettability by water measured

by the captive-bubble and sessile-drop methods (Source: After

Gutierrez-Rodriquez et al ., 1984 with permission of Elsevier), and the

relationship between coal rank and the total oxygen content (Source:

After Bloom et al ., 1957 with permission of Elsevier).

Coal

Min

eral

mat

ter

Coo

h

OH

Por

es

12.3 Schematic representation of coal surface.



the maximum values for contact angles on paraffi nic hydrocarbons can be

as high as 110 ° , while those for aromatic hydrocarbons are only 85 ° . This

explains why the wettability of very aromatic anthracites is lower than that

of bituminous coals. This concept was further developed in the patchwork

assembly model by Keller (1987).

Such an analysis must also include coal porosity. For example, Horsley

and Smith (1951) observed that some petrographic constituents (e.g. fusain),

lose good natural fl oatability after prolonged immersion in water. More

recent results (He and Laskowski, 1992) entirely prove the effect of porosity.

However, while on less hydrophobic surfaces water is sucked into capillaries

by capillary forces and this makes such a coal even more hydrophilic, the

capillaries on the surface of a hydrophobic coal will stay fi lled with air and

this will make such a surface more hydrophobic.

12.2 Coal flotation

Coal fl otation is the only fi ne coal cleaning process that is effective in treat-

ing − 0.15 mm size coal. Because of coal high natural hydrophobicity it may

appear to be easy to fl oat, but the wide range of surface properties of coals

from various ranks, and various degrees of liberation of the treated particles

make the process very often diffi cult. Flotation of low rank/oxidized coals and

desulfurizing fl otation are still challenging problems awaiting for solution.

12.2.1 Effect of rank on fl otability

In accordance with what has been said, coal fl oatability should strongly

depend on rank as has been extensively discussed (Laskowski, 2001). In

1951, Horsely and Smith concluded that in order to obtain equal recoveries

a larger quantity of reagents were required for anthracites and lignites than

for bituminous coals. In practice, this requires the use of different combina-

tions of reagents in fl oating different coals. Xu and Aplan (1993) demon-

strated it in a very simple way. Figure 12.4 shows that while MIBC alone is

suffi cient to fl oat the very hydrophobic bituminous coals, a combination of

MIBC (frother) and an oil (collector) is needed to fl oat lower-rank coals.

Aplan noted a semi-logarithmic relationship between the fuel oil consump-

tion and the carbon content in coal.

As has already been pointed out, coal is heterogeneous and it contains

organic matter and mineral matter. The former appears in the form of mac-

erals, and the latter as minerals. Macerals are classifi ed into three groups:

vitrinite, exinite (liptinite) and inertinite. The vitrinite group comprises the

most abundant macerals in coal. Macerals do not appear in isolation, but

occur in associations in various proportions and with variable amounts of

mineral matter to give rise to the characteristic banded or layered character



of most coals. These associations are referred to as lithotypes and can be

distinguished macroscopically. The lithotypes include vitrain (bright bands

in coal), clarain (bright, lustrous constituent, which in contrast to vitrain has

dull intercalations), durain (dull) and fusain (black or gray in color with

fi brous structure similar to that of charcoal).

Since macerals have different chemical compositions, their surface and

fl otation properties also vary. As Fig. 12.5 taken from Klassen’s monograph

(Klassen, 1963) shows, coal particles varying in size and petrographic com-

position behave differently in the process. Fine bright particles dominate in

the fi rst products and only with time coarse particles and dull constituents

start fl oating. Large particles, including particles that are not liberated, fl oat

only when the fi ne particles are removed from the cell. These data correlate

very well with Horsley and Smith’s observations (Horsley and Smith, 1951),

which indicate that bright petrographic components (vitrain) are more

hydrophobic and fl oat better than dull components (durain).

Arnold and Aplan (1989) claim that hydrophobicity of coal macerals fol-

lows the pattern:

exinite>vitrinite>inertinite.

70 80 90 100

Carbon (%)

0.01

0.1

1

10

100

Min

imum

am

ount

of f

roth

er p

lus

colle

ctor

for

optim

umre

cove

ry (

kg/t)

MIBC only

MIBC and oil

MIBC and naphthenic acid

12.4 Minimum amount of frother and collector for optimum recovery

of coals of various carbon contents. (Source: After Aplan, 1993.)



The conclusions regarding the behavior of coal macerals in fl otation are

further complicated by mineral matter content. The effect of petrographic

composition of coal particles on their fl otation properties can be studied

only for fresh (unoxidized) and low ash samples (Holuszko and Laskowski,

1995). For samples containing more than 15% ash, the surface properties are

predominantly determined by mineral matter.

12.2.2 Flotation reagents

The behavior of coal in the fl otation process is determined not only by a

coal’s natural fl oatability (hydrophobicity), but also by the acquired fl oat-

ability resulting from the use of fl otation reagents. The general classifi cation

of the reagents for coal fl otation is shown in Table 12.1 (Laskowski, 2001).

The use of liquid hydrocarbons (‘oils’) as collectors in fl otation of coal is

characteristic for the group of inherently hydrophobic minerals (graphite,

sulfur, molybdenite, talc, coals are classifi ed in this group). Since oily collec-

tors are water-insoluble, they must be dispersed in water to form an emul-

sion. The feature making emulsion fl otation different from conventional

fl otation is the presence of a collector in the form of oil droplets, which must

2 4 6 8 10 12 14 16 18 20

Flotation time (min)

0

20

40

60

80

100

Yie

ld (

%)

4

2

1

3

12.5 Effect of petrographic composition and particle size on coal

fl otation kinetics. (1) bright coal; (2) dull coal; (3) shale interlocked with

dull constituents; (4) gangue. (Source: After Klassen, 1963.)


Tab

le 1

2.1

C

oal

fl o

tati

on

reag

en

ts

Typ

e

Flo

tati

on

use a

s

Fu

ncti

on

al

gro

up

E

xa

mp

les

Acti

on

No

np

ola

r

(Wate

r-in

so

lub

le)

Co

llecto

rs

_

Ke

rose

ne

Fu

el

oil

Se

lecti

ve

we

ttin

g a

nd

ad

he

sio

n

of

oil

dro

ps t

o c

oa

l p

art

icle

s

Su

rface a

cti

ve

(Wate

r so

lub

le)

Fro

thers

H

yd

rox

yl

Nit

rog

en

ou

s

Ali

ph

ati

c a

lco

ho

ls

Po

lyg

lyco

ls

Fro

the

rs w

ith

so

me

co

lle

cti

ng

ab

ilit

ies. A

lso

im

pro

ve

em

uls

ifi c

ati

on

of

oily

co

lle

cto

rs

Em

uls

ifi e

rs

(So

lub

le in

oil

y

co

llecto

r)

Pro

mo

ters

H

yd

rox

yl

Ca

rbo

xy

l

Nit

rog

en

ou

s

Po

lye

tho

xy

late

d

alc

oh

ols

, fa

tty

acid

s,

etc

.

Fa

cil

ita

te c

oll

ecto

r

em

uls

ifi c

ati

on

an

d

sp

rea

din

g o

ve

r co

al

Ino

rgan

ic

(Wate

r so

lub

le s

alt

s)

Mo

difi

ers

_

Na

Cl,

Na

2 SO

4

H 2 S

O 4 ,

Ca

(OH

) 2

Ca

(OH

) 2

Pro

mo

ters

pH

re

gu

lato

rs

Su

lfi d

e d

ep

ressa

nts

Pro

tecti

ve

Co

llo

ids

Dep

ressan

ts

Hy

dro

xy

l

Ca

rbo

xy

l

Po

lym

ers

: sta

rch

,

de

xtr

in,

ca

rbo

xy

me

thy

l

ce

llu

lose

, e

tc.

Mo

difi

ers

,

Co

al

de

pre

ssa

nts



collide with mineral particles in order to enhance the probability of particle-

to-bubble attachment. The process is based on selective wetting: the drop-

lets of oil can adhere only to particles that are to some extent hydrophobic.

The effect of emulsifi cation on fl otation has been studied, and its benefi cial

effect on fl otation is known (Sun et al ., 1955).

Coal fl otation is commonly carried out with a combination of an oily col-

lector (e.g. fuel oil) and a frother (e.g. MIBC). All coal fl otation systems

require the addition of a frother to generate small bubbles and to create

a stable froth (Table 12.2). Typical addition rates for frothers are in the

order of 0.05–0.3 kg of reagent per tonne of coal feed. Depending on the

hydrophobic character of the coal particles, an oily collector such as diesel

oil or kerosene may or may not be utilized. When required, dosage rates

commonly fall in the range of 0.2–1.0 kg of reagent per tonne of coal feed,

although dosage levels up to 2 kg/t or more have been known to be used for

some oxidized coals that are diffi cult to fl otate.

The benefi cial effect of a frother on fl otation with an oily collector was

demonstrated and explained by Melik-Gaykazian et al . (1967). Frother

adsorbs at the oil/water interface, lowers the oil/water interfacial tension

and hence improves emulsifi cation. However, frother also adsorbs at the

coal/water interface (Frangiskos et al ., 1960; Fuerstenau and Pradip, 1982;

Miller et al ., 1983) and provides anchorage for the oil droplets to the coal

surface. Chander et al . (1994), after studying various non-ionic surfactants,

concluded that the fl otation of coal can be improved in their presence

because of the increased number of droplets, which leads to an increase in

the number of droplet-to-coal particle collisions. While the use of oily col-

lectors and frothers is the most common, also a group of fl otation agents

known as promoters have found application in coal fl otation. In gen-

eral, these are strongly surface-active compounds and are mostly used to

enhance further emulsifi cation of water-insoluble oily collectors in water.

Because of environmental concerns associated with tailing ponds, the

method for disposing of fi ne refuse from coal preparation plants by under-

ground injection has been gaining wide acceptance. Unfortunately, many

common fl otation reagents, including diesel oil, are not permitted when fi ne

refuse is injected underground into old mine works. This is the main driv-

ing force for fi nding replacement for the crude-oil based fl otation collectors

(Skiles, 2003). An alternative to fuel oil may be biodiesel, a product created

by the esterifi cation of free fatty acids generally from soy oil, with an alcohol

such as methanol, and subsequent transesterifi cation of remaining trigly-

cerides. Water, glycerol and other undesirable by-products are removed, to

produce a product that has physical characteristics similar to diesel oil. The

use of some vegetable oils was demonstrated to provide equivalent (and

even superior) fl otation results when compared with diesel fuel (Skiles,

2003). These are the results of commercial scale tests on a circuit that has



4.25 m in diameter columns. The product concentrate ash was 13.5%. The

consumption of the tested vegetable oil was about two times lower from the

consumption of diesel oil in these tests.

12.2.3 Flotation of low rank coals

The subject of fl otation of low-rank coals was tackled by Wheeler (1994) in his

interesting paper on the effect of frothers on coal fl otation. Frothers have by

far the largest effect on coal recovery and they do not only act in their tradi-

tionally accepted function of ‘bubble makers.’ In his tests he used anthracite,

medium-volatile bituminous, high-volatile B bituminous, and subbituminous

A coals, and different frothers in combination with fuel oil. For easy-to-fl oat

medium-volatile bituminous coal the aliphatic alcohols such as MIBC were

found to be excellent. Going down in the order of natural fl oatability, medium-

Table 12.2 Frothers utilized in coal fl otation

Name Formula Solubility

in H 2 O

(1) Aliphatic alcohols

Methyl isobutyl carbinol

(MIBC)

2-ethyl hexanol

2,2,4-trimethyl-

pentanediol

1,3-monoisobutyrate

(TEXANOL)

R − OH

C H3 CH CH2 CH CH3

CH3 OH

CH3 CH2 CH2 CH2 CH CH2 OH

CH2 CH3

CH3 CH3 O CH3

CH3 CH CH C CH2 O C CH

OH CH3 CH3

Low

Low

Insoluble

(2) Polyglycol-type frothers

DF 250

DF1012

Aerofroth 65

DF 400

DF-1263

CH 3 (PO) 4 OH

CH 3 (PO) 6.3 OH

H(PO) 6.5 OH

CH 3 (PO) 4 (BO)OH

Total

32%

Total

Very good

PO stands for propylene oxide (-CH 2 -CH 2 -CH 2 -O-), and BO for butylene oxide

(-CH 2 -CH 2 -CH 2 -CH 2 -O-) Cresylic acids (mixture of cresols and xylenols) that in the

past were commonly used in coal fl otation are not in use any more because of

their toxicity.



volatile bituminous > anthracite > high-volatile B bituminous > subbitumi-

nous A. MIBC quickly loses its effectiveness, fi rst to 2-ethyl hexanol, then to

texanol and glycol frothers (e.g. DF-1012). On the bvBb coal 2-EH fl oated

15% more coal than MIBC. Wheeler’s results confi rm that, while short chain

aliphatic alcohols possess only frothing properties, other frothers also exhibit

collecting properties, and the properties of oil emulsifi ers. So, the fi rst conclu-

sion is that in the fl otation of lower-rank/oxidized coals, one single reagent is

not suffi cient. A combination of properly selected frother and an oil, and good

emulsifi cation, in most cases leads to a satisfactory fl otation. The use of the

specifi cally selected promoters may also be helpful.

The use of properly selected reagents under the best possible conditions

is especially important when coals are diffi cult to process. An emulsifi cation

of the reagents and their stage addition are particularly useful in such cases.

An obvious practical solution is shown in Fig. 12.6.

12.2.4 Desulfurizing fl otation

The need to reduce the sulfur content of coal to low levels is one of the

more pressing needs facing coal preparation engineers if they are to

actively assist power-generating stations in reducing their overall sulfur

emissions. Yet, as of now, effective pyrite depression for a wide variety of

coals remains out of reach.

FrotherOilWater

1/3

1/3 1/3

1/3

Emulsification

2/3

Feed

12.6 A general concept showing a coal fl otation circuit with

emulsifi cation of oily collector and stage addition of reagents. (Source:

After Laskowski, 2001 with permission of Elsevier.)



Coal, a sedimentary organic rock, contains organic matter (macerals) and

inorganic matter (minerals). Coal preparation upgrades raw coal by reduc-

ing its content of mineral matter; the particles with lower ash content are

separated from those with higher ash content. The most common minerals

that occur in coal are clay minerals, carbonates (e.g. dolomite, calcite, sider-

ite), oxides (e.g. quartz) and sulfi des (e.g. pyrite). The last one is especially

important.

Convincing pieces of evidence indicate that most of the mineral matter

in coal down to micron grain sizes is indeed a distinct separable phase that

can be liberated by fi ne crushing and grinding. Keller (1984) measured the

particle-size distribution of mineral matter obtained by low-temperature

ashing of a few samples of coal from the Pittsburgh seam. Most mineral

grains were in the size range from 1 to 10 μ m, while the rest were coarser

than 100 μ m. Cleaning these coals by the Otisca-T oil agglomeration pro-

cess demonstrated that the ash content can be reduced to 1% ash by grind-

ing coal down to a few microns. This indicates that the inherent ash content

may be as low as about 1% if coal is fi nely ground to obtain proper libera-

tion, and then concentrated using a highly selective method.

Sulfur is the constituent of coal that most affects coal marketing. Three

types of sulfur in coals can be distinguished by chemical analysis: sul-

fi des (pyrite and marcasite), sulfates (mostly gypsum) and organic sul-

fur. Organic sulfur in coal appears in different organic compounds, such

as thiophenes, sulfi des (aliphatic R-S-R and aromatic φ -S- φ and thiols

(R-SH and φ -SH). The majority of the organic sulfur in high rank coals is

thiophenic (Attar. 1979; Attar and Dupuis, 1981). Typical sulfur analyses

of coals from different regions throughout the world set out by Mayers

(1977) varied from 0.38% to a high of 5.32%. The pyritic sulfur content of

these selected coals varied from a low of 0.09% to a high 3.97%, while the

organic sulfur content varied from a low of 0.29% to a high of 2.04%. In

general, organic sulfur levels greater than 2% or much less than 0.3% are

almost never encountered, and pyritic sulfur contents greater than 4% are

also uncommon.

The distinction between inorganic and organic sulfur is of great impor-

tance. Inorganic sulfur content in coals can be reduced by physical sepa-

ration methods. In general, coal and pyrite can be separated either by

depressing pyrite and fl oating coal, or by depressing coal and fl oating pyrite.

But since pyrite density is over 5 g/cm 3 and coal density is in the range

from 1.3 to 1.5 g/cm 3 , the rejection of gangue (and thus also pyrite) can be

improved by better circuitry and machinery. Gravity separation methods

are quite effi cient in separating coal particles from pyrite particles. Due to

the very high density of pyrite, even very small amounts of pyrite are suf-

fi cient to increase the density to a point where the coal particles can be



rejected. Therefore, particles containing small amounts of pyrite are more

easily rejected by density processes than by surface-based processes such as

fl otation. However, the effi ciency of separation by gravity methods falls rap-

idly for particles fi ner than 100 μ m. Therefore, coal desulfurization depends

critically on pyrite grain size, and hence on the dissemination of pyrite in

coal, since only pyrite that is liberated can be separated from coal. Statistical

analyses performed by Zitterbart et al . (1985) revealed that the percentage

of liberated pyrite is inversely correlated to the mean particle size for sev-

eral seams. For example, for the Pittsburgh seam coal about 55% of pyrite

was liberated after grinding coal down to 50 μ m.

Since coal is not ground down to liberation sizes before fl otation (as it is

done in the case of ores), the fl otation feed is a mixture of poorly and well

liberated particles. A combination of gravity separators and fl otation cells

in the same fi ne coal processing circuit is therefore essential; such combined

fl owsheets are very characteristic for coal fl otation.

In practice, the maximum particle size for coal fl otation is generally 28

mesh (0.6 mm) for highly fl oatable coals. The conditions required to recover

coarse particles (i.e., high aeration rates and high reagent dosages) must often

be avoided since they also favor the fl otation of impurities. Consequently,

most coal fl otation is applied to minus 100 mesh (0.15 mm) particles as more

cost-effective spiral concentrators or multiple stages of water-only cyclones

can be used to upgrade the plus 100 mesh (0.15 mm) fractions. In such cir-

cuits, the fractions high in sulfur that contain pyrite are rejected with high

density reject.

Pyrite almost always appears in polymetallic sulfi de ores. Since it is

treated as a gangue, which has to be depressed in the differential fl otation

of sulfi des, its fl oatability has been extensively studied. In the fl otation of

sulfi de ores with thio-collectors (e.g. xanthate), it is common to depress iron

sulfi des (pyrite/marcasite/pyrrhotite) by carrying out fl otation in an alkaline

environment using lime. However, similar conditions do not depress pyrite

in the fl otation of coal (Fig. 12.7). The collector–frother system in coal fl o-

tation is dictated by coal rank and fl oatability, and the selected reagents for

best coal fl otation apparently also promote pyrite fl otation. It is known that

some sulfi des display so-called self-induced fl otation in moderately oxidiz-

ing conditions. It is of interest to point out that Baker et al . (1990) observed

the synergistic action of coal and oxygen in coal pulps. This mechanism,

as well as incomplete liberation from organic matter, may be responsible

for the different behaviors of coal-pyrite and ore-pyrite. It is known that

coal-pyrite fl oats well in the presence of hydrocarbon collectors (Olson and

Aplan, 1984).

Since the fi rst option, namely fl otation of coal and depression of pyrite

has not been very successful, the second option, that is the reverse fl otation



in which coal is depressed and pyrite is fl oated, was also investigated. In the

‘two-stage reverse fl otation process,’ developed by the US Department of

Energy (Miller and Deurboruck, 1982), after the fi rst conventional fl ota-

tion stage, the froth product that contains coal but also pyrite particles is

repulped with dextrin and the pyrite is fl oated off with xanthate in acidic pH

from depressed coal. Although the pilot plant tests were very encouraging,

the process has not been commercialized.

12.3 Solid–liquid separation

Unit operations in the coal preparation plant can broadly be divided into

four distinct groups: classifi cation (that is mostly screening), benefi ciation,

dewatering of the benefi ciation products, and water clarifi cation (Fig. 12.8).

As Figs 12.8 and 12.9 demonstrate, with the exception of screening (and

some crushing), all other unit operations are carried out in water. Therefore,

what the coal prep plants are dealing with are aqueous suspensions. As the

term ‘fi ne’ indicates, in the fi ne coal cleaning circuit the solid particles are

fi ne. These fi ne suspensions, whether fl otation froth products or fl otation tail-

ings, are subjected to the solid–liquid separation since clarifi ed water must

be recycled back to the process. The properties of these suspensions, and our

ability to control them, determine the outcome of such unit operations.

Coal-pyrite

Coal

Ore-pyrite

20

20

40

Flo

tatio

n re

cove

ry (

%)

60

80

100

4 6 8

pH

10 12

Fuel oil: 50 mg/tMIBC: 50 mg/t

12.7 Comparison of fl otation recovery of fi ne coal, coal-pyrite, and

ore-pyrite as a function of pH using fuel oil and MIBC. Coal – 74 μ m,

coal- and ore-pyrite – 45 μ m. (Source: After Jiang et al ., 1993 with

permission of Elsevier.)



12.3.1 Stability of mineral suspensions

The solid-in-liquid dispersed systems are commonly classifi ed as colloids

or suspensions. In the former, the particle size is below 1 μ m; in the lat-

ter case, these particles are larger than 1 μ m. Brownian motion of particles

suspended in a liquid, characteristic for colloidal systems, ceases when par-

ticles are coarser than 1 μ m. However, irrespective of their size, particles

ROM coal

Recycled water

Waterclarification

Products

Dewatering

Beneficiation

Classification(Screening)

12.8 Water circuit of coal preparation plant.

Mining

Fine storage

Grinding

Slurry transportation

Classification

Concentration

Thickening

Drying

0 10 20 30 40 50

Water, volume %

60 70 80 90 100

Filtration

Screening

Crushing

Coarse storage

12.9 Variation in water content during various stages of processing

(Holland and Apostolides, 1969).



suspended in a liquid frequently collide, not only because of their Brownian

movement, but also because these particles, depending on their size, settle

at different rates. As a result of such particle–particle collisions, the stability

of the system is determined by the interaction between the particles during

these encounters. The system is considered to be stable when the particles

are dispersed and settle as individual units; the system is unstable, and coag-

ulates, when the particles form aggregates and settle quickly.

In the case of colloidal systems, detection of the coagulation is fairly sim-

ple. In the case of mineral suspensions, the size of the particles is such that

they are not subjected to the Brownian motion and thus the solid particles

settle due to gravity (at the rate given by Stokes’ equation). The state of

aggregation of such a slurry can be judged by visual inspection of samples

left to stand in tall glass cylinders. The typical behavior of A dispersed (sta-

ble) or B coagulating (or fl occulating) polydisperse and multicomponent

slurries is shown schematically in Fig. 12.10.

12.3.2 DLVO theory

In the DLVO theory (developed independently by Deryaguin, Landau,

Verwey and Overbeek) the energy of interaction between solid particles

Micrographs

B

1 2 3

A

(a)

(b)

12.10 Visual appearance of mineral suspensions (schematic). (a)

stable; (b) coagulating. (1) initial; (2) short time; (3) long time. Circles:

appearance of samples in the optical microscope (Kitchener, 1978).



is estimated as a sum of the London-van der Waals attraction, and electro-

static repulsion (when the interacting particles have the same either neg-

ative or positive electrical charge) resulting from overlap of the electrical

double layers surrounding the interacting particles (Fig. 12.11).

In the classical DLVO theory, the total energy of the interaction between

two particles is given by

V = V R − V A [12.6]

where V R is the energy of electrostatic double layer repulsion (positive

value means repulsion) and V A is the van der Waals attraction (negative

value means attraction).

In practical situations, the electrokinetic potential (known as zeta poten-

tial) is measured to characterize the electrical charge of solid particles.

Figure 12.12 shows the calculated value of the total interaction energy ( V )

for the system at two different values of the zeta potential.

The repulsion between two particles depends on the electrical charge

of these particles. Figure 12.12 shows repulsion curves at two different val-

ues of zeta potential of the interacting particles. Since zeta potential can

be changed, for instance by changing pH, and the van der Waals attraction

forces, which depend on the chemical composition of the particles, are not

changed when pH is altered, and the two total interaction curves for this

system are different. In Case 1, the zeta potential is large and the repulsion

curve V R (1) is very positive, giving large positive values of the total inter-

action curve (the maximum on the total interaction curves is referred to as

energy barrier). If the kinetic energy of the interacting particles is not large

enough, the particles will not be able to overcome the energy barrier and

the particles will not attach to each other. The system will then be stable. In

Case 2, the energy barrier does not exist; each collision between the par-

ticles leads to attachment, and the system will coagulate. Hence, coagulation

+++

++ +

++

+++++

++

++

++

++++

+

+ +

+ ++

––

––– – – –

–– – – –

–––

–

–––––

––

––

+

+

+

++

+++

++

++

++

+

+

++

+ + + + ++ +

+ +

+++

++

+++

+

+

+

+++

+

+++

12.11 Schematic picture of the two identically charged solid particles

which double layers overlap when they approach each other (Gregory,

2006).



is the process in which the particles aggregate when the electrical repulsion

between the particles is lower than the energy of attraction.

For particles which are hydrophobic, the attraction energy also includes

the hydrophobic force, which is larger than the van der Waals attraction. As

shown by Xu and Yoon (1989, 1990), the coagulation curve in such a case

looks like that shown in Fig. 12.13 (top part). The experimental curves were

obtained for a bituminous coal. Since with increasing hydrophobicity the

additional hydrophobic attractive force becomes larger, which together with

the van der Waals force is balanced by the electrostatic repulsion dependent

on pH, the pH range of coagulation increases with increasing hydrophobic-

ity of the interacting particles.

As Fig. 12.14 shows, the iso-electric points of coals (the pH at which the

zeta potential values are zero) are more negative for lower-rank and oxi-

dized coals. Such particles acquire more negative zeta potentials in water,

are less hydrophobic (Fig. 12.2), and thus form more stable suspensions.

Fine coal suspensions of hydrophobic bituminous coals coagulate eas-

ily. Honaker et al . (2004) have even shown that the naturally hydrophobic

material (such as coal and graphite) can be selectively coagulated and sepa-

rated from hydrophilic impurities without the use of oily agglomerants and

fl occulants.

VR (1)

VR (2)

V (1)

V (2)

VA

Distance between particles (H )

–

0

Pot

entia

l ene

rgy

of in

tera

ctio

n (V

)

+

12.12 Total interaction energy obtained by summation of an attraction

curve, V A , with two different repulsion curves, V R (Shaw, 1970).



12.3.3 Flocculation

The merit of modern polymeric fl occulants is their ability to produce larger,

stronger fl ocs than those obtained by coagulation. Flocculants are polymers

with high molecular weight that are soluble in water. It is generally accepted

that polymers used as fl occulants aggregate suspensions of fi ne particles by

a bridging mechanism.

The bridging is considered to be a consequence of the adsorption of the

segments of the fl occulant macromolecules onto the surfaces of more than

one particle. Such bridging links the particles into loose fl ocs (Fig. 12.15).

20

40

40

0

–40

1 3 5

pH

pHc

7 9 11

Ec

(%)

Zet

a po

tent

ial (

mV

)

60

80

0

Fresh Oxidized

12.13 Effect of pH, and hydrophobicity, on coagulation of fi ne coal

particles. (Source: Xu and Yoon, 1989 with permission of Elsevier.)



The polymers used in fl occulation can be classifi ed into coagulants, which

are highly charged cationic polyelectrolytes with molecular weights in the

50 000 to 10 6 daltons range, and fl occulants, with molecular weights up to

20 × 10 6 daltons. It is known that fl occulants are not very effective for treat-

ing stable suspensions and, as pointed out by Kitchener (1972), the fl oc-

culation is much more effi cient if the suspension is fi rst destabilized by

coagulation. This can be achieved by changing pH or by addition of some

inorganic coagulants (e.g. lime or alum). Also, low molecular weight cationic

polymers can be used to destabilize suspensions, as most mineral particles in

–

0

Heavily oxidized

Subbituminous

pH

Bituminous

Anthracites

Zet

a po

tent

ial (

mV

)+

12.14 Generalized zeta potential vs pH diagram for coals of various

rank. (Source: Laskowski and Parfi tt, 1989 with permission of Taylor &

Francis)

(a) (b)

12.15 Schematic illustration of (a) bridging fl occulation, (b)

restabilization at high concentrations by adsorbed polymer.



water carry negative electrical charge. In general, the destabilization process

is strongly dependent on process water chemistry.

Adsorption of the polymer is generally necessary for fl occulation to occur.

It is important, however, to realize that adsorption and fl occulation are not

separate sequential processes, but occur simultaneously (Hogg, 1999). There

is general agreement as to the basic mechanism involved in the process; the

optimum fl occulation occurs at fl occulant dosages corresponding to a par-

ticle coverage that is signifi cantly less than complete. Incomplete surface

coverage ensures that there is suffi cient unoccupied surface available on

each particle for the adsorption of segments of the fl occulant chains dur-

ing collision of the particles. The bridging takes place at fl occulant dosages

corresponding to a particle surface coverage that is signifi cantly less than

complete, and thus at higher concentrations, the polymers stabilize suspen-

sions by the mechanism referred to as steric stabilization.

Hogg et al . (1993) showed that the appropriate choice of fl occulants is

determined primarily by chemical factors (mineral composition, solu-

tion chemistry, etc.), but the performance of the fl occulant depends more

on physical variables, such as agitation intensity and the rate of fl occulant

addition.

Flocculants

The vast majority of commercial fl occulants are based on copoly-

mers of acrylic acid and polyacrylamide (also referred to as hydrolyzed

polyacrylamide):

(-CH 2 -CH-) m (-CH 2 -CH-) n

| |

CONH 2 COO - Na +

As a result of hydrolysis even ‘non-ionic’ polyacrylamides contain some

anionic groups. This is expressed as ‘the degree of anionicity’ (the degree

of anionicity of completely hydrolyzed polyacrylamide is 100%, so it is a

polyacrylic acid).

Another important group of fl occulants is polyethylene oxide,

(-CH 2 CH 2 O-) n C (Rubio, 1981; Scheiner and Wilemon, 1987). Cationic poly-

electrolytes such as copolymers of acrylamide and quaternary ammonium

compounds are also available (e.g. Poly-DADMAC). Naturally occurring

materials such as polysaccharides (e.g. carboxymethyl cellulose, starch, guar

gum, etc.) have also been used as fl occulants. According to Kitchener (1978),

the fi rst use of fl occulants involved the application of starch in combination

with lime for the clarifi cation of a coal mine’s effl uent (the patent was fi led

in 1928).



The effectiveness of polymers as fl occulants depends on their molecular

weight, the sign of their charge (e.g. anionic or cationic), and the relative

charge density (for polyacrylamides this is expressed by the degree of anio-

nicity). Depending on molecular weight, the same compounds can operate

as dispersants (e.g. dextrin, low molecular weight) or fl occulants (e.g. starch,

high molecular weight). Low molecular weight copolymers of polyacrylate

type are manufactured as dispersants (e.g., Dispex manufactured by Allied

Colloids (now CIBA), etc.).

Xu and Cymerman’s (1999) data confi rmed that the best fl occulants for

the clay-containing wastes (Syncrude tailings) are moderately anionic high

molecular weight polyacrylamides (optimum around 20–30% anionicity).

Hamza et al . (1988) reported that anionic polyacrylamides were the best for

enhancing the settling rate of fi ne coal.

Polymer molecular weight

The molecular weight of fl occulants is commonly characterized through vis-

cosity measurements. This is based on the Flory-Huggins (Flory, 1953):

[ ]]] = KMa [12.7]

where [ η ] is the intrinsic viscosity of the polymer solution (has units of

reciprocal concentration), M is the polymer molecular weight, and K and a

are constants.

After Henderson and Wheatley (1986) Fig. 12.16 shows the effect of poly-

acrylamide intrinsic viscosity (that is indirectly molecular weight) on the

sedimentation rate of fl occulated tailings for polyarylamides with varying

anionicities.

Because of the relationship between polymer intrinsic viscosity and its

molecular weight (Equation [12.7]), what Fig. 12.16 shows is the effect of

fl occulant molecular weight on fl occulation.

It must be born in mind, however, that the intrinsic viscosity of a polymer

increases with rising solvent quality. This is shown in Fig. 12.17 for various

polymers in various solvents. In a good solvent polymer macromolecules are

in extended form, but coil when the solvent quality decreases. This may hap-

pen when the ionic strength of the system is increased, or the pH is changed.

The exponent a in Equation [12.7] is a measure of the solvent quality and, as

Fig. 12.17 shows, it is large (larger than 0.5) for a good solvent, and smaller

than 0.5 for poor solvents.

Since conformation of polymer macromolecules in solvent depends on

the solvent quality, also polymer adsorption onto solid particles depends on

it. Adsorption is generally higher from a poor solvent than a good solvent

(Koral et al ., 1958).



00

0.5

Log

(Tai

lings

sed

imen

tatio

n ra

te)

(cm

/min

)

1

1.5

2

2.5

10

Polymer intrinsic viscosity (L/g)

20 30

12.16 Effect of polyacrylamide intrinsic viscosity (molecular weight) on

sedimentation rate of the fl occulated tailings. (Source: After Henderson

and Wheatley, 1986). Copyright Taylor & Francis Group, LLC. (http://

www.taylorandfrancis.com), reproduced with permission.)

101

102

[η]/(

ml g

–1)

103

104 105 106 107

M /(g mol–1)

a = 0.0

a = 0.5

a = 1a = 2

12.17 Intrinsic viscosity [ η ] as a function of the molar mass Mw for

different polymer-solvent systems as given by Eq. [7]. In addition to the

straight line relationships indicating the range of good solvents (a > 0.5)

and bed solcents (a<0.5) the experimental curve for polyacrylamide in

water at 25 °C ia also shown after Kulicke and Clasen (2004).



Klein and Conrad (1980) derived the following empirical equation that

can be used to determine polyacrylamide molecular weight. This relation-

ship holds for polyacrylamide samples ranging in molecular weight from 5

× 10 5 daltons to 6 × 10 6 daltons when measurements are conducted in 0.5 M

NaCl solution at 25 ° C.

[ ] .]] = ⋅−7 1. 9 1⋅ 0 3 0 77Mw [12.8]

Testing fl occulation

Solubilities and rates of dissolution of high molecular weight fl occulants in

water are generally low, and preparation of the polymer solution is a very

important fi rst stage (Brown, 1986). The following step, mixing the polymer

solution with the suspension, is critical (Owen et al ., 2009).

Since fl occulants are either used to enhance solids settling rates to maxi-

mize thickener capacity, to enhance dewatering by fi ltration, or to improve

water clarifi cation, various tests are utilized. They include measurements of

solids settling rate, sediment density, fi ltration characteristics, and superna-

tant turbidity.

Several techniques have been proposed to determine the settling velocity

in laboratory experiments, the ‘jar tests’ being the most common (Coe and

Clevenger, 1916; Richardson and Zaki, 1954; Michael and Bolgers, 1962).

Jar testing involves homogenization of the suspensions in settling cylinders,

introduction of the fl occulant, and mixing by moving a plunger up and down

in the cylinders (or by inverting the cylinders several times). This procedure

is claimed not to be satisfactory because of the local over-dosing that can

occur when the relatively concentrated fl occulant solution meets the slurry

(Kitchener 1978); but more important is that the agitation in this method

does not produce the optimum fl occulation. Farrow and Swift (1996) dem-

onstrated that the jar test has several problems. It is important to realize that

adsorption and fl occulation are not separate sequential processes, but occur

simultaneously (Hogg, 1999). The commonly used improved experimental

procedure includes addition of the fl occulant to a vigorously agitated sus-

pension, which is immediately stopped after addition of the reagent (Keys

and Hogg, 1979). Different mixing/polymer addition conditions may result

in very different fl oc sizes and settling rates. Owen et al . (2009) showed that

mixing of the slurry with a dilute fl occulant solution within the feedwell

determines the performance of commercial thickeners. It was also shown

that under certain conditions intense agitation for short times may even

change the nature of fl occulation, from total fl occulation to a selective fl oc-

culation of only some mineral constituents (Ding and Laskowski, 2007).



The use of the shear vessel, as described by Farrow and Swift (1996) and

Rulyov 1999, 2004), in the fl occulation tests was recently tested by Rulyov

et al . (2011) and Concha et al . (2012). Their set-up is shown in Fig. 12.18. The

use of a shear vessel (similar to rotational Couette viscometers) in assessing

fl occulation effi ciency has the advantage of quantifying the mixing intensity

through the shear rate. Using the shear vessel, Rulyov (1999) and Rulyov

et al . (2004) have demonstrated that the mixing time in fl occulation can be

reduced, from minutes to 5–6 s, by the appropriate hydrodynamic treatment

of the suspension at a given shear rate. In the set-up shown in Fig. 12.18 the

Couette-type vessel is rotating in the external vessel, the gap being 1.5 mm.

The reactor is fed continuously with the suspension by a measuring peri-

staltic pump. Before entering the Couette reactor, the pulp receives contin-

uously a diluted fl occulant solution, at a fl ow rate to give a pre-determined

dosage. After 6 s treatment at a pre-determined shear rate in the Couette

reactor, the fl occulated suspension is discharged from the ultra-fl occulator

through a 3 [mm] inner diameter transparent tube equipped with an opto-

electronic sensor that registers the fl uctuation of the intensity of the light

beam passing normally through the tube (in accordance with techniques

proposed by Gregory and Nelson (1984)). The electronic signal is processed

Computer

Motor

Feed slurry

Levelsensor

To waste

Settlingcolumn

Flocculant

Variablespeed motorPeristaltic

pump

Diversionwhen

sampling

P

P

P B

AC

200 mm

12.18 Schematic illustration of the set-up used to test fl occulation.

(Source: After Concha et al ., 2012 with permission of Elsevier.)



and displayed in a three digital format, thus showing, in relative units, the

values of fl occulation effi ciency (or mean fl oc size) and the mean shear rate

γ . In the tests designed to measure the settling rate of the treated suspen-

sion, the suspension from the outlet of the tester is continuously fed to a

14 mm diameter settling cylinder and, as soon as the suspension fi lls the cyl-

inder, the cylinder is inverted once and the initial settling rate is measured.

The fl occulation of fl otation tailings from one of the major Chilean copper

mines and Orifl oc-2010 polycarylamide fl occulant in a Couette-type reactor

was recently reinvestigated by Concha et al . (2012). By varying the shear rate

from 100 to 2000 [s − 1 ], the solid concentration from 1 to 15 % by volume,

and the fl occulant dosage from 0 to 20 g/ton, it was shown that an important

interaction exists between these variables. At the optimal fl occulant dosage,

the optimal suspension concentration and the optimal fl occulation time, an

increase by 50% in the solid fl ux density function is possible when the shear

rate of γ = −100 1s is changed to the optimum value of around γ ≈ −400 1s .

It is worth pointing out that ultrafl occulator-type devices have already

been installed at some coal preparation plants (Rulyov, 2004; Rulyov et al ., 2009).

Testing the use of fl occulants in fi ltration

In present day practice, disposal of the fi ne waste fraction (for instance, coal

fl otation tailings) is usually accomplished by:

1. addition of fl occulant to the slurry and thickening of the resulting fl ocs

in a thickener, with the underfl ow from the thickener pumped to an

impoundment area;

2. treatment of the thickener underfl ow with fl occulant and dewatering on

a mechanical device, such as vacuum fi lter or fi lter press.

The former method requires the availability of enough land environmen-

tally suitable for construction of impoundments and, with the latter, the

dewatered material can be mixed with coarse waste fraction and discarded

by stacking. This latter method requires dewatering by fi ltration.

The dewatering process during fi ltration comprises several stages, as illus-

trated in Fig. 12.19 (Lockhard and Veal, 1996). Flocculants reduce the seg-

regation of fi ne and ultrafi ne particles, which causes blinding of the fi lter

medium, and thus increases dewatering rates.

Experimental determination of the effect of fl occulants on fi ltration usu-

ally involves fl occulation of the tested slurry, transfer of the fl occulated

material to a funnel-type fi lter that is operated under controlled vacuum,

and determination of both the fi ltration rate as well as the cake moisture

content. The fi ltration rate is described by Darcy’s equation:



d

d

Vt

KA P

H=

η [12.9]

where d dtdV / is the volumetric fl ow rate of fi ltrate through a fi lter cake in

time t , A is the fi lter area, Δ P is the pressure drop across the fi lter cake, η

is the liquid viscosity, H is the cake thickness, and K is the rate constant

referred to as permeability.

KkS

=ε 3εε

2 2( )− ε( )ε1 [12.10]

where ε is the cake porosity, S is the specifi c surface area of the particles, and

k is the Kozeny constant.

The effect of fl occulants on the fi ltration rate is related to changes in fi lter-

cake properties; the permeability, which is determined by porosity of the

cake, is the most important. As it is determined by particle size and shape,

the appearance of the fl ocs improves the cake porosity. In the simple fi ltra-

tion tests in a funnel-type fi lter the particles segregate. Figure 12.20 shows

Initial bridging startscake formation and

filtration commences

Cake forming Cake formed

Cake compacted Desaturation by air Air breakthrough

(d) (e) (f)

(a) (b) (c)

12.19 Schematic representation of stages in dewatering by fi ltration.

(Source: After Lokhart and Veal, 1996). Copyright Taylor & Francis

Group, LLC.(http://www.taylorandfrancis.com), reproduced with

permission.)



a better set-up (Tao et al ., 1999), which has been introduced to avoid this

problem.

In the device shown in Fig. 12.20, the slurry is continuously circulated to

avoid sedimentation and the fi lter leaf is submerged in well-mixed slurry.

The tests allow determination of the fi ltration rate vs fi lter-cake thickness,

and the cake moisture content vs the fi lter thickness.

As Fig. 12.19 shows, after the formation of the cake, the cake compres-

sion and water expression follow. The situation depicted in Fig. 12.19d is the

saturated capillary state, with all the pores fi lled with water. Under these

conditions, a capillary pressure opposes air entry, and only when the applied

pressure exceeds the capillary pressure does the removal of water from the

cake commence (Hogg, 1995).

ΔΘΡ =

2γ cos

r [12.11]

where γ is the surface tension of the liquid, Θ is the contact angle at the

solid–liquid interface (receding angle should be used).

Equation [12.11] describes the applied pressure necessary to expel liquid

from the pores in a packed bed. The equation shows that more hydrophobic

coals are easier to dewater, and that the use of surfactants to lower water

Circulationsump

Circulationpump

Stirrer Moisturetrap

Filtercell

Moisturetrap

Vacuumpump

12.20 Illustration of experimental set-up for fi ne coal dewatering by

fi ltration. (Source: After Tao et al ., 1999). Copyright Taylor & Francis


permission.)



surface tension may also have some merit. From this point of view, the fi ltra-

tion aids can be classifi ed into:

1. Flocculants (to increase cake porosity and thus fi ltration rate).

2. Oily hydrocarbons (to make the coal particle more hydrophobic, agglom-

erate fi ne particles, and thus increase not only cake porosity but also par-

ticle hydrophobicity).

3. Surfactants (to lower water surface tension and thus reduce capillary

retention forces).

While polyacrylamide fl occulants are commonly used to improve the set-

tling rate, their application in fi ltration circuits has some specifi c functions.

The use of high molecular weight non-ionic and anionic fl occulants shows

that there exists a fl occulant addition range where there is a signifi cant reduc-

tion in the moisture content of the fi lter cake. Increased polymer addition

results in an increased moisture content. This can be explained by different

fl ow modes of liquid through the fl occulate fi lter cake. While the increased

fl occulation of fi ne particles may improve the interfl oc fl ow, the drainage of

liquid from the fl ocs becomes increasingly diffi cult (Mishra, 1987). This sug-

gests that, while at certain doses the cake moisture content can be reduced,

it may increase again at high fl occulant doses. When moisture reduction is

the objective, for high molecular weight polymer, over-dosing can lead to

an increase in the moisture content. With low molecular weight polymers, a

greater degree of moisture reduction can be achieved (Mishra, 1987). Since

fl occulants are hydrophilic macromolecules, their adsorption makes solid

particles hydrophilic and this also increases the ability of the solid to retain

moisture.

Emulsifi ed oil can also be used as a fi ltration aid. Such emulsion not only

agglomerates fi ne coal particles but also renders them hydrophobic. Nicol

and Rayner’s (1980) results demonstrate that an oil addition of 1% can lead

to a lowering of the fi lter-cake moisture from 26% to 16% (wet basis). This

considerable lowering of fi lter-cake moisture is accomplished by an increase

in both fi ltration rate and solids pick-up rate as oil droplets agglomerate

fi ne coal particles. A dramatic improvement in fi ne coal dewatering with

the use of asphalt emulsions was reported by Wen et al . (1994). The effect

of fl occulants, emulsifi ed oil, and the addition of surfactants on fi ne coal

fi ltration have been studied by Laskowski and Yu (1998, 2000). They have

shown that the emulsifi cation of kerosene in the presence of surfactants

can dramatically reduce the size of oil droplets and, more importantly, can

entirely change the electrokinetic properties of such droplets. The particles

of low-rank/oxidized coals are negatively charged in water and are diffi -

cult to agglomerate with oil. However, when the oil is emulsifi ed in solu-

tion with a cationic surfactant, the obtained cationic-emulsion agglomerates



such a suspension effi ciently, improves fi ltration, and also reduces the cake

moisture content. Novel dewatering aids were patented by Yoon and Basilio

(U.S. Patent 5,670,056). As pointed out by Yoon et al . (2003, 2006), while the

addition of 0.5 kg/t of kerosene could reduce the cake moisture content by

5%, the use of their novel additives was able to reduce the moisture con-

tent from 20–30% down to the 8–14% range using vacuum fi ltration. The

improved performance of new hydrophobizing agents has been attributed

to a simultaneous decrease in fi ltrate surface tension and increase in particle

hydrophobicity. As disclosed in a patent (U.S. Patent 5,670,056) these new

dewatering additives include fatty acid esters; the examples provided indi-

cate that they are mostly used as a solution in butanol. These agents seem

to be similar to the EKT agent patented as a coal fl otation promoter (Polish

Patent 104,569).

Use of fl occulants in coal processing

In the most common applications, anionic polyacrylamide fl occulants are

applied with molecular weight of 10 6 –10 7 daltons.

While fl occulation of coal-clay fi ne waste generated by coal preparation

plants that is then pumped to a pond or impoundment provides a method

for rapidly recovering most of the water contained in the waste slurry, this

technique also requires the availability of enough land. A method of disposal

that minimizes the land requirements is that whereby the underfl ow from

the thickener is treated with a fl occulant and dewatered on a mechanical

device such as fi lter or belt press. Such treatment yields a material that can

be mixed with the coarse waste fraction and disposed by stacking. However,

this technique heavily depends on whether the treated material reslurry

during a wet season. As was shown by the US Bureau of Mines (Brown

and Scheiner, 1983; Scheiner, 1996), an interesting alternative is the use of

polyethylene fl occulant (PEO). In the developed process of coal-clay waste

fl occulation with PEO, addition of calcium or magnesium salts (for instance

lime) is required. Their results indicate that prior coagulation before appli-

cation of PEO is very important for effi cient bridging. In this process lime is

added up to pH 9 or higher and the PEO dosage required to get optimum

results varied from 50 to150 g/t. A dewatering system developed by the US

Bureau of Mines uses PEO to fl occulate the waste, followed by dewatering

on a static screen. The waste is dewatered (using about 0.05 kg/t of PEO) to

produce a material having a solid content ranging from 55% to 60%. It is

important that the dewatered material is quite stable and does not re-slurry

easily when brought into contact with water. Comparison of PEO with PAM

(with about 40 Da) in the dewatering of the fl occulated coal-clay wastes

on a belt fi lter press showed that the required dosages of PEO were much

lower than those with PAM to obtain about 70% solids content.



12.3.4 Oil agglomeration

Insoluble ‘oily’ collectors are utilized in coal fl otation. Such oils appear in

the pulp in the form of droplets when oil is emulsifi ed in aqueous environ-

ment. The droplets attach to the particles that are hydrophobic and then

bridge these particles when the system is vigorously stirred. This is so-called

oil agglomeration. Three major factors control oil agglomeration (Capes,

1991): (a) solid wettability; (b) the amount of the oil; and (c) the type and

intensity of conditioning. The amount of agglomerating oil is critical. At low

oil levels, discrete lenses-shaped rings are formed at the points of contact of

the particles (Fig. 12.21a and 12.21b). At higher doses, the oil rings begin to

coalesce and form a continuous network (Fig. 12.21b and 12.2c).

Since droplets of pure aliphatic hydrocarbons attach only to very hydro-

phobic particles, they are very selective. Droplets of the oils containing

polar hydrocarbons can also attach to the high-ash coal particles. This trans-

lates into higher yields of the agglomerated product and higher ash con-

tent. ‘Heavier’ oils were shown to yield excellent recoveries under the more

intense mixing conditions needed to disperse more viscous oil in water

(Capes and Germain, 1982; Capes 1991). Fuel oils with addition of aromatic

hydrocarbons were found to be very good agglomerants (Agus et al ., 1996;

(a) (b) (c)

(d)

12.21 Oil distribution on moist agglomerates: (a) pendular state; (b)

funicular state; (c) capillary state; (d) oil droplets with particles inside or

at its surface.



Blaschke, 1990). Labuschange (1987) and Good et al . (1994) described how

addition of alcohols to oil can improve oil agglomeration.

New light was shed on the mechanism of the oil agglomeration process

in a series of papers by Wheelock and his co-workers (Drzymala et al ., 1986;

Drzymala and Wheelock 1997; Milana et al ., 1997). Their results revealed

that the agglomeration process of a moderately hydrophobic coal with hep-

tane is triggered by a small amount of air present as a separate phase. The

rate of agglomeration increases as more air is admitted to the system, or as

the amount of agglomerant or agitation increases.

Oil droplets can attach only to the particles that are hydrophobic, and

the oil agglomeration of low-rank, subbituminous coals is not very effi cient.

Pawlak et al . (1985) reported that subbituminous coals can be agglomerated

with heavy refi nery residues. Since particles of low-rank coals are charged

negatively in water, it was also demonstrated that kerosene with 1% addi-

tion of dodecylamine when emulsifi ed in water produces oil droplets charged

positively in a very broad pH range that agglomerate subbituminous coal

very well (Laskowski and Yu, 2000).

Various oil agglomeration processes have been developed (Mehrotra

et al ., 1983). These include Trent, Convertol, NRCC (National Research

Council of Canada), Shell, Olifl oc, CRFI (Central Fuel Research Institute

in Dhanbad, India), and BHP processes. In the process developed at the

National Research Council in Ottawa (Capes, 1991), the light-oil addition

and high-intensity conditioning produces micro-agglomerates, and this is

followed by the addition of heavy oil and low-intensity conditioning to pro-

duce a handleable product that can be recovered by screening. In the more

recent developments, the addition of oil was reduced to a few percent, and

this provides better selectivity. In these more recent fl owsheets, the high-

intensity mixing is followed by a low-intensity mixing, which enables the

coal-oil agglomerates to enlarge and strengthen. The agglomerated slurry

is then passed to a Vor-Siv screen. The recovered agglomerates are further

dewatered in a screen bowl centrifuge. In a fi nal stage, the agglomerates can

be further enlarged by pelletization.

Solid–liquid separation in fi ne coal cleaning circuits is always diffi cult (and

expensive) and the fl owsheet shown in Fig. 12.22 constitutes a very inter-

esting alternative to conventional dewatering and drying. In this process

classifying hydrocyclones are used to split the fl otation feed into –0.1 mm

and +0.1 mm fractions. The former is processed by oil agglomeration while

the latter goes to fl otation. The Oilfolc process offers a means for avoiding

thermal drying. The fl otation concentrate obtained when fl oating coarser

feed dewatered by vacuum fi ltration contains 2% less moisture and the oil

agglomeration of the fi ne fraction produces 6–8% ash clean coal products. It

is worth noting that such a process can dramatically improve handleability

of the fi nal product.



12.3.5 Pelletization

Coal preparation plants rely mostly on gravity separation techniques, and in

the past the coarse products constituted main saleable products. Introduction

of new separation methods such as fl otation resulted in a wide utilization of

fi ne size fractions as well. This in turn has resulted in a considerable increase

in the amount of fi ne clean coal products which poses severe storage, trans-

portation and pollution problems. One possible method by which these

problems can be alleviated is pelletization of fi ne coal.

As reported in literature (Leonard and Newman, 1989; Holuszko and

Laskowski, 2010), the addition of a few percent of water to a dry coal can

substantially increases its bulk density. This indicates that in the presence of

water droplets some sort of consolidation takes place in the bed of fi ne dry

particles. Such a process can be further intensifi ed by tumbling the particles

in a disk or drum (pelletization).

The pelletization process is controlled by capillary phenomena, and the

driving force for the pelletization is the lowering of the total surface free

energy of the system through a reduction of the effective air–water interfacial

area. As shown by Kapur and Fuerstenau (1966) and Sastry and Fuerstenau

(1977), the kinetics of pellet formation proceed through three stages: the

formation of nuclei agglomerates, the transitional stage, and the ball growth

stage. The rolling action of the drum brings the individual particle into prox-

imity with each other, so that the physical forces become operative and cause

–0.5 mm

–0.5 + 0.1 mm

–0.1 mm

Oil

C

C

W

12.22 Flowsheet of the Oilfl oc test set-up in the Monopol Dressing

Plant of Ruhrkohle AG. (Source: After Bogenschneider and Jasulaitis,

1977.)



the particles to rearrange, and surface tension reduction brings about nuclei

formation via the bridges of wetting liquid. In the agglomeration of granu-

lar materials by capillary forces, the wettability of the aggregated particles

by the applied liquid plays an important role (Schubert, 1984). Figure 12.23

shows the coal pelletizing circuit (Sastry and Mehrota, 1981).

Only a limited number of research results on pelletization of fi ne coal are

available, and Sastry and Fuerstenau’s EPRI report (1982) is probably the

most valuable source of information in this area. Some results of this report

will be used to illustrate the principles. Sastry and Fuerstanu (1982) found

that it is possible to pelletize coal of different ranks and widely different size

distributions. Rive et al . (1964) reported that even coal as coarse as 1.2 mm

could be pelletized, provided that it contained at least 50% of − 74- μ m mate-

rial. The rate of the process is governed by the operating conditions and the

type of coal. The rate of pellet growth is strongly affected by the amount of

feed moisture. The amount of moisture required to pelletize is different for

different coal samples, and was found to be mainly determined by the ash

content.

Figure 12.24 shows the pelletization results obtained with different sam-

ples of raw coal.

The coal fi nes for these experiments were prepared by stage-crushing

in a jaw crusher, and dry grinding in a laboratory ball mill. Two hundred

Feedconveyor

Binderbin

Mixer andfluffer

Water supply

Liquid binderstorage

Sprays

Pelletizingdisc

Rev. conveyor

Shredder

Roller-screen

Dry pellets to silo or stockpile

Belt dryer

12.23 Flowsheet of a coal pelletizing circuit with pelletization disk.



gram batches of the dry coal utilized in the pelletization experiments

were premixed with the desired amount of distilled water (10 min) to pre-

pare a moist feed. The moist feed was then placed in a 150-mm length by

225-mm diameter pelletizing drum, enclosed in a humidifi ed chamber and

tumbled at a rotational speed of 40 rpm. At the end of each experiment,

a portion of the pellets was used for determining their moisture content

and compressive strength. As Fig. 12.24 demonstrates, pelletization of fi ne

coal is possible only within a narrow range of moisture additions. Outside

that range, the coal fi nes are either too dry to agglomerate or too wet

and form lumpy and weak agglomerates. Ash content was found to be

the most signifi cant property that determines moisture requirements for

pelletization.

Sastry and Fuerstenau’s (1982) report also provided information on the

effect of binders on fi ne coal pelletization. The dry compressive strength

of the pellets produced without a binder is from 1 to 10 kg strength. Pellet

strength was found to decrease with decreasing ash and sulfur content. Corn

starch was found to be a very good binder, in that it improves the compres-

sive strength of the pellets by several hundred percent; however, such pellets

exhibit poor resistance to moisture penetration. On the other hand, asphalt

emulsion, which was found not to improve pellet qualities much, renders

Dekker

Belle ayre south

Shamokin anthracite

Illinois No. 6

UpperFreeport

Sewickley

Freeport

WestKentucky

LowerFreeport

Ohio no. 9

Pittsburgh no. 8

LowerKittanning

015

20

Moi

stur

e ra

nge

for

pelle

tizat

ion

wt (

%)

30

40

50

20Ash content wt (%) (dry basis)

40 60

12.24 Infl uence of ash content (dry basis) of coal samples on their

moisture requirements for pelletization. (Source: After Sastry and

Fuerstenau, 1982.)



pellets waterproof. The pellets produced with 1% by weight of corn starch

that were subsequently sprayed with asphalt emulsion produced strong pel-

lets which were also water resistant.

12.4 Fine coal handleability

The coal must fl ow by gravity in each operation in the distribution system

in order to keep these operations and the whole transportation process

effi cient. Handleability is commonly defi ned as the ability of the coal to

pass through a handling system without causing blockages and hold-ups

(Brown, 1997).

Preparation plants produce coal as blends of various size fractions.

Metallurgical coal fi nes are treated by fl otation, and fi ne coal products are

recovered from fl otation circuits by fi ltration. The fi lter cake commonly has

a moisture content in excess of 20%. This, in combination with the rest of the

blend, can result in handling problems. In the preparation of thermal coals,

untreated fi nes are usually recombined with the cleaned coarse fractions

from wet gravity separation, which also contain a considerable amount of

moisture. Many handleability studies carried out over the years have related

some of the physical parameters of coal samples, such as amount of fi nes,

mineral matter type, and moisture content, to the handling characteristics.

In the 1950s British coal scientists (Cutress et al ., 1960) developed a method

to provide a means of assessing the ease of discharge of washed coals and

blends from the hopper at the bottom of rail wagons. The Durham Cone

test offered a quick and relatively simple way to determine coal fl owability.

The vibrating cone was designed to imitate train movement (Fig. 12.25); as

a result, the behavior of coal during transport by trains could be reproduced

using the Durham Cone. Over the years the method became used as the

Coal blend

Coal handleability

Collectingbin

Lift springloaded leverto releasedischarge plate

Vibration inhorizontaldirection

Discharge plate

500 mm dia.Stainlesssteelhopper

12.25 Schematic illustration of Durham Cone.



standard test to examine handleability, and was used in many studies with

varied success. A fi rst comprehensive study on handleability of coal using

Durham Cone was published by Hall and Cutress (1960), and was followed

by many others (Vickers, 1982; Mikka and Smitham, 1985; Arnold, 1992;

Brown, 1997). The major drawback of the method was that the reproduc-

ibility of the results was questionable. It was also pointed out that mixing

of the sample before the test was extremely important, and that any mixing

involving rolling produced a balling effect and altered the fl ow properties of

the mixture as measured by Durham Cone.

The test results are expressed as Durham Cone Index, DCI = mass sam-

ple/average time, or as a fl ow rate (or Durham Cone discharge rate), with

the units being kg/s.

The analysis of the data reported by several researchers failed to detect

any simple correlations (Vickers, 1982; Brown et al ., 1996). Pretty good cor-

relation was, however, found between the DCI and the product of the sim-

ple moisture content and the fi nes ash fraction, where the fi nes ash fraction

is defi ned as:

γ 0γγ 038 0 038

100

. .038 0Ash

Ashwhole

γ 0.038 is the weight percent yield of the − 0.038 mm size fraction, Ash 0.038 is the

ash content in the − 0.038 mm material, and Ash whole is the ash content in the

whole sample.

This correlation is shown in Fig. 12.26 (Brown et al ., 1996, 1997, 2000).

R = 0.94

800600400

Moisture * Ash fraction of the fines <0.038 mm

20000

1

Dur

ham

Con

e In

dex

(kg/

s)

2

3

12.26 Durham Cone Index vs moisture x fi nes ash fraction parameter.

(Source: After Brown 1997.)



Handleability problems can easily be solved by removing wet fi nes from

the power station blend, but this solution is not conducive to maximizing the

coal industry saleable output. Better understanding of the factors that con-

trol fi ne coal handleability is vital for fi nding a long term solution that would

allow further dewatering/reconstitution of the fi lter-cake products into any

saleable product.

In the tests described by Jones (1990) it was considered that the recon-

stitution of the − 0.5 mm size fraction into coarser granules capable of with-

standing shear in storage and handling had to be an essential part of any

fi nal solution. The results are given in Fig. 12.27. As this fi gure indicates, the

fi lter cake addition in excess of 8% of the total blend placed the blend han-

dleability below the target level and into the problem zone. By pelletizing

the − 0.5 mm fi nes and adding it in equal parts together with non-pelletized

fi nes the handleability was drastically improved.

Handleability has recently been re-examined and it has been shown that

it critically depends on coal surface properties (Holuszko and Laskowski,

2004).

In this research program, pelletizing was used to characterize fi ne coal

tendency to aggregate when it is transported/tumbled.

As Fig. 12.28 shows, depending on coal surface wettability the adhesion

of water droplets onto the coal surface will be different. While in the case of

weakly hydrophobic (subbituminous or oxidized) coal the water will tend to

00

2

4D

urha

m C

one

disc

harg

e ra

te (

kg/s

)

6

10 20

Filter cake addition (% wt)

30 40

Filter cake only

Target handleability

Filter cake/Pellet mix ratio 1:1

50

12.27 Durham Cone discharge rate for various percentages of fi lter

cake addition to the power station blend. (Source: After Jones, 1990

with the permission of the Mining and Materials Processing Institute of

Japan.)



spread on the surface and strongly bridge coal particles, in the case of very

hydrophobic bituminous coal (upper part of Fig. 12.28) the adhesion of the

droplets to coal surface will be weak and this will lead to a formation of very

frail pellets.

It is rather well established that the content of fi nes and the moisture

content are the most important factors (Wawrzynkiewicz, 2003; Arnold,

2004) determining coal handleability. However, tests on the effect of mois-

ture content clearly indicate that this may be very different depending on

coal surface properties. Figure 12.29 shows the fl ow curves determined with

the use of the Durham Cone for two coal samples.

Medium-volatile bituminous coal samples from British Columbia were

used in these tests. The oxidation degree of the tested samples was deter-

mined using the transmittance measurement of the alkali extracts (ASTM D

5263–93), and their wettability was characterized by the equilibrium mois-

ture (ASTM D 1412–93). The high equilibrium moisture indicates hydro-

philic coal, while low (around 1%) moisture indicates hydrophobic coal. The

high transmittance values indicate non-oxidized coal (hydrophobic), the

values well below 90% pointing to oxidized coal (hydrophilic). The equi-

librium moisture for LC 3 sample was 1.30%, while for the LC 8U sample

it was 7.34%; the transmittance values for these samples were, respectively,

95.25% and 26.06%.

Water droplet

Coal surface

Water layer

Coalsurface

12.28 Schematic illustration of the effect of coal wettability on

the behavior of water droplets on coal surface; upper part very

hydrophobic bituminous coal, bottom part lower-rank/oxidized coal.



As Fig. 12.29 shows, the fl owrate patterns for hydrophobic LC 3 and

hydrophilic LC 8U are different. The fl ow rates were plotted vs surface

moisture for these samples. The surface moisture was obtained as the dif-

ference between actual moisture content of the sample and its equilibrium

moisture. It can be noticed that at the same amount of surface moisture, e.g.

at 10% surface moisture (Fig. 12.29), the LC 8U coal sample ceases to fl ow,

while the LC 3 coal is still handleable.

For lower-rank and oxidized coals, the equilibrium moisture is usually

much higher than for high rank and non-oxidized coals. Therefore, these coals

can tolerate higher moisture levels before they become diffi cult to handle.

However, at moisture levels exceeding their equilibrium moistures, a rapid

deterioration of their handling behavior is observed, leading to practically

non-fl owing conditions. In the case of hydrophobic coals, only the mineral

matter affects handleability signifi cantly. For such coals, an increase in mois-

ture content tends to affect handleability to a certain level, but these coals

do not cease to fl ow completely. Apparently a high amount of clay material,

even in very hydrophobic coals, can have a very damaging effect on handle-

ability. These effects can also be tested using pelletization experiments, as it

was shown that coals that can easily be pelletized may become diffi cult to

handle at a given level of moisture (Holuszko and Laskowski, 2004).

The picture that emerges from the discussed available results indicates

that whenever fi ne coal particles are ready to aggregate (as a result of their

surface properties and the moisture content) they will make the whole blend

of various saleable size fractions aggregate, and that will cause handleability

0.00

0.50

1.00

1.50

2.00

2.50

3.00

0.00 5.00 10.00 15.00 20.00

Surface moisture content (%)

Flo

w r

ate

(kg/

s)

LC8U LC 3

12.29 Durham Cone discharge rate for LC 3 (36% fi nes) and LC 8U (37%

fi nes) coal samples as a function of surface moisture content. (Source:

After Holuszko and Laskowski, 2006.)



problems. The readiness of such a material to aggregate is what creates the

problem and this ‘readiness degree’ can be ‘discharged’ only by either ren-

dering such particles more hydrophobic, or by allowing them to aggregate

by pelletizing this material.

12.5 Rheology effects in fine coal processing

The importance of this topic has further increased with introduction of coal-

water slurries as another way of fi ne coal utilization.

12.5.1 Viscosity of suspensions

The viscous nature of the fl uid, i.e. its internal friction, is manifested only

when one region of the fl uid moves relative to another. The measurements

may involve the fl ow of liquid through capillary under a given pressure,

or movement of the liquid placed in between two cylinders (Couette) and

caused by the rotation of one of them. In all such cases, the liquid fl ow rate

(shear rate) is related to the applied shear stress (pressure) via Newton’s

low of viscosity:

τ ηDη [12.11]

where τ is the shear stress (Pa), η is the dynamic viscosity (Pa · s, Pascal

second = Nm–2s), and D is the shear rate (s − 1 ). In the old system a con-

venient measure was the centipoise (cP = 0.01 poise), because water has

a viscosity of about 1 cP at room temperature (on the SI scale this cor-

responds to 1 mPa · s).

Figure 12.30 is a sketch of the shear stress vs shear rate relationship for

several different fl uids.

For a stable (no aggregation) diluted suspension of solid spherical par-

ticles in Newtonian liquid, Einstein derived the following expression

η η= ηo( )φφ+ k [12.12]

where η is the viscosity of the suspension, ηo s is the viscosity of the sus-

pending medium and φ is the volume fraction of the suspended solid

particles.

Rheological measurements, when relative viscosity (relative viscosity =

η η/ o ) is plotted vs solid concentration (usually v/v %), provide the curve

displayed in Fig. 12.31. As this plot demonstrates, the viscosity of a suspen-

sion of spherical particles signifi cantly deviates from Einstein’s law above

φ = 0 1 , and beyond a volume fraction, φmaφφ x , called the maximum packing



fraction, the dispersed particles lock into a rigid structure and fl ow ceases.

The packing fraction for hexagonally packed monodisperse spheres is 0.74.

As pointed out by Barnes et al. (1989), these values are much smaller for

other particles (only 0.3 for grains, and 0.2 for rods).

0

1

Rel

ativ

e vi

scos

ity (

ηr)

1Ømax

Volumetric solid fraction (Ø)

12.31 Schematic graph showing the concentration dependence of the

relative viscosity of a typical suspension. The slope at the origin equal

intrinsic viscosity.

Plastic

Pseudoplastic

Newtonian

She

ar s

tres

s, τ

Shear rate, D

Dilatant

12.30 Rheological curves for Newtonian and several non-Newtonian

fl uids.



The relative viscosity function depicted in Fig. 12.31 in the dilute region is

determined by the requirement that the slope at the origin equals the intrin-

sic viscosity (2.5 for spheres). The function is best described by the Krieger–

Dougherty equation (Krieger and Dougherty, 1959; Kierger, 1972):

η ηη

φφ

ϕ

rηoη

= = −⎛⎝⎜⎛⎛⎝⎝

⎞⎠⎟⎞⎞⎠⎠

−

1maφφ x

[ ]ηη maϕϕ x

. [12.13]

where ηr is the relative viscosity, η is the viscosity of suspension/colloid, ηo

is the viscosity of suspending medium, φ is the volume fraction of the solid,

φmaφφ x is the maximum packing fraction, and [ ]]] is the intrinsic viscosity.

Taking [ ]]] = 2 5. and φmaφφ x = 0 7. 4 for close-packed uniform spheres the

exponent in Equation [12.13] is 2.5 × 0.74 = 1.85. The value of [ ]]] for non-

spherical particles is much larger than 2.5.

The rheology of fi ne particle systems depends on many factors: particle

size, shape and solids concentration. Another factor that strongly infl uences

rheological behavior is particle–particle interaction. In general, rheologi-

cal behavior becomes more non-Newtonian as the particle size decreases.

While well-dispersed systems of spherical particles at low-to-moderate solid

concentration (below 40%) exhibit Newtonian behavior, the aggregating

systems are non-Newtonian. As Equation [12.13] indicates, the measured

relative viscosity becomes very large when ϕ is near to the close-pack condi-

tion and the effect of the factors discussed above becomes particularly evi-

dent in this high solid content range. The utility of the Krieger–Dougherty

equation is in the fact that it takes into account the effect of particle-size dis-

tribution and particle shape, as both affect the maximum packing fraction.

12.5.2 Coal-water slurries (CWS)

Typically coal that has been mined and cleaned in a coal preparation plant is

shipped either to a power generation plant (thermal coal) or to a coke-mak-

ing plant (metallurgical coal). This needs large areas for dry sample storage,

handling, and transportation, and poses all sorts of environmental problems

(e.g. dusting, spontaneous combustion, etc.).

Coal–water slurries (CWS) (also referred to as coal–water fuels (CWF),

or as coal-water mixtures (CWM)) are highly loaded suspensions of fi ne

coal in water. Since these are mixtures of coal and water, CWS is free from

some of the major problems of solid coal, such as dusting and spontaneous

combustion during storage and transportation. Unlike solid coal, CWS is

a liquid, so it does not require large handling facilities. Utilization of fi ne

coal in the form of CWS also simplifi es the fi ne coal preparation circuit, in



that it does not need deep dewatering and drying. The fi ne coal in a fi lter

cake is directly converted into CWS and is pipelined to a power-generating

plant, where it is burned like a heavy oil; the coarse coal does not have to be

blended with the troublesome fi nes and so its handling is improved as well.

Since CWS is utilized as a fuel to generate power, its calorifi c value is an

important factor. Because of the loss of energy on water evaporation (latent

heat of evaporation for water is about 2300 kJ/kg) the presence of water in

CWS reduces its heating value. For instance, for one kilogram of CWS con-

taining 60% of bituminous high-volatile coal with a heating value of 27 MJ/

kg, the simplifi ed calculation gives:

Evaporation loss

Combustion heat× =

××

× =1000 4 2 3

0 6 27100 5 6

×4 2. %68

Of course, the evaporative heat loss can be reduced by increasing coal

content in the CWS as shown in Fig. 12.32. For a highly loaded CWS with

70% coal the minimum evaporative heat loss will be about 3.65% for high-

volatile bituminous coal (27 MJ/kg), and 4.9% for subbituminous coal (20

MJ/kg). However, the most important requirement for the CWS is that it

must be pumpable. Consequently the problem is how to increase solids

concentration in CWS without raising viscosity above an acceptable level.

The answer to this question has two components: (i) the effect of coal par-

482

4Eva

pora

tive

heat

loss

(%

)

6

8

10

12

52 56 60% coal

64 68 72

: 20 MJ/kg: 27 MJ/kg

12.32 Effect of coal content in CWS and coal heating value on

evaporative heat loss during combustion of CWS. (Source: After

Laskowski, 2001 with permission of Elsevier.)



ticle size and particle-size distribution, and (ii) the effect of coal surface

properties and chemical additives (referred to as viscosity reducers).

Effect of coal particle-size distribution on rheology of CWS

The maximum packing fraction, φmaφφ x in Equation [12.13], increases with

increasing polydispersity of the suspension. Broader particle size distribu-

tions have higher values of φmaφφ x because the small particles fi t into the gaps

between the larger ones. On the other hand, non-spherical particles lead to

poorer space-fi lling and hence lower φmaφφ x . Particle aggregation, for exam-

ple fl occulation, also leads to a low maximum packing fraction because, in

general, the fl ocs are not close-packed. This was experimentally confi rmed

by Farris (1968). Since coal content in CWS is the most important factor

determining its utility, the effect of the coal particle-size distribution on the

rheology of CWS was thoroughly examined.

The top particle size in CWS is determined by boiler requirements; on the

other hand, the size and the yield of the fi nes fractions are determined by

CWS viscosity and grinding cost. Because of the burning process limitations,

it is usually assumed that the top size cannot exceed 250 μ m with a particle-

size distribution of 70–80% below 75 μ m. With these limits in mind, Ferrini

et al . (1984) showed how the coal particle-size distribution could be opti-

mized with regards to viscosity of the highly loaded CWS. Figure 12.33 shows

particle-size distribution of the samples tested by Ferrini et al . (1984).

10

20

40

60

80

Cum

ulat

ive

% r

etai

ned

100

10

7

8

3

12

11109

Particle size (μm)

100

12.33 Particle-size distribution of a family of bimodal particle samples

derived from the continuous reference curve No. 3. (Source: After

Ferrini et al ., 1984.)



Curve 3 in Fig. 12.33 shows the particle-size distribution of the sample

prepared by grinding. The bimodal suspensions were prepared by screen-

ing − 20 μ m and +45 μ m size fractions from this sample and mixing them at

different ratios. Apparent viscosity for the samples varying in particle-size

distribution is shown in Fig. 12.34. As this fi gure demonstrates, the changes

in viscosity with particle-size distribution are quite dramatic. Samples 9, 10,

and 11, that is the samples that contain about 40% of the fi ne fraction, have

much lower viscosities than Sample 3.

Ferrini et al .’s (1984) results clearly demonstrate that a highly loaded

CWS can be obtained only by properly adjusting the coal size distribution.

The benefi cial effects that can be obtained by manipulating the particle-

size distribution are considerable, and therefore grinding and selection of

optimum particle-size distribution is an import aspect in all CWS-patented

technologies.

The Chinese authors (Zhang Rong-Zen et al ., 1984; Zuna Wang et al ., 1993) claim that the Chinese experience indicates that if a fi ne coal is unsuit-

able in its as-received form for CWS preparation, in most cases this is due to

low content of ultrafi ne particles.

Effect of coal surface properties on rheology of CWS

Rheology of fi ne particle systems depends on many factors, particle–particle

interactions being one of them. As shown by Xu and Yoon (1989, 1990),

hydrophobic particles tend to coagulate in aqueous environment (Fig. 12.13).

We studied such effects with the use of rheological methods.

20 30 40

Fine fraction (%)

11

400

800

1200

Vis

cosi

ty (

mP

a .s

)

1600

2000

2400

10

9

83

7

12

50 60

Continuoussize dist.

70

12.34 Effect of − 20 μ m fraction content on apparent viscosity of

bimodal coal-water slurries at constant overall coal content (72% by

wt.). (Source: After Ferrini et al ., 1984.)



The tests were carried out with F4 sample of a bituminous coal (ash 11.7%,

moisture 0.6%, fi xed carbon d.m.m.f 73.3%) from a mine in British Columbia.

The contact angle measurements used to characterize its wettability by

water gave 90 ° (advancing contact angle). The suspensions were prepared

after dry pulverization of the sample below 212 μ m and all the rheological

tests were carried out at 59% solids by wt.

As Fig. 12.35 reveals, for the suspensions prepared from the fresh F4 coal

sample, very high yield stress values were obtained over a very broad pH

range (from about 4 to 9). For the sample oxidized at 85–90 ° C over 24 h,

high yield stress values were recorded only around pH of 4.5, while for the

sample oxidized over 221 h, the yield stress maximum is obviously situated

at pH much lower than 3.5. The point is that the iso-electric points for these

three samples were found to be at pH 8 for the fresh sample, around pH of

4.5 for the sample oxidized for 24 h, and at pH of 2.5 for the heavily oxidized

sample. Thus, while the hydrophobic fresh sample coagulates over a broad

pH range from about 4 to 9 (Fig. 12.35) and it is not much affected by the

position of the iso-electric point (i.e.p.), the oxidized sample (24 h) coagu-

lates only around its i.e.p. The heavily oxidized sample has i.e.p. around pH

2.5 and it can be expected that the maximum coagulation pH is also situated

around this pH value. This agrees remarkably with the data published by Xu

and Yoon (1989, 1990).

30

4

8

Cas

son

yiel

d st

ress

(P

a)

12

16

20

4 5 6 7 8

pH

9 10 11 12 13

F4 Fresh F4 Oxidized (24 h)

F4 Oxidized (221 h)

12.35 Yield stress values for aqueous suspensions of the bituminous

coal as a function of pH. (Source: After Melo et al ., 2004, with

permission of the Metallurgical Society of CIM.)



In the rheological measurements, whenever suspended particles coag-

ulate and the network of the interacting particles is formed, high yield

stress values are recorded. The maximum values correspond well with the

maximum coagulation (Johnson et al ., 2000). While strongly hydrophobic

particles coagulate over a broad pH range (see Fig. 12.13), the hydrophilic

particles coagulate only over the pH range at which these particles do not

carry electrical charge.

In total agreement with these conclusions are our rheological measure-

ments carried out with CWS prepared from the same F4 bituminous coal in

the presence of humic acids.

Some of these results are shown in Fig. 12.36. Direct measurements of the

wettability of the tested coal in the presence of humic acid indicate that this

hydrophobic bituminous coal becomes hydrophilic as a result of humic acid

adsorption. This is not surprising, as humic acids are hydrophilic polymers.

This compound resulting from oxidation of organic matter, or obtained

directly by extraction of low-rank coals, is a complex anionic polyelectro-

lyte that contains both carboxylic and phenolic functional groups. There

is evidence that irrespective of the original coal surface properties, humic

acids impart properties similar to low-rank coals to all coals (Laskowski,

100 10

20

30

Con

tact

ang

le (

°)

40

50

60

70

80

90

200

400Yie

ld s

tres

s (P

a)

600

800

1000

1200

100 1000

Humic acid conc. (ppm)

10 000

F4 advancing F4 receding F4 yield stress

12.36 Effect of humic acids on wettability of bituminous coal and on

the yield stress of 65% (wt.) CWS prepared from this coal. (Source:

After Pawlik et al ., 1997 with permission of the Metallurgical Society of

CIM.)



2001). The electrokinetic measurements revealed that the coal particles in

the humic acids solutions acquire a very negative charge, as evidenced by

very negative zeta potential values (almost − 70 mV) (Pawlik et al ., 1997).

The increased hydrophilicity of coal particles and their increased negative

charge must stabilize such particles against coagulation. As a result, the rhe-

ological properties of such suspension become more Newtonian, as indi-

cated by decreasing values of the yield stress.

Coals varying in rank differ in surface properties. While bituminous coals

are quite hydrophobic, the subbituminous coals are not (Fig. 12.2). The rhe-

ological measurements depicted in Figs 12.35 and 12.36 indicate that these

differences in surface properties must affect the rheological properties of

the suspensions prepared from such coals. Since hydrophobic solid particles

suspended in aqueous environment tend to aggregate, the rheological prop-

erties of such suspensions should be different from the rheological proper-

ties of the suspensions prepared from low-rank coals.

Figure 12.37 shows, after Schwartz (1985), the apparent viscosity of the

coal–water slurry prepared using three different coal samples. This fi gure

indicates that while it is possible to obtain highly loaded CWS from Coal 3

(76% wt.) at the limiting viscosity of 1000 mPa.s, for Coal 1 the maximum

solids content at this viscosity is only about 65%. The wettability tests gave

for the three coal samples the following contact angles: 90 ° for Sample 3, 40 °

for Sample 2, and 20 ° for Sample 1. These samples were also characterized

by chemical analyses that gave the following C/O contents: 4.6 (Sample 1),

550

500

App

aren

t vis

cosi

ty (

mP

a.s)

1000

1500

60

Coal 1

Coal 2

Coal 3

65 70

Concentration of coal (wt. %)

75 80

12.37 Apparent viscosity of coal–water slurry prepared from three

different coal samples with the same stabilizer (0.05% of non-ionic

ethylene oxide/propylene oxide copolymer). (Source: After Schwartz,

1985.)



12.7 (Sample 2), and 20.8 (Sample 3), with the following contents of C/H:

13.6 (Sample 1), 15.8 (Sample 2), and 18.5 (Sample 3). These results reveal

that while Sample 3 is a very hydrophobic bituminous coal, the other two

samples were clearly less hydrophobic, with Sample 1 being a low-rank sub-

bituminous coal.

Figure 12.37 illustrates well the existing controversy. As is seen from this

fi gure, much higher coal content (loading) can be achieved with higher-

rank coals than with low-rank coals, and this is sometimes interpreted that

it is easier to prepare CWS from high rank coal. The problem is that coal is

highly porous and heterogeneous. Coal moisture content (equilibrium mois-

ture) is a function of coal rank. While it is very low for bituminous coals

(around 1%) it can be as high as 20% for subbituminous coals. Since the

maximum packing fraction, maφφ x , is around 0.75, the minimum content of

water in suspension must be at least 25%. This is needed as water is a ‘lubri-

cating’ medium, without which this fi ne particle system cannot behave as a

fl uid. For a low-rank coal, containing 20% equilibrium moisture (coal inter-

nal moisture), the minimum content of water is then around 45%, while for

the bituminous coal this can be around 26%. Since water content in CWS is

the most important factor that determines its calorifi c value, a few processes

have been developed to reduce the ‘internal’ water content in the low-rank

coals used in the preparation of CWS.

Figure 12.38 shows apparent viscosity of CWS, measured at a shear rate of

100 s − 1 , prepared from six different coals. In all cases, 1% of the same non-

ionic additive was utilized. Several conclusions are evident from these plots.

First, the experimentally measured apparent viscosity dramatically increases

at a given solids content, and second, this limiting solids content is different

for different coals. Since the particle-size distributions of these samples were

similar, the conclusion is that the rheology of CWS depends on coal surface

properties. Seki et al . (1985) derived the following relationship:

M CM

f cC c+⎛⎝⎛⎛⎛⎛⎝⎝⎛⎛⎛⎛ ⎞

⎠⎞⎞⎞⎞⎠⎠⎞⎞⎞⎞

1 1100

[12.14]

where Mf is the free water in CWS, Cc is the dry coal content in CWS, and

Mc is the coal equilibrium moisture content.

So, obviously, for a highly loaded CWS (large Cc in Equation [12.14]),

the free water content will sharply decrease if the coal moisture content is

high, and this will raise the viscosity of CWS. The relationship between coal

content in CWS at a viscosity of 1000 mPa.s and coal equilibrium moisture

content is shown in Fig. 12.39.

As these results demonstrate, the rheological properties of CWSs pre-

pared from different coals characterized by similar particle-size distributions



can be correlated with the coal moisture content, and then coal rank. The

relationship between the CWS viscosity and coal moisture content can be

used to estimate CWS’s slurryability (maximum coal content in the CWS)

at an acceptable viscosity.

Additives in preparation of CWS

CWS must be pumpable and this requirement translates into suffi ciently

low viscosity, but CWS may be stored in tanks over a long period of time,

and thus must also exhibit suffi cient sedimentation stability. Different addi-

tives are required to reduce viscosity and to increase stability.

Dispersants. Many low molecular weight polymers (MW < 100 000

daltons) adsorb onto coal surfaces and can be used as dispersants for

coal–water suspensions. All good coal dispersants are also very effi cient

depressants for coal fl otation (Pawlik, 2005). Both anionic (e.g., carboxym-

ethyl cellulose, humic acid, polystyrene sulfonate, etc.) and non-ionic (dex-

trin) depress coal fl otation and, by dispersing coal particles, improve CWS’s

60200

1000

App

aren

t vis

cosi

ty (

mP

a.s)

5000

65

Coal content (%)

70 75

Coal

Under 74 mmparticlesize (%)

78818079817677

1.01.01.01.01.01.01.0

1.42.43.33.34.95.68.6

Additive% concentration

Equilibriummoisture

content (%)

Coal HCoal GCoal ECoal CCoal ACoal DCoal F

×

12.38 Relation between apparent viscosity and coal content in CWS for

different coals. Non-ionic additive 1% on dry coal basis. (Source: After

Seki et al ., 1985.)



stability. According to Hara et al . (1992), while good stability requires steric

stabilization, a good slurryability requires strong electrostatic stabilization.

Figure 12.40 depicts the effect of dispersant concentration on the viscosity

of CWS prepared from three different coals (Higashitani et al ., 1990).

As Fig. 12.40 shows, the CWS viscosity can clearly be reduced only at

rather high dosages of dispersants (around 1000 g/t). Yoshihara (1999) syn-

thesized graft copolymers and tested their effect on CWS rheology. The graft

copolymers of sodium polyacrylic acid with polystyrene side chain showed

higher adsorption onto coal and turned out to be better viscosity reducers

for CWS. Polystyrene sulfonate (PPS) with molecular weight of 14 000 is

commercially used in Japan for the preparation of CWS.

Additives stabilizing CWS against settling (anti-settling additives). CWSs are made of relatively coarse particles. Such particles tend to settle

under gravity and, when stabilized against aggregation, settle as individ-

ual particles and form a compact sediment that may be diffi cult to re-

disperse. Since CWSs may be stored in tanks over a considerable period

of time, such a system must therefore also be stabilized against settling.

Figure 12.41 shows schematically how the effect of the structure that

develops as a result of particle–particle interactions affects the rheology

and stability of CWSs.

First, it must be borne in mind that the properties of CWSs are very

different from the properties of the disperse systems dealt with by colloid

060

65

Coa

l con

tent

at 1

000

mP

a.s

(%)

70

75

2 4 6

Equilibrium moisture (%)

8

Anionic

Coal-type:

Coal B

Coal D

Coal G

Coal ACoal E

Coal F

Coal CCoal H

Additive-Concentration:Anionic; 0.6%Nonionic; 1.0%

Ani

onic

Non

ioni

c A

Non

ioni

c B

Nonionic

10 12

12.39 Relation between coal content in CWS at 100 mPa.s and coal

equilibrium moisture. (Source: After Seki et al ., 1985.)



chemistry. The most obvious difference is the solid content. While in a

very dilute disperse system any aggregation between particles results in

a loss of stability and fast sedimentation, in a heavily loaded suspension

the aggregation may prevent particles from settling and so it may stabi-

lize the system. But aggregation creates a network of interacting parti-

cles, and this will increase viscosity. Therefore, it is obvious that particles

in CWS must be properly dispersed to reduce viscosity, and to increase

maximum packing density. Dispersants are used for this reason. But since

CWS must also be stable, some weak aggregation will also be required.

This is achieved with the use of high molecular weight polyelectrolytes

(for instance various natural gums with MW > 10 6 ) (Saeki et al ., 1999).

CWS from low-rank coals

There are many deposits of subbituminous coals and lignites that can easily

be recovered by strip mining. Very often these coals constitute a very attrac-

tive ‘clean’ energy source, since they may contain 0.2% sulfur and 8% ash

(e.g. Alaskan low-rank coals). However, the high inherent moisture content

of these coals and their low heating value on an as-mined basis compromise

the economics for rail haulage. Consequently, most low-rank coals are con-

sumed by electric utilities located near mines sites.

10–40

2

4η (P

a.s)

6

8

10–3 10–2

Cs, kg-surfactant/kg- coal

10–1

D = 34.0 s–1

D D 59T H 61

W W 63Coal Cc (wt %)

12.40 Effect of concentration of sodium salt of naphthalene

formaldehyde sulfonate on CWS apparent viscosity at 34 s − 1 . Symbols

WW, DD and TH stand for different coal samples. (Source: After

Higashitani et al ., 1990 with permission of the Mining and Materials

Processing Institute of Japan.)



In an effort to upgrade coal and produce a transportable fuel, several

dewatering processes have been developed. Almost all of the coal-inher-

ent moisture content can be removed by thermal drying; the fi nal mois-

ture content achieved in the product is dependent on particle size and

residence time. However, since low-rank coals depreciate due to shrink-

age and loss of structural elasticity when they are dried in hot gases, the

dried product is dusty and is subject to spontaneous combustion. When

the dried coal is slurried in water, a signifi cant portion of the coal mois-

ture removed during drying is reabsorbed onto coal particles. Such slur-

ries have only slightly higher energy densities than similar slurries made

from raw coals (usually less than 11.6 MJ/kg (Potas et al ., 1986)). To min-

imize moisture reabsorption, the dried coal can be coated with a fuel oil

or can be briquetted.

Many so-called low-rank coal (LRC) drying processes have been devel-

oped (Willson et al ., 1997). In one such process, the hot-water drying process

High

Welldispersed

Poor

Low

None

High

GoodSedimentation stability

Viscosity

Yield stress

Maximum solids content

High

High

Low

Weaklyaggregated

Extensivelyaggregated

ZeroSurface charge

12.41 Illustration of concentrated suspensions showing that at a

given solid’s content the volume of sediment depends on particle

aggregation. (Source: After Laskowski and Parfi tt, 1989 with permission

of Taylor & Francis.)



(HWD), coal is heated under pressure in water. Under such conditions,

evolving tars remain on the coal surface and plug micropore entrances. The

results shown for a subbituminous coal C from No. 4 seam of Usibelli Coal

Mine, Alaska, confi rm the very strong effect of temperature on the process;

equilibrium moisture content decreases from 25% down to 9.5%, and calo-

rifi c value increases from 27.3 mJ/kg to almost 31 MJ/kg when the process

temperature is increased from 275 ° C to 325 ° C (Walsh et al ., 1993). Ohki

et al . (1999) used FTIR spectroscopy and showed that in the processing of

an Indonesian LRC, Adaro (16.42% moisture, 45.6% volatile matter, 1.24%

ash, 36.74% FC on as-received basis), the oxygen content of coal was signi-

fi cantly reduced.

This results from decomposition of carboxylic groups (and any reduction

in the content of oxygen in coal translates into lower equilibrium moisture

content). Potas et al . (1986) demonstrated that the maximum solids content

in the CWS prepared from two Alaskan coal samples (lignite and subbitu-

minous) can be increased from the 45% range to about 55–60% range.

In the vacuum drying/tar coating method (Usui et al ., 1999) the coal is

dried under vacuum at 200 ° C, then at 270–350 ° C and is coated with tar (5%

wt./coal), and it is used to prepare CWS.

Coal reverse fl otation

Bituminous coals are hydrophobic (see Fig. 12.2), fl oat easily, and therefore

forward fl otation is a common practice. But since subbituminous coals, and

also the coals stored in old tailings ponds, which are not that hydrophobic,

are diffi cult to fl oat the reverse fl otation may be quite an attractive option

in such cases. The reverse fl otation of coal has recently been shown to be

possible (Ding and Laskowski, 2006).

There are fundamental differences between the forward and reverse coal

fl otation. In the case of the forward fl otation, the clean coal is recovered as

a froth product that is made hydrophobic with the use of various fl otation

reagents. In the case of reverse fl otation, the clean coal product is what is

ending up in the fl otation tailings; this product is kept as hydrophilic as pos-

sible, to make its fl otation impossible (Fig. 12.42). While these differences

may not seem to be very signifi cant, they are extremely important when

these products are utilized to make a CWS. These differences are clearly

seen when rheological curves for the clean coal product from the forward

fl otation (froth product) (Fig. 12.43) are compared with the rheological

curves obtained for the CWS prepared also for the clean coal but obtained

from the reverse fl otation (these are fl otation tailings) (Fig. 12.44).

The fi rst and the most obvious difference is that in the fi rst case, that is

when the froth product from forward fl otation is used to prepare CWS,

the dispersant is needed to increase the CWS loading (Fig. 12.43). PSS10



dispersant (polystyrene sulfonate with MW of 14 000) was from LION

Corp., Japan. In the latter case, when the fl otation tailings from the reverse

fl otation are used, high solids loading can be obtained without any dis-

persant (Fig. 12.44). This product is suffi ciently hydrophilic and does not

require the dispersant to prepare CWS. The reverse fl otation is not as simple

as forward fl otation and in the case discussed required the use of cationic

Raw coal

Raw coal

Forward flotation

Reverse flotation

Concentrate (coal)

Concentrate (gangue)

Tailings (gangue) (reject)

Tailings (coal)

12.42 Schematic illustration of coal forward and reverse fl otation

processes.

460

200

400

App

aren

t vis

cosi

ty (

mP

a.s)

600

800

1000

1200

1400

1600

48 50 52

Solids content (%)

54 56 58 60

Forward in water

Forward with 0.5% PSS10

Forward with 1% PSS10

12.43 Effect of PSS10 dispersant on apparent viscosity (calculated at

a shear rate of 100 s − 1 ) of slurries prepared from the forward fl otation

concentrates. (Source: After Ding and Laskowski, 2009). Copyright

Taylor & Francis Group, LLC. (http://www.taylorandfrancis.com),

reproduced with permission.)



collector (Arquad 12–50 from AKZO Nobel) along with some depressants/

dispersants (dextrin, tannic acid and A-100 polyacrylamide fl occulant from

Cytec). This, however, does not look so bad if compared with the dosages

of dispersing additives that must be used in the preparation of CWS (they

are usually in the range of 1%, that is 10 kg/t) when fl otation froth product

is utilized.

12.5.3 Rheology of magnetite dense media

Dense medium separation

In a dense-medium separation raw coal is introduced into the dense medium,

whose density is higher than the density of the lighter constituent (coal) but

lower than the density of the heavy constituent (inorganic rock). In a per-

fect separator, the particles have an infi nite time to report to either sink or

fl oat products. In practice, a limited time is available for the separation, and

the rate with which the particles move (relative to the medium) determines

the outcome of the process. Particles characterized by a small density dif-

ferential or small size do not move in the medium fast enough and will be

misplaced. Thus, the separation effi ciency is based on the phenomena that

determine sedimentation of solid particles in liquids.

460

200

400

App

aren

t vis

cosi

ty (

mP

a.s)

600

800

1000

1200

1400

48 50 52

Solids content (%)

54 56 58 60

Reverse in water

Reverse with 0.5% PSS10

Reverse with 1% PSS10

12.44 Effect of PSS10 dispersant on apparent viscosity (calculated at

a shear rate of 100 s − 1 ) of slurries prepared from the reverse fl otation

tailings. (Source: After Ding and Laskowski, 2009). Copyright Taylor &

Francis Group, LLC.(http://www.taylorandfrancis.com), reproduced with

permission.)



Settling phenomena

A particle settling in a liquid under gravity is subject to two forces, a driv-

ing force due to gravity and a drag force that opposes motion. The equation

resulting from such considerations that describes the terminal velocity of

solid particle in a liquid can be written in the following form:

vd g

Ctp

=4

3

( ))

ρ [12.15]

where d is the spherical particle diameter, δ is the particle density, ρ is the

liquid density, and C p is the dimensionless drag coeffi cient.

C fd pf (Re )p [12.16]

and since

Rep

vd=

ρvvη

[12.17]

the drag acting upon the moving particle depends also on the medium vis-

cosity ( η ) but to the extent that depends on the prevailing hydrodynamic

conditions.

For small spherical particles ( d < 100 μ m)

Cdp

=24

Re [12.18]

and this, when substituted into Equation [12.15], gives the Stokes equation

vd g

t =2

18

( ))

η [12.19]

As this equation indicates, the settling rate of small particles strongly

depends on medium viscosity.

For large particles ( d > 1 mm) C d ≈ 0.44 and Equation [12.15] gives

Newton’s equation



vgd

t =10

3

( )))−ρ

[12.20]

showing that in a turbulent fl ow regime the settling rate of spherical par-

ticles does not depend on viscosity at all.

There are a few empirical equations (e.g. Allen’s equation) that were

derived for the intermediate Reynolds number range 1 <Re < 10 3 (Concha

and Almendra, 1979). They all indicate only weak dependence of the parti-

cle settling rate on viscosity.

The point is that the settling rate of solid particles in liquid strongly

depends not only on particle size but also on the density differential ( δ - ρ ).

For small values of this differential, that is for so-called near-density par-

ticles, the settling rate will be slow (as in the Stokes equation) also for larger

particles. The movement of these particles in the medium will then also be

described by the Stokes equation, and will strongly depend on the medium

viscosity.

In a static dense-medium separation (dense-medium bath), the gravi-

tational driving force responsible for sedimentation of solid particles in

li quids is given by:

Fd

ggFF =πdd3

6( )δ ρ− [12.21]

where g is the acceleration of gravity.

In a cyclone, the acceleration of gravity is substituted by a centrifugal

acceleration (Sokaski et al ., 1991)

Fd V

rcFF = ( )πdd δ ρ− )3 2V

6 [12.22]

where F c is the centrifugal force, V is the tangential velocity, and r is the

radius of cyclone.

In a typical cyclone, the centrifugal force acting on a particle in the

inlet region is about 20 times greater than the gravitational force in a

static bath, but in the conical section of a cyclone this force is much

larger. These large forces responsible for separating coal from inorganic

gangue are therefore much larger in a cyclone than in a static bath. This

explains the cyclone’s large capacity and its ability to process fi ne coal

particles.



Dense media

The most common media for separation is a fi ne magnetite suspension in

water, ferrosilicon suspension, or a mixture of the two, depending on the

required separation density. Magnetite alone is used when a density is

required in the range from 1.25 to 2.2 g/cm 3 , a mixture between 2.2 and 2.9 g/

cm 3 , and ferrosilicon alone for a density range 2.9–3.4 g/cm 3 . With magnetite

density of about 5.0 g/cm 3 , 500 g of magnetite per liter of suspension, that is

magnetite content of 10% by volume (35 % wt.) the suspension density is

1.4 g/cm 3 , and with doubled magnetite content to 20% by volume (55% wt)

magnetite suspension density will be 1.8 g/cm 3 .

The same calculations for ferrosilicon alloy, assuming that its density is

6.8 g/cm 3 (Collins et al . (1974) claims that density of ferrosilicon containing

14–16% Si falls in the range from 6.6 to 7.0) give for the range of media den-

sities from 2.9 to 3.4 g/cm 3 the ferrosilicon volume percent of 33–41%.

According to Mikhail and Osborne (1990), the following is the standard

particle-size distribution of magnetite used in coal preparation: maximum

5% by mass coarser than 45 μ m, and 30% by mass fi ner than 10 μ m.

Magnetite medium viscosity

Bergh ö fer (1959) was probably the fi rst to obtain full rheological curves for

magnetite dense-medium suspensions. His results clearly demonstrate plas-

tic properties of such suspensions (see Fig. 12.30) and show that the devia-

tion of the rheological properties of magnetite suspensions from Newtonian

behavior quickly increases with fi neness of the magnetite. Meerman (1959),

Whitmore (1958), and Yancey et al . (1958) sampled magnetite dense media

from operating coal preparation plants and confi rmed that such suspensions

behave like Bingham plastic fl uids.

A few samples of magnetite used in our rheological studies (He and

Laskowski, 1995) are listed in Table 12.3.

Particle-size distributions of the tested samples (Table 12.3) are well

described by the Rosin-Rammler-Bennett (RRB) particle-size distribution:

d 63.2 is the size modulus (d 63.2 is that aperture through which 63.2% of the

sample would pass), and m is the distribution modulus (the slope of the

curves on the RRB graph paper), with diminishing value of the distribu-

tion modulus the line representing distribution becomes less steep and that

translates into larger differences between the sizes of fi ne and coarse par-

ticles in the sample.

Mag #1 was the Craigmont Mine commercial magnetite used in western

Canadian coal preparation plants. Mag #2 was prepared by grinding Mag #1.

Mag #3 was obtained by rejecting fi nes from Mag #1 in a classifying cyclone.

Mag #4 and Mag #5 were micronized magnetites (70% and 90% below 5



μ m, respectively), provided by the US Department of Energy in Pittsburgh

(Klima et al ., 1990).

Because of high magnetite density, in spite of fi ne size of the magnetite

particles in the tested samples, these samples settle rather quickly and rhe-

ological measurements require special equipment. A sensor system (ESSP)

that was specially designed and built for magnetite suspensions (Klein et al ., 1995) was employed in the tests, along with a HAAKE Rotovisco RV 20

rheometer. The suspensions were demagnetized with the use of demagnetiz-

ing coil before measurements. The results are shown in Fig. 12.45 (He and

Laskowski, 1999).

As these results indicate, the yield stress for coarse magnetite suspen-

sions (Mag #1) at solids content below 10% by volume is very low and the

existence of yield stress is more clearly manifested in the fi ne magnetite

suspensions (Mag #4 and Mag #5). High yield stress values in these suspen-

sions provide a proof for quite strong particle-particles interactions in such

fi ne disperse systems.

A few equations exist to describe rheological curves. Perhaps the most

common is the Bingham equation, which describes the fl ow curves of plastic

fl uids (see Fig. 12.30).

τ τ η=τB pτ η=ττ lD [12.23]

where τ is the shear stress, τ B is the Bingham extrapolated yield value (see

Fig. 12.46), ηplη is the plastic viscosity, and D is the shear rate.

The equation shown to describe rheological curves of magnetite suspen-

sions particularly well is the Cason equation (Klein et al ., 1990; He and

Laskowski, 1999; He et al ., 2001)

τ τ η= +τ ( )⎡⎣ ⎤⎦⎤⎤C Cη+τ (1 2τ 1 2 2/D [12.24]

where τCτ is the Casson yield stress and ηCη is the Casson viscosity.

Table 12.3 Particle-size distributions

of the tested magnetites

Sample d 63.2 ( μ m) M

Mag #1

Mag #2

Mag #3

Mag #4

Mag #5

30.5

18.0

33.0

4.3

2.7

3.2

1.6

4.1

1.9

2.5



00

0

10

20

30

40

50

60

0

10

2022%

18%

18%

20%

15%

12%

2%

15%

10%5%

30

40

1

2

She

ar s

tres

s (P

a)S

hear

str

ess

(Pa)

3

4

5

6

0

1

2

3

4

5

8

6

7

50 100 150 200 250 300

5%

15%

22%

25%

28%

Solid content 30%

Solid content 28%

Solid content 25% Solid content 22%

25%

22%

20%

18%

12%

5%

Mag #1 Mag #2

Mag #4 Mag #5

0 50 100 150 200 250 300

0 50 100 150 200 250 300

Shear rate (s–1)

Shear rate (s–1)

0 50 100 150 200 250 300

12.45 The fl ow curves of the tested samples of magnetite suspensions.

(Source: After He and Laskowski, 1999). Copyright Taylor & Francis


permission.)



Figure 12.47 shows the Casson yield stress determined from the experi-

mental data displayed in Fig. 12.45.

As Fig. 12.47 reveals, the Casson yield stress is very sensitive to the changes

in the size of magnetite particles and solids content, while the Casson vis-

cosity turned out not to be very sensitive to these changes over the studied

medium density range (it is practically constant over the medium density

range from 1.2 to 1.7 g/cm 3 ). The Casson yield stress and viscosity values for

four tested magnetite samples at a medium relative density of 1.45 are given

in Table 12.4.

As these results demonstrate, the yield stress values are much larger than

the viscosity, and while the yield stress increases with the fi neness of the

tested magnetite samples, the Casson viscosity does not. Since the appar-

ent viscosity of a non-Newtonian system (such as magnetite suspension)

is determined by both yield stress and viscosity (that is either by Bingham

yield stress and plastic viscosity, or Casson yield stress and Casson viscosity),

it can be shown that the contribution coming from the yield stress exceeds

by far the contribution of the viscosity (He et al ., 2001).

Mineral particles that are treated in a dense medium must move in the

medium in order to report to the proper products (low density concentrate

or high density tailings). In order to start moving in the medium, in order to

00

0.5

τB

τC

1She

ar s

tres

s (P

a)

1.5

Bingham model Casson model Data

2

2.5

3

50 100

Shear rate (1/s)

150 200 250

12.46 Example of rheogram of slurry fi tted with either Bingham model

or Casson model; as this example demonstrates, obtained τ B and τ C

values are not identical.



shear the medium (in a static bath this happens under gravity), the particles

must overcome the threshold drag. It is determined by the yield stress. In the

dense-medium cyclone, the acceleration is dominated by centrifugal accel-

eration; the results shown in Figs 12.48 and 12.49 were calculated assuming

centrifugal acceleration of 40 m/s 2 .

As Fig. 12.48 shows, while only particles with size smaller than 1 mm will

not be able to shear the Mag #1 and Mag #2 media (in a dense-medium

cyclone at 40 m/s 2 centrifugal acceleration), the near-density particles must

be larger than 10 mm in order to be able to shear the Mag #4 medium.

For 2 mm particles, the movement in the Mag #1 and Mag #2 media

requires only density differential larger than 0.1 g/cm 3 . In Mag #4 the par-

ticles with a size of 2 mm will be able to move in the medium only if their

1.10.001

0.01

0.1

Cas

son

yiel

d st

ress

(P

a)

1

10

Temp. = 19°C

100

1.2 1.3 1.4 1.5 1.6 1.7

Medium density (g/cm2)

1.8 1.9 2 2.1

Mag#1 Mag#2 Mag#4 Mag#5

2.2 2.3

12.47 The effect of magnetite particle size and medium solid content on

Casson yield stress. (Source: After He and Laskowski, 1995). Copyright

Taylor & Francis Group, LLC.(http://www.taylorandfrancis.com),

reproduced with permission.)

Table 12.4 Casson yield stress and viscosity

for four studied magnetite samples at a

medium relative density of 1.45

Sample τ c (mPa) η c (mPa.s)

Mag #1

Mag #2

Mag #4

Mag #5

62

118

2110

2660

1.85

1.50

0.70

1.50



density is about 0.6 g/cm 3 larger than the medium density (Fig. 12.49). These

results demonstrate how important are dense-medium rheological proper-

ties for separation effi ciency.

The medium capacity to maintain the homogeneity, the constant medium

separation density δ 50 in the separating device, is a function of medium sta-

bility. The medium used in dense-medium separation of coal is an aqueous

Mag #2Mag #1

Mag #4

Mag #5

00

400

800

She

ar r

ate

D (

s–1)

1200

1600

2000

5 10 15

Particle size (mm)

20 25 30

12.48 Effect of particle size on effective shear rate of a particles with

a relative density of 1.55 in various magnetite dense media with a

relative density of 1.45. (Source: After He et al ., 2001 with permission of

Elsevier.)

00

200

400

She

ar r

ate

D (

s–1)

600

800

1000

0.3 0.6

Mag #2

Mag #1

Mag #4

Mag #5

Relative density differential

0.9 1.2

12.49 Effect of medium density differential on shear rate for a 2 mm

particle in various magnetite dense media with relative density of 1.45.

(Source: After He et al ., 2001 with permission of Elsevier.)



suspension of fi ne magnetite particles. Static stability of magnetite suspen-

sion is defi ned as the reciprocal of the settling velocity that is easily deter-

mined in the lab (Bozzato et al ., 1999). In a cyclone, magnetite particles

are also subjected to centrifugal forces and tend to segregate creating den-

sity gradients. This tendency to segregate is a measure of the magnetite

medium stability under dynamic conditions prevailing in a cyclone.

Practically, it is determined by the density differential between the cyclone

underfl ow and overfl ow. As Fig. 12.50 indicates, the dynamic medium stabil-

ity critically depends on magnetite particle size and medium density (that is

magnetite concentration in the medium).

These measurements confi rm that the measured density differential

clearly depends on magnetite particle size, and it was shown to be a func-

tion of the yield stress (He and Laskowski, 1994). As could be expected, Mag

#3 turned out to be very unstable, as characterized by a huge density differ-

ential of 1 g/cm 3 . For Mag #1 the maximum density differential is about 0.4

g/cm 3 . Collins et al . (1983) claim that separation effi ciency is satisfactory as

long as the medium density differential is in the range of 0.2–0.5 g/cm 3 .

All the studies confi rm that the behavior of the medium is process-deter-

mining. In the empirical relationship proposed by Napier-Munn (1990) to

describe the effect of medium rheology on DMC separation, the separation

10

0.05

0.1

0.15

Den

sity

diff

eren

tial (

g/cm

3 )0.2

0.25

0.3

1.1 1.2 1.3 1.4

Feed medium density (g/cm3)

1.5

Mag#4

Mag#1

Mag#2

Mag#3

1.6 1.7 1.8 1.90.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

1.1

12.50 Effect of medium composition on medium differential measured

in a 6” cyclone (D6B-12-S287 Krebs Engineering Int.) that was gravity

fed at 10xD inlet pressure. (Source: After He and Laskowski 1995).

Copyright Taylor & Francis Group, LLC.(http://www.taylorandfrancis.

com), reproduced with permission.)



effi ciency expressed by a value of E p was related to the medium apparent

viscosity by:

E dnp [12.25]

where E p is the probable error, and Φ is a function of apparent viscosity, Φ = ( )p p

f .

Since the yield stress is a dominant factor that determines the apparent

viscosity of fi ne magnetite suspensions, the experimentally measured yield

stress can practically substitute for apparent viscosity in Equation [12.25]

(He et al ., 2001).

Optimizing medium composition is essential to achieving balanced sta-

bility and rheological properties of the dense medium (He and Laskowski,

1995). The effect of medium composition on DMC performance depends

on feed particle size. The separation of coarse particles (larger than 2 mm)

is strongly affected by the medium stability and only slightly by medium

rheology. Thus, the separation effi ciency improves when medium stability is

improved by using fi ner magnetite dense medium. On the other hand, the

separation of fi ne feed particles (below 0.5 mm) is more strongly affected

by the medium rheology.

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